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UNITED STATES

SECURITIES AND EXCHANGE COMMISSION

Washington, D.C. 20549

 

 

FORM 8-K

 

 

CURRENT REPORT

Pursuant to Section 13 or 15(d) of the

Securities Exchange Act of 1934

Date of Report (Date of earliest event reported): March 31, 2025

I-80 GOLD CORP.

 

 

(Exact name of registrant as specified in its charter)

 

British Columbia

 

001-41382

 

Not Applicable

(State or other jurisdiction of

incorporation)

 

(Commission

File Number)

 

(IRS Employer

Identification No.)

 

5190 Neil Road, Suite 460

Reno, Nevada, United States

  89502
(Address of principal executive offices)   (Zip Code)

Registrant’s telephone number, including area code: (775) 525-6450 

Not Applicable

(Former name or former address, if changed since last report)

 

 

Check the appropriate box below if the Form 8-K filing is intended to simultaneously satisfy the filing obligation of the registrant under any of the following provisions (see General Instruction A.2. below):

 

 

Written communications pursuant to Rule 425 under the Securities Act (17 CFR 230.425)

 

 

Soliciting material pursuant to Rule 14a-12 under the Exchange Act (17 CFR 240.14a-12)

 

 

Pre-commencement communications pursuant to Rule 14d-2(b) under the Exchange Act (17 CFR 240.14d-2(b))

 

 

Pre-commencement communications pursuant to Rule 13e-4(c) under the Exchange Act (17 CFR 240.13e-4(c))

Securities registered pursuant to Section 12(b) of the Act:

 

Title of each class

 

Trading Symbol(s)

 

Name of each exchange on which registered

Common Shares   IAUX   NYSE American LLC
Common Shares   IAU   The Toronto Stock Exchange

Indicate by check mark whether the registrant is an emerging growth company as defined in Rule 405 of the Securities Act of 1933 (§230.405 of this chapter) or Rule 12b-2 of the Securities Exchange Act of 1934 (§240.12b-2 of this chapter).

Emerging growth company ☒

If an emerging growth company, indicate by check mark if the registrant has elected not to use the extended transition period for complying with any new or revised financial accounting standards provided pursuant to Section 13(a) of the Exchange Act. ☐

 

 

 


Item 7.01

Regulation FD Disclosure

The Company has prepared the following technical report summaries which are attached hereto as exhibits 99.1, 99.2, 99.3, and 99.4, respectively:

 

   

S-K 1300 Initial Assessment & Technical Report Summary, Ruby Hill Complex, Eureka County, Nevada dated March 27, 2025

 

   

S-K 1300 Initial Assessment & Technical Report Summary, Lone Tree Deposit, Nevada dated February 24, 2025

 

   

S-K 1300 Initial Assessment & Technical Report Summary, Granite Creek Mine, Humboldt County, Nevada dated March 27, 2025

 

   

S-K 1300 Initial Assessment & Technical Report Summary for the Cove Project, Lander County, Nevada dated March 26, 2025

The information in this Item 7.01 and exhibits attached hereto are being furnished and shall not be deemed filed for purposes of Section 18 of the Securities Exchange Act of 1934, as amended (the “Exchange Act”), or otherwise subject to the liability of that section, and shall not be incorporated by reference into any registration statement or other document filed under the Securities Act of 1933, as amended, or the Exchange Act, except as shall be expressly set forth by specific reference in such filing.

 

Item 9.01

Financial Statements and Exhibits.

(d) Exhibits

 

Exhibit
No.
  Description
99.1   S-K 1300 Technical Report Summary, Initial Assessment for the Ruby Hill Complex, Eureka County, Nevada effective December 31, 2024
99.2   S-K 1300 Technical Report Summary, Initial Assessment for the Lone Tree Deposit, Nevada effective December 31, 2024
99.3   S-K 1300 Technical Report Summary, Initial Assessment for the Granite Creek Mine, Humboldt County, Nevada effective December 31, 2024
99.4   S-K 1300 Technical Report Summary, Initial Assessment for the Cove Project, Lander County, Nevada effective December 31, 2024
104   Cover Page Interactive Data File (embedded within the Inline XBRL document).


SIGNATURES

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned hereunto duly authorized.

 

 Date: March 31, 2025

   
    i-80 GOLD CORP.
   

By:

 

/s/ Ryan Snow

   

Name:

 

Ryan Snow

   

Title:

 

Chief Financial Officer

EX-99.1 2 d913666dex991.htm EX-99.1 EX-99.1

Exhibit 99.1

LOGO

S-K 1300 Technical Report Summary Initial Assessment of the Ruby Hill Project, Eureka County NV i-80 Gold Corp. Prepared by: FORTE DYNAMICS, INC 120 Commerce Drive, Units 3 & 4 Fort Collins, Colorado 80524 Prepared for: i-80 Gold Corp. 5190 Neil Road, Suite 460 Reno, Nevada 89502 Others: Practical Mining LLC TR Raponi Consulting Ltd. Effective Date: December 31, 2024 Issue Date: March 29, 2025 QP Firms: Practical Mining LLC TR Raponi Consulting Ltd. Forte Dynamics, Inc.


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   March 29, 2025
 

 

VERSION CONTROL

 

Rev. No     Date    Status    Prepared By    Checked By    Approved By

REV A

   3/19/2025    DRAFT    K. Ollila    J. Heiner    A. Amoroso

REV B

   3/25/2025    DRAFT    K. Ollila    J. Heiner    A. Amoroso

REV C

   3/29/2025    FINAL    K. Ollila    J. Heiner    A. Amoroso
                          

 

 

FORTE DYNAMICS, INC.

   P a g e | 2 of 362    i-80 Gold Corp.


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   March 29, 2025
 

 

DATE AND SIGNATURE PAGE

S-K 1300 Initial Assessment and Technical Report Summary, Ruby Hill Complex, Eureka County, Nevada

Prepared for: i-80 Gold Corporation

Report Date: March 29, 2025

Prepared by the following Firms:

 

       
QP Firm    Responsibilities/Contributions    Signature    Date
       

Practical Mining LLC

   1-9, 11.1, 11.2, 12, 13.1, 15.1, 16, 17, 18.1, 19.1, 20-25         March 28, 2025
       

TR Raponi Consulting Ltd.

   1, 10, 14.1, 14.2, 22-24         March 28, 2025
       

Forte Dynamics, Inc.

   1, 2, 11.1, 11.3, 11.4, 12, 13.2, 13.3, 14.3, 15.2, 18.2, 19.2, 21-25         March 28, 2025

 

 

FORTE DYNAMICS, INC.

   P a g e | 3 of 362    i-80 Gold Corp.


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   March 29, 2025
 

 

Table of Contents

 

Date and Signature Page

     3  

1.  Executive Summary

     17  

1.1   Introduction

     17  

1.2   Property Description

     19  

1.3   Geology and Mineral Deposits

     20  

1.3.1   Distal Disseminated

     20  

1.3.2   Carlin Type

     20  

1.3.3   CRD and Skarn

     21  

1.4   Metallurgical Testing and Processing

     22  

1.4.1   Archimedes Underground

     22  

1.4.2   Mineral Point Open Pit

     22  

1.5   Mineral Resources

     23  

1.5.1   Archimedes Underground

     23  

1.5.2   Archimedes Open Pit

     24  

1.5.3   Mineral Point Open Pit

     25  

1.6   Mining, Infrastructure, and Project Schedule

     26  

1.6.1   Archimedes Underground

     26  

1.6.2   Archimedes Open Pit

     26  

1.6.3   Mineral Point Open Pit

     26  

1.7   Economic Analysis

     27  

1.7.1   Archimedes Underground

     27  

1.7.2   Mineral Point Open Pit

     28  

1.8   Conclusions

     30  

1.8.1   Archimedes Underground

     30  

1.8.2   Archimedes Open Pit

     31  

1.8.3   Mineral Point Open Pit

     31  

1.9   Recommendations

     32  

1.9.1   Archimedes Underground

     32  

1.9.2   Archimedes Open Pit

     32  

1.9.3   Mineral Point Open Pit

     32  

1.9.4   Work Programs

     34  

2.  Introduction

     36  

2.1   Registrant for Whom the Technical Report Summary was Prepared

     36  

2.2   Terms of Reference and Purpose of this Technical Report

     36  

2.3   Qualified Persons

     36  

2.4   Details of Personal Inspection by Qualified Persons

     37  

2.5   Report Version Update

     37  

2.6   Units of Measure

     37  

2.7   Coordinate System

     38  

2.8   Mineral Resource and Mineral Reserve Definitions

     39  

3.  Property Description

     40  

3.1   Property Description

     40  

3.2   Status of Mineral Titles

     40  

3.3   Royalties

     47  

3.4   Environmental Liabilities

     49  

3.5   Permits/Licenses

     49  
  

 

 

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4.  Accessibility, Climate, Local Resources, Infrastructure, and Physiography

     50  

4.1   Accessibility

     50  

4.2   Climate

     50  

4.3   Local Resources

     50  

4.4   Infrastructure

     50  

4.5   Physiography

     50  

5.  History

     52  

5.1   Historic Ownership

     52  

5.2   Historic Mining

     54  

5.3   Historic Exploration

     55  

6.  Geologic Setting, Mineralization and Deposit Types

     60  

6.1   Regional Geology

     60  

6.2   Project Geology

     62  

6.3   Stratigraphy

     65  

6.3.1   Lower Cambrian

    

65

 

6.3.2   Middle Cambrian

    

65

 

6.3.3   Upper Cambrian

    

66

 

6.3.4   Lower-Middle Ordovician

    

66

 

6.3.5   Cretaceous

    

67

 

6.3.6   Tertiary/Quaternary

    

68

 

6.4   Structure

    

69

 

6.4.1   Archimedes Deposit Structure

    

70

 

6.4.2   Mineral Point Trend Structure

    

71

 

6.4.3   Historic Ruby Hill and FAD Structure

    

73

 

6.5   Alteration

    

74

 

6.5.1   Archimedes Deposit Alteration

    

74

 

6.5.2   Mineral Point Trend Alteration

    

74

 

6.5.3   Historic Ruby Hill and FAD Alteration

    

75

 

6.6   Mineralization

    

75

 

6.6.1   Archimedes Deposit Mineralization

    

78

 

6.6.2   Mineral Point Trend Mineralization

    

79

 

6.6.3   Historic Ruby Hill and FAD Mineralization

    

79

 

6.7   Deposit Types

    

80

 

6.7.1   Characteristics of Polymetallic Carbonate Replacement Deposits

    

80

 

6.7.2   Characteristics of Skarn Deposits

    

80

 

6.7.3   Characteristics of Carlin-Type Gold Deposits

    

80

 

6.7.4   Distal-disseminated Mineralization at Ruby Hill

    

81

 

7.  Exploration

     82  

7.1   Geophysical

     82  

7.1.1   Archimedes Area

    

82

 

7.1.2   FAD Area

    

83

 

7.2   Drilling

     85  

7.2.1   Historic Drilling at Ruby Hill

    

85

 

7.2.2   Drilling Methods

    

89

 

7.2.3   Geological Logging

    

91

 

7.2.4   Sample Recovery

    

91

 

7.2.5   Collar Surveys

    

91

 

7.2.6   Downhole Surveys

    

91

 

7.2.7   Metallurgical Drilling

    

92

 

 

 

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7.2.8   Sample Length/True Thickness

     94  

7.2.9   Potential Downhole Contamination

     94  

7.2.10  Summary and Interpretation of All Relevant Drilling Results

     94  

7.2.11  i-80 Drilling

     94  

7.3   Hydrogeology

     95  

7.3.1   Sampling Methods and Laboratory Determinations

     95  

7.3.2   Hydrogeology Investigations

     95  

7.3.3   Hydrogeologic Description

     96  

7.3.4   Mine Dewatering

     101  

7.3.5   Dewatering Discharge

     101  

7.3.6   Groundwater Flow Model

     102  

7.3.7   Model Results

     105  

8.  Sample Preparation, Analysis and Security

     111  

8.1   Sampling Methods

     111  

8.2   Analytical and Test Laboratories

     111  

8.3   Density Determinations

     111  

8.4   Sample Preparation and Analysis

     113  

8.4.1   Barrick

     113  

8.4.2   Homestake

     114  

8.5   Quality Assurance and Quality Control (QA/QC)

     115  

8.5.1   Barrick QA/QC Program

     115  

8.6   Historical Databases

     121  

8.7   Historical Sample Security

     121  

8.8   Comments on Historic Ruby Hill Data

     121  

8.9   i-80 Sample Preparation, Laboratory Analysis, Security, and Quality Control Procedures

    

122

 

8.9.1   i-80 Sample Preparation Procedures

     122  

8.10  i-80 Standards and Blanks

     123  

8.11  i-80 Duplicate Assays

     123  

8.12  QP Opinion

     124  

9.  Data Verification

     126  

9.1   Historical Data Review

     126  

9.2   Wood Data Verification 2021

     126  

9.3   Practical Mining Data Verification 2023

     126  

10.  Mineral Processing and Metallurgical Testing

     128  

10.1 Archimedes Underground

     128  

10.1.1  Refractory Testing Programs

     128  

10.1.2  Deleterious Elements

     133  

10.1.3  Recovery Estimates

     134  

10.2 Mineral Point Open Pit

     135  

10.2.1  Historical Operations

     135  

10.2.2  Historical Test Work

     135  

10.2.3  Mineral Point Leach Cycle Times

     141  

10.2.4  Mineral Point Reagent Consumptions

     141  

10.2.5  Deleterious Elements

     141  

10.2.6  Recovery Estimates

     142  

11.  Mineral Resource Estimates

     144  

11.1 Introduction

     144  
  

 

 

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11.2 Archimedes Underground

     145  

11.2.1  Grade Shells

     146  

11.2.2  Density

     146  

11.2.3  Statistics

     148  

11.2.4  Grade Capping

     152  

11.2.5  Block Model

     153  

11.2.6  Model Validation

     153  

11.2.7  Resource Classification

     159  

11.2.8  Factors That May Affect Mineral Resources

     159  

11.2.9  Reasonable Prospects for Eventual Economic Extraction

     159  

11.2.10  Archimedes Underground Mineral Resource Statement

     159  

11.2.11  QP Opinion

     161  

11.3 Archimedes Open Pit

     161  

11.3.1  Summary Workflow

     161  

11.3.2  Exploratory Data Analysis (EDA)

     161  

11.3.3  Resource Estimation

     165  

11.3.4  Model Validation

     166  

11.3.5  Mineral Resource Classification

     170  

11.3.6  Reasonable Prospects for Eventual Economic Extraction

     170  

11.3.7  Archimedes Open Pit Mineral Resource Statement

     170  

11.3.8  QP Opinion

     171  

11.4 Mineral Point Open Pit

     171  

11.4.1  Summary Workflow

     172  

11.4.2  Geological Modeling

     172  

11.4.3  Exploratory Data Analysis

     174  

11.4.4  Grade Estimation

     177  

11.4.5  Resource Model Validation

     180  

11.4.6  Bulk Density

     183  

11.4.7  Mineral Resource Classification

     183  

11.4.8  Reasonable Prospects for Eventual Economic Extraction

     184  

11.4.9  Mineral Point Open Pit Mineral Resource Statement

     185  

11.4.10  Factors that may Affect Mineral Resources

     185  

11.4.11  QP Opinion

     186  

12.  Mineral Reserve Estimates

     187  

13.  Mining Methods

     188  

13.1 Archimedes Underground

     188  

13.1.1  Mine Development

     188  

13.1.2  Mining Methods

     189  

13.1.3  Geotechnical and Ground Support

     191  

13.1.4  Cemented Rock Fill

     200  

13.1.5  Staffing and Underground Equipment Requirements

     200  

13.1.6  Mine Plan

     201  

13.2 Archimedes Open Pit

     206  

13.3 Mineral Point Open Pit

     206  

13.3.1  Initial Pit Limit Evaluations

     206  

13.3.2  Open Pit Economic Parameters

     207  

13.3.3  Pit Designs

     215  

13.3.4  Haul Road Design

     227  

13.3.5  Economic Evaluation

     228  

13.3.6  Cutoff Grade

     228  

13.3.7  Pit Design Inventories

     228  

13.3.8  Drilling and Blasting

     229  

13.3.9  Production Schedules

     229  

 

 

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13.3.10  Mine Fleet

     231  

13.3.11  Dewatering

     231  

14.  Recovery Methods

     232  

14.1 Archimedes Underground

     232  

14.1.1  Introduction

     232  

14.1.2  Refractory Mineralization Processing

     232  

14.2 Lone Tree Pressure Oxidation Facility

     234  

14.2.1  Lone Tree Mill Historic Processing

     234  

14.2.2  Lone Tree Facility Block Flow Diagram

     235  

14.2.3  Key Design Criteria

     237  

14.2.4  Lone Tree Facility Description

     237  

14.2.5  Utilities Consumption

     241  

14.3 Mineral Point Open Pit

     243  

14.3.1  Summary Process Design Criteria

     243  

14.3.2  Process Descriptions

     245  

14.3.3  Process Water

     247  

14.3.4  Process Flowsheet

     247  

15.  Infrastructure

     249  

15.1 Archimedes Underground

     249  

15.1.1  Operations Dewatering

     249  

15.1.2  Operations Monitoring Wells and VWPs

     249  

15.1.3  Operations RIBs

     249  

15.1.4  Operations Water Supply

     249  

15.1.5  Electrical Power

     249  

15.1.6  Underground Mine Facilities

     249  

15.1.7  Backfill

     250  

15.2 Mineral Point Open Pit

     256  

15.2.1  Site Layout

     256  

15.2.2  Existing Infrastructure

     258  

15.2.3  Planned Infrastructure

     259  

15.2.4  Operations Dewatering

     264  

15.2.5  Operations Monitoring

     267  

15.2.6  Water Supply

     267  

16.  Market Studies and Contracts

     268  

16.1 Precious Metal Markets

     268  

16.2 Contracts

     269  

16.2.1  Financing Agreements

     269  

16.3 Refractory Mineralized Material Sale Agreement

     271  

16.4 Other Contracts

     271  

17.  Environmental Studies, Permitting and Social or Community Impact

     272  

17.1 Closure and Reclamation Requirements

     272  

17.2 Social or Community Impacts

     272  

17.3 Permits

     273  

17.4 Water Use Permits

     274  

17.5 QP Opinion

     274  

18.  Capital and Operating Costs

     275  

18.1 Archimedes Underground

     275  
  

 

 

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18.1.1  Capital Costs

     275  

18.1.2  Operating Costs

     275  

18.1.3  Cutoff Grade

     276  

18.2 Mineral Point Open Pit

     277  

18.2.1  Capital Cost Estimate

     277  

18.2.2  Operating Cost Estimate

     279  

19.  Economic Analysis

     281  

19.1 Archimedes Underground

     281  

19.1.1  Taxes

     281  

19.1.2  Cash Flow

     281  

19.1.3  Sensitivity

     291  

19.2 Mineral Point Open Pit

     293  

19.2.1  Principal Assumptions

     293  

19.2.2  Operating Cost

     293  

19.2.3  Capital Costs

     293  

19.2.4  Cost Summary

     294  

19.2.5  Economic Model

     294  

19.2.6  Economic Analysis Without Inferred Resources

     301  

20.  Adjacent Properties

     302  

21.  Other Relevant Data and Information

     303  

22.  Interpretation and Conclusions

     304  

22.1 Conclusions

     304  

22.1.1  Archimedes Underground

     304  

22.1.2  Archimedes Open Pit

     305  

22.1.3  Mineral Point Open Pit

     305  

22.2 Risks and Opportunities

     305  

23.  Recommendations

     309  

23.1 Archimedes Underground

     309  

23.1.1  Metallurgical Testing

     309  

23.1.2  Permitting and Mine Development

     309  

23.1.3  Resource Conversion and Exploration Drilling

     309  

23.1.4  Dewatering

     309  

23.2 Archimedes Open Pit

     309  

23.2.1  Mineral Resources

     309  

23.3 Mineral Point Open Pit

     309  

23.3.1  Mineral Resources

     309  

23.3.2  Mining and Infrastructure

     310  

23.3.3  Metallurgical Testing

     310  

23.4 Work Program

     311  

23.4.1  Archimedes Underground

     311  

23.4.2  Archimedes Open Pit

     312  

23.4.3  Mineral Point Open Pit

     312  

24.  References

     313  

25.  Reliance on Information Provided by the Registrant

     320  

 

 

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Figures

 

Figure 1-1: Ruby Hill Complex Overview      18  
Figure 3-1: Ruby Hill Complex Location Map      40  
Figure 3-2: Ruby Hill Complex Land Position      41  
Figure 3-3: Ruby Hill Royalty Map      48  
Figure 5-1: Geophysical Surveys in the Ruby Hill Project Area      57  
Figure 5-2: Rock Samples with Gold Grade (opt) within the Ruby Hill Claim Block      58  
Figure 5-3: Soil Samples with Gold Grade (opt) within the Ruby Hill Claim Block      59  
Figure 6-1: Regional Geologic Map      61  
Figure 6-2: Ruby Hill Project Geology and Deposit Locations      64  
Figure 6-3: Ruby Hill Stratigraphic Column      68  
Figure 6-4: Geology of East Archimedes, West Archimedes and Archimedes Underground Including 426, and Ruby Deeps Zones      71  
Figure 6-5: Mineral Point Trend Geology      72  
Figure 6-6: Historic Ruby Hill and FAD Deposit Geology      73  
Figure 6-7: Plan View of Ruby Deeps, 426, 007, 008, Blackjack, and Hilltop Zones      76  
Figure 6-8: Plan View of Mineral Point Trend and Archimedes Deposits      77  
Figure 6-9: Fence Section of Mineral Point Trend and Archimedes Deposits      78  
Figure 7-1: Exploration Targets at Ruby Hill      84  
Figure 7-2: Drill Hole Collar Locations      86  
Figure 7-3: Distribution of Drill Types Included in the 2021 Ruby Hill Project Mineral Resource Estimate      87  
Figure 7-4: Plan View of Drilling by Campaign      88  
Figure 7-5: Fence Section of Drilling by Campaign (Looking North)      89  
Figure 7-6: Diamond Valley Hydrographic Basin and Ruby Hill Mine Permit Area      98  
Figure 7-7: Surface Geology and Pre-Mining Groundwater Level Contours      99  
Figure 7-8: Dewatering Well and Groundwater Monitoring Locations      102  
Figure 7-9: Property Overview showing Plan Operations Boundary, Existing Mine Operations Boundary, and Existing Archimedes Pit with Planned UGWs for the 426 and Blackjack Deposits      103  
Figure 7-10: Schematic Section through the Archimedes Pit Area      104  
Figure 7-11: Ground Water Flow Model Boundary      106  
Figure 7-12: Ground Water Flow Model Grid      107  
Figure 7-13: Mine-Area Hydrogeologic Zones and Flow Barriers, Layer 2      108  
Figure 7-14: Projected Changes in Groundwater Level, End of Mining      109  
Figure 8-1: Control Chart for Standard OREAS 54PA      116  
Figure 8-2: ALS Global (Chemex) Pulps Checked at Inspectorate      118  
Figure 8-3: Mean Versus Half Relative Difference for Field Duplicates      119  
Figure 8-4: Scatter Plot of all Lab Duplicates      120  
Figure 8-5: Mean Versus Half Relative Difference for Pulp Duplicates      120  
Figure 8-6: i-80 Lab Duplicates      124  
Figure 8-7: i-80 Prep Duplicates      125  
Figure 10-1: 2024 FLSmidth Program Preg-Robbing as a Function of Organic Carbon Concentration      131  
Figure 11-1: Block Model Extents      145  
Figure 11-2: Underound Model Extents and Drill Hole Traces      146  
Figure 11-3: Density Box and Whisker Plot by Lithology Formation      147  
Figure 11-4: 426 0.1 Au opt Box and Whisker Plots      148  
Figure 11-5: Ruby Deeps 0.1 Au opt Box and Whisker Plots      149  

Figure 11-6: 0.002 Au opt Box and Whisker Plots

     151  

 

 

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Figure 11-7: Gold Cumulative Frequency      152  
Figure 11-8: Silver Grade Shells Cumulative Frequency      152  
Figure 11-9: 426 Deposit Comparison of Composite and Block Grades      155  
Figure 11-10: Ruby Deeps Deposit Comparison of Composite and Block Grades 120450N      156  
Figure 11-11: Drift Analysis Gold      157  
Figure 11-12: Drift Analysis Silver      158  
Figure 11-13: Graphical Statistical Comparison of Rock Units      162  
Figure 11-14: Statistics for Key Geological Units      163  
Figure 11-15: Composite Study Results      164  
Figure 11-16: Gold and Silver Composite Samples within the Indicator Shell      165  
Figure 11-17: Example Gold Variogram      165  
Figure 11-18: Cross Section of Estimated Block Model and Composites      167  
Figure 11-19: Example Swath Plots      168  
Figure 11-20: Comparison of Cumulative Frequency      169  
Figure 11-21: Fence Section Looking North Showing Main Faults and Stratigraphic Units for the Ruby Hill Project      173  
Figure 11-22: Example Cross Section Showing Modeled Sulfide Domain and Redox Codes in the Drillhole Database      174  
Figure 11-23: Gold and Silver Raw Assay Sample Grade Histograms and Probability Plots      175  
Figure 11-24: Example Cross Section of the Mineral Point Trend Showing Raw Assays (Right of Trace) and Downhole 10 ft. Composites (Left Trace) with the Optimized Pit Shell      176  
Figure 11-25: Box and Whisker Plot for Assay Sample Grades and 10 ft. Composites for Gold and Silver      177  
Figure 11-26: Indicator Threshold Selection – CV of Gold and Silver Assay Composite Grades      178  
Figure 11-27: Au Estimation – Implementation of a Soft Boundary Between LG and HG Composites      179  
Figure 11-28: Area of Au High-Grade Blow-out and Eureka Corp Underground Drilling      180  
Figure 11-29: Estimated Block Grades and 10 Foot Composite Grades for Gold - Section 121200 N Looking N      181  
Figure 11-30: Swath Plots – Gold – Indicated Blocks      182  
Figure 11-31: Bulk Density Values by Lithology      183  
Figure 11-32: Cross Section Showing the Mineral Point Resource, Resource Pit Shell, and Topo      184  
Figure 13-1: Archimedes Underground Isometric View Showing Portals, Main Ramp and Ventilation Development      188  
Figure 13-2: Stope Mining Sequence Part A      190  
Figure 13-3: Stope Mining Sequence Part B      191  
Figure 13-4: RQD Logged Drill Holes (426 - Turquoise, Ruby Deeps - Gold)      192  
Figure 13-5: Cross Section 119625N Showing RQD Values (426 - Turquoise, 426 Fault - Gray, Ruby Deeps - Gold, Holly Fault - Red)      193  
Figure 13-6: RQD Box and Whisker Plot      194  
Figure 13-7: Q-system Support Recommendations      196  
Figure 13-8: Q Logged Drill Holes (426 - Blue, Ruby Deeps - Gold)      197  
Figure 13-9: Cross Section 119625N Showing Q Values (426 - Blue, 426 Fault - Gray, Ruby Deeps - Gold, Holly Fault - Red)      198  
Figure 13-10: Q Value Box and Whisker Plot      199  
Figure 13-11: Permitting Development and Initial Production Schedule      202  
Figure 13-12: LG Shells by Revenue Factor      209  
Figure 13-13: Percentage of Profit, Processed Material, and Recoverable Gold by LG Shell      211  

Figure 13-14: Plan View of LG Pit Shells and Cross Section Locations

     212  

 

 

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Figure 13-15: Pit Optimization Looking West (Section A’ – A)

     213  

Figure 13-16: Pit Optimization Looking North (Section B’ – B)

     213  

Figure 13-17: Pit Optimization Looking North (Section C’ – C)

     214  

Figure 13-18: Pit Optimization Looking North (Section D’ – D)

     214  

Figure 13-19: Pit Optimization Looking North (Section E’ – E)

     215  

Figure 13-20: Pit Phasing and Section Line

     216  

Figure 13-21: Cross Section F’ to F of Pit Phasing

     217  

Figure 13-22: Phase 1 Design

     218  

Figure 13-23: Phase 2 Design

     219  

Figure 13-24: Phase 3 Design

     220  

Figure 13-25: Phase 4 Design

     221  

Figure 13-26: Phase 5 Design (First Phase of Heap Leach Relocation)

     222  

Figure 13-27: Phase 6 Design (Second Phase of Heap Leach Relocation)

     223  

Figure 13-28: Phase 7 Design

     224  

Figure 13-29: Phase 8 Design

     225  

Figure 13-30: Phase 9 Design

     226  

Figure 13-31: Final Pit and Estimated Block Model in Orthogonal View Looking Northwest

     227  

Figure 13-32: LOM Annual Production Schedule

     230  

Figure 14-1: Third Party POX Facility Simplified Flowsheet

     233  

Figure 14-2: Loan Tree Block Flow Diagram

     236  

Figure 14-3: Mineral Point Process Flowsheet

     248  

Figure 15-1: Portal Surface Facilities Conceptual Layout

     250  

Figure 15-2: Site Layout Map

     257  

Figure 15-3: Existing Infrastructure

     258  

Figure 15-4: Hydrologic Blocks of Mineral Point

     265  

Figure 16-1: Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average

     268  

Figure 19-1: Mineralization Mined and Processed with Inferred

     286  

Figure 19-2: Gold Production and Unit Costs with Inferred

     286  

Figure 19-3: Mineralization Mined and Processed without Inferred

     287  

Figure 19-4: Gold Production and Unit Costs without Inferred

     287  

Figure 19-5: Cash Flow Waterfall Chart with Inferred

     289  

Figure 19-6: NPV 5% Sensitivity with Inferred

     291  

Figure 19-7: NPV 8% Sensitivity with Inferred

     291  

Figure 19-8: IRR Sensitivity with Inferred

     292  

Figure 19-9: Profitability Index Sensitivity with Inferred

     292  

Figure 19-10: Pre-Tax LOM Annual Cash Flow

     295  

Figure 19-11: After-Tax LOM Annual Cash Flow

     296  

Figure 19-12: Pre-Tax Sensitivity NPV @5%

     298  

Figure 19-13: Pre-Tax Sensitivity IRR

     298  

Figure 19-14: After-Tax Sensitivity NPV @5%

     299  

Figure 19-15: After-Tax Sensitivity IRR

     299  

 

 

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Tables

 

Table 1-1: Summary of Archimedes Underground Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      23  
Table 1-2: Summary of Archimedes Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      24  
Table 1-3: Summary of Mineral Point Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      25  
Table 1-4: Capital and Operating Cost Summary      27  
Table 1-5: Financial Statistics      28  
Table 1-6: Unit and Total Operating Costs With and Without Inferred Resources      29  
Table 1-7: After-Tax NPV Comparison of With and Without Inferred Resources      30  
Table 1-8: Archimedes Underground Work Program      34  
Table 1-9: Mineral Point Work Program      35  
Table 2-1: Personal Inspections by Qualified Persons      37  
Table 2-2: Units and Abbreviations      38  
Table 3-1: Ruby Hill Project Owned Patented Claims      42  
Table 3-2: Ruby Hill Project Owned Unpatented Claims      43  
Table 3-3: Ruby Hill Project Leased Unpatented Claims      44  
Table 3-4: Golden Hill FAD Property Owned Patented Claims      44  
Table 3-5: Golden Hill FAD Property Leased Patented Claims      45  
Table 3-6: Golden Hill FAD Property Owned Unpatented Claims      46  
Table 3-7: Golden Hill FAD Property Leased Unpatented Claims      46  
Table 3-8: Ruby Hill Complex Property Holding Costs      47  
Table 3-9: Ruby Hill Royalties      47  
Table 3-10: Golden Hill Royalties      47  
Table 5-1: Historic Regional Ownership and Activities      52  
Table 5-2: Production History Summary      55  
Table 5-3: Historic Exploration      56  
Table 6-1: Major Structural Features and Orientations within the Property Area      69  
Table 7-1: Drilling Statistics for Drillholes Included in the 2021 Ruby Hill Project Mineral Resource Estimate      87  
Table 7-2: Distribution of Drilling by Campaign      87  
Table 7-3: 2004 Barrick Metallurgical Holes      92  
Table 7-4: 2009 Metallurgical Holes      93  
Table 7-5: 2010 and 2011 Metallurgical Holes      93  
Table 7-6: 2011 Metallurgical Holes      93  
Table 7-7: Summary of Hydrogeological Surveys Since 2004 (Wood 2021)      96  
Table 8-1: Assay, Density and Metallurgical Laboratories      112  
Table 8-2: Barrick Rock Type Density Values      113  
Table 8-3: ALS Global Gold Analytical Parameters      114  
Table 8-4: Count and Description of QA/QC Samples by Year      116  
Table 8-5: SRM Performance      117  
Table 8-6: Selected i-80 Blank and Standard Reference Results      123  
Table 9-1: Drill Holes in 426 and Ruby Deeps Zones      127  
Table 9-2: Drillhole Data Fields Reviewed      127  
Table 10-1: Ruby Hill Project Refractory Testing Programs      128  

Table 10-2: January 2008 426 Zone Barrick Technology Centre Test Results Summary

     129  

 

 

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Table 10-3: November 2011 426 Zone Barrick Technology Centre Test Results Summary      130  
Table 10-4: 2024 FLSmidth Program Assays Summary      130  
Table 10-5: FLSmidth Program BTAC Conditions Summary      132  
Table 10-6: FLSmidth Program BTAC and Roasting CIL Recovery Summary      132  
Table 10-7: November 2011 426 Zone Barrick Technology Centre Refractory Sample Assays Summary      134  
Table 10-8: Ruby Hill (Archimedes) Summary of Estimated Gold Recoveries      134  
Table 10-9: Ruby Hill Project Historical Metallurgical Testing Programs      135  
Table 10-10: June 2004 KCA Archimedes Column Test Results Summary      136  
Table 10-11: May 2005 KCA Archimedes Column Test Results Summary      136  
Table 10-12: January 2009 426 Zone KCA Column Leach Test Results Summary      137  
Table 10-13: November 2011 426 Zone KCA Column Leach Test Results Summary      137  
Table 10-14: February 2011 Mineral Point Deposit KCA Bottle Rolls Test Results Summary      138  
Table 10-15: February 2011 Mineral Point Deposit KCA Column Leach Test Results Summary      138  
Table 10-16: July 2012 Mineral Point Deposit KCA Bottle Rolls Test Results Summary      139  
Table 10-17: July 2012 Mineral Point Deposit KCA Column Leach Test Results Summary      139  
Table 10-18: February 2014 Mineral Point Deposit KCA Bottle Rolls Test Results Summary      140  
Table 10-19: February 2014 Mineral Point Deposit KCA Column Leach Test Results Summary      141  
Table 10-20: Summary of Column Leach Test Results      142  
Table 10-21: Mineral Point Summary of Estimated Gold and Silver Recoveries      143  
Table 11-1: Univariate Density Statistics by Lithology Formation (tonnes/m3)      147  
Table 11-2: Gold Univariate Statistics for 426 0.1 Au opt Composites      148  
Table 11-3: Silver Univariate Statistics for 426 0.1 Au opt Composites      149  
Table 11-4: Gold Univariate Statistics for Ruby Deeps 0.01 Au opt Composites      150  
Table 11-5: Silver Univariate Statistics for Ruby Deeps 0.01 Au opt Composites      150  
Table 11-6: Gold Univariate Statistics for 0.002 Au opt Composites      151  
Table 11-7: Silver Univariate Statistics for 0.002 Au opt Composites      151  
Table 11-8: Gold and Silver Grade Caps      153  
Table 11-9: Estimation Search Distances and Sample Requirements      153  
Table 11-10: Ellipsoid Search Parameters for each Grade Shell      153  
Table 11-11: Comparison of Composite and Block Model Statistics      154  
Table 11-12: Mineral Resource Classification Scheme      159  
Table 11-13: Summary of Archimedes Underground Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      160  
Table 11-14: Summary Sample Statistics - Archimedes      162  
Table 11-15: Variogram for 0.05 Au ppm Indicator      164  
Table 11-16: Variograms for Au and Ag      166  
Table 11-17: Gold and Silver Search Parameters      166  
Table 11-18: Resource Classification by Sample Density      170  
Table 11-19: Summary of Archimedes Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      171  
Table 11-20: Estimation Parameters      179  
Table 11-21: Global Bias Check within Indicated Resources      181  
Table 11-22: Parameters for Mineral Resource Pit Shell Construction      184  
Table 11-23: Summary of Mineral Point Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024      185  
Table 13-1: Guidelines for the Selection of Primary Support for 20-foot to 40-foot Tunnels in Rock      192  

Table 13-2: RQD Univariate Statistics by Grade Shell

     194  

 

 

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Table 13-3: Q Value Univariate Statistics by Grade Shell      199  
Table 13-4: Personnel Requirements      200  
Table 13-5: Equipment Requirements      201  
Table 13-6: i-80 Support Equipment      201  
Table 13-7: Ruby Hill Development Schedule      202  
Table 13-8: Archimedes Production Mining Plan (Includes Inferred Mineral Resource)      203  
Table 13-9: Mine Production Rates by Excavation Type      203  
Table 13-10: Ruby Hill Processing Plan (Includes Inferred Mineral Resource)      204  
Table 13-11: Ruby Hill Processing Plan (Without Inferred Mineral Resource)      205  
Table 13-12: Pit Slope by Lithology Unit      207  
Table 13-13: Pit Optimization Parameters      207  
Table 13-14: Profit Factor for Optimization Results      210  
Table 13-15: Pit Design Parameters      215  
Table 13-16: Design Metal Prices, Costs, and Recoveries      228  
Table 13-17: In-Pit Mineral Resources by Pit Phase      229  
Table 13-18: LOM Production Schedule      230  
Table 13-19: Mining Equipment List      231  
Table 14-1: Summary of Key Process Statistics      237  
Table 14-2: Lone Tree Facility Water Consumption by Type      242  
Table 14-3: Lone Tree Facility Energy Usage by Area      242  
Table 14-4: Mineral Point Design Criteria      243  
Table 15-1: Ruby Hill Active Dewatering Wells (LRE 2025)      251  
Table 15-2: Summary of Locations, Construction Information, and Water Levels for Dewatering Wells, VWPs, Monitoring Wells, and Piezometers      252  
Table 15-3: Existing Infrastructure Plans      259  
Table 15-4: Heap Leach Pad Phases      260  
Table 15-5: WRSA Parameters      261  
Table 15-6: Ruby Hill Pumping Wells      266  
Table 17-1: Ruby Hill Project Significant Permits      273  
Table 18-1: Mine Development Unit Costs      275  
Table 18-2: Project Capital Costs ($M)      275  
Table 18-3: Underground Mine Operating Costs      276  
Table 18-4: Resource Cutoff Grades by Process      276  
Table 18-5: Mineral Point Project Capital Cost Summary      277  
Table 18-6: Mineral Point Mining Equipment LOM CAPEX      278  
Table 18-7: Mineral Point Process Infrastructure LOM CAPEX      278  
Table 18-8: Mineral Point Pre-Production and Facilities LOM CAPEX      279  
Table 18-9: Mineral Point Owner’s Costs LOM CAPEX      279  
Table 18-10: Mineral Point LOM Operating Cost Summary      280  
Table 18-11: Mineral Point Processing Costs      280  
Table 19-1: Income Statement with Inferred      282  
Table 19-2: Cash Flow Statement with Inferred      283  
Table 19-3: Income Statement without Inferred      284  
Table 19-4: Cash Flow Statement without Inferred      285  
Table 19-5: Capital and Operating Cost Summary With Inferred      288  
Table 19-6: Capital and Operating Cost Summary Without Inferred      288  
Table 19-7: Financial Statistics      290  
Table 19-8: Economic Model Parameters      293  

 

 

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Table 19-9: Cost Summary      294  
Table 19-10: Pre-Tax NPV Summary      295  
Table 19-11: After-Tax NPV Summary      295  
Table 19-12: Sensitivity Summary      297  
Table 19-13: Gold and Silver Price Sensitivity After-Tax Analysis      300  
Table 19-14: Economic Model Parameters Comparison of With and Without Inferred Resources      301  
Table 19-15: After-Tax NPV Comparison of With and Without Inferred Resources      301  
Table 22-1: Risks and Uncertainties      307  
Table 22-2: Opportunities      308  
Table 23-1: Archimedes Underground Work Program      311  
Table 23-2: Mineral Point Work Program      312  

Appendices

Appendix A – Site Visit Report

Appendix B – Mineral Point Open Pit Economic Model with Inferred Resources

 

 

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1.

EXECUTIVE SUMMARY

 

1.1

Introduction

This Technical Report Summary (“TRS”) dated the 29th day of March, 2025 with an effective date of December 31, 2024 provides an updated statement of Mineral Resources for Ruby Hill Mining Company LLC’s Ruby Hill Complex. This Technical Report Summary provides an Initial Assessment (“IA”) for the Archimedes Underground, Archimedes Open Pit, and Mineral Point Open Pit Resource areas at Ruby Hill.

The mining contemplated for each of these areas is independent of the other and there is no interaction between them. Exploitation of one area does not preclude exploitation of the other. This report considers each to be a stand-alone operation, and there has not been any shared benefit assigned to operating or capital costs.

The Ruby Hill property contains several historical mines, current resources, and exploration targets (Figure 1-1). The property is endowed with multiple types of mineralization, including Carlin-style gold, distal disseminated silver-gold, carbonate replacement deposits (CRD), and skarn base metals. i-80 is currently focused on precious metal deposits. The resources considered in this report include the Archimedes Underground Carlin-style gold deposit and the Mineral Point distal disseminated silver-gold deposit.

 

 

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Figure 1-1: Ruby Hill Complex Overview

(Source: i-80 Gold, 2023)

Cautionary Note:

The financial analysis contains certain information that may constitute “forward-looking information” under applicable United States securities legislation. Forward-looking information includes, but is not limited to, statements regarding

 

 

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the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law.

 

1.2

Property Description

The Ruby Hill property is located in the historic Eureka mining district. It lies west of the town of Eureka in Eureka County, central Nevada. It is a large property containing multiple deposit types and several past-producing mines, and as such it has a long history of prior ownership and production. The property takes its name from the most significant historical mine, the Ruby Hill mine, named for the hill it lies beneath roughly 1.2 miles southwest of Eureka. Historic mining generally exploited silver and base metal mineralization with the majority of production coming from the Ruby Hill mine from 1873-1916. Sporadic exploration and production from 1916 through 1959 included discovery and mining of the TL, Holly and Helen deposits roughly 1.2 miles north of Ruby Hill as well as further attempts to re-access the Ruby Hill mine and a largely as-yet unexploited deposit interpreted as a lower offset of the Ruby Hill deposit known as FAD. The FAD and Locan shafts were sunk to target depths but were plagued by unmanageable water inflows when crosscut mining intersected water bearing structures.

Modern mining began in 1992 with the discovery of the Archimedes Carlin-style gold deposit roughly 1.5 miles NNE of Ruby Hill and one mile NNW of Eureka. Archimedes has been mined using open pit methods from 1997-2002, 2007-2013, and 2020-2021. A pit wall failure in 2013 made continued large scale open pit mining unfeasible due to the economic environment at that time, but continued exploration delineated resources exploitable using underground mining methods. These resources are collectively called Archimedes Underground.

The mineral deposits being considered for economic extraction in this TR are the Archimedes Underground and the Mineral Point deposits.

The Mineral Point deposit was delineated by previous owners between 1992 and 2015 but has not been mined with the exception of limited areas at the southern end of the deposit exploited by the historic TL, Holly and Helen underground mines.

The property is located on owned fee land, owned and leased patented mining claims, and owned and leased unpatented mining claims. i-80 Gold purchased the northern portion of the Ruby Hill property, containing the Archimedes and Mineral Point deposits and small historic underground mines including TL, Holly and Helen from Waterton Global in 2021. The southern portion of the property, including the historic Ruby Hill mine and FAD deposit, was acquired by i-80 through a merger with Golden Hill Mining Corporation in 2022. The Ruby Hill complex comprises 10,608 acres from the Ruby Hill purchase and 3,229 acres from the Golden Hill Merger. i-80 differentiates the property for managerial/administrative purposes, referring to

 

 

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the northern portion as Ruby Hill and the southern portion as Golden Hill. Collectively they are known as the Ruby Hill Project or the Ruby Hill Complex.

 

1.3

Geology and Mineral Deposits

The Ruby Hill Project is located along the southeastern end of the Battle Mountain/Eureka gold trend. The Eureka gold mining district exposes a nearly continuous sequence of Cambrian and Ordovician sedimentary rocks approximately 10,000’ thick consisting primarily of carbonate units which are favorable for mineralization with subordinate shale and quartz sandstone.

Mineralization at Ruby Hill is characterized by intrusion-related distal-disseminated silver-gold, carbonate replacement base metal deposits, and skarn deposits that have been overprinted by younger Carlin-type gold mineralization. The main precious metal mineralization at Ruby Hill occurs in favorable lithostratigraphic units bound by high angle structures that are interpreted to have been conduits for hydrothermal fluids responsible for gold and silver mineralization. The earlier carbonate replacement base metal mineralization occurs in metamorphosed and skarn-altered limestone units proximal to Cretaceous intrusions.

 

1.3.1

Distal Disseminated

The Mineral Point deposit consists of gold and silver mineralization hosted by the Cambrian Hamburg dolomite in the nose of a broad anticline that plunges gently to the north-northwest and is bound to the east by the Holly Fault and to the west by the Spring Valley Fault. The Mineral Point Trend is 10,000 ft long, 2,400 ft wide and up to 500 ft thick. The top of the Mineral Point Trend is near surface at its south end and 500 ft below surface at its north end. The majority of the mineralization in the Mineral Point deposit is oxidized and has a high ratio of cyanide soluble to fire assay total gold. This deposit has not been mined and is the largest (and lowest grade) precious metal mineral resource in the Ruby Hill Project.

 

1.3.2

Carlin Type

The Archimedes deposit was discovered by Homestake Mining Company in 1992. The upper portions, called West and East Archimedes, were mined as the Archimedes open pit by Homestake followed by Barrick Gold Corporation from 1998 through 2015, and to a lesser extent by Ruby Hill Mining Company, LLC in 2020 and 2021. The Archimedes Underground remains unmined.

The West Archimedes deposit is hosted in the Ordovician Upper Goodwin limestone unit and is bound to the west by the Holly Fault. The zone strikes north-west and dips shallowly to the north-east. The deposit measures 2,000 ft along strike and 740 ft down dip and is up to 300 ft thick. The majority of West Archimedes was mined as an open pit before mining at East Archimedes. The mineralization in the West Archimedes deposit is oxidized and has a high ratio of cyanide soluble to fire assay total gold.

The East Archimedes Zone occurs east of the Graveyard Fault and proximal to the Graveyard Stock. Mineralization extends eastward from the West Archimedes Zone in the Upper Goodwin Formation and extends downward in the Lower Laminated and Lower Goodwin units along the contact with the Graveyard Stock. Silver and base metal grades are elevated in the East Archimedes zone in comparison with the other zones in the Ruby Hill Project in an envelope around the Blackjack zone replacement-style zinc mineralization described below. Mineralization in East Archimedes is roughly 1,200 ft wide and 1,200 ft long in plan and extends from surface where it is well defined by shallow drilling to several mineralized intersections over 1,800 ft below surface. The upper portion of the East Archimedes deposit, above an elevation of approximately 5,000 ft, is oxidized and transitional oxide-sulfide mineralization with a high ratio of cyanide soluble to total fire assay gold. The upper portion of the East Archimedes zone has been mined.

 

 

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The Archimedes Underground lies below, and extends north of, the Archimedes pit. It includes multiple zones of Carlin-type mineralization characterized by mineralization controlling faults and differing lithologic units with increasing depth. Mineralization is variably oxidized and refractory.

The 426 zone occurs in the Lower Laminated unit of the Goodwin Formation and the upper part of the underlying basal Goodwin unit of the Goodwin Formation in the nose of a fold. The mineralized zone forms a rod-shaped body plunging shallowly to the northeast that is 1,400 ft long, 200 ft wide and 200 ft thick. The top of the zone is approximately 1,000’ below surface, but it is 500’ below the bottom of the current East Archimedes pit bottom. The majority of the higher-grade mineralization occurring in the Goodwin Formation Lower Laminated unit is sulfide-style mineralization with a low ratio of cyanide soluble to total fire assay gold but the lower portion of the zone that is hosted in the basal Goodwin Unit has a moderate cyanide soluble to total fire assay gold mineralization.

The Ruby Deeps zone is a north-south striking, shallowly east dipping zone of mineralization hosted in the Windfall Formation and Dunderberg Shale in proximity to bodies of Bullwhacker Sill intrusive bound by the Graveyard Fault to the east and the Holly Fault to the west. The zone is 2,400 ft long 500 ft wide and 600 ft thick. The top of the zone is 1,600 ft below surface and 1,000 ft below the bottom of the West Archimedes pit. Within the zone there are several tabular horizons of higher-grade mineralization that are 40 ft to 100 ft thick.

The 007 Zone is an exploration target controlled by the NE trending NS Fault. Higher-grade oxide Au mineralization within the fault zone has been intersected by two holes, Barrick’s RC hole P7, 55’ @ 0.291 Au opt and i-80’s core hole iRH22-18A, 43.9’ @ 0.276 Au opt. Three more i-80 holes west of the fault zone intersect mineralization extending west into the Bullwhacker member. The zone is untested to the north and south, currently projecting about 400 ft along strike, 100 ft along dip, and ranges from 10 ft thick where stratigraphically controlled to over 40 ft thick within the NS fault zone.

The 008 Zone is an early exploration stage target. It is stratigraphically controlled, lying near the top of the Windfall Formation in the hinge of an anticline bracketed by the 426 and NS faults. The anticline appears to have formed above an intrusive lens emplaced within the upper member of the Windfall Formation, stratigraphically higher than typical Cretaceous sill material, which typically intruded along the lower contact of the Windfall Fm. The 008 Zone is not well defined but currently is about 350 ft long by 200 ft wide by 15 ft thick.

 

1.3.3

CRD and Skarn

Skarn and CRD mineralization are known to occur on the property but are not being considered for extraction in the current analysis.

Polymetallic (Au-Ag-Pb-Zn) skarn and carbonate replacement deposit (CRD) mineralization is lithologically and structurally controlled. Skarn occurs at Blackjack and the Hilltop Fault-Graveyard Flats stock intersection, primarily within the carbonate-rich Ordovician and mid to upper Cambrian formations adjacent to the Graveyard Flats stock. Minor skarn and CRD mineralization occur within the Cretaceous intrusive units.

Blackjack is a pod of zinc skarn mineralization hosted by the Lower Goodwin Unit proximal to the Graveyard Flats stock within the East Archimedes Zone below the Archimedes pit. It has elevated lead, copper and silver due to CRD overprinting. The base metal-rich CRD and skarn mineralization has been overprinted by later Carlin-style gold mineralization resulting in locally higher-grade gold zones. It is approximately 750 ft wide, 750 ft long and 900 ft high. The Hilltop Fault-Graveyard Flats stock intersection is an exploration stage target and has not been well defined.

 

 

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CRD mineralization tends to occur in the carbonate-rich formations along WNW trending faults. Examples include the historic Ruby Hill mine, the FAD deposit and the Hilltop exploration target.

 

1.4

Metallurgical Testing and Processing

 

1.4.1

Archimedes Underground

The Ruby Hill project encompasses several deposits and mineralization types hosting both precious and base metals. Historical production dates to 1998, primarily under Homestake Mining and Barrick Gold, with intermittent operations up to the current date.

Assumptions are based on historic and current metallurgical performance and the test work reports for oxide gold heap leaching, benchmarks, and the test work reports for zinc sulfide flotation. No detailed process design or production planning has been undertaken at this stage of the Project.

Historically, there have been three destinations for treatment of mineralization from the Ruby Hill Mine: (i) run of mine (ROM) and crushed mineralization to a heap leach pad, (ii) crushing and tank leaching with agglomerated tailings routed to the heap leach pad, and (iii) higher-grade sulfide mineralization (DSO) routed to Nevada Gold Mines Goldstrike Operation for autoclave processing.

For the Archimedes Underground, production with be processed at a third party destination capable of processing refractory ore until such time that i-80 has refurbished the Lone Tree Autoclave facility. The third party destination is an autoclave circuit capable of processing 4 - 5 million tons per year and consists of primary crushing, two parallel semi-autogenous grinding (SAG) Mill-Ball Mill grinding circuits with pebble crushing, five parallel autoclaves capable of acid pressure oxidation (POX) and three of which are capable of alkaline POX, two parallel calcium thiosulphate (CaTS) leaching circuits with resin-in-leach (RIL), electrowinning for gold recovery, and a refinery producing doré bullion from both autoclave and roaster circuits.

The Lone Tree Autoclave Facility is located immediately adjacent to i-80, approximately 12 miles west of Battle Mountain, 50 miles east of Winnemucca, and 120 miles west of Elko. The Lone Tree processing facilities were shut-down at the end of 2007. Since that time, the mills have been rotated on a regular basis to lubricate the bearings. In general, the facility is still in place with most of the equipment sitting idle. i-80 Gold Corp’s objective is to refurbish and restart the POX circuit and associated unit operations, including the existing oxygen plant, as it was operating before the shut-down, while meeting all new regulatory requirements. The flotation circuit is not being considered for restart. The POX circuit will have capability to operate under either acidic or basic conditions.

 

1.4.2

Mineral Point Open Pit

The Mineral Point project encompasses several deposits and mineralization types hosting both precious and base metals. Historical production dates to 1998, primarily under Homestake Mining and Barrick Gold, with intermittent operations up to the current date.

Generally, previous operating experience as well as the metallurgical test work confirms the amenability of oxide material to heap leaching for precious metals extraction. From 2004 to 2014, seven test work programs were carried out, by Kappes Cassiday Associates (KCA) focusing on column leaching and bottle roll leach testing of the oxide deposits, namely Archimedes, 426 and Mineral Point. Mineral Point estimated recoveries are based on alteration type ranging from 83% to 84.4% gold and 40% to 45.2% silver for oxide mineralization. The proposed process for Mineral Point Open Pit material is a two-stage crush conventional heap leach operations with a Merrill-Crowe processing facility.

 

 

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1.5

Mineral Resources

 

1.5.1

Archimedes Underground

Practical Mining LLC estimated the Archimedes Underground mineral resource using all drilling and geological data available through October 31, 2022. Wood Canada Ltd. estimated and reported open pit mineral resources in the inaugural NI 43-101 Technical Report under i-80’s ownership of the Ruby Hill Project. All work, including drilling, done since the time of the inaugural report has targeted the 426, Ruby Deeps and other underground deposits and does not influence the Open Pit mineral resource reported on the October 2021 report. The open pit mineral resources reported in October 2021 are current and are restated herein.

Table 1-1: Summary of Archimedes Underground Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

           
Deposit   

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

 
Indicated Mineral Resources
           

426

   899    6.9    0.8    199    22
           

Ruby Deeps

   892    8.3    2.4    237    69
           

Total Indicated

   1,791    7.6    1.6    436    92
 
Inferred Mineral Resources
           

426

   1,038    6.6    1.2    219    40
           

Ruby Deeps

   3,150    7.6    2.4    769    246
           

Total Inferred

   4,188    7.3    2.1    988    286

Notes:

  1.

Underground mineral resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of $27.25 per ounce (Section 16.1).

  2.

Mineral resources have been estimated using pressure oxidation gold metallurgical recoveries of 96.8% and 89.5% for the 426 and Ruby Deeps deposits respectively.

  3.

Pressure oxidation cutoff grades are 5.06 and 5.48 Au g/t (0.148 and 0.160 opt) for the 426 and Ruby Deeps deposits respectively.

  4.

Detailed input mining, processing, and G&A costs are defined in Section 18.1.

  5.

Units shown are metric.

  6.

The contained gold ounces estimates in the mineral resource table have not been adjusted for metallurgical recoveries.

  7.

Numbers have been rounded as required by reporting guidelines and may result in apparent summation differences.

  8.

A mineral resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

  9.

An inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration.

  10.

Mineral resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors.

  11.

Mineral resources have an effective date of December 31, 2024.

  12.

The reference point for mineral resources is in situ.

 

 

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1.5.2

Archimedes Open Pit

The Archimedes deposit was previously mined by Homestake and Barrick for West Archimedes and East Archimedes respectively. Mining ceased after a pit wall failure. In this study an updated estimation of the Archimedes mineral resource has been developed Forte Dynamics, Inc (Forte), and the mining potential for continuing the surface exploitation of the deposit was evaluated to estimate a current open pit mineral resource estimate.

The Archimedes mineral resources are detailed in Table 1-2. Mineral resources are not Mineral Reserves and have not been demonstrated to have economic viability. There is no certainty that the mineral resource will be converted to Mineral Reserves. Inferred mineral resources do not have sufficient confidence that modifying factors can be applied to convert them to mineral reserves. The quantity and grade or quality is an estimate and is rounded to reflect the fact that it is an approximation. Quantities may not sum due to rounding.

Table 1-2: Summary of Archimedes Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

             
Deposit   

Cutoff Au

(g/t)

  

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

 
Indicated Mineral Resources
             

Archimedes Pit

   0.2    4,280    1.98    10.7    272    1,460
   0.1    4,320    1.96    10.6    272    1,490
   0.05    4,340    1.95    10.6    272    1,480
 
Inferred Mineral Resources
             

Archimedes Pit

   0.2    820    1.18    8.9    31    230
   0.1    870    1.12    8.5    31    250
   0.05    880    1.11    8.5    31    250

Notes:

  1.

Mineral resources have an effective date of December 31, 2024.

  2.

Mineral resources are the portion of Mineral Point that can be mined profitably by open pit mining method and processed by heap leaching.

  3.

Mineral resources are below an updated topographic surface (below Archimedes pit).

  4.

Mineral resources are constrained to economic material inside a conceptual open pit shell. The main parameters for pit shell construction are a gold price of $2,175/oz Au, a silver price of $26.00/oz, average gold recovery of 77%, average silver recovery of 40%, open pit mining costs of $3.31/tonne, heap leach average processing costs of $3.47/tonne, general and administrative cost of $0.83/tonne processed, gold refining cost of $1.85/oz, silver refining cost of $0.50, and a 3% royalty (Section 16.1).

  5.

Mineral resources are reported above a 0.1 g/t Au cutoff grade. Silver revenues were not considered in the cutoff grade.

  6.

Mineral resources are stated as in situ.

  7.

Mineral resources have not been adjusted for metallurgical recoveries.

  8.

Reported units are metric tonnes.

  9.

Reported table numbers have been rounded as required by reporting guidelines and may result in summation discrepancies.

 

 

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1.5.3

Mineral Point Open Pit

Forte reviewed the Mineral Point Open Pit mineral resource estimate completed by Wood in July 2021. The scope of the review included the informing drillhole and sample data, exploratory data analysis (EDA), input models, and the current topography. The scope also included a review of the grade estimation methodology and model validation, bulk density determination, resource classification, reasonable prospects for eventual economic extraction (RPEEE), and the statement of mineral resources.

Upon completion of the Mineral Point open pit resource review, Forte made some slight modifications to the Wood block model. Note that the estimated block grades were not altered or changed. Updates included updating the block model with the current topographic surface, recoding the Wood 2021 lithological model to the block model along with an assigned specific gravity (SG) values based on lithology code, and updated values and conversions for tonnage factor. Forte also used an updated pit shell to constrain and report the mineral resource under the requirements for RPEEE, which was based on a 2024 Scoping Study completed by Forte and used for other work completed in this Technical Report Summary.

The Mineral Point Open Pit mineral resources are detailed in Table 1-3. Mineral resources are not mineral reserves and have not been demonstrated to have economic viability. There is no certainty that the mineral resource will be converted to mineral reserves. Inferred mineral resources do not have sufficient confidence that modifying factors can be applied to convert them to mineral reserves. The quantity and grade or quality is an estimate and is rounded to reflect the fact that it is an approximation. Quantities may not sum due to rounding.

Table 1-3: Summary of Mineral Point Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

           
Deposit   

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

 
Indicated Mineral Resources
           

Mineral Point

   216,982    0.48    15.0    3,376    104,332
           

Total Indicated

   216,982    0.48    15.0    3,376    104,332
 
Inferred Mineral Resources
           

Mineral Point

   194,442    0.34    14.6    2,117    91,473
           

Total Inferred

   194,442    0.34    14.6    2,117    91,473

Notes:

  1.

Mineral resources have an effective date of December 31, 2024.

  2.

Mineral resources are the portion of Mineral Point that can be mined profitably by open pit mining method and processed by heap leaching.

  3.

Mineral resources are below an updated topographic surface.

  4.

Mineral resources are constrained to economic material inside a conceptual open pit shell. The main parameters for pit shell construction are a gold price of $2,175/oz Au, a silver price of $26.00/oz, average gold recovery of 77%, average silver recovery of 40%, open pit mining costs of $3.31/tonne, heap leach average processing costs of $3.47/tonne, general and administrative cost of $0.83/tonne processed, gold refining cost of $1.85/oz, silver refining cost of $0.50, and a 3% royalty (Section 16.1).

  5.

Mineral resources are reported above a 0.1 g/t Au cutoff grade.

  6.

Mineral resources are stated as in situ.

  7.

Mineral resources have not been adjusted for metallurgical recoveries.

  8.

Reported units are metric tonnes.

  9.

Reported table numbers have been rounded as required by reporting guidelines and may result in summation discrepancies.

 

 

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1.6

Mining, Infrastructure, and Project Schedule

 

1.6.1

Archimedes Underground

Permitting approval for development and mining above the 5100 elevation is anticipated by the end of Q2 2025 and underground development will commence immediately thereafter. This is consistent with previously approved permits for mining the Archimedes open pits. Production mining in the 426 deposit will start in 2026 and continue through 2027 with oxide material processed on site in the existing heap leach facility and refractory material sent to a third party for toll processing. Permits for mining below the 5100 elevation are anticipated in the second quarter of 2027 with development mining for the Ruby Deeps deposit beginning shortly thereafter.

Mining conditions anticipated are typical for northern Nevada underground mines. Long hole open stoping will be the primary mining method and will be supplemented with underhand drift and fill mining where deposit geometry dictates. Mining will be undertaken by a qualified contractor, eliminating the need to recruit a workforce and purchase mining equipment.

Transportation, electrical and support infrastructure already exists at Ruby Hill. Additional infrastructure requirements are limited to:

 

   

Overhead power line and transformer at the portal site.

   

Backfill and shotcrete plants.

   

Fuel and oil storage near the portal.

   

Contractor’s maintenance facility and office.

   

Mine water supply tank.

 

1.6.2

Archimedes Open Pit

The Archimedes Open Pit mineral resource has not been evaluated for surface mining.

 

1.6.3

Mineral Point Open Pit

The Mineral Point Project, operated by i-80 Gold, is planned as an open pit mining operation using conventional equipment, targeting a processing rate of 68,000 tons per day. While there is currently no Mineral Reserve Estimate, the project contains indicated and inferred mineral resources. Pit optimization using Hexagon Mine Plan software identified an optimal pit shell (LG72) with a 78% revenue factor, containing 4.98 million ounces of gold and 195.5 million ounces of silver at an average stripping ratio of 2.8:1. Key economic parameters include a gold price of $2,175/toz, silver price of $27.25/toz, and heap leach average recovery rates of 78% for gold and 41% for silver. The calculated cutoff grade for gold is 0.011 oz/ton, ensuring the extraction of economically viable material.

The mine design consists of nine pit phases, with mining benches at 50-foot intervals and a projected Life-of-Mine (LOM) of 17 years. The operation will rely on a mining fleet comprising two rope shovels, (2) hydraulic shovels, (26) haul trucks, and various support equipment. Annual production is expected to average 4.5 million ounces of gold and 177.3 million ounces of silver. Dewatering will be required in later mine stages, though the extent is yet to be determined.

The project will leverage existing infrastructure from previous mining activities at the Ruby Hill site, including site access, haul roads, waste rock storage, and power supply, with necessary upgrades. Key processing facilities include a crushing and stacking system, a heap leach pad, and a Merrill Crowe plant for gold and silver recovery. The heap leach facility will be developed in five phases, with a total capacity of 466.8 million

 

 

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tons, while a Merrill Crowe plant will process pregnant leach solution at a rate of 11,500 gallons per minute, ultimately producing doré bars for off-site refining.

Supporting infrastructure includes an expanded truck shop, warehouse, administration building, and water management systems for process, potable, fire suppression, and stormwater control. Power will be sourced from an existing substation with potential upgrades, and site communications will be maintained through telemetry and radio networks. Waste rock storage will utilize both surface storage and in-pit backfilling. Environmental considerations include stormwater management ponds, lined heap leach pads, and containment systems for fuel and hazardous materials. Road expansions and rerouting will be necessary to accommodate mining activities, ensuring operational efficiency while minimizing environmental impact.

 

1.7

Economic Analysis

Economic analysis relies on many forward-looking assumptions for the estimation of metal prices, capital and operating costs. These are subject to change depending on operating strategy, new information collected through future operations and macroeconomic conditions. Actual economic outcomes often deviate significantly from forecasts.

 

1.7.1

Archimedes Underground

The economic model is based on a mine plan that includes 69% inferred mineral resources. The results obtained excluding inferred material is a gross adjustment. Recalculation of capital and operating costs has not been included in the scenario excluding inferred mineral resources. The values presented are derived from a constant dollar after tax cash flow analysis. Capital and operating costs are summarized below in Table 1-4 and financial statistics are presented in Table 1-5.

Table 1-4: Capital and Operating Cost Summary

 

Category     Total Cost $M      $/ton Processed      $/Au oz 
Costs Without Inferred

Mining

   $223    $147.74    $801

Processing and Transportation

   $204    $135.51    $735

G&A, Royalties and Net Proceeds Tax

   $133    $87.99    $477

By Product Credits

   ($0.2)    ($0.13)    ($1)

Total Cash Cost

   $560    $377.11    $2013

Closure and Reclamation

   $8.9    $5.89    $32

Sustaining Capital

   $106    $70.36    $382

All in Sustaining Costs

   $646    $447.37    $2,427

Construction Capital

   $49    $32.75    $178

All in Costs

   $724    $480.12    $2,604
Cost With Inferred

Mining

   $750    $148.98    $808

Processing and Transportation

   $682    $135.51    $735

G&A, Royalties and Net Proceeds Tax

   $210    $41.80    $227

By Product Credits

   ($0.7)    ($0.1)    ($1)

Total Cash Cost

   $1,642    $326.17    $1,769

Closure and Reclamation

   $8.9    $1.77    $10

Sustaining Capital

   $106    $21.08    $114

All in Sustaining Costs 2

   $1,757    $349.01    $1,893

Construction Capital

   $49    $9.81    $53

All in Costs 3

   $1,806    $358.82    $1,946

 

 

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Table 1-5: Financial Statistics

 

Parameter

 

  

 Value With 

 Inferred 

    Value  Without 
 Inferred 

Gold Price (US$/oz)

   $2,175    $2,175

Silver Price (US$/oz)

   $27.25    $27.25

Mine Life (years)

   10    10

Mining Rate (tons/day)

   1,600    450

Tons Processed Autoclave (kton)

   4,846    1,452

Average Grade Autoclave (Au oz/ton)

   0.209    0.209

Average Gold Recovery (Autoclave %)

   90%    90%

Autoclave Gold Produced (koz)

   910    272

Tons Processed Heap Leach (kton)

   188    56

Average Grade Heap Leach (Au oz/ton)

   0.111    0.111

Average Gold Recovery (Heap Leach %)

   87%    87%

Heap Leach Gold Produced (koz)

   18    5.5

Average Annual Gold Production (koz)

   102    31

Total Recovered Gold (koz)

   928    278

Project After-Tax NPV5% (M$)

   $127    ($113)

Project After-Tax NPV8% (M$)

   $91    ($109)

Project After-Tax IRR

   23%    NA

Payback Period

   7.8 Years    NA Years

Notes:

  1.

Net of byproduct sales.

  2.

Excludes, construction capital, exploration, corporate G&A, interest on debt, and corporate taxes.

  3.

Excludes exploration, corporate G&A, interest on debt, and corporate taxes.

  4.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable United States and Canadian securities legislation. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in the Company’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this report is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. The Company disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law.

 

1.7.2

Mineral Point Open Pit

The economic analysis of the Mineral Point Project is based on the mine schedule, capital and operating costs, metal recovery parameters, and royalties. The project, operated by i-80 Gold, is planned as an open pit operation with a processing rate of 68,000 tons per day. The economic model assumes a gold price of $2,175/oz and a silver price of $27.25/oz, with a total initial capital investment of $708 million and sustaining capital of $388 million. In addition, approximately 115 million tons of stripping is required to gain access to the body of mineralized material, costing $287 million. The life-of-mine (LOM) plan spans approximately 16.5 years, with total recovered gold and silver estimated at 3.5 million ounces and 72 million ounces, respectively. The estimated pre-tax net present value (NPV) at a 5% discount rate is $827.6 million, with

 

 

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an internal rate of return (IRR) of 13.8% and a payback period of 7.6 years. After-tax, the NPV at 5% is reduced to $614.1 million, with an IRR of 12.1% and a payback period of 7.9 years.

A sensitivity analysis indicates that the project is most sensitive to metal prices and recovery rates, followed by capital and operating costs. The inclusion of inferred resources, which constitute 38% of the mine plan, significantly impacts the economic assessment. When excluding inferred resources, the mine life is reduced to 11.5 years, and the after-tax NPV at 5% drops to $157.9 million. Under this scenario, the project becomes marginal at higher discount rates, with an IRR of 7.8% and a longer payback period of 8.9 years. While the Mineral Point Project demonstrates economic potential, additional exploration and refinement of cost estimates are necessary to improve confidence in the resource model and the feasibility of long-term operations. Table 1-6 shows the total and unit operating costs with and without inferred. Table 1-7 shows the financial with and without inferred.

Table 1-6: Unit and Total Operating Costs With and Without Inferred Resources

 

Category     Total Cost $M      $/ton Processed      $/Au oz 
Costs With Inferred

Mining

   $3,874.40    $9.80    $1,097.75

Processing

   $1,542.23    $3.90    $436.97

G&A

   $296.58    $0.75    $84.03

Refining, Royalties & Net Proceeds Tax

   $722.30    $1.83    $204.65

By-Product Credits

   $(1,952.96)    $(4.94)    $(553.34)

Total Operating Cost/Cash Costs

   $4,482.57    $11.34    $1,270.07

Closure & reclamation

   $69.83    $0.18    $19.78

Sustaining Capital

   $388.43    $0.98    $110.05

All-in Sustaining Costs

   $4,940.82    $12.49    $1,399.91
Cost Without Inferred

Mining

   $2,213.49    $11.15    $1,124.31

Processing

   $774.50    $3.90    $393.40

G&A

   $148.94    $0.75    $75.65

Refining, Royalties & Net Proceeds Tax

   $374.27    $1.88    $190.10

By-Product Credits

   $(851.57)    $(4.29)    $(432.54)

Total Operating Cost/Cash Costs

   $2,659.63    $13.39    $1,350.92

Closure & reclamation

   $67.33    $0.34    $34.20

Sustaining Capital

   $131.48    $0.66    $66.78

All-in Sustaining Costs

   $2,858.44    $14.39    $1,451.90

 

 

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Table 1-7: After-Tax NPV Comparison of With and Without Inferred Resources

 

Parameter   Unit     Value With  Inferred     Value Without  Inferred  

Mine Life

year 16.5 11.5

Mining Rate

kton/day 356.2 328.8

Processing Rate

kton/day 68.4 49.3

Total Processed Material

kton 395,444 195,591

Total Mine Material

kton 1,675,243 987,993

Average Processing Grade Au

toz/ton 0.011 0.012

Average Processing Grade Ag

toz/ton 0.448 0.383

Contained Au

ktoz 4,525 2,430

Contained Ag

ktoz 177,293 76,109

Recovered Au

ktoz 3,529 1,969

Recovered Ag

ktoz 72,028 31,407

Heap Leach Recovery Au (average)

% 78% 81%

Heap Leach Recovery Ag (average)

% 41% 41%

Total LOM CAPX

US$M $1,383.2 $941.2

NPV @ 0%

US$M $1,470.0 $574.1

NPV @ 5%

US$M $614.1 $157.9

NPV @ 8%

US$M $295.8 $(10.9)

NPV @ 10%

US$M $134.8 $(100.1)

NPV @ 12%

US$M $4.3 $(174.8)

IRR

% 12.1% 7.8%

Payback Period

Year 7.6 7.9

 

1.8

Conclusions

 

1.8.1

Archimedes Underground

 

1.8.1.1

Mineral Resources

The Archimedes Underground mineral resource contains approximately 70% inferred mineral resources. The planned underground development and drilling program is planned to upgrade inferred mineral resources to indicated.

 

1.8.1.2

Mining and Infrastructure

Mining conditions for the Archimedes underground are typical for sedimentary deposits in the north-east Nevada extensional tectonic environments are anticipated. The Ruby Deeps deposit will require dewatering with anticipated pumping rates of 500 to 1,000 gpm.

 

1.8.1.3

Metallurgical Testing

Metallurgical testing of refractory samples from Archimedes underground deposits has confirmed amenability to grinding followed by pressure oxidation and carbon in leach. Gold recoveries ranged from 80% to 91%. Metallurgical testing programs have identified deleterious elements that are common to deposits in this part of Nevada. Deleterious elements content in the oxide samples are low, while sulfide samples are characterized by high levels of sulfide sulfur, arsenic, and mercury. Processing of Archimedes sulfide mineralization through a third-party or i-80’s Lone Tree autoclave will ensure removal and capture of these deleterious elements.

 

1.8.1.4

Recovery Methods

Metallurgical testing has confirmed that processing of Archimedes underground sulfide mineralization can be processed through Nevada Gold Mines Twin Creeks or the Lone Tree autoclave facilities. The 426 mineralized lenses are more amenable to alkaline conditions while the Ruby Deeps lenses perform better with acidic conditions.

 

 

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1.8.1.5

Financials

 

   

Initial capital requirements total $49.4M with an additional $106.1M in sustaining capital.

   

The project achieves after-tax NPV 5% of $126.8M and NPV 8% of $91.1M.

   

The estimated payback period is 7.8 years with an IRR of 23%.

 

1.8.2

Archimedes Open Pit

 

1.8.2.1

Mineral Resources

The Archimedes deposit was previously mined by Homestake and Barrick for West Archimedes and East Archimedes respectively. Mining ceased after a pit wall failure. An updated mineral resource estimate was completed, with the majority of mineral resources classified as indicated. There is currently potential for additional surface production of the deposit which would add to the value of the overall Ruby Hill project.

As the pit was never restarted after the wall failure, it will be important to understand and mitigate rock mechanics stability and safety issues prior to any decision to restart the project.

Given the current focus on the underground mine and the Mineral Point pit, no additional work in the Archimedes pit has been planned.

 

1.8.3

Mineral Point Open Pit

 

1.8.3.1

Mineral Resources

The Mineral Point Open Pit mineral resource contains approximately 47% inferred mineral resources. Drilling is planned for the deposit to obtain fresh material for additional metallurgical testing. The additional metallurgical test results can be used in future work, along with additional testing for representative bulk density measurements to be used with future updated geological, alteration, redox and structural models. This can be used for future mineral resource updates and potentially upgrading inferred mineral resources to indicated mineral resources.

 

1.8.3.2

Mining and Infrastructure

Mineral point will be a large-scale open pit gold and silver deposit typical of other northern Nevada mines with stripping ratio of 2.9:1, excluding capitalized pre-stripping. Overall average gold grade processed of 0.39 g/tonne with an expected average gold recovery of 78% and an average silver grade processed of 15.37 g/tonne. Most of the current infrastructure on site can be re-used or expanded for the project. Power for the proposed operation will be provided by the power supplier that historically fed the site.

 

1.8.3.3

Metallurgical Testing

Historical metallurgical testing and production have confirmed the amenability of Mineral Point open pit oxide and sulfide mineralization to conventional cyanide heap leaching; Metallurgical testing of samples from the Mineral Point open pit deposit has also shown amenability to crushing for heap leaching. Gold and silver recoveries ranged from 80-85% and 32-45% respectively.

 

1.8.3.4

Recovery Methods

Oxide and sulfide material is amenable for processing by crushed-ore cyanide heap leaching. Gold and silver leach at the heap-leach facility will be extracted by Merrill-Crowe zinc precipitation.

 

1.8.3.5

Financials

 

   

Total capital requirement of $1,383.2M

   

The project achieves an NPV 5% of $614.1M and NPV 10% of $134.8M After-Tax

 

 

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The project has and IRR of 12.1% and a payback period of 7.9 years After-Tax

 

1.9

Recommendations

 

1.9.1

Archimedes Underground

 

1.9.1.1

Metallurgical Testing

 

   

Additional metallurgical testing is recommended from initial Ruby Hill production areas to confirm metallurgical recoveries with Twin Creeks process conditions. Sample selection should be based on available mine production plans and should reflect typical stope dimensions and expected dilution. Testing should include:

   

Comminution testing to confirm throughput through the Sage Mill.

   

Pressure oxidation tests using Twin Creeks conditions.

   

CIL tests on pressure oxidation productions.

 

   

Additional testing on Ruby Hill base metal sulfide zones to investigate flotation parameters to produce saleable lead and zinc concentrates. Detailed assays of lead and zinc concentrates are recommended to determine the extent of deleterious elements that may impair their salability.

 

1.9.1.2

Permitting and Mine Development

 

   

Complete the EA and POO amendment for Mining the 426 deposit above the 5100 elevation.

   

Initiate construction of the haulage portal and decline in Q3 2025.

 

1.9.1.3

Resource Conversion and Exploration Drilling

 

   

Begin Resource Conversion Drilling as soon as decline advance and drill platforms become available.

   

The lower leg of the decline provides a drill platform for exploration of the Blackjack deposit.

 

1.9.1.4

Dewatering

 

   

Initiate a hydrogeologic study of the Windfall formation, drill a deep test well and complete a drawdown test.

 

1.9.2

Archimedes Open Pit

Due to the short-term development plans for Mineral Point Open Pit and Archimedes Underground, additional work for the Archimedes Open Pit is not currently defined. Should resources be available a detailed geotechnical review of the existing pit slopes in Archimedes could help to quantify future potential. In light of current development plans on the property, this is not budgeted at this time.

 

1.9.3

Mineral Point Open Pit

 

1.9.3.1

Mineral Resources

It is recommended that i-80 complete additional resource definition drilling and conduct a review of major and minor rock alteration types, and how they align with overall geology, grade domains, metallurgical recovery and bulk densities. This would also include review of the geological model, including lithological, structural, and alteration controls on overall grade distribution and metallurgical recovery. The additional drilling could be used to better define the limits of mineralization and potentially upgrades block classification.

 

 

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The following points are recommended for additional evaluation:

 

   

Review of the overall (and subsequent low and high-grade) grade distributions to better understand impacts on mineralized domains.

   

Detailed review of deposit wide bulk densities to better define the bulk density for the project, including bulk densities of lithology and alteration type.

   

Additional drilling to increase the resource definition and confidence, along with potential upgrading of resource classification (inferred to indicated, indicated to measured).

   

Additional drilling for potential resource expansion.

Upon completion of the above items, an update to the geological model and mineral resource estimate should be conducted, along with updated metallurgical recovery assumptions.

 

1.9.3.2

Mining and Infrastructure

It is recommended that a site wide water balance be developed for the project to better understand water captured on-site (pit, HLP, WRSA) and evaluate the ability to utilize this water for process make-up water or to provide water for agriculture use. This would include evaluation of climate and available make-up water sources to understand total project requirements for make-up water or discharge as required. The evaluation would include a more accurate reflection of drain down for events, and potentially reduce the event pond volumes required, which could impact capital and sustaining capital costs.

There are several opportunities for infrastructure related components of the project to evaluate, including:

 

   

Conveyor stacking versus truck stacking, reduction of capital and operating costs.

   

Blasting versus crushing and screening, reduction of capital and operating costs.

   

Reduced number of event ponds and utilize larger event ponds to reduce capital costs.

   

Increased Heap Ultimate height of 300 feet, reduction of disturbance area as well as capital costs.

   

Utilization of existing crusher to self-perform overliner manufacturing to reduce capital costs

   

Evaluate all pits for potential for pit dewatering, including water quality evaluation, for ability to utilize this water as process make-up water or for agricultural use.

 

1.9.3.3

Metallurgical Testing

It is recommended that additional metallurgical testing be conducted to further define the predicted recovery for the Mineral Point Open pit project. This includes evaluation of sulfide sulfur content which will assist with determining the various oxidations by lithology as well as understanding recovery and reagent consumptions. This should also be conducted for waste as there may be a need to segregate waste into PAG and NAG facilities.

Next phases of the metallurgical testing program would incorporate additional leach tests, coarse bottle rolls, and column leach tests. This testing is required to support crush size selection, recovery estimates and reagent consumptions for lime and cyanide. Testing is also required to provide comminution design data. Testing and samples to be tested include:

 

   

Samples should focus on weakly-altered alteration of the major formations, the largest component of the Mineral Point resources. Sample selection should address spatial and grade variability within the deposit.

   

Identify samples in transition areas to sulfide mineralization to establish boundary criteria such as sulfide sulfur content.

   

Use of PQ diameter drilling will permit testing up to -2” crush size to evaluate the impact of crush size on recoveries.

 

 

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Evaluate the pilot leach testing of a bulk sample to determine ROM recoveries.

 

   

Testing of composite samples representing the first year and second year mine production once optimal conditions are selected.

 

   

Conduct column leach tests with taller columns and columns in series to replicate actual lift heights and heap leach operations.

 

   

Conduct laboratory tests to determine the crusher work index and abrasion indices to support crushing plant design.

 

   

Geotechnical testing, namely compacted permeability testing, of samples to determine the permeability and stacking characteristics of the mineralized material.

 

   

ABA testing of leach residue under conditions to support environmental permitting.

Additional considerations include metallurgical and geotechnical testing which will further the understanding of the ore’s clay content. This would include particle size distribution analysis, Atterberg limits, plasticity index, by ore type. This would also be coupled with compacted permeability testing to understand long term effects of loading and stacking. It is also recommended that ore decrepitation testing be conducted. Additional evaluation of the outcomes of this testing will verify the proposed application rate, leach cycle, and stack height for the various oxidations and lithologies based on permeability and agglomeration requirements.

It is also recommended that additional testing of proposed overliner material be conducted to evaluate screening requirements as well as stability for geotechnical design. This could also lead to a reduction in the overliner depth requirement, decreasing capital costs for the project.

Additional test work for recovery potential of the relocated HL material from historic operations should be conducted to potentially include revenue from this material.

The program has an estimated cost of $600,000 (excluding drilling costs) based on current conditions.

 

1.9.4

Work Programs

 

1.9.4.1

Archimedes Underground

The work program outlined in Table 1-8 will advance the 426 deposit to production within two years. Project risks are manageable, and opportunities exist to enhance the project economics.

Table 1-8: Archimedes Underground Work Program

 

Description     2025         2026     Estimated
  Costs  (US$M)  

Portal Construction

  0.1     0.1     0.2  

Mine Development

  7.8     21.0     28.8  

Resource Conversion Drilling

  2.1     -     2.1  

Dewatering Well and Hydrogeologic Study

  3.9     -     3.9  

Environmental, Metallurgical Testing and Feasibility Study

  0.5     2.0     2.5  

Ventilation and Electrical

  0.2     2.7     2.9  

Project Administration

  5.0     0.6     5.6  

Contingency

  2.9     4.5     7.4  

Total

  22.5     30.8     53.3  

 

 

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1.9.4.2

Archimedes Open Pit

Due to the short-term development plans for Mineral Point Open Pit and Archimedes Underground, additional work for the Archimedes Open Pit is not currently recommended.

 

1.9.4.3

Mineral Point Open Pit

The work program outlined in Table 1-9 will advance the Mineral Point Open Pit project to a Pre-Feasibility Study (PFS).

 

1.9.4.3.1

Phase 1

A two-phase work program is recommended. The focus of the Phase 1 work program will be additional drilling to obtain new sample material for metallurgical test work, hydro and geotechnical studies. This will include metallurgical test work of sufficient variability samples to support overall recovery assumption prior to moving to Phase 2. The additional drilling will also be used for subsequent resource definition, and potential resource classification upgrade and expansion. Based on the results of Phase 1, Phase 2 may be warranted. Additional metallurgical test work and other studies may be needed to further de-risk the Project.

 

1.9.4.3.2

Phase 2

The focus of the Phase 2 work program will be additional drilling for resource definition and expansion; and will include additional metallurgical test work to refine the process parameters. The Phase 2 drilling will be designed for resource conversion and growth, with the objective of converting inferred resources to indicated resources, as well as converting indicated resources to measured resources. The additional drilling and potential upgrade of inferred resources to indicated resource may lead to mineral reserves.

Table 1-9: Mineral Point Work Program

 

Description

  Estimated Costs  

(US$M)

Phase 1  

Additional Drilling for Metallurgical, Hydro and Geotechnical Test Work

  $ 3.30  

Metallurgical Test Work

  $ 0.25  

Contingency

  $ 0.70  
Phase 1 Total $ 4.25
Phase 2  

Resource Definition & Expansion Drilling

  $ 15.0  

Metallurgical Test Work

  $ 0.20  

Contingency

  $ 1.00  
Phase 2 Total $ 16.20

 

 

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2.

INTRODUCTION

 

2.1

Registrant for Whom the Technical Report Summary was Prepared

This Technical Report Summary (TRS) is as initial assessment Technical Report Summary in accordance with the Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for the registrant i-80 Gold Corporation and its subsidiaries Ruby Hill Mining Company LLC, Premier Gold Mines, USA Inc. and Golden Hill Mining Corporation (collectively “i-80” or the “Company”, or the “Registrant”). This is the initial TRS for i-80’s Ruby Hill Project. The company has previously disclosed information on the project under Canadian Securities National Instrument 43-101 (Wood 2021).

 

2.2

Terms of Reference and Purpose of this Technical Report

This Initial Assessment is a preliminary technical and economic study of the economic potential of all or parts of mineralization to support the disclosure of mineral resources. The Initial Assessment is preliminary in nature. It includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the Initial Assessment will be realized. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

This report is based in part on internal Company reports, previous studies, maps, published government reports, company letters and memoranda, and public information as cited throughout this report and listed in Section 24. Reliance upon information provided by the registrant is listed in Section 25 when applicable.

 

2.3

Qualified Persons

This Initial Assessment was compiled by Practical Mining, Raponi Consulting, and Forte Dynamics. All three firms are third-party firms comprising experts in their respective fields in accordance with 17 CFR § 229.1302(b)(1). i-80 has determined that all five firms meet the qualifications specified under the definition of qualified person in 17 CFR § 229.1300. Additional technical information was provided by the registrant and is detailed in Section 25.

None of the Qualified Persons (QPs) has any beneficial interest in i-80 or any of its subsidiaries, or in the assets of i-80 or any of its subsidiaries or in any property near the Ruby Hill Project. The QPs will be paid a fee for this work in accordance with normal professional consulting practices.

Practical Mining prepared/contributed to the following sections of this report:

 

   

Sections 1-9, 11.1, 11.2, 12, 13.1, 15.1, 16, 17, 18.1, 19.1, 20-25

Raponi Consulting prepared/contributed to the following sections of this report:

 

   

Sections 1, 10, 14.1, 14.2, 22-24

Forte Dynamics prepared/contributed to the following sections of this report:

 

   

Sections 1, 2, 11.1, 11.3, 11.4, 12, 13.2, 13.3, 14.3, 15.2, 18.2, 19.2, 21-25

 

 

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2.4

Details of Personal Inspection by Qualified Persons

Table 2-1 summarizes the details of the personal inspections on the property by each qualified person or, if applicable, the reason why a personal inspection has not been completed.

Table 2-1: Personal Inspections by Qualified Persons

 

QP Firm   Discipline   Dates of Personal
Inspection
   Details of Inspection
Practical Mining LLC  

Mining, Mineral Resources, Mineral Reserves, Geology and Mineralization of Carlin Type Deposits, Drilling, Data Verification

  July 14, 2022    Site specific hazard training, examined core and core logging procedures, examined proposed underground portal location, overview of Archimedes pits and heap leach pad.
TR Raponi Consulting Ltd.   Metallurgical Testing, Mineral Processing   None    Reviewed prior test work and designed and supervised current test work on samples from the various deposits at Ruby Hill.
Forte Dynamics, Inc.   Geology and Mineralization, Exploration, Drilling, Sample Preparation, Analysis and Security, Data Verification, Mining, Mineral Resources, Mineral Reserves, Site Infrastructure   January 16, 2025    Overview of the project history and current status, examined the Archimedes pit, examined site infrastructure, examined the heap leach pad, review of drill core, geology and mineralization, completed check assays from selected available drill core intervals, review of sample preparation, analysis and security, field inspection for drillhole collar locations, review of current geological model, topography and resource, reviewed proposed heap leach facility area and proposed waste rock storage area. See Site Visit Report (Forte, 2025) in Appendix A for additional details.

 

2.5

Report Version Update

This TRS is the initial S-K 1300 report by i-80 for the Ruby Hill Project. In July 2021, an NI 43-101 Mineral Resource Estimate technical report was prepared by Wood for i-80 (Wood 2021).

 

2.6

Units of Measure

U.S. Imperial units of measure are used throughout this document unless otherwise noted. Units and abbreviations are listed in Table 2-2. Currency is expressed as United States Dollars.

 

 

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Table 2-2: Units and Abbreviations

 

      Imperial    Metric
     Units    Description    Units    Description
Time    yr    year    yr    year
   d    day    d    day
   hr    hour    hr    hour
   min    minute    min    minute
   s    seconds    s    seconds
Length    ft    feet    m    meter
   in    inch    cm    centimeter
   mil    thousandth of an inch    mm    millimeter
   mi    miles    µm    micrometer (micron)
Area    ft2, sq ft    square feet    m2    square meters
   ha    hectare    ha    hectare
Mass    st    short ton    mt, t    metric tonne
   kton    kilo ton    ktonne    kilo tonne
   dst    dry short tons    dmt    dry metric tonnes
   kst, kdst    thousand dry short tons    kmt, kdmt    thousand dry metric tonnes
   Mtons    millions of short tons    Mtonnes    millions of metric tonnes
   lb    pound    kg    kilogram
   oz    ounce    g    gram
   koz    kilo-ounce
   toz, troz, troy oz    troy ounce
Grade    opt, opst    troy ounces per short ton    g/t, gpt    grams per tonne
   opmt    troy ounces per metric tonne
Volume    ft3    cubic feet    m3    cubic meter
   gal    gallons    L    liter
Volumetric Flow Rate    gpm    gallons per minute    Lpm    liters per minute
   scfm    standard cubic feet per minute    m3/hr    cubic meters per hour
Density    lb/ft3    pounds per cubic foot    t/m3    tonnes per cubic meter
   sg    specific gravity    sg    specific gravity
Percent Solids    wt%    percent solids by weight    wt%    percent solids by weight
Work Index (Hardness)    kWh/st    kilowatt-hours per short ton    kWh/t    kilowatt-hours per tonne
Elevation    amsl    above mean sea level          
   fasl    feet above sea level    masl    meters above sea level
Throughput    st/h, stph    short tons per hour    t/h, tph    metric tonnes per hour
   st/d, stpd    short tons per day    t/d, tpd, mtpd    metric tonnes per day
   st/y, stpy    short tons per year    t/y, tpy    metric tonnes per year
   kst/y, kstpy    thousand short tons per year          
Temperature    °F    degrees Fahrenheit    °C    degrees Celsius
Concentration    ppm    parts per million    mg/L    milligrams per liter
   g/L    grams per liter
Power    hp    horsepower    kW    kilowatt
   kW-hr    kilowatt hour
   MW    megawatt
Work Index    kWh/st    kilowatt hour per short ton    kWh/t    kilowatt hour per metric tonne
Mill Speed    rpm    revolutions per minute    rpm    revolutions per minute
Pressure    psi    pounds per square inch    kPa    kilopascal
   mPa    megapascal
Voltage    kV    kilovolt    kV    kilovolt
   kVA    kilovolt-amperes    kVA    kilovolt-amperes

 

2.7

Coordinate System

Spatial data utilized in the analysis presented in this PEA are projected in the Ruby Hill Mine Grid (local grid, ft) and UTM NAD83 Zone 11 North (ft). The project centroid location (derived from the geological model) is 9495,115158 in the Ruby Hill Mine Grid, and 1925147,14352286 in UTM NAD83 Z11N.

 

 

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2.8

Mineral Resource and Mineral Reserve Definitions

The terms “mineral resource” and “mineral reserves” as used in this TRS have the following definitions:

Mineral Resources

7 CFR § 229.1300 defines a “mineral resource” as a concentration or occurrence of material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. A mineral resource is a reasonable estimate of mineralization, taking into account relevant factors such as cut-off grade, likely mining dimensions, location or continuity, that, with the assumed and justifiable technical and economic conditions, is likely to, in whole or in part, become economically extractable. It is not merely an inventory of all mineralization drilled or sampled.

A “measured mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of conclusive geological evidence and sampling. The level of geological certainty associated with a measured mineral resource is sufficient to allow a qualified person to apply modifying factors, as defined in this section, in sufficient detail to support detailed mine planning and final evaluation of the economic viability of the deposit. Because a measured mineral resource has a higher level of confidence than the level of confidence of either an indicated mineral resource or an inferred mineral resource, a measured mineral resource may be converted to a proven mineral reserve or to a probable mineral reserve.

An “indicated mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of adequate geological evidence and sampling. The level of geological certainty associated with an indicated mineral resource is sufficient to allow a qualified person to apply modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Because an indicated mineral resource has a lower level of confidence than the level of confidence of a measured mineral resource, an indicated mineral resource may only be converted to a probable mineral reserve.

An “inferred mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. The level of geological uncertainty associated with an inferred mineral resource is too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability. Because an inferred mineral resource has the lowest level of geological confidence of all mineral resources, which prevents the application of the modifying factors in a manner useful for evaluation of economic viability, an inferred mineral resource may not be considered when assessing the economic viability of a mining project and may not be converted to a mineral reserve.

Mineral Reserves

17 CFR § 229.1300 defines a “mineral reserve” as an estimate of tonnage and grade or quality of indicated and measured mineral resources that, in the opinion of the qualified person, can be the basis of an economically viable project. More specifically, it is the economically mineable part of a measured or indicated mineral resource, which includes diluting materials and allowances for losses that may occur when the material is mined or extracted. A “proven mineral reserve” is the economically mineable part of a measured mineral resource and can only result from conversion of a measured mineral resource. A “probable mineral reserve” is the economically mineable part of an indicated and, in some cases, a measured mineral resource.

 

 

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3.

PROPERTY DESCRIPTION

 

3.1

Property Description

The Ruby Hill Complex is in Eureka County, Nevada, 1.5 miles northwest of the town of Eureka, and it is part of the historic Eureka mining district. It is centered at roughly 39°31.5’ N latitude and 115°59’ W longitude. The Complex is owned by Ruby Hill Mining LLC and Golden Hill Mining Corporation, both are wholly owned subsidiaries of i-80. The northern part of the project, including the Archimedes pit, Archimedes underground, and Mineral Point deposit, is referred to as Ruby Hill and the southern part of the project, containing the historic Archimedes Underground mine and the FAD deposit, is referred to as Golden Hill. Ruby Hill encompasses about 10,608 acres and Golden Hill about 3,229 acres, together totaling about 13,837 acres (56,004 hectares) including owned patented and unpatented claims, owned surface fee land, and owned and leased unpatented claims. The federal land is administered by the US Department of Interior - Bureau of Land Management. Figure 3-1 shows the location of the Ruby Hill Project.

 

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Figure 3-1: Ruby Hill Complex Location Map

(Source: i-80 Gold, 2024)

 

3.2

Status of Mineral Titles

The Ruby Hill Complex land position comprises various forms of title. Figure 3-2 shows the Ruby Hill Complex land position. On the northern Ruby Hill portion of the property, i-80, through its wholly owned subsidiaries Ruby Hill Mining Company LLC and Golden Hill Mining Corporation, owns 34 patented claims (Table 3-1), 640 unpatented claims (Table 3-2), and leases seven unpatented lode claims (Table 3-3). The lease expires May 12, 2032, and may be renewed by notice. i-80 also owns a land patent covering about

 

 

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1,644.5 acres (665.6 hectares) in the vicinity of the Archimedes and Mineral Point deposits. The mineral rights underlying the patented land are held by patented and unpatented lode claims.

 

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Figure 3-2: Ruby Hill Complex Land Position

(Source: i-80 Gold, 2024)

 

 

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Table 3-1: Ruby Hill Project Owned Patented Claims

 

Claim Name  Mineral Survey  Number   Patent  Number   Claim Type 

 Number of

Claims

Bullwhacker

  51   1264   Millsite       1

Cyanide

  4686   1753   Lode   1

Vera Cruz and California

  76   1772   Lode   1

Alabama

  106   2075   Lode   1

Hoosac

  60   2115   Lode   1

Wide West

  105   2193   Lode   1

Racine

  89   2485   Lode   1

General Lee

  120   2531   Lode   1

Williamsberg

  117   2618   Lode   1

Holly Lode

  122   3850   Lode   1

Bowman

  175   4228   Lode   1

Little Giant

  192   4304   Lode   1

Price, Price No. 2

  228, 229   4410, 4411   Lode   2

Oriental and Belmont

  196   4511   Lode   1

Europa Consol.

  176   4622   Lode   1

Fredrika

  269   7023   Lode   1

Belle of the West NO. 2

  271   8024   Lode   1

Central Consolidated

  268   8066   Lode   1

Minerva, Silver Bill and Diagonal

  292, 255   9783, 9784   Lode   2

Members No. 2

  281   11490   Lode   1

Protection

  300   11552   Lode   1

Lone Pine

  4686   17513   Lode   1

Morning Star, Macon City

  249, 250   18852, 18853   Lode   2

Democrat

  310   20068   Lode   1

Horizontal, Herculean

  316, 317   22273, 22274   Lode   2

Margarita

  1946   40910   Lode   1

Porphyry, Quartzite

  3596   179187   Lode   2

Silver Lick and Bobbie Burns Consol.

  75   Lode   1

Silver West

  131   Lode   1

Total Owned Patented Claims

  34

 

 

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Table 3-2: Ruby Hill Project Owned Unpatented Claims

 

Claim Name

BLM Serial Number

 Claim 
Type
Number of
Claims

ESPH 1- ESPH 85

NMC1076732 - NMC1076816

Lode       85

TDB 1 - TDB 12

NMC1089497 - NMC1089508

Lode   12

LH-1 - LH-25

NMC483711 - NMC483735

Lode   25

LH-27 - LH-77 

NMC483737 - NMC483787

Lode   51

RH-5

NMC489850

Lode   1

SP#1 - SP#37

NMC604319 - NMC604355

Lode   37

SP#37A, SP#38, SP#38A

NMC604356 - NMC604358

Lode   3

SP#39 - SP#45

NMC604359 - NMC604365

Lode   7

SP#51 - SP#58

NMC604371 - NMC604378

Lode   8

LH#98 - LH#120, LH#140

NMC606475 - NMC606498

Lode   24

LH 130, LH 132, LH 134 - LH 136

NMC615733 - NMC615737

Lode   5

LH 139, LH 141

NMC615740 - NMC615741

Lode   2

PLS#37 - PLS#42

NMC676560 - NMC676565

Lode   6

PLS#66 - PLS#94

NMC676589 - NMC676617

Lode   29

PLS#236 - PLS#245, PLS 246 - PLS 248

NMC676759 - NMC676771

Lode   13

PLS 255, PLS 264

NMC676778 - NMC676787

Lode   2

HMC 11 - HMC 12

NMC677967 - NMC677968

Lode   2

WLH#9 - WLH#42

NMC681558 - NMC681591

Lode   34

WLH#85 - WLH#91

NMC681634 - NMC681640

Lode   7

PLS 265 - PLS 273

NMC682320 - NMC682328

Lode   9

PLS 275, PLS 277

NMC682330, NMC682332

Lode   2

PLS 285 - PLS 292

NMC682340 - NMC682347

Lode   8

HMC 15 - HMC 24, HMC 33 - HMC 38

NMC683512 - NMC683527

Lode   16

LH 78A - LH 87A

NMC683528 - NMC683537

Lode   10

HOPE, HOPE1 - HOPE11

NMC699711 - NMC699722

Lode   12

HOPE13 - HOPE21, HOPE EXTENSION

NMC699724 - NMC699733

Lode   10

HOPE EXTENSION 1 - HOPE EXTENSION 12

NMC699734 - NMC699745

Lode   12

JANUARY, JULY NO. 1, JULY NO. 2

NMC699746 - NMC699748

Lode   3

CUB, CUB NO. 1

NMC699749 - NMC699750

Lode   2

AUGUST 3 - AUGUST 6

NMC699751 - NMC699754

Lode   4

AUGUST 8 - AUGUST 9, ADAMS HILL EXTENSION

NMC699756 - NMC699758

Lode   3

ADAMS HILL EXTENSION NO. 1 through 7

NMC699759 - NMC699765

Lode   7

ADAMS HILL EXT. 8 - ADAMS HILL EXT. 10

NMC699766 - NMC699768

Lode   3

CYANIDE EXTENSION NO. 7, CYANIDE NO. 8

NMC699769 - NMC699770

Lode   2

CYANIDE EXTENSION NO. 13, CYANIDE NO. 14

NMC699771 - NMC699772

Lode   2

CYANIDE EXTENSION NO. 16, 17, 24, 25, 26, 27

NMC699773 - NMC699778

Lode   6

SAGEBRUSH, SAGEBRUSH 1, HOLLY 2

NMC699779 - NMC699781

Lode   3

MARCH EXT. 2 - MARCH EXT. 6, SEPTEMBER

NMC699802 - NMC699807

Lode   6

SEPTEMBER 1 - SEPTEMBER 3

NMC699808 - NMC699810

Lode   3

SEPTEMBER 5 - SEPTEMBER 10

NMC699811 - NMC699816

Lode   6

DECEMBER 7 - DECEMBER 10

NMC699818 - NMC699821

Lode   4

OCTOBER FRACTION, NOVEMBER

NMC699822 - NMC699823

Lode   2

NOVEMBER 1, NOVEMBER 2, NOVEMBER FRACTION

NMC699824 - NMC699826

Lode   3

ARC 1 - ARC 41

NMC699827 - NMC699867

Lode   41

ARC 43 - ARC 58

NMC699869 - NMC699884

Lode   16

R-E 10, R-E 15, R-E 20

NMC699892, NMC699897, NMC699902

Lode   3

R-E 25 - R-E 26

NMC699907 - NMC699908

Lode   2

R-E 31, R-E 34

NMC699911 - NMC699912

Lode   2

JAY 22, JAY 24, JAY 26

NMC699964, NMC699966, NMC699968

Lode   3

SNOW, SNOW 1 - SNOW 5

NMC704357 - NMC704362

Lode   6

MARCH, MARCH #1 - MARCH #3

NMC704363 - NMC704366

Lode   4

MARCH 4 - MARCH 11

NMC704367 - NMC704374

Lode   8

MARCH EXT, MARCH EXTENSION #1

NMC704375 - NMC704376

Lode   2

JAY # 23, JAY # 25, JAY # 27, HOPE # 12

NMC705154 - NMC705157

Lode   4

AUGUST # 7, SEPTEMBER # 11

NMC705158 - NMC705159

Lode   2

ARC 62

NMC713810

Lode   1

PLS # 279, PLS # 281

NMC771503 - NMC771504

Lode   2

LH 137R, LH 138R

NMC832613 - NMC832614

Lode   2

RHMS 300 - RHMS 350

NMC909518 - NMC909568

Millsite   51

Total Owned Unpatented Claims

  640

 

 

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Table 3-3: Ruby Hill Project Leased Unpatented Claims

 

       
Claim Name  BLM Serial Number  Claim 
Type
Number
of
Claims

SWAN

NMC72580

Lode

 

1

MERIT

NMC72581

Lode

 

1

GOLD QUARTZ, GOLD QUARTZ #1, GOLD QUARTZ #2

NMC72582 - NMC72584

Lode

 

3

WEST #1, WEST #2

NMC72586, NMC72587

Lode

 

2

Total Leased Unpatented Claims

 

7

On the Golden Hill portion of the property, i-80 owns 105 patented lode and millsite claims (Table 3-4), leases 5 patented claims (Table 3-5), owns 149 unpatented lode claims (Table 3-6), and leases seven unpatented lode claims (Table 3-7). The lease on the unpatented claims expires May 12, 2032, and may be renewed by notice.

Table 3-4: Golden Hill FAD Property Owned Patented Claims

 

         
Claim Name Mineral Survey
Number
Patent Number  Claim 
 Type 
Number
of
Claims

SENTINEL, MAMMOTH

 

40, 41

 

382, 383

 

Lode

 

    2

BUCKEYE, CHAMPION

 

37, 38

 

389, 390

 

Lode

 

2

SAVAGE, LOOKOUT

 

42, 43

 

391, 392

 

Lode

 

2

CARSON

 

68

 

882

 

Lode

 

1

RICHMOND, TIP-TOP

 

64, 65

 

885, 886

 

Lode

 

2

SKYLARK, CALLOWAY

 

56, 57

 

1120, 1121

 

Lode

 

2

IONE LODE, GRANT LODE

 

74, 73

 

1221, 1222

 

Lode

 

2

SURPLUS LODE, PORTER, BROWN

 

85, 86, 87

 

1581, 1582, 1583

 

Lode

 

3

NUGET

 

46

 

2066

 

Lode

 

1

WILSON, JACKSON

 

97, 98

 

2109, 2110

 

Lode

 

2

LUPITA

 

49

 

2204

 

Lode

 

1

ST. GEORGE

 

66

 

2265

 

Lode

 

1

SILVER STATE MINE, ORIGINAL BALTIC MINE

 

111, 112

 

2296, 2297

 

Lode

 

2

MARCELINA EAST

 

119

 

2830

 

Lode

 

1

AT LAST

 

47

 

2968

 

Lode

 

1

BUCKEYE MILLSITE, CHAMPION MILLSITE

 

113, 114

 

3607, 3608

 

Millsite

 

2

BROWN MILLSITE

 

139

 

3742

 

Millsite

 

1

SILVER REGION, VICTORIA

 

160, 161

 

3751, 3755

 

Lode

 

2

GRAND CENTRAL

 

174

 

4077

 

Lode

 

1

PORTER MILLSITE, CARSON MILLSITE

 

138, 137

 

4197, 4198

 

Millsite

 

2

CONNELL

 

190

 

4310

 

Lode

 

1

DAVIES, DAVIES NO. 2

 

230, 231

 

4414, 4415

 

Lode

 

2

DIAGONAL, GREAT EASTERN

 

200, 165

 

4546, 4555

 

Lode

 

2

PEACH, MARRIAGE AMENDED, LA VETA

 

2869, 2867, 2873

 

4567, 4568, 4569

 

Lode

 

3

T.R., HONEYMOON AMENDED, GULCH

 

2870, 2868, 2872

 

4570, 4571, 4572

 

Lode

 

3

ALBION NO. 1, REMNANTS, FAD

 

2860, 3252, 3223

 

4573, 4574, 4575

 

Lode

 

3

APEX, ACOUCHMENT, BIG TR

 

2865, 2866, 2871

 

4576, 4577, 4578

 

Lode

 

3

ALBION NO. 2, ARCTIC, CLIFF MINE

 

2861, 2857, 2856

 

4579, 4580, 4581

 

Lode

 

3

ALBION NO. 3, LUCKY MAN, RAVINE

 

2862, 2852, 2858

 

4582, 4583, 4584

 

Lode

 

3

MAIN SHAFT, ATLANTIC, ANTARCTIC

 

2864, 2854, 2855

 

4586, 4587, 4588

 

Lode

 

3

ALBION CONSOLIDATED

 

2863

 

4589

 

Lode

 

1

RICHMOND RANCHO

 

211

 

4714

 

Lode

 

1

HOPE CONSOLIDATED

 

206

 

4800

 

Lode

 

1

SURPLUS MILLSITE

 

141

 

4923

 

Millsite

 

1

BADGER

 

218

 

5558

 

Lode

 

1

ISANDULA

 

213

 

5677

 

Lode

 

1

JACK & SCANLAND

 

217

 

6057

 

Lode

 

1

SKYLARK MILLSITE

 

214

 

6093

 

Millsite

 

1

GREEN SEAL

 

167

 

6169

 

Lode

 

1

WESTERN & WINCHESTER

 

216

 

6412

 

Lode

 

1

DON RICARDO

 

274

 

7415

 

Lode

 

1

REAR GUARD

 

225

 

7528

 

Lode

 

1

 

 

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Claim Name Mineral Survey
Number
Patent Number  Claim 
 Type 
Number
of
Claims

REARGUARD MILLSITE

 

225

 

7528

 

Millsite

 

1

PRIDE OF THE WEST

 

267

 

7582

 

Lode

 

1

KEMP & KEEN

 

265

 

7886

 

Lode

 

1

GERALDINE LODE

 

284

 

8023

 

Lode

 

1

CENTRAL HILL

 

273

 

8097

 

Lode

 

1

ST. ANDREW LODE

 

242

 

9451

 

Lode

 

1

ST. PATRICK LODE

 

241

 

9640

 

Lode

 

1

TINNIE

 

195

 

10012

 

Lode

 

1

CHARTER

 

297

 

10344

 

Lode

 

1

PHIL SHERIDAN

 

270

 

15562

 

Lode

 

1

MONARCH 2, MONARCH 3, RICHMOND EXTENSION, RICHMIND EXTENSION NO. 1 through 4, RICHMOND FRACTION, RUBY HILL FRACTION, RUBY HILL NO. 1, RUBY HILL NO. 2

 

4686

 

17531

 

Lode

 

11

FITZGERALD LODE

 

313

 

19065

 

Lode

 

1

MAUD C.

 

307

 

19166

 

Lode

 

1

FRIES, FRANK

 

308, 309

 

19815, 19816

 

Lode

 

2

FEBRUARY, NOVEMBER, SHALE

 

3596

 

179187

 

Lode

 

3

ADAMS AND FERREN AND DEEP MINE

 

116

 

Lode

 

1

HARLEM AND EUREKA BELLE CON.

 

262

 

Lode

 

1

PATROON AND GRAND DELIVERY CONS.

 

261

 

Lode

 

1

ST. ANDREW MILLSITE

 

242

 

Millsite

 

1

ST. DAVID, AKA ST. DAVID MINE

 

2859

 

Lode

 

1

ST. PATRICK MILLSITE

 

241

 

Millsite

 

1

Total FAD Owned Patented Claims

 

105

Table 3-5: Golden Hill FAD Property Leased Patented Claims

 

           
Claim Name Mineral
Survey
Number
Patent Number Claim Type

 Number 
of

Claims

Expiration/
Renewal
Date

CONTINENTAL

 

212

 

5684

 

Lode

1 

 

June 16,
2032


INDEPENDENT

 

248

 

6008

 

Lode

1 

 

June 16,
2032


STAR OF THE WEST

 

7981

 

Lode

1 

 

May 22,
2032


SHOO FLY NO. 2, SHOO FLY NO. 3

 

58, 59

 

2294, 2295

 

Lode

2 

 

June 9,
2032


Total FAD Leased Patented Claims

5 

 

June 9,
2032


 

           

Claim Name

Mineral Survey
Number

Patent

Number

Claim

Type

Number of
Claims
Expiration/
Renewal Date

CONTINENTAL

 

212

 

5684

 

Lode

 

1

 

June 16, 2032

INDEPENDENT

 

248

 

6008

 

Lode

 

1

 

June 16, 2032

STAR OF THE WEST

 

7981

 

Lode

 

1

 

May 22, 2032

SHOO FLY NO. 2, SHOO FLY NO. 3

 

58, 59

 

2294, 2295

 

Lode

 

2

 

June 9, 2032

Total FAD Leased Patented Claims

 

5

 

June 9, 2032

 

 

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Table 3-6: Golden Hill FAD Property Owned Unpatented Claims

 

       
Claim Name    BLM Serial Number    Claim 
Type 
   Number
of Claims
 

ESPH 86- ESPH 96

  

NMC1076817 - NMC1076827

  

Lode 

  

 

11

 

HMC 50

  

NMC1078382

  

Lode 

  

 

1

 

TDB 13 - TDB 57

  

NMC1089509 - NMC1089553

  

Lode 

  

 

45

 

SRH 35 - SRH 36

  

NMC1094131 - NMC1094132

  

Lode 

  

 

2

 

RH – 1 - RH – 4

  

NMC489846 - NMC489849

  

Lode 

  

 

4

 

SP #46 - SP #50

  

NMC604366 - NMC604370

  

Lode 

  

 

5

 

HMC 3 - HMC 4

  

NMC661367 - NMC661368

  

Lode 

  

 

2

 

HMC 6, HMC 8, HMC 9

  

NMC661370 - NMC661372

  

Lode 

  

 

3

 

HMC 39

  

NMC699710

  

Lode 

  

 

1

 

ARC 42

  

NMC699868

  

Lode 

  

 

1

 

ARC 59 - ARC 60

  

NMC699885 - NMC699886

  

Lode 

  

 

2

 

RE-3A, R-E 6 - R-E 9

  

NMC699887 - NMC699891

  

Lode 

  

 

5

 

R-E 11 - R-E 14

  

NMC699893 - NMC699896

  

Lode 

  

 

4

 

R-E 16 - R-E 19

  

NMC699898 - NMC699901

  

Lode 

  

 

4

 

R-E 21 - R-E 24

  

NMC699903 - NMC699906

  

Lode 

  

 

4

 

R-E 27, R-E 30

  

NMC699909 - NMC699910

  

Lode 

  

 

2

 

ANN 16 - ANN 20

  

NMC699913 - NMC699917

  

Lode 

  

 

5

 

SRH 1 - SRH 6

  

NMC699918 - NMC699923

  

Lode 

  

 

6

 

SRH 8

  

NMC699925

  

Lode 

  

 

1

 

SRH 10 - SRH 12

  

NMC699927 - NMC699929

  

Lode 

  

 

3

 

SRH 14 - SRH 26

  

NMC699930 - NMC699942

  

Lode 

  

 

13

 

SRH 28 - SRH 32

  

NMC699943 - NMC699947

  

Lode 

  

 

5

 

SRH 34

  

NMC699948

  

Lode 

  

 

1

 

JAY 1 - JAY 8

  

NMC699949 - NMC699956

  

Lode 

  

 

8

 

JAY 11 - JAY 14

  

NMC699957 - NMC699960

  

Lode 

  

 

4

 

JAY 18 - JAY 19

  

NMC699961 - NMC699962

  

Lode 

  

 

2

 

ARC #61, JAY #9, JAY #20

  

NMC705151, NMC705152, NMC705153

  

Lode 

  

 

3

 

ARC 63

  

NMC713811

  

Lode 

  

 

1

 

SRH 27

  

NMC808229

  

Lode 

  

 

1

 

Total FAD Owned Unpatented Claims

       

 

149

 

Table 3-7: Golden Hill FAD Property Leased Unpatented Claims

 

       
Claim Name    BLM Serial Number    Claim Type    Number
of
Claims

WEST NO. 3 - WEST NO. 5

  

NMC661796 - NMC661798

  

Lode

  

3

WEST, WEST EXTENSION

  

NMC72585, NMC72591

  

Lode

  

2

HMC 2, HMC 5

  

NMC661366, NMC661369

  

Lode

  

2

Total FAD Leased Unpatented Claims

       

7

Patented land is subject to property taxes and lease holding payments to the claim owner if applicable. Unpatented claims have annual maintenance fees of $200 per claim payable to the Bureau of Land Management and a notice of intent to hold (NIH) in the amount of $12 per claim payable to Eureka County. The BLM MLRS mining claim database shows all claim fees paid through September 2025. The NIH was paid to Eureka County on July 10, 2024. All claim fees are current. Annual property holding costs for the Ruby Hill Complex are listed in Table 3-8.

 

 

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Table 3-8: Ruby Hill Complex Property Holding Costs

 

       
Description    Payee          Quantity       Amount    

Unpatented Claim Maintenance Fee

  

BLM

  

803

  

 

$160,600.00

 

Notice of Intent to Hold Unpatented Claims

  

Eureka County

  

803

  

 

$9,660.00

 

Patented Claim Property Taxes

  

Eureka County

  

139

  

 

$1,356.24

 

Patented Claim Property Taxes 5 leased

  

Eureka County

  

5

  

 

$52.35

 

Real Property Taxes

  

Eureka County

  

1

  

 

$103,967.73

 

Collingwood Ranch Property Taxes

  

Eureka County

  

1

  

 

$1,007.88

 

Personal Property Taxes

  

Eureka County

  

1

  

 

$1,216.51

 

Water Leases

  

Lease Holders

  

multiple

  

 

$28,700.00

 

Yearly Mining Claim Lease Payments

  

Lease Holders

  

multiple

  

 

$8,250.00

 

Total

            

 

$314,810.71

 

 

3.3

Royalties

Several royalties are in effect on various areas of the property. Table 3-9 lists the royalties in the Ruby Hill area, and Table 3-10 lists the royalties in the Golden Hill area. Figure 3-3 shows the royalty areas. Some royalties were retained by previous owners upon sale of the property while others were negotiated as lease agreements with claim holders. Royalties are not payable until production occurs in the area covered by the royalty.

Table 3-9: Ruby Hill Royalties

 

   
Lessor/Grantor   Lease Type

ASARCO Incorporated

 

4% NSR

RG Royalties, LLC

 

3% NSR

Arthur A. & Elizabeth O. Biale Trust

 

3% NSR

Placer Dome

 

2.5% NSR

Table 3-10: Golden Hill Royalties

 

   
Lessor/Grantor   Lease Type

ASARCO Incorporated

 

4% NSR

Biale Lease

 

3% NSR

Herrera Lease

 

4% NSR

MacKenzie Lease (50% Interest)

 

2% NSR

Warren Lease

 

4% NSR

RG Royalties

 

3% NSR

Royalty Consolidation Company

 

0.5-1.5% NSR

 

 

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Figure 3-3: Ruby Hill Royalty Map

(Source: i-80 Gold, 2024)

 

 

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3.4

Environmental Liabilities

The closure cost for Ruby Hill is estimated to be $27 million (i-80, 2025). The associated Bond was accepted by the Bureau of Land Management (BLM) on August 8, 2023 and covers authorized disturbance associated with issued permits for Ruby Hill (RHMC 2023). There are no other known environmental liabilities associated with pre-Project operations (RHMC, 2021).

RHMC controls a total of 8,107-acre feet per annum (AFA) of water rights for consumption and occupation (RHMC, 2024).

Due to a history of over pumping in the region based on a heavy agricultural reliance, the Diamond Valley Basin was categorized as a Critical Management Area (CMA) by the Nevada State Engineer’s office in 2015. The designation allowed the State Engineer and the community to agree on certain tools to reduce over-pumping, including the implementation of a Diamond Valley Groundwater Management Plan (GMP). Following resolution of a lengthy legal dispute by senior water rights holders in the Basin, the GMP was reinstated effective January 1, 2023. As a groundwater user within the GMP designated area, RHMC controls sufficient water rights to support its mining operations (RHMC, 2024).

 

3.5

Permits/Licenses

In conjunction with the permitting actions associated with the Archimedes Underground Mine in-pit surface support facilities, a Determination of NEPA Adequacy (DNA) was deemed sufficient for the Plan of Operations (PoO) Amendment NVN-067782 approved by the BLM March 30, 2023. Additionally, on June 23, 2023, the Nevada Division of Environmental Protection – Bureau of Mining Regulation & Reclamation (NDEP-BMRR) approved an Engineering Design Change (EDC) to Water Pollution Control Permit (WPCP) NEV0096103 for the construction of the surface facilities. Permitting actions tied to mining of the underground are currently in progress with the BLM evaluating a PoO Amendment and associated Environmental Assessment (EA) while NDEP-BMRR is analyzing a WPCP Major Modification.

 

 

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4.

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

 

4.1

Accessibility

The Ruby Hill Project area is a 4.5-mile drive from the town of Eureka, Nevada. From the intersection of Clark Street and US Highway 50, travel north on Hwy 50 3.2 miles to the junction of Nevada State Route 278 on the right and the Homestake Road turn-off to the left. Turn left, and travel south 1.3 miles on a well-graded gravel road to the Ruby Hill gate. Eureka is located on Highway 50, about 242 miles east of Reno via Interstate 80 and Hwy 50, or 92 miles south of Carlin, via Nevada State Route 278.

 

4.2

Climate

The climate in Eureka County is typical of the high-desert environment. Typical summer temperatures near Eureka range between 50°F and 82°F while winter temperatures range between 18°F and 38°F. Average precipitation is about 11.8 inches including just under 59 inches of snowfall. Typical snow accumulation is roughly 3 inches on average at lower elevations, although occasional large storms may accumulate significantly more for short durations. The town of Eureka lies at about 6485 ft elevation, while the project area ranges from 6160 ft to 6680 ft. The FAD shaft to the south of the Project sits at about 6900 ft elevation.

Mining operations are able to continue year-round with brief pauses for summer lightning storms or unusually heavy winter snowstorms.

 

4.3

Local Resources

The town of Eureka has a population of about 410. Basic services are available. The Eureka Mining District has a long history of mining activity, and mining suppliers and contractors are accustomed to working in the area. Some experienced and general labor is available locally, and some may be sourced regionally from the towns of Elko (114-mile drive north of the Project), Reno (242 miles west of the Project), Ely (78 miles east of the Project), and other small towns in the region. There are a number of mining operations in the region and as such, there is always competition for employees.

 

4.4

Infrastructure

 

   

Electricity – The local utility company is NVEnergy. There is sufficient electrical energy at the site for all planned operations.

   

Labor – There are numerous operating mines in northern Nevada and a skilled labor force is available.

   

Supplies – Local suppliers can provide all materials necessary to support the planned mining operations.

   

Water – The Ruby Hill project can supply sufficient water from existing wells to support all planned mining operations.

 

4.5

Physiography

The Project lies in the Basin and Range Province, a structural and physiographic province comprised of generally north to north-northeast trending, fault bounded mountain ranges separated by alluvial filled valleys.

 

 

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The Project is located on the northern flank of the Fish Creek Range sloping towards Diamond Valley. Topography is gentle to moderate with steeper hills to the south in the FAD area. Vegetation is typical of the high desert with sagebrush on the alluvial fans, and piñon and juniper on the mountain slopes.

 

 

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5.

HISTORY

 

5.1

Historic Ownership

The Ruby Hill Project is located in the northern portion of the historic Eureka mining district. Prospecting began in the 1860’s with production occurring by the 1870’s from carbonate replacement (CRD) type deposits. Much of the historic work occurred south of the Ruby Hill Project, with historic mines and prospects scattered through the northern Fish Creek mountains west of Eureka extending south several miles towards the Little Smoky Valley. The majority of historical production, estimated at about 80%, was from the original Ruby Hill Mine (Nolan, 1962 and Nolan and Hunt, 1968). The original Ruby Hill mine is located roughly 1 1⁄4 miles south of the Ruby Hill Project. i-80 merged with Golden Hill Corporation and acquired the historic Ruby Hill site (FAD property), consolidating their ownership of the Ruby Hill Complex.

Modern work at the Project began in 1992, when Homestake Mining Company made the Archimedes Carlin-type discovery at the current Project area. Table 5-1 lists the general history of ownership, exploration, and mining of the larger, recently consolidated Ruby Hill complex. History prior to 1992 is focused generally on the FAD portion of the property, while history from 1992 to the present is focused on the Archimedes area. One exception is the TL shaft and associated historic underground mines, which Eureka Corporation mined from 1953-1958 in the vicinity of the current Mineral Point Trend resource area.

Table 5-1: Historic Regional Ownership and Activities

 

Year    Company    Comment

1864

  

Various

  

Oxidized gold-silver CRD mineralization discovered by prospectors

1869

  

Various

  

Ruby Hill CRD mineralization discovered on Prospect Mountain

  

W.W. McCoy devises furnace for recovering metals from oxidized ores

1873-1905

  

Richmond Mining Company

  

Production from the Ruby Hill deposit. Smelting ceases 1890

1873-1916

  

Eureka Consolidated Mining Company

  

Production from the Ruby Hill deposit

  

The Locan shaft was sunk to 1200 level. High water flow encountered in crosscut partially flooding shaft. Shaft dewatering unsuccessful, mine shut down

  

Smelting ceased 1891

1905-1912

  

Richmond-Eureka Mining Company

  

Richmond Mining Company and Eureka Consolidated Mining Company properties consolidated into Richmond-Eureka Mining Company

  

Controlling interest held by Unites States Smelting, Refining, and Mining Company (USSRAM)

  

Rehabilitation of Richmond and Eureka consolidated mines. Processing of stope fill and low-grade ore

1919

  

Ruby Hill Development Company

  

Leased property from Richmond-Eureka Mining Company. Dewatered Locan shaft

  

Project abandoned due to exhaustion of finances

1923

  

Richmond-Eureka Mining Company

  

Dewatered Locan shaft to 1,200 level

  

Drove SE crosscut to Ruby Hill fault, and a drift to SW. SW drift encountered high water flow and work stopped

  

Vertical exploration hole (type unknown) drilled from 900 level. Hole caved, and project abandoned

1920’s -

   Various lessors    Sporadic production

1930’s

1937-1959

  

Eureka Corporation, Ltd.

   Obtained leases on Ruby Hill property from Richmond-Eureka Mining Company
   Completed 4 churn holes (totaling 3,596 feet), 260 surface and underground core holes (87,633.8 feet), 13 mud rotary holes (14,252 feet), and 6 RC holes (9,903 feet)
   Intersection of high-grade polymetallic mineralization in 5 surface core holes led to the FAD shaft being sunk to 2,500’ depth to develop mineralization. Underground development encountered high water flow which flooded shaft
   Rotary drilling in 1953 in Adams Hill area intersected mineralization in Hamburg Dolomite
   Sinking of the T.L. shaft started in 1953 to exploit mineralization and was completed in 1955 to a depth of 1,127 feet *(This lies above, and locally intersects, the current Mineral Point resource.)

 

 

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Year    Company    Comment
        Mining commenced in 1956 and shut down in 1958 due to lack of ore

1989-1991

   American Smelting and Refining Company (ASARCO)    Drilled 12 RC exploration holes totaling 5,314 feet

1960-1992

   Ruby Hill Mining Company    Richmond-Eureka Mining Company (75%) and Eureka Corporation (25%) form Ruby Hill Mining Company
   In June 1960 a consortium was formed consisting of Richmond-Eureka Mining Company, Eureka Corporation, Newmont Mining Company, Cyprus Mines Corporation, and Hecla Mining Company to finance additional drilling and produce a FAD feasibility study
   Collectively, Consortium drilled 148 exploration holes (129,362.3 feet); 13 churn (3,641 feet); 33 Mud Rotary (74,039 feet); 6 percussion (395 feet); 3 RC (1,458 feet); and 93 core holes (50,218.3 feet)
   Fourteen holes drilled in FAD shaft area intersected mineralization. Decision made to dewater FAD shaft to exploit new mineralization
   In 1963 FAD shaft was dewatered to the 2250 level. New crosscut, 1,028’ long, to evaluate mineralized zone completed in 1964. Crosscut used to drill exploration percussion and core holes
   Drilling completed in 1966 and mine placed on inactive status pending economic evaluation
   1966 and 1974 Hecla feasibility studies indicate project not feasible
   In 1974 Newmont withdrew from the consortium followed by Hecla in 1979
   Cyprus remains as surviving partner drilling 39 mud rotary (7,945 feet), and 98 air track (4,983 feet) exploration holes for near-surface, bulk-mineable gold mineralization between 1980-1981
   Exploration unsuccessful and property reverted to Sharon Steel Corporation successor to Ruby Hill Mining Company in 1982
   Sharon Steel Corporation drilled 127 exploration/definition RC holes totaling 31,539 ft between 1982 and 1991

1993-1994

   Placer Dome    Drilled 11 RC exploration holes (12,350 feet) at Ruby Flats

1994

   Unknown    Drilled 1 RC hole for 500 feet

1992-2001

   Homestake Mining Company    Homestake acquired Ruby Hill property from Ruby Hill Mining Company in 1992
   Exploration/definition drilling between 1992-1993 discovered/defined the Archimedes deposit (both West and East) along with the 426 zone
   In 1994 Homestake announced plans to develop an open pit mine and processing facility to exploit West Archimedes mineralization. Construction began in 1997 and production commenced in 1998
   The eastern portion of the Archimedes deposit (East Archimedes) not developed due to low gold prices, high strip ratio, change of mineralization from oxide to sulfide, and mineralization largely below water table creating permitting issues
   Mining ceased in 2002 and reclamation activities started on mine waste dumps and pit area
   Completed 1,502 (1,022,842.5 feet) exploration/definition holes between 1992-2001; 1374 RC holes (875,083 feet), and 128 core holes (147,759.5 feet)
   DIGHEM Surveys conducted an airborne magnetic & electromagnetic survey in 1994 on E-W flight lines at nominal 600’ spacing with mean terrain clearance of 115 feet
   Zonge Geosciences completed ground magnetics survey at 150’ spacing in 2000.
   In 1998, conducted dump sampling program on Diamond Tunnel dump to evaluate grade and tonnage (south of property)
   Between 1999-2000 conducted rock chip sampling program to determine potential for multi element correlation as pathfinder for gold
2001-2015    Barrick Gold Corporation    Barrick acquired Ruby Hill property during 2001 merger with Homestake Mining Company
   In 2002 Chadwick and Russell completed Archimedes pit mapping
   Completed positive feasibility study on East Archimedes deposit in 2004, a mineral reserve audit in 2005, and NI 43-101 Technical Reports in 2008 and 2012
   2005 East Archimedes developed as conventional open pit mining and heap leach operation with initial gold production in 2007
   In 2013 the East Archimedes high wall failed, and mining was suspended pending economic assessment of moving failed material to continue mining

 

 

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Year    Company    Comment
        Barrick completed a pre-feasibility study on the 426 zone in 2009 and a feasibility study in 2012. The 2012 feasibility concluded that the 426 zone needed +$975/oz gold to be economical.
        2003-2015 drilled 674 (811,575 feet) exploration/infill/definition drill holes; 523 RC (630,745 feet) and 151 core (180,830) holes
        2002 Quantec Consulting Inc. conducted a 5-line Titan-24 magnetotelluric survey, added additional 4 lines in 2010
        2006 merged gravity data from multiple sources and various scales
        2007 Magee Geophysics Services LLC conducted a 3,182 station gravity survey on 300’ grid spacing
        Conducted rock chip sampling program in 2002
2015    Waterton Precious Metals Funds II Cayman, LP    Purchased Ruby Hill mine from Barrick. Waterton formed new corporate entity called Ruby Hill Mining Company, LLC
2015-2021    Ruby Hill Mining Company, LLC    Completed 42 sonic drill holes totaling 4,106’ between 2019 - 2020
  

2017 reprocessing of selected historical geophysical datasets, multi-element analysis study of drill core to aid in lithology identification, and structural review by SRK.

Conveyed the Historic Ruby Hill claims and Fad Mine to Golden Hill Mining Corp.

   McCoy Mining was hired to begin mining from the bottom of the East Archimedes Pit in August 2020. The operation mined about 2,599,000 tons of ore containing 40,900 ozs Au. Mining was completed in November of 2021
October 2021- Present    i-80 Gold Corp    Acquired Project October 18, 2021
   Completed East Archimedes mining November 2021
   Residual leaching and gold recovery from the East Archimedes heap leach pad
   IP Survey 2022
   Ongoing drilling (72 holes totaling 135,941 ft (41,435 m) at time of writing. (Not all holes are within the current resource area.)
   February 2023 purchased FAD property from Paycore Minerals. Paycore had initiated drilling programs testing CRD mineralization at depth and a near-surface oxide target proximal to historic Archimedes Underground mine with favorable results.
April 2022- February 2023    Golden Hill Mining Corp.    Acquired FAD property (south of the Project), drilled 33,675 feet (10264 m), sold to i-80

 

5.2

Historic Mining

Historical district production from 1866 through 1964 is estimated at 1.65 Moz of gold at an average grade of 0.83 oz/ton (28.5 g/t Au) and 39.0 Moz of silver at an average grade of 19.5 oz/ton (668.6 g/t Ag) from 2.0 Mtons mined (also reported >625M lbs Pb @ 15.63 %) of which 80% is estimated to be from the original Ruby Hill Mine (Nolan, 1962 and Nolan and Hunt, 1968). The bulk of historical mining was completed by 1891 when the Eureka smelter closed. Sporadic shipments of lower grade ores by lessors continued until about 1940 along with minor production from Adams Hill and Mineral Point, which are in the vicinity of the current resource area. Production from mines on Adams Hill and Mineral Point contributed no more than 125,000 tons of low-grade material, with most of the production, 67,000 tons, coming from the Holly mine (Nolan, 1962).

The Holly mine was accessed via the TL shaft, sunk by Eureka Corporation, LTD in 1953 to a depth of 1,127 feet. The historic workings lie above, and locally intersect, the current Mineral Point resource. Although records are sparse, the TL shaft appears to have been used to access two main working areas: the Holly mine and the Williamsburg/Bullwhacker mine. The Holly mine is located in the footwall of the Holly fault near its juncture with the 150, 426 and Hilltop faults, adjacent to the southwest highwall of the Archimedes pit. Developed levels range from roughly 70 feet to 900 feet below surface. The Williamsburg mine was developed on levels ranging from near-surface to roughly 1,070 feet below surface. Workings tend to follow the contact of the Bullwhacker sill and the Catlin Member of the Windfall Fm (stratigraphically above the host of the current Mineral Point Trend resource). Although the historical mining concept is not well documented, the target of both the Williamsburg and Holly mines was likely CRD mineralization.

 

 

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The FAD shaft was sunk in 1941 to access CRD mineralization intercepted by surface exploration drilling adjacent to the historically mined Ruby Hill deposit. The FAD mineralization is thought to be a continuation of the historic Ruby Hill deposit, down-dropped to the north-east by normal faulting. The shaft reached a depth of 2,500 feet and a drift was driven on the 900 “Locan” level for underground exploration drilling and test mining. 78 holes totaling 12,976 feet are known to have been drilled from the Locan level. The shaft eventually flooded, and little work was completed from 1963 until Paycore Minerals acquired the property in April 2022. As of February 2023, Paycore reported completing 33,675 feet (10,264 m) of drilling at 656-foot (200-m) step-outs, expanding the CRD deposit footprint to almost one square mile (1.5km x 1.5km) open in multiple directions. Paycore also reported a near-surface oxide exploration target above the FAD CRD, adjacent to historic infrastructure.

Modern work at the Project began in 1992, when Homestake Mining Company made the Archimedes Carlin-style discovery at the current Project area. About 1,508,900 oz Au have been produced from the Archimedes pit from roughly 24.3 Mtons of ore. Table 5-2 lists historical production.

Table 5-2: Production History Summary

 

     
Year   Company    Comment
1866-1964   Numerous    Eureka District produced 1.65 Moz Au, 39 Moz Ag, 625 Mlb Pb and 12 Mlb Zn from 2 Mtons of ore (Historical estimate)
   1873-1905 Richmond Mining Company mined 488,081 tons of material valued at $15,209,012.
   1873-1916 Eureka Consolidated Mining Company mined 550,455 tons material valued at $19,242,012,
   1871-1939 Richmond-Eureka Mining Company mined 88,081 tons material valued at $4,021,674.
   Small scale sporadic production from numerous lessors.
1953-1958   Eureka Corporation, LTD.    Sunk TL shaft in 1953, production from historic Williamsburg/Bullwhacker and Holly Mines (underground, these workings lie above, and locally intersect, the current Mineral Point resource.) Production estimates are included in cumulative historic Eureka District totals (Nolan). Subordinate to historic Ruby Hill production (125 Ktons max)
1998-2000   Homestake Mining Company    Produced 365,491 oz Au from 3.7 Mtons of mineralization from West Archimedes Pit
2001-2015   Barrick Gold Corporation    Produced 1,081,458 oz Au from approximately 18 Mtons of ore from West and East Archimedes Pits
2016-2021   Ruby Hill Mining Company, LLC    Produced 21,105 oz Au from residual leaching of pad. Mining in bottom of East Archimedes Pit in August 2020 through August 2021.

 

5.3

Historic Exploration

Exploration for the Ruby Hill Project has a long history which consisted of rock-chip sampling, soil sampling, mapping, drilling, and geophysical surveys. Modern projects conducted by previous owners Homestake Mining Company, Barrick Gold Corporation and RHMC are presented here, and a 2022 IP survey conducted by i-80 is presented in Section 9. All known drilling at the Project is presented in Section 10. A list of all known historical exploration efforts in the district is presented in Table 5-3.

 

 

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Table 5-3: Historic Exploration

 

Year

  Company   Comment

1864

  N/A   Oxidized gold-silver CRD mineralization discovered by prospectors

1923

  Richmond-Eureka Mining 
Company
  Drove SE crosscut to Ruby Hill fault, and a drift to SW.
  Vertical exploration hole (type unknown) drilled from 900 level. Hole caved, and project abandoned

1937-1959

  Eureka Corporation, Ltd.   Completed 4 churn holes (totaling 3,596 feet), 260 surface and underground core holes (87,633.8 feet), 13 mud rotary holes (14,252 feet), and 6 RC holes (9,903 feet)
  Intersection of high-grade polymetallic mineralization in 5 surface core holes led to the FAD shaft being sunk to 2,500’ depth to develop mineralization.
  Rotary drilling in 1953 in Adams Hill area intersected mineralization in Hamburg Dolomite

1989-1991

  American Smelting and
Refining Company
(ASARCO)
  Drilled 12 RC exploration holes totaling 5,314 feet

1960-1992

  Ruby Hill Mining Company   Consortium (Richmond-Eureka, Eureka Corp, Newmont, Cyprus, Hecla) drilled 148 exploration holes (129,362.3 feet); 13 churn (3,641 feet); 33 Mud Rotary (74,039 feet); 6 percussion (395 feet); 3 RC (1,458 feet); and 93 core holes (50,218.3 feet)
  Fourteen holes drilled in FAD shaft area intersected mineralization. Decision made to dewater FAD shaft to exploit new mineralization
  In 1963 FAD shaft was dewatered to the 2250 level. New crosscut, 1,028’ long, to evaluate mineralized zone completed in 1964. Crosscut used to drill exploration percussion and core holes
  Cyprus remains as surviving partner drilling 39 mud rotary (7,945 feet), and 98 air track (4,983 feet) exploration holes for near-surface, bulk-mineable gold mineralization between 1980-1981
  Sharon Steel Corporation drilled 127 exploration/definition RC holes totaling 31,539 ft between 1982 and 1991

1993-1994

  Placer Dome   Drilled 11 RC exploration holes (12,350 feet) at Ruby Flats

1994

  Unknown   Drilled 1 RC hole for 500 feet

1992-2001

  Homestake Mining
Company
  Exploration/definition drilling between 1992-1993 discovered/defined the Archimedes deposit (both West and East) along with the 426 zone
  Completed 1,502 (1,022,842.5 feet) exploration/definition holes between 1992-2001; 1374 RC holes (875,083 feet), and 128 core holes (147,759.5 feet)
  DIGHEM Surveys conducted an airborne magnetic & electromagnetic survey in 1994 on E-W flight lines at nominal 600’ spacing with mean terrain clearance of 115 feet
  Zonge Geosciences completed ground magnetics survey at 150’ spacing in 2000.
  Between 1999-2000 conducted rock chip sampling program to determine potential for multi element correlation as pathfinder for gold

2001-2015

  Barrick Gold Corporation   In 2002 Chadwick and Russell completed Archimedes pit mapping
  2003-2015 drilled 674 (811,575 feet) exploration/infill/definition drill holes; 523 RC (630,745 feet) and 151 core (180,830 feet) holes
  2002 Quantec Consulting Inc. conducted a 5-line Titan-24 magnetotelluric survey, added additional 4 lines in 2010
  2006 merged gravity data from multiple sources and various scales
  2007 Magee Geophysics Services LLC conducted a 3,182 station gravity survey on 300’ grid spacing
  Conducted rock chip sampling program in 2002

2015-2021

  Ruby Hill Mining Company,
LLC
  Completed 42 sonic drill holes totaling 4,106’ between 2019—2020
  2017 reprocessing of selected historical geophysical datasets, multi-element analysis study of drill core to aid in lithology identification, and structural review by SRK

October

2021-Present

  i-80 Gold Corp   IP Survey 2022
  Ongoing drilling (72 holes totaling 135,941 feet (41,435 m) at time of writing. (Not all holes are within the current resource area.)

April 2022-

February 2023 

  Paycore Minerals   Acquired FAD property (south of the Project), drilled 33,675 feet (10264 m), sold to i-80

 

 

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Figure 5-1 shows the location of geophysical surveys completed from 1994 to 2022. Geophysical surveys have been instrumental in locating CRD mineralization.

Figure 5-2 and Figure 5-3 show locations and gold grades of rock samples and soil samples collected by previous operators within the Ruby Hill claim block.

 

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Figure 5-1: Geophysical Surveys in the Ruby Hill Project Area

(Source: i-80 Gold, 2024)

 

 

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Figure 5-2: Rock Samples with Gold Grade (opt) within the Ruby Hill Claim Block

(Source: Wood, 2021)

 

 

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Figure 5-3: Soil Samples with Gold Grade (opt) within the Ruby Hill Claim Block

(Source: Wood, 2021)

 

 

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6.

GEOLOGIC SETTING, MINERALIZATION AND DEPOSIT TYPES

 

6.1

Regional Geology

The Ruby Hill Project is located in the Eureka mining district in east-central Nevada, within the northern part of the Fish Creek Range which is a nearly continuous sequence of Cambrian and Ordovician sedimentary rocks (Figure 6-1) totaling nearly 10,000 ft in thickness (Nolan, 1962). These strata accumulated on a stable continental shelf margin and consisted primarily of carbonate units with subordinate shale and sandstone (Dilles et al., 1996). The Cambrian Eldorado Dolomite, the Hamburg Dolomite and overlying Dunderberg Shale, portions of the Windfall Formation, and the Goodwin-Ninemile transition, host most of the mineralization within the district (Barrick, 2011).

During the Mississippian Antler Orogeny, the Roberts Mountains Allochthon, consisting primarily of deep marine sedimentary rocks, was thrust from the west onto the continental margin (Evans and Theodore, 1978), creating a foreland basin in the vicinity of the present-day location of the town of Eureka, NV (Poole, 1974). Post-Antler Mississippian and Permian strata deposited after the Antler Orogeny filled the basin with carbonaceous silts, sands, and conglomerates represented by the Chainman and Diamond Peak formations (Dilles et al, 1996).

Thrust faulting and significant deformation of the Paleozoic section occurred between Permian and Late Cretaceous time (Taylor et al., 1993), and culminated in the development of the Prospect Mountain duplex of the Early Cretaceous Hoosac thrust fault (Lisenbee, 2001), a major regional scale structure that cuts Permian rocks, and is in turn cut by intrusive units dated 110 to 100 Ma (Dilles et al., 1996). Most of the Eureka district is located in the hanging wall of the Hoosac thrust.

Cretaceous fresh-water sedimentary rocks unconformably overlie the older Paleozoic units east of Eureka, NV (Nolan, 1962). Cretaceous age granodiorite and quartz porphyry intrude the Paleozoic section. These include the Ruby Hill stock, Bullwhacker Sill, and Graveyard Flats intrusive which are interpreted to be genetically linked to the base metal carbonate replacement deposits at Ruby Hill (Barrick, 2011). Oligocene volcanic tuffs and andesite intrusive rocks are also present within the district, primarily to the NE and SE. The youngest deformational event occurred during the Miocene when Basin and Range extension formed regional high- angle N-S trending normal faults.

 

 

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Figure 6-1: Regional Geologic Map

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The Eureka district hosts mid-Cretaceous, igneous-related, polymetallic carbonate replacement deposits that have subsequently been overprinted by Carlin-type gold-silver mineralization. Gold and silver mineralization possibly dates to the early-middle Cenozoic (Eocene) and temporally coincides with the onset of extension and Eocene-Oligocene magmatism. Post mineral uplift exposed portions of the Archimedes gold deposit and likely contributed to the relatively deep level of oxidation. Subsequent Miocene Basin and Range faulting resulted in reburial of the Archimedes system beneath 60 to 500 ft of Tertiary-Quaternary overburden in East Archimedes.

 

6.2

Project Geology

The Ruby Hill Project is located on the southeastern end of the Battle Mountain-Eureka gold trend, in the northern portion of the of the Eureka mining district.

From the late Neoproterozoic to the Devonian, the Cordilleran passive margin sequence, a westward-thickening section of clastic and carbonate rocks, was deposited on the rifted North American continental shelf in what is now eastern Nevada and western Utah. The Eureka district was situated near the distal, western margin of the shelf (Cook and Corboy, 2004). The Project is underlain by a thick (approximately 10,000 feet) sequence of carbonate and siliciclastic units comprised of the Prospect Mountain Quartzite, Pioche Shale, Eldorado Dolomite, Geddes Limestone, Secret Canyon Shale, Hamburg Dolomite, Dunderberg Shale, Windfall Formation, Goodwin Formation, Ninemile Formation, Antelope Valley Formation, and the Eureka Quartzite. During the Mississippian, the Roberts Mountains allochthon, composed of distal slope and basinal sediments, was thrust to the east over the western edge of the continental shelf during the Antler orogeny (Dickinson, 1977). Eureka is immediately to the east of the Antler thrust front, but was the site of synorogenic deposition of Mississippian clastic sediments that were sourced from the Roberts Mountains allochthon (Smith and Ketner, 1977). During the Pennsylvanian and early Permian, eastern Nevada underwent a protracted series of deformation and erosion events recorded by unconformity-bound packages of carbonate and clastic rocks (Trexler et al., 2004). In the Eureka district the Pennsylvanian–Permian section consists of 1.3 km of limestone and conglomerate, including the Ely Limestone and Carbon Ridge Formation. However, these rocks are located east of the Project boundary.

The Eureka area has experienced a multiphase tectonic history of contractional deformation complexly overprinted by extensional deformation. The earliest observable deformation event occurred during the Cretaceous Sevier Orogeny as part of the Central Nevada Thrust Belt. The Sevier Orogeny is defined by subduction of the Farallon Plate beneath the North American Plate resulting in contractional deformation. This deformation resulted in the development of the Eureka culmination, a north-striking anticline with a 20 km wavelength, a 4.5 km amplitude, and limb dips of 25°–35°, which is corroborated by deep Paleogene erosion levels that can be traced for ~100 km along strike (Long et al., 2014). Locally, the Mineral Point anticline and several thrust faults are attributed to this deformational period (Hastings, 2008). During the Cretaceous and post-contractional deformation, the region was subjected to widespread magmatic activity, resulting in emplacement of the Ruby Hill stock and the Graveyard Flats intrusive. Late Cretaceous through Eocene saw high-angle extensional deformation accompanied by felsic magmatism. Basin and Range extension began in the Miocene and continues through present, forming elongate N-trending basins and valleys and regional high-angled generally N-trending faults (Dickinson, 2006). Within the district, dominate structural trends are low and high-angled N-, NE-, and E-trending faults. Major structural features within the Property which control mineralization include the NNW- trending Mineral Point anticline, the west bounding Spring Valley fault, the N-trending Jackson-Lawton-Bowman-Holly fault system, and the WNW-trending Blanchard, Hilltop and Ruby Hill fault zones (Hoge et al., 2015).

 

 

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Mineralization within the Project area is characterized as:

 

   

Au Carlin-type: West Archimedes, East Archimedes, Ruby Deeps, 426 zones and exploration targets including Blue Sky, 007, 008, and 1428.

 

   

Au+Ag distal-disseminated: Mineral Point deposit.

 

   

Zn-Pb-Ag-Au carbonate replacement deposit type (Polymetallic CRD): deposits mined historically throughout the district including FAD and Ruby Hill, exploration potential within the newly identified “Hilltop Corridor” and multiple targets supported by drill intercepts along the newly interpreted Hilltop fault.

 

   

Skarn base metal: Blackjack and Hilltop Fault-Graveyard Flats stock intersection.

Mineralization is lithologically and structurally controlled and is focused primarily within the carbonate-rich Ordovician and mid to upper Cambrian formations. Minor skarn and CRD mineralization occur within the Cretaceous intrusive units.

The northern “Ruby Hill” portion of the Project contains two distinct mineral resources, the Mineral Point Trend and the Archimedes complex (consisting of West Archimedes, East Archimedes, 426, Ruby Deeps, Hilltop, 007 and Blackjack). The Mineral Point Trend and Archimedes are separated by the Holly fault.

The southern “Golden Hill” portion of the Project contains the historic Ruby Hill mine and the unmined FAD deposit, which is interpreted to be a deeper extension of the Ruby Hill deposit down dropped by the northwest-striking, down-to-the-northeast Ruby Hill normal fault (Figure 6-6).

Alteration within the project area consists of skarn, calc-silicate, marble and hornfels, silicic, argillic, decarbonatization, and propylitic styles. Silicic alteration most commonly occurs as jasperoid and is most developed in the northern portion of the Property and associated with Carlin-type and distal disseminated mineralization. Decarbonatization is ubiquitous throughout Carlin-type and distal-disseminated mineralization zones including 426, Ruby Deeps, Mineral Point, and Blackjack where Carlin-type alteration overprints skarn (Hastings, 2008). Skarn alteration is limited to areas adjacent to the Graveyard Flats stock. Calc-silicate and propylitic alteration is also found adjacent to the Graveyard Flats stock, in dikes and sills, and in deeper drilling beneath the Archimedes pit and Hilltop areas. Marble and hornfels are seen adjacent to CRD ore at Hilltop, distal to the Graveyard Flats stock, as well as in deeper drilling beneath the Archimedes pit and Hilltop areas. Decarbonatization and argillic assemblages are the most common form of alteration at Ruby Hill and FAD.

 

 

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Figure 6-2: Ruby Hill Project Geology and Deposit Locations

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6.3

Stratigraphy

A stratigraphic column depicting the stratigraphy at the Project is shown in Figure 6-3.

 

6.3.1

Lower Cambrian

Prospect Mountain Quartzite (Ꞓpm)

Light tan to white well-sorted quartzite. White, pink, tan, and brown when weathered. Commonly cross-laminated with rare pebble conglomerate interbeds. Micaceous to sandy shale interbeds common near base of unit. The unit is not observed within the Property area but within the region it is mapped up to 1,500 ft thick.

Pioche Shale (Ꞓp)

Khaki to green, less commonly red-orange, sandy micaceous, locally calcareous shale. Contains thin interbeds of red-brown micaceous sandstone and quartzite, and mottled, well-bedded, dark-blue limestone with abundant trilobite fragments (Long et al, 2014). The unit unconformably overlies the Prospect Mountain Quartzite. It is not observed within the Property area but within the region it is mapped up to 500 ft thick.

 

6.3.2

Middle Cambrian

Eldorado Dolomite (Ꞓed)

Medium-dark gray, massive weathering dolomite. Forms distinct gray cliffs. Commonly mottled and streaked with white stringers and spots. Dark dolomite locally alternates with lighter gray, rough textured dolomite giving the appearance of alternating light and dark bands up to 1 ft thick, which defines bedding (Long et al., 2014). Fenestral (birds’ eye) structure is common. Alters to a light-gray, coarse-crystalline (sanded), massive, featureless dolomite. Upper contact is interfingered with the Geddes limestone. Within the Project area the unit is up to 2,240 ft thick.

Geddes Limestone (Ꞓp)

Well bedded, thin to medium bedded, dark blue to black carbonaceous limestone, with maroon-weathering silty and shaly partings, and black nodular chert (Long et al., 2014). Forms angular blocky float. Lower contact is interfingered with the Eldorado Dolomite. Commonly folded at the outcrop scale. Black color and well-developed bedding diagnostic of the unit. Within the Project area the unit is up to 550 ft thick.

Secret Canyon Shale

Divided into two distinct interbedded members, the Lower Shale Member and the Clark Springs Member. Within the Project area the unit is up to 1,250 ft thick.

Lower Shale Member (Ꞓss)

Brown, olive to tan, calcareous, argillaceous shale with local interbedded limestone (Nolan, 1974). Weathers to a brown, red, and/or yellow (Nolan et al., 1956). Overlies the Geddes Limestone with a sharp conformable contact.

Clark Springs Member (Ꞓsc)

Thin- to well-bedded, bioturbated, silty, micritic limestone with distinctive mottled yellow or red argillaceous partings (Nolan, 1974; Long, 2014). Gradational contact with Lower Shale Member.

 

 

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Hamburg Dolomite (Ꞓh)

Massive, light- to medium-gray, coarse crystalline dolomite with mottled white stringers that define bedding, and oblong “blue bird” stringers (Long, 2014). Typically porous or vuggy, commonly altered to jasperoid. Lower contact gradational with Clark Springs Member. Within the Project area the unit is up to 1,320 ft thick.

 

6.3.3

Upper Cambrian

Dunderberg Shale (Ꞓd)

Brown, khaki, and gray, fissile, paper thin, generally non-calcareous shale with diagnostic nodular limestone discs, and interbeds of medium-bedded, medium-gray limestone (Long, 2014). Outcrop-scale folding is common. Within the Property the unit is up to 265 ft thick.

Windfall Formation

Formation is divided into two members, the Caitlin (Ꞓwc) and Bullwhacker (Ꞓwb) members. Within the Property the unit is up to 700 ft thick.

Caitlin Member (Ꞓwc)

The Caitlin Member consists of alternating thick-bedded, massive weathering, medium- coarse crystalline, medium-gray limestone (Long, 2014). Interbedded with thin-bedded, sandy-silty limestone with tan to red, sandy-shaly partings. Trilobite fossil hash common in thicker bedded limestone. Sharp conformable contact between Caitlin member and Dunderberg Shale (Nolan et al., 1956).

Bullwhacker Member (Ꞓwb)

The Bullwhacker Member is thin-bedded, tan to light-brown, sandy or shaly, medium gray limestone, with tan-red sandy-shaly partings and interbeds (Long, 2014). It weathers to a diagnostic tan to red color, and trilobite hash and brachiopods are common. Additionally, the unit contains rare gray chert nodules.

 

6.3.4

Lower-Middle Ordovician

Pogonip Group

The Pogonip Group is divided into three formations, the Antelope Valley (Oav), Ninemile (Onn), and Goodwin (Og) described below.

Goodwin Formation (Og1, Ogll, Og2)

The Goodwin Formation is a light- to medium-gray, massive weathering limestone, and medium-gray, medium to thick bedded, silty, well bedded, fine crystalline limestone (Long, 2014). It is divided into three units, the Basal unit (Og1), Lower Laminate unit (Ogll), and Upper Goodwin (Og2). Within the Property the unit is up to 1,100 ft thick.

Basal unit (Og1)

The basal unit consists of massive bedded, fine- to medium-grained, medium to dark gray, chert-bearing calcisiltite and calcarenite (Dilles et al., 1996). The Og1 unit is approximately 350 ft thick.

Lower Laminated unit (Ogll)

Consists of tan to gray, laminated to thin bedded micrite, calcisiltite, and shaly limestone (Dilles et al., 1996). The unit varies in thickness from 150 to 250 ft.

 

 

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Upper Goodwin (Og2)

Composed of thin to medium bedded, chert bearing calcisiltite and calcarenite (Dilles, et al., 1996). Light gray, brown, and black chert nodules common (Long, 2014). It is approximately 500 ft thick.

Ninemile Formation (Onn)

Platy, thin bedded, porcelaneous, carbonaceous, fossiliferous, olive-green limey shale, and shaly medium-grained limestone (Dilles et al., 1996; Long, 2014). Weathers to a distinctive olive and brown color. Within the Property the unit is up to 520 ft thick.

Antelope Valley Formation (Oav)

Thin to medium and locally thick bedded, medium-blue gray, fine crystalline limestone (Long, 2014). Ubiquitous tan to yellow silty partings, and local tan, brown, and white chert nodules. Lower contact interfingers with Ninemile Formation. Within the Property the unit is up to 500 ft thick.

Eureka Quartzite (Oe)

 

Vitreous white to dark gray, fine to medium grained, well sorted quartzite. Exhibits “sugary” quartz texture. Weathers gray to red and is commonly brecciated. The lower portion of the unit is commonly cross-laminated and it unconformably overlies Antelope Valley Formation (Nolan et al., 1956). Within the Property the unit is up to 535 ft thick.

 

6.3.5

Cretaceous

Graveyard Flat Intrusion (Kgf)

The Graveyard Flats intrusion, discovered beneath alluvial cover during drilling at Archimedes, is of Cretaceous age. Primary mineralogy is quartz monzonite (Hastings, 2008). The intrusive consists primarily of quartz, and variably altered plagioclase phenocrysts in a fine-grained, equigranular, plagioclase-dominated groundmass (Dilles et al., 1996). Common alteration products include sericite, kaolinite, calcite, chlorite, epidote, and pyrite (Dilles et al., 1996). Primary ferromagnesian minerals are not preserved. Dilles et al. (1996), based on observed textural variations within the intrusive, suggest that the intrusion may have been emplaced in multiple phases. Mortenson et al. (2000) reports a U-Pb zircon age of 106.2 ± 0.2 Ma for the intrusion.

Bullwhacker Sill (Kbs)

The Bullwhacker sill is located west of the Graveyard Flat intrusion, and dips gently east underneath the Archimedes pit where it may merge with the Graveyard Flat intrusion (Hoge, 2015). It is generally emplaced along the contact between the Windfall Formation and the Dunderberg Shale (Dilles et al., 1996) as far west as the hinge of the Windfall anticline, after which it tends to trend upwards through the Bullwhacker member. The sill is offset by several normal faults. West of the Holly Fault the sill dips more steeply east, conformable with bedding steepened by the Bowman-Williamsburg fault. The disseminated mineralization within the Mineral Point trend lies below the sill, and the western limit of the mineralization generally coincides with the westmost extent of the sill, while the eastern limit is proximal to the sill near the Bowman-Williamsburg fault. In the area of the Ruby Deeps deposit, the sill intruded along multiple planes within the Windfall Formation up to the Catlin-Bullwhacker contact, forming multiple lenses. The Ruby Deeps occurs proximal to, and locally within, the lenses. One conspicuous lens of intrusive was emplaced in the Bullwhacker member between the 426 and NS faults. The 008 deposit is proximal to this lens, and the 007 and 426 deposits are somewhat proximal.

 

 

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6.3.6

Tertiary/Quaternary

Volcanic Units (Trf/Tv)

Tertiary rhyolitic flows, tuff and volcaniclastic rocks are present in the northern part of the district and exposed in eastern and southeastern Archimedes pit wall. Within the Property the unit is at least 200 ft thick.

Sparse intersections of west northwest-trending lamprophyre dikes also have been observed from pit mapping and noted in some East Archimedes drill holes.

Alluvium (Qal)

Within the Property the alluvium unit is up to 535 ft thick and consists of “stream alluvial, piedmont gravels, and slope wash” (Nolan, 1962).

 

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Figure 6-3: Ruby Hill Stratigraphic Column

(Source: i-80 Gold, 2023)

 

 

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6.4

Structure

The Property has undergone a complex tectonic history of deformational, extensional, and intrusive events, producing a series of folds, and high- to low-angled faults. Structures have been defined by a combination of surface mapping including Nolan (1962), Cooper (2002), Hauntz (1999), and Chadwick and Russell (2002), Uken (2017a 2017b), drill hole logging, geologic modeling, and interpretation of geophysical data.

Mesozoic deformational events produced a series of generally N-, NW-, and NE-trending faults and NW- to NE-trending folds within the Property area (Long et al., 2014). Tertiary Basin and Range extension and subsequent high-angled faulting have transected and possibly displaced some portions of the deposits within the Property (Nolan, 1962).

The main structural features within the Property area include early low-angled thrust faults (45° -95°), and apparent low- to high-angle normal faults (20° -45°) in three dominant orientations, which include 345° -015°, 030° -050°, and 080° -110° (Table 6-1). Major faults within the Property include the Holly fault, Bowman-Williamsburg fault, Hilltop fault, Ruby Hill fault, Champion thrust, and the Blanchard fault zone (Table 6-1). A number of the high-angle normal faults are interpreted to have crosscut and reactivated low-angle thrust faults.

Large-scale folds within the Property include the NNW-trending (330° -340°), gently N- plunging (5° -10°) Mineral Point anticline, located in the central and north-west portions of the Property. The Mineral Point anticline is one control to mineralization within the Mineral Point deposit. Small-scale folds throughout the property control mineralization locally.

Table 6-1: Major Structural Features and Orientations within the Property Area

 

Structure Orientation Major Features Kinematics Dip Notes

N-Trending Faults

NNW to NNE

The Bowman Fault, Holly Fault and associated splay faults including the Holly Splay Fault, Armpit Fault, 599 Fault, and 150 Splay Fault Oblique normal slip High-angle The fault surface is typically undulating with up to several feet of gouge fill.

NE-Trending Faults

NE

The 426 Fault, 194 Fault, Jackson fault, and Graveyard Fault Strike-slip and oblique normal slip  Variable dips from steeply dipping to more shallow dipping Faults are gouge filled with up to 4 in of gouge material.

E-Trending Faults

EW-WNW

Blanchard fault zone, Hilltop fault, Ruby Hill fault, and associated unnamed EW and WNW faults Strike-slip High-angle

The Blanchard fault zone may be up to 100 ft wide in portions of the Archimedes pit.

The Hilltop fault is tens of feet thick with gouge and oxidation along it.

Thrust Faults

NS

The Champion thrust, Prospect Mountain thrust, and Ratto Canyon thrust, (off Property to the S), and other possibly reactivated normal faults within the district Reverse Low-angle Commonly associated with folds, including the Mineral Point anticline. Folding typically occurs in well laminated units and varies in amplitude from approximately 20 inches to 3 feet.

Folds

NNW-NNE

The Mineral Point anticline Anticline

Significant mineralization control.

 

 

 

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6.4.1

Archimedes Deposit Structure

East Archimedes, West Archimedes, 426, Ruby Deeps, 007, 008, Hilltop and Blackjack are located on the eastern side of the near north-trending Holly Fault. Chadwick and Russell (2002), Hastings (2008), Morkeh (2011), and Uken (2017a, 2017b), mapped the structure and geology of the Archimedes pit.

On the western margin of West Archimedes is the N-trending high-angle normal Holly fault (east dipping, 79°), and the 150 fault (east dipping in the northern portion and west dipping in the southern; 85°; Chadwick and Russell, 2002; Hastings, 2008; Morkeh 2011). The 150 fault offsets the Bullwhacker Sill to the east by 500 ft (Hastings, 2008). The 194 fault, 426 fault, and Armpit faults are variable N- to NE-trending (345°-020°), east to west dipping, low- to high-angle faults (46° -87°), which transect the center portion of West Archimedes (Chadwick and Russell, 2002; Hastings, 2008; Morkeh 2011). The Blanchard fault zone is a NW-trending (295°), steeply dipping (NE; 75° -85°), fault zone which is reported to be 100 ft wide in some locations (Chadwick and Russell, 2002; Hastings, 2008).

Within the East Archimedes zone the Graveyard fault zone is a N-trending (350° -010°), west dipping (60° -80°), series of faults, which transects the east margin of the pit (Chadwick and Russell, 2002; Hastings, 2008; Morkeh 2011). The Blanchard fault zone continues east from the West Archimedes pit into East Archimedes for an unknown distance into the Graveyard Flat intrusion on the eastern margin of the pit (Chadwick and Russell, 2002; Hastings, 2008; Morkeh 2011).

The 426 zone is spatially associated with the NE-trending 426 fault zone and north of the Blanchard fault zone. The 007 zone is spatially associated with the NNE-trending NS Fault, also lying north of the Blanchard fault zone. The 008 zone lies between the 426 and NS faults, north of the Blanchard fault zone, along the hinge of an anticline formed above and intrusive lens.

Structure within the Ruby Deeps deposit area is a continuation at depth of faulting related to the Archimedes deposit to the east and the Mineral Point deposit to the west. The Ruby Deeps deposit is bounded to the east by the Graveyard Flats fault and the west by the Holly fault. The Blanchard fault zone transects the center portion of the deposit but does not appear to offset mineralization.

The Hilltop Fault has similar orientation to the Blanchard fault. It trends WNW just south of the Archimedes pit, from the Holly Fault towards the Graveyard Flat intrusion. It is undetermined whether the Hilltop fault transects the Ruby Deeps deposit or defines its southern boundary. Several drillholes have intersected CRD mineralization at various elevations along the Hilltop fault. Structures in the Archimedes deposit area are displayed on Figure 6-4.

 

 

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Figure 6-4: Geology of East Archimedes, West Archimedes and Archimedes Underground Including 426, and Ruby Deeps Zones

(Source: i-80 Gold, 2023)

 

6.4.2

Mineral Point Trend Structure

The Mineral Point deposit is in the central portion of the Property, west of the Holly fault. It is situated within the district-scale NNW-trending Mineral Point open anticline that plunges gently to the north. Major structures at Mineral Point represent a horst-like anticlinal dome bounded on the east by the Holly fault, and to the west by the Spring Valley fault (Figure 6-5). The primary lithological host of the Mineral Point mineralization is the Cambrian Hamburg Dolomite. Mineralization outcrops at the southeastern extent and plunges to a depth of about 550 ft at its northern extent, dipping roughly 5° along its 10,000 ft length. Several steeply dipping normal faults of varying apparent displacement are associated with the Mineral Point anticline. From west to east these include the west-dipping West Fault which bounds the west limb of the anticline and defines the western limit of mineralization; the Bowman-Williamsburg Fault which parallels the axial plane of the anticline; and the Holly fault which is an offshoot or northward extension of the district scale Jackson-Lawton Fault system to the south (Loranger, 2013). The Bowman- Williamsburg and Holly Faults both dip steeply to the east. Structures in the Mineral Point Trend are displayed on Figure 6-5.

 

 

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Figure 6-5: Mineral Point Trend Geology

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6.4.3

Historic Ruby Hill and FAD Structure

The historic Ruby Hill and FAD deposits are separated by the northwest-striking, down-to-the-northeast Ruby Hill normal fault. Mineralization at FAD lies within the hanging-wall of the fault with Ruby Hill mineralization in the footwall. East of FAD the Jackson-Holly fault system drops stratigraphy down to the east. Additionally, the Ruby Hill fault cuts and offsets the Ruby Hill stock and mid-Cretaceous carbonate-hosted base metal mineralization. The Jackson branch cuts and offsets the Ruby Hill fault. Thus, the Jackson fault system is probably mid-Cenozoic in age, postdating the Ruby Hill fault. Evidence from zonation in the FAD and Ruby Hill mineralized zones suggests the Ruby Hill fault was pre-mineral, but may also have significant post-mineral offset. The Champion thrust fault, a west dipping fault, is an important control on mineralization at FAD where it forms a basal contact to mineralization. The thrust fault places Eldorado dolomite on Prospect Mountain quartzite with an approximately 100 ft thick gouge and rubble zone and pre-dates mineralization and all normal faulting. Structures in the Archimedes deposit area are displayed on Figure 6-6.

 

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Figure 6-6: Historic Ruby Hill and FAD Deposit Geology

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6.5

Alteration

Within the Project area, four main forms of alteration types have been observed; silicic, argillic, decarbonatization, and reaction skarn/skarnoid. Other types of alteration identified within the Property include skarn, propylitic, and quartz-sericite-pyrite (QSP).

Silicic is characterized by complete (jasperoid) or partial replacement by silica and development of quartz and silica infill of breccias. Silicic altered rocks often appear red to brown in color and are intensely silicified. Jasperoid alteration is commonly observed within the Ninemile Formation, Goodwin Formation, and the Hamburg Dolomite, and within the Mineral Point and Archimedes deposits. Outcrops of jasperoid alteration are common within the Property and are easily distinguished by coloration and resistance to weathering. Dilles et al. (1996) notes that jasperoid consists of quartz with minor late chalcedonic silica filling vugs and veins filling fractures. Iron oxides consist of limonite and hematite pseudomorphs after pyrite and indicate a proto-ore pyrite content ranging from 5% to 20%. Gold occurs on the margins of oxidized pyrite and along hairline fractures in jasperoid (Dilles et al., 1996).

Argillic altered units are predominantly characterized as replacement of feldspar in igneous units by clay minerals (e.g., kaolinite and illite). Argillic alteration has been extensively logged within the carbonate units within the Property and most likely correlates with the removal of carbonate minerals during decarbonatization (Golder, 2012). Argillic altered material often appears white or bleached and may vary from chalky to greasy in texture.

Decarbonatized units are characterized by brecciated and sanded textures associated with dissolution of the carbonate-rich matrix of limestones and dolomites due to the interaction with an acidic fluid. Decarbonatization has been observed across the property.

Reaction skarn/skarnoid alteration forms a halo to garnet-pyroxene alteration and is composed of marble, hornfels, wollastonite, tremolite, and other calc-silicate minerals. This alteration is also present at depth beneath the Hilltop Zone and elsewhere on the property proximal to the Holly fault zone at depth.

Metasomatic garnet-pyroxene skarn and retrograde alteration assemblages are present within Blackjack and East Archimedes at depth, proximal to the Graveyard Flats intrusion. Additionally, propylitic alteration (calcite, chlorite, epidote) and QSP alteration is observed within the Bullwhacker Sill and the Graveyard Flat intrusions.

 

6.5.1

Archimedes Deposit Alteration

Within the East and West Archimedes deposits the three main alteration types are observed along with skarn and propylitic assemblages proximal to the intrusive units.

Silicic alteration is spatially associated with the Blanchard Fault zone, and subsequent intersecting N- to NE-trending faults (Holly, 150, 194, Armpit, 426, and Graveyard Flats). Decarbonatization with breccia textures are observed in carbonaceous sedimentary units. Argillic alteration is logged extensively along the Blanchard fault zone and at the intersections of the Blanchard fault zone with the N- to NE-trending faults.

 

6.5.2

Mineral Point Trend Alteration

Common types of alteration along the Mineral Point Trend include silicic, decarbonatization (sanded and breccia texture development), and argillic assemblages (Golder Associates, 2012; Loranger, 2013). Silicic alteration occurs primarily within the Hamburg Dolomite, and is more prevalent within the SE portion of the deposit area. Silicic altered units are also observed as a series of stacked units that are interpreted to have

 

 

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preferentially developed along intraformational rock units in the folded Hamburg Dolomite. Sanded and brecciated textures are most common in the Hamburg Dolomite and varies from weak to strong.

Hydrothermal alteration of the Bullwhacker Sill consists of propylitic, QSP, and argillic alteration assemblages (Langlois, 1971). Propylitic alteration resulted in the development of a chlorite-calcite-kaolinite assemblage. Argillic alteration consists of kaolinite, sericite, and quartz (Langlois, 1971). Golder Associates (2012) report that the most intense argillic alteration occurs in the upper 5 to 10’ of the sill.

 

6.5.3

Historic Ruby Hill and FAD Alteration

The most common form of alteration at Ruby Hill and FAD are decarbonatization (sanded and breccia texture development) and argillic assemblages. In the Hamburg dolomite above the FAD mineralization, widespread decarbonatization in the form of sanding is present. Proximal to the mineralization the dolomite host rock has been metamorphosed to marble, with local decarbonatization distal to mineralization. The Prospect Mountain quartzite commonly shows argillic alteration in form of clay development below mineralization. Additionally, argillic alteration and decarbonatization at the historic Ruby Hill is more widespread than at FAD, likely due to supergene oxidation of sulfides forming acidic fluids.

 

6.6

Mineralization

Within the Property area, four styles of mineralization occur divided into three groups:

 

   

Polymetallic (Au-Ag-Pb-Zn) skarn or carbonate replacement deposit (CRD) of assumed Cretaceous age: Blackjack, Hilltop, FAD, and the historic Ruby Hill, Helen, Holly, and TL mines.

 

   

Au±Ag distal-disseminated mineralization of assumed Cretaceous age: Mineral Point.

 

   

Au Carlin-type mineralization of assumed Eocene age: East Archimedes, West Archimedes, 426, Ruby Deeps, 007, and 008 zones.

The zinc skarn and polymetallic CRD style is the oldest mineralization event recognized at the Property and is related to emplacement of the Cretaceous intrusive units. The precious metal-rich Carlin style overprints the older CRD event and is interpreted to have developed during early to middle Cenozoic (Eocene) times, similar to other Au-Ag deposits of the Battle Mountain/Eureka Trend. Mineralization is largely controlled by lithology and structure.

Distal disseminated Au-Ag mineralization is located west of the Holly fault in the N-tending, largely oxidized lower-grade Mineral Point Trend. This mineralization contains low-grade lead and zinc in addition to significant quantities of silver and lacks realgar and orpiment in contrast to Carlin-type mineralization.

Carlin-type gold mineralization overprinted the CRD/Skarn mineralization. It is largely confined to the area east of the Holly fault in structurally and lithologically controlled deposits (East and West Archimedes, 426, Ruby Deeps, 007 and 008; Figure 6-7.

Gold occurs as free grains within the oxide portions along with iron oxides, and associated with sulfide minerals (pyrite, arsenopyrite, arsenian pyrite, realgar, and orpiment) within the unoxidized portions of the deposits. Within the oxide horizons, petrographic work for samples from the Archimedes deposits “…indicate(s) that the gold was originally associated with pyrite grains, with no evidence of silica encapsulation. Higher grade gold mineralization occurs in zones of silicification and decarbonatized limestone,” (Resource Evaluations Inc., 2005).

Mineralization including Au, Au-Ag and Au-Ag-Pb-Zn is primarily hosted within the Windfall and Goodwin Formations, and within the Hamburg Dolomite. Combined mineralization spans an area approximately

 

 

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12,000 ft long, 9,000 ft wide, at the maxima, and spans from surface to approximately 2,400 ft below surface.

Mineralization is focused along high- and low-angle faults, lithologic contacts, and fold axes.

A plan and cross section showing the geometry and relationships of the Archimedes Deposit and Mineral Point Trend are shown in Figure 6-8 and Figure 6-9.

 

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Figure 6-7: Plan View of Ruby Deeps, 426, 007, 008, Blackjack, and Hilltop Zones

(Source: i-80 Gold, 2023)

 

 

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Figure 6-8: Plan View of Mineral Point Trend and Archimedes Deposits

(Source: i-80 Gold, 2023)

 

 

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Figure 6-9: Fence Section of Mineral Point Trend and Archimedes Deposits

(Source: Wood, 2021)

 

6.6.1

Archimedes Deposit Mineralization

At East and West Archimedes, gold-rich mineralization is associated with jasperoid and moderately to strongly decarbonatized limestone. Gold occurs in the oxidized ores as discrete grains less than 3 microns in diameter (Barrick, 2004; Barrick, 2012). Mineralization is controlled by structure and lithology. Within both deposits, the main mineralized bodies are focused along the NW-trending Blanchard fault zone. Second order control to mineralization within West Archimedes is focused by steeply dipping, N-trending normal faults (Holly, 150, 194, 426, and Armpit faults; Barrick 2004). Within East Archimedes, second order control to mineralization is by the N-trending Graveyard fault and East Archimedes fault.

East Archimedes mineralization is a NW-trending, roughly tubular shaped mineralized body, approximately 1,350 ft in height, 800 ft in thick, and 1,900 ft wide. The upper portion flattens and flares out to the west and connects to West Archimedes. Mineralization extends from surface to approximately 1,400 ft below surface and the main host rocks include Ogll and Og2 of the Goodwin Formation.

The West Archimedes zone is NW-trending, roughly cigar shaped, 1,700 ft long, 200 ft thick, and varies from 400 ft to 1,200 ft wide. Mineralization extends from surface to approximately 150 ft below surface and the main host rock is Og2 of the Goodwin Formation.

Mineralization at 426 is NE-trending, roughly rod-shaped, 1,300 ft long, 250 ft thick, and 250 ft wide. Mineralization is variably oxidized. Oxidation correlates strongly with proximity to fault structures and secondarily with elevation. The top of mineralization commences approximately 800 ft below surface with the main host rocks being the Og1 (oxide-rich) and Ogll (sulfide-rich) units of the Goodwin Formation.

Mineralization at Ruby Deeps is N-S trending, tabular zone comprised of stacked mineralized bodies developed within favorable lithological horizons. The overall zone is 2,200 ft long, 900 ft thick, and 800 ft wide. Mineralization is locally oxide at higher elevations and predominantly sulfide-bearing at lower elevations. The top of mineralization is approximately 1,200 ft below surface with the main host rock being the Windfall Formation.

Drilling is sparser eastward from Ruby Deeps and 426 towards the NS fault, and 007 and 008 are expressions of similar style mineralization continuing eastward from Ruby Deeps and 426 through favorable units. 007 and 008 lie generally on-trend with 426, but at lower elevation, lying north and east of the upper reaches of Ruby Deeps.

 

 

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The 007 Zone is controlled by the NE trending NS fault. Higher-grade oxide Au mineralization within the fault zone has been intersected by two holes, Barrick’s RC hole P7, 55’ @ 0.291 Au opt and i-80’s core hole iRH22-18A, 43.9’ @ 0.276 Au opt. Thickness and grade appear to be enhanced where the NS Fault intersects the Windfall-Goodwin contact. Three more i-80 holes west of the fault zone intersect thinner, stratigraphically controlled mineralization extending west along the Windfall-Goodwin contact. The zone is untested to the north and south, currently projecting about 400 ft along strike, 100 ft along dip, and ranges from 10 ft thick where stratigraphically controlled to over 40 ft thick within the NS fault zone.

The 008 Zone is stratigraphically controlled, lying near the top of the Windfall Formation in the hinge of an anticline bracketed by the 426 and Graveyard faults. The anticline appears to have formed above an intrusive lens emplaced within the upper member of the Windfall Formation, stratigraphically higher than typical Cretaceous sill material, which typically intruded along the lower contact of the Windfall Formation. The 008 Zone is not well defined but currently is projected about 350 ft long by 200 ft wide by 15 ft thick.

The Blackjack zone (not included in the current resource estimate) is a pod of zinc skarn mineralization hosted by the Lower Goodwin Unit proximal to the Graveyard Flats stock within the East Archimedes Zone below the Archimedes pit. It has elevated lead, copper and silver due to CRD overprinting. The base metal-rich CRD and skarn mineralization has been overprinted by later Carlin-style gold mineralization resulting in locally higher-grade gold zones. It is approximately 750 ft wide, 750 ft long and 900 ft high. The upper part of the Blackjack zone is partially oxidized with a high-to-moderate ratio of cyanide soluble to total fire assay gold, but sphalerite is un-oxidized. The lower portion of the zone is un-oxidized. Sulfide minerals include pyrite, sphalerite, galena, chalcopyrite and arsenopyrite. The top of mineralization is approximately 1,200 ft below surface, however the top of the deposit is partially exposed in the south east wall of the open pit at East Archimedes.

 

6.6.2

Mineral Point Trend Mineralization

Gold-silver mineralization at Mineral Point is dominantly oxide in nature with small, but higher-grade refractory material (Loranger, 2013). Mineralization is predominantly hosted within the Hamburg Dolomite and consists of decarbonatized dolomite and breccias composed of silicified and oxidized clasts of dolomite in a fine grained dolomite and silica matrix. Locally breccias are gossanous where a higher percentage of original pyrite existed. Higher grade breccia zones are cut by late, multistage quartz veins (Loranger, 2013). Mineralization also occurs along the upper contact into the overlying silicic altered Dunderberg Shale which hosts oxide and sulfide minerals.

The main mineralized zone at Mineral Point is roughly elliptical in shape, NNW-trending, and is approximately 10,000 ft in length, 2,400 ft wide, and approximately 500 ft thick. The mineralization extends from approximately 240 to 1,400 ft below surface.

 

6.6.3

Historic Ruby Hill and FAD Mineralization

Lead, zinc, gold, and silver values in oxidized replacement mineralization of the historic Ruby Hill occur in cerrusite, anglesite, and plumbojarosite, and in lesser amounts of mimetite, bindheimite, hemimorphite, and smithsonite (Nolan, 1962; Nolan and Hunt, 1968). These minerals are mixed with limonite, goethite, hematite, dolomite, calcite, aragonite, copper oxides, and small amounts of barite, wulfenite, and unreplaced wall-rock dolomite. All metallic oxide minerals formed from weathering of sulfide minerals, as remnant nodes of galena, pyrite, and sphalerite. The primary host rock is Eldorado dolomite with lesser mineralization in the Hamburg dolomite. Mineralization is likely controlled by fracture sets with structural intersections forming larger mineralized zones. Historic mineralized zones spanned 4000 ft in length NW to SE, 3-50 feet in width over a 500 ft wide zone, and were mined to depths over 1000 ft from surface.

 

 

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At FAD the primary sulfide minerals are pyrite, galena, and sphalerite with lesser chalcopyrite, arsenopyrite, and tennantite-tetrahedrite. These sulfide minerals replace both hydrothermal dolomite ± calcite and Eldorado Dolomite that encloses sulfide masses. On a microscopic scale, pyrite contains inclusions of sphalerite, chalcopyrite, and pyrrhotite. Sphalerite contains inclusions of chalcopyrite, pyrite, tennantite, and rare pyrrhotite, which is sometimes elongated and entrained along cleavages (Vikre, 1998). Silver is contained in solid solution with galena, but silver continued to be added as argentite veinlets after galena deposition had ceased. Gold is mostly contained through solid solution in pyrite. Mineralization may be controlled by a WNW-ESE near vertical fault similar to the Blanchard and Hilltop faults. The Champion thrust is an important lower boundary to mineralization, but mineralized zones may be strongly fracture controlled as is common in other deposits of this type. All mineralization at FAD is contained within the Eldorado dolomite. The FAD mineralized zones comprise multiple sub-horizontal lenses over a length of 1600 ft WNW to ESE. Individual lenses range from 3 to 100 ft thick over a width of approximately 500 ft.

 

6.7

Deposit Types

Mineralization at Ruby Hill is characterized by intrusion-related distal-disseminated, carbonate replacement, and skarn deposits that have been overprinted by younger Carlin-type gold mineralization.

 

6.7.1

Characteristics of Polymetallic Carbonate Replacement Deposits

The carbonate replacement mineralization is similar to other polymetallic (Pb-Zn-Ag ± Au) deposits found worldwide that are spatially associated with Cretaceous age intrusive units (Cox and Singer, 1987; Megaw et al.,1988; Plumlee et al., 1995; Titley 1993 cited in Hammarstrom, 2002; and Kamona, 2011). The carbonate replacement mineralization consists of massive to semi-massive pyrite, galena, sphalerite, and other sulfides typically with sharp boundaries into barren marble. Locally, mineralization is oxidized into gossanous bodies. Fluids are sourced from intrusions, with metals in bisulfide complexes at temperatures of 250° -500°C, with the depositional mechanism typically being a pH change that results in rapid deposition of metals (Beinlich et al., 2019).

 

6.7.2

Characteristics of Skarn Deposits

The skarn deposits at Ruby Hill are consistent with zinc skarns throughout the Cordillera (Meinert, 1987; Dawson, 1996). The Blackjack deposit is located along the margin of the Cretaceous Graveyard Flats stock. However, drilling by i-80 Gold suggests this is a faulted contact. At Blackjack, sphalerite is found disseminated and semi-massive to massive in garnet-pyroxene altered carbonates. In the eastern Hilltop area zinc skarn is contained within carbonates altered to marble and wollastonite and appears to be located more distal to the Graveyard Flats stock. Both zones typically show evidence of brecciation associated with mineralization. Zinc skarns typically form distal to their source intrusions at temperatures of 350-450°C with mineralization subsequent to metamorphism (Williams-Jones et al., 2010).

 

6.7.3

Characteristics of Carlin-Type Gold Deposits

Gold and silver mineralization within the Ruby Hill deposits is predominantly attributed to a Carlin-type overprint interpreted to temporally coincide with the onset of extensional tectonics and Eocene-Oligocene magmatism (Barrick, 2004).

The structural setting, alteration mineralogy, and mineralization characteristics of the Ruby Hill gold deposits are consistent with Carlin-type deposits as defined in Radtke (1985) and Hofstra and Cline (2000).

Carlin-type deposits formed in the mid-Tertiary after the onset of extension in an east- west trending, subduction-related magmatic belt. The deposits are located along long- lived, deep crustal structures

 

 

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inherited from Late Proterozoic rifting and the formation of a passive margin within Paleozoic carbonate sequences composed of silty limestone to calcareous siltstone. High and low-angle faults as well as intrusive rocks acted as conduits for moderately acidic fluids containing gold in bisulfide complexes, likely sourced from intrusions at depth (Muntean et al., 2011). Deposits typically show enrichment in antimony, arsenic, mercury, and thallium, caused by hydrothermal fluids with temperatures up to 250°C. Gold deposition occurs in arsenian pyrite, is predominantly hosted within carbonaceous sequences near major high-angle structural zones and is concentrated in structural traps and/or replacement horizons of reactive and permeable sedimentary beds.

Alteration of host carbonate sequences consists of decarbonatization, argillization, and silicification. Gangue minerals in Carlin-type deposits consist of calcite, siderite, clays, and ferroan dolomites that can occur as geochemical fronts beyond the mineralized zones.

 

6.7.4

Distal-disseminated Mineralization at Ruby Hill

Ore grades of gold and silver with elevated concentrations of zinc, lead, and copper and are found in the Mineral Point Trend. This mineralization is attributed to the earlier Cretaceous age of mineralization and is found predominately in the Hamburg dolomite. Ore fluids were likely similar to Carlin-type fluids and resulted in the formation of collapse breccias and an associated geochemical signature including arsenic, antimony, thallium, and mercury.

Distal-disseminated deposits share many similarities to Carlin-type deposits as the hydrothermal fluids are analogous. However, distal disseminated deposits typically occur within 5 km of an intrusion, have an association with base metals, and show a zonation pattern outward from the intrusive source. Examples include Lone Tree, Cove, and Star Pointer (Nevada), Mercur and Barneys Canyon (Utah), Jeronimo (Chile), Bau (Malaysia), Mesel (Indonesia), and Zarshuran (Iran) (Hill, 2016).

 

 

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7.

EXPLORATION

 

7.1

Geophysical

The Ruby Hill land package extends well beyond the extents of the current analysis (Mineral Point and Archimedes areas), and several exploration targets are being analyzed. Targets have been developed based on historical exploration and drilling projects as described in Section 6 and Section 10, as well as recent work by i-80.

In 2022, i-80 completed an IP/DC Resistivity survey and Transient EM surveys. Discovery Int’l Geophysics Inc. of Saskatoon, SK, S7K 7Z1, Canada completed the work from September 26 to November 6, 2022.

For the IP/DC Resistivity survey, six lines totaling 10.4 mi (16.8 km) were laid out in the vicinity of the Archimedes pit and southward toward the property boundary, covering an area of about 2.1 mi X 1.3 mi (3400m X 2100m). Each line had about 75 nodes with DIAS32 single-channel receivers connected with common voltage reference wire. Current was applied using a 70kW generator and Dias 25 kW transmitter and bi-directional pole-dipole and pole-pole data were simultaneously collected. Geophysicists performed QAQC on the data, analysed and interpreted it using proprietary algorithms. The west side of the grid showed high resistance, the east side showed lower resistance, and the middle showed high chargeability. The high chargeability anomaly might be associated with the presence of sulfides.

For the transient electromagnetic (TEM) surveys, two fixed loops were used (FLTEM) and four boreholes were scanned (BHTEM). One fixed loop was arranged around the Archimedes pit encompassing an area of about 3610 ft X 3280 ft (1100 m X 1000 m), and another roughly rectangular loop was run immediately south of the first, covering about 3610 ft X 2950 ft (1100 m X 900 m). Survey stations were set out in 16 lines totaling about 7 mi (11.4 km). Station spacing was closest along the southwest edge of the pit to maximize resolution around drill targets (the Hilltop fault zone). A 70kVa generator, Phoenix TXU30 transmitter, and DigiAtlantis timing controller were used to generate an upward magnetic field in each loop. Acquisition was with an EMIT Fluxgate magnetometer sensor and EMIT SMARTem24 receiver.

The four drillholes scanned for the BHTEM surveys were iRH22-40, iRH22-41, iRH22-43, and iRH22-51. An EMIT Digi-Atlantis Borehole System probe and controller were used. The probe was lowered down the hole using an electric/hand winch attached to a 1480m, 4-conductor cable. Transmission was with a 65-75KVA/50-58kW generator, Phoenix TXU30 transmitter and EMIT SMARTem24 Tx controller. All holes were surveyed with both north and south loops, and iRH22-40 was additionally surveyed with the perimeter of both loops energized. Data was measured at 16 ft to 66 ft (5 m to 20 m) intervals, with closer spacing in areas of sharp amplitude shifts to accurately characterize conductive response. Holes iRH22-40 and iRH22-41 were scanned between the 4515 ft and 4965 ft elevations, which coincides with the Lower Hilltop zone, while iRH22-43 and iRH22-51 were scanned between the 5675 and 6120 ft elevations, which coincides with the Upper Hilltop zone.

 

7.1.1

Archimedes Area

 

   

The area north of the Ruby Deeps zone named Blue Sky, where sparse historic drilling defined a large arsenic anomaly at the alluvium-bedrock contact.

 

   

The area beneath the Archimedes pit along the contact of the Graveyard Flats stock where skarn mineralization has been intercepted and drilling is still sparse.

 

   

Continuation of the Ruby Deeps to the south along the hanging-wall of the Holly fault.

 

   

The 428 target located beneath the Archimedes Pit near the Blanchard fault that contains two significant drill intercepts at the top of the Hamburg dolomite below hornfelsed Dunderberg Shale.

 

 

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The newly interpreted Hilltop fault zone at the southern boundary of the Archimedes area. i-80 has interpreted a new NW-striking fault structure, similar to major structures controlling the largest past producing mines in the district (Archimedes pit and original Archimedes Underground). A Titan MT survey completed by Barrick in 2010 and an IP survey completed by i-80 in 2022 identified geophysical anomalies including a resistivity high interpreted as a fault zone flanked by conductivity and chargeability highs coincident with massive sulfide mineralization. The conductivity highs closely correlate with CRD and skarn mineralization on the Hilltop fault. Hilltop zones include the high-grade Upper Hilltop near-surface oxide and semi-massive to massive sulfide mineralization, Lower Hilltop deeper high-grade polymetallic CRD mineralization, and East Hilltop high-grade CRD and skarn mineralization.

 

7.1.2

FAD Area

 

   

The Hilltop Corridor stretching from the Hilltop fault zone adjacent to the Archimedes pit south over one mile to the FAD deposit. A 2022 IP survey outlined significant chargeability anomalies within the Hilltop corridor between the Archimedes Pit and FAD. In addition, the 2010 Barrick Titan MT survey contained lines through this corridor that indicated conductivity highs.

 

   

The recently acquired FAD deposit and surrounding area, adjacent to the historic Ruby Hill mine. The FAD CRD mineralization is located in the hanging-wall of the northwest striking Ruby Hill fault. The ore body was discovered in 1937 by Eureka Corp, Ltd. through surface core drilling. The ore zones consist of predominately shallow-dipping bodies of massive sulfide composed of pyrite, galena, and sphalerite, with minor amounts of other sulfides and sulfosalts. The ore is hosted in the Eldorado dolomite with approximate ore body dimensions of 1700 feet NW-SE, 900 feet wide, and 500 feet thick with most ore zones 10-50 feet thick. Carlin style mineralization is thought to overprint the area analogous to Archimedes overprinting Blackjack.

 

 

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Figure 7-1: Exploration Targets at Ruby Hill

(Source: i-80 Gold, 2023)

 

 

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7.2

Drilling

Wood provided a detailed account of historical drilling at the Project in support of their mineral resource estimate carried out in 2021 for Ruby Hill Mining Company (Wood, 2021), prior to i-80 commencing drilling on the property. Wood’s summary of historic drilling at Ruby Hill follows, with figures updated to include i-80 drilling. About 95 percent of the holes being used in the current resource estimation are in the Mineral Point Trend, and the remaining five percent are in the Archimedes area. A description of i-80 drilling procedures is appended as Section 7.2.11.

 

7.2.1

Historic Drilling at Ruby Hill

The RHMC drillhole database consists of data from over 3,600 drillholes and 2.3 million feet of drilling from throughout the southern portion of Eureka County. The database includes holes that have been drilled to test 24 different targets and includes reverse circulation, diamond core, reverse circulation pre-collar with diamond core tail and percussion and churn drill hole types. A total of 2,491 drillholes have been drilled on the current Ruby Hill property and 2,100 drillholes totaling 1.5 million feet of drilling define the Mineral Point Trend and Archimedes deposits. A plan view of the drilling in relationship to the Property boundary, and the drill collars attributed to the Mineral Point Trend and Archimedes Deposits are color coded in Figure 7-2.

The dataset used to produce the Mineral Resource Estimate for the Mineral Point Trend consists of drillhole data compiled from eight companies and work carried out from 1950 to 2015; however, 95% of the drilling was completed from 1992 to 2015 by Homestake, and subsequently by Barrick following completion of its acquisition of Homestake in 2004 (Table 7-1).

Just over 75% of drilling carried out at Ruby hill has been reverse circulation drilling. Diamond drilling has been used to provide drill core for detailed geological and geotechnical logging, metallurgical sampling, to extend reverse circulation holes below the water table to ensure representative sampling for assaying and as twin holes to confirm reverse circulation hole sampling. Mud rotary and other drill types have mainly been used to drill pre-collar holes for diamond drilling. Proportion of drilling by type is charted in Figure 7-3.

The following discussion of drilling, sampling, sample preparation and data verification is sub-divided into five main drill campaigns by owner and type where standards and procedures for data acquisition and confidence in data quality are relatively consistent. The five campaigns are RC and diamond core drilling by Homestake, RC and diamond core drilling by Barrick and the relatively minor amount of drilling carried out by other operators. Table 7-2 lists the distribution of drill footage by campaign and Figure 7-4 and Figure 7-5 show the location of the drilling by campaign in plan and fence section views.

 

 

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Figure 7-2: Drill Hole Collar Locations

(Source: i-80 Gold, 2023)

 

 

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Table 7-1: Drilling Statistics for Drillholes Included in the 2021 Ruby Hill Project Mineral Resource Estimate

 

           
Company   Drill
Holes
  Footage   Proportion of
Footage (%)
  Start Date   End Date
           

Eureka Corp.

  250   55,558   3.5   1950   1956
           

Hecla

  6   5,945   0.4   January 1960   August 1967
           

Newmont

  1   4,666   0.3   1970   1970
           

AMOCO-Cyprus

  27   3,962   0.3   1978   1978
           

Sharon Steel Corp.

  45   8,510   0.5   August 1982   November 1988
           

ASARCO

  2   635   0.0   July 1989   July 1989
           

Homestake

  1,172   771,445  

48.7

  March 1992   September 2004
           

Barrick

  597   733,667   46.3   October 2003   November 2015
           

Total

  2,100   1,584,387   100.0   1950   2015

 

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Figure 7-3: Distribution of Drill Types Included in the 2021 Ruby Hill Project Mineral Resource Estimate

(Source: Wood, 2021)

Table 7-2: Distribution of Drilling by Campaign

 

         
Owner   Type   Campaign   Footage   Proportion (%)
         

Homestake

  RC   1   638,077   40.3
  DDH   2   133,368   8.4
         

Barrick

 

RC

 

3

 

556,650

 

35.1

 

DDH

 

4

 

177,017

 

11.2

         

Other

      5  

79,275

  5.0
         

Total

          1,584,387   100.0

 

 

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Figure 7-4: Plan View of Drilling by Campaign

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Figure 7-5: Fence Section of Drilling by Campaign (Looking North)

(Source: i-80 Gold, 2023)

 

7.2.2

Drilling Methods

Drilling at Mineral Point was 83% by RC with 53% of drill footage drilled by Barrick and 28% drilled by Homestake. Approximately 8% of drilling was diamond core drilling by Barrick and Homestake. Eureka Corporation drilled approximately 46,000 feet of underground and surface drill core accounting for about 6% of total drill footage.

Drilling at Archimedes was 70% RC with 52% of drill footage drilled by Homestake and 18% drilled by Barrick. Approximately 30% of drilling at Archimedes was diamond core drilling and contributions by other operators is negligible.

 

7.2.2.1

Reverse Circulation Drilling

Barrick drilled 336 RC holes at Mineral Point Trend and 119 RC holes at Archimedes. RC holes were both vertical and inclined. Drilling was conducted by Eklund Drilling Company (Elko, NV), and Boart Longyear (Salt Lake City, UT). Where documented drilling was conducted with a TH-75 drill rig. Hole diameters ranged from 5.0 to 6.75 in. Drill logs indicate that for deeper RC holes intersecting the water table, if the RC hole could not be kept dry during drilling it was extended using diamond drilling.

Homestake drilled 381 RC holes at Mineral Point and 671 holes at Archimedes. The majority of RC holes drilled by Homestake were vertical. Drilling was conducted by Eklund Drilling Company (Elko, NV). Where documented holes were drilled with an MPD-1500 drill rig. Hole diameters ranged from 4.75 to 6.0 inches.

Asarco drilled two short RC holes at Archimedes in 1989. Drilling was conducted by Eklund Drilling Company (Elko, NV), and Hackworth Drilling, Inc. (Elko, NV).

Sharon Steel drilled 45 vertical exploration and definition RC holes totaling 8,510 feet. Drilling was conducted by a number of companies including O’Keefe Drilling (Butte, MT), Boyles Brothers, Polar Drilling, Lang Exploratory Drilling (Elko, NV), and Tonto Drilling Services, Inc. (Salt Lake City, UT). Where documented drill rigs used were a Jaswell 2400, Long Year 44 core rig adapted for RC drilling, Drill Systems CSR 1000, Chicago Pneumatic 650 WS, and T4W. Where noted, hole diameters were 5.25 inches.

 

 

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Eureka Corporation completed 2,788’ of RC drilling in two holes at Mineral Point. Drilling was conducted by Sierra Drilling Company (Bakersfield, CA). Drilling equipment, drill procedures, and sampling procedures from the Eureka RC drilling are not documented.

 

7.2.2.2

Core Drilling

Barrick drilled 131,375 feet of diamond drill core holes at Mineral Point and Archimedes. 38,800 feet of the total were diamond drill tails from RC precollars, including the total footage downhole from the collar. Drilling was conducted by a number of companies including Boart Longyear (Salt Lake City, UT), Dynatec Drilling, Inc. (Salt Lake City, UT), Major Drilling (Elko, NV), EMM Core Drilling Services (Winnemucca, NV), National Drilling (Elko, NV), and Connors Drilling, LLC (Montrose, CO). Where documented, core sizes drilled include PQ (3.345 in), HQ3 (2.406 in), HQ (2.5 in), and NQ (1.875 in). Where noted, an LF90 D drill rig was used. Most core holes are inclined.

Homestake drilled 133,368 feet of core holes at Mineral Point and Archimedes. Drilling was conducted by a number of companies including Tonto Drilling Services, Inc. (Salt Lake City, UT), Boart Longyear (Salt Lake City, UT), Connors Drilling LLC (Montrose, CO), Inland Pacific Drilling (Newman Lake, WA), and Westec/Haztec Drilling, Inc. (Meridian, ID). Where documented, drill rigs used were an LS-244 truck mounted rig and an LY44 drill rig. Hole size was HQ (2.5 in), reduced to NQ (1.875 in) when poor ground conditions dictated. Holes were both vertical and inclined, drilled on azimuths of 025° to 357° and inclinations of -45° to -87°.

Hecla drilled two vertical surface core holes totaling 3,511.5 feet. Drilling was conducted by Nichols Universal Drilling Co., Sprague & Henwood Inc., Continental Drilling Company, and Boart Longyear (Salt Lake City, UT). Where documented, the drill rig used was a Longyear 34 diamond drill. Where noted, holes were collared with NX (2.125 in) size core and reduced to BX (1.625 in) or HQ (2.5 in) size core reduced to NQ (1.875 in), dependent on depth and/or ground conditions.

Eureka Corporation drilled 239 exploration and definition core holes totaling 46,123.8 ft with 232 holes drilled underground and 24 collared at surface. Forty-seven were vertical and the remaining 214 were oriented with azimuths that ranged from 006° to 359° and inclinations of -70° to -85°. Drilling was conducted by Boyles Brothers. Holes were typically collared with NX (2.125 in) size core, and reduced to BX (1.625 in), AX (1.125 in) or EX (0.845 in) core size as depth and ground conditions necessitated. Drilling equipment and drill procedures are undocumented.

 

7.2.2.3

Other Drilling Methods

Amoco-Cyprus drilled 25 exploration mud rotary holes totaling 3,830 ft, and 2 exploration air track holes totaling 1,143 ft. All holes were vertical. Drilling equipment, drill procedures, and sampling procedures are undocumented.

Newmont drilled three vertical mud rotary exploration holes totaling 11,697 ft. Collared hole size ranged from 11 to 15 in with reduction to 9.625 and 6.75 in as depth and ground conditions necessitated. Drilling equipment and drill procedures are undocumented.

Hecla drilled five mud rotary holes totaling 2,496 ft, and 3 churn holes totaling 1,143 ft. Mud rotary and churn holes were vertical. Where documented, drilling was conducted by Continental Drilling Company, and Boyles Brothers. Drilling equipment, drill procedures, and sampling procedures are undocumented. Hole size for mud rotary drilling was 5.625 in, whilst hole sizes for churn holes are undocumented.

Eureka Corporation drilled seven mud rotary holes totaling 7,011 ft, and nine churn holes totaling 4,802 ft. All holes were vertical. Drilling equipment, drill procedures, and sampling procedures are undocumented. Mud rotary holes ranged from 8.5 to 9.0 in in diameter, and churn hole sizes ranged from 10 to 15 inches.

 

 

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7.2.3

Geological Logging

 

7.2.3.1

Barrick

Barrick geologists captured RC and core logging data on graphic strip logs on paper. The parameters captured included:

 

   

stratigraphic unit, rock type

 

   

chert intensity and color

 

   

oxidation characteristics, iron oxide occurrence and intensity

 

   

modal percentage of pyrite and total sulfides

 

   

intensity of silicification, decalcification decarbonatization, dolomitization, and skarn alteration

 

   

percentage of vein calcite and quartz

 

   

estimated percentage of downhole contamination (for RC)

 

   

intensity of realgar, orpiment, scorodite, carbon, carbonate mineralization

 

   

structure types and orientation

Graphic logs have been retained in a folder for each hole including original assay sheets, downhole survey reports, daily drill company sheets and notes on performance of quality control samples, database issues and other drilling issues.

 

7.2.3.2

Homestake

Homestake logging was also captured on graphic strip logs, on paper and captured many of the same parameters as the Barrick log sheets. The Homestake log sheets are also retained in drillhole folders and binders.

 

7.2.3.3

Logging by Other Operators

Logging by all other historic operators was also captured on paper and the parameters logged include rock type, structure, alteration, mineralization and oxidation intensity and handwritten notes about drilling including water flow.

 

7.2.4

Sample Recovery

Core recovery for the Barrick drilling programs was 92% and only suffered in broken zones. Core recovery for the Eureka Corporation, Hecla, and Homestake core drill programs are unknown.

Churn, rotary, percussion, air track and RC sample recovery for all drill programs is not documented.

 

7.2.5

Collar Surveys

Collar survey data exists for holes drilled from 1992 to 2015 when Homestake and Barrick were conducting mining operations at Ruby Hill. Collar locations were captured by mine survey personnel using a Trimble 4400 differential GPS survey system with centimeter accuracy.

The method of survey is unknown for drilling conducted prior to 1992.

 

7.2.6

Downhole Surveys

Barrick engaged International Drilling Services (IDS) of Elko, Nevada, to conduct downhole surveys with measurements collected every 50 ft using a Humphrey Gyroscopic System instrument. Dependent on the survey year, declinations used to convert magnetic north to grid north migrated from 13° to 16.25° E.

 

 

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Homestake employed both Silver State Survey, Inc. (NV) and Wellbore Navigation, Inc (CA) to conduct downhole surveys. Surveys were conducted on 50 ft intervals. Surveys conducted by Silver State Surveys, Inc used a Sperry Sun downhole camera survey instrument, and Wellbore Navigation, Inc. used an Inrun Survey Minimum Curvature gyro reference system bearing True North. Declinations are undocumented.

Survey procedures for earlier operators were variable and, in some cases, poorly documented:

 

   

Eureka Corporation holes were surveyed by Houston Oil Field Material Company (HOMCO) of California at 100 to 200 ft intervals. Survey type, equipment and declination are undocumented.

 

   

Newmont engaged HOMCO and Eastman Directional Drilling Oil Well Services (Denver, CO) to conduct downhole surveys at 100 ft intervals. Survey type and equipment are undocumented. Where documented, a declination of 17.5° E was used.

 

   

Hecla captured directional surveys at 100 and 200 ft intervals downhole but the surveyor, survey type, survey equipment and declination are undocumented.

 

   

It is unknown if Amoco-Cyprus, Sharon Steel or ASARCO conducted downhole surveys.

 

7.2.7

Metallurgical Drilling

In 2004 Barrick completed a cyanide soluble assay metallurgical program on mineralized drill intervals from East Archimedes to assist in gold recovery modeling. Material from 12 RC and two core holes were used (Table 7-3). A mineralogical study of 17 select samples was also conducted by Barrick Metallurgical Services Mineralogy Lab.

Table 7-3: 2004 Barrick Metallurgical Holes

 

               
Hole ID   Easting
(ft)
  Northing
(ft)
  Elevation
(ft)
  Azimuth
(degree)
  Inclination
(degree)
  Length
(ft)
  Hole
Type
               

HRH237

  12260.0   117964.0   6509.0   45   -60   1,000.0   RC
               

HRH256

  12336.0   118502.0   6490.0   94.5   -48   1,045.0   RC
               

HRH262

  12350.0   118500.0   6500.0   123.9   -54   905.0   RC
               

HRH335

  11944.9   118171.8   6512.7   0   -90   945.0   RC
               

HRH385

  12016.2   118522.7   6503.9   0   -90   1,000.0   RC
               

HRC271

  12226.2   118310.1   6504.8   88.3   -60   1,983.0   Core
               

HC1408

  12468.8   118515.6   6479.7   0   -90   924.5   Core
               

HRH1387

  12086.7   118879.8   6497.0   0   -90   1,305.0   RC
               

HRH1389

  12787.6   118455.5   6472.5   0   -90   1,400.0   RC
               

HRH1400

  12436.4   118381.6   6483.6   0   -90   1,285.0   RC
               

HRH1402

  12724.0   118074.0   6468.0   0   -90   940.0   RC
               

HRH1407

  12640.2   118673.7   6459.4   0   -90   1,355.0   RC
               

HRH1413

  12661.1   118144.7   6479.9   0   -90   1,100.0   RC
               

HRH1415

  12861.8   118527.1   6464.6   0   -90   1,200.0   RC
               

HRH1416

  12855.6   118670.2   6460.8   0   -90   1,485.0   RC

In 2009 Barrick engaged Kappes, Cassidy & Associates (KCA) of Reno, Nevada to complete metallurgical testwork on Archimedes drill holes. Material from 2 RC (hole size undocumented) and 10 core holes were used (Table 7-4).

In 2010 and 2011 Barrick engaged KCA to complete metallurgical testwork on Mineral Point core (Table 7-5) and RC cuttings identified as “Watertank RC material” (hole number(s) undocumented).

 

 

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In 2011, 16 refractory and two oxide samples from the 426 zone were tested at Barrick Technology Centre. Samples from nine core holes (Table 7-6) were received for the test program.

Table 7-4: 2009 Metallurgical Holes

 

               
Hole ID    Easting
(ft)
   Northing
(ft)
   Elevation
(ft)
   Azimuth
(degree)
   Inclination
(degree)
   Length (ft)    Hole Type
               

HRH1766

   11,552.0    119,810.9    6,440.8    225    -50    1,305.0    Core
               

BRH-36C

   10,639.0    119,759.4    6,453.9    106    -48    1,500.0    Core
               

BRH-37C

   10,626.2    119,757.3    6,453.9    140    -59    1,481.0    Core
               

BRH-38C

   10,864.4    119,628.9    6,445.4    133    -69    1,463.0    Core
               

BRH-41C

   10,855.2    119,644.0    6,444.6    175    -62    1,269.0    Core
               

BRH-67C

   10,979.2    119,697.6    6,448.1    102    -70    1,141.0    Core
               

HRH1767

   11,551.4    119,806.2    6,440.8    213    -69    960.0    RC
               

BRH-08C

   12,563.2    118,542.1    6,466.3    35    -90    2,062.0    Core
               

BRH-06C

   12,804.3    118,663.3    6,464.5    181    -76    2,168.0    Core
               

BRH-12C

   12,936.0    118,662.1    6,453.8    180    -80    2,044.0    Core
               

BRH-18C

   12,797.4    118,667.9    6,464.5    173    -80    2,168.0    Core
               

BRH-17C

   12,556.3    118,712.1    6,473.8    175    -76    1,750.0    RC

Table 7-5: 2010 and 2011 Metallurgical Holes

 

               
Hole ID    Easting
(ft)
   Northing
(ft)
   Elevation
(ft)
   Azimuth
(degree)
   Inclination
(degree)
   Length (ft)    Hole Type
               

BRH-165C

   9,200    119,018    6,464    131.7    -88.6    1,403.0    Core
               

BRH-166C

   8,618    119,550    6,447    173.4    -88.9    682.0    Core
               

BRH-184C

   7,297    118,088    6,497    45.9    -69.7    1,180.0    Core
               

BRH-231C

   8,536    118,703    6,405    42.2    -89.7    1,102.0    Core
               

BRH-235C

   8,709    118,879    6,428    36.3    -89.5    1,093.0    Core

Table 7-6: 2011 Metallurgical Holes

 

               
Hole ID    Easting
(ft)
   Northing
(ft)
   Elevation
(ft)
   Azimuth
(degree)
   Inclination
(degree)
   Length (ft)    Hole Type
               

BRH-95C

   11,361.8    119,737.3    6,453.0    130.4    -70.3    1,672.0    Core
               

BRH-99C

   11,138.3    119,826.4    6,478.7    97.3    -83.5    1,660.0    Core
               

BRH-103C

   10,945.9    119,742.0    6,447.2    134.7    -79.9    1,500.0    Core
               

BRH-210C

   11,319.4    120,089.8    6,505.0    113.0    -74.7    1,380.5    Core
               

BRH-211C

   11,322.5    120,059.2    6,505.5    158.2    -79.0    1,338.0    Core
               

BRH-212C

   11,163.9    119,805.8    6,478.1    126.2    -78.7    1,277.0    Core
               

BRH-213C

   11,128.5    119,810.3    6,477.7    152.0    -70.7    1,202.0    Core
               

BRH-214C

   10,822.3    119,793.5    6,446.8    138.3    -63.4    1,266.0    Core
               

BRH-215C

   10,737.3    119,806.8    6,446.5    145.2    -57.9    1,156.5    Core

 

 

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7.2.8

Sample Length/True Thickness

Approximately 66% of the drilling at Ruby Hill was vertical, producing essentially true- width intercepts through the relatively flat-lying mineralized zones. The remaining holes (34%) have steep inclinations and intersect mineralized units at high angles. Figure 7-5 provides an image of drill hole intersections with the mineralized bodies.

 

7.2.9

Potential Downhole Contamination

Oakley (1997) of the Elko Mining Group, a subsidiary of Waterton Global Resource Management, conducted a study of potential downhole contamination of reverse circulation holes drilled by Barrick. The study included compilation of intervals from 18 drill holes identified as having potential downhole contamination from drill core logging by Barrick, analysis of decay and cyclicity, and comparison of twin RC-diamond core holes including preparation of histograms and Q-Q plots comparing the grade distributions of twin holes, and downhole grade profile plots. The study concluded that the holes identified as being potentially contaminated by Barrick project geologists were likely contaminated and identified additional drillholes and intervals with potential sampling and assaying issues. The study culminated in a list of 30 holes for exclusion, nine holes having depths below which assays were suspected of being contaminated, and were flagged for exclusion, six holes with intervals flagged for exclusion, and six holes with anomalous silver grades that were flagged for exclusion.

A comprehensive review of Barrick, Homestake and other company drilling by Wood, and identification of additional intervals for exclusion is presented in Section 8.

 

7.2.10

Summary and Interpretation of All Relevant Drilling Results

Figure 7-4 and Figure 7-5 provide an example of the Ruby Hill drilling and the outlines of the mineralization in the Mineral Point and Archimedes deposits and illustrates the variability of density of drilling, the widths of mineralized intersections and drillhole intersection angles to mineralization. A discussion of the distribution and types of material intersected in metallurgical drilling, metallurgical test work composites, and an interpretation of the results of the metallurgical test work are presented in Section 10. Examples of the interpretation of the drilling in the construction of geological models and use of the interpretations in mineral resource estimation are presented and discussed in more detail in Section 11.

 

7.2.11

i-80 Drilling

i-80 completed 9,883 feet of drilling in 2021 at Ruby Deeps for infill. Holes were drilled using RC pre-collars followed by HQ core-tails. In 2022, 137,210 feet of drilling was completed with a mix of core and RC. Core drilling was conducted by National Drilling (Salt Lake City, UT) with RC conducted by Envirotech Drilling (Winnemucca, NV). Where documented, core sizes drilled include PQ (3.345 in), HQ3 (2.406 in), HQ (2.5 in), and NQ (1.875 in). Most holes were inclined. Thirty-six i-80 holes totaling 75,546.5 feet contributed to the current resource estimation, representing about 30 percent of holes flagged for use in the Ruby Deeps and 426 deposits. The remainder of the i-80 drilling was in exploration areas including Hilltop, Ruby Deeps expansion, Blackjack definition, 428, and Blue Sky. All of the i-80 holes contributing to the current resource estimation were drilled using core or RC precollar with core tail.

Drill hole collars are surveyed by the Ruby Hill surveyor using Trimble equipment with sub-centimeter accuracy referencing a local base station with GNSS rover. Coordinates are collected in the Ruby Hill mine grid, NAD83(2011), US survey feet. Downhole surveys are performed by IDS using a north seeking gyro.

i-80 logs geological characteristics of drill samples in Excel, filling data fields similar to those recorded by Barrick and Homestake but with additional focus on sulfides to support characterization of CRD

 

 

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mineralization. Data is organized in Excel sheets with tabs for geotechnical, sampling, lithology, alteration, oxide, sulfide, structure, point data, veins, density and water level information. Core recovery averaged about 93%, which is comparable to core recovery of Barrick’s drilling campaigns. Recovery within the modeled mineral envelopes is similar to overall recovery, about 93% for Ruby Deeps and about 95% for 426. The slightly higher recovery within the 426 zone correlates with relatively high RQD values in the OGLL host unit.

Geometry of Carlin type mineralization at Ruby Hill is well understood, and drill spacing is close enough to allow true thickness to be reasonably represented by interpolating between mineralized intercepts in adjacent drill holes. Most holes intersect the mineralization at near-normal orientations.

 

7.3

Hydrogeology

 

7.3.1

Sampling Methods and Laboratory Determinations

Hydrogeological data, including water table measurements, pore pressure distribution, and direction of groundwater flow, were collected in conjunction with exploration and geotechnical investigations in pre-construction studies and later from hydrogeological studies for on-going programs in the pit and planned underground mining areas.

Groundwater dewatering and monitoring wells are the primary method of collecting hydrogeological data in support of mining operations, as well as the collection of pore pressure data, which can be converted to groundwater level elevations, from a network of vibrating wire piezometers (VWPs). Another source of data is hydrologic testing. Most wells that are drilled undergo hydrologic testing to estimate aquifer parameters. These tests include injection (slug) tests, air-lift tests, and short-term and long-term pumping tests. Data obtained from testing operations are analyzed using industry standard analytical methods. Analytical and numerical groundwater flow models have been developed based on 3D geological modeling and supported by the site-specific aquifer test analysis results.

From approximately 1997 through 2024, a total of 17 dewatering wells and 54 monitoring locations were completed in the Project area. In 2006, rapid infiltration basins (RIBs) were constructed east of the Project area to infiltrate groundwater pumped from dewatering operations into downgradient, permeable alluvial sediments. During 2022, 16 vibrating wire piezometers were installed north of the Archimedes Pit to increase monitoring in the planned underground mining operations area (HGL, 2022). Currently, there are six active dewatering wells, nine active monitoring wells, and 47 active vibrating wire piezometers (VWPs) across 33 locations. Current dewatering pumping rates range from 30 gpm to 110 gpm from the six dewatering wells. All dewatering wells are monitored, controlled, and data are logged using a supervisory control data acquisition system (SCADA) or manually collected daily if not equipped.

According to permitting requirements, 13 monitoring wells are sampled on a routine basis and analyses run for the State of Nevada Profile I suite at a certified analytical laboratory, currently Western Environmental Testing Laboratory (WETLAB), Reno, NV. Monitor wells and exploration drill holes that have piezometers installed are monitored for water levels and piezometric heads.

 

7.3.2

Hydrogeology Investigations

Throughout the span of various mine property owners and operators, the Project area has been the subject of multiple studies aimed at characterizing the hydrogeologic properties of the stratigraphy within the Project area and the surrounding region (Table 7-7). Water Management Consultant (WMC, 2004) and Jones (2004) developed early conceptual hydrogeological and groundwater models, as well as characterizing the physical properties of major water bearing geologic units for the East Archimedes Pit Expansion involving deepening of the existing pit below the groundwater table. Continuing from 2005 through 2021, additional hydrogeologic studies were completed by WMC, Schafer Limited, John Shomaker Associates, Inc. (JSAI), WSP,

 

 

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FloSolutions, SRK and Piteau Associates in support of groundwater monitoring, dewatering operations, and water balances (Schafer 2005 and 2010; JSAI 2010, 2012, 2013, 2015, and 2021; WSP 2016; Piteau Associates 2017 and 2018; FloSolutions 2020; and SRK Consulting 2021).

Table 7-7: Summary of Hydrogeological Surveys Since 2004 (Wood 2021)

 

Date    Report/Investigation    Author/Lead Consultant

2004, October

  

East Archimedes Project, Groundwater Flow Model

  

Jones, M.A.

2004, October

  

East Archimedes Project, Assessment of the Hydrogeologic Conditions and Dewatering Feasibility

  

Water Management Consultants,

2005, May

  

Final East Archimedes Pit Lake Water Quality

  

Schafer Limited LLC

2010, August

  

Revised Archimedes Pit Lake Water Quality

  

Schafer Limited LLC

2010, June

  

Final Ruby Hill Mine Groundwater Flow Model 2010 Update

  

J. Shoemaker and Associates Inc,

2011

  

Aquifer Test

  

J. Shoemaker and Associates Inc,

2012

  

Ruby Hill Mine Groundwater Flow Model

  

J. Shoemaker and Associates Inc,

2012, June

  

Bullwhacker Dewatering Evaluation

  

J. Shoemaker and Associates Inc,

2013

  

Spring Investigation

  

J. Shoemaker and Associates Inc,

2015

  

Aquifer Test, Mineral Point Dewatering Projection

  

J. Shoemaker and Associates Inc,

2015

  

Aquifer Test, Base Metals Dewatering Projection

  

J. Shoemaker and Associates Inc,

2016, September

  

Ruby Hill Groundwater Characterization and Dewatering Update – Technical Memorandum

  

WSP Parsons Brinkerhoff

2016, December

  

Pit Lake Water Balance and Evaporation to Validate Water Rights Requirements

  

WSP Parsons Brinkerhoff

2017, July

  

Ruby Hill Mine Pit Lake Study

  

Piteau Associates Engineering Ltd.

2018 July

  

Mineral Point PW-15 Pumping Test and Updated Hydrological Model

  

Piteau Associates Engineering, Ltd.

2020, May

  

Draft Ruby Hill Produced Water Management Plan, Preliminary Hydrogeological Conceptual Model and Alternatives Analysis

  

FloSolutions

2020 June

  

Draft Ruby Hill Rib Characterization Plan

  

FloSolutions

2021, March

  

Ruby Hill Mine Pit Lake Geochemical Model Report

  

SRK Consulting (U.S.), Inc.

2021, March

  

Ruby Hill Project Water Level and Water Balance for Permitted and Existing Pits – Technical Memorandum

  

J. Shoemaker and Associates Inc,

More recently, i-80 contracted HydroGeoLogica Inc. (HGL), now part of LRE Water, to conduct operations for monitoring of groundwater levels and pore pressures, plan and oversee operations of dewatering wells, and groundwater flow modeling for local-scale dewatering and regional scale permitting related to the 426 and Blackjack planned underground operations (HGL, 2023).

 

7.3.3

Hydrogeologic Description

The Ruby Hill Mine is in Eureka County, Nevada. The mine is located at the south end of Diamond Valley, about 1 mile from the town of Eureka. Diamond Valley, delineated as Hydrographic Basin 153 by the U.S. Geological Survey (USGS) and Nevada Division of Water Resources (NDWR), is a narrow north to south-oriented basin with a drainage area of approximately 748 square miles. The basin boundaries are formed by Sulphur Spring Mountain along its western margin, the Diamond Mountains along the eastern margin, and the Fish Creek Range at the southern margin. The basin extends approximately 45 miles along its N-S

 

 

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axis and its average width is roughly 15 miles. All facilities associated with the Ruby Hill Mine are in the basin Figure 7-6.

The high elevation mountain-block areas on the perimeter of the basin receive the majority of the basin’s annual precipitation and are the principal source of groundwater recharge. Recharge enters the subsurface directly into bedrock and as runoff from the mountains that infiltrates to groundwater through alluvial channels and fans. Surface runoff and subsurface flow from upstream basins enters Diamond Valley at Devils Gate, a topographic low on the west margin of the basin.

Most drainages flow intermittently because of seasonal snowmelt or extreme precipitation events. Runoff diminishes rapidly down slope over the alluvial fans, as water flows into the ground. Figure 7-7 shows the pre-mining groundwater level conditions and surface geologic units of the study area, grouped simply into six hydrogeologic units: recent alluvium, older alluvium, carbonate rock, volcanic rock, non-carbonate sedimentary rock, and intrusive rock. Groundwater moves from the perimeter bedrock highlands comprising mostly carbonate rocks, toward the interior of the basin comprising a deep basin-fill aquifer consisting of coarse to fine-grained sediments. Water is removed from the basin as groundwater pumping and as evapotranspiration. (Jones, 2004).

 

7.3.3.1

Surface Water

The Diamond Valley Basin is characterized as a closed watershed (endorheic): the only natural discharge from the basin occurring as evaporation and plant transpiration, primarily at the playa located at the northern end of the basin. There is no surface water or groundwater discharges from the basin.

The Project area is within the Lower Slough Creek-Frontal Diamond Valley subwatershed (Hydrologic Unit Code [HUC] 160600051503) within the larger Diamond-Monitor Valleys watershed (HUC 16060005). Surface water within the Project Area is dependent on seasonal precipitation. Precipitation data from the Eureka, Nevada, Station (Western Regional Climate Center [WRCC]) for a period of record (POR) 1903-2022 indicates average precipitation is 11.64 inches (Nexus, 2022).

 

 

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Figure 7-6: Diamond Valley Hydrographic Basin and Ruby Hill Mine Permit Area

(Source: LRE Water, 2025)

 

 

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Figure 7-7: Surface Geology and Pre-Mining Groundwater Level Contours

(Source: Jones, 2004)

 

 

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Surface hydrology consists of surface seeps and springs, rainfall, and snow melt. There are some seeps and springs upgradient of the Project Area; however, none contribute to channelized surface flow in the Project area. Hydrology at the Project is the result of precipitation runoff and snowmelt that drains generally north from the mountains located to the south of the Project Area. When precipitation events occur or during snow melts, the water flows rapidly off the slopes in the area from the high topographic areas and slows as it reaches shallow valleys toward agricultural fields in Diamond Valley. Ephemeral flows may occur seasonally in the Project area during spring snow melt or after intense storms that produce large amounts of precipitation. If these runoff events are large enough to create flow in the valley channels, they are diverted into the agricultural fields (Nexus, 2022).

JSAI (2013) conducted a desktop study of springs 2012, followed by a field reconnaissance during November 2012 that identified 61 springs within the southern part of Diamond Valley. Results indicated the source of water for the identified springs in the mountain watersheds above Eureka is related to perched or locally sourced groundwater not connected to the groundwater system at Ruby Hill Mine. Springs in the immediate vicinity of Eureka have different water quality than the groundwater at Ruby Hill Mine and are unrelated to the regional aquifer.

 

7.3.3.1

Groundwater

The Project area lies within the southern portion of the Diamond Valley Hydrographic Basin. Diamond Valley is the terminal basin in the flow system, receiving sub-surface groundwater flow from the upgradient basins through basin-fill alluvial and possibly carbonate-rock groundwater systems. Estimates for the inflow rate to Diamond Valley vary between 16,000 and 35,000 acre-feet per year (Berger et al., 2016).

The alluvial aquifer system in Diamond Valley is basin-fill deposits ranging from fine to coarse-grained unconsolidated materials eroded from the adjacent mountain ranges. Geologic logs from oil and gas drilling indicate the total basin fill thickness is up to 7,500 ft with significant portions containing fine-grained sediments. The water table is separated from deeper confined aquifers by clay beds and lower conductivity geologic units in some areas (Tumbusch and Plume, 2006).

Harrill (1968) identified the zones of lowest hydraulic conductivity in Diamond Valley as being along the south, southeast, and west valley margins, and in the north central area. A zone of high hydraulic conductivity is found in the south-central part of the valley where irrigated agriculture occurs. Prior to increased irrigation in recent years, the groundwater flow direction in the basin was generally northward towards the playa. This is consistent with the drainage pattern in the project area, where water flows from the higher elevations in the south, towards Diamond Valley in the north.

The southern Diamond Valley basin fill aquifer has undergone a water level decline of approximately 50 ft or more since irrigation pumping began in the 1950s. The rate of decline increased during the mid-1970s. By 1990, the water levels were declining at rates of 1.5 to 2.5 ft/yr (Arteaga et al, 1995). Locally, some well water levels had dropped by up to 90 ft in 2005 (Tumbusch and Plume, 2006). Groundwater in the basin now flows generally toward the area of most concentrated pumping.

The bedrock aquifer system occurs as groundwater movement primarily in the higher permeability carbonate rocks while the siliciclastic sedimentary formations act predominantly as confining, or lower-permeability units. Over time the carbonate rocks have been extensively fractured and faulted. Carbonate rocks are also subject to dissolution interaction with groundwater. The dissolution features form preferential flow pathways that define and reinforce groundwater flow paths. Faults and igneous intrusions in the carbonate rocks result in compartmentalization of the aquifer system. Thrust faults and normal faults can create conduits for groundwater flow in the carbonate rocks, but they can also impede groundwater movement where they

 

 

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juxtapose hydrogeologic units of differing hydraulic conductivity and/or due to low-permeability fault gouge (Tumbusch and Plume, 2006).

The faulted structural blocks of the Ruby Hill Mine area have formed a network of hydraulic compartments. Where water moves easily within and between compartments, groundwater gradients are shallow. Where flow between compartments is structurally restricted, groundwater levels can show abrupt changes across short lateral distances. Faults typically act as barriers to flow perpendicular to the fault but can act as conduits along the fault plane.

Regionally, groundwater recharge occurs to both the alluvium and the bedrock of the upper piedmont slope elevations and, during years of high run-off, to the alluvium at middle and lower piedmont slope elevations. Groundwater moves towards the center of the basin in the thickening sequences of alluvial deposits. Most natural discharge from the basin occurs through evapotranspiration from the alluvial deposits beneath the valley floor. Locally, historical and current dewatering of the Archimedes Pit has influenced direction of groundwater movement in the vicinity of the mine. Local groundwater movement is also influenced by delivery of water from dewatering operations to two RIBs constructed in the alluvial aquifer system for downgradient recharge to the basin.

 

7.3.4

Mine Dewatering

 

7.3.4.1

Archimedes Pit

The West Archimedes deposit is hosted in Ordovician Upper Goodwin limestone and is bounded by the Holly Fault to the West. The mineralization for this deposit is oxidized and was mined as an open pit above the water table from 1998 through 2002.

The East Archimedes deposit, east of the Graveyard fault and hosted in the Upper Goodwin formation, extends downward through the Lower Goodwin formation nearly 1,800 feet from ground surface. This zone was mined from 2004 through 2013. In 2013, a slope failure on the south wall of the pit caused suspension of mining activities (Wood, 2021). The mine remained in care and maintenance until early 2020 when the remaining accessible benches of the East Archimedes pit were mined through mid-2021.

Active dewatering started in 2006 to lower the water table below the planned pit expansion and has continued though present. Maximum permitted discharge for dewatering operations is 1,000 gpm and historical pumping rates have reached approximately 850 gpm from up to 10 dewatering wells. Figure 7-8 provides a map showing the network of dewatering and monitoring locations. Currently, there are six operational dewatering wells PW-9, PW-10, PW-11, PW-13, PW-16, and PW-17. Production from the dewatering wells is between 30 and 113 gpm. Due to pump efficiency concerns, PW-9, PW-10, and PW-11 are periodically cycled off to allow for groundwater recovery prior to continued pumping. The current combined average pumping rate of 250 gpm within the Archimedes block has maintained groundwater levels below 5,450 ft amsl, approximately at the current bottom pit elevation. PW-17 completed in the Holly hydrologic block, is the only currently operating dewatering well that is not within the Archimedes block.

 

7.3.5

Dewatering Discharge

Water pumped from the dewatering wells not utilized for mining operations is currently discharged to RIBs on the west side of the project area through HDPE pipelines. Two cells, RH-1 and RH-2 are in operation (NEV2005106), with discharge to one of the two cells at any given time. When RIB maintenance is required, discharge is routed to the dormant cell. Current dewatering efforts are well under the permitted 1,000 GPM threshold of the RIBs and the RIB infiltration is sufficiently limiting surface ponding in the active cell.

 

 

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Figure 7-8: Dewatering Well and Groundwater Monitoring Locations

(Source: LRE Water, 2025)

 

7.3.6

Groundwater Flow Model

 

7.3.6.1

Background

The project is proposing a modification to the currently permitted open pit mine plan (open pit plan) that would involve underground mining of the 426 and Blackjack deposits immediately adjacent to and below the existing Archimedes Pit (underground mine plan). The footprint of the 426 and Blackjack deposits are shown in Figure 7-9 relative to the plan of operations boundary together with the footprint of the proposed underground workings (UGWs). The underground mine plan was designed to remain within the ore block of the previously approved open pit mine plan.

 

 

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Figure 7-9: Property Overview showing Plan Operations Boundary, Existing Mine Operations Boundary, and Existing Archimedes Pit with Planned UGWs for the 426 and Blackjack Deposits

(Source: LRE Water, 2025)

The Archimedes Pit has been dewatered from a pre-mining water table of 5,910 ft amsl to a groundwater elevation of approximately 5,400 ft amsl in support of current mining operations. Both the open pit and proposed underground mine plans extend to a minimum elevation of 5,100 ft amsl which will require an additional 300 ft of dewatering. Figure 7-10 illustrates a cross-section through the Archimedes Pit area. The figure presents the current open pit shell together with the permitted ultimate open pit shell and the authorized maximum pit depth. The proposed underground workings are superimposed on this figure to illustrate their location and elevation within the existing pit disturbance and above the permitted maximum pit depth.

The underground mine plan is similar to the open pit mine plan in terms of dewatering and other hydrogeological factors, but avoids the removal of excess waste rock from the pit shell. UGWs would be backfilled with low hydraulic conductivity cemented rock fill. There would be no open hydraulic connections via UGWs post-closure, so the post-closure hydrogeologic flow regime would be nearly identical to pre-mining conditions.

 

 

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Figure 7-10: Schematic Section through the Archimedes Pit Area

(Source: LRE Water, 2025)

 

7.3.6.2

Model Overview

Groundwater flow modeling associated with the current mine plan by Jones (2004) and JSAI between 2010 and 2021 provides relevant dewatering, drawdown and pit lake analyses for the underground mine plan. HGL (2023) provided a summary and assessment of the groundwater flow model and updates which were developed to support dewatering and permitting at Ruby Hill and evaluated any potential differences that may be expected as a result of the proposed underground mine plan. Dewatering predictions were assessed relative to the currently proposed underground expansion of the 426 and Blackjack Deposits.

Since 2004, pumping and water table responses have been recorded and incorporated into a groundwater flow model for the site which has been maintained with regular updates to support permitting, mine dewatering and planning. These efforts are presented in several documents, including:

 

   

Jones (2004) - Initial groundwater flow model was developed to support the East Archimedes pit expansion;

 

   

JSAI (2010) – Groundwater flow modeling was updated for evaluation of a pit expansion. This version of the groundwater flow model simulates an increase in depth of the Archimedes Pit to 5,100 ft amsl;

 

   

JSAI (2011) – The groundwater flow model was updated with recently-acquired aquifer test data to verify model calibration and predictions.

 

   

JSAI (2012) – The groundwater flow model was updated to evaluate a mine expansion to include the 426, Archimedes, and Bullwhacker deposits.

 

   

JSAI (2013) – The groundwater flow model was used to evaluate the potential impacts to local, high-elevation streams and springs.

 

   

JSAI (2021) – The 2012 groundwater flow model was used to develop an updated pit lake water balance model to support the SRK (2021) geochemical model for the Water Pollution Control Permit (WPCP) renewal. The application was reviewed and approved by the Nevada Division of Environmental Protection (NDEP).

 

 

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7.3.7

Model Results

The historical groundwater flow model predictions have been calibrated and corroborated by recent and ongoing observations, and, as such, the groundwater flow model is considered a reliable tool. The design, calibration, and results of the JSAI (2012) groundwater flow model are summarized below.

 

   

The groundwater flow model was constructed in MODFLOW (McDonald and Harbaugh, 1988). The model area is shown in Figure 7-11 together with measured groundwater elevation from the first quarter of 2012. The model covers an area of approximately 180 square miles and comprises a grid of 94 rows, 93 columns, and 4 model layers (Figure 7-12).

 

   

The groundwater flow model incorporates detailed information on geology and structure based on exploration drilling results and hydrogeological investigations at the site and at the regional scale. Hydrogeologic units (geologic units with similar hydrogeological properties and behavior) were defined and represented in the model as ‘zones’, with geologic structures simulated as flow barriers or conduits. An example of the hydrogeologic zones and flow barriers for model layer 2 is provided in Figure 7-13.

 

   

Parameterization of hydrogeologic units, structures, and boundary conditions are based on observed, measured, and interpreted responses associated with long-term operational data and results of hydrogeologic testing programs/investigations. Steady-state and transient model calibrations were conducted to ensure the validity of model predictions.

 

   

The impact assessment involved predicting the pumping requirement to maintain a dry pit (down to a minimum mine elevation of 5,100 ft amsl) then simulating that drawdown through the project life. The pumping rates required for mine operations peaked at approximately 850 gpm at a groundwater and pit floor elevation of approximately 5,400 ft amsl, and then declined to about 600 gpm for the remaining operational period, well below the permitted maximum rate of 1,000 gpm.

 

 

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Figure 7-11: Ground Water Flow Model Boundary

(Source: LRE Water, 2025)

Figure 7-14 shows the projected maximum extent of the 10-foot drawdown contour (isopleth) at the end of the dewatering period (the end of active mining operations). The groundwater drawdown area is predicted to be limited to less than a mile to the north of the mine and extend up to approximately 4 miles to the south. The spatial difference in drawdown expression is due to:

 

   

The presence of thick sequences of alluvium with high hydraulic conductivity and storage to the north and northwest of the project; and,

 

   

The presence of low storage, fractured, and faulted bedrock to the south and southeast of the mine.

 

 

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Figure 7-12: Ground Water Flow Model Grid

(Source: JSAI, 2012)

 

 

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Figure 7-13: Mine-Area Hydrogeologic Zones and Flow Barriers, Layer 2

(Source: JSAI, 2012)

 

 

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Figure 7-14: Projected Changes in Groundwater Level, End of Mining

(Source: JSAI, 2012)

 

7.3.7.1

Model Summary

The groundwater flow model constructed by JSAI (2012-2021) was designed to represent dewatering of an open pit mine. The modeling is based on an extensive hydrological, hydrogeological, geological, structural and climatological data set and includes recent updates and a strong calibration to both steady state and transient conditions.

LRE Water (formerly HGL) has evaluated the modeling and concludes the current groundwater flow model dewatering predictions for the open pit mine plan are representative for the proposed underground mine plan for the following reasons:

 

   

All mining will be performed above the 5,100 ft amsl level under both plans.

 

 

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The underground mining will be within the proposed and permitted mine shell.

 

   

The currently permitted maximum dewatering rate (1,000 gpm) will be maintained for the proposed underground mine plan.

 

   

The same material in the authorized plan will need to be dewatered to the same depth in order to maintain a dry underground mine.

The projected pit dewatering rates of approximately 800 gpm and ultimate post-closure pit lake level of 5,800 ft amsl predicted by the JSAI (2012 and 2021) models reflect both the open pit mine plan and the proposed underground mine plan of the 426 and Blackjack Deposits. There are no significant hydrogeological or spatial differences that warrant additional model changes at this time considering the close proximity of the underground workings relative to the open pit plan. As operations progress, the process of improving the groundwater model is warranted by incorporating new data, adjusting parameters, modifying the model’s conceptual and numerical 3D framework, or refining the grid resolution to better represent the complex dynamics of the groundwater system, ultimately allowing for more accurate predictions and informed decision-making regarding water management strategies.

 

 

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8.

SAMPLE PREPARATION, ANALYSIS AND SECURITY

The following section describes procedures employed by previous operators at Ruby Hill for the security, laboratory preparation, and analysis of reverse circulation (RC) and core samples during the drilling programs completed from 1992 to 2015. The descriptions are largely summarized from previous technical and feasibility reports (Barrick, 2004; REI, 2005; Newman and Mahoney, 2008; Barrick, 2012; RPA, 2012; and Barrick, 2013). A description of i-80’s procedures follows in Section 8.9

 

8.1

Sampling Methods

Homestake and Barrick employed similar sampling procedures for RC drilling. For most RC holes, only bedrock was sampled with the exception select intervals of alluvium saved for waste rock characterization (Barrick, 2004 & 2013). RC cuttings were sampled on 5’ intervals except in 1992 when 10 ft intervals were used on select holes. Coarse and fine fractions of RC cuttings were collected in cloth or plastic sample bags.

Homestake and Barrick also employed similar procedures for sampling drill core. Core was sampled in consistent 5 ft intervals except where shorter intervals were dictated by geologic conditions. Core was cut in half along the long axis using a diamond saw, and a half-split was bagged and submitted to the laboratory for analysis.

Sampling methods are not well documented for the drill campaigns carried out before the Barrick and Homestake campaigns from 1992-2015. The following information has been compiled from the drill logs and interrogation of the drillhole database:

 

   

Eureka Corporation sampled rotary, RC and core holes on 5 ft or shorter intervals based on geologic conditions. Newmont samples were generally collected on 5 ft intervals, although intervals ranged from 1 to 10 ft based on geologic conditions.

 

   

Hecla rotary holes were sampled on 10 ft intervals, percussion holes on 20 to 30 ft intervals, surface RC holes on 10 ft intervals, and underground RC holes on 4 ft intervals. Surface core holes were sampled on 5 ft intervals, and underground core holes on 4 ft intervals although intervals for both hole types ranged from 0.5 to 10 ft based on geologic conditions.

 

   

Amoco-Cyprus sampling for mud rotary holes was conducted on 10 ft intervals. Air track holes were sampled on 6 ft intervals, although intervals ranged from 2 to 10 ft based on geologic conditions.

 

   

Sharon Steel sampling was conducted primarily on 5 ft intervals although 10 ft intervals were used based on geologic conditions.

 

   

Asarco sampling was conducted on 10 ft intervals.

 

8.2

Analytical and Test Laboratories

The Ruby Hill mineral resource estimate database is comprised of gold, silver, base metal and major and trace element geochemistry and density data acquired at independent laboratories. The majority of assaying of samples collected from drilling by Homestake was carried out at Berringer Laboratories in Reno Nevada, and assaying from the Barrick campaigns was carried out at the ALS Global laboratory in Reno Nevada. Details of other work are presented in Table 8-1.

 

8.3

Density Determinations

Density determinations were carried out during programs operated by Barrick with analyses at G&T Metallurgical Services in Kamloops, BC, Canada, McClelland laboratory in Reno, Nevada and at the Bald Mountain mine site in Nevada.

 

 

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8.3.1.1

Barrick

Material densities used for the estimation of mineral resources in the East Archimedes deposit in 2004 were determined by traditional volume displacement procedures using drill core (holes unknown) sealed by acrylics (Barrick, 2004). Average bulk density values obtained by the tests are shown in Table 8-2.

Between 2007 and 2008, G&T Metallurgical Services performed 41 bulk density measurements from four core holes from the East Archimedes deposit. Bulk density measurements were determined using the water immersion volume displacement method.

In 2008, Barrick submitted 49 samples from two core holes to MLI for bulk density determinations. Only 47 samples were analyzed with 2 samples rejected due to being broken. Bulk density measurements were made using a standard volume displacement method on oven dried, coated (spray lacquer finish) pieces of drill core.

Table 8-1: Assay, Density and Metallurgical Laboratories

 

Company    Year    Lab Name and Location    Accreditation    Testwork Performed
Eureka Corporation   

1950’s -

1960’s

   Union Assay Office, Inc, Salt Lake City, UT    Unknown    Au, Ag, Pb assays
Amoco    1980-1981    Unknown    Unknown    Precious and base metal assays
Sharon Steel   

1980’s,

1991

   Sharon Steel Corporation Mining Division    Unknown    Precious and base metal assays
Hecla   

1960,

1969

  

Union Assay Office, Inc, Salt Lake City, UT

Skyline Labs, Wheatridge, CO

   Unknown   

Union Assay: Au, Ag, Pb, Zn (no analysis information)

Skyline labs: multi-element

Homestake    1992-1993   

Barringer Laboratories, Reno, Nevada

Legend Assay Laboratory, Reno, NV

Bondar Clegg, Sparks, NV (acquired by ALS

Chemex, 2001)

   Unknown   

Barringer: Au-FA/AA, Path 7 (Ag, As, Cu, Hg, Pb, Sb, Zn)

Legend: Au-FA/AA, 1AT

Bondar Clegg: Au-FA/AA, Ag, As, Cu, Hg, Pb, Sb, and Zn

     1994-2001   

ALS Global (previously ALS Chemex Labs), Reno, NV

Bondar Clegg,

Vancouver, BC (acquired by ALS Chemex, 2001)

  

ALS Global - ISO Guide 25 moving to adopt ISO 9002

Bondar Clegg

moving to adopt ISO Guide 25

  

ALS Global: Au-FA/AA, Ag, As, Cu, Hg, Pb, Sb, Zn, and CN-Au

Bondar Clegg: Au-FA/AA, 35 multi-element suite, Hg

Barrick    2003-2015   

ALS Global, Reno, Nevada

BSI Inspectorate, Reno, NV

Kappes, Cassiday & Associates (KCA), Reno, NV

McClelland (MLI), Reno, NV

G&T Metallurgical Services, Kamloops, BC

Bald Mountain Mine Site, NV

  

ALS Global - ISO 9001:2000; ISO

17025:2000

BSI Inspectorate - ISO 9001:2000

certified

KCA was working towards ISO 9002 at the time

  

ALS Global: gold assays, multi-element geochemistry, density determinations

BSI Inspectorate: Au check assays

KCA: metallurgical testwork, Au assays

MLI, G&T and Bald Mountain: density determinations

RHMC    2017    ALS Global   

ALS Global - ISO 9001:2000; ISO

17025:2000

   Density determinations

 

 

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Table 8-2: Barrick Rock Type Density Values

 

Unit Density (cu. ft/st)

Alluvium

  14.5

Limestone (Goodwin Formation)

  13.5

Intrusive (Graveyard Flats)

  13.5

Volcanic Tuff/Rhyolite Flow

  13.5

Fill Material

  18.2

In 2007 and 2008 ALS Global conducted bulk density determinations on 38 samples from 10 core holes located in East Archimedes. Bulk density determinations were conducted using the OA-GRA09A method utilizing the following equation:

Bulk Density = A/C – [(B-A)/Dwax]

A = weight of sample; B = weight of waxed sample in air; C = volume of displaced water; Dwax = density of wax.

Between 2009 and 2015, Barrick conducted an additional 878 bulk density determinations from 71 holes located in the East Archimedes and Mineral Point deposit areas. Determinations were conducted by Barrick’s Bald Mountain mine site laboratory. The density determination method is unknown.

 

8.3.1.2

RHMC

RHMC collected samples representative of the different lithological, alteration and redox units for density determination. Twenty-two samples were collected from nine core holes collared in the Mineral Point area and submitted to ALS Global for analysis. Samples ranged from 0.25 to 0.60 ft in length. Bulk density determinations were conducted using the OA-GRA09A method using the following equation:

Bulk Density = A/C – [(B-A)/Dwax]

A= weight of sample; B = weight of waxed sample in air; C = volume of displaced water; Dwax = density of wax

 

8.4

Sample Preparation and Analysis

Sample preparation and analysis procedures for the Barrick and Homestake drilling are reasonably well documented and have been confirmed by reviewing assay certificates from these programs. Details of the sample preparation and assay procedures follow.

Details about sample preparation and analysis procedures for samples analyzed prior to the Homestake campaigns beginning in 1992 are not well documented.

 

8.4.1

Barrick

Exploration RC and core sample preparation and gold assaying were conducted by ALS Global. Sample preparation procedures included:

 

   

Samples were dried and weighed

 

   

Samples were crushed and screened to minus 2 mm

 

   

Samples were split to 500 g then pulverized to minus 75 µm (-200 mesh)

 

   

A 30 g pulp (one assay ton) was split for assay

 

   

Pulp excess and coarse rejects were retained and stored.

 

 

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All samples were assayed using standard 30 g charge, FA digest with AA. Samples with greater than 0.10 oz/st Au were rerun by FA with gravimetric finish. Samples with greater than 0.010 oz/st Au were assayed using cyanide digestion with AA finish. This cut-off was reduced to 0.005 oz/st Au in September 2006 to provide AA assays closer to mine cut-off grades (Barrick, 2013). Table 8-3 lists ALS Global gold analytical parameters.

Table 8-3: ALS Global Gold Analytical Parameters

 

ALS Global

Code

   Sample  Digestion      Assay/Analysis  

  Pulp Weight  

(g)

  Detection Limit  
(g/t Au)

 

   Upper Limit   
(g/t Au)

 Au-AA13

  Cyanide Leach   AAS   30   0.030   50

 Au-AA23

  Fire Assay Fusion   AAS   30   0.005   10

 Au-GRA21

  Fire Assay Fusion   Gravimetric   30   0.050   1,000

Mercury was analyzed using an aqua regia digestion with a cold vapor/AA finish (Hg- CV41). A 48 multi-element package (ME-MS61) included a 4-acid digest and inductively coupled plasma mass spectrometry (ICP-MS) finish. Base metal overlimits (>10,000 ppm) were rerun using an overlimit method with a 0.4 g charge, 4-acid digest and ICP finish.

 

8.4.2

Homestake

Except for approximately 15 RC holes that were prepared at the Ruby Hill mine site assay laboratory, all drill samples from the Homestake drill programs were prepared at independent commercial laboratories including Barringer (1992-1993), Legend (1992-1993), ALS Global (1993-2001), and Bondar Clegg (1992-2001).

Barringer Laboratories (Barringer)

No documentation exists for the preparation procedures used for samples by Barringer. Gold content was determined using a 30 g charge with a fire assay (FA) digest and atomic absorption (AA) finish. Detection limit was 1 ppb. Samples assaying greater than 0.1 oz/st Au (3.43 g/t Au) were rerun using a gravimetric finish. A multi-element “Pathfinder” analysis package was used for Ag, As, Sb, Hg, Cu, Pb, and Zn analyses, although the analytical procedure is undocumented.

Legend Assay Laboratory (Legend)

No documentation exists for preparation protocols used by Legend. Gold was analyzed using a 30 g charge, FA digest and AA finish. Detection limit was 0.001 oz/st Au (0.031 g/t). Samples assaying greater than 0.1 oz/st Au (3.43 g/t Au) were rerun using a gravimetric finish.

ALS Global

Preparation protocols used by ALS Global included samples were crushed to 70% passing minus 2 mm, a 250 g split collected using a riffle splitter, and the split was pulverized to 85% passing -75 µm in a ring and puck mill. Gold was analyzed using a 30 g charge, FA digest and AA finish. Detection limit was 5 ppb. Samples assaying greater than 0.1 oz/st Au (3.43 g/t Au) were rerun using a gravimetric finish. Cold cyanide leach gold analyses (30 g) were also made on select samples. Ag, As, Cu, Pb, and Zn analyses were determined by nitric acid-aqua regia (AR) digest with an AA finish. Antimony analyses were determined using a hydrochloric acid-potassium chloride digestion and an AA finish. Mercury was analyzed using a nitric acid-hydrochloric acid digestion with an AA finish.

 

 

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Bondar Clegg (Bondar)

Preparation protocols used by Bondar included samples were crushed to 75% passing minus 2 mm, a 250 g split collected, and the split was pulverized to 95% passing -150 µm. Gold was analyzed using a 30 g charge, FA digest and AA finish. Detection limit was 5 ppb. Samples assaying greater than 0.1 oz/st Au (3.43 g/t Au) were rerun using a gravimetric finish. Mercury was analyzed using a cold vapor digestion with an AA finish. A six multi-element package (Ag, As, Cu, Pb, Sb, Zn) included an AR digest and AA finish. The 35 multi-element package included an AR digest with an inductively coupled plasma atomic emission spectrometry (ICP-AES) finish. Antimony analyses were determined using a hydrochloric acid-potassium chloride digestion and an AA finish.

 

8.5

Quality Assurance and Quality Control (QA/QC)

Barrick implemented a QA/QC program for its RC and diamond drill programs from 2004 to 2015 and digital results of the QA/QC program are incorporated in the digital database for the project.

Review of drillhole logs, sample submission sheets and notes on assay certificates from the Homestake drilling indicates that a QA/QC program was used for some of the sampling and assaying; however, the extent of the implementation of QA/QC and full detailed results of the program are not available in the digital database for the project.

It is not clear whether operators prior to Homestake implemented QA/QC for data quality assurance prior to 1992.

The results of the Barrick QA/QC program have been reviewed in detail by REI (2005), Waterton (EMG, 2017) and by Wood in 2020.

A description of the QA/QC programs and selected results for the Barrick and Homestake programs follows.

 

8.5.1

Barrick QA/QC Program

The Barrick QA/QC program evolved from analysis of check samples at a secondary laboratory to a more robust program including routine insertion of standard reference materials, coarse blanks, pulp duplicates and field duplicate samples with tolerances for standard reference materials and blank materials used to flag sample batches for re-assay prior to import into the digital database.

Table 8-4 shows the evolution of Barrick’s QA/QC program with the number of control samples of different types shown for each year, and the number of original sample assays analyzed per year. ALS Global also started re-assaying lab pulp duplicates in 2012.

 

 

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Table 8-4: Count and Description of QA/QC Samples by Year

 

Year   No. of
  Standards/Blanks 
  No. of Field  
  Duplicates
  No. of
  Duplicates  
  No. of Lab  Pulp  
  Duplicates

  No. of
  QA/QC

  samples  

  No. of
  Assays  
  Percentage  of  
  Assays

2004

  58   0   0   0   58   576   10.00

2005

  201   0   15   0   216   1,980   10.90

2006

  182   23   53   0   258   4,007   6.40

2007

  165   2   16   0   183   4,877   3.80

2008

  236   41   119   0   396   4,464   8.90

2009

  755   197   401   0   1,353   14,408   9.40

2010

  1,699   438   960   0   3,097   32,227   9.60

2011

  1,220   295   679   0   2,194   22,639   9.70

2012

  1,248   317   696   877   3,138   23,945   13.10

2013

  506   117   225   363   1,211   8,309   14.60

2015

  271   77   152   135   635   2,823   22.50

UNKN 

  21   0   0   0   21   1,900   1.10

Total

  6,562   1,506   3,316   1,375   12,760   122,155   Average: 10.8

Barrick inserted 3,445 standards of 25 different types with best values ranging between 0.214 g/t Au and 8.367 g/t Au between 2004 and 2015. Standards included commercially prepared oxide gold reference from OREAS and Rocklabs and internal oxide gold standards developed by Barrick. All standards were inserted under the guidance of the project geologist.

Barrick’s QA/QC guidelines stated that during the program re-runs were to be requested when the result exceeded ±3 standard deviations (3SD) of the expected value. Failed standards within non-mineralized intervals were reviewed and re-assayed at the discretion of the project geologist. A total of 99 samples (3%) were flagged as failed from 3,445 SRM samples. The weighted average bias of all standards is 1.15% and the biases of OREAS 54PA and BCH-OX-01, BCHOX-02 and BCH-03 standards which were the most commonly inserted standards range from 0.7% to 3.2%. Figure 8-1 presents the results of SRM OREAS 54PA which is one of the most commonly analyzed SRM. Eighty-six percent of samples were within 2 standard deviations (2SD), and 96% within 3SD of the expected value (Table 8-5).

 

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Figure 8-1: Control Chart for Standard OREAS 54PA

(Source: Wood, 2021)

 

 

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Table 8-5: SRM Performance

 

 Standard ID   Sample  
Count
  Au Grade     % Within  
2SD
  % Within  
3SD
  Bias (%)     Relative Standard  
Deviation (%)

BCH-OX-01

  338   0.214   91   97   1.10   5.30

BCH-OX-02

  203   0.338   77   93   3.20   4.00

BCH-OX-03

  541   2.260   85   98   3.20   4.40

BCH-OX-04

  204   6.450   96   100   -0.30   2.30

BCH-OX-06

  108   0.283   96   99   2.00   4.20

OREAS 2PD

  201   0.885   95   100   -1.70   3.00

OREAS 50PB

  29   0.841   83   90   0.70   6.00

OREAS 52c

  190   0.346   99   99   3.60   6.20

OREAS 52PB

  199   0.307   93   99   4.60   3.90

OREAS 53PB

  41   0.623   90   95   0.00   5.70

OREAS 54PA

  429   2.900   98   99   0.70   3.40

OREAS 6PC

  158   1.520   99   100   0.50   3.10

OxD57

  61   0.413   95   98   -0.80   3.00

OxG38

  86   1.031   94   99   -0.30   4.00

OxH29

  33   1.298   73   88   -1.80   5.10

OxH52

  37   1.291   89   95   -1.10   3.70

OxH55

  124   1.282   92   96   1.50   3.60

OxI23

  81   1.844   78   91   -1.20   4.50

OxK48

  31   3.557   58   87   -1.10   3.30

SF12

  78   0.819   88   91   -4.60   11.70

SG14

  71   0.989   96   100   0.90   3.70

SJ10

  36   2.643   72   97   -2.30   3.30

SK11

  51   4.823   82   92   -1.20   3.50

SK21

  77   4.048   70   94   -0.20   3.90

SN16

  38   8.367   53   76   -2.40   6.70

Total

  3,445   86   96   4.50

A total of 3,116 blanks were inserted in the sample stream by Barrick with 51 samples (or 1.6%) plotting above the 0.025 g/t Au. Material used for blank samples was sourced from the Devonian Devils Gate Limestone.

Ninety-eight sample pulps, representing approximately 4% of existing sample pulps from drilling at East Archimedes by Barrick were sent to BSI-Inspectorate Laboratory in Reno, Nevada for check assays. Original assays were performed by ALS Global. Six certified standard samples from OREAS of Australia were also randomly introduced with the pulps. Original ALS Global assays indicated approximately 70% of the 98 pulps consisted of mineralized material, the remainder was classified as waste.

Results from the BSI-Inspectorate check assays have a mean grade slightly lower than the ALS results for the same samples and the relative bias increases slightly with increasing grade (REI, 2005) (Figure 8-2). This relative bias confirms the small positive bias of approximately 1-3% evident in the analyses of the SRM materials analyzed at ALS.

 

 

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Figure 8-2: ALS Global (Chemex) Pulps Checked at Inspectorate

(Source: Newman and Mahoney, 2008)

Field duplicates were added to the QA/QC protocol as part of the 1 in 300 QA/QC samples. For core duplicates the other half of core was taken and analyzed. For RC duplicates, a secondary sample was taken at the splitter on the drill rig. Barrick used a sample ID that was consecutive to the original sample to identify the duplicate sample.

A total of 1,037 field duplicates (230 core and 807 RC) with mean values greater than 0.1 g/t Au were analyzed and 73.4% of the samples plot within ±15% of half the relative difference (Figure 8-3).

 

 

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Figure 8-3: Mean Versus Half Relative Difference for Field Duplicates

(Source: RHMC, 2017)

A plot of lab pulp duplicate samples on a scatter graph (Figure 8-4) indicates good repeatability for the pulp duplicates with 90% plotting within 5% half the relative difference of the original analysis. All samples were assayed by ALS Global between 2012 and 2015.

Pulp duplicates plotted on mean versus half relative difference graphs indicates over 90% of samples plot within 10% of half the relative difference (Figure 8-5). All values greater than 10% of half the relative distance are very low grade (<0.06 g/t).

 

 

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Figure 8-4: Scatter Plot of all Lab Duplicates

(Source: RHMC, 2017)

 

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Figure 8-5: Mean Versus Half Relative Difference for Pulp Duplicates

(Source: RHMC, 2017)

 

 

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8.6

Historical Databases

In early 2004 Barrick prepared the Ruby Hill drill hole database for use in resource modeling efforts. A systematic program was instituted to combine the various disparate databases into an accurate database. The program produced an accurate Ruby Hill drill hole database stored in Microsoft® Access.

More recent Barrick RC and core logging was performed by a company geologist using a logging template. All geologic, structural, geotechnical, metallurgical, and density measurements, taken at 50 ft intervals, were recorded on the template and entered into an acQuire database. It is unknown when Barrick migrated the database from Microsoft Access to acQuire. The acQuire database was maintained by the Barrick Gold Exploration Inc., office in Elko, Nevada.

In April 2016 RHMC contracted the Maxwell Geoservices of Vancouver, Canada to migrate the Ruby Hill acQuire database to Maxwell’s DataShed software. Original digital assay results were directly imported, csv files were generated from pdf or paper versions of each assay lot and then imported. As of the date of this Report, information that has been loaded into DataShed includes collar, downhole survey, assay, lithological and multi-element data.

The database was maintained on the RHMC server in Reno, and nightly back-ups were made at a secure off-site location.

 

8.7

Historical Sample Security

Sample handling procedures and chain of custody for drilling prior to the 2002 closure of the Ruby Hill operation are not well documented. It is assumed samples from earlier drilling were in the custody of the drill contractor, Homestake geologists, or employees of the various laboratories that prepared and assayed the drilling samples. In 2005 REI (2005) notes that examination of remaining historical core was in good order in core boxes with drill run blocks in place and sample intervals clearly marked and was of the opinion that drill core in general was probably well handled, transported, and stored during the course of drilling.

The security procedures and chain of custody employed for drill samples is poorly documented. Newman and Mahoney (2008) report that no officer or employee of the company prepared drill samples, except that core samples were split by a company employee before sending to the assay lab, and a minor number of holes (14) were prepared and assayed at the company’s internal lab. RC drill cutting samples were picked up from the drill rig by the assay lab’s courier service. Core samples were first split in half by company staff, one half was archived, and the other half picked up by the lab courier service. Laboratory chain of custody was typical to commercial labs in Nevada at the time of activity according to Newman and Mahoney (2008).

All remaining pulps were securely stored in locked shipping containers on site. Remaining core is also stacked on pallets and stored on site with more than half of the core covered. Numerous uncovered core boxes have been partially to completely destroyed due to weathering.

 

8.8

Comments on Historic Ruby Hill Data

The Ruby Hill mineral resource dataset has been acquired over many years during which time best practice for drilling, sampling, assaying, sample and data security practices have evolved significantly. The data acquired by Barrick from 2003 to 2015 has been acquired from RC and diamond drill core holes using industry standard practices for surveying, logging, sampling, sample preparation, assaying and assay QA/QC. Review of QC data indicates that the accuracy, sampling and analytical precision and reproducibility of the Barrick assaying for gold and silver is of good standard. Database compilation efforts by Barrick beginning in 2004, and by RHMC in 2016 included direct import of digital files wherever possible to limit the possibility of data transcription issues. The Barrick data has been used to provide data quality

 

 

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assurance for the Homestake data and data acquired by operators before Homestake and is discussed in Section 9 on data verification.

 

8.9

i-80 Sample Preparation, Laboratory Analysis, Security, and Quality Control Procedures

 

8.9.1

i-80 Sample Preparation Procedures

RC samples were collected on 5-foot intervals using an adjustable cyclone splitter. The target weight for each interval was approximately 8 kg of sample caught directly in a sample bag placed below the splitter. Each sample was tied and placed in a sample bin. Sample bins were transported from the drill site to the core yard area by an i-80 employee and then from there samples were picked up by either ALS Minerals, American Assay Laboratories, or Paragon Geochemical.

Core samples were cut by i-80 technicians using core saws at i-80’s core processing facility at Lone Tree. Technicians prepare sample bags for the sample intervals specified in the logging geologist’s cut sheet. Core sample intervals may range from one foot to ten feet. The technician cuts the core in half, places one half into the sample bag and returns one half to the core box. When the entire sample interval is split, the technician ties the bag and places it into a sample bin. When splitting is complete for the hole, the sample bin is picked up by a driver for ALS Minerals, American Assay Laboratories, or Paragon Geochemical and delivered to the respective lab.

 

8.9.1.1

i-80 Laboratory Analysis Procedures

Both Core and RC samples were submitted to either ALS Minerals, American Assay Laboratories, or Paragon Geochemical, all located in Sparks, Nevada. All labs are independent of i-80. Paragon is certified under ISO/IEC 17025:2017. ALS Minerals and American Assay Laboratories are ISO 9001 and 17025:2017 certified. Samples were dried, weighed, screened, crushed to 70% passing 10 mesh, split to 250g with a riffle splitter, then pulverized to 85% passing 200 mesh. Samples submitted through Paragon Geochemical were analysed with a 50 element suite (code 50AR-MS) using 0.5g aqua regia digestion with ICP-MS finish. Samples submitted through ALS Minerals (4977 Energy Way, Reno, NV 89502 or 1345 Water St. Elko, NV 89801) were analysed with a 35 element suite (code ME-ICP41) using 0.5g 4-acid digestion with ICP-AES finish. The ALS ICP-AES facility is located at 2103 Dollarton Hwy, North Vancouver, BC, Canada. Samples submitted through American Assay Laboratories (1506 Glendale Ave, Sparks, NV 89431) were only analysed for gold with pulps sent to ALS Minerals for multi-element analysis. Each sample sent to Paragon Geochemical (1555 Industrial Way, Sparks, NV 89431) was analysed for Au using 30g fire assay, aqua regia digestion with AAS finish (code Au-AA30) with detection range 0.005 to 5 ppm Au. Samples with Au result greater than 5 ppm Au were analysed using 30g fire assay with gravimetric finish (code Au-GR30), detection range 0.14 to 10,000 ppm Au. Each sample sent to ALS Minerals was analysed for Au using 30g fire assay, aqua regia digestion with AAS finish (code Au-AA23), with detection range 0.005 to 10 ppm Au. Samples with Au result greater than 10 ppm Au were analysed using 30g fire assay with gravimetric finish (code Au-GRAV21), detection range 0.05 to 10,000 ppm Au. Each sample sent to American Assay Laboratories was analysed for Au using 30g fire assay, aqua regia digestion with AAS finish (code FA-PB30-ICP), with detection range 0.003 to 10 ppm Au. Samples with Au result greater than 10 ppm Au were analysed using 30g fire assay with gravimetric finish (code GRAVAu30), detection range 0.103 to 10,000 ppm Au.

 

8.9.1.2

i-80 Security

Core is transported from the drill to the Ruby Hill core shed, a rented facility near Ruby Hill which is fenced and locked. It is stored in the core yard until it can be logged. Once logging is complete, core is transported to Lone Tree by i-80 personnel for splitting. Lone Tree is fenced and access is controlled with ID key cards. Once splitting is complete, lab drivers pick up the samples, maintaining chain of custody.

 

 

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RC samples are stored at the drill site under supervision of the drillers until the hole is complete. It is then transported to the Ruby Hill core yard and stored for a short time until it can be picked up by the lab driver, maintaining chain of custody.

Split core retained in the original box is stored at the Lone Tree core yard. Pulps are initially returned to the Lone Tree lab for check analysis before being stored at Lone Tree.

 

8.9.1.3

i-80 QA/QC

Several standards with various characteristics (high, medium and low grade, oxide and refractory) are in use at Ruby Hill. The provenance of most of the standard material is Carlin type deposits. Standards are inserted at a rate of approximately 5% with the goal of matching the standard grade to the nearby grade of the rock. Blanks are inserted at a rate of approximately 5%. Duplicates are made at approximately 5% (2.5% sample duplicates generated during the core splitting phase or at the drill rig with a Y-splitter in the case of RC samples, 2.5% prep dups generated after the pulverizing phase of sample prep by the assay laboratory.) QAQC is inserted at the discretion of the geologist performing the logging on core material. QAQC for RC samples is pre-determined by the sample sheet which is made prior to drilling the hole.

i-80 uses crushed marble for blanks, and for standards purchases certified reference materials from OREAS, a reputable supplier of reference materials for the mining industry.

Check samples are conducted at the Lone Tree assay lab facility once pulps are returned to the site. All samples with values >1 ppm Au have a check assay performed, with a target check rate of about 10% of total sample stream.

 

8.10

i-80 Standards and Blanks

i-80 has used 40 different commercially prepared standard reference materials and blanks in its QA/QC program for the Ruby Hill drilling. The QA/QC data through Dec 31, 2022 contains a total of 3,764 gold assays. Selected results for i-80’s QA/QC program are shown in Table 8-6.

Table 8-6: Selected i-80 Blank and Standard Reference Results

 

Std ID   Blank   Blank
  Marble  Chip  

  CDN-GS-  

1Z

  G919-10     KIP-19  

  OREAS-  

273

  OREAS-  

277

Count

  192   876   99   119   124   330   208

Mean

  0.017   0.006   1.152   7.542   2.489   9.925   3.356

Standard Dev

  0.147   0.026   0.058   0.366   0.313   1.629   1.109

Min

  0.002   0.002   0.950   5.900   2.040   0.306   0.009

Q25

  0.003   0.003   1.120   7.510   2.420   10.000   3.350

Median

  0.003   0.003   1.160   7.600   2.460   10.000   3.420

Q75

  0.005   0.005   1.190   7.695   2.490   10.700   3.480

Max

  2.031   0.611   1.250   8.020   4.930   13.000   10.800

No Rejected

  2   0   0   0   0   38   20

% Pass

  99%   100%   100%   100%   100%   88%   90%

 

8.11

i-80 Duplicate Assays

The database contains 2,145 lab duplicates. ALS assayed 878 and AAL 1,267. The results from both labs are displayed in Figure 8-6. Both labs performed well with regression line slopes of unity and correlation coefficients of 0.999.

 

 

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8.12

QP Opinion

It is the opinion of the QP that sample preparation, security, and analytical procedures meet industry standard practices.

 

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Figure 8-6: i-80 Lab Duplicates

(Source: Practical Mining, 2023)

The database contains 596 prep duplicates. ALS assayed 392 and AAL 244. The results from both labs are displayed in Figure 8-7. Both labs performed well with regression line slopes of unity and correlation coefficients of 0.996.

 

 

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Figure 8-7: i-80 Prep Duplicates

(Source: Practical Mining, 2023)

 

 

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9.

 DATA VERIFICATION

 

9.1

Historical Data Review

A detailed summary of drill hole data analysis undertaken historically at Ruby Hill is provided by Wood in their Technical Report dated 2021. Property tenure has varied over the years, and data analysis has been performed on correspondingly varying data sets, generally covering areas beyond the focus of the current analysis. In summary, Barrick validated 18 holes in support of its East Archimedes Project feasibility study in 2004 (Newman and Mahoney, 2008). In 2005, REI performed an audit of the East Archimedes Project and compared assay values in the estimation database with laboratory certificates for 12 drillholes. They concluded the assay database was valid for mineral resource estimation. In 2011, Barrick updated their Ruby Hill block model to include new drilling from the 426 and Mineral Point areas. The database was checked for overlapping and missing intervals and for excessive azimuth and inclination deviations. Errors in the lithology table (typos and inconsistent naming conventions) were identified and corrected. Barrick deemed the database to be in good condition (Barrick, 2013). In 2016, RHMC performed a detailed data review after migrating the data from acQuire to Datashed. Multiple errors and inconsistencies were identified and corrected.

Very few core twins of RC holes have been drilled on the larger Ruby Hill property. Homestake drilled four twins in the Mineral Point area, and Barrick twinned two holes in the Mineral Point area. Homestake concluded two of its four RC holes were contaminated, while Barrick attributed grade differences to lithology and structural characteristics of the rocks. RHMC agreed with Barrick’s analysis. RHMC also performed statistical analysis of drilling by type and operator (Barrick vs Homestake and RC vs Core) and noted differences in grade but also found that holes of differing type within proximity of each other (200 feet) compare reasonably well, indicating reproducible assay by type and company (Wood, 2021).

 

9.2

Wood Data Verification 2021

Wood completed detailed data verification for their 2021 Mineral Resource Estimate covering the Mineral Point Trend and Archimedes area deposits. Wood analyzed downhole contamination using Quantile-Quantile plots to compare the grade distributions of the core samples and the RC samples. Raw results indicated a slight high bias in the core. The bias was nearly eliminated when filtered to data within the mineralized domains.

Wood checked the digital database against original hardcopy records by selecting 100 holes for collar and assay data audits and 50 holes for downhole survey and lithology audits. The audit focused on holes drilled by Barrick because hardcopy records for Homestake holes tend to be incomplete. Wood observed no discrepancies in the assay data. Original collar data was not available for some holes, but locations correspond well with topography; Wood recommended attempting to recover lost survey reports. Downhole surveys were deemed reasonable. Geology corresponded well with paper logs. Homestake data was supported by comparing Homestake holes with nearby Barrick holes, which demonstrated grade and thickness compare well between drilling campaigns.

Gold grades were also analyzed visually, and Wood identified four holes with mineralized intercepts that do not correlate well with adjacent data. Those four intervals were excluded from the mineral resource estimation. Wood concluded the database was suitable for use in the Mineral Resource Estimate.

 

9.3

Practical Mining Data Verification 2023

In 2023, Practical Mining updated the resource estimate in the Ruby Deeps and 426 areas to include new drilling. 102 drillholes were flagged for use in the estimate, and 15 holes (representing about 15% of the data set) were chosen for detailed review. The holes selected for review were chosen to represent the area

 

 

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of interest in an even spatial distribution as well as represent different operators over time (i-80, Barrick and Homestake.) Table 9-1 summarizes holes drilled in the 426 and Ruby Deeps zones by type and operator.

Table 9-1: Drill Holes in 426 and Ruby Deeps Zones

 

Company    Core    RC    Validated

i-80

   31         4 core

Barrick

   34    18    2 core, 2 RC

Homestake

   12    7    5 core, 2 RC

Totals

   77    25    15

Practical Mining requested original hardcopy data records for the selected holes including collar location surveys, downhole deviation surveys, geology logs, and assay certificates. Collar survey records were available for only four holes, two drilled by Barrick and two drilled by Homestake. Another Barrick hole, BRH2C, was located a significant distance from the planned location recorded on the geology log. i-80 geologists were subsequently able to recover survey data confirming the location of BRH2C. Practical Mining recommends continuing to recover missing survey records, as well as systematically archiving data in digital and hard copy formats as new holes are drilled. Practical Mining viewed holes in Vulcan to confirm collars coincide with topography. Hole HC1399 lies about 32.5 feet above the original topography, but it is located in the topsoil stockpile area and was apparently drilled during construction of the stockpile.

All holes used in the estimation have downhole deviation surveys, although some hard copy records were not archived. Downhole survey records were available for all of the selected i-80 and Barrick holes, and three of the Homestake holes. All of the selected records match the database, except the most recently drilled i-80 hole, iRH22-57, which had an intermediate version of the survey taken before the final 1011 feet were drilled. Updating the final survey will affect the location of a mineralized interval in the Ruby Deeps zone, but should not have a significant effect on the mineral estimation due to its location within the modeled zone and the small magnitude of the change relative to wide drillhole spacing. All hole traces were viewed in Vulcan and no excessive deviation was noted.

Geology logs were available for all requested holes. Logs match the database quite well. Two Homestake holes had logs digitized for the core tail but not the RC pre-collars, and one Homestake hole had not been entered in the digital database at all. Practical Mining recommends digitizing the data for consistency and to make the database more comprehensive, although it is unlikely to have an effect on the geology model since drill spacing is close and the geology data is interpolated between adjacent holes. Practical Mining viewed all drillhole traces coded by lithology in Vulcan and observed that the drill data coincides very well with i-80’s lithological and structural models.

Assay certificates were unavailable for one requested Barrick hole and two Homestake holes. Certificates for 12 holes were compared with the database and only one mismatch was identified, a minor error where the preliminary value was exported instead of the final value. Practical Mining viewed all drillhole traces coded by assay grades in Vulcan and noted that grade and thickness correlate well between adjacent holes and along geological contacts. Table 9-2 summarizes the number of holes reviewed per data field.

Table 9-2: Drillhole Data Fields Reviewed

 

     

Collar

Surveys

  

Downhole

Surveys

  

Geology

Logs

  

Assay

Certificates

Holes Reviewed

   8    11    15    12

Percent of Population

   7.8%    10.8%    14.7%    11.8%

Practical Mining recommends continuing to recover collar survey records and archiving all drilling records properly. Practical Mining concludes the database is suitable for use in the mineral resource estimation.

 

 

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10.

 MINERAL PROCESSING AND METALLURGICAL TESTING

 

10.1

Archimedes Underground

This section summarizes all the relevant test work performed on the Archimedes Underground project. The Archimedes Underground project encompasses several deposits and mineralization types hosting both precious and base metals. Historical production dates to 1998, primarily under Homestake Mining and Barrick Gold, with intermittent operations up to the current date. Characteristics of each deposit, historical production and metallurgical interpretation for the Archimedes Underground deposits are described in this section, based on data provided by Ruby Hill Mining LLC. Generally, metallurgical test work confirms the amenability of oxide mineralization to heap leaching for precious metals extraction. Tests on refractory samples support gold extraction via pressure oxidation.

 

10.1.1

Refractory Testing Programs

A series of testing programs have been completed on refractory samples from 426, Blackjack and Ruby Deeps zones. These programs are summarized in Table 10-1.

Table 10-1: Ruby Hill Project Refractory Testing Programs

 

No.    Document Title    Deposit    Technical Content    Date
1    Barrick Technology Center    426 Zone    Refractory roasting, pressure oxidation, leach tests    2008
2    G&T Metallurgy    426 Zone    Refractory flotation, leach tests    2008
3    Barrick Technology Center    426 Zone    Pressure oxidation, CaTS and standard CIL leach tests    2011
4   

FLSmidth Minerals Testing and Research

Center

   426 and Ruby
Deeps
   Refractory pressure oxidation and roasting    2024

The laboratories used for testing have the following accreditations:

 

   

Kappes Cassiday and Associates, no certifications listed on website.

 

   

Barrick Technology Center, a part of Barrick Gold Inc. at the time. No certifications provided.

 

   

G&T Metallurgy, now part of ALS Metallurgy but no known accreditations at the time programs were completed.

 

10.1.1.1

January 2008 Barrick Technology Centre Program

This report summarizes the testing of three composites consisting of various blends of 426 Zone samples with typical Barrick Goldstrike roaster feed material. Table 10-2 shows the results. In the report, an adverse trend is noted between gold recovery and increasing arsenic concentration. In the table, BTR stands for Bench Top Roaster and BTALK for Bench Top Alkaline autoclave.

 

 

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Table 10-2: January 2008 426 Zone Barrick Technology Centre Test Results Summary

 

Sample    Head Grade(opt)    Leach Extraction (Au %)
   Au    As    BTR    BTALK    Pilot Plant

BGMI Roaster Feed Baseline—Pilot Plant

   0.252    861              87.0

BGMI Roaster Feed Baseline—BTR

   0.262    868    89.6          

RH 426 Composite 1—Pilot Plant

   0.350    9,808              51.0

RH 426 Composite 1 – BTR & BTalk

   0.368    10,211    81.9    90.1     

RH 426 Composite 2

   0.174    2,194    85.3    91.0     

RH 426 Composite 3

   0.125    1,208    88.1    89.6     

RH 426 Composite 2&3 (1:1 blend)

   0.156    1,787              86.1

RH 426 Composite 2&3 (1:1 blend)

   0.152    1,787    83.5    91.3     

Blend 3.6% Comp. 1 Baseline

   0.260    993              87.3

Blend 10% Comp. 1 in Baseline

   0.253    1,751              82.9

Blend 20% Comp. 1 in Baseline

   0.261    1,735              80.6

 

10.1.1.2

February 2008 Barrick Technology Centre Program

The report summarizes a program that investigated recovery of arsenic to a pre-flotation concentrate, while minimizing gold losses. Up to 80% of the arsenic was recovered in the pre-flotation concentrate, with gold losses of approximately 5.0%, with a concentrate mass recovery of 2.7%. This was achieved in a single-stage cleaning step using strongly alkaline conditions. Subsequent gold recovery to a sulfide concentrate was only 66%, with a mass recovery of 32%. Low selectivity and high mass concentrate mass recovery indicated poor gold liberation.

 

10.1.1.3

December 2008 G&T Metallurgical Services Program

The test program was developed to investigate the potential for producing a pre-flotation concentrate with high arsenic and low gold recoveries to this stream. Arsenic occurs mainly as realgar in the 426 Zone samples.

The test flowsheet included, after the arsenic pre-float, a bulk sulfide rougher flotation step. The objective was to recover the sulfide mineralization and gold into a flotation concentrate that could then be further processed to recover the gold.

Gold extraction from the whole ore and flotation product streams was also investigated using cyanidation bottle roll techniques.

Rougher flotation tests failed to produce greater than 50% As recovery into a pre-float concentrate. A single test with an arsenic feed content of 2.4%, achieved about 82% As recovery to the pre-float concentrate.

Gold recovery, to a bulk sulfide rougher concentrate, carried out on the pre-flotation tailing was also limited to about 50%. To achieve this result, about 30 percent of the feed mass needs to be recovered to the bulk sulfide rougher concentrate.

Cyanidation bottle roll tests were carried out on whole ore and flotation products from one sample. Under a variety of test conditions, the best 48 hour gold extraction from any stream was about 30%.

 

10.1.1.4

November 2011 Barrick Technology Centre Program

Sixteen refractory and two oxide samples from the 426 Zone were tested at The Barrick Technology Centre. For the refractory samples, CIL recoveries following alkaline pressure oxidation gave recoveries ranging from 77% to 93%, with an average recovery of 88%. Direct CIL tests on the two oxide samples gave recoveries between 92% and 96%, averaging 94%. The sulfide sulfur (S2-) content of these oxide samples was <0.05%. Table 10-3 shows results of BTALK tests followed by both Calcium Thiosulfate (CaTS) leaching and standard cyanide CIL leaching. On average, CaTS leaching produced comparable recoveries to standard CIL.

 

 

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Table 10-3: November 2011 426 Zone Barrick Technology Centre Test Results Summary

 

Sample    Head Grade    Recovery (Au %)
   Au (opt)    CO3 (%)    CORG (%)    S2- (%)    As (%)    CaTS    CIL

Average

   0.158    26.51    0.05    1.59    0.39    87.61    88.02

Maximum

   0.533    45.40    0.09    3.06    2.58    94.20    93.10

Minimum

   0.026    1.50    0.02    0.40    0.03    70.90    76.50

 

10.1.1.5

2024 FLS Program

This program included thirteen samples (five from 426 Zone and eight from Ruby Deeps) for metallurgical testing included cyanide leach shake/Preg Rob testing, pressure oxidation by benchtop autoclave, calcination by benchtop roast testing, cyanidation testing of POx and roaster calcines, and benchtop flotation scoping along with mineralogical testing. The samples and major assays are shown in Table 10-4.

Table 10-4: 2024 FLSmidth Program Assays Summary

 

 Sample Description   

Assays

 

   Au
(opt)
   Ag
(opt)
  

STOT

(%)

  

S=

(%)

  

SO4

(%)

  

So

(%)

  

CTOT

(%)

  

CORG

(%)

  

As

(%)

   Hg
(ppm)

426 Zone Central Sample

   0.185    0.065    0.87    0.51    0.35    0.00    0.06    5.61    1.20    57.19

426 Zone East Sample

   0.277    0.063    2.49    2.13    0.35    0.02    0.11    1.47    0.63    43.88

426 Zone High Grade Sample

   0.653    0.129    4.01    3.42    0.57    0.02    0.12    3.26    4.77    91.75

426 Zone Low Grade Sample

   0.159    0.056    2.43    2.15    0.28    0.00    0.11    2.70    0.27    27.65

426 Zone Composite Sample

   0.312    0.119    2.95    2.54    0.39    0.02    0.13    3.20    0.70    56.52

Ruby Deeps North Sample

   0.304    0.137    1.42    1.19    0.23    0.00    0.19    3.37    0.11    9.05

Ruby Deeps Mid Sample

   0.212    0.160    1.79    1.40    0.39    0.00    0.16    1.43    0.25    9.99

Ruby Deeps South Sample

   0.239    0.115    2.03    1.53    0.50    0.00    0.16    0.47    0.29    36.48

Ruby Deeps Intrusive Sample

   0.248    0.131    3.51    3.32    0.19    0.00    0.06    1.65    0.66    17.04

Ruby Deeps Dunderberg Shale Sample

   0.318    0.101    4.02    3.71    0.31    0.00    0.07    1.36    1.08    17.52

Ruby Deeps High Grade Sample

   0.524    0.131    2.86    2.13    0.73    0.00    0.14    1.87    1.20    24.40

Ruby Deeps Low Grade Sample

   0.169    0.133    2.25    1.54    0.68    0.03    0.20    1.93    1.25    16.08

Ruby Deeps Composite Sample

   0.439    0.117    3.02    2.18    0.75    0.09    0.15    3.05    0.55    40.58

Average gold grades are similar for the 426 and Ruby Deeps samples are comparable at 0.317 and 0.307 opt respectively. Sulfide sulphur grades are comparable as well at 2.15% and 2.13% respectively. Arsenic and mercury grades are markedly higher in the 426 samples compared to Ruby Deeps, averaging 1.52% As and 55.4 ppm Hg for 426 Zone and 0.67% As and 21.39 ppm for the Ruby Deeps samples.

The two composite samples were subjected to QEMSCAN analysis to characterize their mineralogy. Major rock forming minerals in the samples include quartz, potassium feldspar, and calcite. Samples also contain appreciable amounts of kaolinite (5.9% to 7.3%). Major sulfide minerals include pyrite at 4.8% in the 426 Zone composite and 3.8% in the Ruby Deeps composite. The next most abundant sulfide mineral is realgar at 0.76% and 0.60% in the 426 composite and in the Ruby Deeps composite respectively. Both samples have relatively low concentrations of arsenopyrite at 0.05% and 0.16% in 426 and Ruby Deeps composites respectively. The amounts of arsenopyrite are significantly lower than expected based on the arsenic concentrations, indicating that the pyrite carries significant amounts of arsenic (arsenian pyrite).

 

 

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Preg robbing tests showed a general correlation with organic carbon content as shown in Figure 10-1. Preg robbing occurred in both 426 and Ruby Deeps samples but was generally higher in the 426 samples.

 

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Figure 10-1: 2024 FLSmidth Program Preg-Robbing as a Function of Organic Carbon Concentration

(Source: FLSmidth, 2024)

Baseline CIL tests were done on all samples at two different grinds. Test conditions included:

 

   

Grind sizes of k80 = 100 mesh and 200 mesh.

 

   

Slurry density = 35% solids.

 

   

Cyanide concentration of 1.0 g/L, maintained at 0.5 g/L.

 

   

Carbon concentration of 20 g/L.

 

   

Test duration = 48 hours.

The results showed the refractory nature of the samples with overall average gold recoveries of 31.2% at the 200 mesh grind and 30.8% at the 100 mesh grind. The Ruby Deeps Dunderberg sample had 0% gold recovery in both baseline tests. Several Ruby Deeps samples had recoveries below 10%. Overall average CIL baseline gold recovery was 31.2%; 55% for the 426 samples and 16.3% for the Ruby Deeps samples.

All samples were subjected to BTAC testing with three different sets of conditions outlined in Table 10-5.

 

 

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Table 10-5: FLSmidth Program BTAC Conditions Summary

 

Operating Conditions    A    B    C

Acidulation

   Yes    No    No

POx Condition

   Alkaline    Alkaline    Acid

Trona Addition

   None    None    Stoichiometric

Temperature (oF)

   390    390    437

O2 Overpressure (psig)

   100    100    100

Pulp Density (% solids)

   30    30    30

Grind Size (k80, mesh)

   200    200    200

Retention Time (minutes)

   45    75    45

BTAC products were then subjected to CIL testing, following the same conditions for the baseline tests with the tests run for 24 hours instead of 48 hours.

The samples were also subjected to batch roasting tests. Tests were run for 90 minutes at 932oC with an atmosphere of 40% oxygen. As with the BTAC tests, samples were ground to k80 of 200 mesh.

Results of the BTAC and batch roasting CIL tests are shown in Table 10-6 along with S= oxidation. Baseline CIL test results are included for comparison.

Table 10-6: FLSmidth Program BTAC and Roasting CIL Recovery Summary

 

 Sample Description    Baseline
CIL
  

BTAC Condition

A

  

BTAC Condition

B

  

BTAC Condition

C

   Roasting
   Recovery
(% Au)
   Oxidation
(% S=)
   Recovery
(% Au)
   Oxidation
(%S=)
   Recovery
(% Au)
   Oxidation
(%S=)
   Recovery
(% Au)
   Oxidation
(%S=)
   Recovery
(% Au)

426 Zone Central Sample

   62.0    81    90.2    56    91.0    98    96.4    83    78.0

426 Zone East Sample

   55.0    58    87.8    74    89.2    100    95.0    80    64.0

426 Zone High Grade Sample

   38.0    40    62.7    51    70.0    99    97.7    80    39.0

426 Zone Low Grade Sample

   50.0    63    89.5    76    91.4    100    97.8    89    67.0

426 Zone Composite Sample

   70.0    67    92.9    68    93.4    98    97.0    94    74.0

Ruby Deeps North Sample

   4.0    50    47.9    60    56.0    47    48.9    90    80.0

Ruby Deeps Mid Sample

   20.0    54    69.3    69    80.0    99    95.7    91    80.0

Ruby Deeps South Sample

   53.0    35    76.3    55    74.8    98    92.9    89    83.0

Ruby Deeps Intrusive Sample

   3.0    18    30.4    32    44.4    29    38.8    76    47.0

Ruby Deeps Dunderberg Shale

Sample

   0.0    44    60.6    58    75.0    99    98.1    92    69.0

Ruby Deeps High Grade Sample

   4.0    57    74.4    69    78.5    98    97.2    95    83.0

Ruby Deeps Low Grade Sample

   9.0    48    57.5    57    67.1    79    85.4    94    75.0

Ruby Deeps Composite Sample

   37.0    47    75.5    61    82.2    56    74.1    99    76.0

 

 

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The results show that:

 

   

BTAC condition C produced the highest overall average gold recovery at 85.8%, followed by BTAC condition B at 76.4%. Roasting and BTAC condition C produced comparable gold recoveries at 70.4%.

 

   

Roasting had the highest overall average S= oxidation at 88.6% but this did not result in the highest gold recoveries.

 

   

Ruby Deeps samples are more refractory than the 426 samples with an average gold CIL recovery of 78.9% compared to 96.8% under BTAC condition C.

 

   

Overall Ruby Deeps sample recoveries are lower than 426 samples with all oxidation conditions used.

A series of flotation tests were conducted on the two composite samples using conditions developed in the December 2008 G&T metallurgical program. Two flotation tests were performed on each sample, one at a grind k80 of 200 mesh, and a second test at a k80 of 270 mesh with a lower pH. Both flotation tests on both samples achieved very low (less than 40%) recovery of Au. For the flotation test on the 426 Comp at a finer grind, a lower pH could not be achieved due to very high carbonate content.

It was noted in all four flotation tests that the pre-float contained a significant amount of the Au that floated. This suggests that some of the Au in these samples was present in a form that is self-floating, possibly hosted in the arsenic minerals. Leaching of flotation tailings was not expected to yield significant additional gold recovery and were not performed.

 

10.1.2

Deleterious Elements

A wide range of analyses were carried out on the samples used in the metallurgical testing programs included in this section. Deleterious elements were identified that are common to deposits in this part of Nevada. Deleterious elements content in the oxide samples is low, while sulfide samples are characterized by high levels of sulfide sulfur, arsenic, and mercury. Processing of Ruby Hill sulfide mineralization through the Twin Creeks autoclave at the Nevada Gold Mines Turquoise Ridge Complex initially and the i80 Lone Tree facility in 2028 will ensure removal and capture of these deleterious elements.

 

10.1.2.1

Arsenic and Mercury

The KCA January 2009 report conducted investigations into arsenic and mercury deportment. Although a note was added stating that as multi-acid digestion was specified, the values for arsenic and mercury may be biased low due to partial volatilization upon digestion.

The arsenic contents of the refractory samples were variable up to 2.6% and averaged 0.4%. One of the two oxide samples had a relatively high arsenic content of 0.43%. The 2024 FLSmidth program confirmed the presence of arsenic as arsenopyrite and arsenian pyrite in appreciable concentrations. Processing of arsenical refractory production through either pressure oxidation or roasting results in the capture and sequestration of arsenic in a stable form suitable for tailings disposal.

Mercury contents in the low and high-grade oxide composites from the 426 zone were moderate at 5.7 ppm and 9.6 ppm respectively.

The KCA February 2014 report analyzed six samples for mercury, they were reported as being between 2 and 10 ppm. All 16 refractory samples documented in the BTC November 2011 report had levels of less than 10 ppm Hg. Mercury concentrations at this level require the inclusion of mercury retorting in electrowinning and gold smelting areas of process facilities and mercury capture equipment on carbon reactivation kilns.

The 2024 FLSmidth program confirmed the presence of significant concentrations of mercury.

 

 

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10.1.2.2

Sulfur and Carbon

The 16 refractory and two oxide samples documented in the BTC November 2011 report were analyzed for total carbon and sulfur. Speciation for organic and inorganic carbon and speciation for sulfide and sulphate sulfur was included. Results are summarized in Table 10-7.

Table 10-7: November 2011 426 Zone Barrick Technology Centre Refractory Sample Assays Summary

 

Sample    Sample Assays (%)
   CTotal    CInorg    CO3    CORG    ST    SO42-    S2-

Average

   5.18    5.13    25.64    0.06    1.64    0.10    1.55

Maximum

   9.17    9.08    45.40    0.09    3.20    0.24    3.06

Minimum

   0.39    0.30    1.50    0.02    0.44    0.04    0.40

The sulfide contents of the refractory samples were variable up to 3.03% and averaged 1.55%. Organic carbon concentrations are at levels that do not indicate preg-robbing (active carbonaceous matter that will adsorb dissolved gold as it leaches). The spiked preg-rob shake flask test results showed low preg-robbing for most of the samples and moderate preg robbing for the remainder. Low to moderate preg-robbing is typically overcome with carbon-leach (CIL) that will overcome the effect of natural carbonaceous matter.

The two oxide samples exhibited almost no preg-robbing.

 

10.1.3

Recovery Estimates

 

10.1.3.1

Archimedes Refractory Mineralization

Recoveries for refractory mineralization were estimated using the average leach recovery from tests using alkaline oxidation followed by CIL on sulfide refractory material (16 data points). The average of the 16 BTC refractory samples is 25.3 %CO32- and 1.4 %S2-. This gives a CO32-: S2- ratio of 18. As a general rule, acid autoclaving is preferred when this ratio is less than 5:1, while alkali autoclaving is preferred when the ratio is greater than 5:1. The average of the refractory samples is 23.2, therefore, these samples are firmly in the alkali autoclaving territory. The 2024 FLSmidth program samples had an average ratio of 9.1:1, although without the 426 Zone central sample, the ratio reduces to 5.3:1.

The results from testing of alkaline pressure oxidation followed by CIL indicated an average leach recovery of 88% could be achieved. Recovery from acid pressure oxidation is higher but is expected to have poorer economics due to the amount of sulfuric acid needed to destroy the carbonate ahead of autoclaving.

 

10.1.3.2

Recommended Recoveries

A summary of the gold recoveries is shown in Table 10-8. Autoclave/CIL recoveries are based on acid pressure oxidation conditions (BTAC Condition C) described in 10.1.1.5.

Table 10-8: Ruby Hill (Archimedes) Summary of Estimated Gold Recoveries

 

 Mineralization Type   

Autoclave/CIL Recovery

(Au %)

426 Zone

   96.8

Ruby Deeps Windfall

   96.0

Ruby Deeps Dunderberg Shale

   98.1

Ruby Deeps Intrusive

   38.8

 

 

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The Ruby Deeps North sample was excluded from the recovery estimates as it is on the periphery of the zone and may no longer be representative. The Ruby Deeps Composite sample was also not considered as the sample is largely from two drill holes and not fully representative.

The weighted average (by lithology) Ruby Deeps CIL gold recovery is 89.5% and 94.6% (without the Intrusive zone). The latter assumes that the Intrusive Zone is not mined.

 

10.2

Mineral Point Open Pit

This section summarizes all the relevant test work performed on the Mineral Point open pit project. The Mine project encompasses several deposits and mineralization types hosting both precious and base metals. Historical production dates to 1998, primarily under Homestake Mining and Barrick Gold, with intermittent operations up to the current date. Characteristics of each deposit, historical production and metallurgical interpretation for the Mineral Point deposits are described in this section, based on data provided by Ruby Hill Mining LLC. Generally, metallurgical test work confirms the amenability of oxide mineralization to heap leaching for precious metals extraction.

 

10.2.1

Historical Operations

Historical operations at Ruby Hill have included three process routes for production: run of mine (ROM) and crushed production to heap leaching, crushing and leaching with agglomerated tailings routed to the heap leach pad, and higher-grade sulfide production (DSO) routed to Goldstrike for autoclave processing. Currently there is residual heap leaching of previously stacked material. This heap leach will be replaced by a new heap leach pad and solution management system.

 

10.2.2

Historical Test Work

A series of historical metallurgical test reports previously completed for other studies on the Ruby Hill Project are shown in Table 10-9.

Table 10-9: Ruby Hill Project Historical Metallurgical Testing Programs

 

No.

  

Document Title

  

Deposit

  

Technical Content

  

Date

1

  

Ruby Hill Project, East Archimedes, Report

of Metallurgical Test Work, Kappes

Cassiday Associates

   Archimedes    Column leach tests    2004

2

   Kappes Cassiday Associates    Archimedes    Column leach tests    2005

3

   Barrick Technology Center    426 Zone    Roasting, pressure oxidation, leach tests    2008

4

   G&T Metallurgy    426 Zone    Flotation, leach tests    2008

5

   G&T Metallurgy    Blackjack    Flotation    2008

6

   Kappes Cassiday Associates    426 Zone    Column leach tests    2009

7

   Kappes Cassiday Associates    426 Zone    Column leach tests    2011

8

   Kappes Cassiday Associates    Mineral Point     Column leach tests    2011

9

   Kappes Cassiday Associates    Mineral Point     Column leach tests    2012

10

   Kappes Cassiday Associates    Mineral Point     Column leach tests    2014

 

10.2.2.1

Archimedes Deposit

 

10.2.2.1.1

June 2004 KCA Column Leach Test Program

Nineteen separate column leach tests were conducted on the core composites, sulfide composite and bulk ROM samples received from the Ruby Hill Project at Kappes Cassiday Associates (KCA). Tests were conducted at a crush size approximating ROM material and crushed material at –1.25” Column tests ran between 40 and 62 days of leaching. Results are summarized in Table 10-10.

 

 

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The overall average gold extraction for the samples was 82%, the average sodium cyanide consumption was 0.82 lb/ton, and the average hydrated lime consumption was 3.60 lb/ton. Sample 31624 was labeled sulfide and had a low recovery of only 31%.

Table 10-10: June 2004 KCA Archimedes Column Test Results Summary

 

 Sample   

Crush Size

(inches)

  

Head Grade

(opt Au)

   Leach
Extraction (Au
%)
   Reagent Consumption (lb/ton)
   Cyanide   

Hydrated

Lime

Low Grade Oxide

   -1.25    0.020    84    0.62    4.00

High Grade Oxide

   -1.25    0.156    84    0.74    3.00

Low Grade Oxide

   -1.25    0.029    81    0.62    2.00

Low Grade Oxide

   -1.25    0.015    72    0.30    2.00

High Grade Oxide

   -1.25    0.272    85    0.80    4.02

Medium Grade Oxide

   -1.25    0.089    86    1.08    6.04

Medium Grade Oxide

   -1.25    0.078    88    0.64    2.00

Medium Grade Intrusive

   -1.25    0.063    87    0.98    5.02

Low Grade Intrusive

   -1.25    0.018    78    0.70    5.00

Medium Grade Oxide

   -1.25    0.091    84    0.54    5.02

Low Grade Oxide

   -1.25    0.044    79    0.26    5.02

High Grade Oxide

   -1.25    0.241    88    0.82    2.00

High Grade Oxide

   -1.25    0.387    86    0.86    2.00

Medium Grade Oxide

   -1.25    0.061    87    0.44    2.00

Oxide

   ROM    0.032    90    0.42    2.20

Oxide

   ROM    0.030    91    0.48    2.20

Oxide

   -1.25    0.032    90    0.42    2.20

Oxide

   -1.25    0.030    89    0.70    2.14

High Grade Sulfide

   -1.25    0.357    31    3.42    10.44

 

10.2.2.1.2

May 2005 KCA Program

Eight separate column leach tests were conducted on four samples received from the Archimedes deposit. Two column tests were conducted on each sample, one at the as received size and another set at a crush size of -1.5”. The 80% passing size (k80) of the ROM and -1.5” tests ranged from approximately 0.20” to 0.60” and there was little difference between the average gold extractions for the as received and crushed material. The column tests ran from 41 to 121 days (ROM3 and ROM5 as received). The results are summarized in Table 10-11.

Table 10-11: May 2005 KCA Archimedes Column Test Results Summary

 

 Sample    Crush Size (inches)   

Head Grade

(opt Au)

   Leach
Extraction (Au
%)
   Reagent Consumption (lb/ton)
   Cyanide   

Hydrated

Lime

ROM3

   ROM    0.152    90    0.32    2.00

ROM4

   ROM    0.010    80    0.42    2.00

ROM5

   ROM    0.086    70    0.24    2.00

ROM6

   ROM    0.014    65    0.22    1.76

ROM3 crushed

   -1.5    0.147    93    0.62    2.00

ROM4 crushed

   -1.5    0.012    67    0.20    2.00

ROM5 crushed

   -1.5    0.084    75    0.40    2.00

ROM6 crushed

   -1.5    0.014    71    0.74    2.00

 

10.2.2.2

426 Zone

 

10.2.2.2.1

January 2009 KCA Program

Metallurgical test work completed on two composites (low and high-grade oxide material) included density testing, head analyses, coarse and pulverized bottle roll leach tests, as well as compacted permeability tests and column leach tests.

 

 

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The bottle roll leach tests carried out on the low-grade composite had gold recoveries of between 82% and 91% could be achieved on material crushed to -1” and 89% on pulverized material. Sodium cyanide consumption ranged from <0.02 to 1.32 lb/ton depending on the concentration used in the leach solutions.

Bottle roll leach tests carried out on the high-grade composite indicated that slightly higher gold recoveries of between 86% and 91% could be achieved on material crushed -1”and 95% on pulverized material. Sodium cyanide consumption ranged up to 1.24 lb/ton depending on concentration used in the leach solution.

Column leach gold extraction from the high-grade composite material, crushed to -1” was higher at 90% after 75 days of leaching. Sodium cyanide consumption was 0.07 lb/ton and hydrated lime addition was 3.00 lb/ton. The results are summarized in Table 10-12.

Table 10-12: January 2009 426 Zone KCA Column Leach Test Results Summary

 

 Sample    Crush Size
(inches)
  

Head Grade

(opt Au)

   Leach
Extraction (Au
%)
   Reagent Consumption (lb/ton)
   Cyanide    Hydrated Lime

Low grade

   -1    0.020    85    0.06    3.00

High grade

   -1    0.102    90    0.08    3.00

 

10.2.2.2.2 

November 2011 KCA Program

Metallurgical test work completed on eight samples included, head analyses, size by size analyzes, coarse and pulverized bottle roll leach tests, and column leach tests.

The bottle roll leach tests had gold extractions between 78% and 94% could be achieved when pulverized to a k80 = -200 mesh. When pulverized to -10 mesh, gold extractions ranged from 72% and 89%.

Column leach gold extractions crushed to -1” ranged from 81% to 93%. Sodium cyanide consumptions ranged from 0.52 lb/ton to 3.02 lb/ton. Hydrated lime consumptions were an average of 2.0 lb/ton. The results are summarized in Table 10-13

Table 10-13: November 2011 426 Zone KCA Column Leach Test Results Summary

 

 Sample    Crush Size
(inches)
  

Head Grade

(opt Au)

   Leach
Extraction (Au
%)
   Reagent Consumption (lb/ton)
   Cyanide    Hydrated Lime

BRH-95C, BRH-99C

   -1    0.91    81    0.86    2.02

BRH-99C, BRH-211C

   -1    1.87    93    0.72    2.00

BRH-101C

   -1    2.40    92    0.52    2.06

BRH-210C, BRH-211C

   -1    1.54    93    1.28    2.02

BRH-213C

   -1    2.33    84    0.66    2.02

BRH-214C

   -1    0.93    91    0.98    2.00

BRH-214C

   -1    6.00    84    3.02    2.02

BRH-212C

   -1    1.70    89    2.06    2.00

 

10.2.2.3

Mineral Point Deposit

 

10.2.2.3.1 

February 2011 KCA Program

The Mineral Point Deposit (formerly named the Bullwhacker Deposit) samples were described as:

 

   

BW-1 Hamburg Dolomite – This sample is dominated by hematite altered sanded dolomite containing secondary goethite after pyrite cubes. The entire interval is oxidized.

 

   

BW-2 Hamburg Dolomite – This sample is again dominantly hematite and limonite altered sanded dolomite. The entire zone is oxidized.

 

 

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BW-3 Hamburg Dolomite and Dunderberg Shale – A small part of this sample is composed of a slightly calcareous limonite altered silicified shale. The rest of the interval is composed of a breccia containing clasts of vuggy silicified dolomite in an argillic, hematite, and goethite altered matrix.

The samples were utilized for head analyses, bottle roll cyanide leach, cyanide shake and column leach test work, acid-base accounting (ABA) and meteoric water mobility procedure (MWMP) testing.

For the pulverized bottle roll tests, gold extraction ranged from 77% to 84% with an average of 81%. For the coarse bottle roll tests, gold extraction ranged from 61% to 83% with an average of 72%. Results are shown in Table 10-14.

Table 10-14: February 2011 Mineral Point Deposit KCA Bottle Rolls Test Results Summary

 

Sample    Lithology   

Crush Size
(mesh or

inches)

  

Head Grade

(opt Au)

   Leach Extraction
(Au %)
   Reagent Consumption (lb/ton)
   Cyanide    Hydrated Lime

BW-1, Pulverized

  

Hamburg

Dolomite/Sanded

   -200M    0.013    77    0.76    1.00

BW-1, Coarse

   -10M    0.011    83    0.06    1.00

BW-2, Pulverized

  

Hamburg

Dolomite/Weakly-Altered

   -200M    0.052    84    1.14    1.00

BW-2, Coarse

   -0.225”    0.067    74    0.40    1.00

BW-3, Pulverized

  

Dunderburg Shale and

Hamburg

Dolomite/Silicic

   -200M    0.038    82    1.00    2.00

BW-3, Coarse

   -1.0”    0.037    61    0.62    1.00

BW-3, Coarse

   -0.361”    0.046    70    0.70    1.00

Column leach tests were conducted on samples from each of the composites. Five of the column tests were conducted at a crush size of -0.5” and were run for a period of 91 days. The column leach test average gold recovery was 80%. On one of the samples (BW-3), two columns were run, one at -0.5” and the other at -1.5”, the recovery from the coarser column was only 1% lower. Samples BW-1 and BW-2 were run with and without agglomeration. Results between the two were relatively close, indicating agglomeration is not required. The results are summarized in Table 10-15. The average sodium cyanide consumption was 1.38 lb/ton. Lime and cement consumptions were variable.

Table 10-15: February 2011 Mineral Point Deposit KCA Column Leach Test Results Summary

 

 Sample    Lithology    Crush Size
(inches)
  

Head Grade

(opt Au)

   Leach Extraction    Reagent Consumption (lb/ton)
   (Au %)    (Ag %)    Cyanide    Hydrated
Lime
   Cement

BW-1

  

Hamburg

Dolomite/Sanded

   -0.5    0.010    85    35    1.10    2.0    — 

BW-1

Agglomerated

   -0.5    0.013    84    39    0.70    —     8.0

BW-2

  

Hamburg

Dolomite/

Weakly

-Altered

   -0.5    0.050    82    50    1.30    2.0    — 

BW-2

Agglomerated

   -0.5    0.051    81    46    0.82    —     8.0

BW-3

Coarse Crush

  

Dunderburg Shale and

Hamburg

Dolomite/

Silicic

   -1.5    0.036    74    14    2.06    2.0    — 

BW-3

Fine Crush

   -0.5    0.037    75    15    2.30    2.0    — 

 

10.2.2.3.2 

July 2012 KCA Program

Samples originated from four drill cores from the Mineral Point deposit. The samples for this program were utilized for head analyses, size by size analysis, bottle roll cyanide leach, agglomeration testing and column leach test work, acid-base accounting (ABA) and meteoric water mobility procedure (MWMP) testing.

 

 

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Cyanide bottle roll leach tests were conducted on each of the samples at crush sizes -1.5” and -0.5” and pulverized to k80 = 200 mesh. Results are summarized in Table 10-16. Results from the pulverized tests for two of the samples, RH 231 and RH 235A, showed anomalously low recoveries compared to the coarse crush sizes. All samples had low sulfide sulphur concentrations so are considered as oxide samples.

Table 10-16: July 2012 Mineral Point Deposit KCA Bottle Rolls Test Results Summary

 

 Sample   Lithology    Crush Size (mesh or
inches)
  

Head Grade

(opt Au)

  

Leach

Extraction

(Au %)

  

Reagent Consumption

(lb/ton)

   Cyanide   

Hydrated

Lime

RH 184  

Hamburg

Dolomite/Silicic and

Weakly-Altered

   -1.5”    1.41    70    0.16    0.50
   -0.5”    0.044    74    0.32    1.00
   -200M    0.042    84    1.66    1.50
RH 231  

Hamburg

Dolomite/Silicic

   -1.5”    0.015    71    0.74    1.00
   -0.5”    0.0.16    76    0.78    1.00
   -200M    0.016    57    0.99    1. 50
RH 235A  

Hamburg

Dolomite/Weakly-

Altered

   -1.5”    0.013    73    0.32    1.00
   -0.5”    0.013    72    0.34    1.00
   -200M    0.013    58    0.52    1.00
RH 235B  

Hamburg

Dolomite/Weakly-

Altered

   -1.5”    0.058    77    0.32    1.00
   -0.5”    0.053    77    0.42    1.00
   -200M    0.056    76    0.84    1.50

Column leach tests were conducted at crush sizes of -1.5” and -0.5” for all samples. Sample RH 184 was agglomerated at both crush sizes. Samples RH 231, 235A and RH 235B were agglomerated at the coarse crush size. Results are summarized in Table 10-17.

The overall gold extractions ranged from 81% to 86% over the 93-day leach period. The cyanide consumptions ranged from 1.80 to 4.52 lb/ton. Hydrated lime consumptions were about 1.00 lb/ton, and cement additions ranged from 4.04 to 4.16 lb/ton. Some tests had high sodium cyanide consumptions although there is no apparent reason as samples are low in sulfide sulphur and soluble copper.

Table 10-17: July 2012 Mineral Point Deposit KCA Column Leach Test Results Summary

 

 Sample    Lithology   

Crush

Size
(inches)

  

Head

Grade

(opt Au)

   Leach Extraction    Reagent Consumption (lb/ton)
   (Au %)    (Ag %)    Cyanide    Hydrated
Lime
   Cement
RH 184   

Hamburg

Dolomite/Silicic

and Weakly-Altered

   -1.5    0.040    82    34    0.90    1.02    4.04
   -0.5    0.046    86    58    2.06       4.12
RH 231   

Hamburg

Dolomite/Silicic

   -1.5    0.014    88    34    1.56    1.00    — 
   -0.5    0.014    81    39    1.16       4.10
RH 235A   

Hamburg

Dolomite/Weakly-Altered

   -1.5    0.013    84    27    1.26    1.00    — 
   -0.5    0.013    82    47    1.14       4.16
RH 235B   

Hamburg

Dolomite/Weakly-Altered

   -1.5    0.051    86    48    1.30    1.00    — 
   -0.5    0.046    82    52    2.26       4.04

 

10.2.2.3.3 

February 2014 KCA Program

Samples originated from four drill cores from the Mineral Point deposit. These samples were utilized for head analyses, head screen analyses with assays by size fraction, comminution test work, bottle roll leach test work and column leach test work.

Cyanide bottle roll leach tests were conducted on each of the samples at pulverized to-10 mesh and to k80 = 200 mesh. Results are summarized in Table 10-18. Gold extractions ranged from 22 to 86%. The sodium cyanide consumptions ranged from 0.04 to 3.48 lb/ton. The samples utilized in leaching was blended with

 

 

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2.00 to 10.0 lb/ton hydrated lime. Extraction increased by an average of 6% when the samples were pulverized from a nominal size of -10 mesh to a target size of k80 = 200 mesh. Sample BRH 445C (580-632.2) originates from the Dunderberg Shale Zone within the Mineral Point deposit with high sulfide sulphur and arsenic content, which likely contributed to the low recoveries.

Table 10-18: February 2014 Mineral Point Deposit KCA Bottle Rolls Test Results Summary

 

 Sample    Lithology    Crush Size
(mesh)
  

Head Grade

(opt Au)

  

Leach

Extraction (Au
%)

   Reagent Consumption (lb/ton)
   Cyanide   

Hydrated

Lime

BRH 445C

(580-632.2)

  

Dunderberg

Shale/Weakly-
Altered (sulfide)

   -10M    0.029    22    3.48    7.50
   -200M    0.028    30    13.08    10.00

BRH 445C

(632.2-670)

  

Dunderberg Shale

and Hamburg

Dolomite/Weakly-

Altered (oxide and

sulfide)

   -10M    0.014    67    0.48    2.76
   -200M    0.014    69    0.30    7.00
BRH 266C   

Hamburg

Dolomite/Silicic

and Sanded

   -10M    0.010    74    0.04    2.00
   -200M    0.010    86    0.38    6.00
BRH 317C   

Hamburg

Dolomite/Weakly-

Altered

   -10M    0.029    57    1.68    2.50
   -200M    0.029    59    2.40    7.00
BRH515C   

Hamburg

Dolomite/Weakly-

Altered

   -10M    0.014    80    0.48    2.26
   -200M    0.014    83    0.64    7.00
BH343C   

Hamburg

Dolomite/Weakly-

Altered and

Sanded

   -10M    0.016    67    0.16    2.00
   -200M    0.016    73    1.66    4.00

Column leach tests were conducted at crush sizes of -1.0” and -0.75” for all samples and leached for 69 days. Samples BRH 266C and BRH 343C at both crush sizes failed column percolation tests completed at the end of the leach cycles. However, gold extraction for these columns was consistent with the other column tests. Results are summarized in Table 10-19.

For column leach tests, gold extractions ranged from 29% to 85% based on calculated heads which ranged from 0.010 to 0.034 opt. The sodium cyanide consumptions ranged from 0.62 to 4.84 lb/ton. The samples utilized in leaching were blended with 2.00 to 9.62 lb/ton hydrated lime. Extraction increased by an average of 4% when the crush size was reduced from 100% passing 0.5” to 100% passing 0.75”. The high cyanide consumption from the BRH 445C (580-632.2) leach tests is attributed to high sulfide sulphur content.

 

 

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Table 10-19: February 2014 Mineral Point Deposit KCA Column Leach Test Results Summary

 

Sample    Lithology    Crush Size
(inches)
  

Head

Grade

(opt Au)

   Leach Extraction   

Reagent Consumption

(lb/ton)

   (Au %)    (Ag %)    Cyanide    Hydrated
Lime

BRH 445C

(580-632.2)

   Dunderberg
Shale/Weakly-
Altered (sulfide)
   -1.0    0.031    29    32    4.52    7.52
   -0.75    0.033    31    48    4.84    9.62

BRH 445C

(632.2-670)

   Dunderberg Shale
and Hamburg
Dolomite/Weakly-
Altered (oxide and sulfide)
   -1.0    0.019    71    39    1.24    2.76
   -0.75    0.017    70    43    0.62    6.96

BRH 266C

   Hamburg
Dolomite/Silicic and
Sanded
   -1.0    0.010    76    6    1.08    2.00
   -0.75    0.011    81    15    2.72    6.04

BRH 317C

   Hamburg
Dolomite/Weakly-
Altered
   -1.0    0.034    57    24    1.08    2.05
   -0.75    0.029    62    29    0.74    6.98

BRH 515C

   Hamburg
Dolomite/Weakly-
Altered
   -1.0    0.021    63    20    1.62    2.50
   -0.75    0.014    85    20    0.99    6.98

BRH 343C

  

Hamburg
Dolomite/Weakly-Altered

and
Sanded

   -1.0    0.015    83    25    0.62    2.00
   -0.75    0.016    74    27    0.70    4.02

 

10.2.3

Mineral Point Leach Cycle Times

Leach cycle times for full scale heap leach operations is typically measured in tons of leach solution applied to tons of ore under leach. The full leach cycle is not normally completed with a single continuous application of solution. The cycle is usually broken down into the primary leach cycle where solution is directly applied to the ore under leach and a secondary leach cycle where solution flows throw an area previously leached from a lift above. The primary leach cycle typically is at a solution application rate of 1:1. The remainder of the recovery would be obtained during secondary leaching as ore in subsequent lifts above are leached. The design final solution application rate is typically 4:1.

The Mineral Point column leach tests showed leach times between 6 days and 34 days to achieve the solution application rate of 1:1. Between 80% and 99% of ultimate Au extractions were achieved within this period excluding sulfide and mixed oxide/sulfide samples. Days of leach in column tests are scaled up based on lift height, bulk density and the size of a block under leach. For this technical report, a primary leach time of 90 days is recommended.

Average retained moisture contents for the three Mineral Point column test programs ranged from 18.6 gallons/ton to 28.2 gallons/ton.

 

10.2.4

Mineral Point Reagent Consumptions

Based on the column test results, recommended sodium cyanide and quicklime consumption rates are 1.0 lb/ton and 8 lb/ton respectively.

 

10.2.5

Deleterious Elements

A wide range of analyses were carried out on the samples used in the metallurgical testing programs included in this section. Deleterious elements were identified that are common to deposits in this part of Nevada. Deleterious elements content in the oxide samples are low, while sulfide samples are characterized by high levels of sulfide sulfur, arsenic, and mercury.

 

 

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10.2.5.1

Arsenic and Mercury

The KCA January 2009 report conducted investigations into arsenic and mercury deportment. Although a note was added stating that as multi-acid digestion was specified, the values for arsenic and mercury may be biased low due to partial volatilization upon digestion.

The arsenic contents of the refractory samples were variable up to 2.6% and averaged 0.4%. One of the two oxide samples had a relatively high arsenic content of 0.43%. Processing of arsenical refractory production through either pressure oxidation or roasting results in the capture and sequestration of arsenic in a stable form suitable for tailings disposal.

Mercury contents in the low and high-grade oxide composites from the 426 zone were moderate at 5.7 ppm and 9.6 ppm respectively.

The KCA February 2014 report analyzed six samples for mercury, they were reported as being between 2 and 10 ppm. All 16 refractory samples documented in the BTC November 2011 report had levels of less than 10 ppm Hg. Mercury concentrations at this level require the inclusion of mercury retorting in electrowinning and gold smelting areas of process facilities and mercury capture equipment on carbon reactivation kilns.

 

10.2.6

Recovery Estimates

Gold and silver recovery estimates were completed using the methodologies described in the following sections.

 

10.2.6.1

Oxide Mineralization

The test results from the four KCA reports relevant to Archimedes, 426 and Mineral Point zones are summarized in Table 10-20. The resources for this technical report include only Mineral Point.

Table 10-20: Summary of Column Leach Test Results

 

Test Program    Zone   

Crush Size

(inches)

  

Leach Extraction

 

   No. of Samples
   (Au %)    (Ag %)

2004-06 KCA

   East Archimedes    -1.25    84.0    13.5    15

2004-06 KCA

   East Archimedes    ROM    90.5    1.5    2

2005-05 KCA

   East Archimedes    -1.5    76.0    3.0    4
   East Archimedes    ROM    77.0    1.0    4

2009-01 KCA

   426    -0.75    87.5    10.0    2

2011-11 KCA

   426    -1.0    88.0    42.0    8

2011-02 KCA

   Mineral Point Oxide    -0.5    83.0    42.5    4
   Mineral Point Mixed    -0.5    75.0    15.0    2

2012-07 KCA

   Mineral Point Oxide    -1.5    85.0    36.0    4
   -0.5    82.8    49.0    4

2014-02 KCA

   Mineral Point Oxide    -1.0    74.0    25.0    3
   -0.75    80.0    30.5    3
   Mineral Point Mixed    -1.0    64.0    31.5    2
   -0.75    66.0    36.0    2

Analysis of the results from all programs showed that crush size had minimal impact on recoveries. Results from crushed and ROM samples are considered as one dataset.

The two KCA programs documented in June 2004 and May 2005 reports were carried out on oxide samples from the East Archimedes deposit. The column tests show no variation of gold recovery with gold grade, or crush size, with the two ROM samples having slightly higher recovery than the crushed samples, likely due to these tests running for longer durations. In the 2005 program, the particle size of the ROM, as-received material was only slightly coarser than the crushed material and recoveries were similar.

 

 

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The KCA January 2009 and November 2011 programs were carried out on oxide samples from the 426 zone.

The three KCA programs February 2011, July 2012 and February 2014 were carried out on thirteen Mineral Point samples. One sample was identified as sulfide, another as mixed and two others contained significant amounts of Dunderberg Shale. The remaining eight were identified as Hamburg Dolomite. The selected flowsheet includes two stage crushing to -0.75”. Column test results show minimal response to finer crush sizes; consequently, all crush sizes were included for recovery estimates. Recoveries were assigned based on the alteration (silicic, sanded or weakly altered) and were used to predict recoveries from within the Mineral Point deposit. Mixed lithology/alteration samples were excluded as recoveries were assigned based on coded alteration in the block model. While the sample set and column leach test results are not large; they are sufficient for this report.

 

10.2.6.2

Recommended Recoveries

A summary of the design gold and silver recoveries based on the alteration types is shown in Table 10-21.

Table 10-21: Mineral Point Summary of Estimated Gold and Silver Recoveries

 

Alteration   

Crushed Heap
Leach
Recovery

(Au %)

  

Crushed Heap
Leach
Recovery

(Ag %)

Silicic Oxide

   84.4    45.2

Silicic Sulfide

   31.0    45.2

Sanded Oxide

   83.5    44.0

Sanded Sulfide

   24.0    44.0

Weakly Altered Oxide

   83.0    40.0

Weakly Altered Sulfide

   24.0    40.0

 

 

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11.

MINERAL RESOURCE ESTIMATES

 

11.1

Introduction

The mineral resource estimate presented herein has been prepared following the guidelines of the Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 1300 through 1305).

Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no guarantee that all or any part of the mineral resources will be converted into mineral reserves. Confidence in the estimate of inferred mineral resources is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability sufficient for public disclosure.

Practical Mining LLC (Practical) estimated the Archimedes Underground mineral resource using all drilling and geological data available through October 31, 2022. Wood Canada Ltd. (Wood) completed the Mineral Point open pit Mineral Resource Estimate in the inaugural NI 43-101 Technical Report (July 2021) under i-80’s ownership of the Ruby Hill Project. Forte Dynamics, Inc (Forte) reviewed the Mineral Point open pit mineral resource Estimate completed by Wood (July 2021). Upon completion of the Mineral Point open pit resource review, Forte made some slight modifications to the Wood resource block model (estimated block grades were not changed or altered) along with using an updated constraining pit shell to report the Mineral Point open pit mineral resource Estimate. Forte also completed an updated mineral resource estimate for the Archimedes open pit deposit.

All work, including drilling, completed since the time of the inaugural technical report has targeted the 426 and Ruby Deeps deposits and does not influence the Mineral Point open pit mineral resource. The Archimedes open pit mineral resource was completed using all current drilling and geological data available through December 31, 2024.

Open pit and underground block model horizontal extents are shown in Figure 11-1. The Archimedes open pit model extends vertically from 7,500 to 6,700 feet amsl, the Mineral Point open pit model extends vertically from 4,600 to 6,900 feet amsl and the Archimedes underground model from 4,000 to 5,300 feet amsl.

 

 

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Figure 11-1: Block Model Extents

(Source: Forte Dynamics, 2025)

 

11.2

Archimedes Underground

In 2022, i-80 moved all drillhole data completed by i-80 and previous property owners to an acQuire database, an industry standard relation SQL data management solution. Collar, downhole survey, assay and geological data was exported to comma-separated values files on February 2, 2023. Practical converted the drill hole data to Vulcan version 11.1 format. i-80 created lithologic and structural models using Leapfrog software which were also imported into Vulcan 11 .1

 

 

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Figure 11-2: Underound Model Extents and Drill Hole Traces

(Source: Practical Mining, 2025)

 

11.2.1

Grade Shells

Practical explicitly modeled grade shells at nominal 0.004 Au opt and 0.1 Au opt limits using lithologic boundaries and the Holly, 426, Graveyard Flats, and Blanchard Faults as general guides. Intercept grades below the shell cutoff were included where the intercept fell within the trend of the grade shell. Similarly, intercepts above the shell grade that are distant and discontinuous were excluded. There are eight (8) unique 0.1 Au opt grade shells and one (1) 0.004 Au opt grade shell in the 426 deposit. The Ruby Deeps deposit contains 15 high grade and two (2) low grade shells. Two high grade and one low grade shell lie west of and on the footwall of the Holly Fault.

 

11.2.2

Density

The Ruby Hill database contains 985 density determinations completed by the previous property owners. i-80 has not completed any density measurements. Univariate statistics sorted by lithology formation are listed in Table 11-1 and graphically in Figure 11-3.

A Vulcan script assigned mean density values in tons per cubic foot to the block model.

 

 

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Table 11-1: Univariate Density Statistics by Lithology Formation (tonnes/m3)

 

     Cd    Ch    Csc    Cwb    Cwc    Kint    Og    Op    Qal    Unk

Count

   56    336    5    49    32    22    242    15    1    195

Mean

   2.516    2.604    2.660    2.461    2.523    2.478    2.595    2.520    2.00    2.572

Std Dev

   0.178    0.211    0.055    0.207    0.172    0.098    0.225    0.109         0.328

CV

   0.071    0.081    0.021    0.084    0.068    0.040    0.087    0.043       0.128

Lower 95% CI

   2.470    2.582    2.612    2.403    2.463    2.437    2.567    2.465         2.526

Upper 95% CI

   2.563    2.627    2.708    2.519    2.582    2.519    2.623    2.575       2.618

Min

   1.760    1.450    2.580    1.890    2.030    2.280    1.780    2.320    2.00    1.910

25% Quartile

   2.438    2.538    2.640    2.340    2.463    2.423    2.520    2.445    2.00    2.440

Median

   2.560    2.660    2.660    2.530    2.585    2.450    2.620    2.520    2.00    2.560

75% Quartile

   2.630    2.723    2.700    2.620    2.633    2.550    2.670    2.600    2.00    2.660

Max

   2.790    3.350    2.720    2.700    2.810    2.660    3.950    2.670    2.00    6.120

Note: One tonne/m3 = 0.0312 tons per ft3

 

LOGO

 

 

Figure 11-3: Density Box and Whisker Plot by Lithology Formation

(Source: Practical Mining, 2025)

 

 

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11.2.3

Statistics

Drill holes were composited such that all composites are approximately ten-feet (10 ft) in length and cut at the grade shell boundary. Each composite is flagged by a grade shell name.

Gold and Silver univariate statistics for each grade shell are presented in Table 11-2 through Table 11-7 and are also presented graphically in the Box and Whisker plots of Figure 11-4 through Figure 11-6.

 

LOGO

Figure 11-4: 426 0.1 Au opt Box and Whisker Plots

(Source: Practical Mining, 2025)

Table 11-2: Gold Univariate Statistics for 426 0.1 Au opt Composites

 

Grade Shell    426-02    426-03    426-04    426-05    426-06    426-07    426-10    426-19    007    008

Count

   84    43    43    40    152    99    28    20    12    7

Length

   815.4    427.5    424.2    362.2    1474.8    988.9    279.8    181.1    118.4    56.0

Std_Dev

   0.118    0.118    0.168    0.131    0.174    0.111    0.157    0.098    0.137    0.108

Lower 95% CI

   0.170    0.140    0.154    0.126    0.152    0.117    0.157    0.120    0.189    0.214

Mean

   0.196    0.175    0.204    0.167    0.179    0.139    0.216    0.163    0.266    0.295

Upper 95% CI

   0.221    0.210    0.254    0.207    0.207    0.161    0.274    0.205    0.344    0.375

Minimum

   0.048    0.039    0.000    0.000    0.000    0.000    0.003    0.016    0.124    0.200

25% Quartile

   0.106    0.111    0.111    0.095    0.063    0.063    0.110    0.106    0.168    0.223

Median

   0.154    0.134    0.167    0.138    0.138    0.133    0.194    0.150    0.219    0.238

75% Quartile

   0.271    0.200    0.276    0.197    0.238    0.182    0.259    0.202    0.335    0.339

Maximum

   0.605    0.617    0.882    0.571    1.060    0.628    0.669    0.480    0.513    0.502

 

 

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Table 11-3: Silver Univariate Statistics for 426 0.1 Au opt Composites

 

Grade Shell

   426-02    426-03    426-04    426-05    426-06    426-07    426-10    426-19    007    008

Count

   84    43    43    40    152    99    28    20    12    7

Length

   815.4    427.5    424.2    362.2    1474.    988.9    279.8    181.1    118.4    56.0

Std_Dev

   0.040    0.006    0.012    0.040    0.166    0.050    0.100    0.009    0.003    0.004

Lower 95% CI

   0.023    0.009    0.009    0.012    0.013    0.022    0.034    0.008    (0.000)    0.007

Mean

   0.031    0.011    0.013    0.024    0.040    0.032    0.071    0.012    0.002    0.010

Upper 95% CI

   0.040    0.013    0.016    0.037    0.066    0.042    0.108    0.016    0.004    0.013

Minimum

   0.000    0.003    0.000    0.000    0.000    0.000    0.000    0.003    0.000    0.005

25% Quartile

   0.012    0.007    0.006    0.004    0.005    0.005    0.019    0.004    0.000    0.008

Median

   0.020    0.010    0.010    0.009    0.016    0.010    0.034    0.012    0.000    0.009

75% Quartile

   0.034    0.015    0.016    0.017    0.036    0.021    0.059    0.015    0.002    0.014

Maximum

   0.248    0.030    0.061    0.154    2.038    0.164    0.425    0.041    0.009    0.016

 

LOGO

Figure 11-5: Ruby Deeps 0.1 Au opt Box and Whisker Plots

(Source: Practical Mining, 2025)

 

 

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Table 11-4: Gold Univariate Statistics for Ruby Deeps 0.01 Au opt Composites

 

Grade Shell

     rd-01        rd-08        rd-09        rd-11        rd-12        rd-13        rd-14        rd-15        rd-16        rd-17        rd-18        rd-20        rdhw-01        rdhw-02  

Count

     140        12        46        89        22        31        9        3        16        16        15        12        12        4  

Length

     1361.        107.0        458.7        880.8        217.6        309.3        93.5        25.0        140.7        164.7        143.1        101.5        124.8        41.0  

Std_Dev

     0.123        0.088        0.126        0.224        0.106        0.072        0.118        0.063        0.066        0.131        0.052        0.087        0.070        0.070  

Lower 95% CI

     0.191        0.111        0.146        0.236        0.101        0.124        0.149        0.104        0.109        0.142        0.161        0.162        0.117        0.098  

Mean

     0.212        0.161        0.183        0.283        0.146        0.150        0.226        0.175        0.141        0.206        0.188        0.211        0.157        0.167  

Upper 95% CI

     0.232        0.210        0.219        0.329        0.190        0.175        0.303        0.246        0.173        0.270        0.214        0.260        0.196        0.236  

Minimum

     0.016        0.038        0.000        0.012        0.027        0.042        0.099        0.115        0.000        0.001        0.112        0.109        0.026        0.102  

25% Quartile

     0.131        0.104        0.102        0.149        0.097        0.104        0.177        0.143        0.115        0.108        0.150        0.137        0.110        0.110  

Median

     0.173        0.152        0.163        0.209        0.129        0.128        0.191        0.171        0.139        0.195        0.180        0.201        0.141        0.160  

75% Quartile

     0.261        0.199        0.245        0.339        0.161        0.170        0.218        0.206        0.172        0.276        0.226        0.273        0.206        0.216  

Maximum

     0.650        0.318        0.496        1.343        0.514        0.314        0.473        0.240        0.285        0.445        0.269        0.362        0.261        0.246  

Table 11-5: Silver Univariate Statistics for Ruby Deeps 0.01 Au opt Composites

 

Grade Shell

     rd-01        rd-08        rd-09        rd-11        rd-12        rd-13        rd-14        rd-15       rd-16        rd-17        rd-18       rd-20        rdhw-01       rdhw-02  

Count

     140        12        46        89        22        31        9        3       16        16        15       12        12       4  

Length

     1361.0        107.0        458.7        880.8        217.6        309.3        93.5        25.0       140.7        164.7        143.1       101.5        124.8       41.0  

Std_Dev

     0.072        0.015        0.017        0.105        0.039        0.034        0.254        0.148       0.033        0.062        0.380       0.052        1.727       0.029  

low 95%Ci

     0.042        0.010        0.013        0.051        0.038        0.038        0.065        (0.066)       0.002        0.017        (0.046)       0.027        (0.228)       0.024  

Mean

     0.054        0.019        0.018        0.073        0.054        0.049        0.231        0.101       0.018        0.048        0.147       0.057        0.750       0.052  

Upper 95% CI

     0.066        0.027        0.022        0.095        0.070        0.061        0.398        0.268       0.034        0.078        0.339       0.086        1.727       0.080  

Minimum

     0.000        0.000        0.000        0.000        0.000        0.000        0.032        0.000       0.000        0.000        0.000       0.000        0.000       0.023  

25% Quartile

     0.000        0.009        0.003        0.007        0.027        0.034        0.070        0.016       0.000        0.000        0.000       0.023        0.030       0.037  

Median

     0.040        0.018        0.016        0.041        0.057        0.050        0.082        0.032       0.000        0.017        0.023       0.053        0.059       0.048  

75% Quartile

     0.083        0.031        0.024        0.086        0.071        0.059        0.417        0.151       0.026        0.069        0.070       0.080        0.095       0.063  

Maximum

     0.548        0.043        0.082        0.545        0.143        0.156        0.764        0.270       0.114        0.182        1.493       0.187        5.556       0.091  

 

 

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Figure 11-6: 0.002 Au opt Box and Whisker Plots

(Source: Practical Mining, 2025)

Table 11-6: Gold Univariate Statistics for 0.002 Au opt Composites

 

Grade Shell    1-Low    007-Low    008-Low    2-Low    West-Low

Count

   841    60    12    2159    94

Length

   8201.7    594.8    103.8    21329.0    921.4

Std_Dev

   0.033    0.020    0.032    0.031    0.026

low 95%Ci

   0.023    0.017    0.029    0.026    0.021

Mean

   0.025    0.022    0.047    0.027    0.026

Upper 95% CI

   0.028    0.027    0.065    0.029    0.031

Minimum

   0.000    0.000    0.007    0.000    0.000

25% Quartile

   0.001    0.006    0.021    0.004    0.003

Median

   0.012    0.016    0.039    0.018    0.017

75% Quartile

   0.040    0.037    0.072    0.041    0.045

Maximum

   0.270    0.088    0.099    0.395    0.093

Table 11-7: Silver Univariate Statistics for 0.002 Au opt Composites

 

Grade Shell    1-Low    007-Low    008-Low    2-Low    West-Low

Count

   841    60    12    2159    94

Length

   8201.7    594.8    103.8    21329.0    921.4

Std_Dev

   0.045    0.012    0.015    0.063    0.284

low 95%Ci

   0.020    0.003    (0.000)    0.015    0.033

Mean

   0.024    0.006    0.008    0.018    0.090

Upper 95% CI

   0.027    0.008    0.017    0.020    0.148

Minimum

   0.000    0.000    0.003    0.000    0.000

25% Quartile

   0.003    0.000    0.003    0.003    0.010

Median

   0.005    0.003    0.003    0.006    0.024

75% Quartile

   0.013    0.005    0.006    0.018    0.068

Maximum

   0.188    0.083    0.055    1.705    2.675

 

 

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11.2.4

Grade Capping

Cumulative frequency plots of composite grades were used to determine grade capping values (Figure 11-7 and Figure 11-8). Grade capping values for the high-grade domains were selected to impact no more than 1% of high-grade composites.

 

LOGO

 

Figure 11-7: Gold Cumulative Frequency

(Source: Practical Mining, 2025)

 

LOGO

Figure 11-8: Silver Grade Shells Cumulative Frequency

(Source: Practical Mining, 2025)

Grade cap values for gold and silver are listed in Table 11-8. The range of influence of composites exceeding the grade cap value is restricted to the 25 x 25 x 25 foot block that contains the composite. Within that block the uncapped value is used in the grade estimation and then it is disregarded in the estimation of neighboring blocks.

 

 

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Table 11-8: Gold and Silver Grade Caps

 

Deposit    Assay    0.1 Au opt Grade Shells    0.002 Au opt Grade Shells
   Grade Cap    No. Composites
Capped
   Grade Cap    No. Composites
Capped

426

   Au    0.6    8    0.3    0

Ruby Deep

   Au    0.6    8    0.3    3

426

   Ag    0.2    7    0.2    0

Ruby Deep

   Ag    0.2    19    0.2    22

 

11.2.5

Block Model

Primary block dimensions are 25 x 25 x 25 feet and blocks inside or touching boundaries of the 0.1 opt grade shells are sub blocked to 5 x 5 x 5 feet. Gold, Silver and Cyanide soluble gold grades were estimated for each block using Nearest Neighbor (NN) and Inverse Distance Weighted cubed (IDW3) methodologies. The estimation process was governed by the search ellipsoid dimensions, orientation and sample requirements shown in Table 11-9 and Table 11-10.

Table 11-9: Estimation Search Distances and Sample Requirements

 

Est.

ID

   Grade Shell    Major (ft)    Semi (ft)    Minor (ft)   

Min.

Composites

  

Max.

Composites

   Composites
per DH

Pass 1

   0.1 opt    40    40    40    3    12    2

Pass 2

   0.1 opt    100    100    100    3    12    2

Pass 3

   0.1 opt    300    300    300    3    12    2

Pass 4

   0.1 opt    600    600    600    2    12    2

Pass 5

   0.002 opt    600    600    600    2    12    2

Table 11-10: Ellipsoid Search Parameters for each Grade Shell

 

Grade Shell    Bearing    Plunge    Dip    Grade Shell    Bearing    Plunge    Dip

426-02

   35    0    0    Rd-12    0    0    0

426-03

   35    -12    0    Rd-13    0    -10    0

426-04

   35    -12    0    Rd-14    0    0    0

426-05

   35    -17    0    Rd-15    0    0    0

426-06

   35    -17    0    Rd-16    0    0    0

426-07

   35    -17    0    Rd-17    0    0    0

426-10

   35    0    0    Rd-18    0    0    0

426-19

   0    0    0    Rd-20    0    0    0

Rd-01

   0    -10    0    Rdhw-01    0    0    0

Rd-08

   0    0    0    Rdhw-02    0    0    0

Rd-09

   0    0    0    All Low    0    0    0

Rd-11

   0    -12    0                    

 

11.2.6

Model Validation

A global comparison of composite and block model gold statistics for each grade shell is shown in Table 11-11. Overall, composite and model statistics compare well and are considered acceptable.

 

 

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Table 11-11: Comparison of Composite and Block Model Statistics

 

Shell    Composites    Block Model
      Count    Length    Mean    Std.
Dev
   CV    Max    Upper
Quartile
   Median    Lower
Quartile
   Min    Mean    Std.
Dev.
   CV    Max    Upper
Quartile
   Median    Lower
Quartile
   Min

426-02

   84    815.4    0.196    0.118    0.603    0.605    0.271    0.154    0.106    0.048    0.185    0.087    0.470    0.587    0.220    0.161    0.125    0.007

426-03

   43    427.5    0.175    0.118    0.672    0.617    0.200    0.134    0.111    0.039    0.187    0.076    0.404    0.750    0.222    0.172    0.136    0.030

426-04

   43    424.2    0.204    0.168    0.825    0.882    0.276    0.167    0.111    0.000    0.224    0.102    0.453    0.714    0.292    0.203    0.144    0.003

426-05

   40    362.2    0.167    0.131    0.784    0.571    0.197    0.138    0.095    0.000    0.159    0.099    0.623    0.703    0.181    0.134    0.098    0.024

426-06

   152    1,474.8    0.179    0.174    0.972    1.060    0.238    0.138    0.063    0.000    0.185    0.112    0.608    0.800    0.244    0.162    0.110    0.000

426-07

   99    988.9    0.139    0.111    0.804    0.628    0.182    0.133    0.063    0.000    0.166    0.096    0.580    0.662    0.223    0.140    0.099    0.000

426-10

   28    279.8    0.216    0.157    0.727    0.669    0.259    0.194    0.110    0.003    0.190    0.095    0.502    0.778    0.237    0.189    0.121    0.002

426-19

   20    181.1    0.163    0.098    0.602    0.480    0.202    0.150    0.106    0.016    0.156    0.086    0.551    0.559    0.197    0.151    0.099    0.010

rd-01

   140    1,361.0    0.212    0.123    0.584    0.650    0.261    0.173    0.131    0.016    0.196    0.093    0.471    0.724    0.229    0.176    0.136    0.010

rd-08

   12    107.0    0.161    0.088    0.548    0.318    0.199    0.152    0.104    0.038    0.181    0.063    0.349    0.317    0.226    0.161    0.138    0.024

rd-09

   46    458.7    0.183    0.126    0.692    0.496    0.245    0.163    0.102    0.000    0.164    0.082    0.499    0.578    0.211    0.158    0.108    0.000

rd-11

   88    880.8    0.284    0.225    0.792    1.343    0.339    0.211    0.149    0.012    0.225    0.116    0.513    0.800    0.284    0.194    0.135    0.021

rd-12

   22    217.6    0.146    0.106    0.727    0.514    0.161    0.129    0.097    0.027    0.147    0.074    0.504    0.547    0.151    0.130    0.116    0.022

rd-13

   31    309.3    0.150    0.072    0.480    0.314    0.170    0.128    0.104    0.042    0.139    0.047    0.340    0.467    0.157    0.134    0.110    0.012

rd-14

   9    93.5    0.226    0.118    0.521    0.473    0.218    0.191    0.177    0.099    0.191    0.094    0.489    0.525    0.213    0.171    0.120    0.094

rd-15

   3    25.0    0.175    0.063    0.357    0.240    0.206    0.171    0.143    0.115    0.171    0.047    0.275    0.327    0.227    0.159    0.140    0.110

rd-16

   16    140.7    0.141    0.066    0.469    0.285    0.172    0.139    0.115    0.000    0.131    0.059    0.453    0.334    0.159    0.124    0.100    0.000

rd-17

   16    164.7    0.206    0.131    0.634    0.445    0.276    0.195    0.108    0.001    0.190    0.103    0.538    0.521    0.255    0.193    0.101    0.001

rd-18

   15    143.1    0.188    0.052    0.278    0.269    0.226    0.180    0.150    0.112    0.174    0.046    0.266    0.374    0.201    0.162    0.138    0.064

rd-20

   12    101.5    0.211    0.087    0.412    0.362    0.273    0.201    0.137    0.109    0.176    0.060    0.344    0.461    0.218    0.163    0.118    0.087

rdhw-01

   12    124.8    0.157    0.070    0.448    0.261    0.206    0.141    0.110    0.026    0.128    0.054    0.419    0.314    0.169    0.114    0.092    0.002

rdhw-02

   4    41.0    0.167    0.070    0.423    0.246    0.216    0.160    0.110    0.102    0.159    0.050    0.314    0.311    0.198    0.143    0.116    0.100

 

 

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Visual comparison of drilling and estimated block grades within the 0.1 opt Au grade shells provides a validation on a localized basis. Two (2) examples are shown in Figure 11-9 and Figure 11-10.

 

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Figure 11-9: 426 Deposit Comparison of Composite and Block Grades

(Source: Practical Mining, 2025)

 

 

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Figure 11-10: Ruby Deeps Deposit Comparison of Composite and Block Grades 120450N

(Source: Practical Mining, 2025)

Drift analysis (swath plot) is a localized comparison of model and drilling grades. The drilling data and block model are sliced into a predefined width in the specified direction and the average grade of each variable contained in the slice is calculated. Results are displayed graphically. Model and drilling grades should track closely together. Drift analysis comparing block model Nearest Neighbor (NN) and Inverse Distance Weighted cubed (IDW3) grades to drilling grades is displayed in Figure 11-11 and Figure 11-12 for gold and silver respectively.

 

 

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Figure 11-11: Drift Analysis Gold

(Source: Practical Mining, 2025)

 

 

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Figure 11-12: Drift Analysis Silver

(Source: Practical Mining, 2025)

 

 

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11.2.7

Resource Classification

Individual block model blocks have been classified using the criteria given in Table 11-12. A minimum of two drillholes within the given distance are required to classify a block.

Table 11-12: Mineral Resource Classification Scheme

 

Class    Major (ft)    Semi (ft)    Minor (ft)   

Min.

Composites

  

Max.

Composites

   Composites
per DH

Measured

   40    40    40    3    12    2

Indicated

   100    100    100    3    12    2

Inferred

   300    300    300    3    12    2

 

11.2.8

Factors That May Affect Mineral Resources

Areas of uncertainty that may materially impact the mineral resource estimates include:

 

   

Changes to long term metal price assumptions.

 

   

Changes to the input values for mining, processing, and G&A costs to constrain the estimate.

 

   

Changes to local interpretations of mineralization geometry and continuity of mineralized Domains.

 

   

Changes to the density values applied to the mineralized zones.

 

   

Changes to metallurgical recovery assumptions.

 

   

Variations in geotechnical, hydrogeological and mining assumptions.

 

   

Changes to assumptions with an existing agreement or new agreements.

 

   

Changes to environmental, permitting, and social license assumptions.

 

   

Logistics of securing and moving adequate services, labor, and supplies could be affected by epidemics, pandemics and other public health crises.

 

11.2.9

Reasonable Prospects for Eventual Economic Extraction

S-K 1300 requires mineral resources demonstrate “Reasonable Prospects for Eventual Economic Extraction” (RPEEE). Stope optimizer software is well suited to meet this requirement. The software will produce stope designs that meet minimum minable geometric shapes that exceed the cutoff grade. These shapes will include necessary low grade or waste dilution included with the stope design.

Mineral resources are defined by a mining geometry consistent with the drift and fill or drift and bench mining methods chosen. The dimensions of a minimum minable stope cross section are 20 feet wide x 15 feet high. Individual stope lengths can vary from a minimum of 20 feet to a maximum of 100 feet.

 

11.2.10

Archimedes Underground Mineral Resource Statement

Uncertainties regarding sampling and drilling methods, data processing and handling, geological modeling, and estimation were incorporated into the classifications assigned.

A mineral resource must demonstrate Reasonable Prospects for Eventual Economic Extraction (RPEEE). This was accomplished using the Vulcan 11.1 Mine Stope Optimizer. The stope optimizer creates stope shapes meeting minimum predefined geometrical criteria and cutoff grade. Optimality, this is achieved when metal content is maximized while obeying the cutoff grade and geometrical criteria. Mineral resources in Table 11-13 are constrained by stopes measuring no less than 15 x 10 x 15 feet in width, length and height with an average undiluted grade of 0.174 Au opt.

Mineral resources are not mineral reserves and have not been demonstrated to have economic viability. There is no certainty that the mineral resource will be converted to mineral reserves. The quantity and grade

 

 

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or quality is an estimate and is rounded to reflect the fact that it is an approximation. Quantities may not sum due to rounding.

There is no guarantee that mineral resources can be converted to mineral reserves. Inferred mineral resources do not have sufficient confidence that modifying factors can be applied to convert them to mineral reserves.

Table 11-13: Summary of Archimedes Underground Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

       
 Deposit   

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

Indicated Mineral Resources
       

 426

   899    6.9    0.8    199    22
       

 Ruby Deeps

   892    8.3    2.4    237    69
       

 Total Indicated

   1,791    7.6    1.6    436    92
Inferred Mineral Resources
       

 426

   1,038    6.6    1.2    219    40
       

 Ruby Deeps

   3,150    7.6    2.4    769    246
       

 Total Inferred

   4,188    7.3    2.1    988    286

Notes:

 

  1.

Underground mineral resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of $27.25 per ounce (Section 16.1).

 

  2.

Mineral resources have been estimated using pressure oxidation gold metallurgical recoveries of 96.8% and 89.5% for the 426 and Ruby Deeps deposits respectively.

 

  3.

Pressure oxidation cutoff grades are 5.06 and 5.48 Au g/t (0.148 and 0.160 opt) for the 426 and Ruby Deeps deposits respectively.

 

  4.

Detailed input mining, processing, and G&A costs are defined in Section 18.1.

 

  5.

Units shown are metric.

 

  6.

The contained gold ounces estimates in the mineral resource table have not been adjusted for metallurgical recoveries.

 

  7.

Numbers have been rounded as required by reporting guidelines and may result in apparent summation differences.

 

  8.

A mineral resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

 

  9.

An inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration.

 

  10.

Mineral resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors.

 

  11.

Mineral resources have an effective date of December 31, 2024.

 

  12.

The reference point for mineral resources is in situ.

 

 

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11.2.11

QP Opinion

Practical Mining is not aware of any environmental, legal, title, taxation, socioeconomic, marketing, political, or other relevant factors that would materially affect the estimation of mineral resources that are not discussed in this Technical Report.

Practical Mining is of the opinion that the mineral resources for the Archimedes Underground Project, which were estimated using industry accepted practices, have been prepared and reported using S-K 1300 definitions.

Technical and economic parameters and assumptions applied to the mineral resource estimate are based on parameters received from i-80 and reviewed by Practical Mining to determine if they were appropriate.

The QP considers that all issues relating to all relevant technical and economic factors likely to influence the prospect of economic extraction can be resolved with further work.

 

11.3

Archimedes Open Pit

The Archimedes deposit area is physically separated from the Mineral Point deposit area and was treated independently in this report.

 

11.3.1

Summary Workflow

The mineral resource estimation workflow for the Archimedes open pit deposit area includes:

 

  1.

Data validation and loading into mining software system.

 

  2.

Exploratory data analysis to determine appropriate estimation domains and estimation parameters.

 

  3.

Use of an indicator shell at an 85% probability of grades being above 0.05 Au g/t to define an outer mineralized envelope.

 

  4.

Analysis of statistics and variography within the indicator domain envelope.

 

  5.

Grouping of rock units with similar statistical behavior into an estimation domain.

 

  6.

Variography and development of estimation parameters

 

  7.

Block model grade estimation.

 

  8.

Block model validation consisted of visual and statistical comparisons methods, including a review and comparison to the historical production.

 

  9.

Mineral resource classification into measured, indicated, and inferred mineral resources.

 

  10.

Economic analysis of resources remaining below the former pit limit to determine if there is reasonable prospects for eventual economic extraction.

 

  11.

Reporting of resource estimation results.

 

11.3.2

Exploratory Data Analysis (EDA)

After data loading and cleanup, summary statistics were run on each rock unit with a lower cutoff of 0.001 ppm (to avoid distortions from unmineralized material). The summary statistics are presented in Table 11-14. An initial statistical review of samples was confined to the principal host geologic units (primarily carbonates). There are additional units on the property, however they were not represented in the Archimedes area.

The alluvium and the tertiary volcanics were mostly unmineralized; and the Secret Canyon and Antelope Valley formations contained significantly less gold than the other units, thus they were excluded from the analysis and resource estimation.

 

 

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Table 11-14: Summary Sample Statistics - Archimedes

 

Code            Lith
Code
   Samples    Minimum    Maximum    Mean    Var.    Std.Dev.    CV
  3      Cambrian Secret Canyon Formation    Csc    452    0.0027    2.91    0.059    0.047    0.217    3.644
  4      Cambrian Hamburg Dolomite    Ch    1195    0.0024    51.12    0.400    2.943    1.716    4.285
  5      Cambrian Dunderberg Shale    Cd    1352    0.0024    40.29    1.421    14.014    3.744    2.634
  6      Cambrian Windfall Catlin Member    Cwc    2152    0.0015    29.97    0.976    6.363    2.523    2.583
  7      Cambrian Windfall Bullwhacker Member    Cwb    8518    0.0015    69.70    0.641    5.067    2.251    3.510
  8      Ordovician Lower Goodwin Member    Og1    18838    0.0015    52.40    0.354    2.625    1.620    4.583
  9      Ordovician Lower Laminated Goodwin Member    Ogll    10858    0.0015    83.47    0.953    12.173    3.489    3.660
  10      Ordovician Upper Goodwin Member    Og2    38659    0.0015    66.51    0.781    8.853    2.975    3.810
  11      Ordovician Ninemile Formation    On    7405    0.0015    59.89    0.408    6.922    2.631    6.443
  12      Ordovician Antelope Valley Formation    Oav    1854    0.0017    9.12    0.041    0.100    0.315    7.645
  13      Cretaceous Bullwhacker Sill    Kbs    1388    0.0024    21.84    0.768    3.363    1.834    2.389
  14      Tertiary Volcanics    Tv    68    0.003    0.22    0.028    0.001    0.031    1.139
  15      Quaternary Alluvium    Qal    2486    0.0015    1.74    0.013    0.003    0.055    4.353
  9999      Not Coded    Unk    15159    0.0015    81.67    0.366    2.580    1.606    4.383
  Total                110384    0.0015    83.47    0.607    6.397    2.529    4.170

A box and whisker plot of the logarithms of these grades is shown in Figure 11-13 supporting the exclusion of certain units from the analysis.

 

LOGO

Figure 11-13: Graphical Statistical Comparison of Rock Units

(Source: Forte Dynamics, 2025)

 

 

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Cumulative frequency plots of five (5) of the key geological units are shown is Figure 11-14. There is evidence of an overrepresentation of very low grades within the database, indicating that the rock units are not the mineralogical control.

 

LOGO

Figure 11-14: Statistics for Key Geological Units

(Source: Forte Dynamics, 2025)

A compositing study was performed to determine an appropriate composite length. This is an analysis of the increasing dilution and loss of variability incurred when combining drill hole samples into units of equal length for informing the resource estimate in an unbiased manner. The study results are shown in Figure 11-15. Grades will be very much diluted should a sample the length of a model block (25 ft) be used. There is an inflection in both curves at the 15ft. point, and it was determined that a 1⁄2 block composite of 12.5 ft. was appropriate for this estimate.

 

 

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Figure 11-15: Composite Study Results

(Source: Forte Dynamics, 2025)

The QP suspects that the precious metals may have followed a structural overprint of fractures spanning the geological units. From this it was determined to develop a shell through indicator kriging to confine the analysis to the mineralized geological units and to limit the spread of metals into country rock. From Figure 11-14 it was determined that a threshold value of 0.05 ppm Au would be appropriate to differentiate mineralized from non mineralized rock. Variograms were developed for samples within the indicator shell as shown in Table 11-15. The estimated indicator values were plotted over drill hole sections to determine an appropriate decision value. The 0.85 probability value was selected, and this shell is being considered as the mineralized domain.

Table 11-15: Variogram for 0.05 Au ppm Indicator

 

Nugget    Sill 1    Sill 2    Azimuth    Dip    Range 1    Range 2

0.17

   0.28    0.55    0    0    50    220
   90    0    25    220
   0    90    35    210

After limiting the composites to the indicator shell, cumulative frequency graphs were developed and are shown in Figure 11-16. Although the distributions are not perfectly log-normal, these are much improved in statistical behavior and were used for the gold and silver grade estimation. To avoid over projection of high-grade samples, the gold composites were capped at 15 g/t Au, and silver composites were capped at 200 g/t Ag.

 

 

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Figure 11-16: Gold and Silver Composite Samples within the Indicator Shell

(Source: Forte Dynamics, 2025)

 

11.3.3

Resource Estimation

Variograms were developed using the composites within the indicator shell. An example gold variogram is shown in Figure 11-17.

 

LOGO

Figure 11-17: Example Gold Variogram

(Source: Forte Dynamics, 2025)

 

 

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The variograms were modeled for both gold and silver within the mineralized domain as shown in Table 11-16.

Table 11-16: Variograms for Au and Ag

 

               

Gold
Au

   Nugget    Sill 1    Sill 2    Azimuth    Dip    Range 1    Range 2
   0.17    0.28    0.55    350    0    35    170
   80    80    35    170
   80    10    20    130

Silver
Ag

   Nugget    Sill 1    Sill 2    Azimuth    Dip    Range 1    Range 2
   0.23    0.76         0    0    95     
   90    0    95     
   0    90    80     

As the mineralized domain limit was based on an estimated value, and was well drilled, grade estimation parameters were limited to one variogram range and there were sufficient samples to estimate the volume at this range. A single pass search strategy was conducted using samples inside and outside of the indicator shell to estimate blocks.

Table 11-17: Gold and Silver Search Parameters

 

Gold
Au
   Azimuth    Dip    Range 2    min    max    Max/hole
   350    0    170    9    16    4
   80    80    170
   80    10    130
Silver
Ag
   Azimuth    Dip    Range 2    min    max    min
   0    0    220    9    16    4
   90    0    220
   0    90    210

 

11.3.4

Model Validation

Block model validation consisted of visual and statistical methods, including a comparison to the historical production. According to historical production records the Archimedes Pit produced about 22 million tons of ore at an average grade of about 2.29 g/t (0.067 opt). Testing the current model against the final mined topography gives an estimated mined resource of 21.4 million tons at about 2.25 g/t (0.066).

Numerous sections were reviewed and in general the estimated block grades compare well to the informing composite samples. Figure 11-18 shows an example cross section for the estimated block model and informing composites, including the reporting pit shell, current topo surface and the depleted topo surface, running SW-NE (100 ft window for composites).

 

 

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Figure 11-18: Cross Section of Estimated Block Model and Composites

(Source: Forte Dynamics, 2025)

The block model was checked for global bias by comparing the average grade of Ordinary Kriging (OK) to Nearest Neighbor (NN) at a zero grade g/t Au cutoff. The global bias was below 3% and considered acceptable and within the recommended Forte guidelines of 5%.

Local bias was reviewed using east-west swath plots to compare the estimate with the informing composite data, analyzing local trends. Two examples are shown through richly mineralized areas of the model. There are some slight differences between the OK and NN models grades, but it is within tolerance and considered normal.

 

 

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Figure 11-19: Example Swath Plots

(Source: Forte Dynamics, 2025)

 

 

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Figure 11-20 shows a statistical comparison of the distributions. As the volume of representative material increases from samples to composites to model blocks, the statistical variance decreases as shown in the change of slope. The capped composites and the estimated grades have similar means, and a lower mean value than the raw samples and uncapped composites.

 

LOGO

Figure 11-20: Comparison of Cumulative Frequency

(Source: Forte Dynamics, 2025)

 

 

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11.3.5

Mineral Resource Classification

The mineral resource was classified into indicated and inferred mineral resources (no measured resources). This was done using the average spacing of the closest three (3) drill holes to the block. Since the variogram models had been normalized (total sill =1.0) the distance at which the variogram reaches a proportion of the sill was chosen. This method was compared to the estimation of composite declustering weights and the sample density on drill hole bench intercept maps.

Table 11-18: Resource Classification by Sample Density

 

      % of Sill    Distance

Measured

   <50%    35 ft

Indicated

   <70%    80 ft

Inferred

   >70%    80 ft

 

11.3.6

Reasonable Prospects for Eventual Economic Extraction

The potential for economic extraction was determined by use of an economic pit limit program, MineFlow, from the Colorado School of Mines. This utilizes a unique algorithm, but provides similar results to a Lerchs-Grossman analysis. The economic parameters applied are equal to those used in the more detailed Mineral Point study shown in Table 13-13. The bulk of the surface minable Archimedes deposit has been mined previously, leaving about 5 million tonnes of ore and 300 thousand gold ounces within the optimized pit shell and below the depleted topo surface.

 

11.3.7

Archimedes Open Pit Mineral Resource Statement

Mineral resources are detailed in Table 11-19 for the Archimedes Open Pit mineral resource statement. Mineral resources are not Mineral Reserves and have not been demonstrated to have economic viability. There is no certainty that the mineral resource will be converted to mineral reserves. The quantity and grade or quality is an estimate and is rounded to reflect the fact that it is an approximation. Quantities may not sum due to rounding.

There is no guarantee that mineral resources can be converted to mineral reserves. Inferred mineral resources do not have sufficient confidence that modifying factors can be applied to convert them to mineral reserves.

 

 

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Table 11-19: Summary of Archimedes Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

             
Deposit   

Cutoff Au

(g/t)

  

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

   
Indicated Mineral Resources      
           

Archimedes Pit

   0.2    4,280    1.98    10.7    272    1,460
   0.1    4,320    1.96    10.6    272    1,490
   0.05    4,340    1.95    10.6    272    1,480
   
Inferred Mineral Resources      
           

Archimedes Pit

   0.2    820    1.18    8.9    31    230
   0.1    870    1.12    8.5    31    250
   0.05    880    1.11    8.5    31    250

Notes:

 

  1.

Mineral resources have an effective date of December 31, 2024.

 

  2.

Mineral resources are the portion of Mineral Point that can be mined profitably by open pit mining method and processed by heap leaching.

 

  3.

Mineral resources are below an updated topographic surface (below Archimedes pit).

 

  4.

Mineral resources are constrained to economic material inside a conceptual open pit shell. The main parameters for pit shell construction are a gold price of $2,175/oz Au, a silver price of $26.00/oz, average gold recovery of 77%, average silver recovery of 40%, open pit mining costs of $3.31/tonne, heap leach average processing costs of $3.47/tonne, general and administrative cost of $0.83/tonne processed, gold refining cost of $1.85/oz, silver refining cost of $0.50, and a 3% royalty (Section 16.1).

 

  5.

Mineral resources are reported above a 0.1 g/t Au cutoff grade. Silver revenues were not considered in the cutoff grade.

 

  6.

Mineral resources are stated as in situ.

 

  7.

Mineral resources have not been adjusted for metallurgical recoveries.

 

  8.

Reported units are metric tonnes.

 

  9.

Reported table numbers have been rounded as required by reporting guidelines and may result in summation discrepancies.

 

11.3.8

QP Opinion

The Archimedes Open Pit mineral resource has been estimated using core drill data using industry best practices, and have been prepared and reported under S-K 1300 definitions. Forte believes that the mineral resource estimate is of sufficient quality to support future exploration and mining related work, including future preliminary economic assessment level studies.

Forte is not aware of any other factors or issues not discussed in this technical report that may materially affect the mineral resource estimate other than normal risks faced by mining projects in terms of environmental, permitting, taxation, socioeconomic, marketing and political factors.

 

11.4

Mineral Point Open Pit

Forte Dynamics, Inc (Forte) reviewed the Mineral Point Open Pit mineral resource estimate completed by Wood in July 2021. The scope of the review included the informing drillhole and sample data, exploratory data analysis (EDA), input models (described below), and the current topography. The scope also included a review of the grade estimation methodology and model validation, bulk density determination, resource classification, reasonable prospects for eventual economic extraction (RPEEE), and the statement of mineral resources.

 

 

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Upon completion of the Mineral Point Open Pit resource review, Forte made some slight modifications to the Wood block model. Note that the estimated block grades were not altered or changed. Updates included updating the block model with the current topographic surface, recoding the Wood 2021 lithological model to the block model along with an assigned specific gravity (SG) values based on lithology code, and updated values and conversions for tonnage factor. Forte also used an updated pit shell to constrain and report the mineral resource under the requirements for RPEEE, which was based on a 2024 Scoping Study completed by Forte and used for other work completed in this Technical Report Summary.

No mineral resource depletion has occurred since the Wood 2021 Mineral Resource Estimate. The 2025 Mineral Point mineral resource estimate is comprised of indicated and inferred mineral resources and is presented in Table 11-23.

 

11.4.1

Summary Workflow

The mineral resource estimation workflow for the Mineral Point Trend consisted of three (3) steps:

 

  1.

Exploratory data analysis to understand grade trends and distributions and select an approach and parameters for grade estimation and density determination.

 

  2.

Estimation of block model grades.

 

  3.

Block model validation consisting of visual and statistical comparison methods.

 

  4.

Mineral resource classification into measured, indicated, and inferred mineral resources.

 

  5.

Economic analysis to determine if there are reasonable prospects for eventual economic extraction.

 

  6.

Reporting of resource estimation results.

Uncertainties regarding sampling and drilling methods, data processing and handling, geological modeling, and estimation were incorporated into the classifications assigned.

A mineral resource optimized Lerchs-Grosman (LG) pit shell was constructed to define the portion of the resource model having reasonable prospects for eventual economic extraction (RPEEE) amenable to open pit mining and run of mine heap leaching.

Classified mineral resources blocks were tabulated for above conceptual cut off grades inside the resource pit shell, and resource risks and opportunities were evaluated.

 

11.4.2

Geological Modeling

 

11.4.2.1

Structural Model

A structural model was developed for the Ruby Hill project by SRK (Uken, 2017a, 2017b). The structural model consists of a set of fault surfaces that offset lithological units and an assessment of fold geometry affecting the lithological units hosting gold mineralization. The model was developed from mapping in the open pit and analysis of blasthole, diamond drill and reverse circulation data. The main fault features are shown in Figure 11-21.

 

 

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Figure 11-21: Fence Section Looking North Showing Main Faults and Stratigraphic Units for the Ruby Hill Project

(Source: Wood, 2021)

 

11.4.2.2

Lithology Model

A lithology model consisting of the stratigraphic units hosting gold and base metal mineralization was constructed using the Project structural model faults and fold geometries along with geological logging from diamond drill and reverse circulation drilling to guide interpretation. Figure 11-21 shows the lithology model for the Ruby Hill Project.

 

11.4.2.3

Oxidation Model

An oxidation model was constructed consisting of wireframes interpreted using the logged oxide-sulfide codes and the ratio of cyanide soluble gold to total gold grade (AURAT). The oxidation model was coded to the block model to define sulfide and oxide blocks. An example cross section showing the modeled sulfide zone and Redox coding in the drillhole database running SW-NE is shown in Figure 11-22.

 

 

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Figure 11-22: Example Cross Section Showing Modeled Sulfide Domain and Redox Codes in the Drillhole Database

(Source: Forte Dynamics, 2025)

 

11.4.3

Exploratory Data Analysis

Exploratory data analysis (EDA) was carried out on raw assay samples and assay composites and included construction and review of histograms, probability plots, boxplots, visual review of spatial grade trends in three dimensions, and down hole and directional grade variography to develop the approach for grade estimation and generate parameters for interpolation. A summary of the EDA is presented here.

Visual assessment of gold grades at Mineral Point indicates that grades are moderate compared to the Archimedes Deposit, but on-strike and lateral continuity is good along the broadly folded Hamburg dolomite unit that hosts the majority of the mineralization at Mineral Point. Locally varying anisotropy, using the hanging wall surface of the Hamburg Dolomite to orient the strike and dip of anisotropic search ellipsoids, was identified as a good way to model the folded grade trend evident at Mineral Point.

Figure 11-23 shows a histogram and probability plot for gold and silver assay sample grades for the Mineral Point Trend constrained to the optimized LG pit shell used to report the mineral resource estimate in this section. The gold grade distribution is log-normal with a mean of 0.39 g/t Au and a median grade of 0.05 g/t Au with a long tail to a maximum grade of 128.5 g/t Au. The coefficient of variation (CV) of the gold assay grades is 6.5.

Based on an assessment of the relatively high variance of the assay grade distribution a 10-foot downhole composite length was selected to reduce variance of the majority 5-foot assay sample intervals. The CV of the gold composite grades is 5.6.

Figure 11-24 shows an example East-West cross section (looking North) through the West central part of the Mineral Point Trend showing raw assays (right of trace) and downhole 10 ft. composites (left of trace) with the optimized pit shell. Figure 11-25 shows a box and whisker plot for the raw assay sample grades and 10 ft. composites for gold and silver.

 

 

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To further manage the high variance of the gold grades the Probability Assigned Constrained Kriging (PACK) method was selected and indicator grade thresholds of 0.08 g/t Au and 1.0 g/t Au were selected to define low- and high-grade gold domains for the Mineral Point Trend.

An analysis of the high-grade assays was undertaken, and assay capping thresholds were selected to mitigate over projection of higher-grade samples.

Experimental correlograms were calculated using 10’ composites within the low- and high-grade domains. Down-hole variograms were used to define the nugget effect, and variogram maps were used to determine the directions of best continuity. Variograms were then modeled in the three primary directions.

EDA for silver grades indicated that although silver is not well correlated with gold grades, the grade distribution of silver is similar to that of gold and a similar approach would be suitable for silver grade estimation. Indicator grade thresholds of 4.0 g/t Ag and 40 g/t Ag were selected to define the low- and high-grade domains for silver.

 

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Figure 11-23: Gold and Silver Raw Assay Sample Grade Histograms and Probability Plots

(Source: Forte Dynamics, 2025)

 

 

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Figure 11-24: Example Cross Section of the Mineral Point Trend Showing Raw Assays (Right of Trace) and Downhole 10 ft. Composites (Left Trace) with the Optimized Pit Shell

(Source: Forte Dynamics, 2025)

 

 

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Figure 11-25: Box and Whisker Plot for Assay Sample Grades and 10 ft. Composites for Gold and Silver

(Source: Forte Dynamics, 2025)

 

11.4.4

Grade Estimation

Grade estimation was completed using the PACK methodology using Vulcan commercial mining software for gold and silver. A composite length of 10 feet and block size of 25 ft x 25 ft x 25 ft were selected to reduce sample variance and build models for open pit mining. The selected block size is consistent with the bench height and selectivity of historic mining in the Archimedes pits and the selectivity envisaged for future open pit mining.

Grades for the Mineral Point Trend were estimated into 25 ft x 25 ft x 25 ft blocks using 10 ft assay composites.

Based on an analysis of the Coefficient of Variation (CV) at a range of grade thresholds, thresholds of around 1.0 g/t Au and 40 g/t Ag were selected to define low- and high-grade domains for gold and silver. This threshold allowed reduction of the variance of composite grades within the two grade domains and enough samples to support estimation in both domains.

 

 

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Figure 11-26: Indicator Threshold Selection – CV of Gold and Silver Assay Composite Grades

(Source: Wood, 2021)

Low and high-grade indicators were estimated from the 10 ft composites using inverse distance weighting to the second power with a search of 500 ft x 500 ft x 50 ft, using a minimum of 6 samples, maximum of 15 samples and maximum of 3 samples per drillhole. Based on volumetric review comparing a Nearest Neighbor (NN) model of the high-grade indicator, an estimated indicator probability of 0.37 was selected as the probability threshold to define blocks for the high-grade domain. Estimated indicator probabilities in the block model were then back-flagged into the composites. Composites with back-flagged probabilities ≥ 0.37 were used to estimate blocks with an estimated indicator ≥ 0.37 for the high-grade domain. Composites with back-flagged indicator probabilities < 0.37 were used to estimate blocks in the low-grade domain. Gold and silver grades for blocks within the high-grade domain were interpolated using the estimation parameters shown in Table 11-20. Estimation search ellipse orientation is based on locally varying anisotropy (LVA) in which each block is assigned an orientation based on the tangent plane to the hanging wall contact of the Hamburg dolomite at the point nearest to the block centroid.

 

 

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Table 11-20: Estimation Parameters

 

Estimation Pass

   Min    Max   

Max

Per DH

   X Axis    Y Axis    Z Axis   

%

Estimated

LG Domain

1

   6    15    3    200    200    50.0    39

2

   6    15    3    300    300    75.0    34

3

   6    15    3    450    450    112.5    25

4

   1    15    3    600    600    150.0    2

HG Domain

1

   5    15    3    200    200    50.0    42

2

   5    15    3    300    300    75.0    24

3

   5    15    3    450    450    112.5    20

4

   1    15    2    600    600    150.0    14

A review of the grade tonnage curve and histograms for gold revealed an inflection at the 0.8 g/t Au indicator threshold. To soften the boundary between low- and high-grade domains, a mixing zone was applied by adjusting the composite selection allowed to estimate each domain. For Au estimates for Mineral Point the final gold grade estimate was based on allowing composites with a probability between 0 and 0.45 to estimate blocks in the LG domain and composites with a probability between 0.20 and 1 to estimate blocks in the HG domain. Figure 11-27 shows the reduction in the “valley” by applying this soft boundary.

 

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Figure 11-27: Au Estimation – Implementation of a Soft Boundary Between LG and HG Composites

(Source: Wood, 2021)

A review of the estimated Au grades noted a high-grade blow-out in a limited area with existing underground development, drilling and assaying by Eureka Corp. To constrain the blowout Wood created a small wireframe around the affected area and applied a local cap grade of 5.0 g/t Au to composites within this area (Figure 11-28).

 

 

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Figure 11-28: Area of Au High-Grade Blow-out and Eureka Corp Underground Drilling

(Source: Wood, 2021)

 

11.4.5

Resource Model Validation

Block model validation consisted of visual comparisons of the ordinary kriging (OK) estimated blocks vs the informing composites, statistical comparisons of the OK grade estimates to the nearest neighbor (NN), and swath plot spatial comparisons of the OK grade estimates to NN and IDW^2 to ensure grade trends were maintained.

 

11.4.5.1

Visual

Estimated block model grades and composite grades were visually examined in cross section, longitudinal sections, and plan views. In general, the composites and model blocks compared well. An example section for gold grades and estimated blocks are shown Figure 11-29.

 

 

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Figure 11-29: Estimated Block Grades and 10 Foot Composite Grades for Gold - Section 121200 N Looking N

(Source: Wood, 2021)

 

11.4.5.2

Global Bias

The block model was checked for global bias by comparing the average Au and Ag estimated OK grades (with no cut-off) to the Nearest Neighbor (NN) average estimates. The NN estimator produces a theoretically globally unbiased (declustered) estimate of the average value when no cut-off grade is applied and is a good basis for checking the performance of the different estimation methods. Global biases are within the recommended Forte guidelines of 5%. The comparison is summarized in Table 11-21.

Table 11-21: Global Bias Check within Indicated Resources

 

     Tons         Estimated Mean   

Relative Difference

(%)

Class    Element    (000s)    NN Mean    (OK)    (OK-NN)/NN

Indicated

  

Ag (g/t)

   183546    0.496    0.493    -0.6
    

Au (g/t)

        15.104    15.251    1.0

 

 

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11.4.5.3

Local Bias

Local bias checks for Au and Ag were performed within the mineralized envelope by creating and analyzing local trends in the grade estimates using swath plots as presented in Figure 11-30.

This was done by plotting the mean values from the NN estimate, the ID2 estimates and the OK estimates in east-west, north-south and vertical swaths or increments. Swath intervals are 100 feet in the easterly direction, 150 feet in the northerly direction, and 50 feet vertically. In the upper row of the swath plots, the red line represents the OK model grades, the blue line represents the ID2 model grades, and the black line represents the NN model grades. In the lower row of swath plots, the number of blocks contained in each swath is shown by the red, blue, and black lines. Because the NN model is declustered and the composites are not, the NN model is a better reference model to validate the OK resource model. Swath plots are for indicated blocks only. Both Au and Ag show good agreement, especially in areas supported by large numbers of blocks. There are some slight differences between the OK and NN models grades, but it is within tolerance and considered normal.

 

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Figure 11-30: Swath Plots – Gold – Indicated Blocks

(Source: Wood, 2021)

 

 

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11.4.6

Bulk Density

Bulk density was assigned to blocks based on lithology model using the median of the bulk density measurements for each unit (Figure 11-31). A specific gravity value of 2.0 was assigned to quaternary alluvium, and a default value of 2.6 was assigned to lithologies for which there were no bulk density measurements. Dimensionless specific gravity values were converted to Imperial density in short tons per cubic foot for tabulation of resources and Imperial tonnage was converted to Metric tonnes for reporting.

 

LOGO

Figure 11-31: Bulk Density Values by Lithology

(Source: Wood, 2021)

 

11.4.7

Mineral Resource Classification

Uncertainties regarding sampling and drilling methods, data processing and handling, geological modeling, and estimation were incorporated into the classifications assigned. The parameters evaluated in the development of confidence classification criteria include the quality of the data used for the estimate, input data spacing, continuity of geological features, and grade and geostatistical assessment of estimation error of forecast grade for quarterly and annual production volumes.

A geostatistical drillhole spacing study was carried out as part of the assessment of parameters for mineral resource classification for the mineral resource estimate. The drillhole spacing study used the gold grade variogram and the coefficient of variation (CV) of the assay composite database to calculate estimation error for forecasts of gold grade for quarterly and annual production volumes at a mining rate of 20 ktpd based on a range of drill patterns. The study indicated that based on the variance of the gold grades and their spatial continuity a 100 ft x 100 ft square pattern would allow estimates of quarterly production with an error of approximately ±15% at the 80th confidence interval, and a 200 ft x 200 ft grid would be required to produce estimates within ±15% at the 80th confidence interval for annual production volumes. A portion of the Mineral Point block model is estimated by drillholes spaced closely enough for measured classification but concerns about data quality for the legacy data caused a downgrade of confidence of this material, and all blocks within an average of 140 ft to the nearest three (3) drillholes were classified as indicated. Blocks estimated from drillholes from 140 ft to 500 ft were classified as inferred. A smoothing routine was run to reduce the number of small, isolated patches of measured and indicated blocks in areas of predominantly inferred classification and reduce the number of small, isolated islands of inferred blocks inside areas of predominantly indicated classification.

 

 

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11.4.8

Reasonable Prospects for Eventual Economic Extraction

A mineral resource optimized LG pit shell was constructed to define the portion of the Mineral Point resource having reasonable prospects for eventual economic extraction (RPEEE) amenable to open pit mining and processing by heap leaching using the 25 ft x 25 ft x 25 ft block model. Conceptual mining, processing and economic assumptions for the open pit resource shell are presented in Table 11-22. Open pit mineral resources contained within the pit shell are reported above a fixed cut-off grade of 0.1 g/t Au.

A cross section showing the extents of the Mineral Point Resource Pit (green) and the current topographic surface (white) is shown in Figure 11-32.

Table 11-22: Parameters for Mineral Resource Pit Shell Construction

 

 Parameter

     Unit    Value

Metals Price

Gold

   US$/toz    2,175

Silver

   US$/toz    26

Au Process Recovery

   %    77

Ag Process Recovery

   %    40

Mining Operating Cost

   US$/tonne    3.31

Processing Cost

   US$/tonne    3.47

G&A Cost

   US$/tonne processed    0.83

Royalty

   %    3.0%

Payable Metal

         

Gold

   %    99.90

Silver

   %    99.50

Treatment & Refining Cost - Gold

   US$/toz    1.85

Treatment & Refining Cost - Silver

   US$/toz    0.50

Overall Slope Angles (OSA)

         

Dumps

   degree    30.00

Alluvium

   degree    55.00

Sanded

   degree    45.00

Unsanded

   degree    45.00

Note: Au and Ag presented recoveries are weighted averages for all materials.

 

LOGO

Figure 11-32: Cross Section Showing the Mineral Point Resource, Resource Pit Shell, and Topo

(Source: Forte Dynamics, 2025)

Note: Blocks displayed above 0.1 g/t Au cutoff.

 

 

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11.4.9

Mineral Point Open Pit Mineral Resource Statement

The estimated tonnages and grades in the mineral resource estimate have not been adjusted for mining recovery and dilution. Contained metal estimates in the mineral resource statement table have not been adjusted for metallurgical recoveries.

Mineral resources are reported in Table 11-23 for open pit oxide heap leach at Mineral Point. Mineral resources are not Mineral Reserves and have not been demonstrated to have economic viability. There is no certainty that the mineral resource will be converted to mineral reserves. The quantity and grade or quality is an estimate and is rounded to reflect the fact that it is an approximation. Quantities may not sum due to rounding.

There is no guarantee that mineral resources can be converted to mineral reserves. Inferred mineral resources do not have sufficient confidence that modifying factors can be applied to convert them to mineral reserves.

Table 11-23: Summary of Mineral Point Open Pit Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

Deposit

  

Tonnes

(000)

  

Au

(g/t)

  

Ag

(g/t)

  

Au oz

(000)

  

Ag oz

(000)

Indicated Mineral Resources

Mineral Point

   216,982    0.48    15.0    3,376    104,332

Total Indicated

   216,982    0.48    15.0    3,376    104,332
 
Inferred Mineral Resources

Mineral Point

   194,442    0.34    14.6    2,117    91,473

Total Inferred

   194,442    0.34    14.6    2,117    91,473

Notes:

  1.

Mineral resources have an effective date of December 31, 2024.

 

  2.

Mineral resources are the portion of Mineral Point that can be mined profitably by open pit mining method and processed by heap leaching.

 

  3.

Mineral resources are below an updated topographic surface.

 

  4.

Mineral resources are constrained to economic material inside a conceptual open pit shell. The main parameters for pit shell construction are a gold price of $2,175/oz Au, a silver price of $26.00/oz, average gold recovery of 77%, average silver recovery of 40%, open pit mining costs of $3.31/tonne, heap leach average processing costs of $3.47/tonne, general and administrative cost of $0.83/tonne processed, gold refining cost of $1.85/oz, silver refining cost of $0.50, and a 3% royalty (Section 16.1).

 

  5.

Mineral resources are reported above a 0.1 g/t Au cutoff grade.

 

  6.

Mineral resources are stated as in situ.

 

  7.

Mineral resources have not been adjusted for metallurgical recoveries.

 

  8.

Reported units are metric tonnes.

 

  9.

Reported table numbers have been rounded as required by reporting guidelines and may result in summation discrepancies.

 

11.4.10

Factors that may Affect Mineral Resources

The QP notes the following points as factors that may materially affect the mineral resources.

 

   

Changes and/or updates to the geological model which was used to code lithology (rock) type to the block model.

 

   

Changes and/or updates to the specific gravity values based on lithology.

 

   

Changes to interpretation and grade continuity of resource domains.

 

   

Interpretation of oxidation-sulfide model which affects mining material type and destination.

 

 

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Interpretation of alteration type related to metallurgical recovery.

 

   

Changes to high-grade capping values used in the grade estimation.

 

   

Changes to input cost assumptions.

 

   

Changes in metallurgical testing results and subsequent recoveries.

 

   

Changes to other commonly uses resource estimation and mining assumptions.

 

11.4.11

QP Opinion

The Mineral Point open pit mineral resource has been estimated using core drill data using industry best practices, and have been prepared and reported under S-K 1300 definitions. Forte believes that the mineral resource estimate is of sufficient quality to support future exploration and mining related work, including future preliminary economic assessment level studies.

Forte is not aware of any other factors or issues not discussed in this technical report that may materially affect the mineral resource estimate other than normal risks faced by mining projects in terms of environmental, permitting, taxation, socioeconomic, marketing and political factors.

 

 

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12.

MINERAL RESERVE ESTIMATES

The Ruby Hill Project does not have any Mineral Reserves.

 

 

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13.

MINING METHODS

 

13.1

Archimedes Underground

 

13.1.1

Mine Development

Underground access will be through two portals located in the north wall of the Archimedes Pit adjacent to the pit haulage ramp. The main decline and portal will provide personnel and equipment access to all areas of the mine and will be 15 feet wide and 17 feet high. Decline gradient will not exceed +/- 13%.

Fresh air intake into the mine and secondary egress will be through a series of raises and drifts connecting to the main decline at logical intervals to promote efficient extraction. The intake portal will also be located in the north wall of the Archimedes Pit approximately 450 feet northwest and 140 feet above the main portal. Ventilation drifts will be 15 feet wide and 15 feet high. The first ventilation raise will be 590 feet in length and eight to ten feet in diameter. It will be excavated with a raise bore and lined with shotcrete or steel. This raise will be equipped with an unguided escape capsule that can be called remotely and operated from the underground station, thus not requiring a hoist operator.

The remaining raises will be excavated using raise bore or vertical crater retreat methods. They may be lined or unlined and also equipped with ladders and landings for egress. Optionally, a second smaller raise parallel to the first may be excavated and equipped for egress allowing for greater airflow (Figure 13-1).

 

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Figure 13-1: Archimedes Underground Isometric View Showing Portals, Main Ramp and Ventilation Development

(Source: Practical Mining, 2025)

 

 

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13.1.2

Mining Methods

Long hole open stoping (LHOS) with delayed backfill is the primary mining method planned for Ruby Hill. (Figure 13-2 and Figure 13-3) This will be widely supplemented with sill breasting from the lowest stope development drift. This allows access development to the lowest stope development drift to maintain a uniform elevation profile while extraction can adapt to varying mineralization boundaries. Stope development drifts will be 15 feet high and 15 to 20 feet wide. The gradient of stope development drifts will not exceed +/- 10%. LHOS widths will match stope development widths. Stope heights can vary from 30 to 60 feet from back to sill of the upper and lower stope development drifts. Sills can be up to 30 feet deep and the entry ramp radiant can be up to – 25% as it only need accommodate a loader which may be operated remotely.

The extraction and backfill sequence for a multi height LHOS panel with sill mining as shown in Figure 13-2 and Figure 13-3 is as follows:

 

  1.

Excavate the lowest and middle stope development drifts.

 

  2.

Excavate the sill below the lowest stope development drift.

 

  3.

Backfill the sill.

 

  4.

Excavate the first LHOS between the lower and middle stope development drifts. Stope lengths can be adjusted to accommodate stope wall stability conditions but have a practical upper limit of 100 to 150 feet.

 

  5.

Backfill the first stope.

 

  6.

(6-9) Excavate and backfill the remaining stopes on the level and drift the upper stope development drift.

 

  7.

Excavate the first stope on the next level.

 

  8.

Backfill the first stope and excavate and backfill any additional stopes.

 

  9.

Backfill the stope development drifts and begin development of the adjacent stope panel if present.

Drift and fill mining can be implemented when the mineralization geometry does not have sufficient vertical extent to allow LHOS or sill mining or where ground conditions will not maintain vertical stope walls. Underhand drift and fill mining is preferred since the backfill quality will be better than the rock quality.

 

 

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Figure 13-2: Stope Mining Sequence Part A

(Source: Practical Mining, 2025)

 

 

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Figure 13-3: Stope Mining Sequence Part B

(Source: Practical Mining, 2025)

 

13.1.3

Geotechnical and Ground Support

 

13.1.3.1

Rock Quality Designation (RQD)

Rock Quality Designation is one of the simplest methods of rock mass classification. The RQD number is the percentage of the sum of the length of core pieces whose length is greater than twice the core diameter divided by the total interval length. Drawbacks to the RQD method are that it does not include any information on the rock jointing surfaces, joint filling material, joint orientation and rock strength. RQD numbers will vary depending on the orientation of the drill hole to the prominent jointing. Tunnel support recommendations based on reviews of tunnels constructed in the US prior to 1969 are presented in Table 13-1 (Deere 1969).

 

 

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Table 13-1: Guidelines for the Selection of Primary Support for 20-foot to 40-foot Tunnels in Rock

 

     Support System

RQD

   Steel Sets    Rock Bolts    Shotcrete

Excellent >90

   None to occasional light sets    None to occasional    None to occasional on 2 – 3 inches crown

Good 75 - 90

   Light Sets 5-6 feet c-c    5-6 feet c-c w/mesh of straps as required    Local, 2-3 inches on crown

Fair 50 - 75

   Light to Medium Sets 4-5 feet c-c    3-5 feet c-c with mesh or straps as required    4 inches or more, crown and sides with possible bolts

Poor 25 - 50

   Medium to Heavy Sets 2-4 feet c-c    2-4 feet c-c with mesh or straps, resin anchors may be required    6 inches or more crown and sides, rock bolts as required 4-6 feet c-c

Very Poor < 25

   Medium to Heavy 2 feet c-c    3 feet c-c, 100% mesh or straps required, resin anchors may be required    6 inches or more on whole section, medium to heavy sets as required

Very Poor Squeezing or Swelling

   Very Heavy 2 feet c-c    2-3 feet c-c, 100% mesh or straps required, resin anchors may be required    6 inches or more on whole section, heavy sets as required

During the 2021 and 2022 drill campaigns, i-80 geologists logged 31 drill holes using the RQD method. The locations and logged values are shown in Figure 13-4 and Figure 13-5.

 

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Figure 13-4: RQD Logged Drill Holes (426 - Turquoise, Ruby Deeps - Gold)

(Source: Practical Mining, 2025)

 

 

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Figure 13-5: Cross Section 119625N Showing RQD Values (426 - Turquoise, 426 Fault - Gray, Ruby Deeps - Gold, Holly Fault - Red)

(Source: Practical Mining, 2025)

RQD logging results by mineralized zone are shown graphically in Figure 13-6 and summary statistics in Table 13-2. The majority of RQD logging is classified as Poor (25 – 50) or Very Poor (< 25). Support recommendations would be rock bolts on 2 – 4-foot centers with wire mesh. The recommended bolt length is 1/3 to 1/4 the span.

 

 

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Figure 13-6: RQD Box and Whisker Plot

(Source: Practical Mining, 2025)

Table 13-2: RQD Univariate Statistics by Grade Shell

 

Grade Shell

   426_10     426_07     426_06     426_05     426_04     426_03     426_02     rd_01     rd_08     rd_09     rd_11 

Count

   20    19    27    15    24    14    40    108    5    23    75

Length

   121    113    166    76    148    86    239    730    35    144    489

Std Dev

   14.46    30.31    29.63    18.28    21.11    17.73    14.66    28.83    0.00    20.41    23.81

Lower 95% CI

   6.0    20.0    12.7    6.3    11.4    8.7    8.9    35.0    0.0    5.9    38.9

Average

   12.3    33.7    23.9    15.5    19.8    18.0    13.5    40.5    0.0    14.3    44.3

Upper 95% CI

   18.7    47.3    35.0    24.8    28.3    27.3    18.0    45.9    0.0    22.6    49.7

Minimum

   0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0

25% Quartile

   0.0    3.5    0.0    0.0    0.0    5.7    0.0    14.6    0.0    0.0    26.0

Median

   8.2    35.0    13.7    9.4    9.7    14.2    8.6    43.5    0.0    0.0    47.0

75% Quartile

   17.5    55.5    32.0    22.3    40.0    31.8    24.7    63.4    0.0    31.5    61.5

Maximum

   50.0    88.4    96.9    57.3    57.9    54.0    48.3    92.0    0.0    54.4    82.0

Grade Shell

   rd_12    rd_13    rd_14    rd_16    rd_17    rd_18    rd_19     rd_20     rd_hw1     rd_hw2     low

Count

   21    32    5    18    19    11    17    7    5    3    1123

Length

   85    169    34    100    97    69    50    64    26    19    7170

Std Dev

   28.25    26.44    0.00    15.31    23.91    31.96    19.69    6.88    0.00    25.38    25.32

Lower 95% CI

   6.8    20.9    0.0    14.2    26.0    46.7    2.3    70.5    0.0    21.0    26.4

Average

   18.9    30.1    0.0    21.3    36.8    65.5    11.7    75.6    0.0    49.8    27.9

Upper 95% CI

   30.9    39.2    0.0    28.3    47.5    84.4    21.1    80.7    0.0    78.5    29.4

Minimum

   0.0    0.0    0.0    0.0    0.0    10.6    0.0    65.0    0.0    25.0    0.0

25% Quartile

   0.0    3.8    0.0    7.4    20.0    39.3    0.0    71.7    0.0    36.8    0.0

Median

   0.0    23.5    0.0    22.1    37.0    82.0    0.0    79.0    0.0    48.6    24.0

75% Quartile

   27.5    49.5    0.0    35.9    58.8    88.0    17.1    79.5    0.0    62.1    45.5

Maximum

   92.0    81.0    0.0    45.0    73.0    105.3    66.7    83.0    0.0    75.7    104.4

 

 

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13.1.3.2

Q-system

The Q-system was developed in 1974 by Barton, Lien and Lund of the Norwegian Geotechnical Institute. It was updated in 1993 and 2002 to include advances in ground support fixtures and shotcrete. The support chart (Figure 13-7) is based on the analysis of over 2,000 Scandinavian and Indian case studies. The Q-value gives a description of the rock mass stability of an underground opening in jointed rock masses. High Q-values indicate good stability and low values mean poor stability. Based on 6 parameters the Q-value is calculated using the following equation:

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The six parameters are:

 

   

RQD - Rock Quality Designation

   

Jn – Joint set number

   

Jr – Joint roughness number

   

Ja – Joint alteration number

   

Jw – Joint water reduction factor

   

SRF – Stress Reduction Factor

Individual parameters are determined during logging or mapping using tables that give numerical values. (Norwegian Geotechnical Institute 2022)

The Q-system also introduces a factor for Excavation Support Ratio (ESR). ESR numbers range from 0.5 for very long-lived strategic excavation to 1.6 for permanent mine openings and 3.5 for temporary mine openings (Norwegian Geotechnical Institute 2022).

 

 

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Figure 13-7: Q-system Support Recommendations

(Source: Norwegian Geotechnical Institute, 2022)

During the 2021 and 2022 drill program, i-80 engaged Call and Nicholas to log 19 drillholes using the Q-system. Drill hole traces and logging results are shown in Figure 13-8 and Figure 13-9.

 

 

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Figure 13-8: Q Logged Drill Holes (426 - Blue, Ruby Deeps - Gold)

(Source: Practical Mining, 2025)

 

 

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Figure 13-9: Cross Section 119625N Showing Q Values (426 - Blue, 426 Fault - Gray, Ruby Deeps - Gold, Holly Fault - Red)

(Source: Practical Mining, 2025)

Q-system logging results by mineralized zone are displayed in Figure 13-10 and summary statistics in Table 13-3. All but a few areas have Q values in the zero to ten range. For production excavations the span to ESR ratio is 1.7 and 3.1 for main development. From Figure 13-7, both excavation categories lie in category three or four. Category three recommended support consists of systematic bolting and five to six centimeters of shotcrete. Category 4 recommendations increase the shotcrete thickness to 6-9 centimeters.

 

 

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Figure 13-10: Q Value Box and Whisker Plot

(Source: Practical Mining, 2025)

Table 13-3: Q Value Univariate Statistics by Grade Shell

 

Grade Shell

   426_02     426_03     426_05     426_06     426_07     426_10     426_04     rd_01     rd_08     rd_09  

Count

   26    7    10    24    13    3    15    40    5    4

Length

   163    53    56    150    83    14    97    298    35    28

Std Dev

   19.89    3.52    2.77    7.24    8.05    3.56    0.81    5.74    0.11    0.00

Lower 95% CI

   -9.6    4.0    1.8    4.5    6.2    0.0    0.2    4.1    0.0    0.0

Average

   -2.0    6.6    3.5    7.4    10.6    4.1    0.6    5.8    0.1    0.0

Upper 95% CI

   5.7    9.2    5.2    10.3    15.0    8.1    1.0    7.6    0.2    0.0

Minimum

   -99.0    1.5    0.8    0.1    0.2    1.0    0.1    0.2    0.1    0.0

25% Quartile

   0.1    5.1    1.2    0.7    3.4    2.1    0.1    1.2    0.1    0.0

Median

   0.9    6.3    2.8    7.1    11.0    3.2    0.3    4.3    0.1    0.0

75% Quartile

   2.8    7.7    5.3    11.7    16.0    5.6    0.6    8.9    0.1    0.0

Maximum

   7.4    12.9    8.6    26.5    25.6    8.0    2.5    20.8    0.3    0.0

Grade Shell

   rd_11    rd_12    rd_13    rd_14    rd_16    rd_17    rd_18    rd_20    rd_hw2    low

Count

   44    16    16    5    12    13    7    6    3    566

Length

   330    55    72    34    65    64    54    60    19    3936

Std Dev

   20.18    25.73    5.54    5.01    2.21    4.93    5.30    5.32    1.71    7.47

Lower 95% CI

   -0.2    -16.4    1.1    0.4    1.1    0.6    14.6    6.9    4.8    5.4

Average

   5.8    -3.8    3.8    4.8    2.3    3.3    18.5    11.1    6.7    6.0

Upper 95% CI

   11.7    8.8    6.6    9.2    3.6    6.0    22.4    15.4    8.7    6.6

Minimum

   -99.0    -99.0    0.1    0.0    0.3    0.2    12.0    4.8    4.8    0.0

25% Quartile

   0.2    0.1    0.4    0.8    0.6    0.8    14.8    6.8    6.0    0.8

Median

   5.3    0.1    2.3    2.8    1.9    1.1    18.4    10.9    7.1    3.3

75% Quartile

   9.3    3.1    3.5    9.9    2.9    1.8    21.4    15.9    7.7    8.5

Maximum

   70.7    12.5    20.7    10.4    7.7    17.1    26.6    17.1    8.2    70.7

 

 

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13.1.3.3

Ground Support Requirements

Both the RQD and Q-system classifications obtained from Ruby Hill drill core logging fall within typical ranges seen at northern Nevada mines. Support requirements anticipated for primary development excavation entail 8-foot Swellex ® rock bolts with welded wire mesh and shotcrete installed to within five feet of the sill. Large intersections can be supplemented with longer Swellex bolts and/or fully grouted cable bolts. Support requirements in production excavations are largely the same, however with only spot shotcrete application. Excavations under backfill may require only spot bolting.

13.1.4  Cemented Rock Fill

The aggregate for Cemented Rock Fill (CRF) will be sourced from previously mined open pit waste. Potential sources should have minimal amounts of clay present. Aggregate will be crushed onsite to 100% passing a two-inch screen and will contain fine fractions similar to that of commercial concrete aggregates. Backfill will be mixed onsite with 5-8% type II Portland Cement and transported underground on the return leg of a haul truck cycle. An LHD fitted with a “Jammer” boom will push the material into place ensuing all voids are filled tightly.

During each shift of backfilling operations, concrete test cylinders will be collected for uniaxial compression testing. When test results are below design strengths these areas will be mined with additional bolting and shotcrete for support.

13.1.5  Staffing and Underground Equipment Requirements

Four crews will work a rotating schedule and operate the mine two 12-hour shifts per day. Multiple heading drift advance rates up to 100 feet per day is possible with the crew size and equipment configuration provided. Stope production up to 500 tons per day per stope can offset some of the drift advance when loading and trucking equipment requirements exceed availability. Backfill placement will be at rates to sustain production rates of 1,500 tons of mineralized material per day. Anticipated i-80 and contractor staffing levels for the Archimedes Underground are combined in Table 13-4.

Table 13-4: Personnel Requirements

 

   

Position

   Headcount  
   

Miners

   96
   

Maintenance

   24
   

Production Forman

   4
   

Maintenance Forman and Planner

   2
   

Mine Superintendent

   1
   

Maintenance Superintendent

   1
   

Surveyors

   2
   

Geologist

   6
   

Engineers

   2
   

Manager

   1
   

Total

   139

Note: Includes Contractor Personnel

The underground contractor will provide the equipment necessary for execution of the work. Table 13-5 lists the type and number of each type of equipment necessary to meet the production and development schedule shown in Table 13-7 and Table 13-8. This list is typical of northern Nevada underground mines of similar size and scope.

 

 

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Table 13-5: Equipment Requirements

 

   

Description

   Number of Units  
   

2-Boom Face Jumbo

   2
   

Rock Bolter

   3
   

Production Drill

   1
   

RC Drill

   1
   

Explosives Truck

   1
   

6-yd3 LHD

   3
   

6-yd3 LHD with Jammer

   2
   

30-ton Haul Trucks

   5
   

Water Truck

   1
   

Road Grader

   1
   

Shotcrete Sprayer

   1
   

Shotcrete Remix Trucks

   2
   

Scissor Deck Truck

   1
   

Forklifts

   3
   

Fuel and Lube Truck

   1
   

Backfill Plant

   1
   

Shotcrete Plant

   1
   

Personnel Transport Tractors

   5

i-80 currently has onsite the surface support equipment listed in Table 13-6 for maintaining the surface roads and stockpiles.

Table 13-6: i-80 Support Equipment

 

         

Make

   Model       Description    Condition      Hours  
         

Cat

   D9R    D9R Dozer    fair    57,503
         

Cat

   D10R    D10R Dozer    fair    11,657
         

Cat

   992C    992C Loader (Not in Service)    poor    36,330
         

Cat

   980G    980G Loader    fair    28,397
         

Komatsu

   PC300    Excavator    fair    10,657
         

Cat

   235C    Excavator    fair    5,631
         

Cat

   14H    14H Blade    fair    35,607
         

Cat

   IT28    IT28 Loader    fair    23,299
         

Cat

   785C    785C Haul Truck    Poor    48,681
         

Sterling

   LT 7501    Water Truck    Average    5,135

13.1.6 Mine Plan

Initial mining within the Goodwin Formation will be governed by an amendment to the Ruby Hill Plan of Operations (POO). This amendment will be part of an Environmental Assessment (EA) that has been initiated, and completion is anticipated in Q2 of 2025. Construction of the Portals and Underground development to the 426 deposit will commence on approval of the EA and POO amendment. Concurrent to construction, a second EA to amend the POO and permit mining in the Windfall Formation will be initiated. It is anticipated that approval of the second POO amendment will take 18 months at which time development and mining of the Ruby Deeps deposit can begin. (Figure 13-11) Mining of mineralization above the 5,100 elevation will be authorized under the EA and will continue while the second POO amendment is being processed. Development below the 5100 elevation will begin three months after receipt of the second POO amendment.

 

 

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Figure 13-11: Permitting Development and Initial Production Schedule

(Source: Practical Mining, 2025)

Development footage for the haulage decline and ventilation excavations are listed in Table 13-7. Also tabulated is the waste tonnage from expensed crosscuts connecting the haulage decline to the stopes. Expensed and capitalized waste mining totals 2.4M tons with waste mining peaking at 1,150 to 1,00 tons per day in 2028 and 2039. This corresponds to the development push into the Ruby Deeps mineralized zone.

Table 13-7: Ruby Hill Development Schedule

 

2025 2026 2027 2028 2029 2030

Capital Development

 Primary Drifting (feet)

2,206.4 6,863.9 2,854.1 6,893.1 6,822.9 -

 Secondary Drifting (feet)

300.0 3,336.2 2,594.5 5,163.3 4,010.3 1,288.1

 Raising (feet)

694.7 139.6 472.9 241.2

 Capital Waste (ktons)

52.4 214.1 110.8 246.8 221.7 23.8

Expensed Waste (ktons)

- 12.8 106.3 177.0 213.7 292.1

Total Waste (ktons)

52.4 226.9 217.0 423.8 435.5 316.0

Waste Mining Rate (tons/day)

285 622 595 1,158 1,193 866
2031 2032 2033 2034 2035  - 2036 Total

Capital Development

 Primary Drifting (feet)

- - - - - 25,640

 Secondary Drifting (feet)

1,841.7 1,505.7 1,205.0 - - 21,245

 Raising (feet)

- - 1,548

 Capital Waste (ktons)

34.8 27.1 24.1 - - 955.6

Expensed Waste (ktons)

136.2 281.1 93.3 116.9 62.9 1,492.3

Total Waste (ktons)

171.0 308.2 117.4 116.9 62.9 2,447.9

Waste Mining Rate (tons/day)

469 842 322 320 265 583

The Archimedes Life of Mine (“LOM”) production plan shown in Table 13-8 was extrapolated using the mining rates listed Table 13-9. These rates are typical of those used at similar Nevada underground mines.

The production mining processing schedules presented below contain 69% inferred mineral resources. The confidence in inferred mineral resources is considered too low to be converted to mineral reserves and there is no guarantee that they will be upgraded to measured or indicated mineral resources.

 

 

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Table 13-8: Archimedes Production Mining Plan (Includes Inferred Mineral Resource)

 

2025 2026 2027 2028 2029 2030

Production Mining

 Stope Development (ktons)

- 18.3 150.9 217.9 282.3 309.3

 Stope Mining (ktons)

- 21.3 117.7 283.0 320.6 276.3

 Cemented Rock Fill (ktons)

- 35.3 164.1 419.1 507.2 528.1

Production Mining (tons/day)

- 79 1,546 2,149 2,031 2,124

Total Mining Rate (tons/day)

285 748 1,376 2,583 2,923 2,477
2031 2032 2033 2034 2035-2036 Total

Production Mining

 Stope Development (ktons)

285.6 234.4 225.5 232.1 200.9 2,157.2

 Stope Mining (ktons)

330.0 384.5 393.8 371.1 240.5 2,738.8

 Cemented Rock Fill (ktons)

451.3 540.0 435.0 446.7 318.3 3,845.2

Production Mining (tons/day)

1,688 1,691 1,731 1,720 640 1,380

Total Mining Rate (tons/day)

2,156 2,533 2,052 2,040 727 1781

Table 13-9: Mine Production Rates by Excavation Type

 

     

Type of Excavation

   Mining Rate     Units   
     

Primary Development 15 x 17 ft.

   10    ft./day
     

Secondary Development 15 x 15 ft.

   8    ft./day
     

Expensed Waste Crosscuts 15x15 ft.

   8    ft./day
     

Stope Development Drift 15x20 ft.

   8    ft./day
     

Longhole Stope or Bench

   500    tons/day
     

Cemented Backfill

   400    tons/day

The processing schedule for oxide and refractory mineralization is shown in Table 13-10. Stockpiling of material for processing in a later year is not anticipated at any time during the Archimedes LOM. The third-party facility will purchase up to 1,000 tons/day of refractory mineralization through 2027. The combined production rate during the time for all i-80 mines operations is not planned to exceed 1,000 tpd. Likewise, the capacity of the Lone Tree refractory facility is planned for 2,500 tpd and the production during that time from all i-80 mining operations is not expected to exceed Lone Tree’s capacity.

The production mining and processing schedules presented herein contain 69% inferred mineral resources. The confidence in inferred mineral resources is considered too low to be converted to mineral reserves and there is no guarantee that they will be upgraded to measured or indicated mineral resources.

Table 13-11 presents a processing schedule that excludes inferred mineral resources. This schedule is a factorization of the schedule that includes inferred mineral resources and does reflect any changes in mine design or adjustment to capital development. Likewise, there has not been a recalculation of capital or operating unit costs.

 

 

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Table 13-10: Ruby Hill Processing Plan (Includes Inferred Mineral Resource)

 

20251   20261   20271   20282   2029 2030  

Refractory High Grade (ktons)

0 26 199 390 540 541

Au Grade (opt)

0.000 0.254 0.236 0.229 0.217 0.221

Ag Grade (opt

0.000 0.042 0.035 0.023 0.029 0.036

Au Contained (koz)

0 7 47 89 117 119

Ag Contained (koz)

0 1 7 9 16 19

Refractory Low Grade (ktons)

0 3 25 49 64 47

Au Grade (opt)

0.000 0.099 0.095 0.103 0.103 0.098

Ag Grade (opt

0.000 0.032 0.029 0.017 0.021 0.035

Au Contained (koz)

0 0 2 5 7 5

Ag Contained (koz)

0 0 1 1 1 2

Refractory (ktons)

0 29 224 439 604 587

Au Grade (opt)

0.000 0.236 0.220 0.215 0.205 0.211

Ag Grade (opt

0.000 0.041 0.034 0.022 0.028 0.036

Au Contained (koz)

0 7 49 94 124 124

Ag Contained (koz)

0 1 8 10 17 21

Au Recovered (koz)

0 4 29 89 117 119

Ag Recovered (koz)

0 0 1 1 2 2

Au Recovery

- 58% 58% 95% 94% 96%

Ag Recovery

- 10% 10% 10% 10% 10%

Refractory Throughput (tpd)

0 80 613 1203 1655 1609

Heap Leach (ktons)

0.0 16.9 61.2 82.2 27.4 0.7

Au Grade (opt)

0.000 0.117 0.106 0.110 0.127 0.067

Ag Grade (opt

0.000 0.013 0.015 0.011 0.010 0.044

Au Contained (koz)

0.0 2.0 6.5 9.0 3.5 0.0

Ag Contained (koz)

0.0 0.2 0.9 0.9 0.3 0.0

Au Recovered (koz)

0.0 1.7 5.7 7.9 3.0 0.0

Ag Recovered (koz)

0.0 0.0 0.1 0.2 0.1 0.0

Au Recovery

#DIV/0! 88% 87% 87% 87% 84%

Ag Recovery

#DIV/0! 18% 16% 20% 25% 18%

Leach Stacking Rate (ton/day)

0 46 168 225 75 2
2031 2032 2033 2034 2035  - 2036 Total

Refractory High Grade (ktons)

540.3 540.9 540.0 540.6 408.3 4265.8

Au Grade (opt)

0.215 0.210 0.233 0.237 0.216 0.223

Ag Grade (opt

0.065 0.067 0.073 0.064 0.051 0.051

Au Contained (koz)

116.3 113.6 125.7 127.9 88.1 950.8

Ag Contained (koz)

35.2 36.3 39.5 34.5 20.8 218.4

Refractory Low Grade (ktons)

75.7 77.9 91.7 87.1 59.3 579.9

Au Grade (opt)

0.100 0.105 0.105 0.105 0.103 0.103

Ag Grade (opt

0.045 0.046 0.041 0.038 0.035 0.036

Au Contained (koz)

7.5 8.2 9.6 9.1 6.1 59.5

Ag Contained (koz)

3.4 3.6 3.7 3.3 2.1 20.8

Refractory (ktons)

616.0 618.8 631.7 627.7 467.7 4845.7

Au Grade (opt)

0.201 0.197 0.214 0.218 0.201 0.209

Ag Grade (opt

0.063 0.064 0.068 0.060 0.049 0.049

Au Contained (koz)

123.8 121.8 135.3 137.0 94.2 1010.3

Ag Contained (koz)

38.6 39.9 43.2 37.8 22.9 239.2

Au Recovered (koz)

114.5 109.0 121.1 122.6 84.3 909.6

Ag Recovered (koz)

3.9 4.0 4.3 3.8 2.3 23.9

Au Recovery

93% 90% 90% 90% 90% 90%

Ag Recovery

10% 10% 10% 10% 10% 10%

Refractory Throughput (tpd)

1,688 1,695 1,731 1,720 1,281 1328

Heap Leach (ktons)

- - - -- - 188.4

Au Grade (opt)

- - - - - 0.111

 

 

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20251   20261   20271   20282   2029   2030  

Ag Grade (opt

- - - - - 0.012

Au Contained (koz)

- - - - - 21.0

Ag Contained (koz)

- - - - - 2.3

Au Recovered (koz)

- - - - - 18.3

Au Recovery

0.0% - - - - 0.4

Ag Recovery

0.0% - - - - 87%

Leach Stacking Rate (ton/day)

- - - - - 19%

Notes:

  1.

All refractory mineralization sold to a third-party processing facility in the years 2025 through 2027.

  2.

Beginning in 2028 refractory mineralization will be processed at i-80’s Lone Tree facility.

Table 13-11: Ruby Hill Processing Plan (Without Inferred Mineral Resource)

 

20251   20261   20271   20282   2029   2030  

Refractory High Grade (ktons)

2025 2026 2027 2028 2029 2030

Au Grade (opt)

0.0 7.8 59.5 116.9 161.7 161.9

Ag Grade (opt

0.000 0.254 0.236 0.229 0.217 0.221

Au Containd (koz)

0.000 0.042 0.035 0.023 0.029 0.036

Ag Contained (koz)

0.0 2.0 14.0 26.7 35.1 35.7

Refractory Low Grade (ktons)

0.0 1.0 7.5 14.6 19.2 14.0

Au Grade (opt)

0.000 0.099 0.095 0.103 0.103 0.098

Ag Grade (opt

0.000 0.032 0.029 0.017 0.021 0.035

Au Containd (koz)

0.0 0.1 0.7 1.5 2.0 1.4

Ag Contained (koz)

0.0 0.0 0.2 0.2 0.4 0.5

Refractory (ktons)

0 8.7 67.1 131.6 181.0 175.9

Au Grade (opt)

0.000 0.236 0.220 0.215 0.205 0.211

Ag Grade (opt

0.000 0.041 0.034 0.022 0.028 0.036

Au Containd (koz)

0 2.1 14.8 28.2 37.1 37.1

Ag Contained (koz)

0 0.4 2.3 2.9 5.1 6.3

Au Recovered (koz)

0 1.2 8.6 26.7 35.0 35.8

Ag Recovered (koz)

0 0.0 0.2 0.3 0.5 0.6

Au Recovery

- 58% 58% 95% 94% 96%

Ag Recovery

- 10% 10% 10% 10% 10%

Refractory Throughput (tpd)

0 24 184 360 496 482

Heap Leach (ktons)

0.0 5.1 18.3 24.6 8.2 0.2

Au Grade (opt)

0.000 0.117 0.106 0.110 0.127 0.067

Ag Grade (opt

0.000 0.013 0.015 0.011 0.010 0.044

Au Containd (koz)

0.0 0.6 1.9 2.7 1.0 0.0

Ag Contained (koz)

0.0 0.1 0.3 0.3 0.1 0.0

Au Recovered (koz)

0.0 0.5 1.7 2.4 0.9 0.0

Ag Recovered (koz)

0.0 0.0 0.0 0.1 0.0 0.0

Au Recovery

- 88% 87% 87% 87% 84%

Ag Recovery

- 18% 16% 20% 25% 18%

Leach Stacking Rate (ton/day)

- 14 50 67 22 1
2031 2032 2033 2034 2035 -  2036 Total

Refractory High Grade (ktons)

161.8 162.0 161.7 161.9 122.3 1277.7

Au Grade (opt)

0.215 0.210 0.233 0.237 0.216 0.223

Ag Grade (opt

0.065 0.067 0.073 0.064 0.051 0.051

Au Containd (koz)

34.8 34.0 37.7 38.3 26.4 284.8

Ag Contained (koz)

10.5 10.9 11.8 10.3 6.2 65.4

Refractory Low Grade (ktons)

22.7 23.3 27.5 26.1 17.8 173.7

Au Grade (opt)

0.100 0.105 0.105 0.105 0.103 0.103

Ag Grade (opt

0.045 0.046 0.041 0.038 0.035 0.036

Au Containd (koz)

2.3 2.5 2.9 2.7 1.8 17.8

 

 

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  20251     20261     20271     20282     2029   2030  

Ag Contained (koz)

  1.0   1.1   1.1   1.0   0.6   6.2

Refractory (ktons)

  184.5   185.4   189.2   188.0   140.1   1451.4

Au Grade (opt)

  0.201   0.197   0.214   0.218   0.201   0.209

Ag Grade (opt

  0.063   0.064   0.068   0.060   0.049   0.049

Au Containd (koz)

  37.1   36.5   40.5   41.0   28.2   302.6

Ag Contained (koz)

  11.6   11.9   12.9   11.3   6.9   71.6

Au Recovered (koz)

  34.3   32.7   36.3   36.7   25.2   272.4

Ag Recovered (koz)

  1.2   1.2   1.3   1.1   0.7   7.2

Au Recovery

  93%   90%   90%   90%   90%   90%

Ag Recovery

  10%   10%   10%   10%   10%   10%

Refractory Throughput (tpd)

  505   508   518   515   384   398

Heap Leach (ktons)

  -   -   -   -   -   188.4

Au Grade (opt)

  -   -   -   -   -   0.111

Ag Grade (opt

  -   -   -   -   -   0.012

Au Containd (koz)

  -   -   -   -   -   21.0

Ag Contained (koz)

  -   -   -   -   -   2.3

Au Recovered (koz)

  -   -   -   -   -   18.3

Au Recovery

  -   -   -   -   -   0.4

Ag Recovery

  -   -   -   -   -   87%

Leach Stacking Rate (ton/day)

  -   -   -   -   -   19%

Notes:

  1.

All refractory mineralization sold to a third-party processing facility in the years 2025 through 2027.

  2.

Beginning in 2028 refractory mineralization will be processed at i-80’s Lone Tree facility.

13.2  Archimedes Open Pit

The Archimedes Open Pit mineral resource has not been evaluated for surface mining.

13.3  Mineral Point Open Pit

i-80 Gold’s Mineral Point Project will consist of an open pit mining operation using conventional equipment. The Project is a conventional hard rock open pit, and mining is planned to be self-performed. Mining is planned on 50-foot (15.24-meter) benches using haul trucks, shovels, and conventional drill and blast activities. Processed material is planned to be mined at a rate of 68,000 tons (62,000 tonnes) per day.

13.3.1  Initial Pit Limit Evaluations

The open pit optimization was performed using the Pseudo Flow algorithm in Hexagon Mine Plan software. The pit optimizer delineates an economic pit shell that maximizes the value of the extractable material by incorporating the mining cost, processing cost, selling cost, gold recovery values, and an overall pit slope. The result of the pit optimization also includes a series of pit shells across a range of revenue factors. Revenue factors are defined as reducing the commodity price but leaving the cost the same. The generated pit shells can then be evaluated to determine which pits are relatively insensitive to economic factors.

This process assessed the sensitivity of the pit optimizations to the fluctuation in the revenue generated, as well as the impact of pit size and stripping ratio on the Projects’ NPV. This procedure yields a series of nested pit shells that prioritize the extraction of the most economically viable and robust material. Less profitable material, characterized by lower gold grade, higher stripping ratios, or higher ratios of the tonnage per ounce of gold, may be mined later in the mine life, or not at all. These “robust” pit shells are used to develop the pushback designs.

 

 

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The pit optimizations use reasonable and relevant economic, cost, recovery, and pit slope assumptions. The pit optimizer included only resource blocks classified as indicated and inferred. The resource block model contains no blocks classified as measured.

13.3.2  Open Pit Economic Parameters

During the pit limit analysis phase, the Project was envisioned as a 275 to 330 thousand tons (250 to 300 thousand tonnes) per day operation with a two-stage crusher and heap leach pad. The pit analysis was performed with pit slopes defined by rock type. The pit slope by rock unit is summarized in Table 13-12. The key pit optimization parameters used to generate the economic pit shells for the deposit are summarized in Table 13-13. The processing cost and process recoveries were defined by rock and mineral alteration.

Table 13-12: Pit Slope by Lithology Unit

 

   

Lithology Unit

   Slope (degrees) 
   

Waste Dump

   30
   

Alluvium

   55
   

Sanded

   45
   

Unsanded

   45

Table 13-13: Pit Optimization Parameters

 

     

Modifying Factor

  Units   Value     
     

Gold Price

  US$/toz    $2,175
     

Gold Price

  US$/gr   $69.93
     

Silver Price

  US$/toz   $26.00
     

Silver Price

  US$/gr   $0.84
     

Gold Refining Charges

  US$/toz   $1.85
     

Silver Refining Charges

  US$/toz   $0.50
     

Royalties

  %   3%
     

Payable Au

  %   99.9%
     

Payable Ag

  %   99.5%
     

Costs

       
     

Mining

  US$/ton   $3.00
     

Mining

  US$/tonne    $3.31
     

Processing (average)

  US$/ton   $3.12
     

Processing (average)

  US$/tonne   $3.44
     

G&A

  US$/ton   $0.75
     

G&A

  US$/tonne   $0.83
     

Heap Leach Recovery Au (average)

  %   78%
     

Heap Leach Recovery Ag (average)

  %   41%

The parameters in Table 13-13 were used in Equation 13-1 to calculate the gold and silver cutoff grades. The gold cutoff grade (COG) of 0.011 toz/ton (0.36 g/tonne)1 and an incremental cutoff grade (ICOG) of 0.003 toz/ton (0.10 g/tonne) was calculated. The silver COG is 0.323 toz/ton (11.08 g/tonne), and an ICOG of 0.171 toz/ton (5.86 g/tonne) was calculated.

 

 

1 Troy ounce per ton conversion to metric grams per tonne may be inconsistent due to truncation and rounding. Imperial is reported to three significant figures, and metric is reported to two significant figures.

 

 

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Equation 13-1: Cutoff Grade Equation

 

LOGO

Where:

Process is the total on site processing cost,

Recovery is the metallurgical recovery in percent (%),

Selling cost includes royalties and payable percent (%).

Figure 13-12 shows the results for each revenue factor shell, for processed and waste tonnes, along with profit. The shells selected for pushback designs and the eventual mine scheduling were LG57.1, LG57.8, LG62, LG62.2, LG66, and LG72. Selected shells along with others are presented in Table 13-14. Pit shell LG72 was selected as the optimal pit shell, which corresponds to a 78% Revenue Factor. Pit shell LG72 shell has a total tonnage of 1,683.6 Mton including 440.1 Mton of processed material at an average grade of 0.011 toz/ton Au for 4.98 Mtoz of contained gold and 195.5 Mtoz of contained silver. The average stripping ratio is 2.8:1. Figure 13-13 shows the percentage of profit, processed material, and recoverable gold by LG shell. Figure 13-15 is a plan view of the site with the six nested pit shells and section lines. Figure 13-16 to Figure 13-19 are cross sections showing the LG pit shells and the estimated block grades for gold. For the below section and plan plots, the block model has filtered out all blocks below 0.003 toz/ton (0.1 g/tonne) Au.

 

 

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Figure 13-12: LG Shells by Revenue Factor

(Source: Forte Dynamics, 2025)

 

 

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Table 13-14: Profit Factor for Optimization Results

 

                                 
Revenue
Factor
  LG
Name
  Processed
kton
  Au
oz/ton
  Ag
oz/ton
  Waste
kton
  Total
kton
  Stripping
Ratio
  Au
ktoz
  Ag ktoz   Revenue
000s$
  Mining
Cost
000s$
  Processing
Cost 000s$
  Total Op Ex 000s$   Net
000s$
  Total
Ton/toz
Au
  Profit/ton
                                 
16.1%   LG45   3,232   0.014   1.271   7,335   10,567   2.3:1   45   4,108   $118,011   $31,701   $9,816   $41,517   $76,494   234.4   $7.24
                                 
27.6%   LG50   4,096   0.013   1.116   8,184   12,279   2:1   53   4,572   $135,733   $36,838   $12,412   $49,251   $86,482   232.5   $7.04
                                 
39.1%   LG55   4,926   0.013   1.003   9,445   14,371   1.9:1   62   4,942   $155,545   $43,113   $14,953   $58,066   $97,479   233.3   $6.78
                                 
41.4%   LG56   5,224   0.012   0.970   9,811   15,035   1.9:1   64   5,069   $160,978   $45,105   $15,863   $60,968   $100,010   235.1   $6.65
                                 
43.9%   LG57.1   62,987   0.015   0.628   221,284   284,271   3.5:1   914   39,583   $1,949,282   $852,812   $190,557   $1,043,369   $905,913   310.9   $3.19
                                 
45.5%   LG57.8   127,856   0.013   0.507   386,174   514,030   3:1   1,716   64,777   $3,574,878   $1,542,091   $386,953   $1,929,045   $1,645,833   299.5   $3.20
                                 
50.6%   LG60   136,219   0.013   0.495   401,969   538,188   3:1   1,806   67,388   $3,756,831   $1,614,564   $412,385   $2,026,949   $1,729,881   298.0   $3.21
                                 
55.2%   LG62   222,434   0.013   0.412   634,104   856,538   2.9:1   2,875   91,687   $5,810,935   $2,569,615   $673,838   $3,243,453   $2,567,481   298.0   $3.00
                                 
55.6%   LG62.2   269,825   0.013   0.511   847,733   1,117,558   3.1:1   3,409   137,875   $7,166,960   $3,352,674   $818,183   $4,170,856   $2,996,104   327.9   $2.68
                                 
62.1%   LG65   287,929   0.012   0.499   883,405   1,171,334   3.1:1   3,572   143,641   $7,512,551   $3,514,001   $873,114   $4,387,115   $3,125,436   327.9   $2.67
                                 
64.4%   LG66   337,702   0.012   0.498   1,057,599   1,395,301   3.1:1   4,217   168,019   $8,781,080   $4,185,903   $1,023,596   $5,209,500   $3,571,581   330.9   $2.56
                                 
73.6%   LG70   408,775   0.012   0.460   1,197,358   1,606,133   2.9:1   4,786   187,846   $9,925,573   $4,818,400   $1,237,059   $6,055,459   $3,870,114   335.6   $2.41
                                 
78.2%   LG72   440,089   0.011   0.444   1,243,516   1,683,605   2.8:1   4,982   195,532   $10,313,809   $5,050,815   $1,330,036   $6,380,851   $3,932,958   337.9   $2.34
                                 
85.1%   LG75   470,608   0.011   0.428   1,278,749   1,749,357   2.7:1   5,154   201,391   $10,640,300   $5,248,072   $1,419,598   $6,667,670   $3,972,630   339.4   $2.27
                                 
96.6%   LG80   524,340   0.010   0.404   1,336,075   1,860,416   2.5:1   5,411   211,642   $11,141,075   $5,581,247   $1,576,397   $7,157,644   $3,983,431   343.8   $2.14
                                 
100%   LG82   601,390   0.010   0.375   1,409,969   2,011,359   2.3:1   5,720   225,383   $11,459,297   $6,034,078   $1,828,498   $7,862,577   $3,596,720   351.6   $1.79

 

 

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Figure 13-13: Percentage of Profit, Processed Material, and Recoverable Gold by LG Shell

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-14: Plan View of LG Pit Shells and Cross Section Locations

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-15: Pit Optimization Looking West (Section A’ – A)

(Source: Forte Dynamics, 2025)

 

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Figure 13-16: Pit Optimization Looking North (Section B’ – B)

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-17: Pit Optimization Looking North (Section C’ – C)

(Source: Forte Dynamics, 2025)

 

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Figure 13-18: Pit Optimization Looking North (Section D’ – D)

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-19: Pit Optimization Looking North (Section E’ – E)

(Source: Forte Dynamics, 2025)

 

13.3.3

Pit Designs

The pit shells and the block model were used as a basis for preliminary life of mine (LOM) open pit mine designs. Pit shell LG57.1 was determined to be too large for an initial pit phase and was split into two sub-phases. The current heap leach relocation was done in phases 5 and 6. Table 13-15 shows the pit design parameters used. Figure 13-20 shows all nine pit phases, along with a section line running along the strike of the deposit. Figure 13-21 is a cross-section of all nine phases with the block model showing Au toz/ton. Figure 13-22 to Figure 13-30 show each pit phase design individually. Figure 13-31 shows the final pit design and estimated block model in an orthogonal view looking northwest. For the below section and plan plots, the block model has filtered out all blocks below 0.003 toz/ton (0.1 g/tonne) Au.

Table 13-15: Pit Design Parameters

 

Parameter     Units    

 Waste 

Dump

    Alluvium      Sanded      Unsanded 

Bench Height

   ft    50.0    50.0    50.0    50.0

Bench Face Width

   ft    18.2    18.2    18.2    18.2

Catch Bench Width

   ft    68.4    13.0    26.8    26.8

Ramp Width

   ft    130    130    130    130

Ramp Grade

   %    10    10    10    10

Bench Face Angle

   deg    70    70    70    70

Inter Ramp Angle

   deg    30    58    48    48

Overall Slope Angle

   deg    30    55    45    45

 

 

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Figure 13-20: Pit Phasing and Section Line

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-21: Cross Section F’ to F of Pit Phasing

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-22: Phase 1 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-23: Phase 2 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-24: Phase 3 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-25: Phase 4 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-26: Phase 5 Design (First Phase of Heap Leach Relocation)

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-27: Phase 6 Design (Second Phase of Heap Leach Relocation)

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-28: Phase 7 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-29: Phase 8 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-30: Phase 9 Design

(Source: Forte Dynamics, 2025)

 

 

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Figure 13-31: Final Pit and Estimated Block Model in Orthogonal View Looking Northwest

(Source: Forte Dynamics, 2025)

 

13.3.4

Haul Road Design

Existing roads are planned to be utilized where possible. New haul roads will have to be built to the top of each phase for waste mining. This will require the removal of vegetation and any topsoil for the construction of the planned haul roads.

Haul roads were designed to be wide enough for two-lane traffic, except for the bottom four benches, which were designed for single-lane travel to minimize waste stripping requirements.

 

 

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13.3.5

Economic Evaluation

The economic evaluation parameters are different from the pit limit runs. Additional benchmarking from other sites was conducted, along with a more detailed workup of the processing cost. The silver price was updated to reflect current trends. 

 

13.3.6

Cutoff Grade

The processed/waste cutoff grades for mineable resource reporting were based on the economic parameters and the individual metal grades within each block. The mining and processing cost, along with the silver price, have been updated from the numbers in Table 13-13. The updated costs and price are shown in Table 13-16. All other inputs were held the same as shown in Table 13-13. The prices in Table 13-16 were used in Equation 13-1 to calculate a gold COG of 0.006 oz/ton (0.19 g/tonne) and ICOG of 0.004 oz/ton (0.14 g/tonne). The silver COG of 0.277 toz/ton (9.49 g/tonne) and an ICOG of 0.171 toz/ton (5.85 g/tonne) were calculated.

Table 13-16: Design Metal Prices, Costs, and Recoveries

 

Description    Units    Value      

Mining Cost

   US$/ton    $2.50

Processing Cost

   US$/tonne    $2.76

Processing Cost

   US$/ton    $3.90

Processing Cost

   US$/tonne         $4.30

Silver Price

   US$/toz    $27.25

Silver Price

   US$/gr    $0.88

 

13.3.7

Pit Design Inventories

Indicated and inferred mineral resource inventories of the preliminary open pit designs are tabulated in Table 13-17. In summary, the final pit limit contains a total tonnage of 1,675 Mton (1,520 Mtonne) including 245.7 Mton (222.9 Mtonne) of indicated mineral resource at 0.013 toz/ton (0.45 g/tonne) Au and 0.426 toz/ton (14.61 g/tonne) Ag, and 149.7 Mton (135.8 Mtonne) of inferred Mineral Resource at 0.009 toz/ton (0.31 g/tonne) Au and 0.486 toz/ton (16.66 g/tonne) Ag, for a total of 4.5 Mtoz (139.97 Mgram) of contained gold and 177.3 Mtoz (5,514.6 Mgram) of contained silver. Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. There has been insufficient exploration to define the inferred resources tabulated above as indicated or measured mineral resources. However, it is reasonably expected that the majority of the inferred mineral resources could be upgraded to indicated Mineral resources with continued exploration. There is no guarantee that any part of the mineral resources discussed herein will be converted into a mineral reserve in the future.

 

 

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Table 13-17: In-Pit Mineral Resources by Pit Phase

 

          Processed Resource    Waste    Total

Phase

   Material    kton    Au
toz/ton
   Ag
toz/ton
   Au
Cont
ktoz
   Ag Con
ktoz
   Au
Rec
ktoz
   Ag
Rec
ktoz
   kton    kton    Stripping
Ratio

PH1

  

  Measured  

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   25,258    0.011    0.371    269    9,378    224    3,936    32,276    57,534    1.3
    

Inferred

   5,067    0.009    0.256    46    1,295    38    528    10,918    15,985    2.2
    

Waste

   -    -    -    -    -    -    -    67,980    67,980    -

PH2

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   37,239    0.015    0.695    558    25,894    451    10,542    77,661    114,900    2.1
    

Inferred

   11,853    0.010    0.463    116    5,489    93    2,211    30,717    42,570    2.6
    

Waste

   -    -    -    -    -    -    -    95,933    95,933    -

PH3

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   76,486    0.011    0.308    839    23,587    689    9,798    67,694    144,180    0.9
    

Inferred

   7,772    0.009    0.317    68    2,465    56    1,008    6,119    13,891    0.8
    

Waste

   -    -    -    -    -    -    -    134,950    134,950    -

PH4

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   59,822    0.013    0.290    767    17,324    607    7,162    39,350    99,172    0.7
    

Inferred

   24,742    0.011    0.262    281    6,489    221    2,649    17,987    42,729    0.7
    

Waste

   -    -    -    -    -    -    -    158,381    158,381    -

PH5

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   -    -    -    -    -    -    -    -    -    -
    

Inferred

   -    -    -    -    -    -    -    -    -    -
    

Waste

   -    -    -    -    -    -    -    9,112    9,112    -

PH6

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   -    -    -    -    -    -    -    -    -    -
    

Inferred

   -    -    -    -    -    -    -    -    -    -
    

Waste

   -    -    -    -    -    -    -    17,343    17,343    -

PH7

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   11,079    0.016    1.045    172    11,575    120    4,630    26,399    37,478    2.4
    

Inferred

   29,243    0.010    0.983    290    28,755    224    11,502    39,115    68,359    1.3
    

Waste

   -    -    -    -    -    -    -    192,356    192,356    -

PH8

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   25,370    0.017    0.528    425    13,402    269    5,363    30,172    55,542    1.2
    

Inferred

   47,724    0.008    0.473    397    22,560    293    9,027    54,363    102,087    1.1
    

Waste

   -    -    -    -    -    -    -    152,333    152,333    -

PH9

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   10,452    0.010    0.328    105    3,432    87    1,399    6,278    16,730    0.6
    

Inferred

   23,336    0.008    0.242    191    5,647    157    2,273    11,687    35,023    0.5
    

Waste

   -    -    -    -    -    -    -    673    673    -

Total

  

Measured

   -    -    -    -    -    -    -    -    -    -
    

Indicated

   245,706    0.013    0.426    3,135    104,591    2,448    42,830    279,831    525,537    1.1
    

Inferred

   149,738    0.009    0.486    1,389    72,702    1,081    29,199    170,906    320,644    1.1
    

Waste

   -    -    -    -    -    -    -    829,062    829,062    -

 

13.3.8

Drilling and Blasting

Primary fragmentation for mining will be carried out using traditional drill and blast techniques that are standard in open pit mining. This study used a powder factor of 0.51 lb/ton (0.25 kg/tonne) for mineralized material and waste rock.

Benches are blasted and mined in 50-foot (15.24-meter) benches. Buffer and trim rows are planned to allow controlled blasting and minimize back-breaking damage to the high walls.

 

13.3.9

Production Schedules

The mine designs were used to create a LOM schedule for the site. This schedule considers open pit mining operations. The yearly mine schedule is presented in Table 13-18. The production schedule is driven by the nominal rate of 68,000 ton/day (62,000 tonne/day) processed material which is a 25 Mton/year (23 Mtonne/year), and the average LOM stripping ratio is 3.2:1 waste-to-processed material.

Table 13-18 details the LOM production schedule by year. Figure 13-32 shows the LOM annual production schedule for processed, heap leach relocation, and waste materials, and recovered Au toz.

 

 

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Table 13-18: LOM Production Schedule

 

Year

    Days    

 Processed 

kton

  

HL
Relo

kton

  

Waste

kton

  

Total

kton

  

Stripping

Ratio

   Au
(toz/ton)
   Ag
(toz/ton)
   Au
Cont (ktoz)
   Ag
Cont (ktoz)
   Au
Rec (ktoz)
   Ag
Rec (ktoz)

1

   365    8,132    -    121,783    129,915    15.0    0.007    0.467    60.7    3,795.0    50.5    1,600.9

2

   365    25,000    -    106,842    131,842    4.3    0.010    0.410    253.3    10,250.9    210.9    4,249.7

3

   366    25,068    -    108,536    133,604    4.3    0.011    0.431    277.0    10,800.9    223.7    4,455.7

4

   365    24,323    -    106,552    130,874    4.4    0.018    0.747    433.3    18,164.1    348.3    7,318.1

5

   365    25,000    -    110,495    135,495    4.4    0.010    0.298    246.2    7,462.0    200.1    3,125.0

6

   365    25,000    -    91,216    116,216    3.6    0.010    0.271    249.8    6,767.9    204.0    2,783.4

7

   366    25,068    9,112    76,394    110,574    3.0    0.012    0.316    294.5    7,913.3    243.4    3,284.3

8

   365    25,000    -    78,883    103,883    3.2    0.012    0.336    300.7    8,411.4    244.0    3,471.3

9

   365    25,000    -    80,373    105,373    3.2    0.011    0.250    280.7    6,249.0    215.9    2,577.5

10

   365    25,000    -    88,958    113,958    3.6    0.014    0.318    348.9    7,947.2    275.1    3,272.8

11

   366    25,068    -    88,600    113,669    3.5    0.015    0.553    364.4    13,867.3    274.8    5,578.9

12

   365    25,000    -    78,692    103,692    3.1    0.011    1.053    269.1    26,333.1    209.8    10,533.2

13

   365    25,000    -    80,420    105,420    3.2    0.014    0.650    344.7    16,255.2    207.1    6,506.9

14

   365    25,000    17,343    9,512    51,855    0.4    0.012    0.525    301.2    13,120.6    210.4    5,248.3

15

   366    25,068    -    6,861    31,929    0.3    0.007    0.380    175.3    9,523.6    143.9    3,809.4

16

   365    25,000    -    17,496    42,496    0.7    0.008    0.268    206.0    6,705.5    168.9    2,717.4

17

   365    12,717    -    1,731    14,448    0.1    0.009    0.293    118.7    3,726.1    98.5    1,495.5

Total

        395,444    26,455    1,253,344    1,675,243    3.2    0.0114    0.4483    4,525    177,293    3,529    72,028

 

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Figure 13-32: LOM Annual Production Schedule

(Source: Forte Dynamics, 2025)

 

 

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13.3.10

 Mine Fleet

The Project’s mining fleet will be designed to support the planned open pit operation, with a focus on maximizing efficiency and production rates while maintaining operational flexibility and safety. The primary equipment for the mining operation will consist of four main shovels (two rope shovels and two hydraulic shovels) and up to 26 haul trucks.

The two rope shovels will be used for overburden removal and high-volume digging, offering the advantage of high digging force and efficiency in hard rock conditions. The rope shovels will be equipped with large capacity buckets to facilitate the efficient loading of the haul trucks. The two hydraulic shovels will primarily be employed for more selective digging in ore zones and areas requiring increased precision. The fleet will consist of up to 26 haul trucks, each with a 320-ton capacity, which can transport large volumes of material efficiently from the pit to the processing plant or waste disposal area. The haul trucks will be selected for their reliability, fuel efficiency, and suitability for the operating environment. A wide range of support equipment will support the load and haul fleet. Table 13-19 contains a list of the proposed mining equipment for the Project.

Table 13-19: Mining Equipment List

 

Item     Manufacturer     Model     # of  Units 

Cable Shovels small

   Komatsu    2800XPC    1

Cable Shovels large

   Komatsu    4100XPC    1

Hydraulic Shovel

   Komatsu    PC5500-11      2

Rear Dump Trucks

   Komatsu    930E-5    26

Loader

   Komatsu    WE1850-3    1

Rotary Drills

   Komatsu    ZR77    5

Bulldozers

   Komatsu    D375A-8    5

Wheel Dozer

   CAT    854    2

Graders

   Komatsu    GD955-7    3

Water Tankers

   Komatsu    830E-5    2

 

13.3.11

 Dewatering

Dewatering will be necessary as the pit develops and as is covered in section 15.2.4.

 

 

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14.

RECOVERY METHODS

 

14.1

Archimedes Underground

 

14.1.1

Introduction

Refractory production from the Ruby Hill operation will be processed via milling, pressure oxidation followed by carbon in leach (CIL) or roasting followed by CIL. The most recent metallurgical testing is described in Section 13 Mineral Processing and Metallurgical Testing that will support processing parameters at the Turquoise Ridge Surface Sage autoclave under a Toll Milling Agreement (TMA).

Ruby Hill production will be classified based on gold grade, level of oxidation and refractory characteristics (e.g. presence of preg-robbing components in ore, refractory sulfide components) which contribute to recovery at processing facilities and is routed based on an integrated process production plan is devised for maximum economic returns.

Nevada Gold Mines LLC (Nevada Gold Mines) operates the Turquoise Ridge Complex, located in Humboldt County, Nevada, USA. Nevada Gold Mines is a joint venture between Barrick Gold Corporation (Barrick) and Newmont Corporation (Newmont), Barrick is the operator of the joint venture and owns 61.5%, with Newmont owning the remaining 38.5%. Under the joint venture, Barrick’s Turquoise Ridge Mine and Newmont’s Twin Creeks Complex were combined as a single operation, now known as Turquoise Ridge. The process operations are now known as the Sage Mill complex.

 

14.1.2

Refractory Mineralization Processing

Prior to 2028, refractory mineralization from Archimedes Underground will be shipped to the Turquoise Ridge Complex. Production will be sampled to determine the geochemistry, gold content, and moisture content.

Specifically, the samples are assayed for organic carbon and gold since the gold recovery formula is dependent upon these two parameters. The simplified Sage Mill flowsheet is shown in Figure 14-1.

 

 

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Figure 14-1: Third Party POX Facility Simplified Flowsheet

(Source: Nevada Gold Mines, 2020)

 

14.1.2.1

Sage Mill Process

The Sage Mill processes 4 - 5 million tonnes per year of feed from various sources.

Mill feed is passed through a grizzly and the undersize is fed to a 8.5 m diameter by 3.0 m long 3.0 MW SAG mill. The SAG mill is fitted with a trommel jet with no pebbles discharged from the mill. SAG mill discharge is combined with the primary ball mill discharge (7.9 m diameter by 9.1 m long, 5.6 MW) and classified by 500 mm diameter cyclones. Primary cyclone overflow is further ground by two 5.0 m diameter by 8.8 m long 3.0 MW ball mills operating in closed circuit with 250 mm diameter cyclones. Secondary cyclone overflow reports to a 61 m diameter thickener. Thickener underflow reports to an acidification circuit where sulphuric acid is added as necessary to ensure adequate autoclave free acid solution levels. The free acid concentration for Turquoise Ridge Complex pressure oxidation circuit is maintained at minimum of 30 g/L.

Thickener overflow solution is returned to the milling circuit. There are three surge tanks ahead of the two autoclaves, providing 15 hours autoclave feed storage. After acidification, ore slurry is added to two identical autoclaves that are operated in parallel. Each autoclave is 5.8 m outside diameter and 22.9 m overall length. Each autoclave has four compartments and provides approximately 50 minutes retention time. The autoclaves are operated at 225oC and 3.2 MPa oxygen over pressure. Two stages of flash heat recovery

 

 

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are utilized. Autoclave discharge is cooled before reporting to the lime neutralization circuit. Autoclave waste gas is cooled and scrubbed before discharging to the atmosphere.

Oxide ore and acidic oxidized sulfide ore slurry are combined in the neutralization circuit.

After neutralization with the carbonate oxide ore and supplemental lime, the combined slurry reports to a carbon-in-leach (CIL) circuit where the combined slurry leached in cyanide solution to extract the gold. The CIL circuit provides approximately 18 hours retention time. Final tailings slurry is pumped to the tailings area. Tailings settle and decant solution is reclaimed and reused in the grinding circuit.

Loaded carbon from the CIL circuit is transferred to the recovery plant. After acid washing to remove inorganic contaminants, the carbon is transferred to the pressure Zadra stripping circuit. Gold is stripped from the carbon using caustic and cyanide solution at elevated temperature and pressure. Pregnant solution from the stripping circuit is pumped to an electrowinning circuit where precious metal is removed from the solution as sludge. The sludge is filtered, dried in a mercury retort, mixed with fluxes, and refined into doré bars.

After carbon stripping, the barren carbon reports to the kiln regeneration circuit and returns to the CIL circuit.

Gold recovery estimates are based on both testwork and operational history at both facilities with curves utilized for both depending on operating strategy and mineralization characteristics.

 

14.2

Lone Tree Pressure Oxidation Facility

i-80 Gold plans to process single refractory mineralization from their Nevada mines at their Lone Tree Mill in a hub and spoke arrangement.

 

14.2.1

Lone Tree Mill Historic Processing

The Lone Tree Mine is located immediately adjacent to I-80, approximately 12 miles west of Battle Mountain, 50 miles east of Winnemucca, and 120 miles west of Elko. Mining commenced at Lone Tree in April 1991 with the first gold pour in August of 1991. In 1993, a POX circuit was added to the facility, which included a SAG / ball mill circuit, followed by a thickening circuit, the POX process for refractory gold ores, and finally CIL, carbon stripping, and refining.

In 1997, a 4,500 tpd flotation plant was constructed to make concentrate to supplement the feed to the POX circuit, as well as to ship excess concentrate to Newmont’s Twin Creeks POX plant or to its Carlin roaster. The Lone Tree processing facilities were shut down at the end of 2007. Since that time, the mills have been rotated on a regular basis to lubricate the bearings. In general, the facility is still in place with most of the equipment sitting idle.

i-80 Gold Corp’s objective is to refurbish and restart the POX circuit and associated unit operations, including the existing oxygen plant, as it was operating before the shut-down, while meeting all new regulatory requirements. The flotation circuit is not being considered for restart. The POX circuit will have capability to operate under either acidic or basic conditions.

In order to restart the process plant, new environmental regulations in relation to allowable mercury emissions must be met. In February 2011, the NDEP and the EPA brought about new standards to limit mercury emissions to 127 lb of mercury for every million tons of ore processed. In order to meet this requirement, the Lone Tree facility will require several environmental upgrades prior to restart.

 

 

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14.2.2

Lone Tree Facility Block Flow Diagram

A block flow diagram for the Lone Tree Mill facility is included in Figure 14-2. The block flow diagram contains the follow major processing areas:

 

   

Ore Reclaim, Grinding and Thickening and Acidulation

 

   

Pressure Oxidation

 

   

POX Off-gas Treatment and Quench Water Loop

 

   

Neutralization, Carbon-in-Leach, and Cyanide Destruction

 

   

Tailings Thickening and Filtration

 

   

Acid Wash, Carbon Stripping, and Carbon Regeneration

 

   

Electrowinning and Refinery

 

   

Plant and Instrument Air

 

   

Oxygen Plant

 

   

Reagent Preparation and Storage

Process and Plant Service Cooling Towers

 

   

Water Distributions

 

   

Steam Generating Plant and Propane Storage.

 

 

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Figure 14-2: Loan Tree Block Flow Diagram

(Source: i-80 Gold, 2025)

 

 

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14.2.3

Key Design Criteria

The Lone Tree Pressure Oxidation (POX) Facility restart will have minimal changes made from the 1993 PDC. A new PDC was developed based on the expected production sources as defined by i-80.

Key process design criteria are summarized in Table 14-1.

Table 14-1: Summary of Key Process Statistics

 

Criteria      Units        Value  

Annual Mill Throughput

   tons    912,500

Daily Throughput (per calendar day)

   tons    2,500

Operating Throughput of Ore to Autoclave Circuit (LTH feed)

   tph    122.5

Operating Time / Availability

   %    85

Design Sulfur Treatment Rate

   tph S    2.7

Gold Recovery

   %    Varies

Silver Recovery

   %    Varies

 

14.2.4

Lone Tree Facility Description

 

14.2.4.1

Mill Feed Reclaim

The purpose of the Mill feed reclaim area is to store and reclaim material for processing, which has been shipped to the lone tree processing facility via highway trucks.

Run of mine (ROM) crushed material is delivered to the stockpile area. Material from various mining locations – namely Granite Creek, Cove, and Archimedes – is dumped at designated locations within the storage area and blended into facility feed stockpiles.

The stockpile area will have the capacity to store multiple days worth of mined and crushed material to accommodate the production shipment schedule to site. Additionally, the reclaim area is utilized for feed blending for the POX circuit. This blending will be used to manage the sulfide sulfur concentrations, gold grades, and carbonate grades through the autoclave to ensure stable circuit operation within the design window for the plant.

 

14.2.4.2

Comminution

The purpose of comminution area is to reduce the particle size of the feed mineralization to the target autoclave circuit feed size for sufficient sulfide oxidation kinetics and gold recovery within the autoclave. The comminution area contains an SABC circuit with a dedicated SAG (semi-autogenous grinding mill) and ball mill to reduce the feed particle size to the target grind size. The SAG mill is fed via a conveyor from the dump hopper. The ball mill cyclone overflow is directed to the POX feed thickening conveyor.

 

14.2.4.3

Thickening and Acidulation

The purpose of the thickening area is to prepare the slurry for autoclave process by densifying the product of the grinding circuit to improve storage capacity of the downstream slurry storage tanks, improve the autoclave heat balance by reducing the water transferred to the autoclave and improving the possible solids flow through the autoclave feed pumps. The dense slurry is stored in two acidulation tanks that provide a combined storage / acidulation retention time of 12 hours. The acidulation tanks ensure a continuous feed to the autoclave plant, unaffected by upstream throughput variations.

 

 

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14.2.4.4

Pressure Oxidation

The POX autoclave circuit includes the slurry pre-heaters, autoclave feed, autoclave, and the POX ancillary services: autoclave agitator seal system, oxygen supply, high pressure cooling water, and high-pressure steam. The Lone Tree Facility restart includes provisions to operate the circuit in alkaline or acidic modes depending on the feed carbonate concentration among other factors.

 

14.2.4.4.1

Slurry Heaters

The purpose of the slurry heaters is to capture excess energy discharged from the autoclave and pre-heat the feed slurry prior to the autoclave process reducing the total energy input required to operate the autoclave. The heating is achieved in two stages consisting of a series of two refractory lined counter-current splash slurry heater vessels. The heat source is flashed steam released from the autoclave discharge slurry during the pressure letdown process. The splash slurry heaters are direct contact heat exchanger and provide a means of heat recovery via steam condensation. This reduces the off-gas load on the downstream off-gas equipment and reduces the required input steam.

 

14.2.4.4.2

Autoclave Feed

The purpose of the autoclave feed area is to increase the pressure of the pre heated slurry to above the autoclave operating pressure to facilitate transfer into the autoclave at the required pressure using the autoclave feed pumps.

 

14.2.4.4.3

Autoclave

The purpose of the autoclave is to oxidize the refractory sulfide minerals under acidic or alkaline conditions to liberate the gold trapped in the sulfide sulfur minerals. The autoclave at Lone Tree is designed to operate at 389 °F and 297 PSI(g) with a slurry residence time of 40 - 50 minutes and consists of 4 compartments. The design expects a 78%—97% cumulative sulfide sulfur oxidation through the autoclave depending on operating conditions. In either operating condition high purity oxygen is introduced to all four compartments of the autoclave at controlled rates to oxidize the fed sulfide minerals. Due to the low sulfur grades steam is required to be continuously fed to the autoclave to maintain the kinetically required oxidation rates to achieve the sulfide sulfur oxidation extent. The autoclave slurry is discharged through a level control choke valve and is fed to the high pressure flash vessel.

 

14.2.4.4.4

Flash System

The purpose of the flash system is to reduce the pressure and temperature of the autoclave discharge, making it suitable for subsequent unit operations downstream. The oxidized slurry undergoes a controlled pressure and temperature reduction process as it passes through two stages of flashing vessels located downstream of the last autoclave compartment.

 

14.2.4.5

POX Off-gas Treatment

The purpose of the POX off-gas treatment area is to effectively eliminate particulate matter present in the POX vent stream, while simultaneously reducing the temperature and volume of the vent gas through direct contact condensation. This process serves to alleviate the burden imposed on downstream equipment, ensuring their optimal performance, and mitigates the environmental impact by minimizing emissions. The off-gas treatment circuit also includes a mercury removal step to minimize autoclave mercury emissions to the environment.

 

 

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14.2.4.6

Slurry Coolers

The purpose of slurry coolers is to reduce the temperature of the incoming slurry from the low-pressure flash vessel to prepare it for the downstream neutralization and CIL circuits through a series of water cooled shell and tube heat exchangers.

 

14.2.4.7

Neutralization

The purpose of neutralization circuit is to neutralize all free acid in the slurry, precipitate the heavy metals as their hydroxides and raise the pH to approximately 10 to ensure cyanide stability in the CIL circuit for personnel safety and process optimization. The neutralization circuit is dosed with lime slurry to raise the pH of the autoclave discharge slurry. The neutralized slurry from this circuit is then fed to the CIL circuit for gold recovery.

 

14.2.4.8

Carbon-in-Leach

The purpose of CIL circuit is to leach and extract gold and silver from the oxidized slurry from neutralization using cyanidation and carbon adsorption. The CIL circuit provides retention time of 24 to 28 hours. The CIL circuit consists of 6 mechanically agitated tanks arranged in a series. The agitators prevent solid settlement and maximize contact time to improve gold and silver recovery. The carbon flows counter current to the slurry flows and the loaded carbon is sent to an elution circuit for carbon stripping and regeneration. Unloaded carbon is fed the last tank of the CIL circuit. The leached slurry is transferred from to the cyanide destruction circuit.

 

14.2.4.9

Elution

The purpose of the elution circuit is to elute precious metals from the loaded carbon and transfer the resulting loaded solution of high gold concentration (pregnant eluate) to the refinery to generate doré.

 

14.2.4.9.1

Carbon Acid Wash

The purpose of acid wash is to rinse the loaded carbon form CIL with dilute nitric acid solution prior to the carbon stripping process. Carbonate scale builds up on the activated carbon during the CIL process and fouls the carbon’s adsorption properties by depositing a layer of scale. If left intact, over time the scale will limit the adsorption capacity of the carbon and will cause softening of the carbon in the regeneration kiln. The loaded carbon from CIL is first treated within the carbon acid wash vessel prior to treatment within the carbon stripping vessel.

 

14.2.4.9.2

Carbon Stripping

The purpose of the carbon strip circuit is to strip the cleaned loaded carbon from the acid wash vessel of the adsorbed gold using a Pressure ZADRA Strip scheme. The ZADRA strip uses several bed volumes of a recirculated solution to strip the precious metals off the loaded carbon. The cyanide solution is buffered by caustic to assist with gold elution. The stripped carbon is then sent to carbon regeneration circuits. The loaded solution is next processed in the electrowinning circuit.

 

14.2.4.9.3

Elution Mercury Abatement

The purpose of elution mercury abatement system is to condition the off gas leaving the pregnant and barren solution tank to remove fine particulate, solution aerosols and condensed and gas phase mercury.

 

14.2.4.10

Carbon Regeneration

The purpose of the carbon regeneration circuit is to restore the activated carbon’s ability to recover gold from the cyanidation circuit solutions. The circuit also permits the introduction of new carbon to the process and removes carbon fines from the process.

 

 

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14.2.4.10.1

Carbon Regeneration Kiln

As carbon is used in the CIL and elution circuits, the surface and internal pore structure becomes contaminated with organic species. The organics foul the carbon, slow the gold adsorption rate, and decrease the gold loading capacity of the carbon. The carbon reactivation electric kiln is a horizontal rotary kiln that is specifically designed for this purpose.

 

14.2.4.10.2

Carbon Fines Handling

Carbon fines are transferred by gravity from the reactivated carbon vibrating screen, carbon reactivation feed vibrating screen, kiln feed hopper, and carbon reactivation electric kiln. The carbon fines are dewatered in a filter press and discharged into supersacks for external sale.

 

14.2.4.11

Refinery

The purpose of the refinery circuit is to recover gold cyanide solutions via electrowinning and produce doré bullion bars.

 

14.2.4.11.1

Electrowinning

The purpose of the electrowinning (EW) circuit is to recover gold from the pregnant solution by applying a voltage across electrodes immersed in the pregnant solution. Rich solution from the pregnant solution tank is transferred through the EW cells to electrowin the gold.

 

14.2.4.11.2

Refining

The purpose of the refining process is to produce doré bars void of other contaminants including but not limited to mercury.

The sludge from the EW cells is first processed in a mercury retort oven to remove the co-captured mercury from the precious metals recovery steps. The retorted gold sludge is then processed in a melt furnace to produce the final mine grade doré bars.

 

14.2.4.12

Cyanide Destruction

The purpose of the cyanide destruction circuit is to effectively reduce the concentration of cyanide in the final tail discharge and the recycled process water, ensuring compliance with predefined environmental standards and regulations and improving the safety of the operation by reducing cyanide concentrations outside of the CIL and elution circuits. The circuit targets a specific concentration limit of 2.5 mg/L of residual weakly acid-dissociable cyanide (CNWAD). This reduction is accomplished through the application of the SO2/air cyanide destruction process, which oxidizes the cyanide to meet the required concentration level. The cyanide destruction circuit is fed directly from the slurry discharge from the CIL circuit.

 

14.2.4.13

Tailings Preparation

The purpose of the tailings circuit is to increase the density of the detoxified tailings to aid with dry stacking of tailings residue. Additionally, this circuit produces process water for internal use within the facility. The tailings preparation circuit consists of a thickener as a first stage of solids densification. The thickener underflow is then fed to a tailings filtration circuit which dewaters the tailings sufficiently to support tailings dry stacking. The de-watered tailings from the filter presses are then dry stacked at the tailings storage facility.

The water removed from the tailings slurry is used as process water within the facility to offset water requirements. Excess process water is processed via a reverse osmosis circuit to provide supplemental permeate water to offset fresh water requirements.

 

 

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14.2.4.14

Water Distributions

There are eight types of defined water services at Lone Tree:

 

   

Fresh water – Is generally used for reagent make-up and water washing streams.

 

   

Gland water – Is used to supply gland water to slurry pumps.

 

   

Mill water – Is used to provide dilution water within the milling circuit.

 

   

Potable water – Is used for safety showers and sanitary uses.

 

   

Demineralized water – Is primarily used to supply the steam generating plant.

 

   

Process water – Is used for washing and slurry dilutions. Additionally, generally feeds the reverse osmosis circuit to generate permeate water.

 

   

Quench water – Is used within the POX off-gas circuit as the source of direct cooling water.

 

   

Excess water – Is discharged from the main processing facility to the existing heap leach facility for treatment.

 

14.2.4.15

Solution Cooling

The purpose of the cooling area is to reject heat absorbed within the process to atmosphere. The solution cooling area includes the process service cooling circuit and the plant service cooling circuit. The process cooling circuit rejects the heat from the autoclave cooling circuit and the elution circuit heat exchangers. The plant service cooling circuit provides trim heat rejection from various equipment support systems throughout the design.

 

14.2.4.16

Reagents

Each set of compatible reagent preparation and storage systems is located within dedicated containment areas to prevent erroneous mixing of reagents. Storage tanks are equipped with level indicators, instrumentation, and alarms to reduce the risk of spills during normal operation. Appropriate ventilation, fire and safety protection, safety shower stations and Safety Data Sheet stations are located throughout the facility.

 

14.2.4.16.1

Oxygen Plant

High purity oxygen is primarily used for oxidation of sulfide during the POX process, of iron conversion from ferrous to ferric in the neutralization circuit, and of cyanide to cyanate in cyanide destruction. Furthermore, during cyanidation, the addition of oxygen maximizes the rate of gold dissolution. At Lone Tree, a cryogenic ASU produces high purity oxygen. The unit uses pressure swing adsorption technology for front end purification and production of high-pressure oxygen at 95% purity.

 

14.2.4.17

Instrument and Plant Air

The Lone Tree facility includes separate instrument and plant air systems to support the facilities air requirements.

 

14.2.5

Utilities Consumption

The plant consumptions for water and power are provided for the average processing case below and consider the design blend of material to be processed within the Lone Tree Facility for the design life of operation.

 

14.2.5.1

Water Consumption

Table 14-2 provides a summary of the water consumption by type for the Lone Tree processing facility.

 

 

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Table 14-2: Lone Tree Facility Water Consumption by Type

 

Type     Consumption  (gpm) 

Mill Water

   1,550

Fresh Water

   570

Permeate Water

   195

Low Pressure Gland Water

   105

High Pressure Gland Water

   170

Demineralized Water

   110

Potable Water

   15

 

14.2.5.2

Electrical Power Requirements

The estimated annual electrical energy requirements for the Lone Tree processing facility are summarized by area in Table 14-3.

Table 14-3: Lone Tree Facility Energy Usage by Area

 

Area   

Annual Energy

 Consumption (MWh/y) 

000 – General Plant Wide

   2,250

180 – Water System

   930

181 – Potable Water

   240

182 – Process Water (RO and Process Water Tank)

   4,900

210 – Mineralization Reclaim

   770

240 – Refinery

   2,310

241 – POX Grinding

   26,920

242 – POX Grinding Thickening and Acidulation

   1,890

244 – Neutralization and CIL and Acid Storage

   6,540

245 – Carbon Stripping

   4,090

247 – CND

   690

248 – Reagents

   2,640

249 – Plant Air and Propane

   3,310

250 – Pressure Oxidation (POX) and POX Utilities

   15,540

251 – POX Demineralized Water System

   2,660

275 – Tailings Filtration

   13,690

300 – Plant Wide Electrical and Instrumentation

   4,000

305 – ABS and CN Storage

   160

320 – POX Mercury Abatement

   900

340 – Quench Water Treatment

   4,020

255 – Oxygen Plant

   40,090

099 – Existing Plant Areas

   3,570

Total

   142,090

 

 

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14.3

Mineral Point Open Pit

The proposed processing facilities for the Mineral Point Project will be developed in correspondence to the mining sequence of the deposit. The primary processing methods include primary and secondary crushing, conveyor stacking on an HLP, extraction with cyanide solution, Merrill-Crowe recovery of precious metals, and refining.

 

14.3.1

Summary Process Design Criteria

Table 14-4 lists the preliminary design process for the process facilities and is grouped by ore mineralization type and deposit as required. It should be noted that the processing circuits have not been optimized at this time and require additional test work to be completed in further stages of this project.

Table 14-4: Mineral Point Design Criteria

 

      Units    Nominal Design       Source

ORE CHARACTERISTICS

Dry Bulk Density

              

Mineral Point

   lb/ft3    118    Forte

Historic Heap Leach Relocated Ore

   lb/ft3    118    Forte
 

Leach Pad Stacking Properties

              

Angle of Repose

   Degrees    37    Forte

Crushed Ore Moisture

   %    4    Forte

Historic Heap Leach Relocated Ore

   %    8    Forte

Work Indices and Abrasion

              

Crusher Work Index (CWi)

   kWh/st    12    DRA-2022

Bond Abrasion Index (Ai)

   g    0.3    DRA-2022

Ave. UCS Strength-Hamburg (CH)

   psi    5,000    Golder-2015        
 

Particle Size Passing 80% (P80) Inches

              

RoM Size Passing 80% (P80)

   in    16    Forte

HL Relocated Ore Size Passing 80% (P80)

   in    0.75    Client
 

OPERATING SCHEDULE

              

Mining

              

Operating Schedule

   days/year    365    Client
     days/week      7    Client
     hours/day    24    Client

Crushing/Stacking

              

Operating Schedule

   days/year    365    Client
     days/week    7    Client
     hours/day    24    Client

Crusher Availability

   Hours/day    20    Forte
                

MERILL-CROWE PLANT

              

Operating Schedule

   days/year    365    Client
     days/week    7    Client
     hours/day    24    Client

Plant Availability

   %    98    Forte
                

PRODUCTION DATA

Overall Ore Production Rate

              

LOM Average Mineral Point

   kstpy    24,900    Forte

HL Relocated Ore in active years

   kstpy    2,258    Forte

Yearly Ore Placed

        See Section 6     

Total Mineral Point

   kdst    408,816    Forte

 

 

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     Units    Nominal Design    Source

Total HL Relocated Ore

   kdst    26,290    Forte

Mineral Point Strip Ratio LOM Average

   waste:ore    3.1    Forte

Precious Metal Grades and Recovery

              

Average Head Grade – Au (LOM)

              

Mineral Point

   opt    0.011    Forte

HL Relocated Ore

   opt    0.000    Forte

Average Head Grade – Ag (LOM)

              

Mineral Point

   opt    0.43    Forte

Historic Leached Ore

   opt    0.00    Forte
                

Recovery – Au (LOM)

              

Mineral Point

              

Au Silicic Oxide Crush HL

   %    84.4     

Au Silicic Sulfide Crush HL

   %    31.0     

Au Sanded Oxide Crush HL

   %    83.5     

Au Sanded Sulfide Crush HL

   %    24.0     

Au Weakly-Altered Oxide Crush HL

   %    83.0     

Au Weakly-Altered Sulfide Crush HL

   %    24.0     

Au Heap Leach Relocate

   %    N/A     

Recovery – Ag (LOM)

              

Mineral Point

              

Ag Silicic Oxide Crush HL

   %    45.2     

Ag Silicic Sulfide Crush HL

   %    45.2     

Ag Sanded Oxide Crush HL

   %    44.0     

Ag Sanded Sulfide Crush HL

   %    44.0     

Ag Weakly-Altered Oxide Crush HL

   %    40.0     

Ag Weakly-Altered Sulfide Crush HL

   %    40.0     

Ag Heap Leach Relocate

   %    N/A     
                

LEACH PAD DESIGN

Leach Pad Properties

              

Leach Pad Area – Phase 1

   ft2    8,420,000    Forte

Phase 1 Ore

   kdst    9,336,000    Forte

Leach Pad Area – Phase 2

   ft2    8,420,000    Forte

Phase 2 Ore

   kdst    9,336,000    Forte

Leach Pad Area – Phase 3

   ft2    8,420,000    Forte

Phase 3 Ore

   kdst    9,336,000    Forte

Leach Pad Area – Phase 4

   ft2    8,420,000    Forte

Phase 4 Ore

   kdst    9,336,000    Forte

Leach Pad Area – Phase 5

   ft2    8,420,000    Forte

Phase 5 Ore

   kdst    9,336,000    Forte

Leach Pad Area – All Phases

   ft2    42,100,000    Forte

Total Capacity of Pad – All Phases

   kdst    466,800,000    Forte

Leach Pad Stacking Method

   —     Conveyors    Forte

RoM Haul Truck Capacity

   st    320    Forte

Ultimate Height

   ft    250    Client

Average Lift Height

   ft    30    Forte

Overall Heap Leach Slope

   h:v    3:1    Forte

Reclaimed Heap Leach Slope

   h:v    3:1    Forte
                

LEACHING SOLUTION MANAGEMENT

Solution Application Method

   —     Drip Emitter    Forte

Barren Solution Application Rate

   gpm/ft2    0.003    Forte

Barren Solution Flow Rate

   gpm    11,500    Forte

Primary Leach Cycle

   days    90    Forte

Area Under Leach

   ft2    3,833,000    Calculated

 

 

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     Units    Nominal Design    Source

Tons Under Primary Leach

   dst    6,785,000    Calculated

Barren Solution pH

   pH    10.5    Forte

Pregnant Solution Collection

              

In-pad Collection Piping Layout

   —     Herring Bone    Forte

Pregnant Solution Pond Operating Volume

   Mgal    8.3    Forte. 12-hours at nominal flow

Pregnant Solution Pond Draindown Volume

   Mgal    16.6    Forte. 24-hours draindown at nominal flow

Pregnant Solution Pond Total Volume

   Mgal    24.9    Forte. Excludes 2-foot freeboard

Event Pond

              

Event-100 year-24 hr (depth)

   in    2.94    Forte

Event Pond Volume Phase 1

   Mgal    15.4    Forte, Excludes 2-foot freeboard

Event Pond Volume Phase 2

   Mgal    15.4    Forte, Excludes 2-foot freeboard

Event Pond Volume Phase 3

   Mgal    15.4    Forte, Excludes 2-foot freeboard

Event Pond Volume Phase 4

   Mgal    15.4    Forte, Excludes 2-foot freeboard

Event Pond Volume Phase 5

   Mgal    15.4    Forte, Excludes 2-foot freeboard

Total Event Pond Volume-All Phases

   Mgal    77.2    Forte, Excludes 2-foot freeboard

PROCESSING AND REAGENTS

Crushing

              

Crusher Availability

   %    83.3    Forte

Primary Throughput (nominal/max)

   st/hr    3,450/4,100    Forte

Mineral Point

Product Size Passing 80% (P80)

   in    0.75    Client/Robert Raponi

Processing

              

Quicklime Consumption

   lb/ton    8    Robert Raponi/Forte

Cyanide Consumption

   lb/ton    1    Robert Raponi/Forte

Merrill-Crowe

   gpm    11,500    Forte

 

14.3.2

Process Descriptions

The Mineral Point Project will place approximately 68.0 kstpd of crushed ore for a period of approximately 17 years. Run-of-mine (ROM) ore will undergo primary and secondary crushing operations in open circuit. Crushed ore and heap leach (HL) relocated ore, from historic operations, will also be placed on the HLP. Loading and stacking of the crushed ore will be done utilizing conveyors and a radial stacker. The relocated ore from the historic HLP will be loaded into haul trucks and direct dumped onto the HLP and spread with dozers. The pregnant solution recovered from the HLP will flow into the process pond and be pumped into the Merrill-Crowe zinc precipitation circuit for metals recovery. The Merrill-Crowe process is a zinc precipitation circuit in which the precipitates will be heated in a retort to capture mercury, after which it is fed into a smelting furnace to produce doré. The doré will be sold and shipped off site for further refining. The site also includes all associated infrastructure, facilities, and reagents necessary for the operation. The Merrill-Crowe and refinery will be indoors, and the refinery will be further enclosed for security purposes. Figure 14-3 shows the flowsheet for Mineral Point.

 

14.3.2.1

Crushing

 

14.3.2.1.1

Primary

Run-of-mine (ROM) ore will be transported from the pit to a primary crusher via 320 st haul trucks. The haul trucks will direct dump into a gyratory crusher. The crushing plant will operate 20 hours per day, seven days a week. The primary crusher will provide a product size with 100 percent passing 7”. The primary crusher product is discharged to the secondary crusher feed.

 

 

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14.3.2.1.2

Secondary

The secondary crushing circuit consists of four cones crushers operating in parallel. The secondary crushers will provide a product size with an 80 percent passing size of 0.75”.

 

14.3.2.2

Ore Handling and Stacking

Discharge from the secondary crushers will be stockpiled and/or discharged directly to overland conveyors. The crushed ore stockpile will provide surge capacity to continue pad loading for up to 24 hours during primary crusher maintenance and stacking conveyor moves. A reclaim feeder will feed ore to the overland conveyor when the stockpile bypass is not active. Quicklime will be added for pH control on the HLP at a rate of 8 lb/ton. Relocated material from historic HLP operations will be hauled directly to the HLP and stacked via haul trucks.

The overland conveyors will then discharge onto two parallel jump conveyor strings. At the discharge point of both strings, there will be horizontal index conveyors to move ore to the radial stackers. The horizontal-radial stacking conveyor is coupled so that it can be moved in a retreat stacking mode without shutting down the system. The proposed stacking lift height is 30 feet. The conveyance capacity is currently sized at the primary crusher throughput.

 

14.3.2.3

Heap Leach Pad

The HLP is designed as a double lined system that consists of a layer of geosynthetic clay liner (GCL) and a layer of geosynthetic liner made from high density polyethylene (HDPE). A series of pregnant solution collection pipes will be installed in a “herring bone” arrangement to collect the pregnant leach solution (PLS) and direct it into the process pond. Overliner material will consist of crushed and screened ore and will provide both liner protection and provide adequate drainage for PLS. The overliner will be placed in a three-foot-thick layer over the liner and solution collection piping.

Ore will then be stacked, utilizing the conveyor system and radial stackers, in 30-foot lifts to a maximum of 250 feet. The leach pad will be constructed in five phases. Each phase is relatively similar in footprint size and will be constructed as needed to store additional ore based on the mine plan throughput and operational parameters including application rate and leach cycle.

 

14.3.2.4

Solution Management

After stacking, the piping heads and drip irrigation lines will be added to the HLP surface. Dilute sodium cyanide solution will be applied to the HLP surface via the header/drip system at a proposed application rate of 0.003 gpm/ft2 with a preliminary leach cycle of 90 days. The cyanide solution, at a nominal flow rate of 11,500 gpm, applied to the HLP surface will percolate though the HLP, being collected on the impervious leach pad liner. The PLS solution flows by gravity into the process pond via the solution collection piping system.

It is planned that the HLP solution application rate will be adjusted during the leach cycle to maximize the gold and silver recovery.

 

14.3.2.5

Process Ponds

The Project consists of the process pond and multiple event ponds, which will be constructed in phases in conjunction with the HLP phasing. The process pond and one of the event ponds will be part of initial construction. The process pond receives the PLS from the HLP via the solution collection piping system. Pumps will then transfer the PLS solution into the Merrill-Crowe for processing. The process pond consists

 

 

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of a lined system including leak detection. The process pond is designed to manage a 12-hr operational flow plus a 24-hr draindown event at the nominal flow rate, including freeboard.

The initial event pond and subsequent event ponds are designed as emergency ponds. The event ponds are sized to capture inflow from a 100-yr, 24-hr storm event matching the HLP footprint’s phased progression. The process pond is connected to the event ponds to manage solutions during upset conditions. Overflow would be directed into the event ponds avoiding release to the environment. The event ponds are also lined with geosynthetics.

 

14.3.2.6

Merrill-Crowe Plant and Refinery

Pregnant leach solution from the process pond will be pumped to the clarifier filter feed tank at the Merrill-Crowe plant. Solution clarification will be performed by clarifying filters arranged to operate in parallel. The clarified solution then proceeds to the deaeration tower, where it will be introduced into an evacuated chamber to remove as much dissolved oxygen as possible. After deaeration, powdered zinc, cyanide, and lead nitrate will be added to the solution to initiate an exchange redox reaction where zinc metal loses electrons to gold and silver, thereby reducing gold and silver to their metallic states and oxidizing zinc to form cyanide complexes in solution.

The gold and silver mixture will then be pumped to plate and frame filters operating in parallel. All the precipitated gold and silver will remain in the filter press until they are discharged when the filters are full. The filtrate solutions will report to the barren solution tank. Additional cyanide and caustic will be introduced, as required, into the barren solution tank before it is recycled to the HLP. Gold and silver precipitates collected by the filter presses will be dried in a retort to remove moisture and mercury before they are fluxed and smelted in an induction melting furnace. At the end of smelting, molten metal will be poured into bullion molds to produce doré bars. The doré bars will be shipped off-site for further refining.

 

14.3.2.7

Reagents

Crushed ore will utilize quicklime during stacking process to be utilized for pH control during the leach process. The proposed lime addition rate is 8 lb/ton.

Cyanide will be brought to the site in briquettes and mixed in batches on site utilizing a mixing skid. Cyanide solution will be added to the barren solution for dissolution of the precious metals in the ore during the leaching process. The LOM average cyanide consumption is 1 lb/ton.

The Merrill-Crowe process will utilize lead nitrate, zinc powder, and diatomaceous earth to further extract leached metals from the PLS.

 

14.3.3

Process Water

Process water makeup is estimated to be on the order of 800 gpm, for a total process requirement of 420 Mgal per year, which does not include infrastructure, facilities, or mining use.

 

14.3.4

Process Flowsheet

The Mineral Point process flowsheet is shown below.

 

 

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Figure 14-3: Mineral Point Process Flowsheet

(Source: Forte Dynamics, 2025)

 

 

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15.

INFRASTRUCTURE

 

15.1

Archimedes Underground

 

15.1.1

Operations Dewatering

Five active dewatering wells PW-9, PW-10, PW-11, PW-13, and PW-16 pump groundwater from the Archimedes block hydrogeologic unit at a combined average rate of approximately 250 gpm. Additionally, dewatering well PW-17 pumps approximately 70 gpm from the Holly block south of Archimedes pit (Figure 7-8). Discharge water is routed via a buried HDPE line to the RIBs for infiltration back into the downgradient alluvial basin aquifer. The dewatering well pump parameters are referenced from LRE 2025 and are listed in Table 15-1.

 

15.1.2

Operations Monitoring Wells and VWPs

Monitoring wells and VWPs are used to collect hydrogeological data in support of mining operations. Currently, there are 9 active monitoring wells and 47 active VWPs across 33 locations (Figure 7-8). Construction and recent water level data are provided in Table 15-2.

 

15.1.3

Operations RIBs

Water from the dewatering wells that is not utilized for operations is currently discharged to Rapid Infiltration Basins (RIBs) on the west side of the project area through HDPE pipelines. Two cells, RH-1 and RH-2 are in operation (NEV2005106), with discharge to one of the two cells at any given time. When RIB maintenance is required, discharge is routed to the dormant cell. Current dewatering efforts are well under the permitted 1,000 GPM threshold of the RIBs and the RIB infiltration is sufficiently limiting surface ponding in the active cell.

 

15.1.4

Operations Water Supply

A potable water well is located west of the Four Corners Road and supplies potable water to the Project. The well is completed in basin alluvial deposits to a depth of 265 ft and equipped with a pump capable of supplying 50 gpm.

 

15.1.5

Electrical Power

Ruby Hill is connected to the NVEnergy grid and has excess power available at the main project substation. An overhead power line will connect the underground transformer to the existing project near the East Archimedes Pit rim.

 

15.1.6

Underground Mine Facilities

The proposed location of portal site facilities is shown in Figure 15-1.

 

 

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Figure 15-1: Portal Surface Facilities Conceptual Layout

(Source: i-80 Gold, 2023)

 

15.1.7

Backfill

Backfill material for unconsolidated waste fill (GOB) can be obtained from any suitable source such as development waste, open pit waste dumps, or leach pads.

Backfill material for Cemented Rock Fill (CRF) will need to meet specifications designed to achieve minimum Uniaxial Compressive Strength (UCS) specifications. This specification is designed to provide the pillar strength needed to maintain stability of adjacent underground excavations and may require screening and/or crushing. CRF material will be mixed at a backfill plant located near the portal and transported underground using the same truck fleet used to remove mineralized material and waste from the mine.

 

 

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Table 15-1: Ruby Hill Active Dewatering Wells (LRE 2025)

 

Well ID    Collar Coordinates (Mine  Grid)    Casing
Diameter
   Well
Depth
   Static
Water
Level
   Pumping
Water
Level
   Screened
Interval (s)
   Average
GPM
   Pump
Power
   Pump
Set-Depth
   Northing    Easting    Elevation    in    ft bgs    ft bgs    ft bgs    ft bgs    GPM    HP    ft bgs

PW-9

   120109    12087    6462    12    1720    1002.1    1570    1200 to 1706    50    40    1650

PW-10

   11679    119741    6445    12    1720    986.8    1138    1000 to 1700    25    40    1650

PW-11

   10724    119800    6449    12    1720    982.9    1184    1200 to 1700    50    50    1620

PW-13

   13279    119974    6410.8    12    1816    1030    1548    1337 to 1800    95    75    1756

PW-16

   121261    11199    6510    12    1967    638    1944    800 to 1987    30    50    1911

PW-17

   117156    11557    6548    12    1820    664    1120.3    800 to 1800    75    60    1780

 

 

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Table 15-2: Summary of Locations, Construction Information, and Water Levels for Dewatering Wells, VWPs, Monitoring Wells, and Piezometers

 

Identifier    Coordinates    Surace
Elevation
   Year          Open Interval of Well or
VWP setting
   Static Water Level or
Hydraulic Head
     
   Easting
(ft)
   Northing
(ft)
   Incl    Depth (ft
bls)c
   Elevation (ft
amsl)
   Depth
(ft
bls)d
   Elevation
(ft
amsl)e
   Date    Comment
Dewatering Wells

PW-7

   120642.0    13254.0    6404.9    2008    -90    840 to 1700    5565 to 4705    820.2    5584.7    5/22/2024    Inactive due to high As, completed in limestone

PW-9

   12087.1    120109.4    6462.0    2010    -90    1200 to 1706    5262 to 4756    1649.7    4812.3    5/22/2024    Active, pump and motor replaced 11/2023

PW-10

   119741.0    11679.0    6445.0    2010    -90    1000 to 1700    5445 to 4745    1138.7    5306.3    5/22/2024    Active

PW-11

   119800.0    10724.0    6449.0    2010    -90    1200 to 1700    5249 to 4749    1152.1    5296.9    2/5/2024    Active

PW-13

   119974.0    13279.0    6410.8    2011    -90    1338 to 1800    5073 to 4611    1058.5    5352.3    10/23/2023    Active

PW-14

   117116.1    12467.3    6511.3    2012    -90    1078 to 1860    5433 to 4651    724.4    5786.9    10/9/2024    Inactive due to highwall failure; water level sensor still functioning; power supply no longer connected

PW-15

   8045.2    117833.0    6428.0    2011    -90    595 to 1200    5833 to 5228    652.2    5775.8    7/30/2011    Mineral Point well; inactive due to location outside Archimedes hydrogeologic block

PW-16

   11199.1    121260.6    6510.0    2013    -90    800 to 1980    5710 to 4530    783.0    5727.0    5/30/2023    Active, pump and motor replaced 11/2023

PW-17

   11557.0    117155.5    6548.0    2012    -90    800 to 1800    5748 to 4748    1082.0    5466.0    8/19/2024    Active, pump and motor replaced 1/2024

VWPsa

iRH22-17

   11953.0    121510.0    6505.1    ---    ---    ---    ---                    

iRH22-17D_5274

   ---    ---    ---    2022    -76    1269    5274    ---    ---    ---    Pressure sensor not functioning

iRH22-17C_4115

   ---    ---    ---    2022    -76    1639    4915    1110    5395.1    11/18/2024    Active

iRH22-17B_4352

   ---    ---    ---    2022    -76    2219    4352    1156    5349.1    11/18/2024    Active

iRH22-17A_3670

   ---    ---    ---    2022    -76    2922    3670    ---    ---    11/18/2024    Pressure sensor not functioning

IRH22-18a

   11859.5    119891.4    6447.8    ---    ---    ---    ---                    

IRH22-18aD_5231

   ---    ---    ---    2022    -87    1218    5231    753    5694.8    11/18/2024    Active

IRH22-18aC_4847

   ---    ---    ---    2022    -87    1603    4847    1127    5320.8    11/18/2024    Active

IRH22-18aB_4352

   ---    ---    ---    2022    -87    2099    4352    1123    5324.8    11/18/2024    Active

IRH22-18aA_3971

   ---    ---    ---    2022    -87    2480    3971    722    5725.8    11/18/2024    Active

iRH22-20

   10781.6    119805.4    6454.1    ---    ---    ---    ---                    

iRH22-21D_5108

   ---    ---    ---    2022    -89    1346    5108    769    5685.1    11/18/2024    Active

iRH22-21C_4863

   ---    ---    ---    2022    -89    1591    4863    1122    5332.1    11/18/2024    Active

iRH22-21B_4504

   ---    ---    ---    2022    -89    1950    4504    1306    5148.1    11/18/2024    Active

iRH22-21A_4146

   ---    ---    ---    2022    -89    2308    4146    2025    4429.1    11/18/2024    Active

iRH22-24

   12709.3    122509.7    6448.6    ---    ---    ---    ---                    

 

 

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Identifier    Coordinates    Surace
Elevation
   Year          Open Interval of Well or
VWP setting
   Static Water Level or
Hydraulic Head
     
   Easting
(ft)
   Northing
(ft)
   Incl    Depth
(ft bls)c
   Elevation
(ft amsl)
   Depth
(ft
bls)d
   Elevation
(ft
amsl)e
   Date    Comment

iRH22-24D_5181

   ---     ---     ---     2022    -85    1272    5181    1074    5374.6    11/18/2024    Active

iRH22-24C_4760

   ---     ---     ---     2022    -85    1695    4760    1114    5334.6    11/18/2024    Active

iRH22-24B_4235

   ---     ---     ---     2022    -85    2222    4235    817    5631.6    11/18/2024    Active

iRH22-24A_3883

   ---     ---     ---     2022    -85    2575    3883    ---     ---     11/18/2024    Pressure sensor not functioning

BRH-365

   10389.9    117205.9    6558    2011    -90    1011    5547    780.4    5777.6    5/22/2023    Active

BRH-399A

   9474.0    118027.2    6456    2011    -90    1200    5256    832.6    5623.4    5/22/2024    Active

BRH-403

   11117.4    117513.0    6516.7    2011    -90    1303    5214    1075.4    5441.3    10/9/2024    Active

BRH-405

   117962.5    13945.6    6492.1    2011    -90    1671    4821    659.2    5832.9    1/25/2021    Current status unknown, unable to access due to highwall failure

BRH-409

   10444.4    121429.4    6374    2011    -90    1296    5078    933.1    5440.9    5/22/2024    Active

BRH-411

   11053.5    117758.6    6518.7    2011    -90    1698    4821    1110.7    5408.0    10/9/2024    Active

BRH-435A

   8146.1    118021.0    6410    2011    -90    1200    5210    652.7    5757.4    5/23/2024    Active

BRH-435B

   8146.1    118021.0    6410    2011    -90    1000    5410    705.0    5705.0    5/23/2024    Active

BRH-435C

   8146.1    118021.0    6410    2011    -90    800    5610    638.3    5771.7    5/23/2024    Active

BRH-436C

   8105.0    117799.0    6410    2011    -90    800    5610    638.9    5771.1    5/23/2024    Active

BRH-437A

   8378.5    118155.3    6439    2011    -90    1200    5239    679.5    5759.5    9/25/2023    Active; logger needs to be replaced, but pressure sensor can be read manually

BRH-437B

   8378.5    118155.3    6439    2011    -90    1000    5439    674.4    5764.6    3/19/2024    Active

BRH-437C

   8378.5    118155.3    6439    2011    -90    800    5639    693.7    5745.3    9/25/2023    Active; logger needs to be replaced, but pressure sensor can be read manually

BRH-453

   12668.8    121965.8    6479    2012    -90    1829    4650    1155.6    5323.5    10/29/2024    Active; logger needs to be replaced, but pressure sensor can be read manually

BRH-455

   12815.2    120572.3    6455    2012    -90    1535    4920    1157.0    5298.0    5/22/2024    Active

BRH-517c

   8491.8    116558.0    6502    2013    -90    954    5548    733.1    5768.9    5/22/2024    Active

BRH-582

   8091.9    119049.9    6502    2013    -90    905    5597    740.0    5762.0    5/23/2024    Active

BRH-583

   8082.0    119045.3    6502    2013    -90    1195    5307    754.4    5747.6    5/23/2024    Active

BRH-584

   7369.5    118261.9    6548    2013    -90    883    5665    776.4    5771.6    5/23/2024    Active

BRH-585

   6955.7    118861.6    6548    2013    -90    1174    5374    908.4    5639.6    1/31/2024    Active; pressure sensor only reads periodically; likely failing over time

BRH-586

   6955.7    118861.6    6501    2013    -90    783    5718    596.2    5904.8    5/23/2024    Active

BRH-587

   6967.3    118842.7    6500    2013    -90    1176    5324    729.4    5770.6    5/23/2024    Active

BRH-590

   11281.2    117197.3    6507.5    2013    -90    1204    5303    504.4    6003.1    10/9/2024    Active

BRH-617

   8154.0    119545.0    6462    2013    -90    1159    5303    688.1    5773.9    5/23/2024    Active

BRH-618

   8143.0    119559.0    6462    2013    -90    759    5703    667.0    5795.0    3/19/2024    Active

 

 

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Identifier    Coordinates    Surace
Elevation
   Year          Open Interval of Well or
VWP setting
   Static Water Level or
Hydraulic Head
     
   Easting
(ft)
   Northing
(ft)
   Incl    Depth
(ft bls)c
   Elevation
(ft amsl)
   Depth
(ft
bls)d
   Elevation
(ft
amsl)e
   Date    Comment

PZ17-01 CH1-3

   3697.5    126571.9    6121    2017    -90    N/A    N/A    ---    ---    ---    ---

PZ17-01 CH1

   ---    ---    ---    ---    -90    292    5829    331.8    5789.2    10/9/2023    Active

PZ17-01 CH2

   ---    ---    ---    ---    -90    322    5799    322    5799.0    10/9/2023    Active

PZ17-01 CH3

   ---    ---    ---    ---    -90    442    5679    300    5821.0    10/9/2023    Active

PZ17-02 CH1-3

   2672.6    119186.4    6170    2017    -90    N/A    N/A    ---    ---    ---    ---

PZ17-02 CH1

   ---    ---    ---    ---    -90    971    5199    458.6    5711.4    10/9/2023    Active

PZ17-02 CH2

   ---    ---    ---    ---    -90    800    5370    ---    ---    ---    Pressure sensor not functioning

PZ17-02 CH3

   ---    ---    ---    ---    -90    440    5730    ---    ---    ---    Pressure sensor not functioning

PZ17-03 CH1-4

   10981.5    113605.9    6654    2017    -90    N/A    N/A    ---    ---    ---    Limestone/Dolomite

PZ17-03 CH1

   ---    ---    ---    ---    -90    1351    5304    523    6131.0    5/22/2024    Active

PZ17-03 CH2

   ---    ---    ---    ---    -90    1271    5384    518    6136.0    5/22/2024    Active

PZ17-03 CH3

   ---    ---    ---    ---    -90    1161    5494    518    6136.0    5/22/2024    Active

PZ17-03 CH4

   ---    ---    ---    ---    -90    811    5844    261    6393.0    5/22/2024    Sensor not functioning

PZ17-04 CH1-4

   7395.9    114594.6    6417    2017    -90    N/A    N/A    ---    ---    ---    Limestone/Dolomite

PZ17-04 CH1

   ---    ---    ---    ---    -90    1000    5417    656.2    5760.8    5/22/2024    Active

PZ17-04 CH2

   ---    ---    ---    ---    -90    800    5617    643    5774.0    5/22/2024    Active

PZ17-04 CH3

   ---    ---    ---    ---    -90    730    5687    692    5725.0    5/22/2024    Active

PZ17-04 CH4

   ---    ---    ---    ---    -90    680    5737    765    5652.0    5/22/2024    Active
 

Monitor Wells

 

Fad Shaft

   10890.0    111090.0    6911.5    ---    -90    1050    5862    1012.0    5899.5    3/18/2024    Active

HRH-1734

   12921.4    120577.1    6423.1    2003    -90    590 to 650    5833 to 5773    582.6    5840.5    5/22/2024    Active

HRH-1736

   11561.2    116791.1    6567    2010    -90    740 to 840    5827 to 5727    DRY    DRY    8/25/2022    Inactive

MW-2R

   14021.3    114979.8    6472.1    ---    -90    ---    ---    110.1    6362.0    7/15/2024    Active, located off Hogpen Road offsite

MW-3R

   6631.2    124018.8    6188.4    ---    -90    425 to 525    5763 to 5713    414.9    5773.5    7/16/2024    Active; 10-inch diameter casing

MW-4R

   11037.2    125060.3    6271.8    2013    -90    780 to 800    5492 to 5472    500.4    5771.4    7/15/2024    Active; completed in limestone

MW-7

   5613.0    121868.0    6169.8    ---    -90    ---    ---    281.6    5888.2    7/16/2024    Active

MW-8

   2647.3    121605.3    6159    ---    -90    260 to 300    5899 to 5859    278.5    5880.5    7/16/2024    Active

MW-9

   2576.6    123201.0    6124    1997    -90    260 to 300    5864 to 5824    240.0    5884.0    7/17/2024    Active

Notes:

  1.

feet above mean sea level; for wells, elevation of land surface at surface casing; for VWPs elevation of surface casing at land surface

  2.

degrees from horizontal at bottom of well or depth of VWP along inclined borehole using IDS survey

  3.

feet below land surface for wells; feet along inclined borehole for VWPs based on IDS inclination survey and Leapfrog Geologic Model positioning

  4.

feet below land surface for wells; feet below land surface of collar location for VWPs

  5.

feet above mean sea level

 

 

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  6.

Source LRE 2025

 

 

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15.2

Mineral Point Open Pit

The Mineral Point Project is identified as a 68,500 short tons per day (stpd) gold and silver secondary crush heap leach project with a Merrill Crowe processing plant. The Mineral Point Project located at the Ruby Hill site includes mining and mineral processing infrastructure that has been used in open pit mining and oxide gold heap leaching activities by previous owners.

 

15.2.1

Site Layout

The Project is located on the Battle Mountain/Eureka gold trend approximately 2 miles northwest of the town of Eureka in Eureka County, Nevada, USA, approximately 90 miles south of Elko and approximately 200 miles east of the city of Reno, Nevada. Figure 3-1 identifies the Project’s location. The Project is accessible by way of US-50.

Project infrastructure at Mineral Point is designed to support the mining, heap leaching, and processing facilities. There are sufficient and appropriate areas within the site to accommodate mining facilities to include waste rock storage area (WRSA), processing facilities, and all applicable storage facilities. Infrastructure that is essential to mining and metals production includes a crusher and conveyor circuit, stockpiles, access roads, haul roads, maintenance, storage area, and supporting ancillary facilities. Figure 15-2 displays the overall site map that identifies the Project’s major infrastructure.

 

 

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Figure 15-2: Site Layout Map

(Source: Forte Dynamics, 2025)

 

 

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15.2.2

Existing Infrastructure

The existing infrastructure on site supported previous mining and processing activities when the Archimedes Pit was an active mine. Figure 15-3 shows the locations of the existing infrastructure on site.

 

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Figure 15-3: Existing Infrastructure

(Source: Forte Dynamics, 2025)

The Project is designed to leverage existing infrastructure, aiming to reduce costs, minimize disturbance to new areas, and enhance the construction timeline. Table 15-3 outlines the intended use of the existing infrastructure during operations. The Project includes site access, access roads, and haul roads that can be utilized for future operations.

 

 

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Table 15-3: Existing Infrastructure Plans

 

 Existing Infrastructure    Planned Use Status for Mineral Point Project

Administration Area

   Utilize in Operations and Expand

Mill

   No Planned Use

Primary Crusher

   No Planned Use

Secondary Crusher

   No Planned Use

Tertiary Crusher

   No Planned Use

Heap Leach Pad (HLP)

   Spent material relocated to new HLP

Southwest Energy Building (Core Shack)

   Utilized by Explosives Contractor

Tire Pad

   Utilize in Operations

Warehouse

   Expand in Operations

Truck Shop

   Use for support equipment/expand in Operations

Fuel Island

   Expand in Operations

Waste Rock Storage Facility

   Expand in Operations

Power Supply

   Utilize in Operations/Upgrade & Improve if needed

 

15.2.3

Planned Infrastructure

The primary infrastructure for the Project includes several key components. The process system consists of the crushing and stacking system, the heap leach facility, Merrill Crowe, refinery, reagents, and waste rock storage area. The preproduction and facilities infrastructure covers utilities, mining support facilities, mine dewatering, and site improvements.

 

15.2.3.1

Process Infrastructure

The process infrastructure includes crushing, conveying, stacking, leaching, and Merrill Crowe processing of ore to recover metals. Once fresh mineralized material from the open pit or process material from the existing HLP is scheduled for processing, it is deemed ore.

 

15.2.3.1.1

Crushing, Conveying, and Stacking

Run-of-mine (ROM) ore will be transported from the pit to a primary crusher via 320 short ton (st) haul trucks. The haul trucks will direct dump into a gyratory crusher. The crushing plant will operate on average 20 hours/day and seven days a week. The primary crusher will provide a product size with a 100 percent passing size of 7”. The primary crusher product is discharged to the secondary crusher feed.

The secondary crushers, comprised of a set of four (4) cone crushers operating in parallel will produce material with 80 percent passing 0.75”. The secondary cone crusher product is discharged to the secondary product conveyor and then stockpiled or discharged onto the final product conveyor, where lime is added for pH control on the heap at a rate of 8 lb/ton.

The crushed ore stockpile will have enough capacity to feed the downstream heap stacking circuit, which will continuously operate 24 hours/day and seven days a week. The crushed ore stockpile will provide buffering capacity to minimize production loss during crusher maintenance and stacking conveyor moves. A reclaim feeder will feed ore to the overland conveyor where ore will be stacked, utilizing the conveyor system and radial stackers, in 30-foot lifts to a maximum of 250 feet. Heap Leach (HL) relocated material from historic operations will be hauled directly to the HLP and stacked via haul trucks.

The primary gyratory crusher will handle a maximum throughput of 3,900 short tons per hour (st/hr) and require a total power of 1,275 horsepower (hp). Each of the secondary cone crushers will have a maximum throughput of 1,000 st/hr, with a combined power requirement of 5,000 hp for all four units.

 

 

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15.2.3.1.2

Heap Leach Facility

The Mineral Point Project will include the construction of a new HLP with associated process and event ponds and solution management that together are referred to as the Heap Leach Facility (HLF), which is a closed system. The HLF will be located west of the proposed open pit.

The HLP is designed as a lined system that consists of a layer of geosynthetic clay liner (GCL), having a hydraulic conductivity less than or equal to 1x10-6 cm/s which acts as the secondary liner system. A layer of 80-mil geosynthetic liner made from high density polyethylene (HDPE) will be placed over the GCL to act as the primary liner for the liner system. A series of pregnant solution collection pipes will be installed in a “herring bone” arrangement to collect the pregnant leach solution (PLS) and direct it into the process pond. Overliner material will consist of crushed and screened ore and will provide both liner protection and provide a hydraulic conductivity of at least 1x10-1 cm/s. Over liner will be screened to 100% passing 2” and limited to a maximum of 10% passing 200 Mesh. The overliner will be placed in a three-foot-thick layer over the liner and solution collection piping.

The HLP will be constructed in five (5) phases. The footprint, capacity, and planned year of construction for each phase is presented in Table 15-4 below.

Table 15-4: Heap Leach Pad Phases

 

Phase    Footprint (million ft2)    Capacity (million tons)    Year Constructed

Phase 1 

   10.5    116.7    -1

Phase 2

   10.5    116.7    4

Phase 3

   10.5    116.7    7

Phase 4

   10.5    116.7    10

Totals

   42.1    466.8    13

Solution will be managed by a series of lined ponds. There will be one 200,000 square feet process pond which will be built in Phase 1 and will be able to hold 24.9 million gallons. This ponds liner system will include GCL and two layers of geosynthetics with a geonet in between to provide leak detection. The process pond will also include bird balls as a wildlife deterrent. There will be 5 total event ponds, one will be built each phase of pad expansion. These event ponds will be approximately 166,500 square feet and will hold 19.3 million gallons each. The event ponds have the capacity to capture inflow from a 100-year, 24-hour storm event. The ponds will be connected in series one overflowing into the next to prevent releases into the environment, as the heap leach process is a closed system. All ponds will be inside a fenced area to provide wildlife deterrents.

 

15.2.3.1.3

Merrill Crowe and Refinery

The Mineral Point Project will process the PLS from the HLP through a Merrill Crowe plant. The Merrill Crowe is designed to process the PLS at a rate of 11,500 gpm. Pregnant leach solution from the process pond will be pumped to the clarifier filter feed tank at the Merrill Crowe plant. Solution will be cleaned by clarifying filters arranged to operate in parallel. The clarified solution then proceeds to the deaeration tower, where it will be introduced into an evacuated chamber to remove as much dissolved oxygen as possible. After deaeration, powdered zinc, cyanide, and lead nitrate will be added to the solution to initiate an exchange redox reaction where zinc metal loses electrons to gold and silver, thereby reducing gold and silver to their metallic states and oxidizing zinc to form cyanide complexes in solution.

The gold and silver mixture will then be pumped to plate and frame filters operating in parallel. All the precipitated gold and silver will remain in the filter press until they are discharged when the filters are full. The filtrate solutions will report to the barren solution tank. Additional cyanide and caustic will be introduced

 

 

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to condition the barren solution tank before it is recycled to the HLP. Gold and silver precipitates collected by the filter presses will be dried in a retort to remove moisture and mercury before they are fluxed and smelted in an induction furnace. At the end of smelting, molten metal will be poured into bullion molds to produce doré bars. The doré bars will be shipped off-site for refining.

The Merrill Crowe and Refinery will be located on a concrete foundation providing secondary containment that will overflow into the event pond. The facility will be located inside a pre-engineered metal building.

A system of tanks, pumps, and piping will be installed to provide a cyanide mixing station to allow cyanide to be brought to site in briquette form, dissolved in water, and diluted to the specified concentration for addition to the HLF’s closed system.

 

15.2.3.1.4

Waste Rock Storage Area

The WRSA was designed with a 3:1 final slope ratio. Lift heights for the WRSA have not been finalized, however a strategy for determining them will be developed in subsequent stages of the Project. The current design is conservative given the current understanding. To assure competent foundation, and to salvage media for closure purposes, the growth media will be removed and stockpiled to an estimate 0.5 foot depth. The growth media stockpile will be located adjacent to WRSA, and clearing and grubbing will be completed in phased approaches as needed. The parameters of the WRSA can be found in Table 15-5 below.

Table 15-5: WRSA Parameters

 

WRSA Parameter    Value

Capacity

   886.7 million tons

Footprint

   75.1 million sq ft

Waste Rock Bulk Density

   118 lb/ft
Average Slope Ratio    3:1

The remaining waste tons will be placed into in-pit backfill areas. These areas will be identified as the pit phasing allows. These in-pit backfills will also be designed to a 3:1 slope ratio.

 

15.2.3.2

Preproduction and Facilities

Infrastructure required to support the mine and process, including utilities, ancillary facilities, and site improvements, are described in the below sections.

 

15.2.3.2.1

Utilities

Power

The site is currently connected to a power grid at a substation located along Highway 50. The Project is anticipated to utilize the same substation while requiring upgrades. The process only power load is currently estimated at 10 megawatts, and it is anticipated that a separate substation will serve the open pit and associated electric shovels.

Communications

The necessary communications infrastructure for the Mineral Point Project is assumed to be in place from the existing mine and/or from the Archimedes Project consisting of the following.

The connection to telephone and internet services has not been confirmed at this time; however, telephone service is available in the City of Eureka. A Cellular Telemetry System will be used to communicate data exchange between the Process Plant and administration building. The system will incorporate a Master Telemetry Station, located in a switch room of the Process Plant, and remote Telemetry Stations, located in remote equipment switchboards. The Master Telemetry Station will communicate with the Plant Process

 

 

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Control System via the preferred communications network and will communicate with the remote locations. Control of the remote equipment will be made by the Plant Process Control System, with sufficient data exchange to ensure correct operation of the equipment.

Fiber is an alternative that could be brought to site with relatively low cost by installing it in parallel with the overhead power line servicing this Project.

To ensure effective communication among personnel and equipment, a site-wide VHF radio network will be installed, equipped with multiple channels. Frequencies for this network will be assigned and approved by the Federal Communications Commission (FCC). This system will facilitate radio communication for both routine and emergency purposes, with mobile radios provided for operating and maintenance personnel to use outside office premises.

Potable Water

No improvements of the existing Potable Water system are anticipated for the Mineral Point Project. Eyewash stations will be self-contained units that can be refilled with bottled water.

Waste

Portable toilets will be used on site to accommodate the employees. Cleaning services will be sourced from a local company, which will also manage sewage disposal by transporting it to the local sewage treatment facility. For waste management, dumpster and roll-off bins will be utilized for garbage storage. These containers will be supplied by a local company, responsible for both their provision and the hauling of garbage to a nearby facility as required.

The routine generation of solid and hazardous waste, inherent in mining and processing activities, will be managed in compliance with local and state regulations.

Any hazardous waste generated at site would be placed in drums, on pallets, labelled, and stored in a designated location. The pallets would be placed in an area offering secondary containment where the material would be stored until it could be hauled offsite by a licensed contractor for appropriate disposal.

Fire Water

The Mineral Point Project will require a fire suppression system. It is foreseen that this system would be comprised of a large Fire/Freshwater tank located on site. This water will be used as make-up water for the process water supply, emergency firefighting supply, dust suppression, and water for the reagents make-up. The upper half of this tank will act as freshwater storage, and the lower half of the tank will be held in reserve for the fire suppression system. The fire water system will consist of a jockey pump, diesel pump, and electric pump. Fire water will be distributed to the site buildings through a distribution and sprinkler system. Additionally, strategically positioned fire hydrants on site will ensure easy access for local fire trucks.

 

15.2.3.2.2

Mining Support

Eureka offers standard municipal amenities including lodging and services, and a limited supply of food and hardware. The nearest major supply center is Elko, roughly 90 miles north of the Project area. Commercial air and rail services are both available in Elko. Rail access is also available in the community of Ely, roughly 60 miles east of the Project area. Unskilled and skilled labor can be found in Eureka, Ely, and a variety of other communities throughout the regional area.

 

15.2.3.2.3

Ancillary Facilities

Ancillary buildings necessary to support the Mineral Point Project include administration building, truck shop and warehouse building, an assay laboratory, and the main gatehouse and truck scale. Other facilities include a truck wash bay and an existing diesel storage and dispensing facility located outdoors. The

 

 

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Mineral Point Project will utilize as much of the existing infrastructure as possible, but also recognizes that with the size of the mine fleet, expansion of buildings and facilities will be required.

Water Management

Open pit stormwater management can be accomplished using a series of trenches and sumps from which water can be pumped. Additional dewatering wells and pumps will be required during active mining. The water recovered will be utilized for process make-up water and dust control.

Administration Building

The site has two existing administration buildings that total approximately 5000 square feet. An additional building will be constructed that is approximately 18,000 square feet, which will include change rooms and more admin space.

Truck Shop & Warehouse

The existing truck shop has three bays designed to service Cat 785 (150 short ton) haul trucks, which was built in the late 1990s. There is a small warehouse that is attached to the back of the truck shop, along with some office space.

The Project will utilize Komatsu 930E-5 (320 short ton) haul trucks which will necessitate extending the truck shop another four bays to accommodate the servicing of the additional, larger haul trucks. The three existing bays will be used to service support equipment and light vehicles. The warehouse will be extended once the new bays are added to the truck shop.

The new bays of the truck shop will be approximately 18,000 square feet, and the new warehouse area will be approximately 9,000 square feet.

Truck Wash

An existing outdoor wash pad will be utilized for light vehicles and support equipment, which will include spray monitors and wheel washes, a water heater, a sump for waste wash water, an oil-water separator, and a portable pressure washer. A new outdoor truck wash including spray monitors, water heater, and oil- water separator is included in Mineral Point Project. The footprint of this facility is 11,000 square feet and will be located next to the shop.

Assay Laboratory

The existing assay lab on site will be utilized for the Mineral Point Project and will process daily production blasthole samples from the mine, along with analytical data from samples in the processing plant. The lab building will be located on site.

Security & Truck Scale

The Mineral Point Project assumes that the facilities required for site security, including site access guard shack and gate, as well as perimeter fencing are currently installed or will be in place from the Archimedes Project. Fencing for process areas including refinery and solution ponds are included in the above process infrastructure.

Fuel Storage

An existing fuel supply area, originally designed to accommodate Cat 785 haul trucks (150 short tons), will be upgraded to support 320 short ton haul trucks. Haul trucks and fuel/lube trucks can easily pull onto the modified fuel pad for refueling before returning directly to the haul road.

Additionally, a secondary fuel island will be installed on the west side of the pit, closer to the crusher and waste dump location, minimizing out-of-cycle truck travel for fueling when the pit is in operation. To prioritize environmental safety, both fuel areas will include containment systems designed to capture any leaks or spills.

 

 

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Explosive Storage

Two explosive magazines will be required, one for boosters and high explosives, and one for detonators. There will be 4 Ammonium Nitrate (AN) storage bins with 100-ton capacity. These will occupy an area of 2,000 square feet. Bulk AN will be delivered to the site and AN-fuel mix trucks will be used for blast loading.

15.2.3.2.4 Site Improvements

Site Preparation

The following site preparations are included for the development of the Mineral Point Project surface infrastructure works area:

 

   

Disturbed areas include:

  o

Crusher area

  o

Access road and haul roads

  o

Powerline footprint

  o

Open Pits

   

Clearing and grubbing for disturbed areas as required; soil will be removed and stockpiled for use during site reclamation.

   

Cut and fill to prepare for disturbed areas; cut material will be reused for fill materials wherever possible. Bulk earthworks are designed to minimize the import of fill materials.

   

Site grading and road water management.

   

Installation of powerline and site water supply.

   

Installation of chain link and barbed wire fences on the site.

   

Access gates will be installed at the site entrance.

Stormwater Management

Stormwater run-off will be diverted away from disturbed areas of the Project. The Project will require diversions and ponds to adequately handle stormwater events. Contact stormwater will be collected in ponds, which may be used for makeup water in the processing facilities. The Project water balance will be prepared in the next level of study to design the pond volumes. It is anticipated that culverts will be required on the access and haul roads where drainages cross to prevent washouts. A diversion will need to be installed to collect offsite water and direct it around the WRSA and HLP.

Access & Haul Roads

This Project will require the rerouting of public roads around proposed facilities. The roads are gravel county roads estimated at approximately 4 miles long. The haul roads will need to be expanded to be able to accommodate 320 ton haul trucks and will have an estimated total additional 3.8 miles of roadway within and outside of the pit.

 

15.2.3.3

Geotechnical Review and Analysis

A geotechnical review and analysis for all proposed facilities is recommended for future study work.

 

15.2.4

Operations Dewatering

Previous dewatering operations in the mine area starting in the 1990s have identified multiple hydrologic blocks that segment bedrock groundwater levels in the immediate vicinity of the open pit operations as seen in Figure 15-4. The eastern portion of the Mineral Point open pit will share one hydrologic block with the western portion of the Archimedes open pit (i.e., the Williamsburg block) and then operate in four additional blocks (i.e., BC, Bullwhacker North, Bullwhacker South, and Spring Valley).

 

 

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Figure 15-4: Hydrologic Blocks of Mineral Point

(Source: JSAI, 2015)

Note: Pit contours reflect the design at the time of the dewatering model.

The predicted dewatering for Mineral Point ramps up to a peak rate of approximately 4,800 gallons per minute (gpm) at the end of mining (JSAI 2015). Dewatering will be achieved through a combination of pumping wells located on the pit perimeter and via in-pit groundwater seepage collection that will be pumped out of the pit by an in-pit booster station located at the working pit bottom. Inflows from the Archimedes Pit area towards the Mineral Point Pit will be controlled by continued operation of existing pumping wells PW-9, PW-10, PW-11, PW-13, PW-16, and/or PW-17. Another existing pumping well located near the center of the Mineral Point Pit (PW-15) will be utilized until it is mined out. Mineral Point dewatering simulations utilized four new pumping wells (one per Mineral Point hydrologic blocks; Table 15-6) to supplement in-pit dewatering efforts (JSAI 2015).

 

 

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Table 15-6: Ruby Hill Pumping Wells

 

Well    Status and
Hydrologic Block
   Northing
(mine grid)
   Easting (mine
grid)
   Collar
Elevation (feet
amsl)
   Well Depth
(feet bgs)
   Anticipated
Pumping Rate
(gpm)

PW-9

  

Existing, Archimedes

Block

   120109    12087    6462    1570    50

PW-10

  

Existing, Archimedes

Block

   119741    11679    6445    1138    25

PW-11

  

Existing, Archimedes

Block

   119800    10724    6449    1184    50

PW-13

  

Existing, Archimedes

Block

   119974    13279    6411    1548    95

PW-16

  

Existing, Archimedes

Block

   121261    11199    6510    1944    30

PW-17

  

Existing, Archimedes

Block

   117156    11557    6548    1120    75

PW-15

  

Existing, BC Block, to

be mined out

   1117861    8318    6428    1200    350

BC Well

   New, BC Block             ~2000    ~350

Bullwhacker

North Well

  

New, Bullwhacker

North Block

            ~2000    ~350

Bullwhacker

South Well

   New Bullwhacker South Block             ~2000    ~350

Spring Valley

Well

  

New, Spring Valley

Block

            ~2000    ~350

In-pit Booster

   New In-pit                ~3400

Approximately 10,000 feet of 12-inch to 24-inch diameter pipelines constructed from HDPE or steel will convey pumped water from the individual pumping wells and in-pit booster to a surface collection point for water treatment, as necessary. A portion of the dewatering water will be utilized as make-up water and dust suppression for the mine operations. The balance of the dewatering water will be conveyed from the mine area to rapid infiltrations basins (RIBs) where it will be artificially recharged into the Diamond Valley aquifer. This conveyance will utilize approximately four miles of 24-inch to 30-inch diameter HDPE pipelines to deliver water to one existing RIB site plus two additional new RIB sites.

An existing water treatment plant is used to lower arsenic concentrations in Archimedes Pit dewatering water prior to discharging to an existing RIB location approximately 3,000 feet northwest of the Mineral Point area. The existing RIB location consists of two basins that receive discharge from a conveyance pipeline. Past operations supporting the Archimedes Pit pumped and discharged an average of 300 gpm to the RIB with short-term peak discharge rates up to 900 gpm (FloSolutions 2021). Each individual basin has been able to independently manage discharges at these rates with limited surface ponding within the basin.

To accommodate the increased pumping associated with the Mineral Point Pit (i.e., 4,800 gpm), the water treatment plant capacity will be expanded as necessary to accommodate dewatering production from the new perimeter wells and in-pit sump in instances where those new dewatering sources have arsenic concentrations above regulatory standards. To artificially recharge the increased dewatering production, the existing artificial recharge system will need to be expanded (FloSolutions 2021) by constructing approximately two new RIB locations consisting of two to four basins each. These locations would be northwest and/or north of the Mineral Point Pit area on the alluvial fans that transition from the mine area to the Diamond Valley floor. This system of three RIB locations will be developed as dewatering production ramps up to allow for sustained infiltration for water management at the predicted pumping rates.

Total electrical power requirements to operate the dewatering system are anticipated to be approximately 1.5 megawatts (extrapolated from Piteau 2017 estimates). The pumping wells and in-pit sump will utilize

 

 

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line power or generators. The water treatment plant will utilize line power. The RIB locations do not require an electrical power source.

 

15.2.5

Operations Monitoring

There is an existing network of nine monitoring well locations plus 47 vibrating wire piezometers (VWPs) installed at 33 individual locations. These monitoring wells and VWPs provide water level data across the mine site within each of the hydrologic blocks as well as the local alluvial groundwater. The monitoring wells provide the ability to collect water samples for analytical laboratory testing to quantify groundwater geochemical conditions and metal concentrations.

The existing monitoring network will require minor expansion to account for the dewatering activities for the Mineral Point Pit once its operations commence. Approximately five to 10 additional piezometer locations will be needed to observe water levels in the five hydrologic blocks associated with Mineral Point Pit dewatering.

Prior to construction and operation of the two new RIBs, installation of three monitoring wells per RIB will be required to observe water levels and collect samples for water chemistry analyses. Monitoring will continue throughout RIB operation and closure periods. In accordance with Nevada Division of Environmental Protection Technical Publication WTS-3A (2017), each site will need one alluvial monitoring well hydraulically upgradient of the RIBs location (based on the pre-infiltration groundwater flow direction) and two alluvial monitoring wells hydraulically downgradient.

 

15.2.6

Water Supply

The potable water supply for workers on-site will be obtained from an existing potable water well and supply system. The existing 265-feet deep potable well is completed in alluvium northwest of the Mineral Point area and produces at a pumping rate of approximately 50 gpm.

Water supply for make-up water and dust suppression will be obtained from its existing dewatering well sources and/or new dewatering well installations. The dewatering pumping will be in excess of the planned consumptive use needs and the existing water rights authorization for consumptive use. Therefore, an additional temporary water rights authorization for Mineral Point Pit dewatering will be needed from the Nevada Division of Water Resources for pumping that does not represent a consumptive use of groundwater but involves dewatering pumping from the pit area followed by recharge of the aquifer via the RIBs.

 

 

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16.

MARKET STUDIES AND CONTRACTS

 

16.1

Precious Metal Markets

Gold and silver are fungible commodities with reputable smelters and refiners located throughout the world. The price of gold has reached all-time highs in 2024 with the Decembers price averaging 2,644 per ounce. As of December 2024 the three-year trailing average gold price was $2,044 per ounce and the two-year trailing average price was $2,166 per ounce. The three -year and two-year trailing average prices for silver in December 2024 were $24.50 and $25.88 per ounce respectively. Historical plots for both are shown in Figure 16-1.

 

LOGO

(Source: Practical Mining, 2025)

Figure 16-1: Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average

Issuers may also rely on published forecasts from reputable financial institutions. The current long term price forecast by CIBC is $2,169 and per ounce and $27.61 per ounce for gold and silver respectively (CIBC., 2025).

Commodity prices for Mineral Reserves are chosen not to exceed financial institution forecasts or the three-year trailing average price. Commodity pricing for the estimation of mineral resources can be 10% to 20% higher than that used for Mineral Reserves. The gold price selected for estimating mineral resources disclosed in this technical report is $2,175. The silver price selected is $27.25 per ounce.

 

 

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16.2

Contracts

 

16.2.1

Financing Agreements

Orion and Sprott Financing Package

The Company entered into a financing package with OMF Fund III (F) Ltd. an affiliate of Orion Mine Finance (collectively “Orion“) on December 31, 2021, and a fund managed by Sprott Asset Management USA, Inc. and a fund managed by CNL Strategic Asset Management, LLC (“Sprott”) on December 9, 2021 (together the “Finance Package”).

The Financing Package in its aggregate consists of:

 

  a.

$50 million convertible loan (the “Orion Convertible Loan“)

 

  b.

$10 million convertible loan (the “Sprott Convertible Loan” and together with the Orion Convertible Loan, the “Convertible Loans”)

 

  c.

$45 million gold prepay purchase and sale agreement entered into with affiliates of Orion (the “Gold Prepay Agreement“), including an accordion feature potentially to access up to an additional $50 million at i-80 Gold’s option

 

  d.

$30 million silver purchase and sale agreement entered into with affiliates of Orion (the “Silver Purchase Agreement“), including an accordion feature to potentially access an additional $50 million at i-80 Gold’s option and an amended and restated offtake agreement entered into with affiliates of Orion (the “A&R Offtake Agreement“)

 

  e.

5,500,000 warrants of the Company issued to Orion (the “Orion Warrants“ and together with the Orion Convertible Loan, Gold Prepay Agreement, Silver Purchase Agreement and the A&R Offtake Agreement, the “Orion Finance Package”).

Under the Gold Prepay Agreement, i-80 Gold was due to deliver to Orion 3,000 troy ounces of gold for each of the quarters ending March 31, 2022 and June 30, 2022, and thereafter, 2,000 troy ounces of gold per calendar quarter until September 30, 2025 in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold. i-80 Gold may request an increase in the $45 million prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the Gold Prepay Agreement.

The final Gold Prepay Agreement includes an amendment to adjust the quantity of the quarterly deliveries of gold, but not the aggregate amount of gold, to be delivered by the Company to Orion over the term of the Gold Prepay Agreement. Under the amended Gold Prepay Agreement, commencing on the date of funding, the Company is required to deliver to Orion 1,600 troy ounces of gold for the quarter ending March 31, 2022, 3,100 troy ounces of gold for the quarter ending June 30, 2022, and thereafter 2,100 troy ounces of gold per calendar quarter until September 30, 2025, in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold, subject to adjustment as contemplated by the terms of the Gold Prepay Agreement. As the funding from Orion did not occur until April 2022, payment for the delivery of 1,600 ounces for the quarter ending March 31, 2022 was offset against the $45 million of proceeds received from Orion.

Under the Silver Purchase Agreement, commencing April 30, 2022, i-80 Gold will deliver to Orion 100% of the silver production from the Granite Creek and Ruby Hill projects until the delivery of 1.2 million ounces of silver, after which the delivery will be reduced to 50% until the delivery of an aggregate of 2.5 million ounces of silver, after which the delivery will be reduced to 10% of the silver production solely from the Ruby Hill Project. Orion will pay i-80 Gold an ongoing cash purchase price equal to 20% of the prevailing silver price. Until the delivery of an aggregate of 1.2 million ounces of silver, i-80 Gold is required to deliver the following minimum amounts of silver (the “Annual Minimum Delivery Amount”) in each calendar year:

 

 

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(i) in 2022, 300,000 ounces, (ii) in 2023, 400,000 ounces, (iii) in 2024, 400,000 ounces, and (iv) in 2025, 100,000 ounces. Upon a construction decision for the Ruby Hill project, comprised of one or both of the Ruby Deep or Blackjack Deposits, which construction decision is based on a feasibility study in form and substance satisfactory to Orion, acting reasonably, i-80 Gold will have the right to request an additional deposit from Orion in the amount of $50 million in aggregate in accordance with the terms of the Silver Purchase Agreement.

Both the Gold Prepay Agreement and the Silver Purchase Agreement were funded on April 12, 2022 with i-80 Gold receiving net proceeds of $71.6 million after netting the aforementioned March 31, 2022 gold delivery and closing costs as further described in Note 10 and Note 24 in the Company’s Financial Statements.

The main amendments reflected in the A&R Offtake Agreement include the increase in the term of the agreement to December 31, 2028, the inclusion of the Granite Creek and Ruby Hill projects, and the increase of the annual gold quantity to up to an aggregate of 37,500 ounces in respect of the 2022 and 2023 calendar years and up to an aggregate of 40,000 ounces in any calendar year after 2023. During the year ended December 31, 2022, Orion assigned all of its rights, title and interest under the A&R Offtake Agreement to TRR Offtakes LLC, now Deterra Royalties Limited.

On September 20, 2023, the Company entered into an Amended and Restated (“A&R”) Gold Prepay Agreement with Orion, pursuant to which the Company received aggregate gross proceeds of $20 million (the “2023 Gold Prepay Accordion”) structured as an additional accordion under the existing Gold Prepay Agreement.

The 2023 Gold Prepay Accordion will be repaid through the delivery by the Company to Orion of 13,333 troy ounces of gold over a period of 12 quarters, being 1,110 troy ounces of gold per quarter over the delivery period with the first delivery being 1,123 troy ounces of gold. The first delivery will occur on March 31, 2024, and the last delivery will occur on December 31, 2026. Obligations under the A&R Gold Prepay Agreement, including the 2023 Gold Prepay Accordion, will continue to be senior secured obligations of the Company and its wholly-owned subsidiaries Ruby Hill Mining Company, LLC and Osgood Mining Company, LLC and secured against the Ruby Hill project in Eureka County, Nevada and the Granite Creek project in Humboldt County, Nevada.

The remaining terms of the A&R Gold Prepay Agreement remain substantially the same as the existing Gold Prepay Agreement. The Company may request an increase in the prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the A&R Gold Prepay Agreement.

In connection with the 2023 Gold Prepay Accordion, the Company issued to Orion warrants to purchase up to 3.8 million common shares of the Company at an exercise price of C$3.17 per common share until September 20, 2026, and extended the expiry date of 5.5 million existing warrants by an additional 12 months to December 13, 2025.

Orion Offtake

In February of 2025, i-80 Gold and Orion entered into an offtake agreement (the “Orion Offtake Agreement”). The Orion Offtake Agreement has similar terms to the current A&R Offtake Agreement with Deterra Royalties Limited and will commences upon its expiry. The Orion Offtake Agreement expires on December 31, 2034.

 

 

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South Arturo Purchase and Sale Agreement (Silver)

The Company entered into a Purchase and Sale Agreement (Silver) (the “Stream Agreement”) with Nomad, which was connected to South Arturo, whereby the Company will deliver to Nomad (i) 100% of the refined silver from minerals from the main stream area, and (ii) 50% of the refined silver from the exploration stream area. Nomad will pay an ongoing cash purchase price equal to 20% of the silver market price on the day immediately preceding the date of delivery and will credit the remaining 80% against the liability. Following the delivery of an aggregate amount of refined silver equal to $1.0 million to Nomad under the Stream Agreement, Nomad would continue to purchase the refined silver at an ongoing cash purchase price equal to 20% of the prevailing silver price. The liability for the Stream Agreement was included in the net asset value in connection with the asset exchange with Nevada Gold Mines LLC (“NGM”) discussed in the “Lone Tree and Ruby Hill Acquisition”, and therefore, is no longer impacting the Financial Statements as of December 31, 2021.

 

16.3

Refractory Mineralized Material Sale Agreement

Refractory mineralization mined prior to 2028 will be sold to a third party for processing under an existing agreement. Payment will be made for 58% of the contained gold at the average gold price realized during the month the material was processed. The processing agreement applies to all i-80 projects and allows a maximum purchase rate of 1,000 tons per day from all i-80 operations. The QP’s have reviewed this agreement and find the terms and conditions are in accordance with industry standard practice.

 

16.4

Other Contracts

The company also intends to negotiate contracts for underground mine development, production mining, and over-the-road haulage with reputable contractors doing business in northeast Nevada. At the time of this report these negotiations have not been initiated.

 

 

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17.

ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

 

17.1

Closure and Reclamation Requirements

The intent of the reclamation program for the Ruby Hill Project is to restore the project area to a beneficial post-mining land use, prevent undue or unnecessary degradation of the environment, and reclaim disturbed areas such that they are visually and functionally compatible with the surrounding topography. RHMC may choose to retain some facilities for post-mining use.

The BLM and the NDEP-BMRR are the primary federal and state agencies with regulations for the reclamation of surface mines in Nevada (43 CFR 3809, NRS 519A, and Nevada Administrative Code [NAC] 519A, respectively). These regulations were used in the development of the approved site-specific reclamation procedures.

The current estimated cost to close and reclaim the Project is approximately $27 million. The associated bond was accepted by the BLM on August 8, 2023 (RHMC, 2023).

The bond amount includes closure of all permitted mining and exploration disturbance at the Project, excluding the underground mining activities which are still in the permitting phase, and is calculated using standardized reclamation cost estimators that assess the following:

 

   

Exploration drill hole abandonment

 

   

Exploration roads and pads

 

   

Waste rock dumps

 

   

Heap leach pads

 

   

Roads

 

   

Pits

 

   

Foundations and buildings

 

   

Other demolition and equipment removal

 

   

Sediment and drainage control

 

   

Process ponds

 

   

Landfill

 

   

Yards

 

   

Waste disposal

 

   

Well abandonment

 

   

Underground portals closure

 

   

Miscellaneous costs

 

   

Monitoring

 

   

Construction management

 

   

Mobilization and demobilization.

There are no other known environmental liabilities associated with Project operations (RHMC, 2024).

 

17.2

Social or Community Impacts

The following information on community relations and stakeholder consultation is taken from Ruby Hill Mining Company (RHMC) personnel inputs in 2024.

Mining activity at the property began in the 1860s and has continued with periodic interruptions until the present day. Throughout its history, Ruby Hill has been a constant presence in the town of Eureka and has been an economic benefit to the community by offering employment, direct and indirect benefits.

 

 

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Ruby Hill and its predecessors, including Homestake Mining Company and Barrick Gold Corporation, have each maintained comprehensive community relations programs. Ruby Hill works closely with community and local stakeholders to provide updates on key developments, including:

 

   

Project status (operations and permitting)

 

   

Community program and initiatives.

Due to the proximity of the mine to the town, Ruby Hill diligently monitors:

 

   

Blasting

 

   

Noise

 

   

Light

 

   

Dust

 

   

Water Use

RHMC holds quarterly meetings with the public, landowners, and County officials to discuss operational status, safety and environmental compliance at the Project including monitoring, blasting schedules, and other matters of similar relevance to the Project’s neighbors. Additionally, Eureka is a community that is familiar with and supportive of mining. RHMC continues to have a positive professional relationship with its stakeholders, including its regulators at the federal and state agencies (RHMC, 2024).

 

17.3

Permits

In conjunction with the permitting actions associated with the Archimedes Underground Mine Project in-pit surface support facilities, a DNA was deemed sufficient for the PoO Amendment NVN-067782 approved by the BLM March 30, 2023. Additionally, on June 23, 2023, the NDEP-BMRR approved an EDC to WPCP NEV0096103 for the construction of the surface facilities. Permitting actions tied to mining of the underground are currently in progress with the BLM evaluating a PoO Amendment and associated EA while NDEP-BMRR is analyzing a WPCP Major Modification.

RHMC is currently permitted to carry out mining operations and reclamation activities at the Project site. This permitting allows it to carry out the exploration, geotechnical and metallurgical field work recommended in this Report. Specific permits related to site activities are presented in Table 17-1.

Table 17-1: Ruby Hill Project Significant Permits

 

Permit Name    Agency    Permit Number

Plan of Operations Amendment

   BLM    NVN-067782

Class II Air Quality Operating Permit

   NDEP-BAPC    AP1041-0713.05

Mercury Operating Permit to Construct

   NDEP-BAPC    AP1041-2252 (De Minimis)

Water Pollution Control Permit - Infiltration Project

   NDEP-BMRR    NEV2005106

Water Pollution Control Permit - Ruby Hill Mine

   NDEP-BMRR    NEV0096103

Mine Reclamation Permit

   NDEP-BMRR    0107

Mining Stormwater General Permit

   NDEP-BWPC    NVR300000: MSW-44886

Onsite Sewage Disposal System

   NDEP-BWPC    GNEVOSDS09L0107

Public Drinking Water System

   NDEP-BSDW    EU-0885-NTNC: NV0000885

Nitrate Removal System

   NDEP-BSDW    EU-0885-TP02: NV0000885

RCRA (Small Quantity Generator)

   NDEP-BSMM    RCRA ID / NVR000002899

Class III Wavered Landfill

   NDEP-BSMM    SWW362

Industrial Artificial Pond Permit

   NDOW    S-479016

Hazardous Materials Storage Permit

   Nevada State Fire Marshal    125455

Waters of the United States Jurisdictional Determination

   USACE   

Request for Approved Jurisdictional

Determination (AJD) submitted to

USACE November 2022

 

 

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17.4

Water Use Permits

RHMC controls a total of 8,107 acre feet per annum (AFA) of water rights for consumption and occupation (RHMC, 2024).

Due to a history of over pumping in the region based on a heavy agricultural reliance, the Diamond Valley Basin was categorized as a CMA by the Nevada State Engineer’s office in 2015. The designation allowed the State Engineer and the community to agree on certain tools to reduce over-pumping, including the implementation of a Diamond Valley GMP. Following resolution of a lengthy legal dispute by senior water rights holders in the Basin, the GMP was reinstated effective January 1, 2023. As a groundwater user within the GMP designated area, RHMC controls sufficient water rights to support its mining operations (RHMC, 2024).

 

17.5

QP Opinion

It is the opinion of the QP that the environmental program, will adequately address any issues related to environmental compliance, permitting, and local individuals or groups.

 

 

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18.

CAPITAL AND OPERATING COSTS

 

18.1

Archimedes Underground

 

18.1.1

Capital Costs

The Company intends to execute a contract mining agreement with a reputable firm for development and production mining at the Archimedes Underground Mine. The unit costs listed in this section are derived from similar mining contracts in northern Nevada. Contingencies include 15% on capital mine development and resource delineation drilling and 25% on all other capital.

Because of this and the infrastructure in place from previous mining activity on the property, capital requirements for the project are only for the construction of underground mine infrastructure, and underground development. The latter comprises 83% of total estimated capital expenditures. The unit rates for like development excavations are sourced from the Cove Underground Project Mine Development bids Table 18-1 (George, 2021). Table 18-2 details the timing and total of capital expenditures required for the Archimedes Underground Project. The final payment to Waterton of $20.0M is anticipated to occur in October 2023.

Table 18-1: Mine Development Unit Costs

 

Description   $/foot 1   

Primary Drifting (15’w x17’h)

  $2,000

Secondary Drifting (15’w x17’h)

  $2,000

Lined Raise Bore (10’ dia.)

  $4,000

Note: Excludes 15% Contingency

Table 18-2: Project Capital Costs ($M)

 

Item     Total      2025      2026      2027      2028      2029      2030      2031      2032      2033 

Mine Development

   100.0    7.8    21.0    12.8    25.1    21.7    2.6    3.7    3.0    2.4

Resource Conversion Drilling

   10.6    2.1    0.0    8.5    0.0    0.0    0.0    0.0    0.0    0.0

Facilities

                                                 

Environmental Permitting

   5.0    0.0    2.0    2.0    1.0    0.0    0.0    0.0    0.0    0.0

Feasibility Study

   0.5    0.0    0.0    0.5    0.0    0.0    0.0    0.0    0.0    0.0

Admin and Management

   3.9    3.9    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0

NV Energy

   1.4    0.8    0.6    0.0    0.0    0.0    0.0    0.0    0.0    0.0

Metallurgical Testing

   0.5    0.5    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0

Dewatering Wells

   3.9    3.9    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0

Contractor Mobilization

   0.2    0.2    0.0    0.0    0.0    0.0    0.0    0.0    0.0    0.0

Portal Construction

   0.2    0.1    0.1    0.0    0.0    0.0    0.0    0.0    0.0    0.0

Escape Hoist

   0.5    0.0    0.5    0.0    0.0    0.0    0.0    0.0    0.0    0.0

UG Electrical

   3.7    0.2    1.0    0.5    0.5    0.5    0.5    0.5    0.0    0.0

Fans/Ventilation

   2.8    0.0    1.2    0.2    0.4    0.4    0.3    0.3    0.0    0.0

Facilities Total

   22.6    9.7    5.3    3.2    1.9    0.9    0.8    0.8    0.0    0.0

Contingency

   22.3    2.9    4.5    4.5    4.7    3.5    0.6    0.7    0.5    0.4

Total Capital

   155.4    22.5    30.8    29.0    31.7    26.0    4.0    5.2    3.5    2.8

Note: Items inside the red box are considered sustaining capital.

 

18.1.2

Operating Costs

Underground operating costs are listed in Table 18-3. Underground mining unit costs are from similar northern Nevada mining contracts and include allowances for owner supplied materials and commodities. Other costs are i-80 estimates or supplier quotations.

 

 

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Table 18-3: Underground Mine Operating Costs

 

Item    Unit Cost    Units

Variable Costs

         

Stope Development Mining (15x20)

   $ 100.00    $/ton

Long Hole Stoping

   $ 80.00    $/ton

Sill Breasting (Floor Pull)

   $ 80.00    $/ton

Cemented Rockfill

   $ 37.93    $/fill ton

Unconsolidated Fill

   $ 13.00    $/ton

Lone Tree Pressure Oxidation - Acid

   $ 106.45    $/ton

Lone Tree Pressure Oxidation – Alkaline

   $ 70.81    $/ton

Over the Road Haulage – Lone Tree

   $ 39.15    $/wet ton

Over the Road Haulage – Third Party Sales

   $ 48.81    $/wet ton  

Crush, Screen and Agglomerate Heap Leaching

   $ 8.63    $/ton

Run of Mine Heap Leaching

   $ 2.41    $/ton

Electrical Energy

   $ 0.08    $/kw-hr

Electrical Demand

   $ 10.39    $/kw

Fixed Costs

         

Mine G&A

   $ 7.3M    $/year

Property Holding Costs

   $ 0.3M    $/year

Electrical Power

   $ 2.8M    $/year

Total Fixed Cost

   $ 10.4M    $/year

 

18.1.3

Cutoff Grade

Cutoff grades for pressure oxidation of refractory mineralization at Twin Creeks and on-site crush, screen and agglomerate leaching of oxide mineralization at Ruby Hill are shown in Table 18-4. For both mineralization types the mine is the production rate limiting factor and the mine limited cutoff grade is the correct cutoff grade to use. If mine production were to increase so that processing is the limiting factor, then the cutoff grade calculation must include fixed costs and sustaining capital.

Table 18-4: Resource Cutoff Grades by Process

 

     

 CSA Heap 

Leach

 

   3rd Party
 Sales 2025 - 
2027
   426 Zone
 Lone  Tree 
Acidic
  

 Ruby Deeps 

Zone Lone

Tree Acidic

Gold Price ($/oz)

   $2,175

Nevada Commerce and Excise Tax

   1.151%

Refining and Sales ($/oz)

   $1.85         $1.85     

Royalty

   3%

Recovery1

   88%    58%    96.8%    89.5%

Process Capacity (tpd)

   10,000    1,000    1,600    1,600

Mine Capacity (tpd)

   1,600

Mining Costs ($/ton)

   $145.88

Haulage Cost

   -    $48.81    $39.15     

Process Cost

   $8.63    -    $106.45     

Incremental Cutoff Grade (opt)

   0.005    0.040    0.072    0.078

Dilution Modifier

   5%

Mine Limited Cutoff Grade (opt)

   0.094    0.172    0.153    0.168

Fixed Costs ($ 000’syear)

   $10,404

Process Limited Cutoff Grade (opt)

   0.109    0.193    0.163    0.176

 

 

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18.2

Mineral Point Open Pit

The capital and operating costs used in this report were based on costs from similar project work performed recently by Forte Dynamics, high-level quotes from vendors, and interpolation from CostMine™ models. The QP believes that the estimates are appropriate for inclusion in this report and that these costs comply with the precision requirements for an Initial Assessment (IA).

 

18.2.1

Capital Cost Estimate

Mine construction capital, which includes all pre-production facilities and equipment, is estimated to total $708 million. This includes $299 million in mobile equipment for the initial fleet. In addition, approximately 115 Mtons (104 Mtonnes) of stripping is required in the first year of production to gain access to the mineralized material, with an incurred cost of $287 million. The life of mine (LOM) sustaining capital is estimated at $388 million, primarily for leach pad expansion and mobile equipment maintenance and rebuilds. Capital estimates included a contingency of 15% on all Mine Equipment and 25% on Process, Preproduction & Facilities, and Owner’s Cost.

Table 18-5 provides a summary of the capital costs by category for the Project.

Table 18-5: Mineral Point Project Capital Cost Summary

 

Category    US$M

Mining Equipment

   $420.7

Process

   $316.0

Preproduction & Facilities

   $80.1

Owner’s Cost

   $93.6

CAPEX Waste Stripping

   $287.3

Total Contingency

   $185.5
Total CAPEX    $1,383.2   

 

18.2.1.1

Mine Equipment Costs

The project is planned to be self-performed, requiring the owner to purchase the necessary mining fleet. Forte engaged with Komatsu and interpolated published data from CostMine™ to develop the capital cost for the mining fleet. Table 18-6 has a detailed list of mining equipment and the LOM CAPEX.

 

 

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Table 18-6: Mineral Point Mining Equipment LOM CAPEX

 

Equipment    # of Units     US$M per Unit     Total US$M 

Cable Shovel small

   1    $29.0    $29.0

Cable Shovel large

   1    $34.7    $34.7

Hydraulic Shovel

   2    $11.8    $23.6

Rear Dump Trucks

   26    $6.4    $165.1

Loader

   1    $9.9    $9.9

Rotary Drills

   5    $3.4    $17.0

Bulldozers

   5    $1.5    $7.3

Wheel Dozer

   2    $2.8    $5.6

Graders

   3    $0.2    $0.47

Water Tankers

   2    $5.4    $10.7

Backhoes Hydraulic

   2    $1.4    $2.7

Service/Tire Trucks

   16    $0.28    $4.4

Bulk Trucks

   3    $0.25    $0.75

Light Plants

   5    $0.02    $0.1

Pumps

   6    $0.03    $0.18

Pickups Trucks

   30    $0.06    $1.9

Sustaining CAPEX 

             $107.7

Contingency

             $63.1
Total CAPEX                $483.8

 

18.2.1.2

Process Infrastructure

Capital costs for the process infrastructure were estimated by scaling similar project work performed by Forte, obtaining high-level quotes from vendors, and/or interpolating published data from CostMine™. For costs of the crushers, conveyors, and stackers, the sizing was estimated using the total throughput of the processed material. The HLP includes bulk earthworks, liner systems, and overliner. The ultimate heap footprint sized to accommodate LOM is planned for five phases, each of similar footprints being constructed approximately every three years. The Process Ponds include one process pond, and five event ponds. One of the event ponds will be required at the onset of stacking of the pad, and an event pond will be constructed with associated Phases of the HLP. Ponds for the Merrill Crowe, both barren and pregnant, are included in costs of the Merrill Crowe line as well as the barren and pregnant pumps and the cyanide mixer. The refinery includes mercury retort. Waste Rock Storage Area (WRSA) Foundation preparation includes clearing and grubbing in five phases approximately every third year. Table 18-7 has a detailed list of process infrastructure items and the LOM CAPEX.

Table 18-7: Mineral Point Process Infrastructure LOM CAPEX

 

Category    US$M  

Crushers/Conveyers/Stacker

   $79.9

Heap Leach Pad

   $192.7

Process Ponds

   $8.0

Merrill Crowe

   $25.1

Refinery

   $7.6

WRSA Foundation Prep.

   $2.7

Contingency

   $79.0
Total CAPEX    $474.0

 

 

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18.2.1.3

Pre-Production and Facilities

Capital costs for the Pre-Production and Facilities infrastructure were estimated by scaling similar project work performed by Forte, obtaining high-level quotes from vendors, interpolating published data from CostMine™, and Client input as well as other mines in the area. The total cost for Utilities includes the pit substation and main substation. Mining Support includes the truck shop, truck wash, warehouse, fuel stations, and blasting supply storage. The Ancillary Facilities includes administrative building and dewatering system. The Site Improvements, stormwater management and rerouting public roads were considered. Table 18-8 has a detailed list of supporting infrastructure items and the LOM CAPEX.

Table 18-8: Mineral Point Pre-Production and Facilities LOM CAPEX

 

Category    US$M   

Utilities

   $4.3

Mining Support

   $37.2

Ancillary Facilities

   $30.6

Site Improvements

   $8.0
Total CAPEX    $120.2

 

18.2.1.4

Owner’s Costs

The owner’s costs were estimated to be 23% of the total process costs. This resulted in a total cost of $91.2 million from the IA CAPEX estimation for engineering and management. There were no estimates for permitting, reclamation/closure, and exploration. Table 18-9 provides a breakdown of the owner’s costs for the project.

Table 18-9: Mineral Point Owner’s Costs LOM CAPEX

 

Category    US$M  

Engineering/Management

   $93.6

Permitting

   $0

Reclamation/Closure

   $0

Exploration

   $0

Contingency

   $23.4
Total CAPEX    $114.0

 

18.2.2

Operating Cost Estimate

Operating costs for the mine were benchmarked against other similar Northern Nevada sites. The plant was estimated by scaling other simpler projects and interpolating published data from CostMine™. This gave a total cash cost (net of by-product credit) of $1,270.19 per Au toz produced. Table 18-10 provides a detailed breakdown of operating costs for the Project.

 

 

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Table 18-10: Mineral Point LOM Operating Cost Summary

 

Operating Costs    Unit    LOM (USD $M)    $/oz Au Produced

Mining to Process

   $2.50 per ton    $988.6    $280.11

Mining Heap Leach Relocation

   $1.50 per ton    $39.7    $11.24

Mining Waste

   $2.50 per ton    $2,846.1    $806.40

Processing

   $3.90 per ton    $1,542.2    $436.97

Mine Site G&A

   $0.75 per ton    $296.6    $84.03

Total Operating Costs:

        $5,713.2    $1,618.75

Refining Cost Au

   $1.85 per toz    $6.5    $1.85

Refining Cost Ag

   $0.50 per toz    $36.0    $10.20

Royalties & State Taxes

        $679.8    $192.6

Total Cash Costs:

        $6,435.5    $1,823.4

Silver Revenue (by-product)

   $27.25 per toz    $1,953.0    $553.21

Total Cash Cost (net of by-product credit)

        $4,482.5    $1,270.19

 

18.2.2.1

Mine Operating Costs

Open pit operating costs were developed by benchmarking other Northern Nevada sites of similar size and operation.

 

18.2.2.2

Mineral Processing Costs

Table 18-11 presents the estimated cost per ton of processed material by area. The number of personnel was estimated for each area, and salaries plus benefits typical of the Nevada mining industry were utilized for Labor estimates. Consumables cost was the most significant cost of processing, which is expected for a Nevada heap leach project. The cyanide cost estimated of $1.19/lb of reagent for briquettes totaled $1.18 per ton processed. The conservative dosing rate of 1 lb/ton determined from lab testing was utilized throughout the LOM. Quicklime consumption was estimated at $0.15/lb of reagent at a conservative dosing rate of 8 lb/ton, as determined from lab testing. Total quicklime consumption was $1.20/ton. Reagent consumptions including Zinc, Diatomaceous Earth, fluxes, anti-scalant, and other less significant reagents were estimated utilizing benchmark cost per ton processed. Maintenance costs were estimated by factors of the CAPEX for equipment of 8% per year, except for conveyors, which was 12% of CAPEX per year. The unit cost of power was $0.13/ kW-hr, estimating power was consumed at 85% of the installed name plate power.

Table 18-11: Mineral Point Processing Costs

 

Process    $/ton of Processed Material
Area    Labor    Consumables    Maintenance    Power    Total

Crushing

   $0.06    $0.07    $0.14    $0.16    $0.43

Stacking

   $0.05    $0.02    $0.06    $0.09    $0.22

Leaching Merrill Crowe

   $0.16    $2.93    $0.05    $0.09    $3.23
Total    $0.27    $3.02    $0.25    $0.35    $3.90

 

 

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19.

ECONOMIC ANALYSIS

 

19.1

Archimedes Underground

 

19.1.1

Taxes

 

19.1.1.1

Federal

The United States Government tax rate on corporations is 21% of taxable income. Taxable income is determined by offsetting revenue with depreciation, amortization, and depletion deductions. Unused depreciation and amortization deductions can be carried forward to the following year. The carryforward balance for the Ruby Hill project at the beginning of 2023 is $117.9M and the Fad property will add $60.0M. The net effect of all deductions reduces the federal tax liability to zero over the life of the project.

 

19.1.1.2

Nevada

Nevada does not have an Income tax, however, there are several other taxes that apply to all businesses and the net proceeds tax applies to mining companies specifically. Net mining proceeds are taxed at a rate of up to 5%. Net proceeds are generally defined as revenue less the costs of production. Capital investments are deductible using straight line depreciation over a 20-year period.

The state legislature enacted an excise tax that went into effect in 2022. The tax applies to gross revenue from the extraction of gold and silver. The tax is two tiered. Revenues greater than $20,000,000 and less than $150,000,000 are taxed at 0.75% while revenues above $150,000,000 are taxed at 1.1%.

Equipment and supplies for use in mining is subject to the sales and use tax. The tax rate for Eureka County is 6.85%.

The commerce tax is imposed on businesses with annual revenue exceeding $4,000,000. The commerce tax rate for mining companies is 0.051% of revenue above $4,000,000.

All employers subject to Nevada Unemployment Compensation is also subject to the Modified Business Tax (MBT) on total gross wages less employee healthcare benefits paid. The MBT rate is 1.378%. The first $50,000 of gross wages is exempt from MBT.

 

19.1.1.3

Property Taxes

Property or ad valorem taxes are based on the value of the property, both real and personal. The Nevada constitution caps the property tax rate at five dollars for every $1000 of assessed value. It is also capped by statute at $3.64 per $100 of assessed value. The assessed value in Nevada is 35% of the taxable value. Real and personal property taxes attributable to Ruby Hill Mining LLC and Golden Hill Mining LLC totaled $107,600.71 in 2024.

 

19.1.2

Cash Flow

A constant dollar cash flow analysis combining the mine production schedule presented in Section 13.1.6 combined with the commodity pricing of Section 16.1 and the capital and operating costs of Section 18 is presented in Table 19-1 and Table 19-2.

The Archimedes Underground production plan includes 70% inferred mineral resources. Inferred mineral resources are too speculative to be mineral reserves and the quantity and grade of inferred mineral resources may not be realized. The without inferred scenario presented in the income statement of Table 19-3 and the cash flow statement of Table 19-4 are a gross factorization of the production plan. There has been no adjustment to capital costs, productivities or unit operating costs.

 

 

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Table 19-1: Income Statement with Inferred

 

  Total 2025 2026 2027 2028 2029 2030 2031 2032 2033 2034 2035 2036 

Revenue

Gold Sales

2,018.3 0.0 12.5 74.5 211.1 260.7 259.8 249.1 237.1 263.4 266.7 172.7 10.6

Silver Sales

0.7 0.0 0.0 0.0 0.0 0.0 0.1 0.1 0.1 0.1 0.1 0.1 0.0

Total Revenue

2,018.9 0.0 12.5 74.5 211.2 260.7 259.9 249.2 237.3 263.6 266.8 172.7 10.6

Operating Costs

Mining

(750.0) 0.0 (6.9) (43.8) (80.9) (98.3) (104.1) (86.5) (103.4) (81.5) (84.2) (57.0) (3.4)

Surface Haulage to Mill

(192.2) 0.0 (1.4) (10.9) (17.2) (23.7) (23.0) (24.1) (24.2) (24.7) (24.6) (17.3) (1.0)

Processing

(490.0) 0.0 (0.1) (0.4) (47.3) (64.5) (62.5) (65.6) (65.9) (67.2) (66.8) (46.9) (2.8)

Electrical Power

(29.4) 0.0 (0.6) (1.5) (2.9) (3.4) (3.7) (4.0) (3.4) (2.6) (2.6) (2.5) (2.3)

Site G&A

(79.1) 0.0 (3.9) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5)

Total Operating

(1,540.7) 0.0 (12.9) (64.1) (155.9) (197.4) (200.8) (187.7) (204.4) (183.6) (185.7) (131.2) (17.1)

General & Administrative

Refining & Sales

(1.7) 0.0 (0.0) (0.1) (0.2) (0.2) (0.2) (0.2) (0.2) (0.2) (0.2) (0.1) (0.0)

Nevada Excise Tax

(60.6) 0.0 (0.4) (2.2) (6.3) (7.8) (7.8) (7.5) (7.1) (7.9) (8.0) (5.2) (0.3)

Royalty

(1.0) 0.0 (0.0) (0.0) (0.1) (0.1) (0.1) (0.1) (0.1) (0.1) (0.1) (0.1) (0.0)

Nevada Net Proceeds Tax

(21.7) 0.0 0.0 (0.6) (2.3) (2.9) (2.9) (2.7) (2.6) (2.9) (2.9) (1.9) 0.0

Nevada Commerce Tax

(16.9) 0.0 0.0 (0.2) (2.0) (2.3) (2.0) (2.2) (0.8) (3.1) (3.1) (1.3) 0.0

Total Cash Cost

(1,642.6) 0.0 (13.3) (67.2) (166.8) (210.7) (213.8) (200.4) (215.2) (197.8) (200.1) (139.8) (17.4)

EBITA

376.3 0.0 (0.8) 7.3 44.3 50.0 46.1 48.8 22.0 65.8 66.7 32.9 (6.8)

Reclamation Accrual (UOP)

(8.9) 0.0 (0.1) (0.3) (0.9) (1.1) (1.1) (1.1) (1.0) (1.2) (1.2) (0.8) (0.0)

Depreciation

(270.0) 0.0 (1.0) (7.2) (24.1) (33.7) (34.3) (33.9) (33.2) (37.9) (38.4) (24.8) (1.5)

Total Cost

(1,921.5) 0.0 (14.4) (74.7) (191.9) (245.6) (249.2) (235.4) (249.5) (236.8) (239.7) (165.4) (19.0)

Pre-Tax Income

97.4 0.0 (1.9) (0.2) 19.3 15.2 10.6 13.8 (12.2) 26.7 27.2 7.3 (8.4)

Income Tax

0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Net Income

97.4 0.0 (1.9) (0.2) 19.3 15.2 10.6 13.8 (12.2) 26.7 27.2 7.3 (8.4)

 

 

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Table 19-2: Cash Flow Statement with Inferred

 

  Total 2025 2026 2067 2028 2029 2030 2031 2032 2033 2034 2035 2036 2037-2045

Net Income

97.4 0.0 (1.9) (0.2) 19.3 15.2 10.6 13.8 (12.2) 26.7 27.2 7.3 (8.4) 0.0

Depreciation

270.0 0.0 1.0 7.2 24.1 33.7 34.3 33.9 33.2 37.9 38.4 24.8 1.5 0.0

Reclamation

0.0 (0.4) (0.3) (0.0) 0.5 0.7 0.7 0.6 0.6 0.7 0.7 0.3 (0.5) (3.5)

Working Capital

(0.0) 0.0 (1.5) (6.2) (11.5) (5.1) (0.4) 1.5 (1.7) 2.0 (0.3) 7.0 14.1 2.0

Operating Cash Flow

367.4 (0.4) (2.7) 0.7 32.4 44.5 45.3 49.9 19.8 67.3 66.0 39.4 6.7 (1.5)

Capital Costs

Capitalized Development

(100.0) (7.8) (21.0) (12.8) (25.1) (21.7) (2.6) (3.7) (3.0) (2.4) 0.0 0.0 0.0 0.0

Definition and Conversion Drilling

(10.6) (2.1) 0.0 (8.5) 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Mine Facilities

(22.7) (5.8) (5.3) (5.2) (3.9) (0.9) (0.8) (0.8) 0.0 0.0 0.0 0.0 0.0 0.0

Contingency

(22.3) (2.9) (4.5) (4.5) (4.7) (3.5) (0.6) (0.7) (0.5) (0.4) 0.0 0.0 0.0 0.0

Total Capital

(155.5) (18.6) (30.8) (31.0) (33.7) (26.0) (4.0) (5.2) (3.5) (2.8) 0.0 0.0 0.0 0.0

After Tax Cash Flow

211.9 (19.0) (33.5) (30.2) (1.3) 18.4 41.3 44.7 16.4 64.6 66.0 39.4 6.7 (1.5)

Cumulative Cash Flow

(19.0) (52.5) (82.7) (84.1) (65.6) (24.4) 20.4 36.7 101.3 167.2 206.7 213.4 1,918

 

 

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Table 19-3: Income Statement without Inferred

 

  Total 2025 2026 2027 2028 2029 2030 2031 2032 2033 2034 2035 2036  

Revenue

Gold Sales

604.5 0.0 3.7 22.3 63.2 78.1 77.8 74.6 71.0 78.9 79.9 51.7 3.2

Silver Sales

0.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Total Revenue

604.7 0.0 3.7 22.3 63.2 78.1 77.8 74.6 71.1 78.9 79.9 51.7 3.2

Operating Costs

Mining

(750.0) 0.0 (6.9) (43.8) (80.9) (98.3) (104.1) (86.5) (103.4) (81.5) (84.2) (57.0) (3.4)

Surface Haulage to Mill

(57.6) 0.0 (0.4) (3.3) (5.2) (7.1) (6.9) (7.2) (7.3) (7.4) (7.4) (5.2) (0.3)

Processing

(146.8) 0.0 (0.0) (0.1) (14.2) (19.3) (18.7) (19.6) (19.7) (20.1) (20.0) (14.1) (0.9)

Electrical Power

(29.4) 0.0 (0.6) (1.5) (2.9) (3.4) (3.7) (4.0) (3.4) (2.6) (2.6) (2.5) (2.3)

Site G&A

(79.1) 0.0 (3.9) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5) (7.5)

Total Operating

(1,062.9)  0.0 (11.8) (56.2) (110.7) (135.7) (140.9) (124.9) (141.3) (119.1) (121.7) (86.2) (14.4)

General & Administrative

Refining & Sales

(0.5) 0.0 (0.0) (0.0) (0.1) (0.1) (0.1) (0.1) (0.1) (0.1) (0.1) (0.0) (0.0)

Nevada Excise Tax

(18.1) 0.0 (0.1) (0.7) (1.9) (2.3) (2.3) (2.2) (2.1) (2.4) (2.4) (1.6) (0.1)

Royalty

(0.3) 0.0 0.0 (0.0) (0.0) (0.0) (0.0) (0.0) (0.0) (0.0) (0.0) (0.0) 0.0

Nevada Net Proceeds Tax

(4.5) 0.0 0.0 (0.2) (0.5) (0.6) (0.6) (0.6) (0.5) (0.6) (0.6) (0.4) 0.0

Nevada Commerce Tax

0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Total Cash Cost

(1,086.3) 0.0 (12.0) (57.1) (113.2) (138.7) (143.9) (127.7) (144.1) (122.2) (124.8) (88.2) (14.5)

EBITA

(481.6) 0.0 (8.2) (34.8) (49.9) (60.6) (66.0) (53.1) (73.0) (43.3) (44.9) (36.5) (11.3)

Reclamation Accrual (UOP)

(8.9) 0.0 (0.1) (0.3) (0.9) (1.1) (1.1) (1.1) (1.0) (1.2) (1.2) (0.8) (0.0)

Depreciation

(332.2) 0.0 (1.1) (8.4) (28.9) (41.2) (42.1) (41.9) (41.1) (47.1) (47.6) (30.8) (1.9)

Total Cost

(1,427.4) 0.0 (13.1) (65.8) (143.0) (181.1) (187.1) (170.7) (186.2) (170.4) (173.6) (119.8) (16.4)

Pre-Tax Income

(822.6) 0.0 (9.4) (43.5) (79.8) (103.0) (109.3) (96.1) (115.1) (91.5) (93.7) (68.1) (13.2)

Income Tax

0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Net Income

(822.6) 0.0 (9.4) (43.5) (79.8) (103.0) (109.3) (96.1) (115.1) (91.5) (93.7) (68.1) (13.2)

 

 

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Table 19-4: Cash Flow Statement without Inferred

 

  Total 2025 2026 2067 2028 2029 2030 2031 2032 2033 2034 2035 2036 2037-2045

Net Income

(822.6) 0.0 (9.4) (43.5) (79.8) (103.0) (109.3) (96.1) (115.1) (91.5) (93.7) (68.1) (13.2) 0.0

Depreciation

332.2 0.0 1.1 8.4 28.9 41.2 42.1 41.9 41.1 47.1 47.6 30.8 1.9 0.0

Reclamation

0.0 (0.4) (0.3) (0.0) 0.5 0.7 0.7 0.6 0.6 0.7 0.7 0.3 (0.5) (3.5)

Working Capital

0.0 0.0 (1.4) (5.2) (6.5) (2.9) (0.6) 1.9 (1.9) 2.5 (0.3) 4.2 8.5 1.7

Operating Cash Flow

(490.5) (0.4) (10.0) (40.3) (56.8) (64.0) (67.1) (51.7) (75.4) (41.2) (45.7) (32.7) (3.4) (1.8)

Capital Costs

Capitalized Development (100.0) (7.8) (21.0) (12.8) (25.1) (21.7) (2.6) (3.7) (3.0) (2.4) 0.0 0.0 0.0 0.0
Definition and Conversion Drilling (10.6) (2.1) 0.0 (8.5) 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Mine Facilities

22.7 5.8 5.3 5.2 3.9 0.9 0.8 0.8 0.0 0.0 0.0 0.0 0.0 0.0

Contingency

(22.3) (2.9) (4.5) (4.5) (4.7) (3.5) (0.6) (0.7) (0.5) (0.4) 0.0 0.0 0.0 0.0

Total Capital

(217.7) (26.1) (43.1) (43.4) (47.2) (36.5) (5.5) (7.2) (4.8) (3.9) 0.0 0.0 0.0 0.0

After Tax Cash Flow

(708.2) (26.4) (53.1) (83.7) (104.0) (100.5) (72.6) (59.0) (80.2) (45.1) (45.7) (32.7) (3.4) (1.8)

Cumulative Cash Flow

(26.4) (79.5) (163.2) (267.2) (367.7) (440.3) (499.3) (579.5) (624.6) (670.2) (703.0) (706.3)  (6,362.5)

 

 

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Mine production, processing and average head grade are shown in Figure 19-1. Annual gold production, cash costs and all in costs are displayed in the graph of Figure 19-2. The corresponding charts depicting results without inferred mineral resources are shown in Figure 19-3 and Figure 19-4.

 

LOGO

Figure 19-1: Mineralization Mined and Processed with Inferred

(Source: Practical Mining, 2025)

 

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Figure 19-2: Gold Production and Unit Costs with Inferred

(Source: Practical Mining, 2025)

 

 

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Figure 19-3: Mineralization Mined and Processed without Inferred

(Source: Practical Mining, 2025)

 

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Figure 19-4: Gold Production and Unit Costs without Inferred

(Source: Practical Mining, 2025)

 

 

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Table 19-5 shows life of mine total costs, cost per ton and cost per recovered gold ounce for the scenario containing inferred mineral resources. Table 19-6 presents the without inferred scenario.

Table 19-5: Capital and Operating Cost Summary With Inferred

 

Category    Total Cost      $/ton Processed     $/Au oz   

Mining

   $750    $148.98    $8068

Processing

   $490    $97    $528

Ore Haulage

   $192    $38    $207

Electrical Power

   $29    $5.84    $32

G&A, Prop Holding Costs

   $79    $16    $85

By Product Credits

   ($1)    ($0.13)    ($0.71)

Total Operating Costs

   $1,541    $306    $1,660

Refining

   $2    $0.34    $1.85

Royalty

   $61    $12    $65

Nevada Taxes

   $40    $0.20    $43

Cash Cost

   $1,642    $3.36    $1,769

Closure and Reclamation

   $8.9    $326    $10

Sustaining Capital

   $106    $21    $114

All in Sustaining Costs

   $1,757    $347    $1,893

Construction Capital

   $49    $9.81    $53

All in Costs

   $1,806    $359    $1,946

Table 19-6: Capital and Operating Cost Summary Without Inferred

 

Category    Total Cost    $/ton Processed     $/Au oz

Mining

   $748    $495.99    $2,690.79

Processing

   $147    $97.34    $528.10

Ore Haulage

   $58    $38.17    $207.08

Electrical Power

   $29    $19.51    $105.86

G&A, Prop Holding Costs

   $79    $52.46    $284.62

By Product Credits

   ($0)    ($0.13)    ($0.71)

Total Operating Costs

   $1,063    $704.89    $3,824.09

Refining

   $1    $0.34    $1.85

Royalty

   $18    $12.03    $65.27

Nevada Taxes

   $0    $0.19    $1.03

Closure and Reclamation

   $4    $2.97    $16.13

Cash Cost

   $0    $0.00    $0.00

Income Tax

   $1,086    $720.42    $3,908.37

Sustaining Capital

   $149    $98.52    $534.48

All in Sustaining Costs

   $1,235    $818.94    $4,442.85

Construction Capital

   $69    $45.85    $248.76

All in Costs

   $1,304    $864.80    $4,691.61

Annual undiscounted cash flows are depicted in the waterfall chart of Figure 19-5. The maximum cash draw of $68.1M occurs in 2026 with the project reaching the breakeven point two years later in 2029.

 

 

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Figure 19-5: Cash Flow Waterfall Chart with Inferred

(Source: Practical Mining, 2025)

 

 

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Table 19-7: Financial Statistics

 

Parameter    With Inferred    Without  Inferred

Gold price - base case (US$/oz)

   $2,175    $2,175

Silver price - base case (US$/oz)

   $27.25    $27.25

Mine life (years)

   10    10

Mining Rate (tons/day) 1

   1,600    400

Tons Processed Autoclave (ton)

   4,846    1,451

Average grade Autoclave (Au oz/ton)

   0.209    0.209

Average gold recovery (autoclave %) 2

   91.9%    91.9%

Autoclave Gold Produced (oz)

   910    272

Tons Processed Heap Leach (ton)

   188    56

Average Grade Heap Leach (Au oz/ton)

   0.111    0.111

Average gold recovery (Heap Leach %)

   87.4%    87.4%

Heap Leach Gold Produced (oz)

   18    5

Average annual gold production (koz)

   102    31

Total recovered gold (koz)

   928    272

Cash cost (US$/oz) 1

   $1,769    $3,908

Sustaining Capital (M$)

   $98    $149

All-in sustaining cost (US$/oz)1,3

   $1,893    $4,443

Pre Production Capital (M$)

   $49    $69

All in Costs (US$/oz) 3,4

   $1,938    $4,692

Project after-tax NPV5% (M$)

   $127    ($566.1)

Project after-tax NPV8% (M$)

   $91    ($501.4)

Project after-tax IRR

   23%    NA

Payback Period

   7.8 Years    NA

Profitability Index 8%2

   1.7    -1.8

Notes:    

  1.

Net of byproduct sales;

  2.

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates breakeven;

  3.

Excludes, construction capital, exploration, corporate G&A, interest on debt, and corporate taxes; and

  4.

Excludes exploration, corporate G&A, interest on debt, and corporate taxes; and,

  5.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable United States securities legislation. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. i-80 disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law.

 

 

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19.1.3

Sensitivity

The Ruby Hill Projects economic sensitivity to changes in gold price, operating costs and capital costs are shown in Figure 19-6 through Figure 19-9. The after tax cash flow breakeven gold price is $1,925 per ounce. A 13% increase in operating costs will also result in breakeven economics and a 139% increase in total capital expenditures is required to reduce the project economics to break even.

 

LOGO

Figure 19-6: NPV 5% Sensitivity with Inferred

(Source: Practical Mining, 2025)

 

LOGO

Figure 19-7: NPV 8% Sensitivity with Inferred

(Source: Practical Mining, 2025)

 

 

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Figure 19-8: IRR Sensitivity with Inferred

(Source: Practical Mining, 2025)

 

LOGO

Figure 19-9: Profitability Index Sensitivity with Inferred

(Source: Practical Mining, 2025)

 

 

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19.2

Mineral Point Open Pit

The economic analysis of the Mineral Point Project is based on the mining schedule, capital and operating costs, recovery parameters, and royalties outlined in earlier sections of this report. This is the initial Technical Report Summary (TRS), which incorporates inferred resources in the economic model. Section 19.2.6 shows the results of this economic analysis without inferred resources.

The economic results presented do not define a mineral reserve. Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. While the economic parameters used in this technical report are considered reasonable, additional information could alter these assumptions and affect the analysis. All figures are expressed in constant 2025 US dollars.

 

19.2.1

Principal Assumptions

The mine will utilize surface production only as of the time of this report.

Mineral processing is planned at 68,000 ton/day (62,000 tonne/day). The mine and plant will be operated by i-80 Gold Corp. personnel.

Table 19-8: Economic Model Parameters

 

Parameter      Unit        Value  

Discount Rate

   %    5%

Gold Price

   US$/toz    $2,175

Silver Price

   US$/toz    $27.25

Cash Reclamation

   US$M    $69.8

The Project uses a contingency of 15% for mining equipment and 20% for everything else, which is considered reasonable for an IA.

The model encompasses 1.0 year of production ramp-up with year 1 averaging 22.3 kton/day (20.2 ktonne/day), followed by 15 years at 68 kton/day (62 ktonne/day), ending with year 17 averaging 34.8 kton/day (31.6 ktonne/day) of processed material mined. A key input to the model is the mine schedule, detailed in Table 13-18, which outlines the grade and tonnage of the mined mineralized material. Revenue is derived from the amount of recovered metal, the specified metal price, and royalties incurred.

 

19.2.2

Operating Cost

Operating costs for the mine were benchmarked against other Northern Nevada sites. The plant was estimated by scaling other simpler projects and interpolating published data from CostMine™. The QP believes that these are appropriate for this level of preliminary study.

 

19.2.2.1

General and Administrative

General and Administrative (G&A) or overhead costs are the costs not directly incurred during production.

No camp facility is required at the Project and most overhead will be carried by the corporation, allowing a distribution of the costs between projects. G&A costs are estimated at $0.75/ton of processed material.

 

19.2.3

Capital Costs

Capital costs for the mining equipment, process plant, and facilities were estimated by scaling similar project work performed by Forte, obtaining high-level quotes from vendors, or interpolating published data from CostMine™. Mine construction capital, which includes all pre-production facilities and equipment, is estimated to total $708 million. This includes $299 million in mobile equipment for the initial fleet. In addition,

 

 

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approximately 115 Mtons (104 Mtonnes) of stripping is required in the first year of production to gain access to the body or mineralized material, costing $287 million. Life of mine (LOM) sustaining capital is estimated at $388 million, primarily for leach pad expansion and mobile equipment maintenance and rebuilds. The accuracy of the estimates is ±50%, and capital costs have a contingency of 15% on mining equipment and 25% on process, production, and facilities as well as owner’s costs.

 

19.2.4

Cost Summary

The costs used in the economic model are summarized in Table 19-9.

Table 19-9: Cost Summary

 

Prices      Unit        Value  

Gold Price

   US$/toz    $2,175

Silver Price

   US$/toz    27.25

Initial Capital

   US$M    $708

Sustaining Capital

   US$M    $388

Project Life

   Years    16.5
Production    Unit    Value

Total Mined Processed Material

   ktons    395,444

Total Heap Leach Relocation Material

   ktons    26,455

Total Mined Waste

   ktons    1,253,344

Total Mined Gold

   ktoz    4,525

Total Mined Silver

   ktoz    177,293

Au Grade

   toz/ton    0.0114

Au Grade

   g/tonne    0.391

Ag Grade

   toz/ton    0.4483

Ag Grade

   g/tonne    15.37
Operating Cost    Unit    Value

Open Pit Mining Cost

   US$/ton    $2.50

Process Cost

   US$/ton    $3.90

Heap Leach Relocation Mining Cost

   US$/ton    $1.50

G&A Cost

    US$/ton processed     $0.75

Royalty

   %    3.0%

 

19.2.5

Economic Model

A summary of the economic model is provided in Appendix B. Additionally, a high-level summary of the Pre-Tax Net Present Value (NPV) is provided in Table 19-10, and the After-Tax summary is included in Table 19-11. Figure 19-10 shows the undiscounted Pre-Tax LOM annual cash flow.

Thirty eight percent (38%) of the material considered for mineral processing is classified as inferred mineral resources. This analysis includes inferred mineral resources, which are considered too speculative geologically to apply modifying factors that would enable them to be classified as mineral reserves, and there is no certainty that this economic assessment will be realized. The detailed analysis of Mineral Point without inferred mineral resources is detailed in section 19.2.6.

 

 

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Table 19-10: Pre-Tax NPV Summary

 

Pre-Tax NPV    US$M

NPV @ 0%

   $1,854.50

NPV @ 5%

   $827.58

NPV @ 8%

   $451.23

NPV @ 10%

   $262.46

NPV @ 12%

   $110.22

IRR

   13.8%

Payback Period

   8.7 years

 

LOGO

Figure 19-10: Pre-Tax LOM Annual Cash Flow

(Source: Forte Dynamics, 2025)

 

19.2.5.1

Taxes and Royalties

Royalties are discussed in detail in Section 3.3. Taxes are calculated as required for a project in Nevada. A summary of the After-Tax NPV is included in Table 19-11. The Project will pay a total of US $263.1 million dollars in federal taxes and a total of US $234.8 million in state taxes during the life of mine. Figure 19-11 shows the undiscounted After-Tax LOM annual cash flow.

Table 19-11: After-Tax NPV Summary

 

After-Tax NPV    US $M

NPV @ 0%

   $1,470.0

NPV @ 5%

   $614.1

NPV @ 8%

   $295.8

NPV @ 10%

   $134.8

NPV @ 12%

   $4.3

IRR

   12.1%

Payback Period

   7.9 years

 

 

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Figure 19-11: After-Tax LOM Annual Cash Flow

(Source: Forte Dynamics, 2025)

 

19.2.5.2

Sensitivity Analysis

A sensitivity analysis was conducted on the parameters of capital cost, operating cost, and metal price, all assessed on a Pre-Tax and After-Tax basis. A summary of these sensitivities is shown in Table 19-12. Figure 19-12 and Figure 19-13 show the sensitivity of NPV @ 5% and IRR Pre-Tax. Figure 19-14 and Figure 19-15 show the sensitivity of NPV @ 5% and IRR After-Tax.

 

 

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Table 19-12: Sensitivity Summary

 

Parameter

   Item    Pre-Tax Sensitivity    After-Tax Sensitivity
   -25%    0%    +25%    -25%    0%    +25%

Gold

   Price (US$/toz)    $1,631    $2,175    $2,719    $1,631    $2,175    $2,719
   NPV @5% (US$M)    $(211.2)    $827.6    $2,274.4    $(395.6)    $614.1    $1,523.0
   NPV @8% (US$M)    $(398.9)    $451.2    $1,610.0    $(543.7)    $295.8    $1,002.4
   NPV @10% (US$M)    $(489.0)    $262.5    $1,272.9    $(613.9)    $134.8    $740.0
   NPV @12% (US$M)    $(559.1)    $110.2    $998.4    $(667.8)    $4.3    $527.6
   IRR (%)    2.6%    13.8%    26.3%    0.3%    12.1%    19.9%

Silver

   Price (US$/toz)    $20.44    $27.25    $34.06    $20.44    $27.25    $34.06
   NPV @5% (US$M)    $722.1    $827.6    $1,341.0    $346.5    $614.1    $828.4
   NPV @8% (US$M)    $358.0    $451.2    $853.2    $55.8    $295.8    $439.4
   NPV @10% (US$M)    $176.0    $262.5    $607.9    $(88.2)    $134.8    $245.4
   NPV @12% (US$M)    $29.6    $110.2    $409.8    $(203.2)    $4.3    $89.7
   IRR (%)    12.5%    13.8%    18.1%    8.7%    12.1%    13.4%

CAPEX

   Price (US$M)    $1,037    $1,383    $1,729    $1,037    $1,383    $1,729
   NPV @5% (US$M)    $1,274.1    $827.6    $789.0    $834.1    $614.1    $349.1
   NPV @8% (US$M)    $835.5    $451.2    $375.7    $480.6    $295.8    $20.8
   NPV @10% (US$M)    $615.2    $262.5    $168.7    $304.4    $134.8    $(142.1)
   NPV @12% (US$M)    $437.3    $110.2    $2.0    $163.0    $4.3    $(272.3)
   IRR (%)    19.8%    13.8%    12.0%    15.0%    12.1%    8.2%

Mining Cost

   Price (US$/ton)    $1.88    $2.50    $3.13    $1.88    $2.50    $3.13
   NPV @5% (US$M)    $1,762.1    $827.6    $289.3    $1,263.5    $614.1    $(92.9)
   NPV @8% (US$M)    $1,218.6    $451.2    $(17.3)    $815.5    $295.8    $(324.5)
   NPV @10% (US$M)    $942.5    $262.5    $(167.5)    $589.1    $134.8    $(436.0)
   NPV @12% (US$M)    $717.6    $110.2    $(286.3)    $405.3    $4.3    $(522.7)
   IRR (%)    23.3%    13.8%    7.8%    18.6%    12.1%    4.1%

Processing Cost

   Price (US$/ton)    $2.93    $3.90    $4.88    $2.93    $3.90    $4.88
   NPV @5% (US$M)    $1,283.8    $827.6    $776.7    $787.3    $614.1    $392.0
   NPV @8% (US$M)    $808.1    $451.2    $401.0    $407.2    $295.8    $91.2
   NPV @10% (US$M)    $568.9    $262.5    $213.2    $217.7    $134.8    $(57.8)
   NPV @12% (US$M)    $375.7    $110.2    $62.1    $65.6    $4.3    $(176.8)
   IRR (%)    17.6%    13.8%    13.0%    13.0%    12.1%    9.2%

The Project’s NPV and IRR in relation to fluctuations in the long-term gold and silver price are outlined in Table 19-13. Based on the economic sensitivity study, the Project is robust regarding both capital and operating costs. It is most sensitive to metal price and, by direct correlation, to metal recovery.

 

 

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Figure 19-12: Pre-Tax Sensitivity NPV @5%

(Source: Forte Dynamics, 2025)

 

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Figure 19-13: Pre-Tax Sensitivity IRR

(Source: Forte Dynamics, 2025)

 

 

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Figure 19-14: After-Tax Sensitivity NPV @5%

(Source: Forte Dynamics, 2025)

 

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Figure 19-15: After-Tax Sensitivity IRR

(Source: Forte Dynamics, 2025)

 

 

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Table 19-13: Gold and Silver Price Sensitivity After-Tax Analysis

 

     Gold Price (US$/toz)
   $ 2,000    $ 2,175    $ 2,500    $ 2,750    $ 2,900    $ 3,000
   NPV 5%    IRR    NPV 5%    IRR    NPV 5%    IRR    NPV 5%    IRR    NPV 5%    IRR    NPV 5%     IRR 

Silver Price

(US$/oz)

   $ 25.00    $  218    8%    $  540    11%    $ 1,126    18%    $ 1,573    22%    $ 1,840    25%    $ 2,017    26%
   $ 27.25    $  294    8%    $  614    12%    $ 1,199    18%    $ 1,647    23%    $ 1,913    25%    $ 2,091    27%
   $ 30.00    $  387    10%    $  705    13%    $ 1,286    19%    $ 1,737    24%    $ 2,001    26%    $ 2,181    28%
   $ 32.75    $  479    11%    $  795    14%    $ 1,377    20%    $ 1,826    24%    $ 2,092    27%    $ 2,270    28%
   $ 35.00    $  554    11%    $  869    15%    $ 1,450    21%    $ 1,899    25%    $ 2,164    27%    $ 2,343    29%

 

19.2.6

Economic Analysis Without Inferred Resources

To comply with S-K 1302.d.iii.A.4, “inferred mineral resources may be included in a preliminary analysis to demonstrate economic potential” if the registrant discloses among other items; the percentage of inferred mineral resources, and the economic analysis excluding the inferred mineral resource. Thus, Forte created a second mine schedule, capital cost, and economic analysis without inferred resources, which comprise 38% of the material processed in the full scenario.

The removal of inferred resources resulted in phases 5 through 9 being removed from the mine plan. This removal also decreased the processing tons from 25Mton (22.7Mtonnes) per year to 18Mton (16.3Mtonnes) per year. Table 19-14 shows the difference in parameters between the two mine plans and economic models. If the parameter is not shown, then it was held constant between the two models. A high-level summary of the After-Tax Net Present Value (NPV) of both models is provided in Table 19-15.

Table 19-14: Economic Model Parameters Comparison of With and Without Inferred Resources

 

Parameter      Unit       Value With  Inferred      Value Without  Inferred 

Mine Life

   year    16.5    11.5

Mining Rate

   kton/day    356.2    328.8

Processing Rate

   kton/day    68.4    49.3

Total Processed Material

   kton    395,444    195,591

Total Mine Material

   kton    1,675,243    987,993

Average Processing Grade Au

   toz/ton    0.011    0.012

Average Processing Grade Ag

   toz/ton    0.448    0.383

Contained Au

   ktoz    4,525    2,430

Contained Ag

   Ktoz    177,293    76,109

Recovered Au

   ktoz    3,529    1,969

Recovered Ag

   ktoz    72,028    31,407

Heap Leach Recovery Au (average)

   %    78%    81%

Heap Leach Recovery Ag (average)

   %    41%    41%

Total LOM CAPX

   US$M    $1,383.2    $941.2

Table 19-15: After-Tax NPV Comparison of With and Without Inferred Resources

 

After-Tax NPV      Unit       Value With  Inferred      Value Without  Inferred 

NPV @ 0%

   US$M    $1,470.0    $574.1

NPV @ 5%

   US$M    $614.1    $157.9

NPV @ 8%

   US$M    $295.8    $(10.9)

NPV @ 10%

   US$M    $134.8    $(100.1)

NPV @ 12%

   US$M    $4.3    $(174.8)

IRR

   %    12.1%    7.8%

Payback Period

   Year    7.9    8.9

 

 

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19.2.6

Economic Analysis Without Inferred Resources

To comply with S-K 1302.d.iii.A.4, “inferred mineral resources may be included in a preliminary analysis to demonstrate economic potential” if the registrant discloses among other items; the percentage of inferred mineral resources, and the economic analysis excluding the inferred mineral resource. Thus, Forte created a second mine schedule, capital cost, and economic analysis without inferred resources, which comprise 38% of the material processed in the full scenario.

The removal of inferred resources resulted in phases 5 through 9 being removed from the mine plan. This removal also decreased the processing tons from 25Mton (22.7Mtonnes) per year to 18Mton (16.3Mtonnes) per year. Table 19-14 shows the difference in parameters between the two mine plans and economic models. If the parameter is not shown, then it was held constant between the two models. A high-level summary of the After-Tax Net Present Value (NPV) of both models is provided in Table 19-15.

Table 19-14: Economic Model Parameters Comparison of With and Without Inferred Resources

 

Parameter      Unit       Value With  Inferred      Value Without  Inferred 

Mine Life

   year    16.5    11.5

Mining Rate

   kton/day    356.2    328.8

Processing Rate

   kton/day    68.4    49.3

Total Processed Material

   kton    395,444    195,591

Total Mine Material

   kton    1,675,243    987,993

Average Processing Grade Au

   toz/ton    0.011    0.012

Average Processing Grade Ag

   toz/ton    0.448    0.383

Contained Au

   ktoz    4,525    2,430

Contained Ag

   Ktoz    177,293    76,109

Recovered Au

   ktoz    3,529    1,969

Recovered Ag

   ktoz    72,028    31,407

Heap Leach Recovery Au (average)

   %    78%    81%

Heap Leach Recovery Ag (average)

   %    41%    41%

Total LOM CAPX

   US$M    $1,383.2    $941.2

Table 19-15: After-Tax NPV Comparison of With and Without Inferred Resources

 

After-Tax NPV      Unit       Value With  Inferred      Value Without  Inferred 

NPV @ 0%

   US$M    $1,470.0    $574.1

NPV @ 5%

   US$M    $614.1    $157.9

NPV @ 8%

   US$M    $295.8    $(10.9)

NPV @ 10%

   US$M    $134.8    $(100.1)

NPV @ 12%

   US$M    $4.3    $(174.8)

IRR

   %    12.1%    7.8%

Payback Period

   Year    7.9    8.9

 

 

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20.

ADJACENT PROPERTIES

There are no adjacent properties relevant to The Ruby Hill Project.

 

 

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21.

OTHER RELEVANT DATA AND INFORMATION

The authors are not aware of any other relevant data or information.

 

 

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22.

INTERPRETATION AND CONCLUSIONS

 

22.1

Conclusions

The authors have reviewed the data from the project, which include Archimedes Underground, Archimedes Open Pit, and Mineral Point Open Pit, and undertook verification of the data that are material to this report. Based on the work completed or supervised by the authors, it is the opinion of the authors that the project data are of sufficient quality for the modeling, estimation, and classification of the gold and silver resources disclosed in this report, as well as for the completion of the Technical Report summarized herein. Furthermore, the authors are unaware of any significant risks or uncertainties that could reasonably be expected to affect the reliability of the current mineral resources.

The economic analysis presented in this Initial Assessment is an evaluation of the Archimedes Underground and Mineral Point Open Pit mineral resources and do not have demonstrated economic viability. The Archimedes Open Pit is a mineral resource and does not include economic analysis.

 

22.1.1

Archimedes Underground

 

22.1.1.1

Mineral Resources

The Archimedes Underground mineral resource contains approximately 70% inferred mineral resources. The planned underground development and drilling program is planned to upgrade inferred mineral resources to indicated.

 

22.1.1.2

Mining and Infrastructure

Mining conditions for the Archimedes underground are typical for sedimentary deposits in the north-east Nevada extensional tectonic environments are anticipated. The Ruby Deeps deposit will require dewatering with anticipated pumping rates of 500 to 1,000 gpm.

 

22.1.1.3

Metallurgical Testing

Metallurgical testing of refractory samples from Archimedes underground deposits has confirmed amenability to grinding followed by pressure oxidation and carbon in leach. Gold recoveries ranged from 80% to 91%. Metallurgical testing programs have identified deleterious elements that are common to deposits in this part of Nevada. Deleterious elements content in the oxide samples are low, while sulfide samples are characterized by high levels of sulfide sulfur, arsenic, and mercury. Processing of Archimedes sulfide mineralization through a third-party or i-80’s Lone Tree autoclave will ensure removal and capture of these deleterious elements.

 

22.1.1.4

Recovery Methods

Metallurgical testing has confirmed that processing of Archimedes underground sulfide mineralization can be processed through Nevada Gold Mines Twin Creeks or the Lone Tree autoclave facilities. The 426 mineralized lenses are more amenable to alkaline conditions while the Ruby Deeps lenses perform better with acidic conditions.

 

22.1.1.5

Financials

 

   

Initial capital requirements total $49.4M with an additional $106.1M in sustaining capital.

 

   

The project achieves after-tax NPV 5% of $126.8M and NPV 8% of $91.1M.

 

   

The estimated payback period is 7.8 years with an IRR of 23%.

 

 

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22.1.2

Archimedes Open Pit

 

22.1.2.1

Mineral Resources

The Archimedes deposit was previously mined by Homestake and Barrick for West Archimedes and East Archimedes respectively. Mining ceased after a pit wall failure. An updated mineral resource estimate was completed, with the majority of mineral resources classified as indicated. There is currently potential for additional surface production of the deposit which would add to the value of the overall Ruby Hill project.

As the pit was never restarted after the wall failure, it will be important to understand and mitigate rock mechanics stability and safety issues prior to any decision to restart the project.

Given the current focus on the underground mine and the Mineral Point pit, no additional work in the Archimedes pit has been planned.

 

22.1.3

Mineral Point Open Pit

 

22.1.3.1

Mineral Resources

The Mineral Point Open Pit mineral resource contains approximately 47% inferred mineral resources. Drilling is planned for the deposit to obtain fresh material for additional metallurgical testing. The additional metallurgical test results can be used in future work, along with additional testing for representative bulk density measurements to be used with future updated geological, alteration, redox and structural models. This can be used for future mineral resource updates and potentially upgrading inferred mineral resources to indicated mineral resources.

 

22.1.3.2

Mining and Infrastructure

Mineral point will be a large-scale open pit gold and silver deposit typical of other northern Nevada mines with stripping ratio of 2.9:1, excluding capitalized pre-stripping. Overall average gold grade processed of 0.39 g/tonne with an expected average gold recovery of 78% and an average silver grade processed of 15.37 g/tonne. Most of the current infrastructure on site can be re-used or expanded for the project. Power for the proposed operation will be provided by the power supplier that historically fed the site.

 

22.1.3.3

Metallurgical Testing

Historical metallurgical testing and production have confirmed the amenability of Mineral Point open pit oxide and sulfide mineralization to conventional cyanide heap leaching; Metallurgical testing of samples from the Mineral Point open pit deposit has also shown amenability to crushing for heap leaching. Gold and silver recoveries ranged from 80-85% and 32-45% respectively.

 

22.1.3.4

Recovery Methods

Oxide and sulfide material is amenable for processing by crushed-ore cyanide heap leaching. Gold and silver leach at the heap-leach facility will be extracted by Merrill-Crowe zinc precipitation.

 

22.1.3.5

Financials

 

   

Total capital requirement of $1,383.2M

 

   

The project achieves an NPV 5% of $614.1M and NPV 10% of $134.8M After-Tax

 

   

The project has and IRR of 12.1% and a payback period of 7.9 years After-Tax

 

22.2

Risks and Opportunities

The Project is subject to the risks and uncertainties typical of gold projects, particularly risk in commodity prices and the precious metals equity markets. Lower metal prices or lack of precious metals equity market interest or activity could render the Project uneconomic or reduce access to project financing.

 

 

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The life of mine (LOM) plan includes a significant percentage of inferred mineral resources along with the indicated resources (there are no measured mineral resources). The current mineable resource demonstrates economic viability but will need to be upgraded to become a mineral reserve.

Metallurgical data appears to be of reasonable quality but does require additional test work. Incomplete classification of material types or misunderstanding of the representativeness of metallurgical samples could lead to a change in recovery or process cost assumptions. Further test work is needed to confirm crush sizes for optimal extraction and to refine cost parameters.

This is an initial assessment, which is based on engineering assumptions related to operating cost, capital cost, recovery, and other inputs. Further test work or analysis may modify these assumptions in ways which negatively impact the Project economics.

The Ruby Hill Project is located in a brownfields mining site with good electrical and transportation infrastructure in place. The local labor force is experienced in the type of mining planned, and contractors are available to perform the work. The permitting requirements for the underground mine are minimal and dewatering could provide a benefit to the agricultural users down gradient from the mine. Table 22-1 shows the risks and uncertainties for the Archimedes Underground (AUG), Mineral Point Open Pit (MPOP), and Archimedes Open Pit projects (AOP). Table 22-2 shows the opportunities for the Archimedes Underground (AUG), Mineral Point Open Pit (MPOP) and Archimedes Open Pit (AOP) projects.

 

 

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Table 22-1: Risks and Uncertainties

 

Project

   Risks    Impact    Mitigation Measure

AUG

   Dewatering Requirements Greater than Anticipated    Increased capital costs and operating costs    Hydrogeological study to determine dewatering requirements

AUG

   Delays in Permitting Approval    Production ramp up delays    Increased mine life and reduced life of mine economics

AUG

   Ground Conditions Worse than Expected    Increased operating costs    Install additional ground support, reduce mining widths, convert to underhand drift and fill mining

MPOP

   Proximity to Local Communities    Potential loss of social license    Maintain a pro-active and transparent strategy to identify all stakeholders and maintain a communication plan. The main stakeholders have been identified, and their needs/concerns understood. Continue to organize information sessions, publish information on the mining project, and meet with host communities.

MPOP

   Metallurgical Recovery    Lower recovery decrease in revenue    Additional test work is required to improve understanding of the recovery in different lithologies and target P80. Evaluate leach cycle, application rate, and lift height for final comminution circuit, including geotechnical considerations.

MPOP

   Permitting Challenges    Delay permitting and increase pre-production costs    Additional biological, geochemical, hydrogeological and archaeological baseline studies and follow-up are required.

MPOP

   Overliner Source for Heap Leach Facility has Not Been Explicitly Identified    Inhibit effective solution management, decrease in revenue    Identify and test overliner sources

MPOP

   Poor Foundation (geotechnical) Conditions    Increased capital costs    Complete geotechnical and hydrogeological studies and material testing programs for the heap leach facility and ancillary infrastructure to define foundation conditions and/or shallow ground water.

MPOP

   Power Availability    Increased capital and operating costs    Perform detailed power study and confirm with provider on capacity. Additional generators to provide power.

MPOP

   Water Supply    Constrained throughput, decreased revenue    Perform detailed water supply from ground and water demand study. Include climate analysis and inclusion of available make-up water sources.

MPOP

   Definition of Resource Model Alteration Types and Recoveries    Recoverable metal, decrease in revenue    Complete additional metallurgical test work to build geometallurgical model.

MPOP

   Bulk Density    Changes to Tonnage and Contained Metal Content, change in revenue    Review of key lithological units from existing drill core and/or potential relogging of core to achieve greater confidence in bulk density determination. Incorporate review work into updated geological model. Use commercial lab for umpire analysis of new samples, along with umpire check analysis of existing samples.

AOP

   Pit slope stability    Reduction in potentially minable resources    Additional Geotech mapping and drilling in the pit limit area.

 

 

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Table 22-2: Opportunities

 

Project

   Opportunities    Impact

AUG

   Process Ruby Hill Material at Lone Tree    Lower costs, higher return on investment.

AUG

   Mine Grades Exceed Plan    Increased gold production, lower cash cost.

AUG

   Underground Resource Conversion, Near Mine Exploration Success    Reserve growth, mine expansion, increased production rate, longer mine life.

AUG

   Dewatering    Infiltration of the water may alleviate historic overallocation down gradient from Ruby Hill and increase the water available for process or agricultural uses for MPOP and AOP.

MPOP

   Metallurgical Recovery    Additional metallurgical test work may improve understanding of operating parameters leading to more accurate revenue projections and a more effective production plan.

MPOP

   Geotechnical    Geotechnical drilling may improve understanding of operating parameters leading to more accurate mine designs and a more effective production plan.

MPOP

   Partial Contract Mining    Using a contractor to perform pre-stripping early in the project life may postpone capital spending.

MPOP

   In-Pit Dumping    Reduce haulage distance/time, improve productivity, decrease mining unit costs, and reduce operating costs.

MPOP

   Increase Ultimate Heap Height    Reduce disturbance and capital costs.

MPOP

   Self-Perform Manufacturing of Overliner    Determine if existing crusher at site could be leveraged and utilized in producing overliner, specifically for sustaining capital costs.

MPOP

   Self-Perform Clear and Grub    Evaluate mine fleet for capability to perform clear and grub for areas of future phases of HLP and WRSA, may reduce sustaining capital costs.

MPOP

   Event Ponds Containment    All event ponds in series include costs for secondary containment to utilize them as process ponds with no duration requirement to empty. Remove contingency design and empty ponds within set durations, may reduce capital costs.

MPOP

   Recovery from HL Relocated Processed Material from Historic Operations    Additional metallurgical test work may prove additional recovery from relocated material, improving revenue.

MPOP

   Screening in Comminution Circuit    May reduce capital costs with introduction of scalping screens to reduce volume sent to secondary crushers.

MPOP

   Resource Conversion and Growth    Conversion of inferred resources to indicated resources, and indicated resources to measured resources, leading to greater resource confidence and potential resource and/or reserve growth.

MPOP

   Improved Stormwater Management    Perform hydrology and hydraulics study to reroute existing drainage around proposed WRSA and HLF, may decrease costs.

AOP

   Waste rock storage    Should future expansion potential for the Archimedes as an open pit operation be eliminated, Archimedes could be utilized as storage for future overburden.

 

 

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23.

RECOMMENDATIONS

 

23.1

Archimedes Underground

 

23.1.1

Metallurgical Testing

 

   

Additional metallurgical testing is recommended from initial Ruby Hill production areas to confirm metallurgical recoveries with Twin Creeks process conditions. Sample selection should be based on available mine production plans and should reflect typical stope dimensions and expected dilution. Testing should include:

 

   

Comminution testing to confirm throughput through the Sage Mill.

 

   

Pressure oxidation tests using Twin Creeks conditions.

 

   

CIL tests on pressure oxidation productions.

 

   

Additional testing on Ruby Hill base metal sulfide zones to investigate flotation parameters to produce saleable lead and zinc concentrates. Detailed assays of lead and zinc concentrates are recommended to determine the extent of deleterious elements that may impair their salability.

 

23.1.2

Permitting and Mine Development

 

   

Complete the EA and POO amendment for Mining the 426 deposit above the 5100 elevation.

 

   

Initiate construction of the haulage portal and decline in Q3 2025.

 

23.1.3

Resource Conversion and Exploration Drilling

 

   

Begin Resource Conversion Drilling as soon as decline advance and drill platforms become available.

 

   

The lower leg of the decline provides a drill platform for exploration of the Blackjack deposit.

 

23.1.4

Dewatering

 

   

Initiate a hydrogeologic study of the Windfall formation, drill a deep test well and complete a drawdown test.

 

23.2

Archimedes Open Pit

 

23.2.1

Mineral Resources

Due to the short-term development plans for Mineral Point Open Pit and Archimedes Underground, additional work for the Archimedes Open Pit is not currently defined. Should resources be available a detailed geotechnical review of the existing pit slopes in Archimedes could help to quantify future potential. In light of current development plans on the property, this is not budgeted at this time.

 

23.3

Mineral Point Open Pit

 

23.3.1

Mineral Resources

It is recommended that i-80 complete additional resource definition drilling and conduct a review of major and minor rock alteration types, and how they align with overall geology, grade domains, metallurgical recovery and bulk densities. This would also include review of the geological model, including lithological, structural, and alteration controls on overall grade distribution and metallurgical recovery. The additional drilling could be used to better define the limits of mineralization and potentially upgrades block classification.

 

 

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The following points are recommended for additional evaluation:

 

   

Review of the overall (and subsequent low and high-grade) grade distributions to better understand impacts on mineralized domains.

 

   

Detailed review of deposit wide bulk densities to better define the bulk density for the project, including bulk densities of lithology and alteration type.

 

   

Additional drilling to increase the resource definition and confidence, along with potential upgrading of resource classification (inferred to indicated, indicated to measured).

 

   

Additional drilling for potential resource expansion.

Upon completion of the above items, an update to the geological model and mineral resource estimate should be conducted, along with updated metallurgical recovery assumptions.

 

23.3.2

Mining and Infrastructure

It is recommended that a site wide water balance be developed for the project to better understand water captured on-site (pit, HLP, WRSA) and evaluate the ability to utilize this water for process make-up water or to provide water for agriculture use. This would include evaluation of climate and available make-up water sources to understand total project requirements for make-up water or discharge as required. The evaluation would include a more accurate reflection of drain down for events, and potentially reduce the event pond volumes required, which could impact capital and sustaining capital costs.

There are several opportunities for infrastructure related components of the project to evaluate, including:

 

   

Conveyor stacking versus truck stacking, reduction of capital and operating costs.

 

   

Blasting versus crushing and screening, reduction of capital and operating costs.

 

   

Reduced number of event ponds and utilize larger event ponds to reduce capital costs.

 

   

Increased Heap Ultimate height of 300 feet, reduction of disturbance area as well as capital costs.

 

   

Utilization of existing crusher to self-perform overliner manufacturing to reduce capital costs

 

   

Evaluate all pits for potential for pit dewatering, including water quality evaluation, for ability to utilize this water as process make-up water or for agricultural use.

 

23.3.3

Metallurgical Testing

It is recommended that additional metallurgical testing be conducted to further define the predicted recovery for the Mineral Point Open pit project. This includes evaluation of sulfide sulfur content which will assist with determining the various oxidations by lithology as well as understanding recovery and reagent consumptions. This should also be conducted for waste as there may be a need to segregate waste into PAG and NAG facilities.

Next phases of the metallurgical testing program would incorporate additional leach tests, coarse bottle rolls, and column leach tests. This testing is required to support crush size selection, recovery estimates and reagent consumptions for lime and cyanide. Testing is also required to provide comminution design data. Testing and samples to be tested include:

 

   

Samples should focus on weakly-altered alteration of the major formations, the largest component of the Mineral Point resources. Sample selection should address spatial and grade variability within the deposit.

 

   

Identify samples in transition areas to sulfide mineralization to establish boundary criteria such as sulfide sulfur content.

 

   

Use of PQ diameter drilling will permit testing up to -2” crush size to evaluate the impact of crush size on recoveries.

 

 

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Evaluate the pilot leach testing of a bulk sample to determine ROM recoveries.

 

   

Testing of composite samples representing the first year and second year mine production once optimal conditions are selected.

 

   

Conduct column leach tests with taller columns and columns in series to replicate actual lift heights and heap leach operations.

 

   

Conduct laboratory tests to determine the crusher work index and abrasion indices to support crushing plant design.

 

   

Geotechnical testing, namely compacted permeability testing, of samples to determine the permeability and stacking characteristics of the mineralized material.

 

   

ABA testing of leach residue under conditions to support environmental permitting.

Additional considerations include metallurgical and geotechnical testing which will further the understanding of the ore’s clay content. This would include particle size distribution analysis, Atterberg limits, plasticity index, by ore type. This would also be coupled with compacted permeability testing to understand long term effects of loading and stacking. It is also recommended that ore decrepitation testing be conducted. Additional evaluation of the outcomes of this testing will verify the proposed application rate, leach cycle, and stack height for the various oxidations and lithologies based on permeability and agglomeration requirements.

It is also recommended that additional testing of proposed overliner material be conducted to evaluate screening requirements as well as stability for geotechnical design. This could also lead to a reduction in the overliner depth requirement, decreasing capital costs for the project.

Additional test work for recovery potential of the relocated HL material from historic operations should be conducted to potentially include revenue from this material.

The program has an estimated cost of $600,000 (excluding drilling costs) based on current conditions.

 

23.4

Work Program

 

23.4.1

Archimedes Underground

The work program outlined in Table 23-1 will advance the 426 deposit to production within two years. Project risks are manageable, and opportunities exist to enhance the project economics.

Table 23-1: Archimedes Underground Work Program

 

Description      2025        2026     

Estimated

Costs (US$M)

Portal Construction

   0.1    0.1    0.2

Mine Development

   7.8    21.0    28.8

Resource Conversion Drilling

   2.1    -    2.1

Dewatering Well and Hydrogeologic Study

   3.9    -    3.9

Environmental, Metallurgical Testing and Feasibility Study

   0.5    2.0    2.5

Ventilation and Electrical

   0.2    2.7    2.9

Project Administration

   5.0    0.6    5.6

Contingency

   2.9    4.5    7.4

Total

   22.5    30.8    53.3

 

 

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23.4.2

Archimedes Open Pit

Due to the short-term development plans for Mineral Point Open Pit and Archimedes Underground, additional work for the Archimedes Open Pit is not currently recommended.

 

23.4.3

Mineral Point Open Pit

The work program outlined in Table 23-2 will advance the Mineral Point Open Pit project to a Pre-Feasibility Study (PFS).

 

23.4.3.1

Phase 1

A two-phase work program is recommended. The focus of the Phase 1 work program will be additional drilling to obtain new sample material for metallurgical test work, hydro and geotechnical studies. This will include metallurgical test work of sufficient variability samples to support overall recovery assumption prior to moving to Phase 2. The additional drilling will also be used for subsequent resource definition, and potential resource classification upgrade and expansion. Based on the results of Phase 1, Phase 2 may be warranted. Additional metallurgical test work and other studies may be needed to further de-risk the Project.

 

23.4.3.2

Phase 2

The focus of the Phase 2 work program will be additional drilling for resource definition and expansion; and will include additional metallurgical test work to refine the process parameters. The Phase 2 drilling will be designed for resource conversion and growth, with the objective of converting inferred resources to indicated resources, as well as converting indicated resources to measured resources. The additional drilling and potential upgrade of inferred resources to indicated resource may lead to mineral reserves.

Table 23-2: Mineral Point Work Program

 

Description   

 Estimated Costs 

(US$M)

Phase 1      

Additional Drilling for Metallurgical, Hydro and Geotechnical Test Work

   $ 3.30

Metallurgical Test Work

   $ 0.25

Contingency

   $ 0.70
Phase 1 Total    $ 4.25
Phase 2      

Resource Definition & Expansion Drilling

   $ 15.0

Metallurgical Test Work

   $ 0.20

Contingency

   $ 1.00
Phase 2 Total    $ 16.20

 

 

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24.

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WSP, 2016. 2016 Ruby Hill Groundwater Characterization and Dewatering Update. Prepared for Elko Mining Group, September 9, 2016.

WSP USA Environment & Infrastructure Inc. (WSP), 2023. Ruby Hill Underground Portal Pit Slope Stability Evaluation, Eureka County, Nevada. Technical Memorandum prepared for I-80 Gold Corp. January 25, 2023.

WMC, 2004. East Archimedes Project, Assessment of the Hydrogeologic Conditions and Dewatering Feasibility: consultant’s report, 56 p. October 2004.

 

 

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25.

RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

i-80 has contracted some studies directly with specialist firms, experts in their discipline and provided the information to the QPs for this Technical Report Summary. The following information was provided to the Qualified Persons by i-80 Gold to use in the preparation of this report:

 

   

The technical status for the claims and land holding is reliant on information provided by The US Bureau of Land Management and the Eureka County Assessor’s Office. [Section 3.2]

 

   

Land Title opinions by Parr Brown Gee and Loveless and Erwin Thompson Faillers provided chain of title and land holding positions. [Section 3.2, 3.3]

 

   

Annual property holding costs were provided by i-80 [Section 3.2, 19.1.1]

 

   

The status of i-80’s environmental program and the permitting activities were provided by i-80. [Section 3.4, 3.5, 17]

 

   

Archimedes Underground hydrogeologic modeling, dewatering estimates, and Lone Tree autoclave operating costs were provided by i-80. [Section 7.3, 18.1.2]

 

   

Stantec provided hydrogeologic modeling and dewatering estimates for Mineral Point. [Section 15.2.4]

 

   

Hatch provided information on the refurbishment of the Lone Tree pressure oxidation (POX or autoclave) facility which is needed to recover metals from the sulfide ores in the Archimedes underground. [Section 14.2]

 

   

LRE Water performed hydrogeological modeling and analysis of inflows and water management for the Archimedes Underground. [Section 7.3,15.1]

 

   

Gold Pricing Forecast – CIBC Bank, was used in the metal price analysis section. [Section 16]

These contributions have been reviewed by the authors, and they believe them to be accurate portrayals of the Project at the time of writing this Technical Report Summary.

 

 

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APPENDIX A – SITE VISIT REPORT

 

 

 

 

 

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Site Visit Report Mineral Point Property, Nevada Project No. 195005 Prepared by:    Prepared for: FORTE DYNAMICS, INC     i-80 Gold Corp 120 Commerce Drive    5190 Neil Road Units 3 & 4    Suite 460 Fort Collins, CO 80524    Reno, NV 89502


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Version Control

 

Revision    Date    Status    Prepared By    Checked By    Approved By

REV A

   7-Feb-2025    Draft    J. Heiner    A. Amoroso    A. Amoroso

REV B

   11-Feb-2025    Final    A. Amoroso    A. Amoroso    A. Amoroso

REV C

   27-Mar-2025    Final v2    A. Amoroso    A. Amoroso    A. Amoroso


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Table of Contents

 

Executive Summary

     5  

1.  Introduction

     6  

2.  Areas of Review

     7  

2.1   Office Discussion

     7  

2.2   Previous Technical Reports

     7  

2.3   Project Location and Coordinate System

     7  

2.3.1   Project Location

     7  

2.3.2   Coordinate System

     8  

2.4   Drilling

     9  

2.4.1   Mineral Point Trend

     9  

2.4.2   Core Shack and Drill Core Review

     9  

2.4.3   Core Storage

     13  

2.4.4   Sampling Procedures

     14  

2.4.5   QA/QC Procedures and Protocols

     15  

2.4.6   Specific Gravity

     15  

2.4.7   Check Assays

     16  

2.4.8   Collar Check Field Inspection

     18  

2.5   Geology

     21  

2.6   Topography

     22  

2.7   Resource Block Model

     22  

2.7.1   Summary of Wood Block Models

     22  

2.7.2   Summary of Forte Block Model

     22  

2.7.3   i-80 Gold Drilling

     22  

2.8   Archimedes Pit

     23  

2.9   Infrastructure

     24  

2.9.1   Current

     24  

2.9.2   Proposed

     25  

2.10 Mineral Point Pit Area

     26  

2.11 Proposed Waste Rock Storage Area

     29  

2.12 Proposed Heap Leach Facility Area

     29  

3.  Conclusions and Recommendations

     31  

 

 

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Figures

 

Figure 1: Project Location Map      8  
Figure 2: Eureka District Stratigraphy      10  
Figure 3: Inside of Core Shack with Drill Core Boxes on Tables for Review      11  
Figure 4: Location of Reviewed Core Drill Holes (South Part of Deposit, Plan View)      12  
Figure 5: Drill Core Example from Drill Hole BRH-517C      13  
Figure 6: Core Storage On-Site      14  
Figure 7: Check Assay Sample from Hole BRH-517C      16  
Figure 8: Check Assay Results for Au      17  
Figure 9: Check Assay Results for Ag      18  
Figure 10: Field Inspection Searching for Collar Locations Covered in Snow      19  
Figure 11: Collar Location of Drill Hole with Brown Top Casing and Piezometer      20  
Figure 12: Aerial Photo Image Showing Garmin Handheld GPS Waypoints of Collar Check Locations      21  
Figure 13: Orthogonal Section of Existing Block Model (Forte-LS 2022) with 2023 i-80 Gold Drill Holes (section width +/- 25 ft)      23  
Figure 14: Archimedes Pit Looking Southeast Towards the Failure      24  
Figure 15: Current Infrastructure Locations      25  
Figure 16: Looking Northwest from the Warehouse Door Towards the Tire Pad      26  
Figure 17: Looking East from the Current Waste Rock Storage Area at Pit Area Phase 1 to 4      27  
Figure 18: Old Head Frame in Pit Area Phase 1 to 4      27  
Figure 19: Looking West at Exit Point of Pit Area for Phases 1 to 4 and Access to WRSA to the South (Left)      28  
Figure 20: North End of Pit Showing HL (Center) and WRSA (Right)      28  
Figure 21: pWRSA and Area that Needs to be Adjusted      29  
Figure 22: Heap Leach Adjustment      30  
Tables

 

Table 1: Mineral Point Check Assay Results      17  
Table 2: Field Collar Location Check      21  

 

 

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EXECUTIVE SUMMARY

Forte Dynamics, Inc (Forte) conducted a site visit at i-80 Gold Corp’s (i-80 Gold) Ruby Hill Mine with the focus on the Mineral Point property. The date of the site visit was January 16th, 2025, and served as the required site visit for QP sign-off on the project. Jon Heiner, Director of Mining, and Aaron Amoroso, Senior Resource Geologist, with Forte, performed the site visit. The site visit was conducted as part of the Preliminary Economic Assessment (PEA) on the Mineral Point property that i-80 Gold contracted Forte to complete along with Practical Mining. Forte will complete the surface open pit work and analysis on the Mineral Point project, and Practical Mining (Practical) will complete all underground and other open pit (not Mineral Point) work and analysis on the project. Practical will act as lead author for the NI 43-101 Technical Report (TR) and S-K 1300 Technical Report Summary (TRS).

 

 

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1.

INTRODUCTION

Forte Dynamics, Inc (Forte) conducted a site visit at i-80 Gold Corp’s (i-80 Gold) Ruby Hill Mine with the focus on the Mineral Point property. The date of the site visit was January 16th, 2025, and served as the required site visit for QP sign-off on the project. Jon Heiner, Director of Mining, and Aaron Amoroso, Senior Resource Geologist, with Forte, performed the site visit. The site visit was conducted as part of the Preliminary Economic Assessment (PEA) on the Mineral Point property that i-80 Gold contracted Forte to complete along with Practical Mining. Forte will complete the surface open pit work and analysis on the Mineral Point project, and Practical Mining (Practical) will complete all underground and other open pit (not Mineral Point) work and analysis on the project. Practical will act as lead author for the NI 43-101 Technical Report (TR) and S-K 1300 Technical Report Summary (TRS).

The site visit covered many topics including an office discussion of the property and project, project history, past and current infrastructure, drilling, review of the project geology and Leapfrog geological model, past and present topographic surfaces, the Wood 2021 Mineral Resource Estimate (MRE) and Forte 2024 Scoping Study (including block models), potential site locations for infrastructure and pit/s, future plans for the project, and timelines for the TR and TRS. The site visit also included a tour of the Archimedes Pits (east and west) from down inside the pits and at a lookout station from above, the service shop (truck shop) and connected warehouse, and the current core shack. No labs were visited during the site visit.

 

 

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2.

AREAS OF REVIEW

 

2.1

Office Discussion

The site visit started with an in-office meeting discussing the following topics:

 

   

Project history

 

   

Current project status

 

   

Infrastructure (past, present, planned), including road issue/s

 

   

Waste dump and heap leach pad

 

   

Topographic surfaces

 

   

Imagery

 

   

Wood 2021 MRE on Mineral Point and Forte 2024 Scoping Study

 

   

Drilling (i-80 Gold drilling and historic drilling)

 

   

Geology and review of Leapfrog geological model

 

   

Review of underground resources and mining

 

   

In conjunction with Practical

 

   

Project timelines and deliverables (NI 43-101 TR and S-K 1300 TRS)

 

2.2

Previous Technical Reports

Recent previous technical reports on the project prepared for i-80 Gold include:

 

   

Forte 2024 Scoping Study

 

   

Practical Mining 2023 updated PEA (unpublished)

 

   

Forte 2022 Scoping Study

 

   

Wood 2021 MRE

 

   

RPA 2012 Technical Report (prepared for Barrick Gold)

 

2.3

Project Location and Coordinate System

 

2.3.1

Project Location

The Ruby Hill (Mineral Point) project is located on the Battle Mountain/Eureka gold trend approximately 2 km northwest of the small town of Eureka in Eureka County, Nevada, USA, approximately 145 km south of Elko and approximately 325 km east of Reno.1

 

 

1 NI 43-101 Report on 2021 Ruby Hill Mineral Resource Estimate, Wood, Section 4 pp. 4-1.

 

 

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Figure 1: Project Location Map

(Source: 2021 Ruby Hill Mineral Resource Estimate, Wood)

 

2.3.2

Coordinate System

The project has used a local grid system referred to as the Ruby Hill Mine Grid, which uses the Locan Shaft as its origin. The Ruby Hill Mine Grid is in feet (ft).

The Locan Shaft origin point of 0,0 ft was modified to 10000,110000 to avoid any negative numbers.

The project centroid location (derived from the geological model) is 9495, 115158 in the Ruby Hill Mine Grid, and 1925147, 14352286 in UTM NAD83 Z11N.

Ruby Hill Mining Company made an update to the Ruby Hill Mine Grid in 2017, applying NAD83_2011 Geodetic Datum (Lat/Long).2

 

 

2 NI 43-101 Report on 2021 Ruby Hill Mineral Resource Estimate, Wood, Section 9.1 pp. 9-1.

 

 

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2.4

Drilling

Since acquiring the project, i-80 Gold has completed drilling of 184 drill holes from 2021-2024. None of the drill holes from 2021-2024 drilling campaigns specifically targeting the Mineral Point deposit; however, approximately seven (7) drill holes intersected the Mineral Point deposit, as part of a drilling program designed to target the CRD style mineralization below and/or adjacent to the Mineral Point deposit.

 

2.4.1

Mineral Point Trend

The Mineral Point deposit is the largest precious metal Mineral Resource in the i-80 portfolio. It consists of gold and silver mineralization hosted by the Cambrian Hamburg dolomite in the nose of a broad anticline that plunges gently to the north-northwest and is bound to the east by the Holly Fault and to the west by the Spring Valley Fault. The Mineral Point deposit is 10,000 ft long, 2,400 ft wide and up to 500 ft thick. The top of the Mineral Point deposit is near surface at its south end and 500 ft below surface at its north end. Majority of the mineralization in the Mineral Point Trend deposit is oxidized and has a high ratio of cyanide soluble to fire assay total gold. This deposit has not been mined and is the largest precious metal Mineral Resource in the Ruby Hill Project.3

 

2.4.2

Core Shack and Drill Core Review

After the joint tour of the pits and surface infrastructure, Tyler Hill (i-80 Gold) accompanied Aaron Amoroso (Forte) around the project site with a focus on the geology and drilling. The first stop on the tour was a visit to the core shack to review relevant drill core to the Mineral Point project resource. The core shack was very nice looking from the outside and inside, clean, very well lit from artificial light, and also heated for comfortable use during winter in Nevada. There were two (2) large tables capable of holding many core boxes for logging with sufficient indoor artificial lighting. Per Aaron’s request, Tyler pulled out portions of three (3) drill holes from i-80 Gold drilling between 2021-2024 that was relevant to the existing Mineral Point resource and located in the south portion of the deposit. The drill holes reviewed were BRH-166C, BRH-184C and BRH-517C (holes drilled by Barrick). Only portions of the drill core near the mineralized zones were pulled out for review, as the mineralization of the available drill holes tends to start at depth (~500ft at depth).

Aaron reviewed the core library and stratigraphic column, focusing on the geology of the mineralized zone for Mineral Point, which included the hanging wall of Dunderberg Shale, then entering the Hamburg Dolomite which hosts the majority of the mineralization with a semi-hard boundary to the footwall Secret Canyon Formation. Tyler stated that the higher grades in the dolomite came via a Cretaceous Intrusion (KI) unit. Aaron reviewed a number of occurrences noting the logged alteration, decarbonation, silicification and brecciation, which were consistent with the Wood technical report and core photos. Some zones of mineralization were heavily fractured and/or altered, and extremely fine grained, resembling a fine-grained beach sand.

 

 

3 NI 43-101 Report on 2021 Ruby Hill Mineral Resource Estimate, Wood, Section 1.5.1 pp. 1-4.

 

 

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Figure 2: Eureka District Stratigraphy

(Source: Photo taken by Forte staff from inside the core shack)

Aaron specifically reviewed the higher-grade intervals available in the pulled core from the three (3) drill holes. It should be noted that a number of the higher/highest grade samples had been fully removed for previous metallurgical test work and thus not available for review or check assay analysis. Some higher-grade remaining intervals had elevated amounts of galena along with elevated levels of oxidation and darker coloring, and higher traces of visible sulfides. Tyler made the comment that the style of mineralization resembles carbonate replacement deposit (CRD) mineralization. Aaron then also reviewed the footwall Secret Canyon Formation (shale), which carries almost no elevated Au grades.

Aaron also completed a general review of the different lithological units, dominant minerals, alteration types, contacts (lithology and mineralized), common structures observed in the core, mineralized zones defined from assays, and mineralization styles.

 

 

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Figure 3: Inside of Core Shack with Drill Core Boxes on Tables for Review

(Source: Photo taken by Forte staff from inside the core shack)

 

 

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Figure 4: Location of Reviewed Core Drill Holes (South Part of Deposit, Plan View)

(Source: Forte Dynamics, Inc. 2024)

 

 

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Figure 5: Drill Core Example from Drill Hole BRH-517C

(Source: Photo taken by Forte staff from inside the core shack)

 

2.4.3

Core Storage

The core is stored outside on the north waste dump, on pallets, under a weather resistant tarp which is strapped down.

The coarse rejects are stored in multiple locations. Historic rejects are stored with core on-site at the Project in barrels. More recent drilling rejects are stored in barrels and/or on pallets under a tarp at the Lone Tree project (Nevada).

Pulps are stored at an indoor warehouse at the Lone Tree project.

 

 

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Figure 6: Core Storage On-Site

(Source: Forte Dynamics, Inc. 2024)

 

2.4.4

Sampling Procedures

 

2.4.4.1

RC Drilling

Sample intervals from the RC drill holes were collected in five (5) feet (ft) intervals, where samples were inserted into sequentially numbered sample bags by the drilling crew with the outside of the bags marked with the drill hole and sample number. Samples were allowed to drain/dry at the sample site and were then reviewed by the geologist in charge of the program to ensure accurate numbering/sequencing of the samples. Once drained and/or dried, the samples were re-located from the drill site to the shipment staging area, where personnel relabeled the bags containing the duplicate samples by assigning the correct sequential number. The samples were then loaded into 4 x 4 x 3-foot wooden or plastic crates in preparation for pickup by the assay lab.

 

2.4.4.2

Diamond Drilling

Sample intervals are chosen by the geologist based on detailed geologic observations. Sample intervals may range from ten feet to a minimum of one foot, with a maximum of five feet in areas of interest. The geologist marks sample intervals on the core and staples a sample ticket double-stub in the core box at the end of the sample interval. Sample IDs are automatically generated in AcQuire, with a prefix that designates the project. Sample tickets are then printed out with sample IDs. Logged core boxes are photographed with

 

 

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a high-resolution camera while wet and then stacked on a wooden pallet prior to being transported to the Lone Tree mine site for cutting of the core and shipping to an assay lab.

The geologist prints a cut-sheet from AcQuire software with the sample numbers and intervals and provides a cut-sheet to the geotechnician. The geotechnician puts one sample bag in a five-gallon plastic bucket on the floor next to the core saw. The core is sawed in half, the left piece is placed into the sample bag, and the right piece goes back into the core box. In the case of broken core, the sampler does their best to divide the sample equally. Once the interval is split, the geotechnician takes one part of the double sample stub from the core box and staples it to the sample bag. The remaining sample stub remains in the core box for future reference. The geotechnician then ties the sample bag shut and marks the sample off the cut-sheet. The tied sample bags are stored in a sample bin for the lab driver to pick up.

Drill hole status, such as splitting, sample dispatch date, batch ID, and dates of both preliminary and final results, are tracked in AcQuire as well as on ALS Mineral’s online portal.

Samples were submitted to three (3) different labs – ALS Minerals (ALS), American Assay Labs (AAL) and Paragon Geochemical Assay Laboratories (PAL) – all located in Sparks, NV.

 

2.4.5

QA/QC Procedures and Protocols

 

2.4.5.1

RC Drilling

Blanks and standards were inserted into the sample stream for every tenth sample. Duplicate samples were collected every 100 ft. i-80 Gold targets a ~20% QC sample insertion rate for their drilling and sampling programs.

Note, there was no active drilling at the time of the site visit, with the last drilling completed by i-80 Gold from early 2024. During the office discussion, Aaron discussed the QA/QC procedures and protocols used by i-80 Gold for their drilling since owning the project. Their QAQC program includes standards, duplicates and blanks, trying to achieve a ~20% QC insertion rate. Based on the discussion, it sounds like i-80 Gold is employing a robust QAQC program that follows industry best practices.

 

2.4.5.2

Diamond Drilling

Similar QA/QC procedures and protocols used for the RC drilling were used for the diamond drilling.

The geologist assigns QAQC samples while logging targeting 5% blanks, 5% standards, and 2.5% field duplicates. The geologist attempts to place blanks after high-grade samples where available. The geologist also attempts to place standards proximal to mineralized zones with standard gold values approximately that of the mineralized zone gold values. However, since the gold value of the rock cannot be known prior to assay, the standard value may not always compare well to the mineralized zone. The geotechnician places the blanks and duplicates with their sample tags in the sample bin with the regular core samples. The standards are placed in a small sample bag with the corresponding sample ID. The standards corresponding to a single hole are then placed in a larger bag prior to shipment to the assay lab.

The geologist completes a sample submittal sheet and randomly designates 2.5% of samples to have a prep duplicate prepared by the assay lab as an additional QAQC measure. The assay lab driver picks up the samples from the Lone Tree core shed and is given a chain of custody form with sample ID’s for the shipment. An electronic copy of the sample submittal form is emailed to the assay lab.

 

2.4.6

Specific Gravity

SG measurements were taken by i-80 Gold staff internally. No samples for the Mineral Point deposit have been sent to a commercial lab for analysis and verification of internal measurements.

 

 

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2.4.7

Check Assays

As part of the data verification process, Aaron collected five (5) samples from three (3) different drill holes to submit to a commercial lab for check assay analysis. As noted above, almost all of the higher-grade samples had been removed for previous met testing analysis. Thus, Aaron had to take the next best samples that had somewhat elevated grades. Table 1 shows the check assays samples selected for umpire lab analysis. The sample were collected by Aaron and submitted to ALS in Sparks, NV by Tyler. The assay results certificate was requested to be sent directly to Aaron from ALS to ensure chain of custody was followed the best that it could be given the circumstances.

 

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Figure 7: Check Assay Sample from Hole BRH-517C

(Source: Photo taken by Forte staff from inside the core shack)

 

 

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Table 1: Mineral Point Check Assay Results

 

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Lab Sample # Field Sample # Drill Hole Interval (ft) Check Assay Au g/t Database Assay Au g/t Check Assay Agg/t Database Assay Agg/t Difference Au Difference Ag RE25015355M001 MP001-CAFD BRH-166C 635-640 RE25015355M002 MP002-CAFD BRH-184C 840-845 RE25015355M003 MP003-CAFD BRH-184C 930-935 RE25015355M004 MP004-CAFD BRH-517C 685-690 RE25015355M005 MP005-CAFD BRH-517C 990-995 Mineral Point Check Assays 0.095 0.146 2.3 3.0 0.65 0.76 0.043 0.666 1.8 7.0 0.06 0.2 0.234 1.250 1.9 41.0 0.19 0.05 14.350 12.400 235.0 111.0 1.16 2.12 1.080 1.135 135.0 169.0 0.95 0.80

 

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Figure 8: Check Assay Results for Au

(Source: Forte Dynamics, Inc. 2024)

 

 

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Figure 9: Check Assay Results for Ag

(Source: Forte Dynamics, Inc. 2024)

 

2.4.8

Collar Check Field Inspection

Aaron did a modified field inspection reviewing existing collar locations near the main infrastructure. This was done is part due to time constraints as well as the snow that fell prior to the site visit, which made finding some existing collar locations difficult. Not many existing collar locations had casing sticking out from the ground. Tyler asked another i-80 geologist to go to the field earlier in the day to find/flag some collar locations for the field inspection. Tyler and Aaron were able to find a few collars to take a waypoint using a handheld Garmin GPS.

 

 

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Figure 10: Field Inspection Searching for Collar Locations Covered in Snow

(Source: Photo taken by Forte staff on the south side of the Mineral Point deposit)

 

 

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Figure 11: Collar Location of Drill Hole with Brown Top Casing and Piezometer

(Source: Photo taken by Forte staff on the south side of the Mineral Point deposit)

Of the five (5) collars check with the handheld GPS, one was a water well and was not in the drilling database (PW5, which is an active dewatering well). The other four (4) collars checked for X and Y location were all within the acceptable difference range when using a handheld GPS. Note that drill hole BRH-317C had a bigger difference in X and Y to the database compared to other checked collars, but due to snow cover, we were unable to find the exact collar and had to estimate its location. It should also be noted that the Z elevation in the database was slightly different than the elevations taken from the current topo surface. These differences are most likely due to different survey methods over the life of the database/s from different owners and drilling campaigns, and/or perhaps elevations taken from a different topo surface than was used for the collar verification. Regardless, the difference in elevation was minimal and within tolerance.

 

 

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Figure 12: Aerial Photo Image Showing Garmin Handheld GPS Waypoints of Collar Check Locations

(Source: Forte Dynamics, Inc. 2024)

Table 2: Field Collar Location Check

 

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Field Inspection Collar Check Diff Subtracted By (ft) # Drill Hole ID 1 BRH-317C* 2 IRH23-28 3 BRH-436 Check X Check Y Topo Elev Database X Database Y Database Elev 8,445.00 116,374.00 6,502.98 9,275.00 116,351.00 6,560.91 8,099.00 117,800.00 6,407.50 Diff X Diff Y Diff Z 8,449.54 116,364.24 6,502.38 4.54 (9.76) (0.60) 9,277.66 116,347.45 6,559.12 2.66 (3.55) (1.79) 5 BRH-435 8.143.00 118.028.00 6.410.00 8,105.20 117,799.31 8.146.05 118.028.09 6,409.74 6.20 (0.69) 2.23 6,409.02 3.05 0.09 (0.98)

 

2.5

Geology

Aaron discussed the project and deposit geology with Tyler in detail in the office and out in the field, including the core shack while reviewing the three (3) core holes pulled for review. Aaron also spent time going through and reviewing the geological model, created by i-80 Gold using Leapfrog software. The geological model appears to be reasonable and makes sense geologically when compared to the drill core, mineralization style and observed macro and micro geological and structural data available. Tyler sent Aaron a copy of the current Leapfrog project which includes the geological model. The main geological units (Hamburg Dolomite) will be reviewed along the modeled mineralized domain zones.

 

 

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2.6

Topography

The current topo surface is believed to be from 2021, which topo Jon used for his work in the 2022 and 2024 Forte Scoping Studies.

i-80 Gold completed a LiDAR survey in 2023 and generated a topo surface. 

 

2.7

Resource Block Model

 

2.7.1

Summary of Wood Block Models

Wood (Wood) completed an MRE in 2021 for the Ruby Hill Complex which included the Mineral Point project resource. The MRE used a probability assigned constrained kriging (PACK) methodology for the resource estimate using thresholds of 1.0 g/t Au and 40 g/t Ag to define low-grade and high-grade domains (composites), and an estimated indicator probability of 0.37 to define blocks for the high-grade domain (ore-waste definition).

The Two (2) Wood block models were created in Vulcan and named Vulcan block models 252525_global_simplified.bmf and 252525_expanded_11dec.bmf.

 

2.7.2

Summary of Forte Block Model

Forte then completed a scoping study on Mineral Point in 2022 where the Wood 2021 resource was reviewed/audited and slighted modified by Larry Snider. The modifications included combining certain data/fields from the two (2) above Wood block models (Vulcan block models 252525_global_simplified.bmf and 252525_expanded_11dec.bmf) into a single updated block model (expanded).

Updates were made to selected fields for oxide-transition-sulfide definition, re-flagging (block coding) of the lithology model using a version of the Wood 3-D wireframe lithological model as well as assigning density values based on the lithology, updates to tonnage factor values (density values in imperial units appeared to be inconsistent along with the conversion from imperial tonnage factors), as well as updating the block model for the updated 2021 topo surface and waste dumps.4 The resulting updated Vulcan block model was named 252525_expanded_11dec_forte10-17-22.bmf.

 

2.7.3

i-80 Gold Drilling

Although no i-80 Gold drill holes from the 2021-2024 drilling campaigns specifically targeted the Mineral Point resource, seven (7) drill holes intersected the resource. Figure 13 shows an orthogonal section of the current block model (50 ft section window) from the south part of the deposit with two (2) 2023 i-80 Gold drill holes and associated mineralized zones. In this section, one (1) of the mineralized zones appears to reasonably follow the current estimated block grades, where the other two (2) mineralized zones would potentially contribute a positive impact to an updated mineral resource estimate.

 

 

4 Technical Memorandum on Open Pit Mineral Reserves Parameters, Forte Dynamics, Dec. 2022, pp. 2-3.

 

 

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Figure 13: Orthogonal Section of Existing Block Model (Forte-LS 2022) with 2023 i-80 Gold Drill Holes (section width +/- 25 ft)

(Source: Forte Dynamics, Inc. 2024)

 

2.8

Archimedes Pit

The Archimedes pit has a large alluvium failure in the southeast corner of the pit, as seen in Figure . The pit is blocked off after the switchbacks on the northwest corner of the pit as the ramp leads into the main pit area. At this location, i-80 Gold is preparing two portals to start the Ruby Deeps underground project. They are currently scaling the wall and hoping to start development in late Q1 or early Q2 2025. Practical Mining is currently working on the two surface deposits (West and East Archimedes) in the Archimedes pit for the upcoming PEA.

 

 

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Figure 14: Archimedes Pit Looking Southeast Towards the Failure

(Source: Forte Dynamics photo looking southeast over Archimedes pit. Underground portals are directly below where the photo was taken)

 

2.9

Infrastructure

 

2.9.1

Current

After viewing the Archimedes pits the first stop to review on-site infrastructure was the current truck shop and warehouse. The shop was built in the late 1990s for 777 haul trucks. It is a three-bay truck shop with a wash pad on the southeast end. A small warehouse is attached to the back of the truck shop, along with some office space. There is also another storage building to the northwest of the truck shop that does not have heating for overflow items. The main fuel island is located to the southeast of the truck shop. This fuel island is set up for 785 haul trucks. There are two administration buildings located next to the old mill. There are currently three crushers on site. The primary crusher is a jaw crusher, the secondary is a cone crusher (the cone has been removed), and a small tertiary crusher. Refer to the DRA report (DRA—H6975-0000-PM-REP-001—Ruby Hill Scoping Study Report_10242022_SA.pdf). All these items are shown in Figure . The main power comes from the substation to the northeast along Highway 50. We were unable to see any nameplates on the transformers. This power supply will be used for the planned underground. The current underground infrastructure will not be inside the Mineral Point pit area and should not have an impact.

 

 

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Figure 15: Current Infrastructure Locations

(Source: Forte Dynamics, Inc. 2024)

 

2.9.2

Proposed

After discussion with personnel on-site, we came to these recommendations:

 

   

Expand the current truck shop to the Northwest, adding four (4) bays to handle 320-ton trucks. Figure 16 shows the area for potential shop expansion.

 

   

Use the current three (3) bays for support equipment.

 

   

Extend the warehouse behind the four (4) new bays for extra storage.

 

   

If needed, expand the storage building.

 

   

If possible, turn the storage building into a light vehicle truck shop if not needed for storage.

 

   

Leave tire pad in current location if possible.

 

   

If need to move, place over in the current Primary Crusher dump pocket area where there is space.

 

   

May need to add or upgrade the main fuel island for the larger fleet.

 

   

Add a secondary fuel island on the west side of the pit closer to the crusher or waste dump location to limit out-of-cycle travel for fuel when the pit is running.

 

   

The current crushers (primary, secondary, and tertiary) are too small for the mineral point project and cannot be reused. A budget review is required for the cost of all new crushers. The old heap leach was only a primary and secondary crush. The tertiary was for the mill.

 

   

The proposed new crusher location to the west of the proposed Mineral Point pit on the current waste dump is good. Access will need to be developed to reach this area.

 

   

We will need to add admin and/or line-out space.

 

   

There is room up and around the truck shop, or can you look around the current admin area?

 

 

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It is possible that the current core shack will be taken back by Southwest Energy (explosives company), and it so, it will need to find a new location.

 

   

There is no plan to use the current mill.

 

 

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Figure 16: Looking Northwest from the Warehouse Door Towards the Tire Pad

(Source: Forte Dynamics, Inc. 2024 photo looking northwest from ware house door behind current truck shop)

 

2.10

Mineral Point Pit Area

The currently proposed Mineral Point pit area can be divided into three (3) main areas. The first is on the south end of the pit, where phases 1 to 4 are mined. The area is split into two (2) sub-areas. The first is the native ground, which is covered in shrubs, trees, drill pads, and some old mining infrastructure. The area has some elevation changes as it is on the foothills of Ruby Hill. This area will also have to deal with some current Waste Rock Storage Area (cWRSA) removal. Figure 17 and Figure 18 show the area and old mining infrastructure in Area 1. Figure shows the exit point for phases 1 to 4 next to the current WRSA.

Area two is at the far north end of the pit and is on flat land of the valley floor. There is minimal elevation change in the area. This area will have to deal with removing the current heap leach pad. It is assumed that the bulk of the removal would be done with mining equipment, and a third-party contractor would do the last 20-ish feet and the liner removal. Sampling of the soil below the pad would have to be done to confirm that no leeks/spills needed to be cleaned up. Figure 20 shows the north end of the pit.

 

 

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Area three is the pit’s center section, mostly removing cWRSA. There is little native ground in this area, and it looks to already be disturbed by other mining.

 

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Figure 17: Looking East from the Current Waste Rock Storage Area at Pit Area Phase 1 to 4

(Source: Forte Dynamics, photo looking east over first four phase of mining)

 

 

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Figure 18: Old Head Frame in Pit Area Phase 1 to 4

(Source: Forte Dynamics, photo looking west towards current waste dumps and old head frame in phase 1 of mining)

 

 

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Figure 19: Looking West at Exit Point of Pit Area for Phases 1 to 4 and Access to WRSA to the South (Left)

(Source: Forte Dynamics, Inc. 2024: photo looking east from first phases of mining)

 

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Figure 20: North End of Pit Showing HL (Center) and WRSA (Right)

(Source: Forte Dynamics, Inc. 2024: photo looking east over new Heap Leach location)

 

 

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2.11

Proposed Waste Rock Storage Area

The proposed Waste Rock Storage Area (pWRSA) is to the west and south of the proposed pit. When on the tour, Carol Olsen, who lives in Eureka, NV, informed us that the pWRSA covers some county roads. One comes off Rubby Hill into the valley, and is extremely close to impacting a paved road leading to radio towers. We will need to look at moving the dump to the west and south into the valley more to limit the impact on these roads as seen in Figure 21. They would like to keep the dump on the west side of the ridge to limit visibility from town. Some cultural sites in this area will have to be remediated. There is also a county road at the bottom of the valley that will need to be relocated to the west along the foothills of the valley. Overall, the group thinks this is a good location for the pWRSA.

 

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Figure 21: pWRSA and Area that Needs to be Adjusted

(Source: Forte Dynamics, Inc. 2024)

 

2.12

Proposed Heap Leach Facility Area

The proposed heap leach is on the flat of the valley. The current ponds are pushed against the current property bound, which places them very close to some homes and farms. The conversation in the field is to expand the pad to the west allowing the pad to be pulled to the south away from the homes and farms as shown in Figure 22. The heap leach also covers the same county road in the bottom of the valley that the pWRSA covers. A new road to the west around the HLF will have to be established for public access. i-80 Gold is working on getting an expanded topo out to the property boundary for this work.

 

 

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Figure 22: Heap Leach Adjustment

(Source: Forte Dynamics, Inc. 2024)

 

 

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3.

CONCLUSIONS AND RECOMMENDATIONS

The site visit was very useful to review the project with i-80 staff and discuss key components to the upcoming PEA. There were no major issues identified during the site visit and at the time of writing this report.

Selected recommendations from the site visit under the scope of the upcoming PEA include:

 

   

Review of SG measurements throughout the deposit, including from historical drill holes to better estimate SG values for use in the project and the resource estimation models.

 

   

Submit samples to a commercial lab for SG analysis and verification of existing internal measurements

 

   

Consider the HW and FW lithological units as well for the analysis.

 

   

Review of magnetic declination and potential adjustment to existing drill holes

 

   

Past drilling and certainly future drilling.

 

   

Additional metallurgical test work

 

   

Note this is planned for later in 2025 under a planned drilling program to get additional fresh material for met testing.

 

   

Review the potential impact of the seven (7) drill holes that intersected the Mineral Point resource from the 2021-2024 i-80 Gold drilling campaigns.

 

   

Consider updating the resource using i-80 Gold’s drilling results from 2021-2024 and the planned 2025 Mineral Point drilling campaign.

 

   

Review alternative resource domain and estimation techniques to focus on a more geologically constrained resource rather than a more statistical PACK estimation workflow.

 

   

Continue to refine the Leapfrog geological model

 

   

Use updated model for future resource updates along with updated SG determinations.

 

 

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www.fortedynamics.com

120 Commerce Drive, Unit 3-4, Fort Collins, CO 80524

Phone: +1 (720) 642-9359   info@fortedynamics.com

 

 

 

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APPENDIX B – MINERAL POINT OPEN PIT ECONOMIC MODEL WITH INFERRED RESOURCES

 

 

 

 

 

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FORTE DYNAMICS, INC
Company    i-80 Gold Corp.
Project Name    Mineral Point
Project Number    195005
Date    3/31/2025
Economic Model    DS_1v2 Sch_v1 With Inferred

 

Inputs

 

Mining and Processing Cost      Metal Price, Deduct, & Recovery  

                                                                   

Mining Ore

     US$/ton          $  2.50        Gold Price        US$/toz        $  2,175.00        Payable Au      %      100%  

Mining Waste

     US$/ton        $  2.50        Silver Price        US$/toz        $ 27.25        Payable Ag      %      100%  

Mining HeapLeach

     US$/ton        $  1.50        Royalties        %        3%        Gold Recovery      %      78%  

Processing Cost

     US$/ton        $  3.90        Refining Cost Au         US$/toz          1.85        Silver Recovery        %          41%  

G&A Cost

     US$/ton        $  0.75        Refining Cost Ag        US$/toz        0.5                         

 

Period     (1)      1      2      3      4      5      6      7      8      9      10      11      12      13      14      15      16      17      Total  
Year     2029      2030      2031      2032      2033      2034      2035      2036      2037      2038      2039      2040      2041      2042      2043      2044      2045      2046  

Production

   Processing Material    tons      -        8,131,522        24,999,945        25,068,438        24,322,575        24,999,945        24,999,945        25,068,438        24,999,945        24,999,945        24,999,945        25,068,438        24,999,945        24,999,945        24,999,945        25,068,438        24,999,945        12,716,826        395,444,125  
   OPEX Waste    tons      -        6,882,728        106,842,126        108,535,995        106,551,803        110,495,080        91,216,491        76,393,718        78,883,005        80,373,042        88,957,968        88,600,089        78,692,481        80,419,828        9,511,676        6,860,676        17,496,102        1,731,192        1,138,443,998  
   CAPEX Waste    tons      -        114,900,417        -        -        -        -        -        -        -        -        -        -        -        -        -        -        -        -        114,900,417  
   Heap Leach Relocation    tons      -        -        -        -        -        -        -        9,111,725        -        -        -        -        -        -        17,343,218        -        -        -        26,454,942  
   Total Material    tons      -        129,914,667        131,842,071        133,604,434        130,874,378        135,495,025        116,216,435        110,573,881        103,882,950        105,372,987        113,957,912        113,668,527        103,692,426        105,419,773        51,854,839        31,929,114        42,496,047        14,448,018        1,675,243,483  
   Processing Material    Tonne      -        7,376,794        22,679,575        22,741,711        22,065,075        22,679,575        22,679,575        22,741,711        22,679,575        22,679,575        22,679,575        22,741,711        22,679,575        22,679,575        22,679,575        22,741,711        22,679,575        11,536,514        358,740,979  
   OPEX Waste    Tonne      -        6,243,908        96,925,574        98,462,227        96,662,197        100,239,479        82,750,232        69,303,235        71,561,479        72,913,218        80,701,334        80,376,672        71,388,638        72,955,662        8,628,850        6,223,902        15,872,202        1,570,511        1,032,779,319  
   CAPEX Waste    Tonne      -        104,235,935        -        -        -        -        -        -        -        -        -        -        -        -        -        -        -        -        104,235,935  
   Heap Leach Relocation    Tonne      -        -        -        -        -        -        -        8,266,020        -        -        -        -        -        -        15,733,507        -        -        -        23,999,527  
   Total Material    Tonne      -        117,856,637        119,605,149        121,203,938        118,727,272        122,919,054        105,429,807        100,310,966        94,241,054        95,592,793        103,380,909        103,118,383        94,068,214        95,635,236        47,041,932        28,965,613        38,551,777        13,107,025        1,519,755,759  

Metal

   Contained Au    toz      -        60,698        253,260        277,039        433,268        246,246        249,812        294,535        300,650        280,749        348,919        364,353        269,114        344,656        301,153        175,310        206,036        118,744        4,524,542  
   Au Grade    toz/ton               0.0075        0.0101        0.0111        0.0178        0.0098        0.0100        0.0117        0.0120        0.0112        0.0140        0.0145        0.0108        0.0138        0.0120        0.0070        0.0082        0.0093        0.0114  
   Recovered Au    toz      -        50,485        210,936        223,656        348,326        200,058        204,021        243,386        243,994        215,920        275,100        274,809        209,808        207,108        210,398        143,929        168,933        98,526        3,529,392  
   Sellable Au    toz      -        50,435        210,725        223,432        347,977        199,858        203,817        243,143        243,750        215,704        274,825        274,535        209,598        206,901        210,188        143,785        168,764        98,427        3,525,863  
   Contained Ag    toz      -        3,795,038        10,250,946        10,800,857        18,164,122        7,462,025        6,767,883        7,913,326        8,411,357        6,248,963        7,947,194        13,867,275        26,333,087        16,255,199        13,120,648        9,523,562        6,705,453        3,726,136        177,293,070  
   Ag Grade    toz/ton               0.4667        0.4100        0.4309        0.7468        0.2985        0.2707        0.3157        0.3365        0.2500        0.3179        0.5532        1.0533        0.6502        0.5248        0.3799        0.2682        0.2930        0.4483  
   Recovered Ag    toz      -        1,600,880        4,249,672        4,455,719        7,318,117        3,125,044        2,783,374        3,284,291        3,471,255        2,577,538        3,272,777        5,578,944        10,533,235        6,506,857        5,248,259        3,809,425        2,717,429        1,495,470        72,028,286  
   Sellable Ag    toz      -        1,592,876        4,228,424        4,433,440        7,281,527        3,109,419        2,769,457        3,267,870        3,453,899        2,564,651        3,256,413        5,551,050        10,480,568        6,474,322        5,222,018        3,790,378        2,703,842        1,487,992        71,668,145  
   Contained Au    grams      -        1,887,926        7,877,263        8,616,870        13,476,165        7,659,116        7,770,029        9,161,061        9,351,267        8,732,273        10,852,589        11,332,640        8,370,400        10,720,021        9,366,902        5,452,754        6,408,455        3,693,367        140,729,097  
   Au Grade    g/tonne               0.2559        0.3473        0.3789        0.6107        0.3377        0.3426        0.4028        0.4123        0.3850        0.4785        0.4983        0.3691        0.4727        0.4130        0.2398        0.2826        0.3201        0.3923  
   Recovered Au    grams      -        1,570,274        6,560,834        6,956,489        10,834,150        6,222,497        6,345,765        7,570,157        7,589,067        6,715,863        8,556,563        8,547,532        6,525,749        6,441,782        6,544,127        4,476,700        5,254,394        3,064,503        109,776,446  
   Sellable Au    grams      -        1,568,704        6,554,273        6,949,533        10,823,316        6,216,275        6,339,419        7,562,587        7,581,478        6,709,147        8,548,006        8,538,984        6,519,223        6,435,341        6,537,583        4,472,223        5,249,140        3,061,439        109,666,670  
   Contained Ag    grams      -        118,038,958        318,840,291        335,944,470        564,967,765        232,095,081        210,504,843        246,132,151        261,622,648        194,364,634        247,185,545        431,320,786        819,051,161        505,593,585        408,098,081        296,216,107        208,563,052        115,895,860        5,514,435,018  
   Ag Grade    g/tonne               16.0014        14.0585        14.7722        25.6046        10.2337        9.2817        10.8229        11.5356        8.5700        10.8990        18.9661        36.1140        22.2929        17.9941        13.0252        9.1961        10.0460        15.3716  
   Recovered Ag    grams      -        49,792,980        132,179,670        138,588,442        227,619,055        97,199,810        86,572,682        102,152,960        107,968,175        80,170,461        101,794,825        173,524,699        327,620,464        202,386,014        163,239,232        118,486,443        84,521,547        46,514,336        2,240,331,795  
   Sellable Ag    grams      -        49,544,015        131,518,772        137,895,499        226,480,960        96,713,811        86,139,818        101,642,195        107,428,334        79,769,609        101,285,850        172,657,075        325,982,362        201,374,084        162,423,036        117,894,011        84,098,939        46,281,764        2,229,130,136  

 

 

FORTE DYNAMICS, INC.

   P a g e | 355 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

FORTE DYNAMICS, INC
Company    i-80 Gold Corp.
Project Name    Mineral Point
Project Number    195005
Date    3/31/2025
Economic Model    DS_1v2 Sch_v1 With Inferred

 

Period     (1)      1      2      3      4      5      6      7      8      9      10      11      12      13      14      15      16      17      Total  
Year     2029      2030      2031      2032      2033      2034      2035      2036      2037      2038      2039      2040      2041      2042      2043      2044      2045      2046  

Revenue & Royalties

   Revenue Au    US$    $ -      $ 109,696,067      $ 458,326,033      $ 485,965,674      $ 756,850,900      $ 434,690,569      $ 443,301,766      $ 528,835,200      $ 530,156,249      $ 469,156,011      $ 597,743,462      $ 597,112,563      $ 455,875,093      $ 450,009,344      $ 457,158,920      $ 312,732,830      $ 367,060,893      $ 214,079,750      $ 7,668,751,322  
   Revenue Ag    US$    $ -      $ 43,405,868      $ 115,224,542      $ 120,811,238      $ 198,421,598      $ 84,731,665      $ 75,467,714      $ 89,049,458      $ 94,118,736      $ 69,886,728      $ 88,737,262      $ 151,266,105      $ 285,595,491      $ 176,425,283      $ 142,299,990      $ 103,287,790      $ 73,679,685      $ 40,547,787      $ 1,952,956,940  
   Royalties Au    US$    $ -      $ 3,290,882      $ 13,749,781      $ 14,578,970      $ 22,705,527      $ 13,040,717      $ 13,299,053      $ 15,865,056      $ 15,904,687      $ 14,074,680      $ 17,932,304      $ 17,913,377      $ 13,676,253      $ 13,500,280      $ 13,714,768      $ 9,381,985      $ 11,011,827      $ 6,422,393      $ 230,062,540  
   Royalties Ag    US$    $ -      $ 1,302,176      $ 3,456,736      $ 3,624,337      $ 5,952,648      $ 2,541,950      $ 2,264,031      $ 2,671,484      $ 2,823,562      $ 2,096,602      $ 2,662,118      $ 4,537,983      $ 8,567,865      $ 5,292,758      $ 4,269,000      $ 3,098,634      $ 2,210,391      $ 1,216,434      $ 58,588,708  
   Total Revenue    US$    $ -      $ 145,036,407      $ 547,126,095      $ 578,908,706      $ 910,740,594      $ 497,061,034      $ 497,168,978      $ 592,224,162      $ 598,017,236      $ 517,280,518      $ 658,787,321      $ 713,826,020      $ 696,378,827      $ 593,527,566      $ 570,091,143      $ 395,276,978      $ 421,623,986      $ 243,744,888      $ 9,176,820,460  

Capital Costs

   CAPEX Waste    US$    $ -      $ 287,251,043      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ 287,251,043  
   Mining Equipment & Sustaining CAPEX    US$    $ 263,794,346      $ 37,588,289      $ 18,611,419      $ 7,262,164      $ 7,683,192      $ 9,200,991      $ 7,984,491      $ 8,174,991      $ 7,793,991      $ 8,174,991      $ 8,365,491      $ 7,793,991      $ 8,174,991      $ 6,079,491      $ 4,078,480      $ 4,442,980      $ 3,490,480      $ 2,011,480      $ 420,706,249  
   Process    US$    $ 158,646,667      $ -      $ -      $ -      $ 40,413,333      $ -      $ -      $ 40,413,333      $ -      $ -      $ 40,413,333      $ -      $ -      $ 40,413,333      $ -      $ -      $ -      $ -      $ 320,300,000  
   Preproduction & Facilities    US$    $ 75,830,000      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ 75,830,000  
   Owner’s Cost    US$    $ 56,400,000      $ -      $ -      $ -      $ 9,300,000      $ -      $ -      $ 9,300,000      $ -      $ -      $ 9,300,000      $ -      $ -      $ 9,300,000      $ -      $ -      $ -      $ -      $ 93,600,000  
   Capex summary    US$    $ 554,671,013      $ 37,588,289      $ 18,611,419      $ 7,262,164      $ 57,396,526      $ 9,200,991      $ 7,984,491      $ 57,888,324      $ 7,793,991      $ 8,174,991      $ 58,078,824      $ 7,793,991      $ 8,174,991      $ 55,792,824      $ 4,078,480      $ 4,442,980      $ 3,490,480      $ 2,011,480      $ 910,436,249  
   Mine Equipment Contingency    US$    $ 39,569,152      $ 5,638,243      $ 2,791,713      $ 1,089,325      $ 1,152,479      $ 1,380,149      $ 1,197,674      $ 1,226,249      $ 1,169,099      $ 1,226,249      $ 1,254,824      $ 1,169,099      $ 1,226,249      $ 911,924      $ 611,772      $ 666,447      $ 523,572      $ 301,722      $ 63,105,937  
   Other Contingency    US$    $ 72,719,167      $ -      $ -      $ -      $ 12,428,333      $ -      $ -      $ 12,428,333      $ -      $ -      $ 12,428,333      $ -      $ -      $ 12,428,333      $ -      $ -      $ -      $ -      $ 122,432,500  
   Total CAPEX    US$    $ 666,959,331      $ 330,477,576      $ 21,403,131      $ 8,351,488      $ 70,977,338      $ 10,581,139      $ 9,182,164      $ 71,542,906      $ 8,963,089      $ 9,401,239      $ 71,761,981      $ 8,963,089      $ 9,401,239      $ 69,133,081      $ 4,690,252      $ 5,109,427      $ 4,014,052      $ 2,313,202      $ 1,383,225,729  

 

 

FORTE DYNAMICS, INC.

   P a g e | 356 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

FORTE DYNAMICS, INC
Company    i-80 Gold Corp.
Project Name    Mineral Point
Project Number    195005
Date    3/31/2025
Economic Model    DS_1v2 Sch_v1 With Inferred

 

Period     (1)     1     2     3     4     5     6     7     8      9      10      11      12      13      14      15      16      17      Total  
Year     2029     2030     2031     2032     2033     2034     2035     2036     2037      2038      2039      2040      2041      2042      2043      2044      2045      2046  
Operating Costs    Surface Ore    US$             $ 20,328,804     $ 62,499,863     $ 62,671,096     $ 60,806,437     $ 62,499,862     $ 62,499,862     $ 62,671,096     $ 62,499,863      $ 62,499,863      $ 62,499,862      $ 62,671,095      $ 62,499,863      $ 62,499,862      $ 62,499,863      $ 62,671,095      $ 62,499,863      $ 31,792,064      $ 988,610,312  
   Surface HL
Relocation
   US$            $ -     $ -     $ -     $ -     $ -     $ -     $ 13,667,587     $ -      $ -      $ -      $ -      $ -      $ -      $ 26,014,827      $ -      $ -      $ -      $ 39,682,414  
   Surface Waste    US$            $ 17,206,820     $ 267,105,314     $ 271,339,988     $ 266,379,507     $ 276,237,700     $ 228,041,227     $ 190,984,294     $ 197,207,512      $ 200,932,604      $ 222,394,919      $ 221,500,223      $ 196,731,203      $ 201,049,570      $ 23,779,190      $ 17,151,690      $ 43,740,256      $ 4,327,980      $
 
 
2,846,109,996
 
 
   Power    US$            $ -     $ -     $ -     $ -     $ -     $ -     $ -     $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -  
   Processing    US$            $ 31,712,935     $ 97,499,786     $ 97,766,909     $ 94,858,042     $ 97,499,785     $ 97,499,784     $ 97,766,910     $ 97,499,787      $ 97,499,786      $ 97,499,784      $ 97,766,909      $ 97,499,786      $ 97,499,784      $ 97,499,786      $ 97,766,908      $ 97,499,786      $ 49,595,621      $
 
 
1,542,232,087
 
 
   Transportation
and Refining
   US$            $ -     $ -     $ -     $ -     $ -     $ -     $ -     $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -      $ -  
   G&A    US$            $ 6,098,641     $ 18,749,959     $ 18,801,329     $ 18,241,931     $ 18,749,959     $ 18,749,959     $ 18,801,329     $ 18,749,959      $ 18,749,959      $ 18,749,958      $ 18,801,329      $ 18,749,959      $ 18,749,959      $ 18,749,959      $ 18,801,328      $ 18,749,959      $ 9,537,619      $ 296,583,094  
    Refining Cost Au     US$            $ 93,398     $ 390,231     $ 413,764     $ 644,403     $ 370,107     $ 377,439     $ 450,264     $ 451,389      $ 399,452      $ 508,934      $ 508,397      $ 388,144      $ 383,150      $ 389,237      $ 266,269      $ 312,525      $ 182,273      $ 6,529,375  
   Refining Cost Ag    US$            $ 800,440     $ 2,124,836     $ 2,227,859     $ 3,659,059     $ 1,562,522     $ 1,391,687     $ 1,642,146     $ 1,735,627      $ 1,288,769      $ 1,636,389      $ 2,789,472      $ 5,266,617      $ 3,253,428      $ 2,624,130      $ 1,904,712      $ 1,358,714      $ 747,735      $ 36,014,143  
   Total Operating
Cost
   US$            $ 76,241,038     $ 448,369,989     $ 453,220,945     $ 444,589,378     $ 456,919,935     $ 408,559,957     $ 385,983,625     $ 378,144,137      $ 381,370,432      $ 403,289,846      $ 404,037,425      $ 381,135,571      $ 383,435,752      $ 231,556,992      $ 198,562,003      $ 224,161,103      $ 96,183,292      $ 5,755,761,422  
Cash Flow Pre-Tax    Discounted @
0% (Net Cash
Flow)
   US$    $ (666,959,346   $ (261,088,672   $ 113,610,437     $ 110,080,520     $ 365,501,441     $ 25,357,785     $ 67,178,469     $ 120,631,147     $ 198,352,566      $ 118,942,186      $ 172,254,448      $ 284,788,781      $ 285,022,196      $ 130,380,521      $ 300,096,508      $ 178,778,993      $ 187,545,870      $ 124,026,042      $ 1,854,499,891  
   Discounted @
5%
   US$    $ (666,959,346   $ (248,655,878   $ 103,048,015     $ 95,091,692     $ 300,698,940     $ 19,868,488     $ 50,129,608     $ 85,730,304     $ 134,252,825      $ 76,671,194      $ 105,749,289      $ 166,510,102      $ 158,711,023      $ 69,143,574      $ 151,569,129      $ 85,995,752      $ 85,916,924      $ 54,112,151      $ 827,583,786  
   Discounted @
8%
   US$    $ (666,959,346   $ (241,748,771   $ 97,402,638     $ 87,385,466     $ 268,654,470     $ 17,258,082     $ 42,333,831     $ 70,387,116     $ 107,163,720      $ 59,500,706      $ 79,787,139      $ 122,141,027      $ 113,186,235      $ 47,940,647      $ 102,171,169      $ 56,358,595      $ 54,742,852      $ 33,520,388      $ 451,225,963  
   Discounted @
10%
   US$    $ (666,959,346   $ (237,353,338   $ 93,892,923     $ 82,705,124     $ 249,642,402     $ 15,745,189     $ 37,920,494     $ 61,902,852     $ 92,532,936      $ 50,443,098      $ 66,411,547      $ 99,816,730      $ 90,816,855      $ 37,766,593      $ 79,024,790      $ 42,798,270      $ 40,815,446      $ 24,537,891      $ 262,460,456  
   Discounted @
12%
   US$    $ (666,959,346   $ (233,114,886   $ 90,569,545     $ 78,353,140     $ 232,282,773     $ 14,388,688     $ 34,034,703     $ 54,567,405     $ 80,111,275      $ 42,891,745      $ 55,461,322      $ 81,869,969      $ 73,158,099      $ 29,879,850      $ 61,405,691      $ 32,662,254      $ 30,592,794      $ 18,063,691      $ 110,218,712  
   Cumulative Cash
Flow @0%
   US$    $ (666,959,346   $ (928,048,018   $ (814,437,581   $ (704,357,061   $ (338,855,620   $ (313,497,835   $ (246,319,367   $ (125,688,220   $ 72,664,347      $ 191,606,533      $ 363,860,981      $ 648,649,762      $ 933,671,957      $ 1,064,052,478      $ 1,364,148,986      $ 1,542,927,979      $ 1,730,473,849      $ 1,854,499,891           

 

 

FORTE DYNAMICS, INC.

   P a g e | 357 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

FORTE DYNAMICS, INC
Company    i-80 Gold Corp.
Project Name    Mineral Point
Project Number    195005
Date    3/31/2025
Economic Model    DS_1v2 Sch_v1 With Inferred

 

Period     (1)     1     2     3     4     5     6     7     8      9      10      11      12      13      14      15      16      17      Total  
Year     2029     2030     2031     2032     2033     2034     2035     2036     2037      2038      2039      2040      2041      2042      2043      2044      2045      2046  

Cash Flow After-Tax

   Discounted
@ 0% (Net
Cash Flow)
   US$     $ (666,959,331   $ (262,945,851   $ 107,406,046     $ 102,740,082     $ 347,914,491     $ 20,425,075     $ 61,562,267     $ 112,623,945     $ 190,359,486      $ 108,174,035      $ 141,464,532      $ 245,453,852      $ 244,796,708      $ 107,194,375      $ 252,402,612      $ 153,111,325      $ 150,595,310      $ 95,866,506      $ 1,470,031,733  
    Discounted 
@ 5%
   US$    $ (666,959,331   $ (250,424,620   $ 97,420,450     $ 88,750,746     $ 286,230,113     $ 16,003,580     $ 45,938,711     $ 80,039,735     $ 128,842,793      $ 69,729,947      $ 86,846,951      $ 143,511,784      $ 136,311,967      $ 56,847,465      $ 127,480,470      $ 73,649,165      $ 68,989,447      $ 41,826,239      $ 614,063,213  
   Discounted
@ 8%
   US$    $ (666,959,331   $ (243,468,380   $ 92,083,372     $ 81,558,389     $ 255,727,537     $ 13,900,963     $ 38,794,671     $ 65,714,990     $ 102,845,307      $ 54,113,949      $ 65,525,450      $ 105,270,950      $ 97,212,141      $ 39,415,149      $ 85,933,256      $ 48,267,075      $ 43,957,335      $ 25,909,740      $ 295,756,735  
   Discounted
@ 10%
   US$    $ (666,959,331   $ (239,041,683   $ 88,765,327     $ 77,190,144     $ 237,630,278     $ 12,682,364     $ 34,750,295     $ 57,793,892     $ 88,804,105      $ 45,876,351      $ 54,540,701      $ 86,030,078      $ 77,999,775      $ 31,050,392      $ 66,465,496      $ 36,653,634      $ 32,773,927      $ 18,966,677      $ 134,831,182  
   Discounted
@ 12%
   US$    $ (666,959,331   $ (234,773,081   $ 85,623,442     $ 73,128,361     $ 221,105,949     $ 11,589,736     $ 31,189,360     $ 50,945,353     $ 76,883,004      $ 39,008,641      $ 45,547,793      $ 70,562,117      $ 62,833,218      $ 24,566,184      $ 51,646,575      $ 27,972,867      $ 24,565,357      $ 13,962,414      $ 4,288,964  
   Cumulative
Cash Flow
@0%
   US$    $ (666,959,331   $ (929,905,182   $ (822,499,136   $ (719,759,055   $ (371,844,564   $ (351,419,489   $ (289,857,223   $ (177,233,277   $ 13,126,208      $ 121,300,243      $ 262,764,776      $ 508,218,627      $ 753,015,335      $ 860,209,710      $ 1,112,612,322      $ 1,265,723,646      $ 1,416,318,956      $ 1,512,185,462           

 

 

FORTE DYNAMICS, INC.

   P a g e | 358 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

 

FORTE DYNAMICS, INC

 

   
Company    i-80 Gold Corp.
   
Project Name    Mineral Point
   
Project Number    195005
   
Date    3/31/2025
   
Economic Model    DS_1v2 Sch_v1 With Inferred

 

 

 

LOGO

  Item   Unit   Pre-Taxes     After-Taxes     Delta (PRE-AFTER)  
                               
  NPV @ 0% (x1,000,000)  

US$

  $ 1,854.50      $ 1,470.03     $ 384.47   
  NPV @ 5% (x1,000,000)  

US$

  $ 827.58      $ 614.06      $ 213.52   
  NPV @ 8% (x1,000,000)  

US$

  $ 451.23      $ 295.76      $ 155.47   
  NPV @ 10% (x1,000,000)  

US$

  $ 262.46      $ 134.83      $ 127.63   
  NPV @ 12% (x1,000,000)  

US$

  $ 110.22      $ 4.29      $ 105.93   
  IRR   %     13.8%       12.1%       1.7%  
  Pay Back Period  

Years

    7.63        7.93        (0.30)  

*Pre-tax also does not include reclamation costs

 

         

LOGO

  Category   Total Costs (US$M)    

Unit Cost

(US$/Process ton)

   

Cost Per Ounce

(US$/Recovered toz

Au)

 
                           
  Mining   $ 3,874.40     $ 9.80     $ 1,097.75  
  Processing   $ 1,542.23     $ 3.90     $ 436.97  
  G&A   $ 296.58     $ 0.75     $ 84.03  
  Refining, Royalties & Net Proceeds Tax   $ 722.30     $ 1.83     $ 204.65  
  By-Product Credits   $ (1,952.96   $ (4.94   $ (553.34
  Total Operating Cost/Cash Costs   $ 4,482.57     $ 11.34     $ 1,270.07  
  Closure & reclamation   $ 69.83     $ 0.18     $ 19.78  
  Sustaining Capital   $ 388.43     $ 0.98     $ 110.05  
    All-in Sustaining Costs   $ 4,940.82     $ 12.49     $ 1,399.91  

 

 

FORTE DYNAMICS, INC.

   P a g e | 359 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

 

FORTE DYNAMICS, INC

 

   
Company    i-80 Gold Corp.
   
Project Name    Mineral Point
   
Project Number    195005
   
Date    3/31/2025
   
Economic Model    DS_1v2 Sch_v1 With Inferred

 

 

 

LOGO    Item    Unit    Pre-Tax      After-Tax  
   -25%      0%      25%      -25%      0%      25%  
                                                                    
   Gold Price     Price    US$        $   1,631.25       $   2,175.00       $   2,718.75       $   1,631.25       $   2,175.00       $   2,718.75   
   NPV @ 5% (x1,000,000)    US$    $ (211.24)      $ 827.58       $ 2,274.36       $ (395.58)      $ 614.06       $ 1,523.01   
   NPV @ 8% (x1,000,000)    US$    $ (398.88)      $ 451.23       $ 1,610.03       $ (543.65)      $ 295.76       $ 1,002.44   
   NPV @ 10% (x1,000,000)    US$    $ (489.02)      $ 262.46       $ 1,272.92       $ (613.88)      $ 134.83       $ 740.04   
   NPV @ 12% (x1,000,000)    US$    $ (559.10)      $ 110.22       $ 998.45       $ (667.80)      $ 4.29       $ 527.55   
   IRR    %      2.6%        13.8%        26.3%        0.3%        12.1%        19.9%  
                                                                    
   Silver Price     Price    US$    $ 20.44       $ 27.25       $ 34.06       $ 20.44       $ 27.25       $ 34.06   
   NPV @ 5% (x1,000,000)    US$    $ 722.07       $ 827.58       $ 1,341.05       $ 346.48       $ 614.06       $ 828.39   
   NPV @ 8% (x1,000,000)    US$    $ 357.97       $ 451.23       $ 853.19       $ 55.79       $ 295.76       $ 439.39   
   NPV @ 10% (x1,000,000)    US$    $ 175.96       $ 262.46       $ 607.94       $ (88.24)      $ 134.83       $ 245.38   
   NPV @ 12% (x1,000,000)    US$    $ 29.57       $ 110.22       $ 409.78       $ (203.19)      $ 4.29       $ 89.65   
   IRR    %      12.5%        13.8%        18.1%        8.7%        12.1%        13.4%  
   CAPEX     Price (x1,000,000)    US$    $ 0.00       $ 0.00       $ 0.00       $ 0.00       $ 0.00       $ 0.00   
   NPV @ 5% (x1,000,000)    US$    $ 1,274.09       $ 827.58       $ 789.03       $ 834.13       $ 614.06       $ 349.07   
   NPV @ 8% (x1,000,000)    US$    $ 835.49       $ 451.23       $ 375.67       $ 480.58       $ 295.76       $ 20.75   
   NPV @ 10% (x1,000,000)    US$    $ 615.19       $ 262.46       $ 168.71       $ 304.35       $ 134.83       $ (142.12)  
   NPV @ 12% (x1,000,000)    US$    $ 437.30       $ 110.22       $ 2.04       $ 162.99       $ 4.29       $ (272.27)  
   IRR    %      19.8%        13.8%        12.0%        15.0%        12.1%        8.2%  
   Mining Cost     Price    US$/ton    $ 1.88       $ 2.50       $ 3.13       $ 1.88       $ 2.50       $ 3.13   
   NPV @ 5% (x1,000,000)    US$    $ 1,762.08       $ 827.58       $ 289.26       $ 1,263.47       $ 614.06       $ (92.93)  
   NPV @ 8% (x1,000,000)    US$    $ 1,218.56       $ 451.23       $ (17.28)      $ 815.50       $ 295.76       $ (324.50)  
   NPV @ 10% (x1,000,000)    US$    $ 942.49       $ 262.46       $ (167.47)      $ 589.06       $ 134.83       $ (435.96)  
   NPV @ 12% (x1,000,000)    US$    $ 717.57       $ 110.22       $ (286.25)      $ 405.33       $ 4.29       $ (522.73)  
   IRR    %      23.3%        13.8%        7.8%        18.6%        12.1%        4.1%  
   Processing Cost     Price    US$/ton    $ 2.93       $ 3.90       $ 4.88       $ 2.93       $ 3.90       $ 4.88   
   NPV @ 5% (x1,000,000)    US$    $ 1,283.81       $ 827.58       $ 776.71       $ 787.30       $ 614.06       $ 391.96   
   NPV @ 8% (x1,000,000)    US$    $ 808.07       $ 451.23       $ 401.00       $ 407.25       $ 295.76       $ 91.23   
   NPV @ 10% (x1,000,000)    US$    $ 568.91       $ 262.46       $ 213.16       $ 217.74       $ 134.83       $ (57.82)  
   NPV @ 12% (x1,000,000)    US$    $ 375.67       $ 110.22       $ 62.07       $ 65.64       $ 4.29       $ (176.82)  
   IRR    %      17.6%        13.8%        13.0%        13.0%        12.1%        9.2%  

 

 

FORTE DYNAMICS, INC.

   P a g e | 360 of 362    i-80 Gold Corp.


LOGO

 

   March 29, 2025
 

 

FORTE DYNAMICS, INC

 

Company    i-80 Gold Corp.
Project Name    Mineral Point
Project Number    195005
Date    3/31/2025
Economic Model    DS_1v2 Sch_v1 With Inferred

 

 

 

LOGO   Item   Unit    Imperial      Metric  
                         
  Processing Material   Ton -> Tonne      395,444,125        358,740,979  
  OPEX Waste   Ton -> Tonne      1,138,443,998        1,032,779,319  
  CAPEX Waste   Ton -> Tonne      114,900,417        104,235,935  
  Heap Leach Relocation   Ton -> Tonne      26,454,942        23,999,527  
  Total Material   Ton -> Tonne      1,675,243,483        1,519,755,759  
  Contained Au   Toz -> grams      4,524,542        140,729,097  
  Au Grade   Toz/ton -> g/tonne       0.0114        0.392  
  Recovered Au   Toz -> grams      3,529,392        109,776,446  
  Sellable Au   Toz -> grams      3,525,863        109,666,670  
  Contained Ag   Toz -> grams      177,293,070        5,514,435,018  
  Ag Grade   Toz/ton -> g/tonne       0.448        15.372  
  Recovered Ag   Toz -> grams      72,028,286        2,240,331,795  
    Sellable Ag   Toz -> grams      71,668,145        2,229,130,136  

 

 

FORTE DYNAMICS, INC.

   P a g e | 361 of 362    i-80 Gold Corp.


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   March 29, 2025
 

 

LOGO

 

 

FORTE DYNAMICS, INC.

   P a g e | 362 of 362    i-80 Gold Corp.
EX-99.2 3 d913666dex992.htm EX-99.2 EX-99.2

Exhibit 99.2

 

LOGO

S-K 1300 Technical Report Summary on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada Prepared for i-80 Gold Corp. https://www.i-80gold.com Prepared by Abani R. Samal, RM-SME Brian Arthur, RM-SME Paul Gates, PE Effective Date: December 31, 2024 Publication date: March 24, 2025


Lone Tree Property

Battle Mountain, Nevada, U.S.A.

NI 43-101 Technical Report

October 20, 2021

 

 

NOTES ON FORWARD-LOOKING INFORMATION

This Technical Report Summary (TRS) contains “forward-looking statements” within the meaning of Section 27A of the Securities Act of 1933, as amended, and Section 21E of the Securities Exchange Act of 1934, as amended (and the equivalent under Canadian securities laws), which are intended to be covered by the safe harbor created by such sections. Words such as “may”, “will”, “should”, “expects”, “intends”, “projects”, “believes”, “estimates”, “targets”, “anticipates” and similar expressions are used to identify these forward-looking statements. Such forward-looking statements include, without limitation, statements regarding i-80 ’s expectation for its mines and any related development or expansions, including estimated cash flows, production, revenue, costs, taxes, capital, rates of return, mine plans, material mined and processed, recoveries and grade, future mineralization, future adjustments and sensitivities and other statements that are not historical facts. Other forward-looking statements in this Report may involve, without limitation, the following

 

   

Assumed commodity prices and exchange rates

 

   

Proposed mining and process production plan

 

   

Projected mining and process recovery rates

 

   

Sustaining capital costs and proposed operating costs

 

   

Assumptions as to dewatering processes

 

   

Assumptions about environmental, permitting, and social risks

For a more detailed discussion of such risks and other factors, see the latest Annual Information Form and Consolidated Financial Statements, available on the i-80 website. i-80 does not undertake any obligation to release publicly, revisions to any “forward-looking statement,” including, without limitation, outlook, to reflect events or circumstances after the date of this presentation, or to reflect the occurrence of unanticipated events, except as may be required under applicable securities laws. Investors should not assume that any lack of update to a previously issued “forward-looking statement” constitutes a reaffirmation of that statement. Continued reliance on “forward-looking statements” is at investors’ own risk.

 

 

2


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T A B L E O F C O N T E N T S

 

1.    SUMMARY

     8  

1.1   Mineral Resource Estimates

     8  

1.2   Location of the Lone Tree Mine

     8  

1.3   Site Infrastructure

     9  

1.4   Land Tenure and Ownership

     10  

1.5   Geology and Mineralization

     11  

1.6   On-site Verification of Data and Information

     12  

1.7   Historical Summary of the Deposit

     12  

2.    INTRODUCTION

     13  

2.1   Terms of Reference and Purpose of this Report

     14  

2.2   Qualified Persons and Firms Engaged in this Project

     14  

2.2.1  GeoGlobal LLC

     14  

2.2.2  Dr. Abani R. Samal – Project Manager & Principal Geologist

     15  

2.2.3  Mr. Brian Arthur – Principal Metallurgist

     15  

2.2.4  Mr. Paul Gates – Principal Mining Engineer

     16  

2.2.5  Responsibilities of the Qualified Persons

     16  

2.3   Scope of this Project

     16  

2.4   Units of Measure

     18  

2.4.1  Coordinate Reference System (CRS) Used in the Maps:

     18  

2.4.2  Conversion Factors Used

     18  

2.5   Definitions

     18  

3.    PROPERTY DESCRIPTION

     19  

3.1   Location of the Project

     19  

3.2   Land Status

     19  

4.    ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

     21  

4.1   Climate

     21  

4.2   Physiography

     22  

4.3   Vegetation

     22  

4.4   Site Infrastructure

     22  

4.4.1  Water Wells:

     22  

4.4.2  Electric Power

     23  

4.4.3  Mineral Processing & Metallurgy Facilities

     23  

4.4.4  Laboratory Facilities:

     25  

5.    HISTORY

     26  

5.1   Historical Summary of the District

     26  

6.    GEOLOGICAL SETTING, MINERALIZATION & DEPOSIT

     26  

6.1   Regional Geology

     26  

6.2   Deposit Geology and Mineralization

     29  

6.3   Controls of Mineralization

     32  

6.4   Alteration

     35  

6.5   Gold Mineralogy

     35  

6.6   Deposit Type

     35  

7.    EXPLORATION

     36  

7.1   Exploration History

     36  

7.1.1  Drilling and Sampling

     37  

7.1.2  Recent Exploration Drilling

     37  

7.1.3  Rotary Drilling

     39  

7.1.4  Reverse Circulation Drilling

     40  

7.1.5  Core Drilling

     41  

7.2   Collar Surveys/Locations

     43  

 

 

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7.3   Down-Hole Surveys

     43  

7.4   Opinion of the Qualified Person

     44  

8.    SAMPLE PREPARATION, SECURITY, AND ANALYSES

     44  

8.1   Sampling

     44  

8.2   Sample Preparation and Analysis

     44  

8.3   Data Security

     45  

8.4   QA/QC Procedures

     45  

8.4.1  Standards

     46  

8.4.2  Blanks

     46  

8.4.3  Check Samples

     46  

8.5   Opinion of the Qualified Person

     46  

9.    DATA VERIFICATION

     46  

9.1   On-site Verification of Data and Information

     46  

9.2   Review of the QA/QC procedure

     51  

9.2.1  Re-assay of pulps and drill cores

     51  

9.2.2  Database Review

     53  

9.3   Opinion of the Qualified Person

     54  

10.   MINERAL PROCESSING AND METALLURGICAL TESTING

     54  

10.1   Opinion of the Qualified Person

     55  

11.   MINERAL RESOURCE ESTIMATES

     56  

11.1   Datasets

     56  

11.1.1 Drill Holes

     56  

11.1.2 Other Data Sets

     56  

11.2   Geological Interpretation and Lithological Models

     58  

11.3   Lithology Model

     60  

11.4   Exploratory Data Analysis and Compositing

     60  

11.4.1 Compositing

     61  

11.4.2 Statistical Analyses of Composited Data

     61  

11.4.3 Contact Plots

     64  

11.5   Geological Domains

     66  

11.6   Variogram Analysis

     66  

11.7   Density /Tonnage Factor Model

     68  

11.8   Grade Interpolation

     69  

11.9   Resource Classification

     71  

11.10 Criteria for Reasonable Prospect for Economic Extraction

     71  

11.10.1  Inferred blocks

     72  

11.10.2  Indicated blocks

     73  

11.10.3  Inventory of Mineral Resources

     73  

11.11 Model Validation

     75  

11.11.1  Cross Sections

     75  

11.11.2  Statistical Validation

     76  

11.12 Tabulation of Estimated Resources

     83  

11.13 Opinion of the Qualified Person

     83  

12.   MINERAL RESERVE ESTIMATES

     83  

13.   MINING METHODS

     84  

14.   PROCESSING AND RECOVERY METHODS

     84  

14.1   Heap Leach

     85  

14.2   Oxide Milling

     85  

14.3   Flotation

     86  

14.4   Autoclaving

     86  

14.5   Process Operating Costs

     87  

14.5.1 Oxide Leach Cost Factors

     88  

14.5.2 Oxide Mill Cost

     88  

 

 

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14.5.3 Flotation Cost updated with PPI factors

     88  

14.5.4 Acidic Autoclave

     88  

14.6   Model Estimate

     89  

14.7   Opinion of the Qualified Person

     89  

15.   PROJECT INFRASTRUCTURE

     89  

16.   MARKETING STUDIES AND CONTRACTS

     89  

17.   ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

     89  

17.1   Environmental Liabilities

     89  

17.2   Dewatering

     90  

17.3   Current Permits

     90  

18.   CAPITAL AND OPERATING COSTS

     91  

19.   ECONOMIC ANALYSIS

     91  

20.   ADJACENT PROPERTIES

     91  

21.   OTHER RELEVANT DATA AND INFORMATION

     94  

22.   TECHNICAL REPORT INTERPRETATION AND CONCLUSIONS

     94  

23.   RECOMMENDATIONS

     94  

23.1   Resource Model update

     94  

23.2   Risk Analyses in Resource Eestimates

     94  

23.3   Future Exploration

     94  

23.4   Geometallurgical Study

     95  

24.   REFERENCES

     96  

25.   RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

     99  

26.   QUALIFIED PERSONS CERTIFICATES

     100  

27.   APPENDIX A:DETAILS OF THE LAND CLAIMS OF THE LONE TREE PROJECT

     104  

28.   APPENDIX B:VARIOGRAM MAPS

     114  

29.   APPENDIX C:VARIOGRAM MODELS

     119  

T A B L E S

 

Table 1-1:Estimated Mineral Resources at 0.62 g/t Au Cut-off Grade

     8  

Table 2-1:QPs and Responsibilities for the Contents of the Report

     16  

Table 3-1:The Area of the Lone Tree Claims

     19  

Table 7-1:Summary of Drilling by Hole Type

     37  

Table 9-1:Details of the Samples Selected for Re-assay

     51  

Table 9-2:Descriptions of the CRMs Used

     52  

Table 10-1:Recoveries and Material Types or Lone Tree

     55  

Table 11-1:Block Model Geometry

     60  

Table 11-2:The Lithology Codes Used in the Block Model

     60  

Table 11-3:General Statistical Characteristics of the Gold (AuFA, opt)

     62  

Table 11-4:Variogram Parameters of AuFA

     68  

Table 11-5:Tonnage Factor by Geologic Unit (ft3/ton)

     68  

Table 11-6:Interpolation Parameters Used for Estimation of Gold Grades

     70  

Table 11-7:Optimum Pit Criteria Applied to Resource Estimate

     72  

Table 11-8:Inventory of Mineral Resources Within $2,175 Pit Shell

     74  

Table 11-9:Comparison of Gold Grades for Composite Data by Lithologic Unit

     80  

Table 11-10: Estimated Mineral Resources at 0.62 g/T Cut-off Grade

     83  

 

 

5


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Table 14-1:Suggested operating Costs for Lone Tree Ores

     88  

Table 20-1:Mineral Resources for Nearby Properties

     92  

Table 27-1:Lone Tree Unpatented Claims, Claimant: Goldcorp Dee LLC

     104  

Table 27-2:Claimant: VEK/ANDRUS ASSOCIATES & Goldcorp Dee LLC

     108  

Table 27-3:Claimant: Larie K. Richardson, Lessee: Goldcorp Dee LLC

     109  

Table 27-4:Lone Tree Brooks Project, Claimant: Goldcorp Dee LLC

     111  

Table 27-5:Lone Tree Buffalo Mountain Project, Claimant: Goldcorp Dee LLC

     112  

Table 27-6:Lone Tree Patented Lands

     113  

F I G U R E S

 

Figure 1-1:Location of the Lone Tree Gold Deposit

     9  

Figure 1-2:The Location of Lone Tree Deposit and Infrastructure

     10  

Figure 1-3:The Property Boundary for the Lone Tree Deposit Project

     11  

Figure 1-4:The Mineralized Zone of the Lone Tree Deposit.

     12  

Figure 2-1:The Property Boundary of the Lone Tree Project

     13  

Figure 3-1:Location of Lone Tree Property and Land Position

     20  

Figure 4-1:Location and Accessibility of the Lone Tree Mining Project

     21  

Figure 4-2:Autoclave Agitator Access Ports

     24  

Figure 4-3:Pulverizing and Preparation Stations

     24  

Figure 4-4:The LECO Facility

     25  

Figure 4-5:ICP Laboratory Facility at Lone Tree

     25  

Figure 6-1:Location of the Lone Tree Mine Relative to Major Mineral Trends in Nevada

     27  

Figure 6-2:Location of Lone Tree Deposit in the Battle Mountain Mining District

     28  

Figure 6-3:General Stratigraphic Sequence of Lone Tree

     30  

Figure 6-4:Local Geology of Lone Tree Deposit and Controls of MIneralizaton

     31  

Figure 6-5:Diagrammatic Model of Geology of Distal-Disseminated Ag-Au Deposit

     36  

Figure 7-1:Location of Selected 1840 Drill Holes for Resource Estimation

     38  

Figure 7-2:Exploration Drill Hole LTE-20001 (Source: NGM)

     39  

Figure 9-1:Location of LTE-20001 on the West Side of the Lone Tree Pit

     47  

Figure 9-2:Location Tag of an Exploration Drill Hole in Sequoia Zone

     48  

Figure 9-3:An Example of a Mineralized Core Stored in a Core Box

     49  

Figure 9-4:Half Core of LTE-20001 from 2544 ft to 2549 ft with 1.215 ppm Gold Grade

     49  

Figure 9-5:The Chip Samples from the Upper Portion of the Hole (LTE-20001)

     50  

Figure 9-6:The Pulps Preserved in a Storage Area of i-80

     50  

Figure 9-7:Comparison of the Re-assayed Analytical Results with the Original Data

     52  

Figure 9-8:The Relationship Between the % Differences Original Au Assay Data

     53  

Figure 11-1:Location of the Drill Holes at the Lone Tree Project

     57  

Figure 11-2:The Geological Codes Used In Creating Solid Model

     58  

Figure 11-3:A Vertical Cross Section at 28000N (Looking North)

     59  

Figure 11-4:A Vertical Cross Section with Rock-type Models (27300N)

     59  

Figure 11-5:The Lithology Block Model (28000 N, Looking North)

     60  

Figure 11-6:Histogram of Drill Hole Assay Lengths

     61  

Figure 11-7:Comparison of Data Statistics of All Lithology Types

     62  

Figure 11-8:Comparison of Sample Means and Confidence Intervals

     63  

Figure 11-9:Histograms of Gold Assay Values (AuFA) by Lithology

     63  

Figure 11-10:Cumulative Frequency Plots of Gold Composites (AuFA)

     64  

Figure 11-11:Contact Plot of AuFA for the Boundary Between Qal and Phv

     65  

Figure 11-12:Contact Plot of AuFA for the Boundary Between Phv and Pem

     65  

Figure 11-13:Contact Plot of AuFA for the Boundary Between Pem and Pb

     65  

 

 

6


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Figure 11-14:Contact Plot of AuFA for the Boundary Between Pb and Ova

     66  

Figure 11-15:Steps Followed in Variogram Analysis

     66  

Figure 11-16:Blocks Below the Current Pit Used for the Optimized Pit shell

     71  

Figure 11-17:The Blocks Classified as Indicated and Inferred Category Resources

     73  

Figure 11-18:The Grade Tonnage Curve of Indicated Category Resources

     74  

Figure 11-19:The Grade Tonnage Curve of Inferred Category Resources

     75  

Figure 11-20:Estimate Blocks with Assay Data Within the $2,175 Pit Shell

     75  

Figure 11-21:Histogram of Holes Used for Estimating Indicated and Inferred Blocks

     76  

Figure 11-22:Histogram of the Holes Used for Estimating Indicated and Inferred Blocks

     77  

Figure 11-23:Number of Composites Used to Estimate Blocks with Gold Grade

     77  

Figure 11-24:Number of Composites Used to Estimate Blocks

     78  

Figure 11-25 Average Distance of Samples From Block Centers

     78  

Figure 11-26:Minimum Distance of Composites From Blocks

     79  

Figure 11-27:Comparison of Composites and Block Grade Estimates in the Ova

     80  

Figure 11-28:Comparison of Composites and Block Grade Estimates in the Pb

     81  

Figure 11-29:Comparison of Composites and Block Grade Estimates in the Pem

     81  

Figure 11-30:Comparison of Composites and Block Grade Estimates in the Phv

     82  

Figure 11-31:Comparison of Composites and Block Grade Estimates in the Qal

     82  

Figure 20-1:Mineral Deposits Adjacent to Lone Tree

     93  

Figure 23-1:A Vertical Cross Section (Looking East) Along East 83700 North-South

     95  

Figure 28-1:Variogram map of AuFA in Qal

     114  

Figure 28-2:Variogram Map of AuFA in Phv

     115  

Figure 28-3:Variogram Map of AuFA in Pem

     116  

Figure 28-4:Variogram Map of AuFA in Pb

     117  

Figure 28-5:Variogram Map of AuFA in Ova

     118  

Figure 29-1:Variogram Model of Lithology code 1 (Qal)

     119  

Figure 29-2:Variogram Model of Lithology Code 2 (Phv)

     120  

Figure 29-3:Variogram Model of Lithology Code 3 (Pem)

     121  

Figure 29-4:Variogram Model of Lithology Code 4 (Pb)

     122  

Figure 29-5:Variogram Modle of Lithology Code 5 (Ova)

     123  

 

 

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1.

SUMMARY

The acquisition of the Lone Tree Property from Nevada Gold Mines (NGM) in 2021 provided i-80 access to processing infrastructure including an autoclave, CIL (carbon-in-leach) mill, a flotation plant, and a heap leach facility complete with an assay lab, and gold refinery. In addition to processing facilities, the land package includes the Lone Tree Mine, the Buffalo Mountain deposit, and the Brooks open pit mine, which are currently on care and maintenance. During this acquisition, independent verification of the quantity of mineral resources at Lone Tree was conducted by GeoGlobal using the available drill hole data and information shared by NGM at that time. A technical report was produced as per the NI 43-101 guidelines and published in October 2021. The acquisition of the Lone Tree mine (along with the processing facilities mentioned above) was completed after the publication of this report. Since then the market conditions have also changed, and there has been an increase in the percentage of shareholders in the United States and the requirement to file an S-K 1300 Technical Report.

GeoGlobal was engaged by i-80 to review the resource estimates and update them as required based on current market conditions and produce this report conforming to the United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 229.1300 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300) and Item 601 (b)(96) Technical Report Summary (TRS).

The effective date of published mineral resources is December 31, 2024.

 

1.1

Mineral Resource Estimates

The in-situ Lone Tree mineral deposit hosts substantial gold mineral resources as shown below. The summary of estimated resources at the end of the fiscal year ended December 31, 2024, is shown in Table 1-1. These mineral resources are estimated using a gold price of $2,175/oz Au and an open-pit cut-off grade of 0.62 g/T Au. More details about the estimated mineral resources are presented in section 11.12. Mineral resources are not mineral reserves and do not have demonstrated economic viability.

Table 1-1: Estimated Mineral Resources at 0.62 g/t Au Cut-off Grade

 

      Million
Tonnes (MT)
   Au (g/T)      Au (K ozs)  

 Indicated Mineral Resources

   7.69    1.73    428 

 Inferred Mineral Resources

   52.94    1.64    2,789 

Resource expansion potential exists down-plunge of the main Lone Tree deposit and in the unmined Sequoia zone where previous drilling returned multiple wide, high-grade, intercepts.

 

1.2

Location of the Lone Tree Mine

The Lone Tree mine project is approximately 30 miles east of Winnemucca, Nevada, and 20 miles northwest of Battle Mountain, Nevada at 40° 50’ 19” N, 117° 12’ 37” W. The land package includes the process area, the Lone Tree Pit, and the Buffalo Mountain Property.

 

 

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The mine office is accessible from Interstate 80 by a paved highway (Figure 1-1).

Figure 1-1: Location of the Lone Tree Gold Deposit

 

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1.3

Site Infrastructure

The Lone Tree site has an employee parking lot with secure access to the administration building and a security gate for vehicles entering the mine property. The site has many buildings on the property. Processing infrastructure at Lone Tree includes an autoclave, carbon-in-leach mill, flotation mill, and heap leach facility. The list of processing facilities includes the following:

 

 

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Lone Tree Autoclave, which processes higher-grade refractory ore.

 

   

Lone Tree Flotation Plant, which processes lower-grade refractory ore.

 

   

Lone Tree Leach pad (Phases 1-4, Figure 1-2), which treats oxide ore in a cyanide heap-leach process.

 

   

Lone Tree Leach pad (Phase 5, Figure 1-2), which treats oxide ore in a cyanide heap-leach process.

 

   

Lone Tree Leach pad (Phase 6, Figure 1-2), which treats oxide ore in a cyanide heap-leach process.

The autoclave and flotation mill, which are currently in care and maintenance.

Details of the site infrastructure are provided in section 4.4.

Figure 1-2: The Location of Lone Tree Deposit and Infrastructure

 

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1.4

Land Tenure and Ownership

Figure 1-3 shows the property boundary of the Lone Tree project. It is significant to note that the land package includes the process area, the Lone Tree Pit and Buffalo Mountain exploration project.

The Lone Tree Properties include interests in fee lands, mineral rights in fee lands, patented mining claims, and unpatented mining claims which are leased or owned by i-80 as of the effective date of this report. All interests were acquired from Nevada Gold Mines and a full description of land and interests is shown in Figure 3-1. Details of all claims are provided in Appendix A. These properties are identified as part of the Lone Tree Mine Plan of Operations (PoO) in Sections 11, 12, 13 and 14.

 

 

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More details are available in Section 4.

Figure 1-3: The Property Boundary for the Lone Tree Deposit Project

 

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1.5

Geology and Mineralization

Mineralization is structurally controlled within three Paleozoic rock sequences at the Lone Tree deposit. The oldest of these three is the Valmy Formation which is unconformably overlain by rocks of the Pennsylvanian Antler Sequence of the Battle and Edna Mountain Formations. The Pennsylvanian-Permian Havallah sequence rocks were thrust over the Antler Sequence rocks in the mine area. The Havallah Sequence is dominated by siltstones, chert and basalts with lesser sandstones and conglomerates. Amongst the three mineralized Paleozoic sequences, Antler Sequence rocks appear to have been preferentially mineralized within the structural zones.

Out of three principal mineralized zones, the Wayne Zone, the Sequoia Zone, and the Antler High Zone, the Wayne Zone is the most preferred zone with a higher amount of mineralized material. The main structural component of the Wayne zone (Figure 1-4) is the north-south trending Powerline Fault, shown in Figure 1-4. While the pit bottom is currently underwater, the footwall of the Powerline fault appears to be exposed on the east wall of the Lone Tree mine.

 

 

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Figure 1-4: The Mineralized Zone of the Lone Tree Deposit.

 

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1.6

On-site Verification of Data and Information

As part of the S-K 1300 report work, Mr. Brian Arthur and Dr. Abani Samal visited the site on Wednesday, August 28th, and Thursday, August 29th, 2024. Dr. Abani Samal also completed a site visit on July 7th and 8th, 2021, as part of the 2021 NI 43-101 report. During the site visit, the following tasks were completed.

 

   

Field review of the general geology of the deposit.

 

   

Physical verification of selected drill hole intercepts.

 

   

Verification of available core, logging, and cutting facilities at the new Lone Tree storage facility.

 

   

A visit to the leach pads and the mill facilities.

 

   

Detailed discussions with site personnel were held.

 

   

A visit to the on-site laboratory facilities.

The locations of selected holes were completed in 2021 and have not changed.

 

1.7

Historical Summary of the Deposit

In the early days, the Lone Tree Hill area was explored for copper. Cordex discovered the southern extension of the Lone Tree gold deposit in 1988, which was referred to as the Stonehouse deposit at the time. Santa Fe Pacific Gold discovered the main part of the Lone Tree deposit in the pediment on the west flank of the hill in 1989 and acquired the Stonehouse portion of the deposit. Santa Fe Pacific Gold began producing gold from the site in 1991. Newmont acquired the deposit from Santa Fe Pacific Gold through a merger and continued operations in 1997. Newmont completed mining operations in 2006.

 

 

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Reclamation activities began in 2007 and residual leaching continues into the present. In July 2019, the non-operating Lone Tree project became part of Nevada Gold Mines, a joint venture between Barrick and Newmont.

In 2021, i-80 Gold acquired the Lone Tree mine along with the processing facilities including an autoclave, CIL (carbon-in-leach) mill, a flotation plant, an assay lab and a heap leach facility. Gold is still being recovered from the leach pad.

 

2.

INTRODUCTION

The Lone Tree Property was acquired on October 14, 2021, by i-80 Gold Corp (i-80) from Nevada Gold Mines (NGM). This acquisition provided i-80 with important processing infrastructure including an autoclave, a flotation mill, CIL (carbon-in-leach) mill, and a heap leach facility complete with an assay lab, carbon columns, and interim toll processing agreements for i-80 at NGM facilities. The property boundary contains the existing mine and the process facilities as shown in Figure 2-1.

Figure 2-1: The Property Boundary of the Lone Tree Project

 

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In addition to the processing facilities, the land package also includes the Buffalo Mountain deposit and the Brooks open pit mine (Figure 2-1), which is currently on care and maintenance.

 

 

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2.1

Terms of Reference and Purpose of this Report

During this acquisition, independent verification of the quantity of mineral resources at Lone Tree was conducted by GeoGlobal using the available drill hole data and information shared by NGM at that time. A technical report with the title ‘Technical Report on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada’ was produced as per the NI 43-101 guidelines and published in October 2021 (henceforth referred to as the 2021 report).

The acquisition of the Lone Tree mine (along with the processing facilities mentioned above) was completed after the publication of the 2021 report. Additionally, the composition of the shareholders of i-80 has recently changed. Therefore, i-80 engaged GeoGlobal again to review the resource estimates and update as required based on current market conditions and produce this report conforming to the United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 229.13001 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300) and Item 601 (b)(96)2 Technical Report Summary (TRS).

The effective date of the published mineral resources is December 31, 2024.

 

2.2

Qualified Persons and Firms Engaged in this Project

This project was led and managed by Dr. Abani Samal, a geologist working for GeoGlobal LLC. The following serve as the qualified persons or qualified firms for this Report in compliance with 17 CFR § 229.1302 (b)(1)(i) and (ii) qualified person definition.

 

2.2.1

GeoGlobal LLC

GeoGlobal, LLC (GeoGlobal) is an international service-providing company to the mineral industry based in the Salt Lake City area in the state of Utah, USA. GeoGlobal provides services to mineral exploration and mining projects around the world. Website: https://geoglobal.co/

Dr. Abani R. Samal is the Principal and owner of the company. The GeoGlobal retains a team of highly experienced professionals as Principal Associates, who together have more than 20 years of experience. The company specializes in mineral deposit evaluation for various commodities globally.

Projects were conducted for various entities including Waterton Copper, First Majestic, i-80 Gold Corp, Equinox Gold Corp, Nevada Copper, World Bank / Government of Nigeria, Rio Tinto, Resource Capital Funds, Orion Mining Finance / Aquila Resources, and Forbes & Manhattan.

 

 

1 Title 17—Commodity and Securities Exchanges, Chapter II—Securities and Exchange Commission, Part 229—Standard Instructions for Filing Forms Under Securities Act of 1933, Securities Exchange Act of 1934 and Energy Policy and Conservation Act of 1975—Regulation S-K The GeoGlobal team has provided support and services to many exploration and mining projects in the USA, Canada, Brazil, Mexico, Kazakhstan, and most recently in Namibia and Nigeria.

2 https://www.ecfr.gov/current/title-17/chapter-II/part-229/subpart-229.600/section-229.601

 

 

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The types of projects include strategic project development, resource estimation, mine-mill reconciliation, etc.

The following professionals from GeoGlobal conducted this project for i-80 Gold Corp.

 

2.2.2

Dr. Abani R. Samal – Project Manager & Principal Geologist

Dr. Abani Samal has extensive training (M.S. and Ph.D.) in economic geology, mineral resource estimation, and geostatistics. He is the owner of GeoGlobal LLC where he works as a Principal.

Prior to GeoGlobal, Dr. Samal worked as the lead resource geologist with Rio Tinto at their corporate regional HQ based in Salt Lake City, Utah, USA, where he led and contributed to various major projects. He also worked as the lead geologist and geostatistician at an international mining consulting firm, Pincock Allen and Holt (now RPM Global), based in Denver, USA.

Dr. Samal has more than 20 years of experience in the mineral industry. He has experience in various commodities including gold-silver, base metals such as Cu-Mo-Au, and ferrous deposits around the world. Dr. Samal’s gold-silver experience includes his contribution to various projects such as Jerritt Canyon, Lone Tree, Las Brisas Venezuela, Pascua-Lama (Chile), Chapada (Brazil), and Black-Fox (Canada). He has led/completed many other confidential projects for M&A, corporate assurance, and corporate financing.

Dr. Samal is a Registered Member of the Society for Mining Metallurgy & Exploration (RM-SME), a Certified Professional Geologist (CPG), and a Fellow of the Society for Economic Geologists (SEG).

Dr. Samal is the project manager for this project and a Qualified Person (QP) for the sections mentioned in Section 2.2.5. He is also the lead QP on behalf of GeoGlobal LLC.

 

2.2.3

Mr. Brian Arthur – Principal Metallurgist

Mr. Brian Arthur is an independent consulting metallurgist who gained more than 30 years of professional experience in precious metal process operations and development with Newmont before becoming independent. He designed hydrometallurgical flowsheets to recover copper, nickel, cobalt, chromium, and zinc from various bi-product streams before joining Newmont.

Brian managed metallurgical laboratories, production teams, process design initiatives, metallurgical test programs, metallurgical accounting activities, and long-range process planning activities. Brian has experience with whole ore roasting, SAG milling, rod milling, and flotation, gold leaching, heap leaching, gold refining, and bio-oxidation.

He holds B.S. and M.S. degrees in Metallurgical Engineering from Montana College of Mineral Science and Technology and an MBA from the University of Nevada – Reno. He is also a QP as a registered member of the Society for Mining Metallurgy & Exploration (SME). Mr. Brian Arthur is working as a Principal Associate with GeoGloball LLC for this project.

 

 

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2.2.4

Mr. Paul Gates – Principal Mining Engineer

Mr. Paul Gates has more than 37 years of experience in the mining industry, including extensive experience in field exploration, long-term mine planning, project feasibility and business case analysis, mine development and operations, mine valuation, and mine optimization. He supervised loading and haulage fleets in large open-pit copper mines with crews in excess of 60 operators.

He is skilled at planning, coordinating, and supervising operations at gold, copper, and silver mines. Paul also has consulting experience at uranium, coal, platinum, and iron ore mines and has a solid understanding of permitting and environmental challenges facing today’s mining industry.

Paul holds a B.S. degree in Mining Engineering from Montana College of Mineral Science and Technology and an M.S. in Business Administration degree from Western New Mexico University.

Paul is a Registered Professional Engineer (PE). He is working as a Principal Associate with GeoGloball LLC for this project.

 

2.2.5

Responsibilities of the Qualified Persons

The QPs have supervised the preparation of this report and take responsibility for the contents of the Report as set out in Table 2-1. The QP Certificates can be found in Section 26.

Table 2-1: QPs and Responsibilities for the Contents of the Report

 

   Qualified Persons    Report Sections    Topics
   Abani Samal   

1, 2, 3, 5, 6, 7, 8, 9, 11,

15, 22, 23, 24, 25, 26

   Summary, Introduction, Geology, Resource Estimation etc.
   Paul Gates    1, 2, 11 and 13    Criteria for Reasonable Prospect for Economic Extraction (Optimum Pit shell and Economic Analyses), Resource Tabulation) and Mining Methods
   Brian Arthur    1.3, 4, 10, 14    Mineral Processing and Metallurgy, Site Infrastructures

Sections 12, 16, 18, and 19 are beyond the scope of this project.

The data and information used in sections 4, 17, 20, and 21 were provided by the registrant.

 

2.3

Scope of this Project

It is understood that no exploration or mining activity has happened at Lone Tree Deposit since the above-mentioned NI 43-101 report was published in October, 2021 (2021 report). However, since then three years have passed and i-80 remains as the 100% owner of the Lone Tree deposit. This project provides an opportunity to revisit the existing data and information and cross-check various aspects of existing resource models and

 

 

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estimates including criteria to demonstrate that there are reasonable prospects for economic extraction which may affect mineral resource classification for the purposes of an S-K 1300 Report. The scope of this project is discussed below.

 

  i.

Onsite Verifications: As per the requirement of the S-K 1300 guidelines, the qualified persons (QPs) comprising the lead geologist and process engineer completed the site visit to do the onsite investigation. The following tasks were completed during the site visit.

 

   

Verification of the geology of the deposit, examination of available drill cores, pulps, and rejects acquired by i-80 from NGM.

   

Verification of the drill hole logs and sampling methods.

   

Selection of pulps and drill cores for re-assay.

   

On-site verification of leach pads, and processing facilities.

   

Verification of processes followed in the on-site laboratory facilities.

   

Verification of the quality of data includes QA/QC procedures, checking original assay records, etc.

 

  ii.

Metallurgical Recovery: The Principal Metallurgist & Process engineer completed the site visit and later reviewed various sets of data as listed below.

 

   

Datasets for determining the appropriate metallurgical processing suitable for the remaining mineralized materials.

   

Recovery factors for various material types and process methods.

   

Processing costs for autoclave and flotation options.

 

  iii.

Geological Model Verification: The geological models are to be reviewed for correctness.

 

  iv.

Mineral Resource Estimation: The mineral resource estimation exercise conducted to produce the above-mentioned NI 43-101 report (the effective date of July 31, 2021) was reviewed and updated as required. The following items were considered.

   

Exploration Data Analyses (EDA).

   

Variogram analyses.

   

The density (or tonnage factor).

   

Grade estimation for gold.

   

Resource classification.

   

Development of an optimized pit shell by a qualified mining engineer with updated criteria.

   

Model validation review and update if required.

 

  v.

Resource Estimates and Tabulation: The resource estimates were updated with new pit-optimization parameters. The classifications of the mineral resources were updated accordingly and the mineral resources were tabulated with an effective date of December 31, 2024.

 

 

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2.4

Units of Measure

 

2.4.1

Coordinate Reference System (CRS) Used in the Maps:

Universal Transverse Mercator Zone 11, North American Datum of 1983. EPSG: 26911. For this report, maps were generated using a combination of Google Earth Pro and QGIS version 3.16.0-Hannover mapping tools.

The block model and drill hole coordinates are in the local Lone Tree mine grid system as defined by Newmont/NGM.

 

2.4.2

Conversion Factors Used

 

     i.

1 ounce (oz) per short ton (t) = 34.2857 parts per million (ppm) = 0.0034286 %

 

    ii.

1 part per million (ppm) = 0.0291667 ounce (oz) per short ton (t)

 

   iii.

1 g/T = Opt*31.1035*1.102 (Opt: 1 ounce per short ton)

 

   iv.

Million Tonnes (MT) = short tons (t) /1.102/1000000

 

    v.

K Ozs = Oz/ 1000

 

2.5

Definitions

 

AMSL:

  

Above Mean Seal Level

AOI:

  

Area of Influence

AuFA:

  

Gold (Au) Fire Assay

BLM:

  

United States Bureau of Land Management

BAS:

  

Bureau of Administrative Services

BAQP:

  

Bureau of Air Quality Planning

BAPC:

  

Bureau of Air Pollution Control

BWPC:

  

Bureau of Water Pollution Control

BWQP:

  

Bureau of Water Quality Planning

BSDW:

  

Bureau of Safe Drinking Water

BMRR:

  

Bureau of Mining Regulation and Reclamation

BCA:

  

Bureau of Corrective Actions

BISC:

  

Bureau of Industrial Site Cleanup

BMM:

  

Bureau of Materials Management

BFF:

  

Bureau of Federal Facilities

BSMM:

  

Bureau of Sustainable Materials Management

CFR:

  

Code of Federal Regulations (United States Federal Code)

CIM:

  

The Canadian Institute of Mining, Metallurgy and Petroleum

FA/AA:

  

Fire Assay with Atomic Absorption finish

GPS:

  

Global Positioning System

ICP:

  

Inductively Coupled Plasma (geochemical analytical method)

MT:

  

Million Tonnes

NDEP:

  

Nevada Division of Environmental Protection

NGM:

  

Nevada Gold Mines

NSR:

  

Net Smelter Royalties

 

 

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NMC#:

  

Nevada Mining Claim Number

NDOW:

  

Nevada Department of Wildlife

Opt:

  

ounce (oz) per short ton (t)

PoO:

  

Plan of Operation

RC:

  

Reverse Circulation (Drill hole)

RPEE:

  

Reasonable Prospect for Economic Extraction

SEC:

  

Securities and Exchange Commission

t:

  

Imperial Tons (2,000 pounds)

T:

  

Metric Tonnes (2,200 pounds)

Tpd:

  

tons per day

TRS:

  

Technical Report Summary

USGS:

  

United States Geological Survey

 

3.

PROPERTY DESCRIPTION

 

3.1

Location of the Project

The Lone Tree gold deposit is in the Battle Mountain district in Humboldt County, Nevada. As shown in Figure 1-1, the nearest town with full services is Winnemucca, which is a historic mining town in northwestern Nevada. The Lone Tree Mine is located approximately 30 miles east of Winnemucca, Nevada, and 20 miles northwest of Battle Mountain, Nevada. The mine office is accessible from Interstate 80 by a paved highway (Figure 3-1). The reference coordinates of the mine are 40° 50’ 19” N, 117° 12’ 37” W. The Lone Tree facilities include an autoclave, carbon-in-leach (CIL) mill, flotation plant, and heap leach facility.

 

3.2

Land Status

The Lone Tree Properties include interests in fee lands, mineral rights in fee lands, patented mining claims, and unpatented mining claims which are leased or owned by i-80 as of the effective date of this report. All interests were acquired from Nevada Gold Mines and a full description of land and interests is shown in Figure 3-1. Details of all claims are provided in Appendix A. The information on the Lone Tree/Buffalo Mountain map was prepared by the Land Department of i-80 (refer to Figure 3-1). The areas of unpatented and patented claims are shown in Table 3-1.

Table 3-1: The Area of the Lone Tree Claims

 

Lone Tree Land Position

     Acres     

Lone Tree Unpatented

   6207

Buffalo Unpatented

   892

Private Property Sections

   5589

Total

   12688

It should be noted that The Brooks unpatented claims are encapsulated in the Lone Tree Unpatented acreage, as they are adjacent to the Lone Tree property.

 

 

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Figure 3-1:Location of Lone Tree Property and Land Position

 

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These properties are identified as part of the Lone Tree Mine Plan of Operations (PoO) in Sections 11, 12, 13 and 14.

This report focuses only on the Lone Tree Mine properties. i-80 has been informed by the Clerk of the Eleventh Judicial District Court, Humboldt County, Nevada that there are no pending actions which relate to the Lone Tree Mine properties in which the Company or the Company’s subsidiaries are named as parties.

 

 

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4.

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The Lone Tree Mine project is located immediately adjacent to Interstate Highway 80 (i-80), approximately 35 miles east of Winnemucca, and approximately 12 miles west of Battle Mountain, Nevada. The project office is easily accessible from Interstate 80 as shown in Figure 4-1. From the mine office, the mine and process facilities are accessible via paved roads. The nearest town with full services is Battle Mountain, Nevada.

Figure 4-1:Location and Accessibility of the Lone Tree Mining Project

 

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4.1

Climate

The climate of the Lone Tree mine area is cold and semi-arid typical of eastern Nevada. The temperatures range from highs of upper 40°C in summer to lows of –7°C in winter. The area experiences low annual precipitation of 15 to 20 cm per year and 50% of the precipitation occurs as snowfall during winter. The climate condition presents no restrictions to the operation of the Lone Tree mine and facilities.

 

 

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4.2

Physiography

Elevations at the Lone Tree property range from approximately 1,490 metres above mean sea level (AMSL) to 1,350 metres AMSL for total relief of approximately 140 metres. The historic open pit bottom is at approximately 1,080 metres AMSL. The terrain varies from a relatively flat alluvial plain to sloped foothills in the area.

 

4.3

Vegetation

Vegetation in this area mainly comprises sagebrush, rabbit brush, and a variety of grasses and forbs. Fauna is not abundant on the Property primarily due to the lack of surface water and limited forage. No threatened or endangered plant or animal species have been noted within the Property’s operating area.

 

4.4

Site Infrastructure

The Lone Tree site has an employee parking lot with secure access to the administration building and a security gate for vehicles entering the mine property. All active areas of the mine are accessed by well-maintained gravel and dirt roads. The access roads are engineered at 60-70 feet wide to accommodate haul trucks and mining equipment traffic.

The site has many buildings on the property, including the following facilities.

   

Administration building

 

   

Haul truck maintenance shop and warehouse

 

   

Dewatering shop

 

   

Potable and fire water pump house

 

   

Mill and autoclave building

 

   

Flotation building

 

   

Filter building

 

   

Maintenance shop

 

   

Geology building

 

   

Assay laboratory building

 

   

CIC process building

 

   

Oxygen plant

 

   

Gold refinery

 

   

Two warehouses in the process area

 

4.4.1

Water Wells:

There are three active water wells on site. Well # NWW1 supplies fresh water to the fire water pump house storage tank and the potable water storage tank at an average pumping rate of 5.3 gallons per minute. Well #SS13 supplies water to the truck shop area fire water pump house storage tank at a rate of 14 gallons per minute. Well #SS14 supplies water to the truck shop area fire water pump house storage tank at a rate of 10 gallons per minute.

 

 

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4.4.2

Electric Power

Electrical power is supplied by Nevada Energy via a main transmission line that runs through the property and delivers 20KV to the main site substation. Transformers step the power down to 24.9KV and then to 4160V for further distribution. Additional transformers are in place to step down the power to the level needed in each area of the site.

The main source of heating comes from three propane tanks, each holding 12,000 gallons. Out of three, one is used in the administration building area and two in the mill processing area.

 

4.4.3

Mineral Processing & Metallurgy Facilities

The site has an autoclave and a flotation mill and a leach pad. The autoclave and flotation mill are in care and maintenance. The leach pad is still producing remaining significant amounts of gold and has a remaining capacity of 10,000,000 tons. The list of processing plants includes the facilities as discussed below.

 

4.4.3.1

Mill Facilities

The mill site was once used to process high grade oxide ore and refractory ores from the Lone Tree mine. The mill had two grinding lines, with one line that prepared feed for an autoclave and the other prepared feed for a flotation plant or oxide leach ores. The SAG and ball mills for both lines are intact.

 

   i.

Lone Tree Autoclave (Figure 4-2), which processes high-grade refractory ore.

 

    ii.

Lone Tree Flotation Plant, which processes low-grade refractory ore.

 

    iii.

Oxygen Plant to supply oxygen to the autoclave and nitrogen to flotation.

 

4.4.3.2

Lone Tree Leach Pads:

The leach pads treat oxide ore in a cyanide heap-leach process. The leach pads (Phases 1-4 and Phases 5-7 as shown in Figure 1-2) are still in service and operation. i-80 has recently stacked some oxide ore from Granite Creek and it is now under leach. The carbon columns were located on the leach pad and looked to be in good condition.

 

4.4.3.3

Carbon Stripping Facilities

Carbon stripping facilities were used to recover the gold loaded on carbon from both the CIL and CIC circuits. Currently, i-80 sends loaded carbon from CIC off-site for stripping.

 

 

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Figure 4-2:Autoclave Agitator Access Ports

 

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Figure 4-3:Pulverizing and Preparation Stations

 

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4.4.4

Laboratory Facilities:

The laboratory facilities are fully functional. The laboratory facility includes the areas for sample receiving, sample preparation area, weighing area, fire assay, wet chemistry area, LECO, and ICP areas. The sample pulverizing machine and preparation area are shown in Figure 4-3. The LECO and ICP labs are shown in the figures (Figure 4-4) and Figure 4-5 respectively.

Figure 4-4:The LECO Facility

 

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Figure 4-5:ICP Laboratory Facility at Lone Tree In the early days, the Lone Tree area was explored for copper, but no significant resources were discovered.

 

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5.

HISTORY

 

5.1

Historical Summary of the District

The initial discovery hole at Lone Tree was drilled in July 1989 by Cordex Exploration Co. on the southern extension of what was to become the Lone Tree gold deposit. This southern portion of the deposit was referred to as the Stonehouse deposit. Santa Fe Pacific Gold discovered the main part of the Lone Tree deposit in the pediment on the west flank of the hill in 1989 and acquired the Stonehouse portion of the deposit from Cordex. Sante Fe Gold began producing gold from the site in 1991. Newmont acquired the deposit from Santa Fe Pacific Gold through a merger and began operations in 1997. Newmont completed mining operations in 2006.

Operations were discontinued in 2006 due to increased production costs, largely resulting from the influx of groundwater into the deepening pit. The pit was allowed to flood and create a lake within the pit. Approximately 4.6 million ounces of gold were produced from the Lone Tree Mine and approximately 5.2 million ounces of gold were produced at the Lone Tree processing facilities during this time.

Mining on the Brooks deposit which lies to the southwest of the main Lone Tree pit was conducted from 2015 to 2019. Approximately 52,000 ounces were placed on the heap leach pad and residual leaching is continuing.

Residual leaching and ongoing reclamation activities from the Lone Tree Mine have continued since 2007. In July 2019, the non-operating Lone Tree project became part of Nevada Gold Mines, a joint venture between Barrick and Newmont.

i-80 Gold Corp acquired the Lone Tree property and processing facilities from NGM on October 14, 2021.

 

6.

GEOLOGICAL SETTING, MINERALIZATION & DEPOSIT

 

6.1

Regional Geology

The Lone Tree deposit occurs in Humboldt County, Nevada, within the Basin and Range physiographic province, in the northern part of the Battle Mountain mining district. The Battle Mountain mining district is dominated by Late Cretaceous and Eocene age magmatism with various ore deposit types including porphyry Cu-Au, porphyry Mo, skarn, and distal disseminated +/- Carlin-type deposits. Holley et al 2019, list several Cu-Mo porphyry along with sedimentary rock-hosted gold deposits, such as Lone Tree, Buffalo Valley, Marigold, North Peak, and Trenton Canyon, which have been classified as distal disseminated and Carlin-type deposits (Doebrich and Theodore, 1996, Theodore, 1998, 2000, Reid et al. 2010). The location of the Lone Tree Mine Relative to Major Mineral Trends in Nevada are shown in Figure 6-1 (modified from Wallace et al. 2004 and Fithian

 

 

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et al. 2018). The location of the Lone Tree deposit is shown along with other deposits in the Battle Mountain trend in Figure 6-2 (Source: NGM); and Au-skarn deposits, such as those at Buckingham, Copper Canyon, Copper Basin, and Elder Creek.

Figure 6-1: Location of the Lone Tree Mine Relative to Major Mineral Trends in Nevada.

 

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Ag:Au ratios are consistent with those in most other Carlin-type deposits, although the lower ratios of some ores are similar to the distal-disseminated Au-Ag deposits such as Lone Tree, Nevada. (Ressel, 2005).

The low Ag:Au ratios and lack of base metals have been used to differentiate Carlin-type Deposits from other sedimentary rock-hosted deposits in northern Nevada such as Lone Tree, which are classified as pluton-related or distal-disseminated Ag- Au (e.g., Cox, 1992; Mosier et al., 1992; Doebrich and Theodore, 1998, Wallace et al 2004).

 

 

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Figure 6-2: Location of Lone Tree Deposit in the Battle Mountain Mining District

 

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Wallace et al (2004) provide a detailed account of the regional tectonic activities in northern Nevada, occurring over a period of 2 billion years starting with Precambrian rocks occurring in the East Humboldt. Paleozoic rocks in this region generally comprise the four distinct tectonostratigraphic assemblages described below (Source: Holley, 2019).

 

   

Cambrian-Ordovician miogeoclinal carbonate shelf-slope rocks identified through deep drilling in the district but not exposed at the surface (Fithian et al., 2018).

 

 

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Ordovician-Mississippian eugeoclinal siliciclastic rocks of the Roberts Mountain allochthon, including the Valmy Formation.

 

   

Autochthonous Pennsylvanian to Permian shallow-water facies of the Antler overlap sequence.

 

   

Mississippian to Permian deep-water siliciclastic rocks and basalts of the Golconda allochthon, which were thrust on top of the Antler overlap sequence by the Golconda thrust during the Permian-Triassic Sonoma orogeny (Theodore, 2000), constituting the Havallah sequence many of the clastic constituents of these rocks appear to be sourced from the Antler highlands (Whiteford, 1990).

Gold deposits are hosted in a variable stratigraphic package of Ordovician through lower Mississippian shallow-water rocks that have been overthrust by deep-water, siliciclastic allochthonous rocks along the Roberts Mountains Thrust during the late Devonian to Early Mississippian Antler orogeny (Roberts et al., 1958, Roberts 1960). Subsequent orogenic shortening during the Pennsylvanian and Permian (Humboldt disturbance) (Ketner, 1977), Early Triassic (Sonoma orogeny) (Silberling and Roberts, 1962), Middle Jurassic (Elko orogeny) (Thorman et al, 1990) and Early Cretaceous (Sevier orogeny) (Armstrong, 1968) have reactivated earlier basement and Antler-related faults. The sedimentary rocks are intruded or unconformably overlain by igneous rocks of three magmatic episodes: Cretaceous, Eocene, and Miocene age.

The current regional physiography is the result of extensional tectonics during the Tertiary. High angle faults formed during this period are interpreted as the main pathways for ore forming fluids. Economic concentrations of gold typically occur near the intersections of northeast and north-south faults, along the margins of intrusive bodies, or at contacts between siliceous and carbonate lithologies. Geochemical enrichment in trace elements such as silver, arsenic, antimony, mercury, and thallium are common to nearly all trend deposits.

 

6.2

Deposit Geology and Mineralization

Mineralization is hosted within structures which crosscut all three Paleozoic rock sequences present in the mine area. The oldest of these three sequences is the Ordovician Valmy Formation, which is a part of the Roberts Mountain Allochthon.

In the mine area, the Valmy consists primarily of quartzite, with lesser amounts of chert, argillite, and minor basalt. The Valmy rocks are unconformably overlain by rocks of the Pennsylvanian Antler Sequence, which belong to the Battle and Edna Mountain Formations. A geological map of the Lone Tree mine along with vertical Cross Section are presented in Figure 6-4. The Edna Mountain Formation at Lone Tree is typified by a sandy siltstone unit grading downward into a lithic sandstone unit. The Battle Formation is observed as a poorly sorted cobble conglomerate of varying thickness. A thin calcareous sandstone tentatively identified as a lateral equivalent of the Antler Formation rocks present at the Marigold Mine has been encountered in drill holes on the southeastern margin of the mine area. Rocks of the Pennsylvanian-Permian Havallah sequence were

 

 

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thrust over the Antler Sequence rocks in the mine area during the Sonoma Orogeny. The Havallah Sequence at Lone Tree encompasses several rock types within at least three packages, but is dominated by siltstones, chert, and basalts with lesser sandstones and conglomerates.

Figure 6-3:General Stratigraphic Sequence of Lone Tree Figure 6-4:Local Geology of Lone Tree Deposit and Controls of MIneralizaton

 

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Although gold mineralization is present in all three Paleozoic sequences, Antler Sequence rocks appear to have been preferentially mineralized within the structural zones. Alluvial cover over the deposit ranges from a minimum of two feet to a maximum in excess of 400 feet. Bedrock has been sharply down-dropped to the north and to the southeast by post-mineral faulting, creating alluvium-filled basins in excess of 1,000 feet deep. See Figure 6-3 for the local stratigraphic interpretation Mine (K.C Raabe, 1995. The Lone Tree Extension Project, Humboldt County, Nevada).

 

6.3

Controls of Mineralization

Gold mineralization at Lone Tree is primarily controlled by structure as seen in Figure 6-4. This figure was developed by modifying the figures developed by NGM staff and the figures shown in Holley et al. (2019). Three principal mineralized structural zones and at least one lesser zone are currently recognized. The three principal structural zones are known as the Wayne Zone, the Sequoia Zone, and the Antler High Zone. The Wayne Zone is the most significant of the three major zones in terms of strike length, mineralized tons, and contained ounces. The Wayne Zone encompasses more than 50 percent of the contained tons and ounces within the overall deposit. The most widely recognized of the lesser zones is known as the Chaotic Zone, aptly named for the structural complexity associated with it.

The Wayne Zone has been described as a system of relatively narrow north-northwest and north-northeast trending faults forming an anastomosing complex of brittle shears enveloping rhomboid blocks of relatively competent but highly fractured domains of lesser strain (Bloomstein et al, 1992). With few exceptions, ore-grade mineralization does not extend along the north-northeast and north-northwest faults beyond the margins of the Wayne Zone. Detailed examination of blast hole data clearly demonstrates a “zig-zag” pattern of mineralization within the principal component structure of the Wayne Zone, known as the Powerline Fault. Higher gold grades within the Powerline Fault are commonly associated with the hanging wall and footwall margins of the fault, which averages 50 feet in width.

The Powerline fault zone is a North - South trending high angle fault zone, extending at least 2,500 m along strike. Mineralization is truncated to the north by the NE trending Poplar Fault. Mineralization in the Wayne Zone is hosted in all 3 rock packages (Valmy, Antler, Havallah) as breccia within the complex structure.

The southern zones of mineralization (Sequoia, Antler High zones) are primarily hosted in the Edna Mountain Formation of the Antler sequence. This mineralization is a combination of structural (Sequoia Fault) and stratiform control. Unlike the Carlin Trend where gold is primarily hosted in arsenian pyrite, at Lone Tree gold is primarily hosted in both arsenopyrite and arsenian pyrite. Lone Tree has always been considered a horst block cored by the Valmy Formation. Siliciclastic sediments with the Powerline Fault on the west side and Sequoia Fault on east side are the main controls to mineralization (Figure 6-4).

 

 

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The Wayne Zone itself ranges in width from 150 to 300 feet. It trends north-south, dips to the west at an average of 65 degrees and has a drill-defined strike length in excess of 9,000 feet. The northern half of the zone essentially bounds Lone Tree Hill on the west. Drilling has encountered Wayne Zone mineralization down-dip nearly 1,000 feet from the original ground surface sub-economic mineralization remains open below this depth.

The Sequoia Zone is located to the southeast of Lone Tree Hill and essentially describes the southeastern margin of the deposit. The structural fabric of the Sequoia Zone is quite similar to that of the Wayne Zone, with some important differences. The known strike length of the Sequoia Zone is 2,000 feet, significantly less than that of the Wayne Zone.

The overall dip of the Sequoia Zone is 75 degrees, as opposed to the 65 degrees of the Wayne Zone. Post-mineral faulting has displaced or cut out mineralization within the Sequoia Zone and has had a significant effect on the continuity of mineralization, both down-dip and along strike.

The Antler High Zone is located within a horst block of Antler sequence rocks between the Wayne Zone and Sequoia Zone and is limited to the southern third of the deposit. The Antler High mineralization is primarily developed within rocks of the Edna Mountain, Battle, and underlying Valmy Formations, and commonly appears parallel or sub-parallel to bedding. Evidence suggests that the Antler High gold mineralization is hosted within a dense network of very narrow fractures similar to a stockwork. Several high-angle mineralized structures are known to cut through the Antler High and may have served as feeder structures. The trends and dip angles of these latter structures are sub-parallel to those of the Wayne Zone and Sequoia Zone. Along the northern margin of the Antler High, several of these structures trend upward through the Golconda thrust and into the overlying highly siliceous Havallah rocks. The combination of these specific high-angle structures and local, lower-angle mineralized structures is known as the Chaotic Zone. In addition to the high-angle structural control in the Antler High, a low-angle (45 degrees east) west-vergent compressional feature known as the Redwood Fault controls a substantial portion of the mineralization in that zone. The Redwood Fault effectively doubles the thickness of the Edna Mountain host rocks within the Antler High.

Mineralization is hosted both within the fault plane itself, and within the highly shattered rocks of the adjacent hanging wall block. The age of the Redwood Fault is not known, but certain evidence suggests that it pre-dates the Sonoma Orogeny.

Mineralized structures have been identified in the hanging wall of the Wayne Zone, and within the footwall of both the Wayne Zone and the Sequoia Zone. Many structures controlling gold mineralization are moderate to high angle, west- or east-dipping normal faults or fractures. Some lower-angle mineralized structures, which are thought to have been re-activated during extension, have been noted. As within the Wayne Zone, mineralization most often occurs at the intersection of NNW and NNE-trending faults of varying dip angles. Strike-slip or oblique-slip motion has been noted on some structures, although kinematic indicators are essentially non-existent in the highly silicified, brittle rocks of the Edna Mountain Formation, or in the Valmy quartzite.

 

 

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A principal characteristic of the Lone Tree deposit is the spatial coincidence of several structurally controlled episodes of mineralization. Hydrothermal breccias, with as much as 25% matrix expansion, host a significant portion of the gold mineralization. High-grade ore occurs at fault or fracture intersections, or at jogs in the faults, which form dilatant zones.

Silicified, multiple-phase breccias have been noted along the margins of the principal mineralized zones. These appear to be early, and in general, are lower in grade. Later tectonic breccias were superposed on the hydrothermal breccias. The most recent structures tend to be milled-breccia post-mineral faults and shears, which often possess >50% clay gouge, and display a crude lamination produced by streaks of iron oxide, pyrite, or angular clasts. Reactivation of high-angle faults is demonstrated by barren, vuggy silica-cemented structures overprinting similarly oriented mineralized zones.

Mineralization is also known to occur in crackle breccias within the more brittle rocks of the Edna Mountain and Valmy Formations, which are crosscut by the Wayne Zone. Zones of intense micro-fracturing noted in the highly silicified Edna Mountain rocks are the closest approximation to “classic” disseminated mineralization yet noted at Lone Tree.

Numerous cross-structures have been identified at Lone Tree. Significant gold mineralization has not been observed in association with any of these structures. The Wayne Zone is cut on the north by a major northeast-trending fault zone known as the Poplar Fault Zone. While the Wayne Zone as a structural zone does not appear to be terminated by the Poplar Fault zone, the down drop of the bedrock surface, thinning of the mineralized faults, and decreased grade all currently limit the economic potential of the Wayne Zone north of the Poplar. Other northeast-trending faults, such as the Willow Fault, have significant effects on the mineralization even though they do not offset the Wayne Zone.

A west-northwest-trending zone of southerly dipping normal faults known as the Pinon Fault zone truncates Lone Tree Hill to the south and is associated with a change in the strike direction of the Wayne Zone at that location. At the extreme southern end of the known mineralization, the Wayne Zone and Sequoia Fault converge. Drilling has identified at least one major northeast-trending structural zone in this area which appears to have some effect on mineralization.

As a result of the fact that the Lone Tree deposit occurs at the margin of a bedrock block essentially surrounded by alluvium, the relationship of the deposit to regional structure is not well understood. It has been speculated that the deposit may have formed in response to strike-slip and normal faulting related to regional wrench faulting. An alternate hypothesis suggests that the faults which control and host mineralization at Lone Tree may be dominantly extensional in nature, with little relationship to strike-slip and wrench faults. The age of the mineralization is constrained to the Eocene based on dating of mineralized intrusive rocks (Holley et al., 2019).

 

 

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It’s clear many of the faults were active at the time of mineralization, although their age of first movement is uncertain.

 

6.4

Alteration

The principal alteration process associated with gold mineralization at Lone Tree is potassic alteration (Bloomstein et al., 1992). Other alteration types noted in the mine area are argillization, silicification, propylitization, and skarnification. A general progression from oxidized argillic alteration in the Havallah sediments down into unoxidized argillization, silicification and potassic alteration in the Antler and Valmy rocks has been noted. Alteration assemblages are commonly mixed within the fault zones as a result of the structural control of mineralization. Pervasive pre-mineral silicification is common in portions of the Havallah Sequence, and throughout most of the Antler Sequence rocks at Lone Tree Mine (K.C Raabe, 1995. The Lone Tree Extension Project, Humboldt County, Nevada).

 

6.5

Gold Mineralogy

Gold mineralization occurs as sub-micron sized inclusions within a distinct generation of very fine-grained pyrite and arsenopyrite in the sulfide zone. Evidence gathered to date suggests that the main gold deposition event occurred in a temperature range of 200o to 450o (epithermal to mesothermal). The ore mineralogy shows evidence of two overprinted assemblages reflecting at least two hydrothermal episodes at Lone Tree. Partial oxidation of the main stage mineralization occurred prior to a later, epithermal event characterized by open-space filling textures and weakly auriferous pyrite and marcasite. In the oxidized portions of the deposit, and particularly in the Havallah rocks, gold occurs as micron-sized particles in goethite and limonite. Post-mineral oxidation extends as much as 700 feet down major structures such as the Wayne Zone. No supergene effects or gold remobilization have been proven or documented at the Lone Tree Mine (K.C. Raabe, 1995). The Lone Tree Extension Project, Humboldt County, Nevada).

 

6.6

Deposit Type

The Lone Tree deposit is characterized as a pluton-related or distal-disseminated Ag- Au deposit. This deposit type is discussed by Wallace et al., 2004 and Munteen and Cline, 2018).

As discussed by Peters et al. (in Wallace et al. 2004) the Lone Tree deposit among others in the Battle Mountain district appears to be related genetically to porphyry systems, even though many deposits do not contain obvious near-surface features that would indicate this connection, mainly because the gold-silver mineralization in these deposits may be over one km away from the causative intrusions. This is why the deposit has been characterized as “distal disseminated”. Due to complex tectonic and extension in the region, the mineralization in these deposit types may have substantially different geometric relations to the intrusive centers and hosted in different stratigraphic horizons as shown in the Figure 6-5 (Wallace et al., 2004). The mineralization at Lone Tree occurs

 

 

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in intensely fractured three stratigraphic horizons. These horizons are similar to but significantly different to the horizons found in other deposits in the region. Gold is associated with a low Ag:Au ratio (<2:1), As, Sb, Hg and Tl as well as elevated Bi, Mo and W. Gold is hosted in arsenopyrite indicating higher temperatures of ore formation in comparison to typical Carlin-type deposits where gold is hosted in arsenian pyrite.

Figure 6-5:Diagrammatic Model of Geology of Distal-Disseminated Ag-Au Deposit.

 

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7.

EXPLORATION

 

7.1

Exploration History

The exploration history of the Lone Tree deposit is documented by Bloomstein et al. 1993. Part of this section is extracted from that publication. Prospecting around Lone Tree Hill is believed to have started in the middle 1860’s when the construction of the Central Pacific portion of the Transcontinental Railroad started about three kilometers northeast of Lone Tree. Sporadic exploration activities continued for copper and gold without much success until Duval Corp and Bear Creek explored the area in 1960’s and 1970’s for porphyry copper. These exploration activities aided in the discovery of low-grade gold mineralization in the area.

Exploration activities in the 1980s by Nerco, Freeport and several Canadian junior companies yielded intercepts of narrow, fracture filled gold mineralization. In 1989 Cordex Exploration and Santa Fe Mining formed a joint venture for exploration of the Lone Tree deposit resulting in a discovery of substantial gold mineralization about one km from Lone Tree Hill.

 

 

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Subsequently, 12 additional holes were drilled and a north-south fault system controlling mineralization was discovered. The first gold was poured in 1991.

Later, Newmont acquired the deposit from Santa Fe Pacific Gold through a merger and assumed operations in 1997. Extensive surface mapping, sub-surface drill hole modeling, geochemical sampling, geophysics, and drilling resulted in the discovery of seven separate ore bodies of similar structural and lithologic genesis (Zacarias, P. A., 2006). All but one deposit located to the south of the active mine were actively mined. Newmont completed mining operations in 2006. Residual leaching continued through 2019. Reclamation began in 2007. In July 2019 the non-operating Lone Tree project became part of Nevada Gold Mines, a joint venture between Barrick and Newmont.

GeoGlobal is aware that various exploration activities were completed in this area. Even though geophysical and other exploration activities such as trenching, pitting, and trial mining were conducted in this area, supporting documents were not available to verify.

 

7.1.1

Drilling and Sampling

Between 1980 and 2015 a total of 1,904 drill holes, summarized in Table 7-1, were completed in and around the Lone Tree mine.

Table 7-1:Summary of Drilling by Hole Type

 

Hole Type    Number Drill holes        Total Footage       

Unknown

   241    197,561

CORE

   108    66,263

CORE;RC

   176    139,912

RC

   1,379    865,613

Out of these holes, 1,840 holes were selected for the resource estimation which are close to the Lone Tree mine (Figure 7-1). The 1,840 holes had 2,19,214 assay data, 113,862 lithological logs. Out of 1840 holes 362 holes had no lithological logs.

 

7.1.2

Recent Exploration Drilling

In 2020, a drill hole (LTE-20001) was drilled on the west side of the mine which tested for the existence of the Comus Formation below the Lone Tree Mine. The Comus Formation is significant because it is the host rock for the Turquoise Ridge, Twin Creeks, and Granite Creek mines. The drill hole intercept is shown in the Figure 7-2. Four zones of mineralization were encountered as listed below.

 

 

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Figure 7-1:Location of Selected 1840 Drill Holes for Resource Estimation

 

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  i.

Upper zone of 10.7m @ 4.49 g/t above a QFP (quartz–feldspar porphyry) dike on the contact of the Permian Havallah Formation (P and Edna Mtn Formation. The Upper zones of mineralization are consistent with stratiform mineralization identified in wide-spaced drilling through the hanging wall to the Powerline Fault; this zone is open in three directions.

  ii.

Zone along the contact of Edna sandstone and Valmy quartzite (Phv) (7.6m @ 6.04 g/t including 1.5m @ 13.5 g/t).

  iii.

Zone of sulfide breccia in Valmy quartzite (38.1m @ 2.15 g/t w/ grades up to 18.95 g/t Au).

  iv.

Lower zone of mineralization hosted within a QFP (Quartz Feldspar porphyry) dike with sooty pyrite on fractures and in the groundmass of the intrusive (40.3m @ 1.22 g/t).

The Lower Plate Ord. Comus Formation was intercepted at 1155 m (3790’). The Comus is characterized by strong calc-silicate hornfels intruded by fine grained diabase sills.

 

 

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A narrow zone of mineralization was encountered down-dip on the Powerline Fault in the Comus Formation (3.0m @ 1.84 g/t). Additional drilling is warranted to vector from the strong calc-silicate alteration to intersect ore controlling structures in more reactive host rocks.

Figure 7-2:Exploration Drill Hole LTE-20001 (Source: NGM)

 

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7.1.3

Rotary Drilling

Historic drilling, including rotary drilling, started in 1980. Samples were typically collected at the drill site after traversing through a rotary wet splitter attached to the return air hose. Most splitters allow for sample size changes by blocking some of the internal rotating vane chambers, thus causing sample material excess to be discarded.

 

 

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The normal sample interval is every five feet, with dry sample weights ranging from 5 to 20 pounds.

Rotary air samples are normally produced by either a down hole percussion hammer bit or a rotary tricone roller bit, with the sample traversing from the bit face up the annulus between the bit and sub or hammer assembly, then into an opening into the drill pipe (“interchange”) center tube and then up to the surface. In the past ten years more use has been made of drill bits that direct the sample into the center tube through an opening in the drill bit face.

Typically, the sample bag (13” by 26” Tyvek 1680 series porous fabric) is clamped onto the splitter outlet. Note that early (circa mid 1980’s) rotary air sampling may have been accomplished in dry conditions using non-porous plastic bags.

The rotary drilling technique includes clearing the bottom of the hole after every rod change and before the next sample chips are collected and washing the splitter if any material is noted sticking to the sampling surfaces. Some early rotary holes were drilled using the conventional air circulation method wherein the sample returned in the annulus between the drill pipe and rock.

Rotary mud drilling includes conventional water-based mud systems in which the sample chips return up the annulus between the drill string and the rock suspended in a ‘mud’ solution. At the surface, the liquid either runs through a settling trough, and the chips manually scooped out of the trough into bags or is directed over a vibrating screen which allows the fluid to fall drain off while the chips progress into a random vane stationary (‘pinball’) splitter and then into sample bags.

 

7.1.4

Reverse Circulation Drilling

The Lone Tree Complex followed a standard procedure for Reverse Circulation (RC) drilling executed by a drilling contractor (Zacarias, 2006).

 

  1.

Samples are collected by the drill contractors through a rotating splitter attached to the drill rig by the drilling contractor.

 

  2.

Samples are collected in five-foot intervals and chip trays are simultaneously filled for later geologic interpretation by the drilling contractor.

 

  3.

Nominal sample weight is between 8 and 12 pounds as collected by the drilling contractor

 

  4.

Samples are collected in micro-pore bags to minimize loss of the fine fraction of sample. These bags are provided to the drill contractor by the Newmont drill services department. Bags are tagged with a bar code to track status and for ease of processing and marked with the hole number and sample footage interval for the lab and the project geologist by the drilling contractor.

 

 

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  5.

Problems with sample contamination in the rotating splitter (cyclone) are minimized by the strict practice of cleaning the inside of the cyclone regularly by the drilling contractor.

 

  6.

All drilling problems, including lost circulation, poor sample recovery, and high water flow are discussed with the project geologist and personnel from drilling services and remedied, if possible, by the drilling contractor.

 

  7.

Samples are prepared for shipment to the assay lab and placed in multi-sample bins by the drilling contractor.

 

  8.

The geologist consults historical data and elects an assay procedure that is appropriate for the style of mineralization (e.g., whether there is a coarse gold issue or “nugget”, and what is the nature of the gold mineralization and gold digestion techniques) in the Lone Tree Complex geology.

 

  9.

The geologist completes the sample submittal with all necessary analytical requests, and assay packages and submits quality-check standards (blanks) by the Lone Tree Complex geology.

 

  10.

The geologist notifies the accredited assay lab to request a sample pick-up.

 

  11.

Assay results are relayed to the database department and the project geologist upon completion.

 

  12.

Sample pulps and coarse rejects are temporarily stored at the assay lab and then returned for storage at the Twin Creeks warehouse or the Winnemucca hangar-Independent assay lab by Newmont Drill Services.

 

  13.

Significant drill intercepts or intercepts that appear anomalously low are often reanalyzed at a different lab as a quality control and verification measure as determined by Lone Tree Complex Geology personnel.

 

  14.

Hard copies of the assay results are filed with the completed geology log for the respective hole in the geology logging facility at the Lone Tree offices by the Lone Tree Complex geology.

 

  15.

Assay data are computerized and available for extraction by Database management.

 

7.1.5

Core Drilling

The following procedure pertains to core drilling and sampling at the Lone Tree Complex (Zacarias, 2006).

 

  1.

The core is cut by the contractor by a diamond bit in 5 to 10-foot runs. The standard diameter for exploration drilling is HQ, 2 3⁄4-inch diameter.

 

  2.

Samples are laid in boxes containing approximately 10-foot capacities by the drilling contractor.

 

  3.

Records are maintained concerning core recovery, run length, core loss, rig time and hole conditioning, and drilling contractor.

 

  4.

Blocks are placed in the boxes which mark the end of a core run and record the length of the run and the length of the core recovered by the drilling contractor.

 

  5.

All drilling problems, including lost circulation, poor sample recovery, and high water flow are discussed with the project geologist and Drilling Services and remedied, if possible, by the drilling contractor.

 

 

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  6.

Any core loss is treated as serious and the proper remedies including fluid modification are implemented by the contractors and the drill services representative by the Drilling contractor and Newmont drill services.

 

  7.

Boxes are stacked when filled and taken by geology to the logging facility by Lone Tree Complex geology.

 

  8.

Core is washed (minimally) and logged for detailed geologic interpretation. Geotechnical and geology logging is done at the same time. Core loss is noted on the log by Lone Tree Complex geologists.

 

  9.

Sample intervals are marked out in the core boxes with aluminum tags for later core cutting/sampling. Sample breaks are based on the geologist’s interpretation and lithology/structure/alteration contacts. In general samples in homogenous intervals are nominally 5 feet in length by Lone Tree Complex geology.

 

  10.

The geologist consults historical data and elects an assay procedure that is appropriate for the style of mineralization (e.g., whether there is a coarse gold issue or “nugget”, and what is the nature of the gold mineralization and gold digestion techniques) by Lone Tree Complex Geology.

 

  11.

The geologist completes the sample submittal with all necessary analytical requests, assay packages and submits quality-check standards (blanks) by Lone Tree Complex Geology team.

 

  12.

The core is picked up by the drill services group and taken to Twin Creeks Mine for cutting and shipment to the assay lab. It is standard procedure to saw the core in half lengthwise and send half to the accredited assay lab and store half in the Twin Creeks warehouse. The geologist can request that the core be cut down a specific “cut line” marked and denoted on the piece of core but this is rare. Whole core (as in the 2003 program) has been sent for assay without cutting in areas where sample integrity must be ensured by Drill Services.

 

  13.

Metallurgical/petrographic/geochemical/density testing may occur at this stage depending on the maturity of the project as determined by the Lone Tree Complex geology team, or by the One Tree Process group.

 

  14.

The remaining half of the core is stored in the Twin Creeks warehouse or a company-rented hangar in Winnemucca by Drilling Services.

 

  15.

Sample pulps and coarse rejects are temporarily stored at the assay lab and then returned for storage at the Twin Creeks warehouse or the Winnemucca hangar by Drill services.

 

  16.

Assay results are relayed to the project geologist and the database manager, and a hard copy of the results is filed with the geologic log in the geology logging facilities at the Lone Tree offices by Lone Tree Complex Geology personnel.

 

  17.

Significant drill intercepts or intercepts that appear anomalously low may are often reanalyzed at a different lab as a quality control and verification measure by Lone Tree Complex Geology personnel and an independent assay lab.

 

  18.

Assay data is entered into an electronic database using computers and made available for extraction and geologic modeling, database management and Lone Tree Complex geology team to proceed to the data quality control and validation flowsheet.

 

 

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Once core is collected, the footage blocks and cut list are checked for accuracy. The core is then laid out, washed, and logged for lithology, formation, alteration, mineralization, and structural measurements on a standardized Lone Tree Complex log form. Samples are then selected based on geologic changes or approximately every 5 feet in geologically homogenous rock. Samples are marked with aluminum tags. Core was then photographed and processed.

 

7.2

Collar Surveys/Locations

Collar grid coordinates have been determined by optical surveys (1960’s through late 1980’s), field estimates, Brunton compass and pacing, compass, and string distance, and most recently the use of laser survey or global positioning system measurements. Modern hole locations were transferred electronically to the database and loaded using automated data programs. Hole locations were field checked by geologists and support staff, then plotted on maps, and visually checked for reasonableness in the database.

Drills were oriented on-site using a foresight and backsight set of survey stakes. Normally these stakes are placed by the geologists using a compass to determine orientation.

Prior to the next higher-level study, additional work will be required to better understand the quality and completeness of the drill hole database.

.

 

7.3

Down-Hole Surveys

Determination of the hole trace was accomplished historically by projection of the initial collar orientation, using a down-hole single-shot or multi-shot film camera.

The most recent downhole survey practice includes the use of gyroscopic surveys, the results of which are automatically loaded to the drill hole database using a direct import function. Gyroscopic surveys are normally reported at 25-foot intervals. Readings are taken with reference to true north (adjustments for declination are made on-site). Magnetic interference is not generally a problem for most of the drill sites in Nevada. Care is taken to reduce the effects of nearby metal objects when compasses are used for survey tool orientation.

Standard procedure at Lone Tree was to perform a downhole survey on all holes greater than 300 feet in length. In some cases (e.g., important angle holes) shorter holes are surveyed as well. An independent contractor performs the survey. The azimuth of the drilled hole is determined using a correction from magnetic north to true north with a standard Brunton pocket transit/compass. The angle correction used for 2003 was 14.5 degrees west of magnetic north as read on the compass. This correction was standard for the contractors and the geologist lining up the drill rig. The downhole survey is done by lowering a gyro through the intact drilling steel and measuring the deviation of the original angle and the variance of the original azimuth. The survey data were recorded, and the geologist received a hard (paper) copy immediately after the survey.

 

 

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An electronic copy of the data was sent to Newmont database managers for inclusion in the database. Possible errors were screened by the geologist and the database managers at this stage before the data became final.

 

7.4

Opinion of the Qualified Person

The drilling information available for this resource estimation is adequate for estimating mineral resources for this report.

 

8.

SAMPLE PREPARATION, SECURITY, AND ANALYSES

 

8.1

Sampling

Sampling methodology and security are discussed in Section 7.1.1 as part of the drilling procedures practiced by Newmont Mining Company for RC and core drilling programs.

 

8.2

Sample Preparation and Analysis

Exploration drill holes were assayed at various accredited laboratories throughout the life of the Lone Tree Mine. The most commonly used internal labs include the internal company labs of Newmont, Santa Fe, and Battle Mountain. The Chemex (now “ALS Chemex”) was the most used commercial laboratory.

Sample preparation occurs at analytical laboratories, and techniques vary depending upon the laboratory and the type of analysis to be performed. Two methods are commonly used to perform gold assays. The first is crushing the entire sample, pulverizing a sample split to minus 100 to 200 mesh, subjecting a 5 to 30-gram split of the pulp to acid or cyanide, and taking readings using an atomic absorption machine. The second method, a fire assay, is to pulp the sample, add a lead litharge charge, and fire the sample in a furnace (“Fire assay”). The resulting metal bead containing gold is then weighed and dissolved in acid for analysis.

In general, fire assays with an atomic absorption (AA) or gravimetric finish were standard using 1-assay ton samples. Fire assay methods account for 99.97% of the ‘best assays’ reported in the NGM database. Multi-element ICP geochemical analyses were common but not run on every sample. All gold assay certificates and geochemical reports were copied and filed with the geologic logs. These logs are available for review in the geology logging facilities at the Lone Tree offices.

Multi-elemental analysis contained in the source database includes ICP and wet geochemistry multi-element suites analyzed by commercial laboratories, consisting of several elements determined from one sample, and XRD/XRF semi-quantitative X-ray determinations. Most X-ray analyses were accomplished in-house by the Newmont Metallurgical Services Department.

 

 

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8.3

Data Security

Newmont implemented the use of an acQuire database in 2002 to store all drilling related data including assays. The database is secured by Oracle permissions, user ODBC connections across a Novell Network, and user license permissions and is maintained by designated database managers.

The Newmont Laboratory at Gold Quarry was electronically connected to the acQuire database, and an automated process transferred data every two hours. Data from the Lone Tree lab (rare) is loaded using the acQuire data input forms.

Outside lab data, primarily from ALS Chemex, was loaded using an acQuire direct import protocol. The import program also generates quality control reports for standards and check samples. Data was normally downloaded from a secure ALS Chemex web site. Access to the site was restricted to three Newmont Nevada employees via a username/password scheme. The ALS Chemex internal QA samples and results are available to Newmont Data staff. Regular audits were conducted by ALS Chemex at the request of Newmont.

Survey data was loaded using emailed survey certificates. Sample intervals are electronically created via an automated form at the Newmont sample prep facilities. These intervals update the acQuire Sample table, and contain the sample ID, footages, and sample types.

Collar creation is accomplished using form inputs. Collar creation for surface holes is restricted to Newmont data staff. The coordinates and depths are left blank until an (normally) electronic survey is sent via email or placed on the network. Depths are taken from the Geologists email, the Drill cost report, from the last assay interval, or driller’s logs.

As a result of the loss of paper copies due to rodent infestation in the storage facility, starting in 2005 the certificates from Chemex have been sent in the form of non-editable, digitally signed, PDF files. These are archived on the network. No certificates are, or have ever been, available from the internal Newmont labs, nor is QA data generally shared.

Data extractions are accomplished either using the acQuire software interface, or by use of an in-house program. Extractions are normally done by one of the two database administrators.

 

8.4

QA/QC Procedures

Internal check assays are performed at all labs. Pulps are retained for all assays where pulps are returned by the lab. Either pulps or coarse rejects can be re-assayed.

 

 

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8.4.1

Standards

A combination of in-house Standard Reference Material (SRM) and commercially prepared SRM’s were used to control assay accuracy. In-house SRMs have been developed over many years, mainly from gold deposits on the Carlin Trend. Commercial SRMs were obtained from Geostats Pty Ltd in Australia. SRMs represent all grade bins very high-grade, high-grade, medium-grade, and low-grade gold in oxide and refractory mineralization. Values have been established for the in-house SRMs for gold assays only, using round robin analysis. Earlier Standard reference materials (SRMs) were submitted at a nominal frequency of one every 60 meters (200 feet), or one SRM for every 40 samples.

 

8.4.2

Blanks

Generally, for RC drilling, blanks are inserted at intervals of 15 meters (50 ft) and multiples of 15 meters (50 ft). For core drilling samples, blanks are inserted at nominal 60 meters (200 feet) intervals. This results in a frequency of SRM insertion of between 2% to 5%. The actual rate of insertion depends on the time of operation.

 

8.4.3

Check Samples

Approximately 5% of the total material is dispatched to umpire laboratories as part of the check assay program. Typical checks will be conducted on pulps and coarse reject samples to test the analytical processes and preparation procedure. Overall, each sample batch submitted for analysis will contain between three to seven check samples.

 

8.5

Opinion of the Qualified Person

The QA/QC processes followed on-site for the collection of samples, and preparation for chemical analyses of industry standards. The use of blanks and check samples to cross-check the laboratories are similar to industry practices globally. Overall the QA/QC processes followed provide sufficient assurance for making the geological data reliable for use in the estimation of mineral resources.

 

9.

DATA VERIFICATION

 

9.1

On-site Verification of Data and Information

As a part of the S-K 1300 report work, Mr. Brian Arthur and Dr. Abani Samal made a site visit on Wednesday, August 28th, and Thursday, August 29th, 2024. A prior site visit was also completed on July 7th and 8th, 2021 as a part of the 2021 NI 43-101 report. During the site visit the following tasks were completed.

 

   

The general geology of the deposit was reviewed in the field.

 

   

Select drill hole intercepts were physically verified.

 

 

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Verification of available core and core-logging, and cutting facilities at the new Lone Tree storage facility.

 

   

Visited to the leach pads and the mill facilities and had detailed discussions with site personnel.

 

   

Verification of the on-site laboratory procedures.

Drill Hole Locations: The verification of locations of select holes was completed in 2021 and those locations have not changed. No new holes have been drilled since the 2021 technical report was published. Therefore, the outcomes of the 2021 review of the drill hole locations are still valid.

The location of the most recent drill hole LTE-20001 drilled in 2020 is shown in Figure 9-1. Additionally, existence of other drill holes in the Sequoia zone were also verified during the 2021 site visit. Location of the LTE-20001 hole was cross-checked with the collar data and found to be accurate. The hole locations are preserved with a wooden stick and an aluminum plate (Figure 9-1). It should be noted that the numbers shown on the aluminum plate are not the actual drill hole numbers, but rather reference numbers that link to the drill hole numbers.

Figure 9-1:Location of LTE-20001 on the West Side of the Lone Tree Pit Exploration holes were drilled over time while the Lone Tree mine was being developed and mining was progressing.

 

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Figure 9-2:Location Tag of an Exploration Drill Hole in Sequoia

 

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This makes it impossible to cross check the actual locations of the drill holes inside the current pit limit.

Verification of Drill Core Logs: During the 2021 trip, the drill cores were personally checked and mineralizations on certain core intercepts were verified at the core-processing center of NGM at their Battle Mountain location. The drill cores were processed as per industry best practices. The drill core processing and logging facility of NGM was well equipped with core cutting tools and logging tables. Half of the mineralized portion of each drill core was sent to the laboratory for assay. The remaining halves of the core were well preserved in core trays, as shown in the Figure 9-3.

After the acquisition of the Lone Tree property, portions of the drill cores and RC chips of selected intercepts of the latest hole drilled by Newmont/NGM (LT 20001) were transferred to i-80. Half cores and chips of selected intercepts of this hole were laid on tables as shown in Figure 9-4 and Figure 9-5. During this onsite verification of drill cores and chip samples, a portion of the lithological logs for this hole was also reviewed. This core logging was done by NGM geologists.

 

 

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Figure 9-3:An Example of a Mineralized Core Stored in a Core Box

 

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Figure 9-4:Half Core of LTE-20001 from 2544 ft to 2549 ft with 1.215 ppm Gold Grade Figure 9-5:The Chip Samples from the Upper Portion of the Hole (LTE-20001)

 

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During both the 2021 and recent visits, discussions were held on the characteristics of mineralization and textural features with the geologists in charge. A few pictures of core boxes and some core logs were taken as examples of the detailed logging procedures. It was noted that the processes followed have been consistent throughout the life of the Lone Tree deposit.

Figure 9-6:The Pulps Preserved in a Storage Area of i-80

 

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During this site visit, the storage facility of i-80 was inspected. The pulps of old drill holes were stored with a digital database and map of the storage area. Figure 9-6 is an example of how the pulps are labeled. No drill core from the older holes was available for inspection. Some pulps will be sent for re-analysis with selected cores from the LTE-20001 holes.

 

 

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The rejects from geochemical analysis were stored in barrels. These were not well labelled and documented.

 

9.2

Review of the QA/QC procedure

During the 2021 site visit, the QP was given access to the core processing facilities and a review of the available documents on the QA/QC procedures. The onsite exploration QA/QC procedures were reviewed and discussed during meetings held during the above-mentioned visits. Nothing has changed since then.

 

9.2.1

Re-assay of pulps and drill cores

To cross-check the assay data of Lone Tree that was acQuired by i-80 for trustworthiness, it was decided to re-assay selected drill cores and pulps available to i-80. A list of selected intercepts of drill holes with available pulps was provided to the geologist in charge of i-80 for re-assay. A total of 194 samples were selected for re-assay. Out of 194 samples, 183 pulps were selected from 19 holes, and 11 core samples were selected from the drill hole LTE-20001. These samples were randomly chosen from various depths and with various known assay values as shown in Table 9-1:Details of the Samples Selected for Re-assay.

Table 9-1:Details of the Samples Selected for Re-assay

 

Hole

Number

  

Number of re-

assayed data

  

Depth_min

(ft)

  

Depth_max

(ft)

  

Type of

samples

 

  

LTE-01473

  

12

  

180

  

320

   Pulp

LTE-01474

  

13

  

95

  

280

   Pulp

LTE-01475

  

8

  

190

  

260

   Pulp

LTE-01476

  

11

  

60

  

260

   Pulp

LTE-01478

  

9

  

195

  

340

   Pulp

LTE-01479

  

9

  

180

  

340

   Pulp

LTE-01480

  

13

  

115

  

420

   Pulp

LTE-01481

  

10

  

120

  

315

   Pulp

LTE-01482

  

10

  

30

  

300

   Pulp

LTE-01484

  

11

  

175

  

250

   Pulp

LTE-01485

  

7

  

25

  

245

   Pulp

LTE-01487

  

9

  

140

  

280

   Pulp

LTE-01488

  

7

  

160

  

255

   Pulp

LTE-01489

  

9

  

205

  

280

   Pulp

LTE-01490

  

7

  

260

  

330

   Pulp

LTE-01491

  

7

  

70

  

320

   Pulp

LTE-01492

  

13

  

140

  

265

   Pulp

LTE-01500

  

9

  

65

  

200

   Pulp

LTE-01502

  

9

  

130

  

270

   Pulp

LTE-20001

  

11

  

1642

  

2572

   Core

 

 

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9.2.1.1

Summary of Re-assay Data

The samples were sent to the ALS laboratories, where certified reference materials (CRM) and blank samples were inserted. A CRM was inserted at every 8 to 10 samples. Four different certified reference materials (CRM) were used: Oreas 277, Oreas 264, Oreas 279, and CDN-GS-P6E. Details of the CRMs are shown in Table 9-2.

Table 9-2:Descriptions of the CRMs Used

 

CRMs    Unit    

Statistical 

mean

  

Standard

Deviations  

  

Analytical Method (summary)

Oreas 277    ppm    3.39    0.12   

Fire assay (15-50g charge weight) with AAS (ICP-OES or ICP-MS finish;

Oreas 264    ppm    0.307    0.011   

Fire assay (15-40g charge weight) with AAS, ICP-OES or ICP-MS finish

Oreas 279    ppm    6.55    0.218   

Gold by fire assay (15-50g charge weight) with AAS, ICP-OES or ICP-MS finish

CDN-GS-P6E    g/t    0.572    0.0155   

30g FA / AA or ICP Finish

The analytical results were compared to the original assay data. Figure 9-7 suggests that the re-assayed data were very close to the original assay data. However, there are some assay data with substantially large -differences from the original data, out of which most are very low-grade data (Figure 9-8). These results suggest that the original drill hole assay data acquired by i-80 from Newmont / NGM are of reliable quality without any major concerns.

Figure 9-7:Comparison of the Re-assayed Analytical Results with the Original Data.

 

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Figure 9-8:The Relationship Between the % Differences Original Au Assay Data

 

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9.2.2

Database Review

A detailed database review was conducted during 2021. The drill hole data was shared by the NGM staff via an electronic cloud-based data room facility and organized in multiple Excel files. Dr. Samal (the lead author of this report) compiled the data using Vulcan software. A total of 1839 drill hole data was selected for independent assessment of mineral resources. Using Vulcan software, drill hole data were checked for errors.

 

   

For this report, a small set of drill holes was selected to cross-verify the assay data. No errors in the drill hole database were found.

   

A significant number of holes had lithological logs. However, out of 1840 selected holes 362 holes had no lithological logs (Refer to Section 7.1). This had no impact on mineral resource estimates for this report.

The processes followed for drill core collection, storage, and sample preparation are well documented in the form of standard operational procedures. The procedures meet industry best practice guidelines for exploration data, which adds to the reliability of the geological interpretation and assay data.

 

   

Bias Study: A study was conducted by NGM (then Newmont) to examine potential bias in the database due to various types of drilling. In this study, the exploration data above

 

 

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and below the water table were considered. The study found bias between rotary hole samples below the water table and core or blast holes. However, as the mine deepens, the proportion of core to rotary composites will increase making it unnecessary to correct for the defined bias. No bias adjustments were used in the resource model (Zacarias, 2006).

 

9.3

Opinion of the Qualified Person

The QP responsible for producing an independent assessment of mineral resources below the current pit reviewed the data and information available for the Lone Tree deposit. These included the topographic data, the drill hole data, the geological interpretation data, the density data, and documents supporting the processes and procedures followed for collection, compilation, storage, security, and quality control.

The above-mentioned reviews concluded that data collected by Newmont/NGM on the Lone Tree Project was acquired using adequate quality control procedures that generally meet industry best practices for an operating mine. The Quality Assurance and Quality Control (QA/QC) processes followed for maintaining the quality of the data meet the best practices guidelines as outlined in § 229.13053 (Item 1305) of Regulation S-K (Subpart 1300). The data is adequate for use in undertaking a mineral resource estimate.

It is the opinion of the QP that the data provided by NGM in 2021 is suitable to be used as the basis of a mineral resource estimate in this project and can also be used in the future studies on Lone Tree. Therefore, no changes are required in the resource model parameters as discussed in Section 11.

 

10.

MINERAL PROCESSING AND METALLURGICAL TESTING

Newmont did a metallurgical study on ore from the Brooks deposit which was mined and processed between 2015 and 2020. During the ten years of operation prior to the acquisition by i-80, Newmont did not conduct any comprehensive metallurgical studies on ore from the Lone Tree deposit. Since its acquisition in 2021, i-80 has not conducted any metallurgical tests or pilot studies on Lone Tree ore(s).

The remaining mineralization in the Lone Tree deposit is located in extensions of the mineralized zones and structures that were mined and processed by Newmont between 1991 and 2019. Therefore, processing data drawn from historical production reported by Newmont/NGM was used to support the metallurgical performances reported herein.

Most of the remaining mineralization in Lone Tree is expected to be sulfidic and require autoclave pretreatment to facilitate gold leaching. Some low grade oxide and high grade

 

 

3 17 CFR Subpart 229.1300 —Disclosure by Registrants Engaged in Mining Operations (Electronic Code of Federal Regulations (e-CFR) Title 17—Commodity and Securities Exchanges CHAPTER II—SECURITIES AND EXCHANGE COMMISSION PART 229—STANDARD INSTRUCTIONS FOR FILING FORMS UNDER SECURITIES ACT OF 1933, SECURITIES EXCHANGE ACT OF 1934 AND ENERGY POLICY AND CONSERVATION ACT OF 1975—REGULATION S-K)

 

 

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oxide ores are also expected but not valued as significant in this study. The historical and proposed recoveries for each material type are shown in Table 10-1 and are expected to prevail in the remaining material.

Table 10-1:Recoveries and Material Types or Lone Tree

 

Mine    Definition    Process   

2005

Actual4

(%)

  

Life of

Property

Actual

(%)

  

Proposed Au

Recovery1

(%)

Lone Tree    High Grade Oxide ore    CIP/CIL   

94.2

  

NA

  

60.0

Lone Tree    High Grade Sulfidic Ore    Autoclave/CIL2   

94.6

  

93.055%

  

94.9

Lone Tree    Concentrates    Autoclave/CIL   

91

  

NA

  

93.9

Lone Tree    Low Grade Sulfides    Flotation3   

77.52

  

81.355%

  

78.59

Lone Tree    Low Grade Oxides    Heap Leach   

NA

  

816%

  

67.3

Lone Tree    Leach Grade Sulfides    Heap Leach   

NA

  

NA

  

63.6

  1.

Source - Zacarias, P., A., February 28, 2006, 2005 Mineral Resource and Ore Reserve Report as of December 31, 2005, pp65.

  2.

Autoclave recovery based on acid autoclave.

  3.

Flotation Recovery - Recovery is to Concentrate= 83.7% and then 93.9% recovery from the Concentrate results in a - Combined Recovery - 83.7(93.9) = 78.59%.

  4.

Source - 2005 LT Summary.xls

  5.

Source - Lone Tree Statistics (1998 - H1 2019).xlsx.

  6.

Source - 2.4.4.1.3 Brooks leach curve EA.25.xlsx and i-80 Pad Tracking LoneTree.xlsx

Details of metal recovery and cost estimates for various processing methods are discussed in Section14.

 

10.1

Opinion of the Qualified Person

It is the opinion of the qualified person that the extension of the historical factors to ores that have yet to be mined is sufficient to support this Mineral Resource Estimate. The following comments in support of the declaration are believed to be true.

   

The same ore types and process methods are projected for future mining and process activities as Newmont exploited historically.

   

The historical production values reported by Newmont are believed to be accurate.

 

 

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11.

MINERAL RESOURCE ESTIMATES

GeoGlobal was mandated to review the estimates of in-situ mineral resources of the Lone Tree deposit made in 2021 and update them as appropriate. The following data was available for use in the resource estimation process. Maptek’s Vulcan Version 2023.4.1 software tool was used for this purpose.

 

11.1

Datasets

As mentioned in Section 9, the database provided by Newmont/NGM in 2021 and used in the 2021 resource estimates contains reliable data and is valid for use in this project.

 

11.1.1

Drill Holes

The dataset contains 3,396 drill holes. This dataset includes holes from Lone Tree, Buffalo Mountain, Lynn, and Second Chance exploration properties. Drill holes close to Lone Tree were selected. This selection contains 1840 drill holes (Figure 11-1). The coordinate system used is the local Lone Tree mine-grid system.

 

11.1.2

Other Data Sets

 

   

Triangulations representing topography including the latest pit-outline and pre-mining topography.

 

   

Three dimensional geological interpretations of rock types and interpretation of fault planes were created using Leapfrog software tool. During this process geological units were combined as shown in Figure 11-2.

 

   

A block model was created using Newmont’s in-house software TSS 3-dimensional modeling Geomodel software. Details are discussed later in this section of the report.

 

 

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Figure 11-1:Location of the Drill Holes at the Lone Tree Project Figure 11-2:The Geological Codes Used In Creating Solid Model

 

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11.2

Geological Interpretation and Lithological Models

As discussed in Section 7, the mineralization at Lone Tree is primarily controlled by fault systems within certain rock types. It is clear that not all rock types are mineralized equally; rather mineralization occurs variably within three Paleozoic rock sequences at the Lone Tree deposit: the Valmy Formation, the Antler Sequence and the Pennsylvanian-Permian Havallah sequence rocks. The Wayne zone is known for rich-mineralization due to the Powerline fault cutting through the favorable rock-types. The rock-type models include five lithologic groups as listed below.

 

  o

Quaternary/tertiary alluvium, colluvium etc (QAL).

  o

Mississippian Permian Havallah (Phv).

  o

Permian Edna (Pem).

  o

Pennsylvanian Battle Mountain (Pb).

  o

Ordovician Valmy (Ova).

 

 

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Figure 11-3 is a vertical Cross Section along 28000 N. This figure shows the geological model created by the NGM geologists that has been adopted for this resource estimation process.

Figure 11-3:A Vertical Cross Section at 28000N (Looking North)

 

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The Power Line fault is the most important structural zone at the Lone Tree deposit. As shown in Figure 11-4, the mineralization appears to be controlled largely by these structural elements. The other major fault at the mine is the Sequoia fault on the south side of the deposit.

Figure 11-4:A Vertical Cross Section with Rock-type Models (27300N)

 

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A block model with dimensions listed in Table 11-1, was created to estimate mineral resources of the Lone Tree deposit. The block model is oriented parallel to the Lone Tree mine grid coordinate system.

 

 

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Table 11-1:Block Model Geometry

 

     East (X)    North (Y)    Elevation (Z)
Origin    78575     20475     -1200 
Block dimensions    50     50     20 
Number of blocks    200     300     310 

 

11.3

Lithology Model

The lithology model was updated using the new lithology solids provided by NGM. The block model was coded based on the geology triangulation model. The variable is ‘newlith’. The codes assigned are presented in Table 11-2.

Table 11-2:The Lithology Codes Used in the Block Model

 

Lithology

  

‘newlith’ code in

the block model

  

NEWGEOL variable

in the composite

data

QAL: Quarternary/ tertiary alluvium, colluvium, etc.

   1    1

Phv: Missi., Perm. Havallah

   2    2

Pem: Perm. Edna

   3    3

Pb: Pennsylvanian Battle Mountain

   4    4

Ova: Ordovician Valmy

   5    5

The lithology block model shown in Figure 11-5 has been coded with the drill hole lithology codes as shown in Figure 11-3.

Figure 11-5:The Lithology Block Model (28000 N, Looking North)

 

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11.4

Exploratory Data Analysis and Compositing

The purpose of the exploratory data analyses (EDA) is to characterize the dataset so that it can be effectively used in grade interpolation and resource estimation. The results of the EDA process help in developing parameters for grade interpolations as discussed below.

 

 

 

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In this project, the EDA section was reviewed. No changes were made to the results of the EDA.

Figure 11-6:Histogram of Drill Hole Assay Lengths

 

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11.4.1

Compositing

Nearly 93% of the drill hole intercepts are five feet long (Figure 11-6). To minimize grade dilution due to compositing and create a database with sample intervals with equal lengths the drill hole data were composited on five-foot (5.00 ft) intervals.

 

11.4.2

Statistical Analyses of Composited Data

The general statistics of the gold composited data (ounces per ton) are presented in Table 11-3. Figure 11-7 compares the means and standard deviations of all five lithology types.

Figure 11-7 shows that the Havallah formation (Phv) has the dominant number of composites. The figure also suggests that the Edna formation (Pem) data hosts higher gold assays compared to other lithology types. The variability of the gold assays (measured by two standard deviations in this figure) is also greatest in the Edna formation.

The use of confidence interval (CI) is a better measure of variances for comparing sample sets of different sizes, as in the CI, the standard deviation is normalized by the number of samples that compares the means and variabilities of the composites in different rock types. Figure 11-8 further confirms that the gold assays in the Pem (Edna formation) are of highest grade followed by Pb (Battle Formation) and Ova (Valmy).

 

 

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Table 11-3:General Statistical Characteristics of the Gold (AuFA, opt)

 

 Lithology    Num
Samples 
   Min     Q1     Median     Q3     Max     Mean     Standard
Dev.

 Newgeol:1 (QAL)

   9711    0.000    0.001    0.001    0.002    1.379    0.005    0.029

 Newgeol:2 (Phv)

   119337    0.000    0.001    0.001    0.001    2.410    0.009    0.049

 Newgeol:3 (Pem)

   23193    0.000    0.001    0.006    0.028    5.652    0.036    0.089

 Newgeol:4 (Pb)

   12401    0.000    0.001    0.004    0.014    2.074    0.020    0.059

 Newgeol:5 (Ova)

   27940    0.000    0.001    0.003    0.011    1.014    0.016    0.045

 Newgeol 0 (not tagged)

   11212    0.000    0.001    0.001    0.001    0.078    0.001    0.002

 All

   203794    0.000    0.001    0.001    0.004    5.652    0.013    0.054

The histograms of all composites are presented in the Figure 11-9. Histograms show that the Havallah (code 3, Phv) are relatively low-grade rocks. In the cumulative frequency plot of the same dataset

Figure 11-10), it appears that the gold composites of the Edna formation (Pem) have higher proportions of higher-grade intercepts compared to the other rock types.

Figure 11-7:Comparison of Data Statistics of All Lithology Types Figure 11-8:Comparison of Sample Means and Confidence Intervals

 

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Figure 11-9:Histograms of Gold Assay Values (AuFA) by Lithology Figure 11-10:Cumulative Frequency Plots of Gold Composites (AuFA)

 

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11.4.3

Contact Plots

Contact plot analyses of AuFA across different lithological boundaries are done to determine the behavior of AuFA near the lithological boundary. This information is important in deciding whether to treat the geological areas as formal ‘domains’ for grade estimation and whether to treat the boundaries as hard or soft boundaries. In this project, Vulcan software was used for contact plot analysis.

The contact plot of AuFA for the boundary between Qal (Quarternary Alluvium) and (Phv Havallah) shown in Figure 11-11 indicates that the Qal is relatively low grade compared to Phv. The gold values change gradually across the boundary. This interpretation is non-conclusive.

The contact plot of AuFA for the boundary between (Phv Havallah) and Pem (Edna) is presented in Figure 11-12. The Pem is of higher grade than the Phv. The contact between Phv and Pem is not sharp, but distinct.

The contact plot analyses shown in Figure 11-11 through Figure 11-14, indicate that the gold assays behave differently for different rock types. Even though the contacts are not sharp they are distinct.

 

 

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Figure 11-11:Contact Plot of AuFA for the Boundary Between Qal and Phv

 

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Figure 11-12:Contact Plot of AuFA for the Boundary Between Phv and Pem

 

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Figure 11-13:Contact Plot of AuFA for the Boundary Between Pem and Pb Figure 11-14:Contact Plot of AuFA for the Boundary Between Pb and Ova

 

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11.5

Geological Domains

Based on the statistical analysis above the lithology units are considered as domains within which the mineralization is assumed to behave consistently throughout. The variogram analyses and interpolation parameters are derived for the gold grade estimation within the lithological domains.

 

11.6

Variogram Analysis

The statistical analysis, as discussed earlier, provides the information used for variogram analysis. The variogram analysis process is presented in the Figure 11-15. Variography is a stepwise process.

Figure 11-15:Steps Followed in Variogram Analysis A detailed review of the variogram process and results used in the 2021 resource estimation project was conducted.

 

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The results of variogram analysis are valid for use in the resource estimation process.

 

  i.

Declustering: When the drill hole data shows clustering effects, at certain locations the drill hole intercepts are closer to each other compared to other locations. A cell declustering process corrects potential bias due to preferential sampling of high-grade zones.

 

  ii.

Variogram Map: To find the inherent spatial anisotropy of the database, variogram maps of each variable lithologic domain are calculated using declustered data.

 

   

Variogram maps are calculated for various lag lengths, tolerance angles and distances of tolerances. Angles of tolerances are ideally kept at half of the angular sectors in a variogram map. The maximum distances of tolerances are kept at half of the lag distance. The variogram maps are shown in Appendix B.

 

   

The variogram maps indicate a North-South and East-West orthogonal set of anisotropies. These anisotropy interpretations correspond to the structural controls of mineralization and hence are geologically valid.

 

  iii.

Experimental variogram calculation: The experimental variograms are calculated along the three orthogonal directions as selected from the variogram maps. The angles of tolerance are ideally kept at half of the angular sectors in a variogram map. The maximum distances of tolerance are kept at half of the lag distance. The experimental variograms are saved for modelling.

 

  iv.

Variogram model: An approved type of variogram model is selected for fitting to the experimental variograms calculated in the prior stage.

 

  v.

A set of three orthogonal variogram models is fitted for each of the elements within each of the lithological domains. As variograms are fitted for the one domain, they do not show regional anisotropy and display geometric anisotropy. This implies same sill and nugget effects for all three variogram models in a set. The variogram ranges may change (anisotropy) in three directions. In this exercise, multiple structures for variogram models were not observed. Variogram model parameters are presented in the Table 11-4. The Variogram models are presented in Appendix C.

 

 

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Table 11-4:Variogram Parameters of AuFA

 

 Lithology units         Azimuth     Dip        Ranges  

Newgeol: 1 (QAL)

  Major Axis    0       0        90  
  Semi-Major Axis    90       0        70  
  Minor Axis    90       -90         200  
  Nugget    0.00002                   
  Sill    0.000155                   

Newgeol: 2 (Phv)

  Major Axis    0       0        90  
  Semi-Major Axis    90       0        90  
  Minor Axis    90       -90         200  
  Nugget    0.0005                   
  Sill    0.001                   

Newgeol: 3 (Pem)

  Major Axis    0       0        100  
  Semi-Major Axis    90       0        60  
  Minor Axis    90       -90         150  
  Nugget    0.002         
  Sill    0.0048                   

Newgeol: 4 (Pb)

  Major Axis    0       0        100  
  Semi-Major Axis    90       0        200  
  Minor Axis    90       -90         120  
  Nugget    0.0005         
  Sill    0.004                   

Newgeol: 5 (Ova)

  Major Axis    0       0        75  
  Semi-Major Axis    90       0        75  
  Minor Axis    90       -90         150  
  Nugget    0.0003         
  Sill    0.000995                   

 

11.7

Density /Tonnage Factor Model

NGM determined density factors for the rock types in the Lone Tree deposit in 2003. This work was reviewed and deemed satisfactory for use in this resource estimate. The block model used tonnage factors for various lithological units. This variable was imported from the NGM Block model. The statistics of the tonnage factors are shown in Table 11-5.

Table 11-5:Tonnage Factor by Geologic Unit (ft3/ton)

 

Lith Units    Num Samples        Min       Median       Max       Mean       Standard
Deviation
 

Newgeol:1 (QAL)

     163658        12.20        16.00        16.00        15.66        0.93  

Newgeol:2 (Phv)

     414357        12.20        13.10        16.00        13.21        0.55  

Newgeol:3 (Pem)

     69318        12.20        13.10        16.00        12.91        0.50  

Newgeol:4 (Pb)

     28926        12.20        13.10        16.00        13.18        0.55  

Newgeol:5 (Ova)

     632029        12.20        13.10        16.00        13.12        0.29  

Newgeol 0 (not tagged)

     1308288        12.20        13.10        16.00        13.46        0.98  

 

 

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11.8

Grade Interpolation

Variogram models could be fitted for AuFA for all geological domains (Table 11-4 and Appendix C). Therefore, ordinary kriging (OK) was chosen as the interpolation technique over non-geostatistical techniques such as inverse distance power (IDP).

The different lithology units were considered unique geological domains. For the estimation of gold grades within a particular domain, the composites within that domain were used. The boundaries between various domains were considered hard boundaries.

The key points in the interpolation of AuFA grade are listed below.

 

i.   

No data are treated as an absence of data.

ii.   

Value zero is treated as zero value.

iii.   

The search ellipsoids are designed according to the orientations of the variogram models. Multi-pass search ellipsoids are used for estimating copper grades in each domain.

iv.   

The first search ellipsoid is the smallest one.

v.   

The axes of the next consecutive passes are relatively larger.

vi.   

A restricted search ellipsoid of 50ft X 50 ft X 30ft is used for the inclusion of high-grade values. A 0.25 opt Au was considered the high-grade value in the grade interpolation.

The following additional variables are saved along with the kriged gold values.

 

i.   

Interpolation pass numbers.

ii.   

Blocks estimated during the first pass are tagged as 1.

iii.   

Blocks estimated during the second pass are tagged as 2.

iv.   

Blocks estimated during the third pass are tagged as 3.

v.   

Blocks estimated during the fourth pass are tagged as 4.

vi.   

Blocks estimated during the fifth pass are tagged as 5.

vii.   

The average distance from samples from block centers.

viii.   

The distance of the closest sample from block centers.

ix.   

Number of samples used to estimate each block.

x.   

Number of holes.

xi.   

Nearest Neighbor Estimate (closest sample value).

xii.   

Kriging variance, which is a measure of error of estimation.

  

Note: This parameter should be used in conjunction with other variables.

xiii.   

Kriging efficiency: A measure of quality of grade estimation using ordinary kriging.

Note: This parameter should be used carefully as the mathematics are not well proven to be robust.

The interpolation parameters for AuFA estimation are shown in Table 11-6.

 

 

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Table 11-6:Interpolation Parameters Used for Estimation of Gold Grades

 

      Drill hole limits      Sample Criteria    Axes      High Yield Limits  
ID   

Holes /

 estimate 

  

Minimum 

/ hole

    

Maximum 

/ holes

     Min       Max      

Max. /  

hole

   Major      

Semi- 

Major

     Minor       Threshold      

Major 

axis

    

Semi- 

major

    

HYL - 

Minor

 

ova5_1

   No      1        10        2        20      X      50        50        20        0        50        50        50  

ova5_2

   Yes      2        10        8        24      6      90        90        100        0.25        50        50        30  

ova5_3

   Yes      2        10        6        24      6      130        130        150        0.25        50        50        30  

ova5_4

   Yes      2        10        4        24      6      160        160        200        0.25        50        50        30  

ova5_5

   No      1        10        2        24      X      180        180        200        0.25        50        50        30  

pb4_1

   No      1        10        2        20      X      50        50        20        0        50        50        50  

pb4_2

   Yes      2        10        8        24      6      90        90        100        0.25        50        50        30  

pb4_3

   Yes      2        10        6        24      6      130        130        150        0.25        50        50        30  

pb4_4

   Yes      2        10        4        24      6      160        160        200        0.25        50        50        30  

pb4_5

   No      1        10        2        24      X      180        180        200        0.25        50        50        30  

pem3_1

   No      1        10        2        20      X      50        50        20        0        50        50        50  

pem3_2

   Yes      2        10        8        24      6      90        90        100        0.25        50        50        30  

pem3_3

   Yes      2        10        6        24      6      130        130        150        0.25        50        50        30  

pem3_4

   Yes      2        10        4        24      6      160        160        200        0.25        50        50        30  

pem3_5

   No      1        10        2        24      X      180        180        200        0.25        50        50        30  

phv2_1

   No      1        10        2        20      X      50        50        20        0        50        50        50  

phv2_2

   Yes      2        10        8        24      6      90        90        100        0.25        50        50        30  

phv2_3

   Yes      2        10        6        24      6      130        130        150        0.25        50        50        30  

phv2_4

   Yes      2        10        4        24      6      160        160        200        0.25        50        50        30  

phv2_5

   No      1        10        2        24      X      180        180        200        0.25        50        50        30  

qal1_1

   No      1        10        2        20      X      50        50        20        0        50        50        50  

qal1_2

   Yes      2        10        8        24      6      90        90        100        0.25        50        50        30  

qal1_3

   Yes      2        10        6        24      6      130        130        150        0.25        50        50        30  

qal1_4

   Yes      2        10        4        24      6      160        160        200        0.25        50        50        30  

qal1_5

   No      1        10        2        24      X      180        180        200        0.25        50        50        30  

 

 

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11.9

Resource Classification

Resource classification is a process of assigning confidence to the mineralized body by a Qualified Person (QP) based on geological continuity, quality of data, and the quality of resource estimation. As discussed in Section 9, NGM has followed the QA/QC programs which meet the industry best practices guidelines.

The geological interpretation of the lithological units produced by the NGM Geologists is still valid and ensures reliability for use in grade estimation and assigning confidence to classify the inventory of materials that meets the criteria for reasonable prospects for economic extraction (RPEE) of the mineral resource as defined in the SK 1300 (§ 229.1302 17 CFR Ch. II, 4–1–24 Edition). The processes followed for estimating gold grades for the Lone Tree deposit meet the industry best practice guidelines as referred to in the SK 1300. Using sample and distance criteria, all blocks contained below the current pit limit (Figure 11-16) are considered candidates for the tests of RPEE.

Figure 11-16:Blocks Below the Current Pit Used for the Optimized Pit shell

 

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11.10

Criteria for Reasonable Prospect for Economic Extraction

To meet the RPEE criteria, an optimized pit shell was created within which all blocks will be considered as resources. An optimized pit shell was generated using a $2,175 gold price, $44.50 per ton milled Autoclave processing cost, $29.91 per ton milled flotation processing cost (Refer Section 14.5), and 94.9% recovery factor for autoclave and 78.6% recovery factor for flotation (Refer Section 10, Table 10-1) based on the operational data from i-80 Gold. The gold price of $2,175 was provided by i-80 using the consensus forward prices as of December 2024 as provided by major Canadian financial institutions.

 

 

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Gold Price as of December 31, 2024:

Gold is a fungible commodity with reputable smelters and refiners located throughout the world. The price of gold has reached all-time highs in 2024 with December’s price averaging 2,644 per ounce. As of December 31, 2024, the three-year trailing average gold price was $2,044 per ounce and the two-year trailing average price was $2,166 per ounce.

The registrant may also rely on published forecasts from reputable financial institutions. The current long-term price forecast by the Canadian Imperial Bank of Commerce (CIBC) is $2,169 per ounce. The commodity prices are chosen not to materially exceed financial institution forecasts or the two-year trailing average price. The gold price selected for estimating Mineral Resources disclosed in this technical report is $2,175 per ounce.

This pit shell meets the RPEE criteria and, hence, will be used for resource statements. The cutoff grade is the minimum gold grade at which the value for ore processing is positive and applying a 3% royalty. The cutoff grade used is 0.018 ounces per ton or 0.62 grams per metric tonne.

The following parameters (Table 11-7) are used for generating the optimized pit limit. The optimum pit shell was developed by Mr. Paul Gates, an associate of GeoGlobal. This pit shell will be referred to as the $2,175 pit shell.

Table 11-7:Optimum Pit Criteria Applied to Resource Estimate

 

     
Variables    Value    Notes
     
Au Price per ounce    $2,175   

Provided by i-80 Gold

     
Mine Cost ($/ton)    $3.00   

Rock (Provided by i-80 Gold)

   $2.75   

Fill material (Provided by i-80 Gold)

     

Processing + G&A

Cost ($/ton)

  

$44.50 for autoclave

$29.91 for flotation

  

Assume 2,500 tons per day autoclave and 5,000 tons per day Flotation (Refer Section 14)

     
Recovery   

94.9% for autoclave

78.6% for flotation

  

Refer to Section 10, Table 10-1

     
Royalty    3.0%   

NSR

     
Cutoff Grade    0.018 Opt   

(i.e. 0.62 g/Tonne) Refer discussions above

     
Slope Angles    40°-45°   

Azimuth from 20o to 219o slope angle 45o

Azimuth from 220o to 19o slope angle 40o

 

11.10.1

Inferred blocks

All blocks within the $2,175 pit shell were tagged as inferred category (auok_mii = 3). All blocks within the $2,175 pit shell were estimated using at least two holes, a minimum of 4 composites. Additionally, 90% of these blocks have composites within 120 feet, and more than 77% of these blocks have composites within an average distance of 120 feet. It should be noted that the variogram models show ranges of 100 to 120 feet. Any portion of these blocks contained within a Resource pit shell is well qualified to be reported as inferred resources.

 

 

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11.10.2

Indicated blocks

All blocks within 40 to 50 feet (one bench height) from the current pit can be indicated based on additional criteria considered:

   

The geology and mineralization in blocks are partially exposed at the current pit and supported by previously drilled blast hole / infill drilling data

   

Continuity of geology and mineralization exposed on the pit-surface can be comfortably projected one bench below

The blocks classified as inferred and indicated category resources are shown in the Figure 11-17. At this time, no measured resources are defined for the Lone Tree deposit.

 

11.10.3

Inventory of Mineral Resources

The estimates of indicated and inferred category mineral resources within the optimized pit shell are provided in Table 11-8 for various cutoff grades.

Figure 11-17:The Blocks Classified as Indicated and Inferred Category Resources Table 11-8:Inventory of Mineral Resources Within $2,175 Pit Shell

 

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Indicated Resources

(auok_mii=2)

  Inferred Resources (auok_mii=3)

 Cutoff (Au 

(OpT)

  Au(OpT)   Tonnage   Oz Gold   Au (OpT)    Tonnage    Oz Gold

0.01 

  0.0379     12,838,420     486,191   0.0383     80,822,109     3,096,295 

0.011 

  0.0399     11,953,766     476,955   0.0396     77,316,387     3,059,409 

0.012 

  0.0414     11,340,975     469,857   0.0408     74,159,800     3,022,753 

0.013 

  0.0430     10,749,461     462,442   0.0421     70,926,196     2,982,447 

0.014 

  0.0445     10,232,365     455,545   0.0433     68,085,392     2,944,693 

0.015 

  0.0459     9,781,243     448,959   0.0444     65,374,935     2,905,262 

0.016 

  0.0475     9,308,452     441,686   0.0456     62,821,268     2,865,278 

0.017 

  0.0489     8,886,209     434,713   0.0468     60,422,522     2,825,961 

0.018 

  0.0504     8,478,659     427,579   0.0478     58,340,344     2,789,252 

0.019 

  0.0517     8,159,931     421,705   0.0489     56,318,134     2,752,267 

0.02 

  0.0528     7,878,278     416,209   0.0499     54,495,014     2,716,576 

0.022 

  0.0554     7,292,916     403,955   0.0521     50,574,558     2,634,429 

0.024 

  0.0578     6,792,444     392,399   0.0542     47,127,926     2,554,805 

0.026 

  0.0606     6,255,499     379,021   0.0565     43,713,643     2,469,384 

0.028 

  0.0634     5,770,842     365,987   0.0588     40,550,374     2,384,362 

0.03 

  0.0659     5,378,369     354,596   0.0608     37,958,241     2,309,000 

Figure 11-18 and Figure 11-19 represent the grade tonnage curves of indicated and inferred category mineral resources.

Figure 11-18:The Grade Tonnage Curve of Indicated Category Resources Figure 11-19:The Grade Tonnage Curve of Inferred Category Resources

 

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11.11

Model Validation

 

11.11.1

Cross Sections

Figure 11-20 presents a vertical Cross Section at 29000N (facing north) with gold grades of the drill holes overlain by the estimated gold grades in the block model using the same color code.

Figure 11-20:Estimate Blocks with Assay Data Within the $2,175 Pit Shell This figure is presented to visually validate the grade estimates in the block model with the drill hole data.

 

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The gold estimate grades in the block model represent the spatial distribution of gold assay values in the drill hole intercepts reasonably well.

 

11.11.2

Statistical Validation

Statistical validation of the final estimates of gold (auok) includes charts and table of various variables as discussed below.

 

   

Number of holes used: All blocks classified as indicated and inferred were estimated using at least two drill holes with approximately 55% of blocks being estimated using more than two holes (Figure 11-21 and Figure 11-22).

Figure 11-21:Histogram of Holes Used for Estimating Indicated and Inferred Blocks Figure 11-22:Histogram of the Holes Used for Estimating Indicated and Inferred Blocks

 

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Number of samples used: All blocks are estimated using at least two holes and six samples. As shown in Figure 11-23, approximately 88% of the blocks were estimated using more than eight samples. Additionally, 64% of all indicated blocks and 61.5 % of inferred blocks were estimated using up to 14 samples (Figure 11-24)

Figure 11-23:Number of Composites Used to Estimate Blocks with Gold Grade Figure 11-24:Number of Composites Used to Estimate Blocks

 

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Average distance and minimum distances of samples: As stated earlier, all blocks were estimated using at least two holes and six samples. As shown in Figure 11-25, 69% of indicated blocks and 58% of inferred blocks were estimated using samples with an average distance of 100 feet or less.

Figure 11-25 Average Distance of Samples From Block Centers Figure 11-26:Minimum Distance of Composites From Blocks

 

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Similarly, 84.5% of indicated blocks and 77% of the inferred blocks have at least one composite at a distance of 100 feet or less (Figure 11-26). For comparison, it is noted that the variogram ranges vary between 100 and 120 feet.

Table 11-9 compares the gold grades of the composite data in each lithology unit with the statistics of the estimated grades in the block model. The coefficients of variation of estimated gold grades in each lithology type are relatively low compared to the composites. This likely results from interpolation. The statistics in this table indicate no fatal flaw or serious mistake in the grade estimates.

In Figures Figure 11-27 to Figure 11-31 the histogram and cumulative frequency plots of the block grade estimates are compared with the input composited data. These figures show that the statistical structures of the block grade estimates are very similar to the input composites. The grade estimates are statistically similar to the composited data, further validating the grade estimates in the block model.

 

 

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Table 11-9:Comparison of Gold Grades for Composite Data by Lithologic Unit

 

Composite Statistics

Lithological
Units
 

Min

(Opt)

 

Max

(Opt)

 

Mean

(Opt)

 

Standard

Dev.

(Opt)

 

Num

Samples

  CV

Comp1: QAL

  0.0001    1.3789    0.0045    0.0287    9711    6.36 

Comp2: Phv

  0.0000    2.4100    0.0088    0.0490    119337    5.54 

Comp3: Pem

  0.0001    5.6520    0.0356    0.0888    23193    2.50 

Comp4: Pb

  0.0001    2.0740    0.0202    0.0590    12401    2.91 

Comp5: Ova

  0.0001    1.0140    0.0159    0.0453    27940    2.85 

Block Model Statistics

Lithological
Units
 

Min

(Opt)

 

Max

(Opt)

 

Mean

(Opt)

 

Standard

Dev.

(Opt)

 

Num

Samples

  CV

Blk1: QAL

  0.0000    0.6246    0.0014    0.0069    68212    4.93 

Blk2: Phv

  0.0000    0.9812    0.0030    0.0149    428823    4.92 

Blk3: Pem

  0.0001    1.0185    0.0224    0.0387    50662    1.73 

Blk4: Pb

  0.0001    1.4541    0.0143    0.0293    31570    2.05 

Blk5: Ova

  0.0000    0.5550    0.0089    0.0160    209277    1.79 

Figure 11-27:Comparison of Composites and Block Grade Estimates in the Ova Figure 11-28:Comparison of Composites and Block Grade Estimates in the Pb

 

 

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Figure 11-29:Comparison of Composites and Block Grade Estimates in the Pem Figure 11-30:Comparison of Composites and Block Grade Estimates in the Phv

 

 

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Figure 11-31:Comparison of Composites and Block Grade Estimates in the Qal The estimated mineral resources at the end of the fiscal year ending on December 31, 2024, are presented in Table 11-10.

 

 

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11.12

Tabulation of Estimated Resources

Table 11-10: Estimated Mineral Resources at 0.62 g/T Cut-off Grade

 

      Million
Tonnes (MT)
   Au (g/T)    Au (K ozs)

 Indicated Mineral Resources

   7.69    1.73    428

 Inferred Mineral Resources

   52.94    1.64    2,789

Notes to accompany the Mineral Resource table for the Lone Tree deposit:

 

     i.

Mineral Resources have an effective date of December 31, 2024

    ii.

Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

    iii.

Mineral resources are shown above a 0.62 g/T (0.018 opt) Au cut-off grade.

    iv.

Mineral Resources are constrained to oxide and transitional oxide-sulfide mineralization inside a conceptual open pit shell. A gold price of $2,175/oz Au. A 94.9% recovery for gold in the autoclave process and 78.6% recovery for flotation is used as parameters for pit shell construction. Open pit mining costs of $3.00 per ton for rocks and $2.75 for fill materials, average processing 2,500 tpd autoclave at $44.50 per ton processed and 5,000 tpd flotation at $29.91 per ton including general and administrative costs are considered. A 3% NSR royalty and pit slopes of 40° to 45° was used.

    v.

Mineral Resources are stated as in-situ with no consideration for planned or unplanned external mining dilution.

    vi.

The contained gold estimates in the Mineral Resource table have not been adjusted for metallurgical recoveries.

   vii.

Units shown are in Million Tonnes (MT), grams per metric tonne (g/T), and thousands of ounces of contained gold (K ozs).

  viii.

Numbers have been rounded as required by reporting guidelines and may result in apparent summation differences.

 

11.13

Opinion of the Qualified Person

The QP is of the opinion that the resource estimation parameters (statistical parameters and variogram models) and techniques (interpolation and resource classification) used are appropriate for the Lone Tree database. The resource classification considers the uncertainties due to data quantity and quality as appropriate for the Lone Tree deposit for the preliminary assessment stage. The parameters used in the assessment of reasonable prospects for economic assessment are based on the most reliable data available at this stage. The final cutoff grade and pit shell were selected after multiple iterations of analyses. The estimated resources are reliable based on the various parameters and criteria as discussed in this section.

 

 12.

MINERAL RESERVE ESTIMATES

As per S-K 1300 guidelines (§ 229.1300, Item 1300), a mineral reserve is an estimate of tonnage and grade or quality of indicated and measured mineral resources that, in the opinion of the qualified person, can be the basis of an economically viable project.

 

 

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More specifically, it is the economically mineable part of a measured or indicated mineral resource, which includes diluting materials and allowances for losses that may occur when the material is mined or extracted. Further, it states that an initial assessment is required for disclosure of mineral resources but cannot be used as the basis for disclosure of mineral reserves.

“Mineral reserves” differ from “Mineral Resources” in that Mineral Reserves are known to be economically feasible for extraction. The S-K 1300 definitions require the completion of a Preliminary Feasibility Study (PFS) as the minimum prerequisite for the conversion of a portion of indicated and /or measured mineral resources to Mineral Reserves. At this time, a PFS has not been completed for the Lone Tree Project. Therefore, reserve estimates have not been made.

 

 13.

MINING METHODS

Lone Tree has historically been mined using open pit mining methods by its previous operators as reported in Zacarias, 2005 and Zacarias, 2006. Currently, there is no known plan for active mining by i-80. However, when Lone Tree decides to resume mining operations in the future, open pit mining methods would be appropriate. The mine design and operational criteria presented in this report may still be suitable for the initial start of the operation.

The bench height of 20 ft. can be used for drill and blasting. The haul road ramp grades of 10% exiting the pit and 11% for short access ramps to the bottom of the pit, and the geotechnical parameters of 40 degree slopes on the west pit walls and 45 degrees on the east pit walls will be incorporated in future pit designs.

For restarting the mining operation, the mining equipment will be of similar size and numbers to achieve the mine production needed to meet the ore milling rate and waste rock stripping requirements for Lone Tree. The equipment used in the past for open pit mining were one hydraulic shovel, a blast hole drill rig, road graders, a track dozer a rubber tire dozer, and a water truck for dust control.

A detailed mine plan with a production schedule is not required for reporting mineral resources in this report. However, this work is highly recommended for preparing the Lone Tree mine for a detailed mine study in the near future.

 

 14.

PROCESSING AND RECOVERY METHODS

i-80 possesses a heap leach, a flotation plant, an autoclave and the necessary support equipment to process both oxide and sulfidic ores. The autoclave and flotation mill are in care and maintenance. The leach pad is still producing significant amounts of gold and has a remaining capacity of 10,000,000 tons. Each process is described below.

 

 

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14.1

Heap Leach

Heap leaching is a method used to process low-grade ores that are readily amenable to cyanide leaching. The major unit operations are listed below.

   

A pad is prepared with a synthetic liner that slopes towards a pond.

   

Crushed or run-of-mine ore are stacked onto the pad.

   

The stacked ore is irrigated with a dilute cyanide solution to dissolve the gold and other precious metals.

   

The solution that drains from the stacked ore and is transferred to the pregnant solution pond.

   

The solution is pumped from the pond to a series of carbon in columns

   

The dissolved gold is loaded onto an activated carbon which is pumped counter-current to the solution flow.

   

The barren solution is refortified with cyanide and other reagents and returned to the stacked ore.

   

The carbon is pumped out of the lead CIC and trucked to a carbon stripping where the gold is stripped from the carbon and loaded onto a cathode by an electrowinning process.

   

The gold is washed off of the cathode, retorted to remove mercury, and melted into dorè bars for sale.

   

The stripped carbon is also returned to the process where it is reused to extinction. Currently i-80 contracts a company to strip the loaded carbon from Lone Tree at another location.

Since the Lone Tree operation was started in 1991, more than 1.9 million ounces of gold have been placed on the heapleach pad and 1.5 million ounces of gold have been recovered. The leach pad is still producing gold from the legacy operation and has a significant capacity to receive more tons.

 

14.2

Oxide Milling

Oxide milling is used to recover gold from high grade oxide ores that is amenable to cyanide leaching. The advantage over the heap leach process is that the ore is ground fine and subjected to intense agitation to promote faster and more complete leaching which results in higher recoveries and a faster conversion of gold to a commercial product. The disadvantage is the cost, hence the need for higher grades to support the additional efforts. The oxide milling circuit at Lone Tree was initially sized to process 2,500 short tons per day (~2,250 metric tons per day). The major unit operations are listed below.

 

   

Crushing and grinding the run of mine ore from a P80 between 150,000 and 300,000 micrometers (µ) to about 75 µ in a wet environment to create a pumpable and mixable slurry as well as liberate a large amount of the gold.

   

Pumping the slurry to a 6 stage Carbon in Leach Circuit (CIL) with a nominal 24-hour retention time.

   

Adding lime to buffer the pH above 10.0.

   

Adding cyanide to initiate the leaching process (sometimes done in the grind circuit).

 

 

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Adding coconut shell activated carbon to the 6th stage and pumping counter current to the slurry stream to facilitate efficient transfer of gold onto carbon as soon as it dissolves into solution.

   

The slurry exiting the CIL circuit is considered tailings and is pumped to a permanent storage facility.

   

The carbon is pumped from the lead CIL tank to a carbon stripping facility where the gold is stripped from the carbon and loaded onto a cathode by an electrowinning process.

   

The gold is washed off of the cathode, retorted to remove mercury and melted into dorè bars for sale.

 

14.3

Flotation

Flotation is a process used to recover gold associated with pyrite or some other floatable mineral. The process does not make a pure gold product as the leach processes did, but makes a concentrate which has a much higher gold and pyrite grade than the ore it was sourced from. The legacy flotation process at Lone Tree was designed to process 4,500 short tons/day (4,085 metric tons/day). During the life of the facility this process treated 18,000,000 tons (16,000,000 metric tons) containing 1,861,000 ounces of gold and recovered 1,497,000 ounces of gold to be processed in autoclaves or roasters. The process is summarized below.

   

Crushing and grinding the run of mine ore from a P80 between 150,000 and 300,000 micrometers (µ) to about 53 µ in a wet environment to create a pumpable and mixable slurry as well as liberate a large amount of the gold.

   

Dosing activators (CuSO4, or PbNO3),

   

Dosing collectors AERO 404 (trademarked dithiophosphate blend) and potassium amyl xanthate (PAX).

   

Dosing frothers (various light weight alcohols).

   

Transferring the ore to flotation process in a 6-stage rougher flotation circuit.

   

Adding nitrogen to the flotation cells to facilitate flotation. Note: this is a patented process developed specifically for Lone Tree trademarked as N2TEC. The nitrogen was sourced as a biproduct of the oxygen plant supporting the autoclave.

   

Transferring the tailings from flotation to the tailings dam for permanent storage. Note: These tailings were initially subjected to cyanide leaching in the same manner as the oxide mill ore but as the flotation process improved, the value of the flotation tailings decreased to a point where the CIL process was eventually shut down.

   

Cleaning the concentrate recovered from the rougher cells in a bank of four cleaning flotation cells.

   

Dewatering the concentrate for transport to an autoclave or roaster circuit for gold extraction.

   

Subjecting the concentrate to the autoclave and subsequent CIL process for gold recovery.

 

14.4

Autoclaving

Autoclaving is a process used to make gold that is locked in sulfide matrixes amenable to cyanide leaching. The autoclave is a vessel that permits the operator to subject the ore to extreme temperatures and pressures to rapidly drive the reactions that covert pyrite (FeS2) to hematite (Fe2O3).

 

 

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Once this conversion takes place, the gold is cyanide soluble. Since 1998 Lone Tree has processed 8,947,000 tons (8,117,000 metric tons) containing 2,522,000 ounces of gold and recovered 2,329,000 ounces of gold. The process is summarized below.

   

Crushing and grinding the run of mine ore from a P80 between 150,000 and 300,000 micrometers (µ) to about 75 µ in a wet environment to create a pumpable and mixable slurry as well as liberate a large amount of the gold.

   

Acidifying the ore with sulfuric acid.

   

Autoclaving the ore at temperatures between 180oC and 196oC (356oF - 385oF) and 240 to 270 pounds per square inch (1.65 – 1.86 megapascals) for about 48 minutes. These temperatures and pressures are a little lower than other autoclaves in Nevada as the gold recoveries from in excess of 90% were achieved with only 70-80% sulfur oxidation. The lower temperatures and pressures reflect that parameter.

   

Slurry cooling and depressurizing.

   

Neutralizing the slurry with lime.

   

Adding extra lime to buffer the pH above 10.0.

   

Adding cyanide to initiate the leaching process.

   

Adding coconut shell activated carbon to the 6th stage and pumping counter-current to the slurry stream to facilitate efficient transfer of gold onto carbon as soon as it dissolves into solution.

   

Pumping the tailings from the CIL circuit to a permanent storage facility.

   

Pumping the carbon from the lead CIL tank to the carbon stripping facility.

   

Stripping the gold from the carbon and loading it onto a cathode in an electrowinning process.

   

Washing the gold off of the cathodes and transferring it to a retort furnace to remove mercury.

   

Melting the remaining metal into dore bars for sale.

i-80 is considering converting the autoclave operation to an alkaline process for ores from the Granite Creek, McCoy Cove, and Ruby Hill deposits. Lone Tree ore was historically processed in an acidic environment and would most likely continue to be processed in an acidic environment. If alkaline autoclaving were the only option, a recovery reduction of about 5 percent would be expected.

 

14.5

Process Operating Costs

Operating costs (Table 14-1) were estimated for each process option available at Lone Tree selecting process options in resource declarations. They were based on current toll mill arrangements with a local facility, published operating costs from similar operations in Northern Nevada, and the 2021 Technical Report updated to account for inflation.

The autoclave values provided for 2024 were provided by i-80 based on current toll mill arrangements. Flotation costs are based on values provided in i-80-Gold-Lone-Tree-Technical Report 2021 factored for inflation based on the Producer Price Index (PPI) factors for Total Mining, Utilities and Manufacturing Industries published by the St. Louis Federal Reserve Bank and benchmarked against estimates collected from other published data for similar projects in the region. Oxide leach and mill data were referenced to published data from other Northern Nevada Mines after applying capacity correction factors.

 

 

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Table 14-1:Suggested operating Costs for Lone Tree Ores

 

Process  

Capacity

(tons/day)

 

2021 Cost

($/ton)

 

2024 Costs

($/ton)

Oxide Leach

  6,000   NA   2.91

Oxide Mill

  2,500   NA   292

Flotation

  4,800   23   29.913

Acid Autoclave

  2,500   38   44.54

Alkaline Autoclave

  2,500   NA   49

Note: Values are rounded and footnotes are defined below.

 

14.5.1

Oxide Leach Cost Factors

 

   

Lone Tree capacity = 6,000 t

   

Reference capacity = 11,500 t

   

Reference Cost = $2.00/t

   

Lone Tree Leach Costs estimate = 2.00(11,500/6,000)0.6 =$2.95/t

 

14.5.2

Oxide Mill Cost

 

   

Lone Tree capacity = 2,500 t/day

   

Reference capacity = 15,000 t/day

   

Reference Cost = $10.00/t

   

Lone Tree Leach Costs estimate = 10.00(15,000/2,500)0.6 =$29/t

 

14.5.3

Flotation Cost updated with PPI factors

 

   

Reference Cost = $23/t

   

Base Factor (July 2021) = 140.765

   

Current Factor (August 2024) = 157.743

   

Current factored cost estimate Flotation only = 23(157.743/140.765) = $26/t

   

Flotation cost including autoclave finish

 

Direct Flotation cost      $26/t
% of Flotation Feed to Concentrate      10%
Total Autoclave Costs – Grinding cost      0.10 x ($44.5 – $5.45) = $3.91/t
Total Cost of Flotation      $29.91/t

 

14.5.4

Acidic Autoclave

 

   

Acid Autoclave provided by i-80 Gold based on toll milling charges they incur at another operation.

 

 

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14.6

Model Estimate

i-80 models do not contain enough information to differentiate tonnage by process type. The nominal average process costs were estimated based on the current operating costs and the capacity for each process. All future ore is expected to be sulfidic and require either flotation or autoclaving.

 

14.7

Opinion of the Qualified Person

It is the opinion that the processing methods and costs are sufficient to support the Mineral Resource Estimate. The following comments are appropriate.

 

   

The equipment and facilities (leach pads, mills, autoclaves, leach tanks, etc.) needed to operate all of the processes proposed in the Mineral Resource Estimate are on site, have been operated in the past, and can be retrofitted to be operated again.

   

The costs are based on recent detailed studies and factored into expected current costs using reasonable approaches.

   

Additional efforts are suggested to develop a means to differentiate tonnages and grades by ore type or process destination.

 

 15.

PROJECT INFRASTRUCTURE

Refer to Section 4.

 

 16.

MARKETING STUDIES AND CONTRACTS

Marketing studies are beyond the scope of this project.

 

 17.

ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

Lone Tree is an operational site with gold production from the existing heap leach pad, an active assay laboratory, ongoing reclamation from past mining and production, environmental monitoring, and treatment of the pit lake.

 

17.1

Environmental Liabilities

Reclamation activities from past mining and processing at the Lone Tree project are ongoing. A reclamation cost estimate prepared in March 2022 estimated cost to close and reclaim the project is $87 M. This amount includes closure of all permitted mining and exploration disturbance at the project and is calculated using standardized reclamation cost estimators that assess the following:

 

   

Exploration drill hole abandonment

 

   

Exploration roads and pads

 

 

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Waste rock dumps

 

   

Heap leach pads

 

   

Roads

 

   

Pits

 

   

Foundations and buildings

 

   

Other demolition and equipment removal

 

   

Sediment and drainage control

 

   

Process ponds

 

   

Landfill

 

   

Yards

 

   

Waste disposal

 

   

Well abandonment

 

   

Miscellaneous costs

 

   

Monitoring

 

   

Construction management

 

   

Mobilization and demobilization

 

17.2

Dewatering

During the mining of the historic Lone Tree pit, dewatering operations were conducted 24 hours per day with an average daily production of 30,000 gallons per minute (gpm) dewatering at its peak was 75,000 gpm. At the end of mining operations in 2006, dewatering wells were turned off and the pit lake began to form. In March 2018, the pit lake elevation was approximately 4,307 ft and water levels continued to rebound. The pit currently functions as a sink and groundwater flow toward the Lone Tree Pit from all directions is anticipated to continue for another 40-50 years to reach equilibrium. Regional groundwater flow is also influenced by local users including agricultural producers and dewatering at the Marigold Mine.

Developing the current resource would require draining the historic Lone Tree pit and extending the cone of depression deeper than previously attempted. This represents one of the greatest challenges in advancing the current resource and will require extensive investigations into the technical, environmental, and social impacts.

 

17.3

Current Permits

Several permits are in place at the Lone Tree site. The following is a list of key permits.

 

 

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Plan of Operations (BLM)

 

   

Nevada State Fire Marshal Hazardous Materials Permit No. 112115 valid until Feb 28, 2025)

 

   

Onsite Sewage Disposal System General Septic Permit (NDEP - BWPC)

 

   

General Permit for Stormwater Discharges Associated with Industrial Activity from Metals Mining Activities (NDEP - BWPC)

 

   

Permits for heap leach facility and tailings decant pond (NDOW)

 

   

Permit to Operate a Public Water System (NDEP - BSDW)

 

   

Water Pollution Control Permits (NDEP - BMRR)

 

   

Reclamation Permit (NDEP - BMRR)

 

   

Tailings Dam Construction and Safety Permit (Nevada Division of Water Resources)

 

   

Class II Air Quality Operating Permits (NDEP - BAPC)

 

   

Class III Waivered Landfill Landfill No. 3 (NDEP – BSMM)

 

   

Two LPG Licenses (Nevada Liquefied Petroleum Gas Board)

 

   

The Mercury Operating Permit to Construct (NDEP-BAPC).

No major environmental study has been conducted by i-80 to address various liabilities including dewatering.

 

 18.

CAPITAL AND OPERATING COSTS

Detailed studies have not been carried out to provide this information. However, various cost factors have been assumed for creating an optimized pit shell to meet the Reasonable Prospect for Economic Extraction criteria. These cost factors have been disclosed in Section 11.

 

 19.

ECONOMIC ANALYSIS

A detailed economic analysis is beyond the scope of this project.

 

 20.

ADJACENT PROPERTIES

Most of the mineral rights surrounding Lone Tree are owned or controlled by NGM. The area’s only other active mining project is Marigold, operated by SSR Mining. There are several inactive mines and exploration or development projects in the area. Other smaller deposits in the area are Trenton Canyon (Figure 20-1) & North Peak and Converse Complex. Mineral Resources for area properties are summarized in Table 20-1.

 

 

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Table 20-1:Mineral Resources for Nearby Properties

 

Property   Owner  

Produced

(Au Oz)

 

Stated

Mineral

Reserves

(M. Au Oz)

 

Stated

Measured

and

Indicated
Mineral
Resources

(M.Au Oz)

 

Stated

Inferred

Mineral
Resources

(M. Au Oz)

Marigold1   SSR   --   3.19   5.66   0.63

Buffalo Valley

Complex2

  Newmont   39,688   n/a   0.47   Unknown
Trenton Canyon & North Peak3   Newmont   n/a   n/a   n/a   n/a
Converse4   Waterton   --   n/a   6.10   0.59

Notes:

 

  1.

SSR, 2017, Marigold Mine, NI 43-101 Technical Report.

  2.

Newmont, 2014; Newmont’s 2013 Annual Report filed February 20, 2014.

  3.

The Nevada Mineral Industry, 2012; Nevada Bureau of Mines and Geology Special Publication MI-2012 less Valmy which was purchased by SSR Mining in 2015.

  4.

Chaparral Gold, October 21, 2014; website, deposit sold to Waterton Global Resource Management in 2014.

 

 

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Figure 20-1:Mineral Deposits Adjacent to Lone Tree A mine grid coordinate system was used by Newmont/ NGM at the Lone Tree deposit for drilling.

 

 

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 21.

OTHER RELEVANT DATA AND INFORMATION

The drill hole collars use a mine grid coordinate system and are mapped using a GPS. The GPS has 0.4-inch (1 cm) accuracy. From 1985 through 1998, a Topcon total station instrument was used, accurate to within five seconds of a degree.

Each of the surveyed data points were recorded digitally and manually, before it is saved in an acQuire™ Data Management data-storage system

 

 22.

TECHNICAL REPORT INTERPRETATION AND CONCLUSIONS

 

  1.

Geological Interpretations: As per the definition of the Inferred resources, geological continuity can be assumed. However, in this Lone Tree resource estimation project, contact analyses proved that the geological units interpreted by NGM are reasonable; Various geological units (as designated and grouped by NGM) had different levels of favorability for mineralization (different AuFA grade distributions).

 

  2.

The RPEE criteria: The resource classification is based on a pit shell that uses operational parameters Section 11.12

 

   

Inventory of all blocks below the current pit limit and within 2,175 pit shell is classified as inferred resources and a subset of these blocks within 50 feet from the current pit surface are classified as indicated blocks.

 

 23.

RECOMMENDATIONS

 

23.1

Resource Model update

The resource estimates presented in this report are valid in a deposit scale, which may be appropriate for long-term mine planning. However, prior to mine production a resource model should be updated using more detailed data analyses in order to achieve the accuracy required at the scale of weekly or monthly production.

 

23.2

Risk Analyses in Resource Eestimates

A simulation-based resource model with risk factors built into it may be beneficial for estimating risks and opportunities for future production. Such an approach may also be useful for strategic exploration planning.

 

23.3

Future Exploration

The Lone Tree deposit provides a potential for improving currently inferred category resources into the indicated category and additional inferred resources through infill drilling.

 

 

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These infill drilling programs should be strategically designed to maximize benefits of drilling. An advanced geostatistical approach to strategic drill hole planning is advised.

Lone Tree has the potential for substantial mineral resources in the Sequoia area. This is the current drill hole intercepts as shown in Figure 23-1. Further deep drilling in the Sequoia zone has the potential to add more mineral resources to the project.

Figure 23-1:A Vertical Cross Section (Looking East) Along East 83700 North-South

 

 

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23.4

Geometallurgical Study

It is highly recommended to develop a geometallurgical model of the Lone Tree deposit by drilling core and RC holes throughout the deposit and analysing the rocks for geometallurgical characterization. This model, when complete, should be included in the future resource and reserve estimates.

 

 

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 24.

REFERENCES

 

  1.

Autoclave Restart Feasibility – Final Report_2023Oct.pdf

 

  2.

Bloomstein, E., Braginton, B., Owen, R., Parratt, R., Thompson, W., 1991, Geology and geochemistry of the Lone Tree gold deposit, Humboldt County, Nevada

 

  3.

Carter, G., 2003, GCarter_assay_data, Inter-company report and memo

 

  4.

Carter, G., 2003, GCarter_data_assay_drilling-5Star- Inter-company report and memo

 

  5.

Cox, D.P., 1992, Descriptive model of distal-disseminated Ag-Au, in Bliss, J.D., ed., Developments in mineral deposits modeling: U. S. Geological Survey Bulletin 2004, p. 19.

 

  6.

Cole, J.A., Lenz, J., C., Janhunen, W., J., (1995) “One Year of Pressure Oxidation at Lone Tree Gol Mine”, Mine Engineering, June 1995, article No 300-000-109-769.

 

  7.

Doebrich, J.L., and Theodore, T.G., 1996, Geologic history of the Battle Mountain mining district, Nevada, and regional controls on the distribution of mineral systems, in Coyner, A.L., and Fahey, P.L., eds., Geology and ore deposits of the American Cordillera: Geological Society of Nevada Symposium Proceedings, Reno/Sparks, Nevada, April 1955, v. 1, p. 453-483.

 

  8.

Fisher (2003): Lone Tree Hydrology (Jay Fischer), 2003 Drawdown performance vs pumping with modeled pumping to achieve LOM drawdown requirements

 

  9.

Fithian, M.T., Holley, E.A., and Kelly, N.M., 2018, Geology of gold deposits at the Marigold mine, Battle Mountain district, Nevada: Reviews in Economic Geology, v. 20, p. 121–155.

 

  10.

Holley, E.A., Lowe, J.A., Johnson, C.A., Pribil, M.J., 2019. Magmatic-hydrothermal gold mineralization at the Lone Tree Mine, Battle Mountain, Nevada. Economic Geology, v. 114, p. 811-856. https://doi.org/10.5382/econgeo.4665.

 

  11.

https://fred.stlouisfed.org/series/PCUAMUMAMUM

 

  12.

HYDRO-SEARCH, INC., 1991, Predicted Post Mining Hydrogeochemistry of the Lone Tree Pit, Valmy, Nevada, Hydro-Search, Inc., Reno, NV. Prepared for Santa Fe Pacific Gold.

 

  13.

Hydrologic Consultants, Inc (HCI) December 2002, Update of Numerical Ground-Water Flow

 

  14.

Kunkel, K., 1998, Idealized Stratigraphic Column of Lone Tree Area, Internal stratigraphy interpretation

 

  15.

Lowe, J. 2019; Petrographic, geochemical, and geochronological investigation of gold mineralization at the Lone Tree gold mine, Battle Mountain, Nevada, MS Thesis, Colorado School of Mines. (https://hdl.handle.net/11124/172843)

 

 

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  16.

Mosier, D.L., Singer, D.A., Bagby, W.C., and Menzie, W.D., 1992, Grade and tonnage model of sediment-hosted Au, in Bliss, J.D., ed., Developments in mineral deposits modeling: U. S. Geological Survey Bulletin 2004, p. 26.

 

  17.

Modeling for Newmont Mining Corporation’s Lone Tree Mine Humboldt County, Nevada

 

  18.

Mosier, D.L., Singer, D.A., Bagby, W.C., and Menzie, W.D., 1992, Grade and tonnage model of sediment-hosted Au, in Bliss, J.D., ed., Developments in mineral deposits modeling: U. S. Geological Survey Bulletin 2004, p. 26.

 

  19.

Moss, K., Saderholm, E., 2003 Lone Tree Density Report 03, Inter-company report

 

  20.

Moss, K., 2003, Lone Tree Model vs. Production Check, Intra-company report

 

  21.

Muntean, J. L., and Cline. J. S, 2018, Diversity in Carlin-Style Gold Deposits, Reviews in Economic Geology, v. 20, pp. 1–5

 

  22.

Nevada Gold Mines LLC – Carlin Complex Technical Report NI 43-101 – March 25, 2020

 

  23.

PTI Environmental Services, 1995, Assessment of Pit Lake Chemogenesis and Waste-rock.

 

  24.

Raabe, K., 1995, The Lone Tree Extension Project, Humboldt County, Nevada

 

  25.

Ressel, M.W., Jr., 2005, Igneous geology of the Carlin trend, Nevada—The importance of Eocene magmatism in gold mineralization: Reno, Nev., University of Nevada Ph.D. thesis, 266 p.

 

  26.

Ressel, M.W., Jr., 2005, Igneous geology of the Carlin trend, Nevada—The importance of Eocene magmatism in gold mineralization: Reno, Nev., University of Nevada Ph.D. thesis, 266 p.

 

  27.

Samal, A., R., 2021, Technical Report on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada, NI 43-101 TECHNICAL REPORT, 10-21-2021.

 

  28.

Theodore, T.G., 1998, Pluton-related Au in the Battle Mountain mining district _ An overview, in Tosdal, R.M., ed., Contributions to the Gold Metallogeny of Northern Nevada: U.S. Geological Survey Open-File Report 98-338, p. 251-252.

 

  29.

Simon Hydro-Search, December 10, 1991, Preliminary Hydrogeology Evaluation Lone Tree Mine Valmy, Nevada

 

  30.

Simon Hydro-Search, May 23, 1991, Results of field Program and Dewatering Model Lone Tree Mine Valmy, Nevada

 

  31.

Simmons, G., L. (1997) “Flotation of Auriferous Pyrite using Santa Fe Gold’s N2TEC Flotation Process,” SME Annual Meeting, February 24-27, 1997, SME article No. 300-000-122-384.

 

 

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  32.

Simmons, G., L., Orlich, J., N., Lenz, J., C., Cole, J., A., (1999) “Implementation and Start-up of N2TEC Flotation at the Lone Tree Mine,” Society for Mining and Metallurgy Exploration, January 1999. SME article No 300-000-114-786.

 

  33.

Simon Hydro-Search, August 20, 1992, Expanded Dewatering Ground-Water Flow Model Lone Tree Mine Valmy, Nevada

 

  34.

Theodore, T.G., 1998, Pluton-related Au in the Battle Mountain mining district _ An overview, in Tosdal, R.M., ed., Contributions to the Gold Metallogeny of Northern Nevada: U.S. Geological Survey Open-File Report 98-338, p. 251-252.

 

  35.

Turner, D., 2003, 2003 Lone Tree Resource Model and Reserve Document, Inter-company report from Consultant Dean Turner, Denver, Colorado

 

  36.

Turner, D., 2003, Summary of Lone Tree Exploration Drill Sample Bias Study, Inter-company report from Consultant Dean Turner, Denver, Colorado

 

  37.

Wallace, A.R., Ludington, S., Mihalasky, M.J., Peters, S.G., Theodore, T.G., Ponce, D.A., John, D.A., and Berger, B.R., 2004, Assessment of metallic mineral resources in the Humboldt River basin, northern Nevada: U.S. Geological Survey Bulletin 2218, 309 p.

 

  38.

Welsh Engineering Inc and Lone Tree Mining Inc, 1991, Ore and Waste Material Characterization Project

 

  39.

Wallace, A.R., Ludington, S., Mihalasky, M.J., Peters, S.G., Theodore, T.G., Ponce, D.A., John, D.A., and Berger, B.R., 2004, Assessment of metallic mineral resources in the Humboldt River basin, northern Nevada: U.S. Geological Survey Bulletin 2218, 309 p.

 

  40.

Zacarias, P. A., 2006, Mineral Resource and Ore Reserve Report as of December 31, 2005, Lone Tree Complex Mine/Project, February 28, 2006

 

 

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 25.

RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

This report has been prepared by the Qualified Persons (QPs) for i-80 Gold (the Registrant). The QPs have relied on the following information provided by the Registrant to inform this Technical Report Summary (TRS).

 

  i.

The gold price as used by i-80 for government filings.

   

The QPs have reviewed the gold price of $2,175.00 used in the pit shell with representatives of the registrant. The chosen price does not materially exceed financial institution forecasts or the two-year trailing average price. More details are available in the section 11.10.

  ii.

The costs of mining operations for mining the rock and fill materials.

   

Costs of mining operations for rock and fill materials were based on i-80’s compilation of recent contracts and mining costs in Nevada. These numbers align with the QP’s experience and knowledge of current costs.

  iii.

i-80 is the owner of the information for land, site infrastructure (Refer to Section 3), data topography, drill-hole data, and geochemical data used in the resource estimation. The geochemical analysis of the re-assay exercise was done by ALS Laboratories and managed by i-80 (Refer to Section 9). The QPs have reviewed this information and are satisfied that it is accurate and relevant to this project.

  iv.

As the owner of the Lone Tree property, i-80 has the information discussed in Section 4 on accessibility, climate, local resources, infrastructure and physiography. The QPs have reviewed the information used in this section and are satisfied that it is accurate and relevant to this project.

  v.

Similarly, the information provided in Section 17 was compiled by i-80 team members as the owner of the information and was reviewed by the QPs. QPs find the information provided in this section correct and relevant to the project.

  vi.

The information used in Section 21 was provided by the i-80 team. The QPs have reviewed the information and found it satisfactory for the purpose of this project.

 

 

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 26.

QUALIFIED PERSONS CERTIFICATES

Certificate of the Qualified Person

ABANI R. SAMAL, RM – SME

I, Abani R Samal, as an author of this report entitled “S-K 1300 Technical Report Summary on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada” (TRS) with an effective date of December 31, 2024 prepared for Premier Gold, a wholly owned subsidiary of i-80 Gold (Registrant), do hereby certify that:

 

I am the Principal of GeoGlobal LLC (GeoGlobal) – a Utah-based consulting company.

I graduated with a PhD degree in 2005 from the Southern Illinois University, Carbondale Illinois where I studied mineral deposits extensively. My PhD research was focused on Florida Canyon gold deposit, Nevada.

I am also a graduate of the Imperial College, London (2000) with Master’s degree in Mineral Exploration.

I am a registered member of the Society for Mining, Metallurgy & Exploration (SME) with membership number as 4136879.

I have worked as a geologist for over 20 years since 1996. I have broad experience in various commodities including gold-silver, base metals and uranium where mineralization is structurally and lithologically controlled similar and comparable to the control of mineralization at the Lone Tree Gold deposit.

I have relevant experience for the purpose of this Technical Report Summary (TRS).

I have more than 15 years of experience in Mineral Resource Estimation and applied geostatistics.

I was the lead author of the NI 43-101 technical report published in 2021 (2021 report) with title Technical Report on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada (Effective Date: July 30 2021, Report date: October 20 2021).

I visited the Lone Tree deposit and verified the digital data and drill-core-logs at the core-shed and core-processing facility at Battle Mountain, Nevada in 2021. I also checked the data and ordered a re-assay of selected drill core intersections in August 2024.

I have read the definition of “qualified person” set out in § 229.1302 (Item 1302)4 Qualified person, technical report summary, and technical studies and certify that by reason of my education, affiliation with a professional association and relevant work experiences, I fulfill the requirements to be a “Qualified Person” in compliance with 17 CFR § 229.1302 (b)(1)(i) and (ii)4 qualified person definition.

 

4 Title 17—Commodity and Securities Exchanges, Chapter II—Securities and Exchange Commission, Part 229—Standard Instructions for Filing Forms Under Securities Act of 1933, Securities Exchange Act of 1934 and Energy Policy and Conservation Act of 1975—Regulation S-K At the effective date, to the best of my knowledge, information, and belief, this Technical Report Summary contains all scientific and technical information that is required to be disclosed to make the Technical Report Summary not misleading.

 

 

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GeoGlobal was engaged by i-80 to review the resource estimates published in the 2021 report and update as required based on current market conditions and produce this report conforming to the United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 229.1300 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300) and Item 601 (b)(96) Technical Report Summary.

I am an employee of GeoGlobal LLC and independent of the registrant to prepare this report for the registrant.

Report sections for which I am responsible for: 1, 2, 3, 4, 5, 6, 7, 8, 9, 11, 15, 17, 21, 22, 23, 24, 25, 26

Signed

Abani R Samal

RM-SME #04136879

March 24, 2025

 

 

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Certificate of the Qualified Person

Brian Arthur, RM – SME

I, Brian W. Arthur, as an author of this report entitled “S-K 1300 Technical Report Summary on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada” (TRS) ), with an effective date of December 31, 2024, prepared for Premier Gold, a wholly owned subsidiary of i-80 Gold (Registrant), do hereby certify that:

 

I am the Principal Brian Arthur Consulting Metallurgy LLC – a Nevada-based consulting company.

I graduated with Bachelors of Science degree in Metallurgical Engineering in 1985 and a Master of Science degree in Metallurgical Engineering in 1987 from Montana College of Mineral Science and Technology, Butte Montana.

I am also a graduate of the University of Nevada with a Master of Business Administration in 2007.

I am a registered member of the Society for Mining, Metallurgy & Exploration (SME) with membership number as 93800RM.

I have worked as a metallurgist for over 35 years since 1987. I have broad experience in metallurgical process used to recover various commodities including gold, silver, copper and zinc.

I have relevant experience for the purpose of this Technical Report Summary (TRS)

I visited the Lone Tree deposit and verified the condition of the process equipment and assay laboratory in August 2024.

I have no prior encounters with i80 Gold.

I have not written or contributed to any prior technical reports for the Lone Tree Project.

I have read the definition of “qualified person” set out in § 229.1302 (Item 1302)5 Qualified person, technical report summary, and technical studies and certify that by reason of my education, affiliation with a professional association and relevant work experiences, I fulfill the requirements to be a “Qualified Person” in compliance with 17 CFR § 229.1302 (b)(1)(i) and (ii)4 qualified person definition.

I am an independent contractor (Associate) for GeoGlobal LLC and independent of the registrant to prepare this report for the registrant.

I am responsible for sections 10 and 14 and contributed to Section 4.

At the effective date, to the best of my knowledge, information, and belief, the Technical Report Summary contains all scientific and technical information that is required to be disclosed to make this Technical Report Summary not misleading.

Signed

Brian Arthur,

RM-SME #93800RM

March 24, 2025

 

 

5 Title 17—Commodity and Securities Exchanges, Chapter II—Securities and Exchange Commission, Part 229—Standard Instructions for Filing Forms Under Securities Act of 1933, Securities Exchange Act of 1934 and Energy Policy and Conservation Act of 1975—Regulation S-K I, Paul A Gates, as an author of this report entitled “S-K 1300 Technical Report Summary on the Mineral Resource Estimates for the Lone Tree Deposit, Nevada” (TRS) with an effective date of December 31, 2024, prepared for Premier Gold, a wholly owned subsidiary of i-80 Gold (Registrant) do hereby certify that:

 

 

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Certificate of the Qualified Person

Paul A Gates, P.E., Member SME

 

I am an associate of GeoGlobal LLC (GeoGlobal) – a Utah-based consulting company.

I graduated with a B.S. degree in 1984 from the Montana Technical University, Butte Montana where I studied mining and mineral economics of mineral deposits.

I am a graduate of Western New Mexico College, of Silver City New Mexico with a Master’s degree in Business Administration (1997).

I am a Registered Professional Engineer in the State of Colorado. PE -#0043794

I am a member of the Society for Mining, Metallurgy & Exploration (SME).

I have worked as a mining engineer for over 39 years since 1985. I have broad experience in various commodities including gold-silver, base metals, uranium and coal.

I have relevant experience for this Technical Report Summary (TRS).

I have more than 35 years of experience in Mineral Resource Estimation, Pit optimization and Strategic mine planning.

I have read the definition of “qualified person” set out in § 229.1302 (Item 1302)6 Qualified person, technical report summary, and technical studies and certify that because of my education, affiliation with a professional association and relevant work experiences, I fulfill the requirements to be a “Qualified Person” in compliance with 17 CFR § 229.1302 (b)(1)(i) and (ii)4 qualified person definition.

I am an associate independent contractor (Associate) of GeoGlobal LLC and independent of the registrant to prepare this report for the registrant.

I am responsible for section 13 and I contributed to sections 1, 2 and 11.

At the effective date, to the best of my knowledge, information, and belief, the Technical Report Summary contains all scientific and technical information that is required to be disclosed to make the Technical Report Summary not misleading.

Paul A Gates

P.E.- #0043794

March 24, 2025

 

6 Title 17—Commodity and Securities Exchanges, Chapter II—Securities and Exchange Commission, Part 229—Standard Instructions for Filing Forms Under Securities Act of 1933, Securities Exchange Act of 1934 and Energy Policy and Conservation Act of 1975—Regulation S-K Table 27-1:Lone Tree Unpatented Claims, Claimant: Goldcorp Dee LLC

 

 

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 27.

APPENDIX A:DETAILS OF THE LAND CLAIMS OF THE LONE TREE PROJECT.

 

LONE TREE PROJECT
Claimant: Goldcorp Dee LLC

220 Claims

Humboldt County, Nevada

 Claim Count      Claim Name    BLM Serial Number   

BLM Legacy

Number

1   

RP # 1

  

NV101496979

  

NMC349047   

2   

RP # 2

  

NV101403912

  

NMC349048

3   

RP # 3

  

NV101345613

  

NMC349049

4   

RP # 4

  

NV101600986

  

NMC349050

5   

RP # 5

  

NV101508065

  

NMC349051

6   

RP # 6

  

NV101348212

  

NMC349052

7   

RP # 7

  

NV101509406

  

NMC349053

8   

RP # 8

  

NV101521081

  

NMC349054

9   

RP # 9

  

NV101605977

  

NMC349055

10   

RP # 10

  

NV101525928

  

NMC349056

11   

RP # 11

  

NV101346868

  

NMC349057

12   

RP # 12

  

NV101349446

  

NMC349058

13   

RP # 13

  

NV101605069

  

NMC349059

14   

RP # 14

  

NV101604954

  

NMC349060

15   

RP # 15

  

NV101492078

  

NMC349061

16   

RP # 16

  

NV101605518

  

NMC349062

17   

RP # 17

  

NV101460048

  

NMC349063

18   

RP # 18

  

NV101758271

  

NMC349064

19   

VAL #163

  

NV101491690

  

NMC361074

20   

VAL #164

  

NV101345767

  

NMC361075

21   

VAL #165

  

NV101497337

  

NMC361076

22   

VAL #166

  

NV101361807

  

NMC361077

23   

VAL #167

  

NV101603873

  

NMC361078

24   

VAL #168

  

NV101454881

  

NMC361079

25   

VAL 169

  

NV101504448

  

NMC361080

26   

VAL #170

  

NV101454919

  

NMC361081

27   

VAL #171

  

NV101503237

  

NMC361082

28   

VAL #172

  

NV101454995

  

NMC361083

29   

VAL #173

  

NV101408718

  

NMC361084

30   

VAL #174

  

NV101504467

  

NMC361085

31   

VAL #175

  

NV101403947

  

NMC361086

32   

VAL #176

  

NV101504526

  

NMC361087

33   

VAL #177

  

NV101406561

  

NMC361088

34   

VAL #178

  

NV101454077

  

NMC361089

35   

VAL #179

  

NV101305311

  

NMC361090

36   

VAL #180

  

NV101451735

  

NMC361091

37   

VAL #181

  

NV101302082

  

NMC361092

38   

VAL #182

  

NV101452924

  

NMC361093

39   

VAL #183

  

NV101303298

  

NMC361094

 

 

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    40       

VAL #184

  

NV101730884

  

NMC361095    

41   

VAL #185

  

NV101303326

  

NMC361096

42   

VAL #186

  

NV101609536

  

NMC361097

43   

VAL #187

  

NV101303411

  

NMC361098

44   

VAL #188

  

NV101494109

  

NMC361099

45   

VAL #201

  

NV101477133

  

NMC361100

46   

VAL #203

  

NV101301202

  

NMC361102

47   

VAL #204

  

NV101500691

  

NMC361103

48   

VAL #205

  

NV101526503

  

NMC361104

49   

VAL #206

  

NV101609874

  

NMC361105

50   

VAL #207

  

NV101523468

  

NMC361106

51   

VAL #208

  

NV101505616

  

NMC361107

52   

VAL #209

  

NV101303248

  

NMC361108

53   

VAL #210

  

NV101608604

  

NMC361109

54   

VAL #211

  

NV101730776

  

NMC361110

55   

VAL #212

  

NV101491549

  

NMC361111

56   

VAL #213

  

NV101732055

  

NMC361112

57   

VAL #214

  

NV101492636

  

NMC361113

58   

VAL #215

  

NV101780894

  

NMC361114

59   

VAL #216

  

NV101610238

  

NMC361115

60   

VAL #217

  

NV101780999

  

NMC361116

61   

VAL #218

  

NV101605723

  

NMC361117

62   

VAL #223

  

NV101497576

  

NMC361122

63   

VAL #224

  

NV101459234

  

NMC361123

64   

VAL #225

  

NV101300539

  

NMC361124

65   

VAL #226

  

NV101494548

  

NMC361125

66   

VAL #227

  

NV102521080

  

NMC361126

67   

VAL #228

  

NV101522020

  

NMC361127

68   

VAL #229

  

NV101404418

  

NMC361128

69   

VAL #230

  

NV101505643

  

NMC361129

70   

VAL #231

  

NV101405177

  

NMC361130

71   

VAL #232

  

NV101509280

  

NMC361131

72   

VAL #233

  

NV101402307

  

NMC361132

73   

VAL #234

  

NV101343113

  

NMC361133

74   

VAL #235

  

NV101409380

  

NMC361134

75   

VAL #236

  

NV101455249

  

NMC361135

76   

VAL #289

  

NV101405725

  

NMC361162

77   

SAR# 1

  

NV101459301

  

NMC373613

78   

SAR# 2

  

NV102521081

  

NMC373614

79   

SAR# 3

  

NV101522021

  

NMC373615

80   

SAR# 4

  

NV101303578

  

NMC373616

81   

SAR# 5

  

NV101609005

  

NMC373617

82   

SAR# 6

  

NV101403623

  

NMC373618

83   

SAR# 7

  

NV101504496

  

NMC373619

84   

SAR# 8

  

NV101402308

  

NMC373620

85   

SAR# 9

  

NV101343114

  

NMC373621

86   

SAR# 10

  

NV101402578

  

NMC373622

87   

SAR# 11

  

NV101343253

  

NMC373623

88   

SAR# 12

  

NV101405750

  

NMC373624

89   

SAR# 13

  

NV101454317

  

NMC373625

 

 

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    90       

SAR# 14

  

NV101406768

  

NMC373626    

91   

SAR# 15

  

NV101456564

  

NMC373627

92   

SAR# 16

  

NV101401785

  

NMC373628

93   

SAR# 17

  

NV101730452

  

NMC373629

94   

SAR# 18

  

NV101731148

  

NMC373630

95   

SAR# 19

  

NV101491560

  

NMC373631

96   

SAR# 20

  

NV101732064

  

NMC373632

97   

SAR# 21

  

NV101491111

  

NMC373633

98   

SAR# 22

  

NV101548834

  

NMC373634

99   

SAR# 23

  

NV101610246

  

NMC373635

100   

SAR# 24

  

NV101478218

  

NMC373636

101   

SAR# 25

  

NV101606462

  

NMC373637

102   

SAR# 26

  

NV101731079

  

NMC373638

103   

SAR# 27

  

NV101344521

  

NMC373639

104   

SAR# 28

  

NV101460156

  

NMC373640

105   

SAR# 29

  

NV101343304

  

NMC373641

106   

SAR# 30

  

NV101459185

  

NMC373642

107   

SAR# 31

  

NV101459238

  

NMC373643

108   

SAR# 32

  

NV101300548

  

NMC373644

109   

SAR# 33

  

NV101495122

  

NMC373645

110   

SAR# 34

  

NV102521091

  

NMC373646

111   

SAR# 35

  

NV101522028

  

NMC373647

112   

SAR# 36

  

NV101405026

  

NMC373648

113   

LTH # 1

  

NV101479414

  

NMC390406

114   

LTH # 2

  

NV101496685

  

NMC390407

115   

LTH # 3

  

NV101520523

  

NMC390408

116   

LTH # 4

  

NV101609543

  

NMC390409

117   

LTH # 5

  

NV101758246

  

NMC390410

118   

LTH # 6

  

NV101528225

  

NMC390411

119   

LTH # 7

  

NV101543630

  

NMC390412

120   

LTH # 8

  

NV101500667

  

NMC390413

121   

VG # 1

  

NV101525297

  

NMC397890

122   

VG # 2

  

NV101604536

  

NMC397891

123   

VG # 3

  

NV101521646

  

NMC397892

124   

VG # 4

  

NV101751245

  

NMC397893

125   

VG # 5

  

NV101490570

  

NMC397894

126   

VG # 6

  

NV101540752

  

NMC397895

127   

VG # 7

  

NV101454021

  

NMC397896

128   

VG # 8

  

NV101525066

  

NMC397897

129   

VG # 9

  

NV101525236

  

NMC397898

130   

VG # 10

  

NV101604022

  

NMC397899

131   

VG # 11

  

NV101495588

  

NMC397900

132   

VG # 12

  

NV101752969

  

NMC397901

133   

VG # 13

  

NV101526207

  

NMC397902

134   

VG # 14

  

NV101452105

  

NMC397903

135   

VG # 15

  

NV101605328

  

NMC397904

136   

VG # 16

  

NV101453790

  

NMC397905

137   

VG # 17

  

NV101340407

  

NMC397906

138   

VG # 18

  

NV101480346

  

NMC397907

139   

SH # 1

  

NV101604023

  

NMC404603

 

 

106


LOGO

 

 

    140       

SH # 2

  

NV101349400

  

NMC404604    

141   

SH # 3

  

NV101546039

  

NMC404605

142   

SH # 4

  

NV101496217

  

NMC404606

143   

SH # 5

  

NV101543490

  

NMC404607

144   

SH # 6

  

NV101454745

  

NMC404608

145   

SH # 7

  

NV101458840

  

NMC404609

146   

SH # 8

  

NV101302889

  

NMC404610

147   

SH # 9

  

NV101456153

  

NMC404611

148   

SH # 10

  

NV101301724

  

NMC404612

149   

SH # 11

  

NV101459540

  

NMC404613

150   

SH # 12

  

NV101349821

  

NMC404614

151   

SH # 13

  

NV101600450

  

NMC404615

152   

SH # 14

  

NV101303622

  

NMC404616

153   

SH # 15

  

NV101601671

  

NMC404617

154   

SH # 16

  

NV101303045

  

NMC404618

155   

SH # 17

  

NV101542214

  

NMC404619

156   

SH # 18

  

NV102520792

  

NMC404620

157   

SH # 19

  

NV101546285

  

NMC404621

158   

SH # 20

  

NV101349935

  

NMC404622

159   

SH # 21

  

NV101453263

  

NMC404623

160   

SH # 22

  

NV101491961

  

NMC404624

161   

SH # 23

  

NV101544749

  

NMC404625

162   

LONE TREE 1

  

NV101609891

  

NMC587843

163   

LONE TREE 2

  

NV101347181

  

NMC587844

164   

LONE TREE 3

  

NV101508006

  

NMC587845

165   

LONE TREE 4

  

NV101347905

  

NMC587846

166   

LONE TREE 5

  

NV101494818

  

NMC587847

167   

LONE TREE 6

  

NV101349811

  

NMC587848

168   

LONE TREE 7

  

NV101454278

  

NMC587849

169   

LONE TREE 8

  

NV101495094

  

NMC587850

170   

LONE TREE 9

  

NV101601633

  

NMC587851

171   

LONE TREE 10

  

NV101460325

  

NMC587852

172   

LONE TREE 12

  

NV101495848

  

NMC587853

173   

LONE TREE 13

  

NV101605491

  

NMC587854

174   

LONE TREE 14

  

NV101496461

  

NMC587855

175   

LONE TREE 18

  

NV101479697

  

NMC587859

176   

LONE TREE 20

  

NV101601778

  

NMC587861

177   

LONE TREE 22

  

NV101503313

  

NMC587863

178   

LONE TREE 23

  

NV101404525

  

NMC587864

179   

LONE TREE 24

  

NV101508264

  

NMC587865

180   

LONE TREE 25

  

NV101406121

  

NMC587866

181   

LONE TREE 26

  

NV101453787

  

NMC587867

182   

LONE TREE 27

  

NV101340403

  

NMC587868

183   

RPA 1

  

NV101348284

  

NMC591404

184   

RPA 2

  

NV101604347

  

NMC591405

185   

SH FRACTION # 1

  

NV101456391

  

NMC593447

186   

SH FRACTION # 2

  

NV101401762

  

NMC593448

187   

SH FRACTION # 3

  

NV101609558

  

NMC593449

188   

VAL #1001

  

NV101456079

  

NMC600381

189   

VAL #1002

  

NV101525018

  

NMC600382

 

 

107


LOGO

 

 

    190       

VAL #1003

  

NV101456108

  

NMC600383    

191   

VAL #1004

  

NV101603869

  

NMC600384

192   

VAL #1005

  

NV101454878

  

NMC600385

193   

VAL #1006

  

NV101608925

  

NMC600386

194   

VAL #1007

  

NV101457882

  

NMC600387

195   

VAL #1008

  

NV101490766

  

NMC600388

196   

VAL #1009

  

NV101755290

  

NMC600389

197   

VAL #1010

  

NV101491106

  

NMC600390

198   

VAL 205A

  

NV101548952

  

NMC662859

199   

VAL 206A

  

NV101480169

  

NMC662860

200   

VAL 207A

  

NV101547608

  

NMC662861

201   

VAL 208A

  

NV101303434

  

NMC662862

202   

VAL 221A

  

NV101304705

  

NMC662864

203   

VAL 223A

  

NV101454639

  

NMC662865

204   

VAL 225A

  

NV102520402

  

NMC662866

205   

VAL 227A

  

NV101478899

  

NMC662867

206   

VAL 229A

  

NV101301693

  

NMC662868

207   

VAL 231A

  

NV101479591

  

NMC662869

208   

VAL 233A

  

NV101302577

  

NMC662870

209   

VAL 235A

  

NV101547229

  

NMC662871

210   

VAL 290A

  

NV102521542

  

NMC662872

211   

LONE TREE NO 100

  

NV101525645

  

NV101525645

212   

LONE TREE NO 101

  

NV101456984

  

NV101456984

213   

LONE TREE NO 102

  

NV101526938

  

NV101526938

214   

LONE TREE NO 103

  

NV101491286

  

NV101491286

215   

LONE TREE NO 104

  

NV101341888

  

NV101341888

216   

LONE TREE NO. 200

  

NV101454275

  

NMC681251

217   

LONE TREE NO. 202

  

NV101457596

  

NMC681253

218   

LONE TREE NO. 203

  

NV101602410

  

NMC681254

219   

LONE TREE NO. 204

  

NV101452277

  

NMC681255

220   

LONE TREE NO. 205

  

NV101605841

  

NMC681256

 Table 27-2:Claimant: VEK/ANDRUS ASSOCIATES & Goldcorp Dee LLC

Claimant: GOLDCORP DEE LLC & VEK/ANDRUS ASSOCIATES

Lessee: Goldcorp Dee LLC

38 Claims

Humboldt County, Nevada

 Claim Count      Claim Name    BLM Serial Number   

BLM Legacy

Number

    1       

VAL # 73

  

NV101454058

  

NMC298449    

2   

VAL # 74

  

NV101609240

  

NMC298450

3   

VAL # 75

  

NV101451611

  

NMC298451

4   

VAL # 76

  

NV101605860

  

NMC298452

5   

VAL # 77

  

NV101452602

  

NMC298453

6   

VAL # 78

  

NV101404509

  

NMC298454

7   

VAL # 79

  

NV101609812

  

NMC298455

8   

VAL # 80

  

NV101405658

  

NMC298456

9   

VAL # 81

  

NV101606641

  

NMC298457

10   

VAL # 82

  

NV101406701

  

NMC298458

11   

VAL # 83

  

NV101495870

  

NMC298459

 

 

108


LOGO

 

 

    12       

VAL # 84

  

NV101301810

  

NMC298460    

13   

VAL # 85

  

NV101491523

  

NMC298461

14   

VAL # 86

  

NV101303230

  

NMC298462

15   

VAL # 87

  

NV101492627

  

NMC298463

16   

VAL # 88

  

NV101478200

  

NMC298464

17   

VAL # 89

  

NV101607692

  

NMC298465

18   

VAL # 90

  

NV101347014

  

NMC298466

19   

VAL # 91

  

NV101602268

  

NMC298467

20   

VAL # 92

  

NV102520872

  

NMC298468

21   

VAL # 93

  

NV101347196

  

NMC298469

22   

VAL # 94

  

NV101349299

  

NMC298470

23   

VAL # 95

  

NV101601472

  

NMC298471

24   

VAL # 96

  

NV101303501

  

NMC298472

25   

VAL # 97

  

NV101479051

  

NMC298473

26   

VAL # 98

  

NV101500025

  

NMC298474

27   

VAL # 99

  

NV101508115

  

NMC298475

28   

VAL #100

  

NV101610068

  

NMC298476

29   

VAL #101

  

NV101340669

  

NMC298477

30   

VAL #102

  

NV101602418

  

NMC298478

31   

VAL #103

  

NV101452680

  

NMC298479

32   

VAL #104

  

NV101603131

  

NMC298480

33   

VAL #105

  

NV101451036

  

NMC298481

34   

VAL #106

  

NV101603862

  

NMC298482

35   

VAL #107

  

NV101456444

  

NMC298483

36   

VAL #108

  

NV102521109

  

NMC298484

37   

BIG MACK # 3

  

NV101609708

  

NMC650813

38   

BIG MACK # 4

  

NV101303989

  

NMC650814

                
Table 27-3:Claimant: Larie K. Richardson, Lessee: Goldcorp Dee LLC
Claimant: Larie K. Richardson

Lessee: Goldcorp Dee LLC

56 Claims

Humboldt County, Nevada

 Claim Count      Claim Name    BLM Serial Number   

BLM Legacy

Number

1   

RAM #217

  

NV101780952

  

NMC506813    

2   

RAM #218

  

NV101495576

  

NMC506814

3   

RAM #219

  

NV101752960

  

NMC506815

4   

RAM #220

  

NV101522001

  

NMC506816

5   

RAM #221

  

NV101607850

  

NMC506817

6   

RAM #222

  

NV101523436

  

NMC506818

7   

RAM #223

  

NV101604326

  

NMC506819

8   

RAM #224

  

NV101525877

  

NMC506820

9   

RAM #225

  

NV101602344

  

NMC506821

10   

RAM #226

  

NV101524668

  

NMC506822

11   

RAM #227

  

NV101601281

  

NMC506823

12   

RAM #228

  

NV101495409

  

NMC506824

13   

RAM #229

  

NV101477936

  

NMC506825

14   

RAM #230

  

NV101495485

  

NMC506826

 

 

109


LOGO

 

 

    15       

RAM #231

  

NV101460191

  

NMC506827    

16   

RAM #232

  

NV101756626

  

NMC506828

17   

RAM #233

  

NV101605402

  

NMC506829

18   

RAM #234

  

NV101349868

  

NMC506830

19   

YEN # 37

  

NV101404341

  

NMC532272

20   

YEN # 38

  

NV101452539

  

NMC532273

21   

YEN # 39

  

NV101407127

  

NMC532274

22   

YEN # 40

  

NV101602400

  

NMC532275

23   

YEN # 41

  

NV101401788

  

NMC532276

24   

YEN # 42

  

NV101609906

  

NMC532277

25   

YEN # 43

  

NV101479156

  

NMC532278

26   

YEN # 44

  

NV101492517

  

NMC532279

27   

YEN # 45

  

NV102520427

  

NMC532280

28   

YEN # 46

  

NV101493681

  

NMC532281

29   

YEN # 47

  

NV101304216

  

NMC532282

30   

YEN # 48

  

NV101494295

  

NMC532283

31   

YEN # 49

  

NV101609446

  

NMC532284

32   

YEN # 50

  

NV101496828

  

NMC532285

33   

YEN # 51

  

NV101604635

  

NMC532286

34   

YEN # 52

  

NV101490404

  

NMC532287

35   

YEN # 53

  

NV101300975

  

NMC532288

36   

YEN # 54

  

NV101497233

  

NMC532289

37   

YEN # 55

  

NV101730896

  

NMC532290

38   

YEN # 56

  

NV101601029

  

NMC532291

39   

YEN # 57

  

NV101607357

  

NMC532292

40   

YEN # 58

  

NV101478875

  

NMC532293

41   

YEN # 59

  

NV101458201

  

NMC532294

42   

YEN # 60

  

NV101340646

  

NMC532295

43   

YEN # 61

  

NV101455113

  

NMC532296

44   

YEN # 62

  

NV101341953

  

NMC532297

45   

YEN # 63

  

NV101603738

  

NMC532298

46   

YEN # 64

  

NV101452199

  

NMC532299

47   

YEN # 65

  

NV101404208

  

NMC532300

48   

YEN # 66

  

NV101340429

  

NMC532301

49   

YEN # 67

  

NV101407801

  

NMC532302

50   

YEN # 68

  

NV101340465

  

NMC532303

51   

YEN # 69

  

NV101349924

  

NMC532304

52   

YEN # 70

  

NV101453531

  

NMC532305

53   

YEN # 71

  

NV101405052

  

NMC532306

54   

YEN # 72

  

NV101456750

  

NMC532307

55   

YEN # 72A

  

NV101341628

  

NMC603367

56   

YEN # 71A

  

NV101349942

  

NMC603366

 

 

110


LOGO

 

 

Table 27-4:Lone Tree Brooks Project, Claimant: Goldcorp Dee LLC

 

LONE TREE - BROOKS PROJECT

 

Claimant: Goldcorp Dee LLC

36 Claims

Humboldt County, Nevada

  Claim Count      Claim Name    BLM Serial Number   

BLM Legacy

Number

1   

VG # 19

  

NV101491011

  

NMC397908    

2   

VG # 20

  

NV101780612

  

NMC397909

3   

VG # 21

  

NV101495423

  

NMC397910

4   

VG # 22

  

NV101759203

  

NMC397911

5   

VG # 23

  

NV101496032

  

NMC397912

6   

VG # 24

  

NV101493165

  

NMC397913

7   

VG # 25

  

NV101302615

  

NMC397914

8   

VG # 26

  

NV101492668

  

NMC397915

9   

VG # 27

  

NV101347666

  

NMC397916

10   

VG # 28

  

NV101758008

  

NMC397917

11   

VG # 29

  

NV101300334

  

NMC397918

12   

VG # 30

  

NV101752753

  

NMC397919

13   

VG # 31

  

NV101305064

  

NMC397920

14   

VG # 32

  

NV101547497

  

NMC397921

15   

VG # 33

  

NV101348705

  

NMC397922

16   

VG # 34

  

NV101479036

  

NMC397923

17   

VG # 35

  

NV101347781

  

NMC397924

18   

VG # 36

  

NV101756612

  

NMC397925

19   

VG # 37

  

NV101458368

  

NMC397926

20   

VG # 38

  

NV101455973

  

NMC397927

21   

VG # 39

  

NV101303760

  

NMC397928

22   

VG # 40

  

NV101601410

  

NMC397929

23   

VG # 41

  

NV101496995

  

NMC397930

24   

VG # 42

  

NV101547574

  

NMC397931

25   

VG # 43

  

NV101495598

  

NMC397932

26   

VG # 44

  

NV101451603

  

NMC397933

27   

VG # 45

  

NV101605173

  

NMC397934

28   

VG # 46

  

NV101452512

  

NMC397935

29   

VG # 47

  

NV101605336

  

NMC397936

30   

VG # 48

  

NV101540889

  

NMC397937

31   

VG # 49

  

NV101459313

  

NMC397938

32   

VG # 50

  

NV101480354

  

NMC397939

33   

VG # 51

  

NV101495344

  

NMC397940

34   

VG # 52

  

NV101477234

  

NMC397941

35   

VG # 53

  

NV101495962

  

NMC397942

36   

VG # 54

  

NV101491519

  

NMC397943

 

 

111


LOGO

 

 

Table 27-5:Lone Tree Buffalo Mountain Project, Claimant: Goldcorp Dee LLC

 

LONE TREE - BUFFALO MOUNTAIN PROJECT

Claimant: Goldcorp Dee LLC

44 Claims

Humboldt County, Nevada

  Claim Count      Claim Name    BLM Serial Number    BLM Legacy Number
1   

BUFFALO # 1

  

NV101605213

  

NMC175220    

2   

BUFFALO # 4

  

NV102521200

  

NMC175223

3   

BUFFALO # 5

  

NV101479832

  

NMC175224

4   

BUFFALO # 6

  

NV101492275

  

NMC175225

5   

BUFFALO # 7

  

NV101479639

  

NMC175226

6   

BUFFALO # 8

  

NV101460394

  

NMC175227

7   

BUFFALO # 9

  

NV101498406

  

NMC175228

8   

BUFFALO # 10

  

NV101503037

  

NMC175229

9   

BUFFALO # 11

  

NV101403937

  

NMC175230

10   

BUFFALO # 12

  

NV101602570

  

NMC175231

11   

BUFFALO # 13

  

NV101302994

  

NMC175232

12   

BUFFALO # 14

  

NV101605956

  

NMC175233

13   

BUFFALO # 15

  

NV101349312

  

NMC175234

14   

BUFFALO # 16

  

NV101759591

  

NMC175235

15   

BUFFALO # 17

  

NV101300600

  

NMC175236

16   

BUFFALO # 18

  

NV101544967

  

NMC175237

17   

BUFFALO # 19

  

NV101400817

  

NMC175238

18   

BUFFALO # 20

  

NV101602244

  

NMC175239

19   

BUFFALO # 33

  

NV101731430

  

NMC175252

20   

BUFFALO # 34

  

NV101529682

  

NMC175253

21   

BUFFALO # 35

  

NV101457161

  

NMC175254

22   

BUFFALO # 36

  

NV101457293

  

NMC175255

23   

BUFFALO # 37

  

NV101479840

  

NMC175256

24   

BUFFALO # 38

  

NV101492279

  

NMC175257

25   

BUFFALO # 39

  

NV101522289

  

NMC175258

26   

BUFFALO # 40

  

NV101605272

  

NMC175259

27   

BUFFALO # 41

  

NV101606425

  

NMC175260

28   

BUFFALO # 42

  

NV101343022

  

NMC175261

29   

BUFFALO # 43

  

NV101781034

  

NMC175262

30   

BUFFALO # 52

  

NV101755252

  

NMC264481

31   

BUFFALO # 53

  

NV101493850

  

NMC264482

32   

BUFFALO # 54

  

NV101755446

  

NMC264483

33   

BUFFALO # 55

  

NV101459689

  

NMC264484

34   

BUFFALO # 56

  

NV101342012

  

NMC264485

35   

BUFFALO # 57

  

NV101497569

  

NMC264486

36   

BUFFALO # 58

  

NV101459227

  

NMC264487

37   

BG # 1

  

NV101344551

  

NMC402331

38   

BG # 2

  

NV101497799

  

NMC402332

39   

BG # 3

  

NV101459639

  

NMC402333

40   

BVJV # 1

  

NV101460147

  

NMC576734

41   

BVJV # 2

  

NV101342019

  

NMC576735

42   

BVJV # 3

  

NV101497579

  

NMC576736

43   

BVJV # 4

  

NV101458035

  

NMC576737

44   

COW 19

  

NV101489629

  

NMC937406

 

 

112


LOGO

 

 

 Table 27-6:Lone Tree Patented Lands
                   

PATENTED LANDS

LONE TREE

HUMBOLDT COUNTY

                     
                            Patented Claims                            
APN    Name    MS      Sec      Qtr    T      R      Acres    Owner
07-0611-08    VAL 202      5079      14      SE      34N        42E      19.57    Goldcorp Dee LLC
07-0611-06    VAL 219      5079      14      NE      34N        42E      17.09    Goldcorp Dee LLC
07-0611-07    VAL 219A      5079      14      NE      34N        42E      1.02    Goldcorp Dee LLC
07-0611-03    VAL 220      5079      14      NE      34N        42E      10.3    Goldcorp Dee LLC
07-0611-05    VAL 221      5079      14      NE      34N        42E      20.65    Goldcorp Dee LLC
07-0611-02    VAL 222      5079      14      NE      34N        42E      10.47    Goldcorp Dee LLC
                         Surface and Minerals                         
APN                  Sec      Qtr    T      R      Acres    Owner
07-0381-21                  1      Ptn SW      34N        42E      175.323    Goldcorp Dee LLC
07-0381-02                  3      All      34N        42E      641.76    Goldcorp Dee LLC
07-0611-01                  11      All      34N        42E      641.34    Goldcorp Dee LLC
07-0611-10                  13      All      34N        42E      625.4    Goldcorp Dee LLC
07-0611-10                  14     

E2NE

E2E2NESE

     34N        42E           Goldcorp Dee LLC
07-0381-07                  15      All      34N        42E      640    Goldcorp Dee LLC
07-0381-10                  23      All      34N        42E      640    Goldcorp Dee LLC
                         Surface Only                         
APN                  Sec      Qtr    T      R      Acres    Owner
07-0391-02                  12     

S2N2NWNENW

N2S2NWNENW

     34N        42E      5.03    Goldcorp Dee LLC
                         Minerals Only                         
APN                  Sec      Qtr    T      R      Acres    Owner
07-0441-03                  1      All      33N        41E      645.93    Goldcorp Dee LLC
07-0381-01                  17      All      34N        42E      640    Goldcorp Dee LLC
07-0381-01                  21      All      34N        42E      640    Goldcorp Dee LLC
                         Leased                         
APN                  Sec      Qtr    T      R      Acres    Owner
07-0381-15                  31      All      34N        42E      310.4    BTF Properties

All Lone Tree properties including the Buffalo Mountain and Brooks project areas are controlled by Goldcorp Dee LLC, a wholly owned subsidiary of i-80 Gold Corp, as claimant, lessee or owner.

 

 

113


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 28.

APPENDIX B:VARIOGRAM MAPS

Figure 28-1:Variogram map of AuFA in Qal Figure 28-2:Variogram Map of AuFA in Phv

 

 

LOGO

 

 

114


LOGO

 

 

 

 

LOGO

 

 

115


LOGO

 

 

Figure 28-3:Variogram Map of AuFA in Pem Figure 28-4:Variogram Map of AuFA in Pb

 

 

LOGO

 

 

116


LOGO

 

 

 

 

LOGO

 

 

117


LOGO

 

 

Figure 28-5:Variogram Map of AuFA in Ova Figure 29-1:Variogram Model of Lithology code 1 (Qal)

 

 

LOGO

 

 

118


LOGO

 

 

 29.

APPENDIX C:VARIOGRAM MODELS

 

 

LOGO

 

 

119


LOGO

 

 

Figure 29-2:Variogram Model of Lithology Code 2 (Phv)

 

 

LOGO

 

 

120


LOGO

 

 

Figure 29-3:Variogram Model of Lithology Code 3 (Pem)

 

 

LOGO

 

 

121


LOGO

 

 

Figure 29-4:Variogram Model of Lithology Code 4 (Pb)

 

 

LOGO

 

 

122


LOGO

 

 

Figure 29-5:Variogram Modle of Lithology Code 5 (Ova)

 

 

LOGO

 

 

123

EX-99.3 4 d913666dex993.htm EX-99.3 EX-99.3

Exhibit 99.3

 

LOGO

Initial Assessment of the Granite Creek Mine, Humboldt County, NV Prepared for: i-80 Gold Corp. 5190 Neil Road Suite 460 Reno, Nevada 89502 Prepared By: Practical Mining LLC. TR Raponi Consulting Ltd. Global Resource Engineering Effective Date: December 31, 2024 Report Date: March 26, 2025


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  Osgood Mining  Company LLC. 

 

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  Osgood Mining  Company LLC 

 

Date and Signature Page

The undersigned prepared this Technical Report Summary (TRS) report, titled: Initial Assessment of the Granite Creek Mine, Humboldt County, NV, dated the 26th day of March, 2025, with an effective date of December 31, 2024, in support of the public disclosure of Mineral Resource and Mineral Reserve estimates for the Granite Creek Mine. The format and content of the TRS has been prepared in accordance with Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items and 1300 through 1305).

Dated this March 26, 2025

/s/ Practical Mining LLC

Practical Mining LLC

/s/ T.R. Raponi Consulting Ltd.

T.R. Raponi Consulting Ltd.

/s/ Global Resources Engineering

Global Resources Engineering

 

 

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  Osgood Mining  Company LLC. 

 

Table of Contents

 

Date and Signature Page

     iii  

Table of Contents

     iv  

List of Tables

     xvi  

List of Figures

     xxi  

List of Abbreviations

     xxvi  

1   Summary

     27  

1.1  Introduction

     27  

1.2  Property Description

     27  

1.3  Geology and Mineral Deposits

     27  

1.4  Metallurgical Testing and Processing

     28  

1.5  Geology and Mineralization

     29  

1.6  History

     30  

1.7  Exploration, Drilling, and Sampling

     30  

1.8  Data Verification

     31  

1.9  Granite Creek Underground

     32  

1.9.1   Mineral Resources

     32  

1.9.2   Mining, Infrastructure, and Project Schedule

     33  

1.9.3   Economic Analysis

     33  

1.9.4   Conclusions

     35  

1.9.5   Recommendations

     36  

1.10  Open Pit

     37  

1.10.1   Mineral Resources

     37  

1.10.2   Mining Methods

     38  

1.10.3   Economic Analysis

     38  

2   Introduction

     41  

2.1  Registrant for Whom the Technical Report Summary was Prepared

     41  

2.2  Terms of Reference and Purpose of this Technical Report

     41  

2.3  Details of Inspection

     41  

 

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  Osgood Mining  Company LLC 

 

2.4  Sources of Information

     42  

2.5  Report Version

     42  

2.6  Qualification of the Authors

     43  

2.7  Sources of Information

     45  

2.8  Units of Measure

     45  

2.9  Coordinate Datum

     46  

3   Property Description and Location

     47  

3.1  Property Description

     47  

3.2  Status of Mineral Titles

     47  

3.2.1   Royalties

     50  

3.3  Environmental Liabilities

     52  

3.4  Permits/Licenses

     52  

4   Accessibility, Climate, Local Resources, Infrastructure, and Physiography

     53  

4.1  Accessibility

     53  

4.2  Climate

     53  

4.3  Local Resources

     53  

4.4  Infrastructure

     54  

4.5  Physiography

     54  

5   History

     55  

5.1  Historic Ownership

     55  

5.1.1   Cordex I Syndicate

     55  

5.1.2   Pinson Mining Company

     55  

5.1.3   Homestake – Barrick

     56  

5.1.4   Atna Resources Ltd. Earn-in and PMC Back-in

     56  

5.1.5   Atna 2011 – 2013 Underground Development

     56  

5.1.6   Osgood Mining Company LLC Acquisition

     57  

5.1.7   i-80

     58  

5.2  Historical Mineral Reserve and Production

     58  

6   Geologic Setting, Mineralization and Deposit

     59  

 

 

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  Osgood Mining  Company LLC. 

 

6.1  Regional Geology

     59  

6.2  Local and Property Geology

     62  

6.3  Structural Framework

     68  

6.3.1   Structural Overview

     68  

6.3.2   Faults

     70  

6.4  Mineralization

     72  

6.4.1   Mag Pit Mineralization

     73  

6.4.2   Underground Mineralized Zones

     74  

6.4.3   Rangefront Zone

     74  

6.4.4   CX Zone

     75  

6.4.5   South Pacific Zone

     75  

6.5  Alteration

     75  

6.6  Deposit Types

     78  

7   Exploration

     80  

7.1  Exploration

     80  

7.1.1   Geologic Mapping and Geochemical Sampling

     80  

7.1.2   Osgood Mining Geologic/Structural Mapping

     81  

7.1.3   Geophysical Surveys

     82  

7.1.4   Underground Drifting/Evaluation

     94  

7.1.5   Trenching and Sampling

     94  

7.2  Drilling

     95  

7.2.1   Drilling Campaigns Overview

     95  

7.2.2   Representative Drill Sections and Plan

     101  

7.2.3   Drilling, Sampling, and Recovery Factors

     108  

7.3  Update to Drilling Statistics to Include i-80 Drilling and Land Package Expansion

     108  

7.3.1   i-80 Drilling

     111  

7.3.2   Representative Cross Sections

     112  

7.4  Hydrogeology

     115  

7.4.1   Sampling Methods and Laboratory Determinations

     115  

 

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  Osgood Mining  Company LLC 

 

7.4.2   Hydrogeology Investigations

     116  

7.4.3   Hydrogeologic Description

     116  

7.4.4   Mine Dewatering

     121  

7.4.5   Dewatering Treatment and Discharge

     124  

7.4.6   Groundwater Flow Model

     124  

8   Sample Preparation, Analysis and Security

     127  

8.1   Sampling Methods and Approach

     127  

8.1.1   Reverse Circulation Drilling

     127  

8.1.2   Sampling Methods

     127  

8.1.3   Recovery

     127  

8.1.4   Sample Intervals

     128  

8.1.5   Logging

     128  

8.2   Diamond Drilling

     128  

8.2.1   Sampling Methods

     128  

8.2.2   Recovery

     129  

8.2.3   Sample Intervals

     129  

8.2.4   Logging

     130  

8.3   Sample Security

     131  

8.4   Sample Preparation and Analysis

     132  

8.4.1   PMC 1970 – 1996

     132  

8.4.2   PMC - Homestake 1997 – 2000

     132  

8.4.3   PMC Barrick 2000 – 2008

     132  

8.4.4   Atna 2004 – 2013

     133  

8.4.5   Atna Underground 2011 – 2016

     134  

8.4.6   i-80 Gold 2021 – 2025

     134  

8.5   Data Validation

     135  

8.5.1   Summary

     135  

8.5.2   Atna Review of Prior Data

     135  

8.5.3   Barrick Review of Prior Data

     136  

 

 

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8.5.4   OMC Data Compilation and Validation

     137  

8.6   Quality Assurance/Quality Control Overview

     139  

8.7   Certified Reference Materials

     140  

8.8   GRE Discussion on QA/QC

     142  

8.8.1   GRE Discussion on CRMs

     142  

8.8.2   GRE Discussion on Blanks

     146  

8.8.3   GRE Discussion on Duplicates

     146  

8.9   Quality Assurance/Quality Control Overview by PM (2021-2025)

     147  

8.10   PM Discussion on QA/QC 2021

     148  

8.10.1   PM Discussion on CRMs

     148  

8.10.2   GRE Discussion on Blanks

     151  

8.10.3   GRE Discussion on Duplicates

     152  

8.11   PM Discussion on QA/QC 2022

     153  

8.11.1   PM Discussion on CRMs

     154  

8.11.2   GRE Discussion on Blanks

     155  

8.11.3   PM Discussion on Duplicates

     156  

8.12   Conclusions

     157  

9   Data Verification

     160  

9.1  GRE Site Inspection (2021)

     160  

9.2  Visual Sample Inspection and Check Sampling

     160  

9.3  Database Audits

     164  

9.4  QP Opinions on Adequacy

     164  

9.5  Practical Mining Drillhole Database Verification

     165  

10   Mineral Processing and Metallurgical Testing

     169  

10.1   Introduction

     169  

10.2   Metallurgical Test Work

     170  

10.2.1   McClelland Laboratories, Inc. March and June 1999

     170  

10.2.2   McClelland Laboratories Inc 2013 & 2014

     178  

10.2.3   Dawson Metallurgical Program 2005 and 2006

     184  

 

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10.2.4   FLS Metallurgical Program 2023

     187  

10.3   Sample Representativity

     202  

10.3.1   Overview

     202  

10.3.2   Bulk Samples

     204  

10.3.3   Drillhole Samples

     204  

10.3.4   Metallurgical Composite Assembly

     206  

10.4   Deleterious Elements

     206  

10.4.1   Homestake Mining 1999

     207  

10.4.2   Atna Resources 2005

     207  

10.4.3   Atna Resources 2013

     208  

10.4.4   Osgood 2023

     208  

10.5   Geometallurgical Modeling

     208  

10.5.1   Cyanide Solubility for Different Zones

     210  

10.5.2   Cyanide Solubility Estimation in the Block Model

     222  

10.5.3   Metallurgical Test and Recovery

     222  

10.5.4   Recovery in the Block Model

     231  

10.6   Conclusions

     231  

10.6.1   Sample Representativity

     231  

10.6.2   Test Work on Open Pit Samples

     231  

10.7   Recommendations

     231  

10.7.1   Test Work Recommendations

     232  

10.7.2   Geometallurgy Recommendations

     232  

11   Mineral Resource Estimates

     233  

11.1   Introduction

     233  

11.2   Drill Hole Database

     235  

11.3   Topography

     237  

11.4   Geologic Model

     238  

11.5   Open Pit Estimation

     239  

11.5.1   Estimation Domains

     239  

 

 

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  Osgood Mining  Company LLC. 

 

11.5.2   Assay Compositing

     246  

11.5.3   Evaluation of Outliers

     249  

11.5.4   Density

     251  

11.5.5   Variography

     251  

11.5.6   Block Model Parameters

     252  

11.5.7   Estimation Domains

     253  

11.5.8   Estimation Parameters

     253  

11.5.9   Geometallurgical Modeling

     255  

11.6  Open Pit Resource

     255  

11.6.1   Block Model Validation

     255  

11.6.2   Mineral Resource Classification

     265  

11.6.3   Mineral Resource Statement

     266  

11.6.4   Calculation of Cutoff Grade

     269  

11.6.5   Sources of Uncertainty

     270  

11.6.6   Mineral Resource Sensitivity

     271  

11.7  Underground Mineral Resources

     274  

11.7.1   Structural and Mineralized Grade Shell Modelling

     275  

11.7.2   Model Validation

     283  

11.7.3   Reasonable Prospects for Economic Extraction

     292  

11.7.4   QP Opinion

     292  

11.7.5   Underground Mineral Resources

     292  

12   Mineral Reserve Estimates

     294  

13   Mining Methods

     295  

13.1  Open Pit

     295  

13.1.1   Introduction

     295  

13.1.2   Whittle Pit Shell Analysis

     295  

13.1.3   Pit Design

     299  

13.1.4   Block Model Coding

     300  

13.1.5   Mining Sequence

     301  

 

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  Osgood Mining  Company LLC 

 

13.1.6   Base Case

     302  

13.1.7   Mine Scheduling

     302  

13.1.8   Mine Operation and Layout

     306  

13.1.9   Drilling and Blasting

     315  

13.1.10  Loading and Hauling

     315  

13.1.11  Haul Roads

     315  

13.1.12  Mining Mobile Equipment

     316  

13.2  Underground

     316  

13.2.1   Development

     318  

13.2.2   Production

     320  

13.2.3   Ground Support

     320  

13.2.4   Granite Creek Mineralization Control Procedures

     322  

13.2.5   Mine Production Plan

     324  

14   Recovery Methods

     326  

14.1  Introduction

     326  

14.2  Process Description

     326  

14.2.1   Crusher Circuit

     328  

14.2.2   Grinding Circuit

     329  

14.2.3   Carbon in Leach (CIL) Circuit

     329  

14.2.4   CIL Strip Circuit

     330  

14.3  Refractory Processing

     330  

14.3.1   Third Party Processing

     330  

14.3.2   Lone Tree Pressure Oxidation Facility

     331  

14.3.3   Key Design Criteria

     335  

14.3.4   Lone Tree Facility Description

     335  

14.3.5   Slurry Heaters

     336  

14.3.6   Autoclave Feed

     337  

14.3.7   Autoclave

     337  

14.3.8   Flash System

     337  

 

 

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  Osgood Mining  Company LLC. 

 

14.3.9   Carbon Acid Wash

     338  

14.3.10  Carbon Stripping

     338  

14.3.11  Elution Mercury Abatement

     339  

14.3.12  Carbon Regeneration Kiln

     339  

14.3.13  Carbon Fines Handling

     339  

14.3.14  Electrowinning

     339  

14.3.15  Refining

     339  

14.3.16  Oxygen Plant

     341  

14.3.17  Utilities Consumption

     341  

15   Infrastructure

     344  

15.1  Operations Dewatering

     344  

15.2  Operations Monitoring Wells and VWPs

     344  

15.3  Operations RIBs

     344  

15.4  Operations Water Supply

     345  

15.5  Underground Development

     345  

15.6  Other Infrastructure

     345  

16   Market Studies and Contracts

     350  

16.1  Precious Metal Markets

     350  

16.2  Contracts

     351  

16.2.1   Private Placement Offering

     351  

16.2.2   Orion and Sprott Financing Package

     353  

16.2.3   Equinox Investment

     351  

16.2.4   Private Placement

     351  

16.3  Previous Financing Agreements

     354  

16.3.1   Offtake Agreement

     354  

16.3.2   South Arturo Purchase and Sale Agreement (Silver)

     354  

16.3.3   Autoclave Mineralized Material Purchase Agreement

     354  

16.3.4   Roaster Toll Milling Agreement

     354  

16.3.5   Contract Mining

     355  

 

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16.3.6   Other Contracts

     354  

17   Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups

     355  

17.1 Environmental Setting

     355  

17.2 Geochemistry

     355  

17.2.1   Onsite Water Quality

     357  

17.2.2   Pit Lake Future Water Quality

     359  

17.2.3   Water Treatment Plant

     359  

17.3 Environmental Studies and Issues

     360  

17.4 Social or Community Impacts

     361  

17.5 Permits

     362  

17.6 Water Use Permits

     362  

17.7 Environmental Permits

     362  

17.7.1   National Environmental Policy Act (NEPA)

     363  

17.7.2   State Permits

     364  

17.7.3   Monitoring Requirements

     364  

17.8 Mine Closure

     364  

17.8.1   Mine Closure Design Criteria

     365  

17.8.2   Closure Costs

     366  

17.8.3   Closure Cost Limitations

     367  

17.9 Local Procurement and Hiring

     367  

18   Capital and Operating Costs

     368  

18.1 Open Pit Capital Cost Estimate

     368  

18.1.1   Sustaining

     369  

18.1.2   Facilities

     369  

18.1.3   Process Plant

     370  

18.1.4   Mine Equipment

     370  

18.1.5   G&A Capital

     370  

18.1.6   Working Capital

     371  

18.1.7   Closure

     371  

 

 

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18.2 Open Pit Operating Cost Estimate

     371  

18.2.1   Labor

     371  

18.2.2   Mining Equipment and Consumables

     375  

18.2.3   Process Plant

     376  

18.2.4   Taxes and Royalties

     376  

18.2.5   General and Administrative

     376  

18.3 Granite Creek Underground

     377  

18.3.1   Capital Costs

     377  

18.3.2   Closure and Reclamation

     377  

18.3.3   Underground Mine Operating Costs

     378  

18.3.4   Cutoff Grade

     379  

19   Economic Analysis

     381  

19.1 Taxes

     381  

19.1.1   Federal

     381  

19.1.2   Nevada

     381  

19.1.3   Property Taxes

     382  

19.2 Granite Creek Underground

     382  

19.2.1   Cash Flow

     382  

19.2.2   Sensitivity Analysis

     390  

19.3 Open Pit

     393  

19.3.1   Model Cases

     393  

19.3.2   Results

     394  

19.3.3   Sensitivity Analyses

     396  

19.3.4   Inferred Mineral Resource Impacts on Economics

     397  

19.3.5   Conclusions of Economic Model

     398  

20   Adjacent Properties

     399  

21   Other Relevant Data and Information

     400  

22   Interpretation and Conclusions

     401  

22.1 Drilling

     401  

 

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Humboldt County, NV

  Osgood Mining  Company LLC 

 

22.2 Environmental

     402  

22.3 Open Pit Conclusions

     402  

22.3.1   Metallurgical Conclusions

     402  

22.3.2   Mineral Resource Conclusions

     403  

22.3.3   Mining

     403  

22.3.4   Economics

     404  

22.4 Underground Conclusions

     404  

22.4.1   Metallurgy

     404  

22.4.2   Mining and Infrastructure

     405  

22.4.3   Economics

     405  

23   Recommendations

     406  

23.1 Metallurgical Recommendations

     406  

23.1.1   Test Work Recommendations

     406  

23.1.2   Geometallurgy Recommendations

     407  

23.2 Environmental Recommendations

     407  

23.2.1   Requirements for the EIS

     407  

23.3 Granite Creek Underground

     408  

23.3.1   Recommendations

     408  

23.3.2   Underground Feasibility Study Work Program

     409  

24   References

     410  

25   Reliance on Information Provided by the Registrant

     411  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xvi  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC. 

 

List of Tables

 

Table 1-1 Statement of Mineral Resources      32  
Table 1-2 Underground Mine Financial Statistics      33  
Table 1-3: Granite Creek Mine Project Open Pit Mineral Resource      37  
Table 1-4 Key Economic Indicators      39  
Table 2-1 Personal Inspections by Qualified Professionals      42  
Table 2-2 QP Section Responsibility      43  
Table 2-3 Units of Measure Conversion Factors      45  
Table 3-1 Holding Costs      48  
Table 3-2 Granite Creek Owned Unpatented Claims      50  
Table 3-3 Granite Creek Leased Unpatented Claims      50  
Table 3-4 Granite Creek Royalties      50  
Table 7-1: Salient Results of the Ogee Zone Channel Sample Assays      94  
Table 7-2: Summary of Drilling on the Granite Creek Property Since 1970      96  
Table 7-3: PMC Drilling through 1996      97  
Table 7-4: Homestake Drilling      97  
Table 7-5: Barrick Drilling 2003      97  
Table 7-6: Atna Drilling 2004      98  
Table 7-7: Atna Drilling 2005-2006      98  
Table 7-8: PMC - Barrick Drilling 2007      99  
Table 7-9: PMC – Barrick Drilling 2008      100  
Table 7-10: Atna Drilling 2012      101  
Table 7-11: Atna Drilling 2013 – 2015      101  
Table 7-12 Drillholes Within the Current Property Boundary by Type and Operator      108  
Table 7-13: Timeline for Hydrogeologic Characterization with Relationship to Mining      118  
Table 8-1 Summary of Errors Within the Granite Creek Project Database      137  
Table 8-2 Initial Data Set and 18 April 2019 Data Subset      138  
Table 8-3 Assay Certificates and Samples Uploaded by Laboratory      138  
Table 8-4 QA/QC 2005 – 2015      139  
Table 8-5 QA/QC 2005 – 2015 Insertion Rates      139  
Table 8-6: CRMs used in each year      141  
Table 8-7: CRMs Used by Year and Company (2005 – 2015)      141  
Table 8-8: CRMs Selected by GRE for Control Charts      143  
Table 8-9: QA/QC 2021 and 2022      147  
Table 8-10: QA/QC 2021 and 2022 Insertion Rates      147  
Table 9-1: Summary Table of Hazen Results with Original Assays      163  
Table 9-2 Excluded Drillholes      166  

 

 Practical Mining LLC   March 26, 2025 


 Page xvii  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

Table 9-3 Drill Holes Selected for Review by Type and Operator      166  
Table 9-4 Drillhole Data Fields Reviewed      168  
Table 10-1: Mag Pit Composites for 1999 Test Work Program      171  
Table 10-2: Preg-Robbing Test Results from the 1999 Test Work Program      172  
Table 10-3: NaOH Bottle Roll Tests from 1999 Test Work Program      174  
Table 10-4: CIL Tests from 1999 Test Work Program      175  
Table 10-5: Column Leach Tests from 1999 Test Work Program      176  
Table 10-6 Column Leach Tests from June 1999 Test Work Program      178  
Table 10-7 Sample Composite List from 2013 Test Work Program      179  
Table 10-8 Bottle Roll Tests Results from 2013 Test Work Program      180  
Table 10-9 Results from Bottle Roll Tests Using NaOH from 2013 Testwork Program      182  
Table 10-10 Sample Composition for Column Leach Tests from 2013 Test Work Program      183  
Table 10-11: Bottle Roll and Column Test Results from 2013 Test Work Program      184  
Table 10-12: Autoclave Pre-treatment Tests from Dawson Test Work Program      187  
Table 10-13: Underground Samples Head Assays from FLS Program      188  
Table 10-14: Underground Samples Batch Autoclave Conditions from FLS Program      190  
Table 10-15: Underground Samples Baseline and Batch Pressure Oxidation CIL Results from FLS Program      191  
Table 10-16: Granite Creek Underground Metallurgical Sample Blends for Additional Testing      194  
Table 10-17: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Test Conditions      194  

•  Table 10-18: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Sulfide Oxidation Results Compared to Predicted Results

     195  

•  Table 10-19: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Gold Recovery Results Compared to Predicted Results

     196  
Table 10-20: Granite Creek Underground Metallurgical Testing Program Continuous POx Run Test Conditions      197  
Table 10-21: Underground Samples (OAPC) Continuous Autoclave Tests from FLS Program      199  
Table 10-22: Underground Blend Samples Cyanide Detox Conditions from FLS Program      200  
Table 10-23: Underground Cyanide Detox Reagent Consumption from FLS Program      200  
Table 10-24: Underground Cyanide Detox WAD from FLS Program      200  
Table 10-25: Drillhole Sample Selection and Testing Matrix      205  
Table 10-26: Composite Assays      206  
Table 10-27: Gold and Arsenic Assays CX Pit      207  
Table 10-28: Mag Pit Drill Core Composite Assays      208  
Table 10-29: Available Cyanide Solubility Data in Different Zones      209  
Table 10-30: Numerical Equivalent Alteration Codes.      209  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xviii  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC. 

 

Table 10-31: Column Test and CIL Test (McClelland April 1999 Report)      223  
Table 10-32: DML Wilmot 2005 -2006, Memo Autoclave Test Results (Samples from 2005)      229  
Table 10-33: DML Wilmot 2005 -2006, Memo Autoclave Test Results (Samples from 2006)      229  
Table 10-34: Autoclave Test Results (Samples from 2023)      230  
Table 11-1: Negative Values in Drill Hole Database      237  
Table 11-2: Open Pit Estimation Zone and Pit Name      240  
Table 11-3: Open Pit Numeric Indicator Model Parameters      240  
Table 11-4: Open Pit Compositing Interval Statistics      247  
Table 11-5: Open Pit Compositing Comparison 20 Foot Intervals      248  
Table 11-6: Open Pit Upper Clipping Au ppm Values by Domain      250  
Table 11-7: Open Pit Domain Density Summary      251  
Table 11-8: Open Pit Variogram Parameters      252  
Table 11-9: Block Model Parameters Open Pit      252  
Table 11-10: Open Pit ID2 Estimation Parameters      254  
Table 11-11: Open Pit Combined Estimator Hierarchy      254  
Table 11-12: Open Pit Comparison of Composite Values to Grade Estimation Methods      262  
Table 11-13: Open Pit Mineral Resource Classification Parameters      265  
Table 11-14: Open Pit Parameters for Resource Class Numeric Indicator Model      266  
Table 11-15: Granite Creek Resource Parameters for Open Pit Optimization      267  
Table 11-16: Granite Creek Open Pit Mineral Resource      268  
Table 11-17: Granite Creek Mineral Resource Sensitivity to Cutoff Grade      271  
Table 11-18 Summary of Drilling Within Block Model Extents      277  
Table 11-19 Composite Statistics      278  
Table 11-20 Density Values Used in the Underground Model      281  
Table 11-21 Underground Grade Capping Values      281  
Table 11-22 Ellipsoid Search Parameters      277  
Table 11-23 Resource Classification Parameters      277  
Table 11-24 Mineralization Processed in 2024      282  
Table 11-25 2024 High Grade Block Model Predicted and High Grade Mill - Model Variance      283  
Table 11-26 All Block Model Predicted and Mill - Model Variance in 2024      284  
Table 11-27 Statement of Mineral Resources      287  
Table 13-1: Granite Creek Open Pit Mine Project Whittle Pitshell Analysis Parameters      290  
Table 13-2: Selected Whittle Pit Shells for Resource Areas      294  
Table 13-3: Pit Parameters      294  
Table 13-4: Summary of Pit Phases      296  
Table 13-5: Base Case Pit Resource      297  
Table 13-6: Granite Creek Mine Project Open Pit Base Case Mine Schedule Summary      299  
Table 13-7: Granite Creek Mine Project Open Pit Mobile Equipment Sizes and Quantities      311  

 

 Practical Mining LLC   March 26, 2025 


 Page xix  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

Table 13-8: Contractors Personnel      311  
Table 13-9: Contractors Underground Equipment      312  
Table 13-10: i-80 Personnel      313  
Table 13-11 Mineralization Routing Criteria      318  
Table 13-12: Annual Production and Development Schedule (Including Inferred Mineral Resources)      319  
Table 13-13: Annual Production and Development Schedule (Excluding Inferred Mineral Resources)      320  
Table 14-1: Summary of Key Process Statistics      330  
Table 14-2: Lone Tree Facility Water Consumption by Type      337  
Table 14-3: Lone Tree Facility Energy Usage by Area      337  
Table 15-1 Dewatering Well Completion Details      341  
Table 15-2: Summary of Locations, Construction Information and Water Levels for Dewatering Wells, VWPs, Monitoring Wells and Piezometers      342  
Table 17-1 Weighted Average Concentrations of MWMP Results of Rock Placed in CX Pit 2005 - 2022      355  
Table 17-2 Water Quality April 2023-Jan 2025      356  
Table 18-1: Granite Creek Open Pit Mine Project Capital Costs      368  
Table 18-2: Granite Creek Open Pit Mine Project Initial Capital Costs      369  
Table 18-3: Granite Creek Open Pit Mine Project Facilities Capital Cost      369  
Table 18-4: Granite Creek Open Pit Mine Project Plant Capital Costs      370  
Table 18-5: Granite Creek Open Pit Mine Project G&A Capital Costs      370  
Table 18-6: Granite Creek Open Pit Mine Project Operating Cost Summary      371  
Table 18-7: Granite Creek Open Pit Mine Project Hourly Laborers by Year      371  
Table 18-8: Granite Creek Mine Open Pit Mine Project Salaried Workers, Mine Management      372  
Table 18-9: Granite Creek Mine Open Pit Mine Project General and Administrative Positions by Year      372  
Table 18-10: Granite Creek Mine Open Pit Mine Project Processing Positions by Year      373  
Table 18-11: Granite Creek Open Pit Mine Project Labor Costs by Year (millions)      374  
Table 18-12: Granite Creek Open Pit Mine Project Mining Equipment Costs by Year (millions)      375  
Table 18-13: Granite Creek Open Pit Mine Project Blasting Costs by Year (millions)      375  
Table 18-14: Granite Creek Mine Project Processing Costs by Year (1000s)      376  
Table 18-155 Capital Cost Estimates ($000’s)      377  
Table 19-1 Real and Personal Property Taxes      382  
Table 19-2 Income Statement (Includes Inferred Mineral Resources)      383  
Table 19-3 Cash Flow Statement (Includes Inferred Mineral Resources)      384  
Table 19-4 Income Statement (without Inferred Mineral Resources)      385  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xx  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC. 

 

Table 19-5 Cash Flow Statement (Without Inferred Mineral Resources)      386  
Table 19-6 Financial Statistics      387  
Table 23-1 Feasibility Study Work Program      408  

 

 Practical Mining LLC   March 26, 2025 


 Page xxi  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

List of Figures

 

Figure 1-1 NPV@5% Sensitivity to Varying Gold Price, Gold Grade, Capital Costs, and Operating Costs      40  
Figure 3-1 Granite Creek Project Location      47  
Figure 3-2 Granite Creek Land Position Map      49  
Figure 3-3 Granite Creek Royalty Map      51  
Figure 6-1 Regional Geologic Map of a Portion of the Osgood Mountains including Granite Creek      60  
Figure 6-2 Granite Creek Stratigraphic Column      64  
Figure 6-3 Geology and Structural Map      67  
Figure 6-4 Geology and Structural Map of the Granite Creek Property      69  
Figure 6-5 Cross-section A-A’ looking Northeast showing Structure, Lithology and Mineralization (Section Line is shown on Figure 6-4)      73  
Figure 6-6 Alteration of the Mag Pit      77  
Figure 7-1: Gravity Survey, 2,587 Stations, Magee Geophysical Services, 2006      84  
Figure 7-2: Pinson Local Gravity Interpretation      85  
Figure 7-3: Location of the MT Survey Lines on the Geology and Pit locations (Left) and on the Residual Gravity (Right)      87  
Figure 7-4: MT Resistivity Depth Inversion for Line 6090      88  
Figure 7-5: MT Resistivity Depth Inversion for Line 12300      89  
Figure 7-6: MT Resistivity Depth Inversion for Line 13860      90  
Figure 7-7: MT Resistivity Depth Inversion for Line 15300      91  
Figure 7-8: MT Resistivity Depth Inversion for Line 17160      92  
Figure 7-9: MT Resistivity Depth Inversion for Line 19230      93  
Figure 7-10: Granite Creek Project Drill Plan by Operator      96  
Figure 7-11: Plan View Section Lines of Granite Creek Mine Project      102  
Figure 7-12: Vertical Section A-A1 of the Mag Pit Area      103  
Figure 7-13: Vertical Section B-B1 of the Pit CX and C Area      104  
Figure 7-14: Vertical Section C-C1 of the Pit A Area      105  
Figure 7-15: Vertical Section D-D1 of the Pit B Area      106  
Figure 7-16: Vertical Section E-E1 of the Underground Resource Area      107  
Figure 7-17 Drilling Completed by PMC      109  
Figure 7-18 Drilling Completed by PMC with Barrick as Operator      110  
Figure 7-19 Drilling Completed by Atna      111  
Figure 7-20 Drilling Completed by i-80      112  
Figure 7-21 Plan View Showing Section Locations through the Underground Resource Area      113  
Figure 7-22 Section A-A’ Showing Drilling in the CX Zone, 100 ft thick, looking North      114  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xxii  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC. 

 

Figure 7-23 Section B-B’ Showing Drilling in the Otto and Ogee Zones, 25 ft thick, looking North      114  
Figure 7-24 Section C-C Showing Drilling in the South Pacific Zone, 50 ft thick, looking North      115  
Figure 7-25 Well Locations      117  
Figure 7-26 Predictive and Passive Inflows from Scenarios One and Two      126  
Figure 8-1: Assay Standard Results (2005-2015)      143  
Figure 8-2: CRM OxG60 (2007 – 2009) FA-ICP-ES      144  
Figure 8-3: CRM OxI54 (2007 – 2009) FA-ICP-ES      144  
Figure 8-4: CRM OXL25 (2005 – 2006) FA-GRAV      144  
Figure 8-5: CRM SG31 (2007 – 2009) FA-ICP-ES      145  
Figure 8-6: CRM SJ32 (2007 – 2009) FA-ICP-ES      145  
Figure 8-7: CRM SQ18 (2005 – 2006) FA-AAS      145  
Figure 8-8: Fire Assay Blank Samples (2005-2015)      146  
Figure 8-9: Laboratory Duplicate Comparison (2005-2015)      147  
Figure 8-10: CRM CDN-GS-7J for the 2021 Drilling Program      149  
Figure 8-11: CRM CDN-GS-8C for the 2021 Drilling Program      149  
Figure 8-12: CRM CDN-GS-30C for the 2021 Drilling Program      150  
Figure 8-13: CRM CDN-GS-P1A for the 2021 Drilling Program      150  
Figure 8-14: CRM CDN-GS-P6E for the 2021 Drilling Program      151  
Figure 8-15: Blank Results for the 2021 Drilling Program      152  
Figure 8-16: Field Duplicate Samples for the 2021 Drilling Program      153  
Figure 8-17: Preparation Duplicate Samples for the 2021 Drilling Program      153  
Figure 8-18: CRM CDN-GS-7J for the 2022 Drilling Program      154  
Figure 8-19: CRM CDN-GS-30C for the 2022 Drilling Program      155  
Figure 8-20: CRM CDN-GS-P6E for the 2022 Drilling Program      155  
Figure 8-21: Blank Results for the 2022 Drilling Program      156  
Figure 8-22 Field Duplicate Samples for the 2022 Drilling Program      157  
Figure 8-23: Field Duplicate Samples for the 2022 Drilling Program      157  
Figure 9-1: Sample Correlation Plot      163  
Figure 10-1: Gold Cyanide Solubility and Total Organic Carbon Influence      173  
Figure 10-2: CIL Recovery and Head Grade Influence      175  
Figure 10-3: Column Recovery and Solubility Influence      177  
Figure 10-4: Bottle Roll Recovery and Solubility Influence      181  
Figure 10-5: Column and Bottle Roll Recovery and Solubility Influence      184  
Figure 10-6: Solubility and Sulfide Influence – Ogee Samples      186  
Figure 10-7: CIL Gold Recovery as a Function of Sulfide Sulfur Oxidation – Underground Samples      193  

 

 Practical Mining LLC   March 26, 2025 


 Page xxiii  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

Figure 10-8 Granite Creek POx Pilot Plant Sulfide Oxidation Profile      198  
Figure 10-9 Plan View Showing Metallurgical Sample Locations      203  
Figure 10-10 Isometric View Showing Metallurgical Sample Locations      204  
Figure 10-11 PCA- Scree Plot for Mag Pit      211  
Figure 10-12 PCA – Biplot for Mag Pit      212  
Figure 10-13 Regression Tree Model for Mag Pit      212  
Figure 10-14 Observed and Predicted Cyanide Solubility for Gold (ppm)      213  
Figure 10-15 Observed and Predicted Cyanide Solubility for Gold (ppm)      214  
Figure 10-16 PCA- Scree Plot for C and CX Pit      214  
Figure 10-17: PCA – Biplot for C and CX Pit      215  
Figure 10-18: Regression Tree Model for C and CX Pit      216  
Figure 10-19: Observed and Predicted Cyanide Solubility for Gold (ppm)      217  
Figure 10-20: PCA- Scree Plot for A Pit      218  
Figure 10-21: PCA – Biplot for A Pit      218  
Figure 10-22: Regression Tree Model for A Pit      219  
Figure 10-23 Observed and Predicted Cyanide Solubility for Gold (ppm)      219  
Figure 10-24 PCA- Scree Plot for B Pit      220  
Figure 10-25: PCA – Biplot for B Pit      221  
Figure 10-26: Regression Tree Model for B Pit      221  
Figure 10-27: Observed and Predicted Cyanide Solubility for Gold (ppm)      222  
Figure 10-28: Cyanide Solubility vs Column Recovery      226  
Figure 10-29: Calculated Head Grade vs Carbon in Leach Recovery      226  
Figure 10-30: Solubility vs BRT Recovery      227  
Figure 10-31 Calculated Head Grade vs Carbon in Leach Recovery (Outlier Removed)      228  
Figure 11-1: Drill Holes Used Plan View on Topography      236  
Figure 11-2: Current Topography Used for Resource Estimation      238  
Figure 11-3: Geologic Model Oblique View      239  
Figure 11-4: Open Pit Estimation Zones      240  
Figure 11-5: Example of Numeric Indicator High Grade Trend Analysis Mag Pit      242  
Figure 11-6: Open Pit Zone 3 Sub-Domains      243  
Figure 11-7: High Grade and Low Grade Open Pit Domains in CX Fault      243  
Figure 11-8: Box and Whisker Plot of Open Pit Estimation Domains      244  
Figure 11-9: HG and LG Distributions in Zones 1 and 2      245  
Figure 11-10: HG and LG Distributions in Zones 3 and 4      246  
Figure 11-11: Open Pit Interval Length Statistics of Au ppm Assays      247  
Figure 11-12: Open Pit Compositing Comparison 20 Foot Intervals      248  
Figure 11-13: Example of Open Pit Cumulative Log Probability Plot Zone 1 HG      250  
Figure 11-14: Open Pit Numeric Indicator Models      253  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xxiv  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC. 

 

Figure 11-15: Open Pit Zone 1 Visual Comparison Composite to Block Model Grade Plan View      256  
Figure 11-16: Open Pit Zone 2 Visual Comparison Composite to Block Model Grade Plan View      257  
Figure 11-17: Open Pit Zone 3 Visual Comparison Composite to Block Model Grade Plan View      258  
Figure 11-18: Open Pit Zone 4 Visual Comparison Composite to Block Model Grade Plan View      259  
Figure 11-19: Open Pit Zone 1 Section Composites and Block Model Cross Section      260  
Figure 11-20: Open Pit Zone 2 Section Composites and Block Model Cross Section      260  
Figure 11-21: Open Pit Zone 3 Section Composites and Block Model Cross Section      261  
Figure 11-22: Open Pit Zone 4 Section Composites and Block Model Cross Section      262  
Figure 11-23: Cumulative Frequency of Composite and Block Data      263  
Figure 11-24: Open Pit Swath Plot X axis, Zone 1 High Grade Domain      264  
Figure 11-25: Open Pit Swath Plot Y axis, Zone 1 High Grade Domain      264  
Figure 11-26: Open Pit Swath Plot Z axis, Zone 1 High Grade Domain      265  
Figure 11-27: Open Pit Constrained Resource Class All Areas Plan View      266  
Figure 11-28 Major Faulting and Underground 0.10 opt Grade Shells      276  
Figure 11-29 Histogram of 0.004 Au opt Composites      279  
Figure 11-30 Histogram of 0.10 Au opt Composites      280  
Figure 11-31 Comparative Cross Section Through Otto and Ogee Zones      283  
Figure 11-32 Comparative Cross Section Through the CX Zone      283  
Figure 11-33 Comparative Cross Section through the South Pacific Zone      285  
Figure 11-34 Easterly Drift Analysis      286  
Figure 11-35 Elevation Drift Analysis      286  
Figure 11-36 Monthly High Grade Mill to Model Au Ounce Variance in 2024      288  
Figure 11-37: 2024 Cumulative High Grade Mill to Model Variance      289  
Figure 11-38 2024 Monthly All Processed to Model Au Ounce Variance      290  
Figure 11-39 2024 Cumulative All Processed to Model Variance      291  
Figure 13-1 Marginal Impact Undiscounted Cashflow Mag Pit      297  
Figure 13-2 Marginal Impact Undiscounted Cashflow CX Pit      297  
Figure 13-3 Marginal Impact Undiscounted Cashflow Pit B      298  
Figure 13-4 Marginal Impact Undiscounted Cashflow Pit A      298  
Figure 13-5 Cross-Section of Typical Pit Slope      300  
Figure 13-6 Granite Creek Mine Project Base Case Mine Schedule      306  
Figure 13-7 Conceptual Project Layout      307  
Figure 13-8 Phased Pit and Site Plan Layout B Pit      308  
Figure 13-9 Phased Pit and Site Plan Layout B Pit & CX Phase 1      309  

 

 Practical Mining LLC   March 26, 2025 


 Page xxv  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

Figure 13-10 Phased Pit and Site Plan Layout B Pit & CX Phases 1 & 2      310  
Figure 13-11 Phased Pit and Site Plan Layout B Pit & CX Phases 1, 2, & 3      311  
Figure 13-12 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase1      312  
Figure 13-13 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase 1 & 2      313  
Figure 13-14 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase 1, 2, & 3      314  
Figure 13-15 Existing (Shaded Blue) and Planned Mine Development      319  
Figure 13-16 Typical 3-Cut Stope and Hanging Wall Attack Ramp      320  
Figure 13-17 Primary Ground Support Installation      321  
Figure 13-18 Cemented Rock Fill in Adjacent Cut      322  
Figure 14-1: Conceptual Flowsheet      328  
Figure 14-2 Third Party POX Simplified Flowsheet      331  
Figure 14-3 Lone Tree Facility Block Flow Diagram      334  
Figure 15-1 Water Treatment Plant      345  
Figure 16-1 Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average      350  
Figure 19-1 Gold Production and Cost per Ounce (With Inferred)      388  
Figure 19-2 Cash Flow Waterfall Chart (With Inferred)      389  
Figure 19-3 Gold Production and Cost per Ounce (Without Inferred)      389  
Figure 19-4 Cash Flow Waterfall Chart (Without Inferred)      390  
Figure 19-5 NPV 5% Sensitivity      391  
Figure 19-6 NPV 8% Sensitivity      391  
Figure 19-7 IRR Sensitivity      392  
Figure 19-8 Profitability Index Sensitivity      392  

 

 

 Practical Mining LLC   March 26, 2025 


 Page xxvi  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

List of Abbreviations

 

       

A

  

Ampere

  

kA

  

kiloamperes

   

AA

  

atomic absorption

  

kCFM

  

thousand cubic feet per minute

   

AGP

  

Acid Generation Potential

  

Kg

  

Kilograms

   

Ag

  

Silver

  

km

  

kilometer

   

ANFO

  

ammonium nitrate fuel oil

  

km2

  

square kilometer

   

ANP

  

Acid Neutralization Potential

  

kWh/t

  

kilowatt-hour per ton

   

Au

  

Gold

  

LoM

  

Life-of-Mine

   

AuEq

  

gold equivalent

  

m

  

meter

   

btu

  

British Thermal Unit

  

m2

  

square meter

   

°C

  

degrees Celsius

  

m3

  

cubic meter

   

CCD

  

counter-current decantation

  

masl

  

meters above sea level

   

CIL

  

carbon-in-leach

  

mg/L

  

milligrams/liter

   

CoG

  

Cut off grade

  

mm3

  

cubic millimeter

   

cm

  

centimeter

  

MME

  

Mine & Mill Engineering

   

cm2

  

square centimeter

  

Moz

  

million troy ounces

   

cm3

  

cubic centimeter

  

Mt

  

million tonnes

   

cfm

  

cubic feet per minute

  

MTW

  

measured true width

   

CRec

  

core recovery

  

MW

  

million watts

   

CSS

  

closed-side setting

  

m.y.

  

million years

   

CTW

  

calculated true width

  

NGO

  

non-governmental organization

   

°

  

degree (degrees)

  

NI 43-101

  

Canadian National Instrument 43-101

   

dia.

  

diameter

  

oz

  

Troy Ounce

   

EA

  

Environmental Assesment

  

opt

  

Troy Ounce per short ton

   

EIS

  

Environmental Impact Statement

  

oz/ton

  

Troy Ounce per short ton

   

EMP

  

Environmental Management Plan

  

%

  

percent

   

FA

  

fire assay

  

PLC

  

Programmable Logic Controller

   

Ft

  

Foot

  

PLS

  

Pregnant Leach Solution

   

Ft2

  

Square foot

  

PMF

  

probable maximum flood

   

Ft3

  

Cubic foot

  

POO

  

Plan of Operations

   

g

  

Gram

  

ppb

  

parts per billion

   

g/L

  

gram per liter

  

ppm

  

parts per million

   

g-mol

  

gram-mole

  

POX

  

Pressure Oxidation

   

g/t

  

grams per metric tonne

  

QAQC

  

Quality Assurance/Quality Control

   

ha

  

hectares

  

RC

  

reverse circulation drilling

   

HDPE

  

Height Density Polyethylene

  

ROM

  

Run-of-Mine

   

HTW

  

horizontal true width

  

RQD

  

Rock Quality Description

   

ICP

  

induced couple plasma

  

SEC

  

U.S. Securities & Exchange Commission

   

ID2

  

inverse-distance squared

  

Sec

  

second

   

ID3

  

inverse-distance cubed

  

SG

  

specific gravity

   

mm

  

millimeter

  

SPT

  

Standard penetration test

   

mm2

  

square millimeter

  

ton

  

US Short Ton

 

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 i-80 Gold Corp   Summary   Page  27  

 

1

Summary

1.1 Introduction

This Technical Report Summary (TRS) dated the 26th day of March, 2025 with an effective date of December 31, 2024 provides an updated statement of Mineral Reserves and Mineral Resources for Osgood Mining Company’s Granite Creek mine. This TRS also provides an Initial Assessment and statement of mineral resources at Granite Creek.

Cautionary Notes:

 

  1.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable Canadian securities legislation. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law.

1.2 Property Description

The Granite Creek Project is located in Humboldt County, Nevada, 28 miles northeast of the town of Winnemucca, and it is part of the historic Potosi mining district. It is centered at roughly 41° 8’ N latitude and 117° 15.5’ W longitude. It encompasses about 4,506 acres (1,823.5 hectares) including owned unpatented claims, leased unpatented claims and owned surface fee land. i-80 Gold purchased the Granite Creek property from Waterton Global in June 2020.

1.3 Geology and Mineral Deposits

The Property is located on the eastern flank of the Osgood Mountains within the Basin and Range tectonic province of northern Nevada. The Granite Creek Mine occurs within a northeast-trending structural corridor known as the Getchell gold trend. This trend also encompasses a number of gold deposits located outside the Property including the Preble, Getchell, Turquoise Ridge, and

 

 

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 Page  28  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Twin Creeks. These deposits are hosted in Paleozoic marine sedimentary rocks. Gold mineralization at the Property is described as a Carlin-type, sedimentary-rock hosted system.

The Property geology comprises a sequence of Cambrian to Ordovician sedimentary rocks that form part of the Osgood Mountain Terrane and the Osgood Mountains. Much of the Property comprises shales, hornfels sedimentary rocks and limestone interbeds of the Preble Formation, and an overlying (or juxtaposed), alternating sequence of limestone, shale, and dolomite with tuffaceous shale and intraformational conglomerates belonging to the Comus Formation. The Preble and Comus Formations have been folded into a broad north-plunging anticline and have been intruded by large Cretaceous granodiorite stocks, resulting in irregular contact metamorphism.

Gold mineralization at the Property is strongly structurally controlled, occurring at favorable sites within a fault network occurring around the eastern edge of the Osgood granodiorite and predominantly within Comus Formation host rocks. Mineralization is commonly associated with the decarbonatization of carbonate rocks and the introduction of silica, fine grained pyrite, arsenian pyrite, and remobilized carbon. Continuity of mineralization is highly variable, ranging from 40 to 4,500 feet (12 to 1,372 meters) in strike extent, 250 to 1,800 feet (76 to 550 meters) in down-dip extent and 5 to 400 feet (1.5 to 122 meters) in thickness. The underground mineralization has a variable thickness between 5 and 30 feet (1.5 and 9 meters).

Oxidation reaches depths of up to 1,800 feet (550 meters) within shear zones. Oxide mineralization includes pervasive limonite, hematite, along with other iron and arsenic oxides. Historical production from the open pits was focused on oxidized material.

Underground mineralization displays pervasive argillization and decarbonatization of host lithologies, along with the formation of dissolution collapse breccias and intense shearing. Where the alteration is strongest, the altered zones consist of punky, spongy decarbonatized limestone in an argillically altered fine-grained, carbon-rich matrix (Gustavson, 2012). Silicification is minor and occurs as a broad overprint on the zone. Underground production includes both sulfide and oxide material.

1.4 Metallurgical Testing and Processing

A wide range of metallurgical testing has been conducted on the Granite Creek deposit from 1999 to 2024. These tests were conducted on material from the oxide, transition and sulfide domains (sulfide being primarily underground). Test work focused on the primary variables in a typical gold project including: comminution, heap leaching, tank leaching (direct cyanidation and carbon in leach), solid/liquid separation, cyanide destruction, and refractory ore treatment (underground material).

 

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 i-80 Gold Corp   Summary   Page  29  

 

The results of the test work program showed that the deposit (open pit resource) was amenable to both heap leach and carbon-in-leach (CIL) processes. The heap leach gold extraction is proportional to the cyanide soluble gold content and the CIL extraction is proportional to the feed grade averaging 86.6% over the life of mine. The Granite Creek material contains organic carbon and shows varying degrees of “preg-robbing,” as such the CIL process route was selected for the process. The underground material is currently toll treated in an offsite autoclave process.

1.5 Geology and Mineralization

The Property is located on the eastern flank of the Osgood Mountains within the Basin and Range tectonic province of northern Nevada. The Granite Creek Mine occurs within a northeast-trending structural corridor known as the Getchell gold trend. This trend also encompasses a number of gold deposits located outside the Property including the Preble, Getchell, Turquoise Ridge, and Twin Creeks. These deposits are hosted in Paleozoic marine sedimentary rocks. Gold mineralization at the Property is described as a Carlin-type, sedimentary-rock hosted system.

The Property geology comprises a sequence of Cambrian to Ordovician sedimentary rocks that form part of the Osgood Mountain Terrane and the Osgood Mountains. Much of the Property comprises shales, hornfels sedimentary rocks and limestone interbeds of the Preble Formation, and an overlying (or juxtaposed), alternating sequence of limestone, shale, and dolomite with tuffaceous shale and intraformational conglomerates belonging to the Comus Formation. The Preble and Comus Formations have been folded into a broad north-plunging anticline and have been intruded by large Cretaceous granodiorite stocks, resulting in irregular contact metamorphism.

Gold mineralization at the Property is strongly structurally controlled, occurring at favorable sites within a fault network occurring around the eastern edge of the Osgood granodiorite and predominantly within Comus Formation host rocks. Mineralization is commonly associated with the dec---arbonatization of carbonate rocks and the introduction of silica, fine grained pyrite, arsenian pyrite, and remobilized carbon. Continuity of mineralization is highly variable, ranging from 40 to 4,500 feet (12 to 1,372 meters) in strike extent, 250 to 1,800 feet (76 to 550 meters) in down-dip extent and 5 to 400 feet (1.5 to 122 meters) in thickness. The underground mineralization has a variable thickness between 5 and 30 feet (1.5 and 9 meters).

Oxidation reaches depths of up to 1,800 feet (550 meters) within shear zones. Oxide mineralization includes pervasive limonite, hematite, along with other iron and arsenic oxides. Historical production from the open pits was focused on oxidized material.

 

 

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 Page  30  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Underground mineralization displays pervasive argillization and decarbonatization of host lithologies, along with the formation of dissolution collapse breccias and intense shearing. Where the alteration is strongest, the altered zones consist of punky, spongy, decarbonatized limestone in an argillically altered fine-grained, carbon-rich matrix (Gustavson, 2012). Silicification is minor and occurs as a broad overprint on the zone. Historical underground production included both sulfide and oxide material.

1.6 History

The Property has been explored by a number of individuals and mining/exploration companies since the late 1930s. The original discovery on the Property was made by Clovis Pinson and Charles Ogee in the mid to late 1930s, but production did not occur until after World War II, when ore from the original discovery was shipped to and processed at the Getchell mine mill. In 1949 and 1950, total production from the Granite Creek Mine amounted to approximately 10,000 short tons (9,071 tonnes) grading approximately 0.14 ounces per ton (opt) (4.8 g/t).

1.7 Exploration, Drilling, and Sampling

Cordex completed ground-based magnetics over the CX Zone in 1970. In 1983, Cordex conducted a 1:6000-scale mapping program of the Property. In 2016, OMC contracted Mr. Robert Leonardson to complete a geological study on the Property that focused on advancing OMC’s understanding of the structural framework and providing guidance on exploration targeting. By the time of preparation of this technical report, i-80 had drilled 17 surface holes within the open pit areas.

The Property has been historically drilled using a combination of reverse circulation (RC) and diamond drilling. The majority of drilling was completed from surface. More recent drilling was completed as underground diamond core drilling. Sampling protocols adopted by former operators were similar and generally followed industry best practices of the time.

RC samples were collected from the drill cyclone in 5-foot (1.5-meter) intervals. Diamond core was sampled predominantly as 5-foot (1.5-meter) intervals but were locally adjusted based on geological alteration and oxidation contacts. RC and core recovery were recorded and considered to be excellent.

Samples were prepared and analyzed by a number of accredited laboratories throughout the Project history, including ALS Chemex, Inspectorate American Laboratories (IAL), and American Assay Laboratories (AAL).

 

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 i-80 Gold Corp   Summary   Page  31  

 

1.8 Data Verification

Data validation has been completed by various operators throughout the Project’s history. This process comprised the checking of original assay certificates and drillhole records against the digital database. This was completed most recently in April 2019 by OMC.

Quality assurance/quality control (QA/QC) samples including Certified Reference Materials (CRMs), coarse blanks, and field duplicate samples were included regularly with samples submitted between 2005 and 2008. A limited number of CRMs were included with drilling completed in 2012.

In 2021, GRE reviewed all prior work on available QA/QC data between 2005 and 2015. GRE also reviewed and checked QA/QC procedures and the database provided by i-80 Gold Corp.

In general, the QA/QC sample insertion rates used fall below general accepted industry standards. For future exploration campaigns, standards, blanks, and duplicates, including one standard, one duplicate, and one blank sample, should be inserted every 20 interval samples, as is common within industry standards.

CRM samples show a reasonable level of accuracy, but poor to moderate precision when using standard deviations provided by the CRM supplier. A maximum of three to five different CRM samples would be adequate to monitor laboratory performance at the approximate cutoff grades, average grades, and higher grades of the deposits.

Blank sample results are considered acceptable and suggest no systematic contamination has occurred throughout the analytical process.

Duplicate sample results show suboptimal performance, which may be a result of the heterogenous nature of mineralization, uncrushed samples, and sampling variance. Overall, duplicate samples appear to be positively biased, with duplicate results returning higher grade than original samples.

Based on the review of the project database and all existing project documents and the author’s observations of the geology and mineralization at the project during the site visit, GRE’s QP considers the lithology, mineralization, and assay data contained in the project database to be reasonably accurate and suitable for use in estimating mineral resources.

 

 

 Practical Mining LLC   March 26, 2025 


 Page 32  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

1.9 Granite Creek Underground

 

1.9.1

Mineral Resources

Table 1-1 Summary of Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

 

Zone   ktons   ktonnes   Au opt   Au g/t   Au  koz
Measured

Ogee

  88   80   0.244   8.4   22

Otto

  59   53   0.256   8.8   15

Meas Total

  147   133   0.249   8.5   37
Indicated

CX

  8   7   0.391   13.4   3

Ogee

  181   164   0.352   12.1   64

Otto

  295   268   0.316   10.8   93

South Pacific

  223   203   0.286   9.8   64

Ind Total

  707   641   0.317   10.9   224
Measured and Indicated

CX

  8   7   0.391   13.4   3

Ogee

  269   244   0.317   10.9   85

Otto

  354   321   0.306   10.5   108

South Pacific

  223   203   0.286   9.8   64

M&I Total

  854   775   0.305   10.5   261
Inferred

CX

  97   88   0.351   12.0   34

Ogee

  42   38   0.563   19.3   24

Otto

  187   170   0.401   13.7   75

South Pacific

  536   486   0.361   12.4   194

Inf Total

  862   782   0.378   13.0   326

Notes Pertaining to Underground Mineral Resources:

 

  1.

Mineral Resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of $27.25 per ounce. Refer to section 16.1 for price selection details.

  2.

Mineral Resources have been estimated using gold metallurgical recoveries of 85.2% to 94.2% for pressure oxidation. Payment for refractory mineralization sold to a third party is 58%. Oxide CIL mineralization payments vary from 40% to 70% based upon the grade of the mineralization.

  3.

The cutoff grade for refractory Mineral Resources varies from 0.151 to 0.184 opt. for acidic conditions. The cutoff grade for oxide mineral resources is 0.075 opt;

  4.

The contained gold estimates in the Mineral Resource table have not been adjusted for metallurgical recoveries.

  5.

Numbers have been rounded as required by reporting guidelines and may result in apparent summation differences.

  6.

A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or

 

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 i-80 Gold Corp   Summary   Page  33  

 

 

quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling;

  7.

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

  8.

Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors,

  9.

Mineral Resources have an effective date of December 31, 2024, and

 

The reference point for mineral resources is in situ.

 

1.9.2

Mining, Infrastructure, and Project Schedule

The Granite Creek Underground mine is fully permitted and has entered the production phase of operations. All infrastructure is in place to support the anticipated production rate and duration of mine operations.

 

1.9.3

Economic Analysis

The mineral resource at the Granite Creek Underground Mine contains 50% by weight inferred mineral resources. On a contained ounce basis, inferred mineral resources account for 56% of the contained gold ounces. The without inferred case presented is a gross factorization of the mine plan and has not been modified to reflect accompanying changes to capital development, productivities or unit operating costs. The results of a constant dollar cash flow analysis of the planned underground mining operation are shown in Table 1-2.

Only the case that includes inferred mineral resources provides a positive cash flow with a NPV 5% of $155M and 84% IRR. The high IRR is attributed to the fact that all mine infrastructure has been completed and production ramp up has started.

Table 1-2 Underground Mine Financial Statistics

 

        With Inferred        Without  Inferred  

Gold price (US$/oz)

   $2,175

Silver price (US$/oz)

   $27.25

Mine life (years)

   8

Average mineralized mining rate (tons/day)

   435    225    

Average grade (oz/t Au)

   0.339    0.292    

 

 

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 Page  34  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

        With Inferred       Without  Inferred  

Average gold recovery (autoclave %)

   78%   78%    

Average annual gold production (koz)

   52   23    

Total recovered gold (koz)

   418   186    

Sustaining capital (M$)

   $88.8   $88.8    

Cash cost (US$/oz) 1

   $1,366   $1,699    

All-in sustaining cost (US$/oz) 1,2

   $1,597   $2,217    

Project after-tax NPV5% (M$)

   $155   ($30)    

Project after-tax NPV8% (M$)

   $135   ($33)    

Project after-tax IRR

   84%   -12.7%    

Payback Period

   3.2 Years   NA    

Profitability Index 5%3

   12.6   0.7    

Notes:

  1.

Net of byproduct sales;

 

  2.

Excluding income taxes, resource conversion drilling, corporate G&A, corporate taxes and interest on debt;

 

  3.

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates breakeven;

 

  4.

This IA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the IA will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability;

 

  5.

Inferred mineral resources constitute 50% of mass and 56% of gold ounces of all mineral resources. The “Without Inferred” statistics presented are a gross factorization of the mine plan without any redesign of mine excavations or recalculation of productivities and costs. Capital costs are the same for the “With Inferred” and “Without Inferred” scenarios. The “Without Inferred” scenario is presented solely to illustrate the project’s dependence on inferred mineral resources.

 

  6.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable Canadian and United States securities regulations. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law;

 

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 i-80 Gold Corp   Summary   Page  35  

 

1.9.4

Conclusions

Metallurgy

 

  1.

Granite Creek underground samples were refractory with baseline CIL gold recoveries ranging from 9% to 46%, averaging 31%;

 

  2.

Shake flask tests with gold cyanide spikes were used to determine preg robbing index. The average preg-robbing index was 17.9%, ranging from 4.4% to 54.1%;

 

  3.

Bench top autoclave batch pressure oxidation tests were completed on all samples with 2 sets of acid conditions and four sets of alkaline conditions. Acid conditions resulted in resulted in higher sulfur oxidations and higher gold recoveries;

 

  4.

Three continuous pressure oxidation runs were completed with two acid and one alkaline sets of conditions based on the batch results. The continuous results followed the results of the batch tests with the acid conditions producing the higher sulfur oxidations and gold recoveries;

 

  5.

Overall gold recoveries increased with increasing sulfur oxidation;

 

  6.

Cyanide destruction tests on CIL tailings using the SO2/air process reduced weak acid dissociable cyanide concentrations to below 50 ppm using established reagent addition rates and retention time;

 

  7.

Thickening and filtration tests on CIL tailings showed unacceptable thickening properties and filtration rates. Thickening and filtration of pressure oxidation streams is not recommended.

 

  8.

Arsenic concentrations in the samples averaged 0.29%, largely occurring as arsenian pyrite with only trace amounts of arsenopyrite.

 

  9.

Sulfide minerals were predominantly pyrite with some marcasite.

 

  10.

 Mercury concentrations ranged from 31 ppm to 138 ppm, averaging 81 ppm. These concentrations will require mercury capture and abatement equipment in the process flowsheet.

Mining and Infrastructure

 

  1.

The mine infrastructure has been completed.

 

  2.

Production ramp up has reached approximately 400 tons per day.

 

  3.

The mining contractor is in place with the full complement of equipment and personnel.

 

  4.

Decline development has accessed 700 vertical feet of mineralization of the Otto and Ogee zones. Development has reached the top of the South Pacific zone allowing additional active production stopes.

 

  5.

The drill lateral drift over the South Pacific zone has been completed.

 

  6.

Reconciliation of the model to mill indicates process head ounces exceed model by 19%. This appears to be from mining in a larger low grade halo around high grade core.

 

 

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 Page  36  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

  7.

Processed grade is lower than the life-of-mine planned grade due to extensive mining of marginal mineralization below the economic cutoff grade.

 

1.9.5

Recommendations

Metallurgical Testing

 

  1.

Establish sampling using the most recent mine plan to select samples to evaluate pressure oxidation with CIL cyanidation under Lone Tree conditions. Testing should also include baseline CIL tests and roasting testing as a comparison.

 

  2.

Testing should attempt to establish head grade and extraction relationships for use in more detailed resource modelling.

 

  3.

Mineralogy impacts need to be established and geologic domains within each resource need to be determined.

 

  4.

Additional comminution data should be collected to assess hardness variability within the zones and any potential impacts on throughput in the Lone Tree process plant.

 

  5.

The resource model should be advanced to include arsenic, TCM, TOC, mercury, as these will be important for predicting grades if toll process offsite is used and potentially for estimating extractions within the resources.

 

  6.

The estimated cost for the suggested next phase metallurgical program is to $350,000 based on current market pricing.

Resource Conversion and Exploration Drilling

 

  1.

Initiate and complete the infill drilling program in the South Pacific zone.

 

  2.

Update the mineralization model with the new drilling results.

Dewatering

 

  1.

Complete the planned dewatering well and re-evaluate the ground water model and inflow into the underground workings.

Mining

 

  1.

Stopes should be designed to mine parallel to strike wherever possible. Mining across strike results in additional planned dilution.

The width of stope production drifts in narrow mineralized areas should be decreased to 13 feet and smaller equipment mobilized to accommodate mineral resources.

 

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 i-80 Gold Corp   Summary   Page  37  

 

1.10

Open Pit

 

1.10.1

Mineral Resources

Table 1-3 shows the pit-constrained open pit Mineral Resource at a gold grade cutoff of 0.30 g/t.

Table 1-3: Granite Creek Mine Project Open Pit Mineral Resource

Class   Zone    Total Process
Material
(1000s
Tonnes)
   Total
Process
Material
(1000s
Tons)
   Au
Grade
(g/t)
   Au
Grade
(opt)
   Total Contained
Au (1000s t. oz)

Measured

  Pit B    2,910    3,207    1.32    0.042    123.41
  Pit A    563    620    1.07    0.034    19.30
  CX    10,889    12,003    1.30    0.042    455.27
  MAG    12,000    13,228    1.21    0.039    467.97
  Total    26,362    29,059    1.26    0.040    1,065.95

Indicated

  Pit B    360    397    1.10    0.035    12.73
  Pit A    689    760    0.80    0.026    17.78
  CX    2,973    3,277    1.25    0.040    119.62
  MAG    7,317    8,066    0.93    0.030    219.16
  Total    11,339    12,499    1.01    0.033    369.29

Measured

+

Indicated

  Pit B    3,270    3,604    1.29    0.042    136.14
  Pit A    1,252    1,380    0.92    0.030    37.08
  CX    13,862    15,280    1.29    0.041    574.89
  MAG    19,317    21,293    1.11    0.036    687.13
  Total    37,701    41,558    1.18    0.038    1,435.24

Inferred

  Pit B    32    36    0.64    0.021    0.67
  Pit A    205    226    0.59    0.019    3.88
  CX    1,347    1,485    1.16    0.037    50.24
  MAG    563    620    1.11    0.036    20.17
  Total    2,148    2,367    1.09    0.035    74.95

 

 

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Initial Assessment of the Granite Creek Mine,

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Please note that mineral resources are not mineral reserves and do not have demonstrated economic viability.

1.10.2 Mining Methods

Mine plans for the resource areas were designed and planned using conventional open pit mining method for the low grade, widely distributed gold. The open pit areas are suitable for phased designs.

1.10.3 Economic Analysis

GRE performed an economic analysis of the project by building an economic model based on the following assumptions:

 

   

Federal corporate income tax rate of 21%

   

Nevada taxes:

   

Proceeds of Minerals Tax – variable, with a maximum of 5% of Net Proceeds

   

Property tax – 2.5605%

   

Nevada gold and silver mine royalty – variable, with a maximum of 1.1% of gross revenue

   

Sales and use taxes are not included in the model

   

Equipment depreciated over a straight 7 or 15 years and has no salvage value at the end of mine life

   

Loss carried forward

   

Depletion allowance, lesser of 15% of net revenue or 50% of operating costs

   

Gold price of $2,175 per troy ounce

   

Gold recovery calculated per block as detailed in Section 13

   

Royalties on individual claims calculated by block, ranging from 0.02% to 7.5%, averaging 5.7%. There also is a 10% royalty applied to net profit.

 

1.10.3.1

Base Case

After analyzing the economic results of all cases considered, GRE selected the CIL only case with 0.85 g/t high grade cutoff, contractor operation, conventional tailings, and 133-tonne haul trucks and 21.9-tonne loaders as the base case as it results in the best overall economic results.

The economic model assumes a 1-year construction period. The time for permitting has not been included in the economic model, but the permitting for the open pit mine is likely to take three to five years and occur during the underground mining portion of the project.

Table 1-4 lists the key economic results for the selected scenario.

 

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 i-80 Gold Corp   Summary   Page  39  

 

Table 1-4 Key Economic Indicators

 

After Tax Economic Measure      Value  
   

After Tax NPV@5% (millions)

   $417.2
   

After Tax IRR

   28.7%
   

Initial Capital (millions)

   $254.7
   

Payback Period (years)

   3.72
   
All-in Sustaining Cost ($/oz Au Produced)    $1,227.4
   

Cash Cost ($/oz Au Produced)

   $1,180.5

Readers are advised that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability under S-K 1300. This IA is preliminary in nature and includes inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves under CIM Definition Standards. Readers are advised that there is no certainty that the results projected in this preliminary economic assessment will be realized.

 

1.10.3.2

Sensitivity Analyses

GRE evaluated the after-tax NPV@5% sensitivity to changes in gold price, gold grade, capital costs, and operating costs. The results indicate that the after-tax NPV@5% is most sensitive to gold price, moderately sensitive to operating cost, and least sensitive to capital cost (see Figure 1-1).

 

 

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Initial Assessment of the Granite Creek Mine,

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  Osgood Mining  Company LLC 

 

Figure 1-1 NPV@5% Sensitivity to Varying Gold Price, Gold Grade, Capital Costs, and Operating Costs

 

LOGO

 

1.10.3.3

Conclusions of Economic Model

The open pit project economics shown in the IA are favorable, providing positive NPV values at varying gold prices, gold grades, capital costs, and operating costs.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Introduction   Page 41 

 

2 Introduction

2.1 Registrant for Whom the Technical Report Summary was Prepared

This TRS was prepared for i-80 Gold Corporation and its subsidiaries Premier Gold Mines USA, Inc. and Osgood Mining Company (collectively i-80) in accordance with the requirements of the Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for i-80.

2.2 Terms of Reference and Purpose of this Technical Report

This TRS provides an initial statement of Mineral Resources for the Granite Creek Mine under regulation S-K 1300 and presents an Initial Assessment (IA) of the indicated and inferred mineral resources.

The quality of information, conclusions, and estimates contained herein are based on: i) information available at the time of preparation and ii) the assumptions, conditions, and qualifications set forth in this report. This IA is a preliminary technical and economic study of the economic potential of all or parts of mineralization to support the disclosure of mineral resources. This IA is preliminary in nature. It includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the Initial Assessment will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

2.3 Details of Inspection

This TR includes technical evaluations from three independent consultants. The consultants are specialists in the fields of geology, exploration, metallurgy, open pit and underground mining.

None of the Qualified Professionals (QPs) has any beneficial interest in i-80 or any of its subsidiaries, or in the assets of i-80 or any of its subsidiaries or in any property near the Granite Creek Project. The QPs will be paid a fee for this work in accordance with normal professional consulting practices.

The QP’s and the sections of this report each contributed to are listed in Table 2-2.

Table 2-2Table 2-1 summarizes the details of the personal inspections on the property by each qualified firm or, if applicable, the reason why a personal inspection has not been completed.

 

 

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  Osgood Mining  Company LLC 

 

Table 2-1 Personal Inspections by Qualified Professionals

 

 Company    Discipline    Dates of Personal
Inspection
   Details of Inspection
       
Practical Mining    Mining, Mineral Resources and Mineral Reserves    October 10, 2024    Site specific hazard training, examined core and core logging procedures, examined underground mine workings, observed core drilling operations, observed mining operations.
       
Raponi Engineering    Metallurgical Testing and Mineral Processing    None    The Granite Creek Mine does not have facilities for mineral processing.
       
Global Resource Engineering    Geology    April 20, 2021    General geological inspection of the Granite Creek area, including visual inspection of key geologic formations, lithologies, structural geology, and mineralization.
       
Global Resource Engineering    Mining, Mineral Resources    April 20, 2021    Examined infrastructure, pit walls, haul roads, examined core storage; examined conditions of underground workings
       
Global Resource Engineering    Metallurgical Testing and Mineral Processing    April 20, 2021    Examined infrastructure, pit walls, haul roads, examined core storage; examined conditions of underground workings

2.4 Sources of Information

This report is based in part on internal Company technical reports, previous studies, maps, published government reports, Company letters and memoranda, and public information as cited throughout this report and listed in 24 References.

2.5 Report Version

This TRS presents the inaugural statement of the Granite Creek Project mineral resources by i-80 under 17 CFR § 229.1300. The Company has most recently disclosed mineral resources for the project under Canadian Securities NI 43-101 regulations with the report Titled “Preliminary Economic Assessment NI 43-101 Technical Report Granite Creek Mine Project Humboldt County, Nevada, USA” dated September 27,2021.

 

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 i-80 Gold Corp   Introduction   Page  43  

 

2.6 Qualification of the Authors

This TR includes technical evaluations from three independent consultants. The consultants are specialists in the fields of geology, exploration, metallurgy, open pit and underground mining.

None of the Qualified Professionals (QPs) has any beneficial interest in i-80 or any of its subsidiaries, or in the assets of i-80 or any of its subsidiaries or in any property near the Granite Creek Project. The QPs will be paid a fee for this work in accordance with normal professional consulting practices.

The QP’s and the sections of this report each contributed to are listed in Table 2-2.

Table 2-2 QP Section Responsibility

 

Section

No.

   Title    Responsible Firm
     

1.1.

  

Introduction

   Practical
     

1.2.

  

Property Description

   Practical
     

1.3.

  

Geology and Mineral Deposits

   Practical
     

1.4.

  

Metallurgical Testing and Processing

   Raponi
     

1.5

  

Geology and Mineralization

   GRE
     

1.6

  

History

   GRE
     

1.7

  

Exploration, Drilling, and Sampling

   GRE
     

1.8

  

Data Verification

   GRE
     

1.9

  

Granite Creek Underground

   Practical
     

1.10

  

Granite Creek Open Pit

   GRE
     

2

  

Introduction

   Practical, Raponi, GRE
     

3

  

Property Description and Location

   Practical
     

4

   Accessibility, Climate, Local Resources, Infrastructure, and Physiography    Practical
     

5

   History    Practical
     

6

   Geologic Setting, Mineralization and Deposit    Practical
     

7.1.

   Exploration    GRE
     

7.2.

   Drilling    GRE
     

7.3

   Update to Drilling Statistics to Include i-80 Drilling and Land Package Expansion    Practical
     

7.4.

  

Hydrogeology

   Practical
     
8.1 – 8.9   

Sample Preparation, Analysis and Security

   GRE

 

 

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  Osgood Mining  Company LLC 

 

Section

No.

   Title    Responsible Firm
     

8.10 & 8.11

  

PM Discussions on QA/QC 2021 & 2022

  

Practical

     

8.12

  

Conclusions

  

GRE

     

9.1.

  

GRE Site Inspection (2021)

  

GRE

     

9.2.

  

Visual Sample Inspection and Check Sampling

  

GRE

     

9.3.

  

Database Audits

  

GRE

     

9.4.

  

QP Opinions on Adequacy

  

GRE

     

9.5.

  

Practical Mining Drillhole Database Verification

  

Practical

     

10.1.

  

Introduction

  

Raponi

     

10.2.

  

Metallurgical Test Work

  

Raponi

     

10.3.

  

Sample Representativity

  

Raponi

     

10.4.

  

Deleterious Elements

  

Raponi

     

10.5.

  

Geometallurgical Modeling

  

GRE

     

10.6.

  

Conclusions

  

GRE

     

10.7.

  

Recommendations

  

GRE

     

11.1.

  

Introduction

  

GRE

     

11.2.

  

Drill Hole Database

  

GRE

     

11.3.

  

Topography

  

GRE

     

11.4.

  

Geologic Model

  

GRE

     

11.5.

  

Open Pit Estimation

  

GRE

     

11.6.

  

Open Pit Resource

  

GRE

     

11.7.

  

Underground Mineral Resources

  

Practical

     

13.1.

  

Open Pit

  

GRE

     

13.2.

  

Underground

  

Practical

     

14.1.

  

Introduction

  

GRE

     

14.2.

  

Process Description

  

GRE

     

14.3.

  

Refractory Processing

  

Raponi

     

15

  

Infrastructure

  

Practical

     

16

  

Market Studies and Contracts

  

Practical

     

17.1.

  

Environmental Setting

  

GRE

     

17.2.

  

Environmental Studies and Issues

  

Practical

     

17.3.

  

Social or Community Impacts

  

Practical

     

17.4.

  

Permits

  

Practical

     

17.5.

  

Water Use Permits

  

Practical

     

17.6.

  

Environmental Permits

  

GRE

 

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 i-80 Gold Corp   Introduction   Page  45  

 

Section

No.

   Title    Responsible Firm
     

17.7.

  

Mine Closure

   GRE
     

17.8.

  

Local Procurement and Hiring

   GRE
     

18.1

  

Open Pit Capital Cost Estimate

   GRE
     

18.2

  

Open Pit Operating Cost Estimate

   GRE
     

18.3.

  

Granite Creek Underground

   Practical
     

19.1.

  

Taxes

   Practical
     

19.2.

  

Granite Creek Underground

   Practical
     

19.3.

  

Open Pit

   GRE
     

20

  

Adjacent Properties

   Practical
     

21

  

Other Relevant Data and Information

   Practical, Raponi, GRE
     

22

  

Interpretation and Conclusions

   Practical, Raponi, GRE
     

23

  

Recommendations

   Practical, Raponi, GRE
     

24

  

References

   Practical, Raponi, GRE
     

25

  

Reliance on Information Provided by the Registrant

   Practical, Raponi, GRE
     

25.1.1

  

Environmental Recommendations

   GRE

2.7 Sources of Information

Information sources are documented either within the text and cited in references or are cited in references only. The authors believe the information provided by i-80 and to be accurate based on their work on the project.

2.8 Units of Measure

US Imperial units of measure are used throughout this document unless otherwise noted. US/Metric conversion factors are listed in Table 2-3. Currency is expressed as United States Dollars unless otherwise noted.

Table 2-3 Units of Measure Conversion Factors

 

US Imperial to Metric Conversions
   

Linear Measure

 

Weight

   

1 inch = 2.54 cm

 

1 short ton = 2,000 lbs = 0.9071 metric tons

   

1 foot = 0.3048 m

 

1 lb = 0.454 kg = 14.5833 troy oz.

   

1 mile = 1.6 km

 

Assay values

   

Area Measure

 

1 ounce per short ton = 34.2857 g/t

   

1 acre = 0.4047 ha

 

1 part per million = 0.0292 opt

   

1 square mile = 640 acres = 259 ha

 

1 troy oz. = 31.10348 g

   

Density

   
   

1 tonne/m3 = 0.0312 tons per ft3

   

 

 

 

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Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

  Osgood Mining  Company LLC 

 

2.9 Coordinate Datum

Spatial data utilized in analysis presented in this TR are projected in the Granite Creek local mine grid unless noted otherwise.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Property Description and Location   Page  47  

 

3 Property Description and Location

3.1 Property Description

The Granite Creek Project is located in Humboldt County, Nevada, 28 miles northeast of the town of Winnemucca, and it is part of the historic Potosi mining district. It is centered at roughly 41° 8’ N latitude and 117° 15.5’ W longitude. It encompasses about 4,506 acres (1,823.5 hectares) including owned unpatented claims, leased unpatented claims and owned surface fee land. The federal land is administered by the BLM. Figure 3-1 shows the location of the Granite Creek Project.

Figure 3-1 Granite Creek Project Location

 

LOGO

3.2 Status of Mineral Titles

Ownership of the Granite Creek Project land position comprises various forms of title. Figure 3-2 shows the Granite Creek land position. i-80 owns 48 unpatented lode claims covering about 897 acres (Table 3-2), and leases 56 unpatented lode claims covering about 1,007 acres (Table 3-3).

 

 

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Initial Assessment of the Granite Creek Mine,

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Osgood Mining 

Company LLC 

 

The lease expires May 9, 2040. i-80 also owns, through its subsidiaries, fee surface land parcels covering about 2,602 acres.

Unpatented claims have annual maintenance fees of $200 per claim payable to the BLM and a notice of intent to hold (NIH) in the amount of $12 per claim plus $12 filing fee per document payable to Humboldt County. Claim maintenance fees are paid through September 2025 with the BLM. The NIH was paid to Humboldt County on July 9, 2024; payments are current at the time of this report. Fee land is subject to Nevada state real property tax, and certain mine infrastructure is subject to Nevada state personal property tax. Leased unpatented claims are subject to yearly lease fees. Holding costs for 2025 are listed in Table 3-1.

Table 3-1 Holding Costs

 

Description    Payee    Quantity     Amount  
Unpatented Claim Maintenance Fee    BLM    104      $20,800.00  
Notice of Intent to Hold Unpatented Claims    Humboldt County       104      $1,260.00  
Real Property Taxes    Humboldt County    5 parcels      $7,234.63  
Personal Property Taxes    Humboldt County    various
infrastructure
     $94,594.15  
Lease fees, annually adjusted by CPI    Lease Holders    56 unpatented
claims
     $122,495.24  
       
Total            $246,384.02  

 

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 i-80 Gold Corp   Property Description and Location   Page  49  

 

Figure 3-2 Granite Creek Land Position Map

LOGO

 

 

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Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Table 3-2 Granite Creek Owned Unpatented Claims

       
Claim Name    BLM Legacy Number      Claim Type     

Number of

Claims

 
   
PACIFIC # 1A - PACIFIC # 7A      NMC319814 - NMC319820        Lode        7  
   
CX # 1A - CX # 23A      NMC319833 - NMC319855        Lode        23  
   
PINSON # 1A - PINSON # 18A      NMC319856 - NMC319873        Lode        18  
       
Total Owned Patented Claims                        48  

Table 3-3 Granite Creek Leased Unpatented Claims

       
Claim Name    BLM Legacy Number      Claim Type     

Number of

Claims

 
   
NEW BEE DEE FRAC. #1, #2      NMC282121, NMC282122        Lode        2  
   
BEE DEE # 21 - BEE DEE # 56      NMC282123 - NMC282158        Lode        36  
   
BEE DEE # 1A      NMC319892 - NMC319909        Lode        18  
       
Total Leased Patented Claims                        56  

 

3.2.1

Royalties

Several royalties are in effect on various areas of the property. Table 3-4 lists the royalties and Figure 3-3 shows the royalty areas. Some royalties were retained by previous owners upon sale of the property while others were negotiated as lease agreements with claim holders.

Table 3-4 Granite Creek Royalties

 

   
Lessor/Grantor    Lease Type
   
Gold Royalty U.S. Corp.    10% NPI
   
Franco-Nevada    2% NSR
   
Franco-Nevada and S&G Pinson    2% NSR
   
proportionate to individual interests Noceto/Phillips/K. Murphy    3.1249% NSR
   
proportionate to individual interests Noceto/Phillips/K. Murphy/D. Christison/J. Christison/M. Murphy    2% NSR
   
0.166667% (0.5 of 1/12th of 2%) NSR to Franco-Nevada and S&G Pinson    0.166667% NSR  
   
Nevada Gold Mines    0.5% NSR
   
Royal Gold    3-5% NSR
   
1996 Sliding scale royalty with successors to agreement: Royal Gold and D.M. Duncan    1-4% NSR

 

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 i-80 Gold Corp   Property Description and Location   Page  51  

 

Figure 3-3 Granite Creek Royalty Map

LOGO

 

 

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Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

3.3 Environmental Liabilities

The reclamation and closure cost for the Granite Creek Mine surface and underground is currently estimated to be approximately $3 million (Osgood Mining Company LLC [Osgood], 2023). A bond in the amount of approximately $2 million is held by the Bureau of Land Management (BLM) to address reclamation activities associated with the underground mine while an additional approximately $1 million bond is held for surface reclamation activities. There are no other known environmental liabilities associated with pre-Project operations (Osgood, 2024).

3.4 Permits/Licenses

The existing permits for the Granite Creek Mine are outlined in Section 17.3 and 17.4. Surface Reclamation Permit #0047 authorizes 448.6 acres of disturbance on private land and 490.4 acres of disturbance on public land while Underground Reclamation Permit #0242 authorizes 88.9 acres of disturbance on private land and 0.6 acres of disturbance on public land. These authorized disturbance acres allow Osgood to conduct the exploration, geotechnical and metallurgical field work to support the study work recommended in this Report as long as the amount of new surface disturbance remains less than that available under the existing authorizations.

Section 17 describes the permits required to execute the mine plan discussed in this report.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Accessibility, Climate, Local Resources, Infrastructure, and Physiography   Page  53  

 

4

Accessibility, Climate, Local Resources, Infrastructure, and Physiography

4.1 Accessibility

The project can be reached by traveling east on US Interstate Highway 80 from Winnemucca to the Golconda exit, about 15 miles, then following Nevada State Route 789 and the Getchell Mine Road northeast about 20.5 miles to the Granite Creek Mine access road. The last 4.5 miles is unpaved, well maintained gravel road. Areas of Route 789 are designated open range, and travelers must watch for cattle. The Granite Creek security facility is located about 0.2 miles west of the intersection with Route 789. Traveling from the east, the Golconda exit lies west of Battle Mountain about 36.5 miles on Interstate Hwy 80. The route from Battle Mountain passes i-80’s Lone Tree facility, which lies south of the interstate about 19 miles west of Battle Mountain.

4.2 Climate

The climate in Humboldt County is typical of the high-desert environment. Typical summer temperatures average roughly 75°F with occasional day/night extremes of 105°F/40°F. Winter temperatures average roughly 30°F with occasional day/night extremes ranging 60°F/-10°F. Average annual precipitation is about 8 inches, the majority of which accumulates as snowfall during the winter months. Typical snow accumulation is roughly 3 inches on average at lower elevations, although occasional large storms may accumulate significantly more for short durations.

Mining operations are able to continue year-round with brief pauses for summer lightning storms or unusually heavy winter snowstorms.

4.3 Local Resources

The town of Winnemucca has a population of about 8,400. Basic services are available. Local mining districts have been active since the 1980’s, and mining suppliers and contractors are accustomed to working in the area. Some experienced and general labor is available locally and from other small towns in the region. There are a number of mining operations in the region and as such there is always competition for employees.

 

 

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Osgood Mining 

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4.4 Infrastructure

Existing infrastructure at the Project includes an office building, dry and warehouse facilities, and a lined stockpile area on the surface. Over 9,000 feet (2,743 meters) of underground workings have been completed, and four deep dewatering wells were drilled and cased, two of which are currently being operated.

Electrical infrastructure suitable for mine operations is installed, and two re-infiltration basins and associated pipelines have been constructed to re-infiltrate water produced in mine dewatering into the valley aquifer.

The mine is accessed through either of two portals, and dual egress has been established for most areas of the mine. Where dual egress is not possible, rescue chambers have been installed. Equipment is repaired in an underground mine shop. Air doors and a ventilation fan provide required air supply to the workings in compliance with Mine Safety and Health Administration (MSHA) standards.

Landline telephone and digital subscriber line service are available at the Project site. Cellular phone service is also available, but is dependent on the strength of receiving antennas, topography, and lines of sight.

4.5 Physiography

The Project lies in the Basin and Range Province, a structural and physiographic province comprised of generally north to north-northeast trending, fault bounded mountain ranges separated by alluvial filled valleys.

The Project is located on the eastern flank of the Osgood Mountains. Topography is gentle to moderate at the Project, ranging from an elevation of 4,840 feet at the mine offices to 5,500 feet at the crest of the historic CX pit highwall, and rises steeply to the west over 3,800 feet to Adam Peak at the top of the Granite Creek drainage. Vegetation is typical of the high desert with sagebrush on the alluvial fans, and juniper on the mountain slopes.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   History   Page  55  

 

5

History

The Property has been explored by a number of individuals and mining/exploration companies since the late 1930s. The original discovery on the Property was made by Clovis Pinson and Charles Ogee in the mid to late-1930s, but production did not occur until after World War II, when ore from the original discovery was shipped to and processed at the Getchell mine mill. In 1949 and 1950, total production from the Granite Creek mine amounted to approximately 10,000 short tons (9,071 tonnes) grading approximately 0.14 ounces per ton (opt) (4.8 g/t).

5.1 Historic Ownership

5.1.1 Cordex I Syndicate

The Property remained functionally dormant from 1950 until 1970, when an exploration group known as the Cordex I Syndicate (John Livermore, Peter Galli, Don Duncan, and Rayrock Resources) leased the Property from the Christison Family (descendants of Mr. Pinson and Property owners), on the strength of its similarity to the Getchell Property and structural position along the range-front fault zone bordering the Osgood Mountains. Following a surface mapping and sampling program in 1971, 17 reverse circulation (RC) drillholes were completed in and around the 1940s era Granite Creek Mine pit, confirming low- grade gold values. An 18th step-out hole encountered a 90-foot (27.4-meter) intercept of 0.17 opt (5.8 g/t) gold (Au). This intercept was interpreted as a subcropping extension of known mineralization northeast of the original pit and was the basis for delineation of what would become the “A” Zone at the Property, a 60-foot (18-meter) by 1,000-foot (305-meter) shear zone. During the late 1970s, the Cordex I Syndicate reorganized into a Nevada Partnership known as PMC, with Rayrock Resources as the Project operator, and began production at the Property.

Cordex Syndicate (Cordex), and its successor, PMC, explored the Property largely through mapping and geochemical sampling. There are three known mapping programs:

•  A regional mapping program from Preble to Getchell by Pete Chapman in the late 1970s

•  A 1:6000-scale mapping program of the Property in 1983

•  A 1:2400-scale mapping program of the Pit areas through the active life of the mine

5.1.2 Pinson Mining Company

PMC began developing the A Pit in 1980 and produced gold the following year. Production from the B Pit began in 1982. Step-out drilling in 1982 to 1983 to the northeast of the A Zone intersected two more discrete zones: the C Zone extending east-northeast from the A Zone and the CX Zone

 

 

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extending northeast from the C Zone. Step-out drilling northeast of the CX Zone in 1984 located an apparently independent fault system (striking north-northwest), dipping steeply east that became the core of the Mag deposit, which went into production in 1987. PMC produced from the CX, CX West and Mag Pits into the mid to late 1990s, until a combination of falling gold prices and erratic mill feed forced closure of the oxide mill in early 1998. Continued attempts to expand production of oxide ore failed, and all active mining ceased on 28 January 1999 (McLachlan, et al., 2000). The project was officially closed in May 2000.

5.1.3 Homestake – Barrick

In the 1990s, Homestake and Barrick became 50/50 partners in PMC through purchase of minority interests (McLachlan, et al., 2000). Homestake and Barrick conducted an exploration program from 1996 to 2000 through PMC, expending some $12M on the Project. The joint venture explored the deeper feeder fault zones of the Property, exploring for a large, high-grade gold system that would support a refractory mill complex. This work, while successful in identifying gold mineralization with underground grades, failed to identify a deposit of sufficient size to be of development interest to Homestake or Barrick, and the partners concluded the exploration program. Subsequent to that decision, in 2003, Barrick acquired Homestake and drilled an additional three exploration drillholes.

5.1.4 Atna Resources Ltd. Earn-in and PMC Back-in

In August 2004, Atna acquired an option to earn 70% Joint Venture interest in the Property from PMC, a wholly owned subsidiary of Barrick, and commenced additional follow-up exploration and development of the Property. Atna completed its earn-in in 2006 and vested in its 70% interest in the Project after expending the required $12M in exploration and development expenditures. PMC elected to back-in to the Project and re-earn an additional 40% interest (bringing PMC’s interest to 70% and Atna’s to 30%) on 5 April 2006. PMC spent over $30M on the Project during the next three-year period and completed its “claw-back” in early 2009. Their work included surface and underground diamond core drilling, RC drilling, underground drifting, and surface infrastructure construction (rapid infiltration basins, mineralized material stockpile pad, underground electrical service upgrades, etc.). A new mining joint venture was formed in 2009 reflecting the Project’s ownership, with PMC owning a 70% interest in the venture and Atna owning a 30% interest. PMC, as the majority interest owner, was the operator of the joint venture.

5.1.5 Atna 2011 – 2013 Underground Development

In September 2011, Atna negotiated the acquisition of PMC’s 70% joint venture interest in the core property position at the Granite Creek Mine Project. The asset purchase and sale agreement included all right title and interest to the core property described above as well as an evergreen processing agreement with Barrick for the processing of underground refractory ores from Granite Creek Mine at Barrick’s Goldstrike facilities.

 

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Development of the Granite Creek Mine underground commenced in early 2012, and mine ramp-up began in late 2012. In total, 6,011 feet (1,832 meters) of primary and secondary development were completed during 2012 and 2013. The primary spiral ramp was driven to the 4530 level from the 4650 adit level, and both top cut and underhand ore mining occurred in three Ogee-zone stope blocks during development. Additional secondary access drifts were in progress when the mine was placed on care and maintenance to access the Range Front and Adams Peak mineral zones but were not completed prior to cessation of underground work. Mining was performed by contract miners using underground mining equipment owned by the contractor. Approximately 30,000 short tons (27,216 tonnes) of ore containing 7,900 oz of gold were mined and shipped to off-site processing facilities.

Work on the Project continued until June of 2013, when the mine was placed on care and maintenance. This decision was driven by a number of factors, including the steep decline in the gold prices in 2013.

In May 2014, the status of the underground mine was changed to an intermittent production status. Under this status, periodic mining of ores from stoping areas developed in 2013 was conducted to develop and test revised stoping methods for the underground and to prove mining economics at small production rates.

5.1.6 Osgood Mining Company LLC Acquisition

In 2016, OMC, a wholly owned subsidiary of Waterton Global Resources Management, acquired the Project. OMC completed numerous drillhole database compilation and verification campaigns, beginning with migration of the ATNA database to Maxwell Datashed Database software in 2017 and database verification and improvement efforts in 2018. In 2016, OMC, with an external consultant, completed a project-scale structural geology study that included surface and underground mapping, historical data review, and cross section interpretation aimed at defining the main structural architecture at Granite Creek Mine and developing exploration and resource drilling targets. This work formed the basis of an updated 3-dimensional (3D) litho-structural model that was used for the 2020 Mineral Resource estimation (AMC, 2020). From 2017 to 2018, OMC also completed an extensive drill material inventory and salvage program that secured the available drill core and RC chips on the property.

OMC continued to maintain compliance and keep all environmental permits for the site in good standing. This included performing permit-related sampling and reporting, as well as renewing permits. In addition, OMC performed regular inspections of the site. During the ownership period,

 

 

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OMC worked with the State of Nevada to close out a Water Pollution Control Permit for a reclaimed portion of the mine, reducing the overall compliance monitoring and reporting liabilities for the operator. In addition, OMC received approval from the State to remove portions of the reclaimed site from the bond.

In addition to these geology and compliance activities, OMC continued to maintain and improve site infrastructure, including a third-party review of hydrology and dewatering requirements that resulted in the replacement of pumps (2019) and the upgrading of two dewatering well process controls. Rapid infiltration basins have been maintained as needed, with water flows being tracked and monitored.

5.1.7 i-80

In April 2021, i-80 Gold Corp was created as a spinout of Premier Gold Mines Limited’s assets located in Nevada concurrent to Equinox Gold Corp’s acquisition of most of the balance of Premier Gold Mines Limited’s assets in Canada and Mexico. The same month, the newly created i-80 Gold Corp completed acquisition of OMC from Waterton Global Resources Management. In May 2021, additional land was purchased by i-80, further increasing the size and ownership in the land package.

In June 2021, Section 31 fee land was acquired by Premier Gold Mines USA, Inc. from Seven Dot Cattle Co., LLC., as well as Christison interest in Section 28 fee lands, and in the PINSON unpatented claims. (Note these are still held in the Premier name, not Osgood.) In 2022, Section 21 fee land T. 38 N., R. 42 E. was acquired by Osgood Mining Company from Nevada Gold Mines, and lessee interest on the BEE DEE unpatented claims in Section 6, T. 37 N., R. 42 E.

5.2 Historical Mine Production

Historically, the Granite Creek Mine Project, with small additions from the nearby Preble and Kramer Hill mines, was credited with gold production in excess of 1 million ounces and less than 100,000 oz of silver (Tingley, 1998). PMC independently compiled a record of production and credited the Granite Creek Mine Property with production of 986,000 oz of gold through 1999.

 

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6

Geologic Setting, Mineralization and Deposit

6.1 Regional Geology

The Property is located on the eastern flank of the Osgood Mountains within the Basin and Range tectonic province of northern Nevada. The Granite Creek Mine, together with the Preble, Getchell, Turquoise Ridge, and Twin Creeks mines, are on what is referred to as the Getchell gold trend (Getchell trend). The main Getchell trend generally strikes northeast-southwest and has been cross-cut by secondary north-south and northwest-southeast-trending structures. The deposits are hosted in Paleozoic marine sedimentary rocks. The rocks are exposed in the Osgood Mountains and have been complexly thrust faulted (Hotz, et al., 1964) and intruded by the Cretaceous-aged (92 Ma) (Silberman, et al., 1974) Osgood Mountains granodiorite stock. These units are unconformably overlain by Miocene volcanic rocks. Figure 6-1 is a regional geologic map of the Osgood Range.

 

 

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Figure 6-1 Regional Geologic Map of a Portion of the Osgood Mountains including Granite Creek

 

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The Osgood Mountains Range is underlain by Cambrian Osgood Mountain Quartzite, Cambrian Preble Formation, Ordovician “Comus” Formation and the “upper plate” Valmy Formation. These units are unconformably overlain by the Permian Etchart Formation (Antler Peak Equivalent) of the Roberts Mountains overlap assemblage, and by the Triassic Golconda allochthon. These uppermost units form a belt of outcrops flanking the western and northern sides of the Osgood Range. These rocks have been intruded by the Cretaceous-aged Osgood Mountains granodiorite stock, which forms the core of the Osgood Mountains. Stratigraphy throughout the Osgood Mountains plunges north (Chevillon, et al., 2000). A significant thermal metamorphic aureole surrounds the stock. At least four Paleozoic units, defined by structure, lithology, and age comprise the Osgood Mountains (McLachlan, et al., 2000). These include the:

•  Autochthonous Cambrian Osgood Mountains Quartzite and Preble Formation and Cambrian to Ordovician Comus Formation

•  Allochthonous Ordovician Valmy Formation, part of the Roberts Mountains allochthon

•  Antler overlap sequence including the Mississippian Goughs Canyon Formation, Pennsylvanian Battle Formation, and Pennsylvanian-Permian Etchart Limestone

•  Allochthonous Pennsylvanian-Permian Farrel Canyon Formation, part of the Golconda allochthon

The autochthonous Cambrian-Ordovician package has been described by Jones (1991, cited in McLachlan et al. 2000) and is comprised of the Osgood Mountains Quartzite, Preble Formation, and Comus Formation. All of these units have undergone regional metamorphism and intense, northwest-directed folding (McLachlan, et al., 2000). At the Getchell Project, these two units are folded together to form the northwest-verging Pinson anticline. The Comus and Preble Formations show distinct facies changes across the district. These units at Turquoise Ridge and Twin Creeks contain tuffs, pillow basalts, and mafic sills, none of which are present in the same units at Granite Creek.

The Roberts Mountains allochthon described by Stenger et al. (1998) is exposed at the Turquoise Ridge and Twin Creeks mines where it has been mapped as the Valmy Formation. The Roberts Mountains allochthon is composed of a thick (>980-foot [299-meter]) sequence of mid-ocean ridge basalts and intercalated pelagic sediments that have been thrust over the Twin Creeks member of the Comus Formation (Stenger, et al., 1998). This sequence has not been identified at the Granite Creek Mine but was likely present and eroded prior to the present day.

 

 

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The Antler overlap sequence in the Osgood Mountains consists of the Pennsylvanian Battle Formation, and Pennsylvanian-Permian Etchart and Adam Peak formations (McLachlan, et al., 2000). The Battle conglomerate consists of cobbles and pebbles of quartzite. The Etchart lies conformably on the Battle and consists of calcareous sandstone underlying fossiliferous limestone. South of the Getchell Project, the Battle and Etchart lie unconformably on the Preble Formation and Osgood Mountain Quartzite (McLachlan, et al., 2000). These units are not present at the Granite Creek Mine Project.

The Golconda allochthon comprises the Mississippian Goughs Canyon and the Pennsylvanian Permian Farrel Canyon formations present along the northwest flank of the Osgood Mountains. The thrust strikes north to northeast from the central part of the range to the Dry Hills in the north (McLachlan, et al., 2000). These units are not present at the Granite Creek Mine Project.

6.2 Local and Property Geology

The geology throughout the Osgood Mountains is typified by folded Cambrian to Ordovician sedimentary rocks that have been intruded by Cretaceous stocks, which have been cross-cut by later high-angle structural deformation. Hotz and Willden (Geology and Mineral Deposits of the Osgood Mountains Quadrangle, Humboldt County, Nevada: U.S. Geological Survey Professional Paper 431, 1964) suggest the high angle faulting is related to the Basin and Range extension. The older rocks are overlain by Miocene andesitic basalt and the surrounding fault bounded basins are filled with quaternary alluvial (Qal) gravel. The Osgood Mountains have a general northeast trend, although, at a structural hinge in the vicinity of the Granite Creek Mine, the east flank of the range rotates and trends north towards the Getchell mine. Gold mineralization is primarily hosted by fine-grained marine sedimentary rocks that overlie a large stock of Cretaceous granodiorite.

Throughout the district Cambrian to Ordovician siliciclastic and carbonate rocks have been intruded by the Cretaceous Osgood Mountains granodiorite, resulting in the formation of large, metamorphosed aureoles with development of several tungsten-bearing skarns. The lowest stratigraphic units recognized locally are the Cambrian Osgood Mountains Quartzite, which is overlain by phyllitic shales, limestone interbeds, and various hornfelsed sedimentary rocks of the Preble Formation. The Preble is overlain by Ordovician sedimentary rocks of the Comus Formation, both of which have been folded into a broad, north-plunging anticline. The west flank of the anticline has been over-thrust by the Ordovician Valmy Formation, which consists of deep-water siliceous shales and cherts. The core of the anticline and scattered localities along the east side of the Osgood Mountains are unconformably overlain or in fault contact with sandstones and conglomerates of the Battle Formation and limestones of the Etchart Formation. The Golconda and Humboldt thrusts displaced Mississippian volcanics and Pennsylvanian shales eastward along the northwest and southern flanks of the Osgood Mountains. Extension during the Tertiary resulted

 

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in outflows of Miocene rhyolitic tuffs, basalts, andesite flows, and younger Quaternary basalt flows.

Gold mineralization at the Property is primarily hosted in the Comus Formation, as shown in Figure 6-2.

 

 

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Figure 6-2 Granite Creek Stratigraphic Column

 

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The stratigraphy of the Osgood Mountains from youngest to oldest is:

 

   

Quaternary: Qal / Qb – Alluvium and basalt

 

   

Tertiary:

 

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Tba – Andesite and basalt flows. Dark green to black aphanitic and weakly porphyritic flows, flow breccia

 

   

Tr – Rhyolitic tuffs. Pumice, welded, reworked, tan to white

 

   

Tcg – Chert, shale, rhyolite clasts in a sandy matrix

 

   

Tbi - Dacite and andesite dikes

 

   

Cretaceous: Kgd – Granodiorite, quartz diorite. Equigranular, medium grain intergrowths of feldspar, quartz, biotite, and hornblende.

 

   

Permian / Pennsylvanian:

 

   

PPmh – Havallah Formation. Interbedded sandstone, chert, shale, siltstone with minor volcanic flows and pyroclastics. Chert, interbedded with sandstone composes up to 50% of the unit.

 

   

PPe – Etchart Limestone. Limestone, sandy limestone, dolomite. Lower portion is sandy limestone with local pebble conglomerate. Upper portion is pure limestone with interbedded dolomite and sandy dolomite. Minor calcareous shale.

 

   

Pennsylvanian: Pb – Battle Formation. Poorly bedded, poorly sorted boulder and pebble conglomerate with coarse-grained sandstone and minor limestone clasts composed of Osgood Quartzite and chert in a shaley to sandy matrix.

 

   

Ordovician:

 

   

Ov – Valmy Formation – Chert, shale, quartzite, volcanics (greenstone). Interbedded chert and shale with quartzite greenstone bed on the east side of the Osgood Mountains. Quartzite is dominant in the lower portion and chert and shale in upper portion.

 

   

Oc – Comus Formation – Upper unit composed of black argillite generally lacking bedding. Lower units composed of alternating thin to medium beds of limestone and argillite. In the Twin Creeks mine area, pillow basalts, mafic igneous sills and dikes exist within the sequence. Mafic igneous rocks are not present in the Comus Formation at the Granite Creek Mine.

 

   

Cambrian:

 

   

Cp – Preble Formation – Dominantly sandstone phyllitic shale. Maroon, light olive, and brown. Upper part contains thin interbeds of limestone rhythmically bedded with shale.

 

   

Com – Osgood Mountain Quartzite. White, gray, light brown, purple-brown to green- gray, medium to thick-bedded quartzite. Impure quartzite, silty sandstone, phyllitic shale.

The Granite Creek Mine is located on the eastern flank of a large Cretaceous granodiorite stock that forms the southern core of the Osgood Mountains. Rocks adjacent to the eastern side of the stock have a general east dip and strike sub-parallel to the trend of the Osgood Mountains. The oldest units exposed against the granodiorite are Cambrian Preble sandstone, phyllitic shales, and

 

 

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interbedded limestones all of which are often metamorphosed. Overlying the Preble is a thick package limestone and argillite of the Ordovician Comus Formation. The Lower Comus is composed of thin to medium interbeds of limestone and argillite. The Upper Comus consists of black argillite typically lacking bedding.

A Cretaceous aged (90 – 92 million years [Ma]) (Silberman, Berger, & Koski, 1974) granodiorite stock intrudes the Paleozoic section in the southern half of the Osgood Mountains. Emplacement of the stock resulted in the formation of an irregular contact metamorphic aureole, which extends as much as 10,000 feet (3,048 meters) from the intrusive contact. The metamorphic event resulted in the formation of maroon-colored, biotite-cordierite- hornfels in the Preble Formation and chiastolite hornfels in the Upper Comus Formation within much of the Property area (McLachlan, Struhsacker, & Thompson, 2000). In addition, carbonate rocks were metamorphosed to marble and calc-silicates (wollastonite, garnet, diopside, and vesuvianite). Several tungsten-bearing skarn deposits were also formed along the margins of the stock (Silberman, Berger, & Koski, 1974). Two tungsten skarns are on the Property.

Outcrop mapping and historic drilling has revealed the presence of extensive folding of the Paleozoic section in the Osgood Mountains. The most prominent of these folds is the Pinson Anticline. The fold is northeast-plunging and northwest-verging and extends for a distance of approximately three miles southwest from the Granite Creek Mine (McLachlan, Struhsacker, & Thompson, 2000). Numerous parasitic folds have also been noted along the limbs of the anticline. Where exposed, the Pinson Anticline is cored by the Cambrian Preble Formation and flanked on the northwest and southeast by sediments of the Ordovician Comus Formation.

Mineralization on the Property exhibits strong structural control. A wide variety of mineralized structural orientations have been documented. The most important structural feature on the Property is the network of faults that border the escarpment marking the southern and eastern edge of the Osgood granodiorite (Sim, 2005). This fault system has been variably interpreted as a single master fault (RFF) (McLachlan, Struhsacker, & Thompson, 2000) that curves around the stock, or more likely, a network of shorter, straighter segments that collectively accommodate several thousand feet of displacement while making a 50° bend around the southeast corner of the stock (Sim, 2005). The fault system can be divided into three structural and stratigraphically mineralized zones, with each mineralized zone defined by one or more major structural elements. These are referred to as the Rangefront, CX, and Mag Zones. Sedimentary rocks in the vicinity of this system generally dip steeply (easterly) away from the contacts of the granodiorite (Sim, 2005).

Figure 6-3 Geology and Structural Map shows the structural and geology map of the Property with the mined-out pits outlined for reference.

 

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Figure 6-3 Geology and Structural Map

 

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In addition to this large-scale fault system, there are numerous northwest and a few east-west structures that have been identified by past mapping and drilling (McLachlan, Struhsacker, & Thompson, 2000). In general, these appear to be mostly older than, and truncated by, the main system. Some of these faults have been re-activated and disrupt the continuity of the main Granite Creek system (Sim, 2005).

6.3 Structural Framework

 

6.3.1

Structural Overview

In 2022, i-80 Gold Corp. created an updated comprehensive geologic model for the Granite Creek Mine Project. This work included surface and underground mapping, structural analyses, geologic interpretation, and the creation of a complete 3D geologic model using Leapfrog Geo. The 3D model utilizes all available data including: drillholes (lithology, assays, etc.), surface mapping (pit and regional), underground mapping (current and historic), televiewer data (interpreted in-house), and structural analyses (stereonet, cross section, etc.).

Structure at the Granite Creek Project is highly complex and indicative of multiple deformation events. Regional deformation events, such as the Antler, Sonoma, and Elko orogenies, are likely responsible for the numerous overprinted fabrics and compressional structures observed at Granite Creek. These compressional structures appear to have been dissected and/or reactivated by subsequent Basin and Range extension. In 2016, Robert Leonardson generated a geologic model for the property that describes a west-northwest-verging imbricate thrust system deflected around the Osgood Stock. i-80 Gold Corp recognizes this thrust system comprised of the Rangefront, Adam Peak, Otto, and CX faults but interprets most of the compression and westward transport along these faults to have pre-dated the emplacement of the stock. The main structural element on the property is the Rangefront fault with a variable strike of 045o in the south to 010o in the north, and a dip of 60o near-surface that shallows with depth. The Adam Peak and Otto faults act as hanging wall splays off the northern extent of the Rangefront fault. The CX fault has a strike of 050o with a 55o near-surface dip that shallows with depth. Bedding generally dips steeply to the northeast, however fold geometries in the hanging wall of the Rangefront fault are complex and polyphase resulting in non-cylindrical interference folds. A property scale northeast trending doubly plunging upright anticline in the hanging wall of the CX fault is interpreted as a fault propagation fold related to west-northwestward compression along the Rangefront and CX faults. The current fault and fold geometries are interpreted to be the result of displacement and top-to-the-east rotation of the country rock by the emplacement of the Osgood Stock. Rotating the main thrust faults from their current ~60o dips back to an assumed syn-compressional dip of ~30o rotates the upright anticline’s axial plane orientation to ~045o, 60o, resulting in typically observed compressional geometries. Reactivation of these thrust faults in their current orientation during Basin and Range extension has resulted in normal-sense down-to-the-east displacement across the

 

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property. The Mag fault system on the eastern portion of the property trends 335o and appears to be a younger fault system associated with Tertiary extension. The two main faults of the system are the Mag and Mag West faults that form a horst, with the Mag fault having significant down-to-the-east displacement.

Figure 6-4 Geology and Structural Map of the Granite Creek Property

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The following subsections give details on significant structural features observed across the Property. Pit structural mapping by Chadwick (2002) collected orientation data and cross-cutting relationships.

 

  7.3.2

 Faults

Rangefront Zone

The Rangefront Zone (RFZ) is a northeast trending fault zone that forms a broad persistent zone of shearing and brecciation along the RFF that bounds the eastern margin of the Osgood Mountains. The RFZ involves the entire stratigraphic sequence at the Property, including the Cambrian Preble, Ordovician Comus, and Cretaceous granodiorite.

Rangefront Fault

The RFF is a prominent 010° to 045° striking normal fault that defines the eastern front of the Osgood Mountains. For much of its length, the fault juxtaposes the Comus Formation in the hanging wall against the Preble Formation in the footwall (McLachlan, et al., 2000). The fault originated as a west-verging thrust fault and has since been reactivated as a down-to-the-east normal fault of significant yet unknown displacement. The hanging wall zone is intensely brecciated and pervasively argillized. The fault has a near-surface dip of 60° that shallows with depth. The hanging wall delineates the lower boundary of the RFZ.

Adam Peak Fault

The Adam Peak fault is a 048° striking hanging wall splay off the Rangefront fault with a near-surface dip of 72° that shallows with depth. The fault has been reactivated as a normal fault with unknown displacement. The footwall delineates the upper boundary of the RFZ.

Otto Fault

The Otto fault is a 040° striking hanging wall splay off the Rangefront fault with a near-surface dip of 80° that shallows with depth. The fault has been reactivated as a normal fault with unknown displacement. The fault is defined by a zone of discrete anastomosing splays. The on-strike and down-dip extent of the Otto fault defines the South Pacific Zone (SPZ) fault system.

CX West Fault

The CX West fault is a younger offsetting fault with a strike of 245° and dip of 70°. This normal fault has a displacement of approximately 150 ft and offsets stratigraphy as well as the Adam Peak and Otto faults in a down-to-the-northwest direction.

 

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Ogee Fault

The Ogee fault is a 10-100 ft wide fault zone with a strike of 070° and dip of 85° that juxtaposes the Upper and Lower Comus formations. The fault is defined by an anastomosing system of splays along strike and down dip. The fault is interpreted as a long-lived accommodation zone that has experienced multiple phases of reactivation. Recent underground mapping data indicates recent right-lateral oblique-normal motion of unknown displacement.

Linehole Fault

The Linehole fault is a through-going southwest trending normal fault that dips 85° to the northwest. The fault has displacement of approximately 100 ft and appears to act as a structural and mineralization boundary on the property.

LH Fault

The LH fault is a splay off the Linehole fault with a strike of 030° and dip of 80°. The fault has normal sense offset with approximately 50 ft of displacement. The intersection of the LH and Ogee faults is one of the most important and prolific structural intersections on the property.

CX Fault

The CX Fault is a complex zone of brittle fracturing that juxtaposes Upper Comus argillite against limestone beds of the Lower Comus. The fault strikes approximately 035° to 045° and dips 55° to 65° southeast Chadwick (2002), as shown in Figure 6-5. The fault originated as a west-verging thrust and has since been reactivated as a normal fault with an unknown amount of down-to-the-southeast displacement.

SPZ Fault System

The SPZ fault system is comprised of the along-strike and down-dip extent of the Otto fault and its associated splays (SPZ and Otto suite of faults). The zone trends northeast with a dip of approximately 50° to the southeast. The Upper and Lower Comus formations are often juxtaposed along the Otto suite of faults in this zone. However, the defining characteristic of this fault zone is a transition from weakly metamorphosed rock in the hanging-wall to more strongly metamorphosed rock in the footwall.

 

 

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Mag Fault

The Mag fault is a younger through-going normal fault with a strike of 340° and dip of 75°. The fault appears to be related to Basin and Range extension. Displacement is unknown but interpreted to be significant in a down-to-the-east direction.

Mag Fault System

The Mag Fault system is a northwest trending suite of brittle faults that define the Mag pit. The two main faults, the east-northeast dipping Mag fault and the west-southwest dipping Mag West fault, form a horst block within which mineralization is concentrated.

6.4 Mineralization

Mineralization at Granite Creek is structurally controlled. Faults are the primary control of mineralization, especially in high-grade underground zones. Lithologic contacts, bedding, and folds also play an important role, especially in near surface (open pit) mineralization. High-grade mineralized zones are moderately continuous along faults with the most prolific zones occurring at structural intersections. Gold mineralization is found within pyrite that consists of two stages of development, an early non-ore- pyrite stage and a gold-bearing arsenian pyrite stage (Ridgley, Edmondo, MacKerrow, & Stanley, 2005). Megascopically, the gold-bearing pyrite is typically dull brassy to black in color and very fine-grained. Pyrite may also be associated with remobilized carbon, imparting a “sooty” appearance to the pyrite. Gold is primarily contained in pyrite as microscopic inclusions or found in solid solution within arsenian-pyrite rims around fine pyrite grains (Wallace & Wittkopp, 1983; Foster, Gold in Arsenian Framboidal Pyrite in Deep CX Core Hole DDH-1541: Unpublished Pinson Mining Company Report, 1994; Ridgley, Edmondo, MacKerrow, & Stanley, 2005). Gold mineralization shows a correlation with arsenic, antimony, mercury, and thallium.

Gold mineralization at the Property is primarily hosted by the Upper and Lower Comus Formations, which consist of argillite and interbedded argillite and limestone, respectively. The Upper Comus is the primary host lithology in the Mag Zone and currently is host to the majority of surface resources at the Pinson (Granite Creek) deposit (Gustavson, 2012). The Upper Comus is also locally mineralized within the B, C, CX, CX-West, and portions of the RFZ. The Lower Comus hosts the majority of the high-grade underground resources. In areas proximal to the Osgood Mountains stock including the underground resources, most of the host rock has been metamorphosed. In these areas argillite has been metamorphosed to hornfels with limestone altered to garnet, pyroxene, wollastonite, and marble. Higher gold grades are typically located in these metamorphosed rocks along fault zones due to the lack of wall rock permeability.

Rocks of the Preble Formation are a poor host for gold mineralization but do contain localized gold concentrations where they have been brecciated and are adjacent to major fluid conduits.

 

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Figure 6-5 is a representative cross section of the property illustrating the geometry of mineralization controlling structural features, such as faults and the lithologic contact between the Upper and Lower Comus Formation.

Figure 6-5 Cross-section A-A’ looking Northeast showing Structure, Lithology and Mineralization (Section Line is shown on Figure 6-4)

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Oxide mineralization includes pervasive limonite and hematite, along with other iron and arsenic oxides. Oxidation is extensive in the Ogee Zone and CX Fault system, occurring along the entire length of the zones and penetrating to a depth of 1,500 feet (457 meters). Within the RFF system, oxidation is more variable. In some fault and shear zones, oxidation may be present to depths of 1,800 feet (549 meters), whereas in others it may only reach to depths of < 500 feet (152 meters) (Ridgley, Edmondo, MacKerrow, & Stanley, 2005).

 

6.4.1

Mag Pit Mineralization

Gold mineralization within the Mag Pit is hosted by argillite of the Upper Comus Formation. The mineralized zone has a north-northwest orientation, sub-parallel to the Mag Fault, dips to the east-

 

 

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northeast and plunges to the south-southeast (McLachlan, Struhsacker, & Thompson, 2000). The mineralized body is tabular, has a strike length of approximately 4,000 feet (1,219 meters), varies from 200 to 400 feet (61 to 122 meters) in width, and has an average down dip extent of 450 feet (137 meters) (Kretschmer, 1985; Foster & Kretschmer, Geology of the Mag Deposit, Pinson Mine, Humboldt County, Nevada, 1991). Bedding within the Upper Comus Formation is the primary control of mineralization. High-grade zones in the southern portion of the deposit are localized along northwest trending faults within the Mag horst block. Mineralization within the Mag deposit is more disseminated and lower grade than the Rangefront, CX, and Ogee zones (Gustavson, 2012). Gold mineralization is spatially associated with decarbonatization, kaolinization, white kaolinite fracture filling, silicification, and quartz veinlets (McLachlan, Struhsacker, & Thompson, 2000).

 

6.4.2

Underground Mineralized Zones

Multiple areas of high-grade gold mineralization at the Granite Creek deposit are amenable to underground mining methods, as shown by previous operators. These include the Rangefront, Otto-Adam Peak, Ogee CX, and South Pacific Zones. All of these zones show strong structural control.

 

6.4.3

Rangefront Zone

The RFZ consists of pervasive argillization and decarbonatization with intense brecciation along the lower bounding RFF. Structural/mineralization trends are difficult to discern in this zone with mineralization occurring as discontinuous amorphous bodies within the Comus Formation. High-grade zones are concentrated in the Lower Comus with anomalous mineralization present in the Preble Formation, proximal to the RFF. Silicification is minor, with calcite veins occurring along the margins of fault zones. Structural and dissolution breccias that occur along bedding and structural intersections within the Lower Comus Formation are particularly receptive to mineralization. The zone has a strike length of approximately 950 feet (290 meters), a down dip extent of 1,100 feet (335 meters), and an average width of 100 feet (30 meters).

 

6.4.3.1

Otto-Adam Peak Zone

The Otto-Adam Peak zone is defined by the Otto and Adam Peak faults and their associated splays. The zone trends northeast, dips southeast, plunges to east-northeast and is offset down-dip by the CX West fault. The zone is pervasively argillized with intense brecciation occurring along faults. Mineralization is moderately continuous, controlled by a network of discrete anastomosing faults and splays within the Lower Comus Formation. High grade mineralization occurs along fault intersections throughout the zone. The mineralization has a strike length of approximately 500 feet (152 meters), a vertical extent of 700 feet (213 meters), and an average width of 75 feet (23 meters).

 

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6.4.3.2

Ogee Zone

The Ogee zone is an east-northeast trending near vertical mineralized zone controlled by the Ogee Fault and associated splays. The zone is argillized, decarbonatized, and intensely brecciated along faults. The upper portion, defined by the intersection of the Ogee Fault and the contact between the Upper and Lower Comus Formation, plunges to the east-northeast at 55o. The lower portion is near vertical and controlled by faults and structural intersections. The upper portion is strongly oxidized while the lower portion is mostly oxidized but contains more comingled sulfide. The mineralization has a strike length of 400 feet (122 meters), a vertical extent of 1,500 feet (457 meters), and an average width of 75 feet (23 meters).

 

6.4.4

CX Zone

The CX zone consists of both near-surface (open pit) and higher-grade underground mineralization. The lower-grade open pit mineralization is controlled by the through-going CX fault and its associated hanging wall and footwall splays. Mineralization is discontinuous and associated with pervasive argillization and decarbonatization within structural and dissolution breccias in the Lower Comus Formation. The near-surface portion of the zone has a strike length of 3,500 feet (1,066 meters), a down dip extent of 400 feet (122 meters), and an average width of 75 feet (23 meters). The higher-grade underground portion of the mineralization is more tightly structurally controlled along the down-dip section of the CX Fault with an average width of 40 feet (12 meters). The underground portion of the zone has a strike length of 1,000 feet (305 meters) and a down dip extent of 1,200 feet (366 meters).

 

6.4.5

South Pacific Zone

The South Pacific Zone (SPZ) is a northeast trending and southeast dipping zone of high-grade fault-bound mineralization with a northeast plunge of 45o. The mineralization is controlled by the along-strike and down-dip extent of the Otto fault. The zone is defined by a suite of northeast striking moderately southeast dipping anastomosing fault splays with the highest grades concentrated along faults that juxtapose the Upper and Lower Comus Formations. The mineralization has a strike length of 1,250 feet (381 meters), a down dip extent of 900 feet (274 meters), and average fault-bound mineralization widths of 25 feet (7.6 meters).

6.5 Alteration

Alteration assemblages observed at the Granite Creek include silicification, decarbonatization, pyrite, and remobilization of carbon. Alteration mapping by Chadwick outlined the distribution of these assemblages within the pits.

 

 

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In the CX Zone, which follows the strike of the CX Fault and includes the A, B, C, CX, and CX-West pits, McLachlan et al. (The Gold Deposits of Pinson Mining Company: A Review of the Geology and Mining History through 1999, Humboldt County, Nevada, 2000) documented gradational changes in the style and intensity of observed alteration. In the southwest, within the B Pit, gold mineralization occurs in strongly fractured shale and silty carbonate that has been weakly silicified and clay altered. In the nearby A Pit, alteration consists of intense silicification of carbonate lithologies and formation of gold-rich jasperoid along structures. Gold grains within the jasperoid are typically <5 microns in size and are found as inclusions in arsenian pyrite (McLachlan, Struhsacker, & Thompson, 2000). Within the C Pit, located northeast of the A Pit, high-grade material is hosted in decarbonatized carbonates that have been crosscut by small faults.

Within the CX Pit, mineralization consisted of silica and pyrite replacing carbonate along narrow structures, resulting in the formation of intermittent jasperoid and locally silicified wallrock. A large volume of the adjacent hanging wall carbonate-bearing siltstone is decarbonatized, but barren. Within the CX-West Pit, mineralization is hosted in strongly calc-silicate carbonates, which exhibit strong argillic alteration.

Mineralization in the Mag Pit is associated with decarbonatization, kaolinization, white kaolinite fracture filling, silicification and quartz veining (McLachlan, Struhsacker, & Thompson, 2000). Except for some massive limestone units, the original carbonate content of the calcareous host lithologies was removed during decarbonatization, resulting in a porous silty textured rock. Silicification occurs as replacement of the decalcified lithologies and healing fault gouge and breccia. Quartz veining and drusy open space coatings are common throughout the deposit. White kaolinite is commonly formed along fractures within the central portion of the deposit and elsewhere occurs as an argillic replacement of the host lithologies (McLachlan, Struhsacker, & Thompson, 2000). Lithology and alteration relationships can be observed in Chadwick’s 2002 pit maps. Chadwick’s alteration map of the Mag Pit is shown in Figure 6-6.

 

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Figure 6-6 Alteration of the Mag Pit

 

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Source: Chadwick 2002

 

 

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The RFF Zone displays pervasive argillization and decarbonatization of host lithologies along with the formation of dissolution collapse breccias and intense shearing. Where the alteration is strongest, the altered zones consist of punky, spongy decarbonatized limestone in an argillically altered fine grained, carbon-rich matrix (Gustavson, 2012). Silicification is minor and occurs as a broad overprint on the zone. Calcite veining is also prevalent along the margins of the RFF.

6.6 Deposit Types

The structural setting, alteration mineralogy, and mineralization characteristics of Granite Creek are consistent with Carlin-type deposits as defined in Radtke (Geology of the Carlin Gold Deposit, Nevada, 1985: USGS Professional Paper 1267, 1985) and Hofstra and Cline (Characteristics and Models for Carlin-Type Gold Deposits, 2000).

Carlin-type deposits formed in the mid-Tertiary after the onset of extension in an east-west-trending, subduction-related magmatic belt. The deposits are located along long-lived, deep crustal structures inherited from Late Proterozoic rifting and the formation of a passive margin within Paleozoic carbonate sequences composed of silty limestone to calcareous siltstone. The carbonate sequences are overlain by either structurally controlled siliciclastic sequences controlled by the Early Mississippian-aged Roberts Mountain allochthon or by stratigraphically controlled siliciclastic sequences. The siliciclastic rocks are less permeable than the underlying carbonate rocks, which traps fluids along major structures, causing them to flow laterally into the permeable and reactive carbonate sequences.

Alteration of host carbonate sequences consists of decarbonatization, argillization, and selective silicification, forming jasperoid and causing carbon flooding. Gangue minerals in Carlin-type deposits consist of calcite, siderite, and ferroan dolomites that can occur as geochemical fronts beyond the mineralized zones.

Gold deposition occurs in arsenian pyrite, is hosted within carbonaceous sequences near major high angle structural zones, and is concentrated in structural traps and/or replacement horizons of reactive and permeable sedimentary beds.

The Carlin-type deposits typically show enrichment in antimony, arsenic, mercury, and thallium, caused by hydrothermal fluids with temperatures ranging from 180-230°C. The source of fluids is likely deep-seated magmas that released gold bearing fluids at depths of 10 to 12 km. These magmas formed during Eocene slab-rollback of the Farallon plate as upwelling asthenosphere impinged on a strongly metasomatized sub-continental lithospheric mantle (Muntean, Cline, Simon, & Longo, 2011). Tertiary dikes associated with mineralization and radiometric age dates between 39 to 42 Ma along with isotopic data provide evidence toward the above hypothesis.

 

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Structural pathways, reactive rocks, and sources of heat, gold, sulfur, and iron are required for Carlin-type deposits to form. Large regional structures transecting reactive rocks create contacts, faults, and shears. These secondary structures create pathways and traps for hydrothermal and metalliferous fluids.

 

 

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7

Exploration

No exploration has been conducted at Granite Creek by i-80. Granite Creek is a production stage project and as such the geologic focus is on drilling to convert resources and extend known mineralization trends.

7.1 Exploration

No exploration work has been conducted by i-80. This section discusses exploration undertaken by previous owners.

Exploration techniques employed on the Property to define additional gold resources have consisted primarily of mapping, geochemical sampling, and drilling. Use of these methods has resulted in the discovery of approximately one million ounces of gold in several open pit deposits. Several geophysical techniques have also been used to aid in the delineation of gold resources, albeit with limited success. The geophysical programs have mostly been applied to exploration programs along strike of known mineralization and as grass-roots applications to locate additional mineralized zones.

Atna became involved in Project planning in July 2004 and began drilling the Property in August 2004 after execution of the earn-in agreement with PMC on 12 August 2004. Atna continued work through April 2006. Atna vested a 70% interest by completing $12M in exploration and development expenditures and completing an NI 43-101 Technical Report of the Project’s resources (Atna Resources Ltd., 2007).

 

7.1.1

Geologic Mapping and Geochemical Sampling

Cordex, and its successor, PMC, explored the Property through geologic mapping and geochemical sampling. There are three known mapping programs:

 

   

A regional mapping program from the Preble to the Getchell mines conducted in the late 1970s

 

   

A 1:6000-scale mapping program of the Property in 1983

 

   

A 1:2400-scale mapping program of the Pinson pit area through the active life of the mine

Bench mapping in the pits occurred during mining and was followed up by detailed 1:1200-scale mapping of the A, B, C, CX, MAG, CXW, and Blue Bell pits by Tom Chadwick starting in 2000, after mining ceased. These maps were completed under the Homestake/Barrick partnership agreement.

 

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Several geochemical programs were also completed by Cordex and PMC during the active mine life of the Granite Creek Mine, and by Homestake. These included programs:

 

   

Cordex took rock chip samples in conjunction with mapping programs. A total of 737 rock chip samples were collected. Samples were assayed for gold, silver, arsenic, antimony, and mercury. Select samples were also analyzed for lead, zinc, copper, and manganese. The combined mapping/sampling programs were responsible for the discoveries of the Blue Bell and Felix Canyon deposits (Sim, 2005).

 

   

PMC completed six float chip geochemical grids consisting of 8,756 samples. These grids covered the MAG deposit and along strike south of the A and B Pits.

 

   

A biogeochemical sagebrush sampling program was conducted in the 1990s with inconclusive results.

 

   

Under the Homestake/Barrick JV, an additional 312 rock samples and 273 soil samples were collected. These programs were completed on strike south of the existing pit areas and west of the A, B, C, and CX Pits.

 

7.1.2

Osgood Mining Geologic/Structural Mapping

In 2016, OMC contracted Mr. Robert Leonardson to complete a geological study on the Property that focused on advancing OMC’s understanding of the structural framework and on providing guidance on exploration targeting. This work included structural and geologic mapping of the open pits and underground exposures, construction of Property-wide cross-sections, and report writing that included the identification of exploration targets on the Project.

Mr. Leonardson concluded that potential targets to discover additional gold mineralization are at intersections of the east-dipping, north–south faults (Rangefront/Mag) with the southeast-dipping CX-type faults. Other areas include the intersection of the sub-vertical northwest-striking faults with the CX-type faults. Examples of the first type are the CX hanging wall splays where they intersect the Mag Fault in the north half of the Mag Pit. The second example is exemplified by the intersection of the Bluebird Fault Zone with the Delaney thrust in the Blue Bell East pit and the intersection of the Bluebell 2 Fault with the CX thrust in the CX B Pit. Zones of limestone decarbonization such as seen in the CX Pit are also potential hosts for gold mineralization. These zones indicate strong fluid/vapor flow through the rock mass. Specific areas for exploration include:

 

   

The intersection of the SOS and JP dikes on the south wall of the CX Pit. This area contains the largest block of decarbonization on the Property, and the hydrothermal alteration may represent an “exhaust plume” emanating from depth.

 

   

The Ogee pipe extension is located between 1,500 feet and 1,800 feet below the CX-C Pit. A historical hole, HPC-070A intersected a 760-foot interval of low to moderate gold grades

 

 

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above 3,160 feet and high-grade mineralization from 3,160 feet to 3,130 feet near, and just south of, the proposed Ogee high-grade down-dip extension.

 

   

The northern continuation of the fault-propagated anticline in the western portion of the Mag Pit between the Mag Fault and CX Fault and to the north of the Mag Pit. The anticline steepens to the south, and the best chance to intersect high-grade mineralization would be at the intersection with the Disturbed Fault.

 

   

The intersection of the Adam Peak Fault and the Mag Fault suite north of the Mag Pit.

 

   

The CX-B Pit decarbonatization zone at the intersection of the CX and Bluebell 2 faults on the west limb of the Pinson anticline.

 

   

The Mag Pit decarbonatization on the west wall along a section of the Mag Fault intersection with the CX and HW faults and the Disturbed Fault.

 

   

The Mag Pit decarbonatization on the west wall along a portion of the Mag Fault intersections with the Disturbed Fault.

 

   

Bluebell east pit decarbonatization at the intersection of the Bluebell and Delaney faults.

 

   

Traps and fault intersections along the north-northwest-trending Mag Fault suite and the northeast-trending CX type faults.

 

   

Flat to ramp traps down dip extension of fluids that mineralized the Bluebell, CX(?) between South Mountain Fault, and the southern Mag suite of faults. Flat to ramp traps along the Adam Peak detachment and subsequent faults (CX, Disturbed, and South Mountain).

 

7.1.3

Geophysical Surveys

Numerous geophysical surveys have been conducted on the Property. These include both regional and detailed surveys. The regional surveys included gravity and aeromagnetics. Detailed surveys involved mostly electromagnetic techniques and included Induced Polarization (IP), Electromagnetics (EM), Magnetotellurics (MT), and Controlled Source Audio-frequency Magneto Tellurics (CSAMT) surveys. A summary of these techniques includes:

 

   

Airborne EM and magnetics by the U.S. Geological Survey (USGS) at quarter-mile line spacing throughout much of the Getchell Trend

 

   

Ground-based magnetics over the CX Zone completed in 1970 by Cordex

 

   

Regional gravity surveys, both public and private, compiled by Homestake in 1997

 

   

Ground-based magnetic survey at the north edge of the Mag Pit completed in 1998 by Homestake

 

   

Several generations of AMT (EM, IP, CSAMT) completed by PMC

 

   

Several CSAMT lines completed by Homestake between 1998 and 2000

 

   

Several EM lines completed by Homestake in 2000

 

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A detailed gravity survey over the Property conducted by Magee Geophysical Services, LLC of Reno, Nevada in October 2006 (Magee Geophysical Services, 2006), during which a total of 2,587 gravity readings were acquired using a 100-meter (328-foot) station spacing covering approximately 27 square km (10 square miles) (Figure 7-1). The results were interpreted by Fritz Geophysics in 2007 (Fritz Geophysics, 2007). The existence of about 1,700 drill holes within the gravity survey area allowed a novel approach to be attempted for the detailed gravity data. Typically, gravity surveys are conducted to attempt to determine the thickness of alluvial cover over bedrock, as well as structures, etc., for possible targets of interest. The final gravity response on the bedrock surface is shown in Figure 7-2, overlain on the topography, with interpreted structures, bedrock rock types and drill hole collar locations. Also included are plots of the original surface gravity field measurements and the thickness of alluvium from drill data. The basement rock types defined by the basement gravity response correlate with the general mapped geology. To the north-west the high density is related to the large intrusive, TKg. This intrusive is magnetic as well. To the south-west is a lower density unit that correlates with a mapped Oc, probably Valmy. Through the center of the survey area there is an even lower density unit that trends reasonably north south and is defined by northerly, northwesterly, and northeasterly structures. Further to the east, there is another higher density unit, possibly Valmy again. This unit is at an alluvial thickness of greater than 1,500 feet (457 meters) and is not as well defined as the other units. As the thickness of alluvium increases, the resolution of the surface gravity data decreases. Finally, at the southeast edge of the survey, there is a large basin fault that appears to drop the bedrock to depths greater than 3,000 feet (914 meters).

 

 

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Figure 7-1: Gravity Survey, 2,587 Stations, Magee Geophysical Services, 2006

 

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Figure 7-2: Pinson Local Gravity Interpretation

 

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In 2008, Barrick interpreted the geophysical survey data at Pinson (Barrick, 2008). For that work, the 2002 MT survey and 2006 gravity survey and all available geological/geochemical information were combined, and a couple of target areas defined requiring a drillhole test. In 2002, Quantec Geoscience were contracted to acquire TITAN

 

 

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24 MT data over the Pinson property. Six east-west lines were collected along the Rangefront, spaced on average around 2,000 feet (610 meters) apart. The dipole spacing along line was 300 feet (91 meters). Quantec ran regular 2 dimensional (2D) inversions on the MT data to create resistivity depth sections which have been deemed sufficient for this targeting exercise.

The location of the MT survey lines has been plotted on the geology and pit locations and on the residual gravity with Bourne’s structural interpretation and targets annotated (Figure 7-3). Figure 7-4 to Figure 7-9 show MT resistivity depth inversions for each of the survey lines, with all drilling, surface geological mapping Chadwick’s interpreted sectional geology, and Bourne’s structural targets annotated (Barrick, 2008).

 

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Figure 7-3: Location of the MT Survey Lines on the Geology and Pit locations (Left) and on the Residual Gravity (Right)

 

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Figure 7-4: MT Resistivity Depth Inversion for Line 6090

 

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Figure 7-5: MT Resistivity Depth Inversion for Line 12300

 

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Figure 7-6: MT Resistivity Depth Inversion for Line 13860

 

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Figure 7-7: MT Resistivity Depth Inversion for Line 15300

 

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Figure 7-8: MT Resistivity Depth Inversion for Line 17160

 

 

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Figure 7-9: MT Resistivity Depth Inversion for Line 19230

 

 

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7.1.4

Underground Drifting/Evaluation

A small exploration drifting program was conducted on the upper “B” zone by Cordex in the 1970s to conduct bulk testing. Results from this program are unavailable.

In May of 2005, Small Mine Development (SMD) of Boise, Idaho, was contracted by Atna to drive exploration drifts, crosscuts, and develop drill stations to complete Atna’s evaluation of the Range Front resource area. Both the Range Front and CX resource areas were of interest in Atna’s program.

The underground development work completed 1,988 feet (606 meters) of 14-foot (4.3-meter) by 16-foot (4.9 meter) adit, 378 feet (115 meters) of decline, and six diamond drill stations (Gustavson, 2012). A small mineability test was also carried out on the newly defined Ogee Zone to evaluate the potential conditions for future stoping. Approximately 400 short tons (363 tonnes) of material were extracted during this test. The results indicated the possibility of drift and fill as a potential mining method.

During 2008, approximately 693 feet (211 meters) of development drifting was completed, and significant geological data was recorded in the RFZ. However, no data on ground conditions was acquired. This data was not collected because it was anticipated that ground conditions would be similar to those encountered at the Getchell Mine, and mineralization would be exploitable by underhand drift and fill stoping methods (Gustavson, 2012).

 

7.1.5

Trenching and Sampling

Atna channel sampled 14 ribs in the Ogee Zone and sent 74 rib and face samples out for assay (Edmondo, McDonald, & Stanley, 2007). Salient results are summarized in Table 7-1. Assays from the samples indicated that no high-grade mineralization was encountered except where the main drift intersected the Ogee Zone on the 4770 elevation.

Table 7-1: Salient Results of the Ogee Zone Channel Sample Assays

 

         
Sample No.   From feet (meters)   To feet (meters)   Length feet (meters)   Gold Grade opt (g/t)
 
North Rib
         

RFUG-055

  76 (23.1)   81 (24.7)   5 (1.5)   0.144 (4.94)
         

RFUG-056

  81 (24.7)   85 (25.9)   4 (1.2)   0.445 (15.26)
         

RFUG-059

  85 (25.9)   88 (26.8 (   3 (0.9)   0.274 (9.39)
         

RFUG-061

  88 (26.8)   93 (28.3)   5 (1.5)   1.448 (49.65)
         

RFUG-063

  93 (28.3)   97 (29.6)   4 (1.2)   0.176 (6.03)
         

RFUG-064

  97 (29.6)   101 (30.8)   4 (1.2)   0.739 (25.34

 

 

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Sample No.   From feet (meters)   To feet (meters)   Length feet (meters)   Gold Grade opt (g/t)
 
North Rib
         

RFUG-067

  101 (30.8)   110 (33.5)   9 (2.7)   0.996 (34.15)
         

Weighted Average

          34 (10.4)   0.682 (23.38)
 
South Rib
         

RFUG-081

  77 (23.5)   80 (24.4   3 (0.9)   0.106 (3.63)
         

RFUG-082

  80 (24.4)   83 (25.3)   3 (0.9)   0.065 (2.23)
         

RFUG-083

  83 (25.3)   93 (28.3)   10 (3)   1.082 (37.10)
         

RFUG-084

  93 (28.3)   96 (29.3)   3 (0.9)   0.894 (30.65)
         

RFUG-086

  96 (29.3)   99 (30.2)   3 (0.9)   0.355 (12.17)
         

RFUG-087

  99 (30.2)   107 (32.6)   8 (2.4)   0.028 (0.96)
         

RFUG-088

  107 (32.6)   112 (34.1)   5 (1.5)   0.228 (7.82)
         

Weighted Average

          35 (10.7)    

 

7.2 Drilling

 

7.2.1

Drilling Campaigns Overview

Numerous holes have been drilled in and around the Property prior to 1970. Unfortunately, this drillhole data is no longer available. Since 1970, a total of 2,083 drillholes totaling 955,747.9 feet (291,312 meters) have been drilled within the Property area. Figure 7-10 shows the drilling by each operator and significant time period. PMC and its predecessors, Rayrock Mines and the Cordex Syndicate, account for most of these holes: 1,434 holes totaling 554,435 feet (168,991.8 meters). Homestake drilled 165 holes totaling 160,207.7 feet (48,831.3 meters), and Barrick drilled 106 holes totaling 101,345.1 feet (30,890 meters). Both companies acted as operators for PMC. Atna, the last company to operate at the Granite Creek Mine, drilled 318 holes totaling 119,074.1 feet (36,293.8 meters).

Table 7-2 presents a summary of the drilling at the Property.

 

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Figure 7-10: Granite Creek Project Drill Plan by Operator

 

 

LOGO

Source: AMC Mining Consultants (Canada) Ltd, 2019

Table 7-2: Summary of Drilling on the Granite Creek Property Since 1970

 

             
Company   Surface RC   Surface Core   UG RC   UG Core   Total
Holes
  Total
Footage
  #
Holes
  Footage
(feet)
  #
Holes
  Footage
(feet)
  #
Holes
  Footage
(feet)
  #
Holes
  Footage
(feet)
                     

PMC

  1,426   546,313.0   8   8,122.0                   1,434   554,435.0
                     

PMC (Homestake)

  136   108,335.0   29   51,872.7                   165   160,207.7
                     

PMC (Barrick)

  39   35,645.0   67   65,700.1   4   930.0   56   19,756.0   106   101,345.1
                     

Atna

  29   18,672.0   65   52,847.6   176   32,068.0   48   15,486.5   318   119,074.1
                     

Total

   1,630     708,965.0     169     178,542.4     180     32,998.0     104     35,242.5     2,083     955,747.9 

Note: RC=reverse circulation, UG=underground.

Source: Osgood Mining Company LLC.

Each period of drilling is described in further detail in Sections 7.2.1.1 to 7.2.1.9.

 

 

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7.2.1.1

PMC Drilling 1970 to 1996

Many holes drilled by PMC during this time period were development holes drilled in and adjacent to existing pits. Over 1,400 holes were drilled within the A, B, C, CX, Mag, CX-West, Felix, and Blue Bell pit areas. Many of these holes were drilled vertically, and all but eight were either conventional rotary or RC. The eight core holes that were drilled (8,122 feet [2,475.6 meters]) were in the B, C, CX, and Mag Pit areas to test stratigraphy, metallurgy, or deep mineralized structures (Golder Associates, 2014). Table 7-3 summarizes the drilling PMC conducted through 1996.

Table 7-3: PMC Drilling through 1996

 

         
Company   Surface RC   Surface Core   Total Holes   Total Footage
  # Holes   Footage (feet)   # Holes   Footage (feet)
             

PMC

  1,426   546,313.0   8   8,122.0   1,434   554,435.0

Source: Osgood Mining Company LLC

 

  7.2.1.2

PMC – Homestake Drilling 1997 to 2000

Between 1997 and 2000, Homestake, as the operator for PMC, drilled 165 holes, as shown in Table 7-4. Of the 165 holes drilled, 136 (108,335 feet [33,020.5 meters]) were directed into the CX and RFF system.

Table 7-4: Homestake Drilling

 

         
Company   Surface RC   Surface Core   Total Holes   Total Footage
  # Holes   Footage (feet)   # Holes   Footage (feet)
             

PMC (Homestake)

  136   108,335.0   29   51,872.7   165   160,207.7

Source: Osgood Mining Company LLC

 

  7.2.1.3

PMC – Barrick Drilling 2003

Four exploration holes were drilled by Barrick, operator at the time for PMC, to test extensions of the CX Fault Zone near its projected intersection with the Mag Pit fault system. The drilling did not identify significant mineralized zones, and no additional work was conducted by Barrick (Golder Associates, 2014). Table 7-5 shows a summary of the Barrick drilling.

Table 7-5: Barrick Drilling 2003

 

         
Company   Surface RC   Surface Core   Total Holes   Total Footage
  # Holes   Footage (feet)   # Holes   Footage (feet)
             

PMC (Barrick)

  3   3,340.0   1   3,003.3   4   6,343.3

Source: Golder Associates 2014.

 

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  7.2.1.4

Atna Drilling 2004

The drilling by Atna in 2004 followed up on mineralized zones previously identified by PMC and Homestake. Thirty-one holes totaling 29,739.5 feet (9,064.6 meters) were drilled. These holes were comprised of four RC holes totaling 2,217 feet (675.7 meters) and 27 core holes totaling 27,522.5 feet (8,388.9 meters) (Table 7-6). This drilling program had five objectives:

 

   

Improve the grade and thickness of mineralized zones, especially in areas where drilling consisted of only RC drilling.

 

   

Infill drilling, especially where previous drill spacing was greater than 400 feet (121.9 meters).

 

   

Expand mineralized zones both laterally and down-dip.

 

   

Obtain rock quality data on hanging wall, footwall, and mineralized zones.

 

   

Evaluate previously identified targets.

Table 7-6 shows a summary of the Atna drilling.

Table 7-6: Atna Drilling 2004

 

         
Company   Surface RC   Surface Core   Total Holes   Total Footage
  # Holes   Footage (feet)   # Holes   Footage (feet)
             

Atna

  4   2,217.0   27   27,522.5   31   29,739.5

Source: Osgood Mining Company LLC.

Of the 31 holes drilled, 13 holes (13,000 feet [3,962.4 meters]) were drilled into the CX Fault Zone and 18 holes (16,739.5 feet [5,102.2 meters]) were drilled into the RFF Zone (Golder Associates, 2014).

 

  7.2.1.5

Atna Drilling 2005 – 2006

The objective of the 2005 to 2006 drilling program was to define and delineate Measured and Indicated gold Mineral Resources in the upper portions of the RFF Zone where Atna had outlined a 1,000-foot (305-meter) long by 200- to 500-foot (61- to 152.4-meter) thick mineralized zone during its 2004 drilling program. The drilling program was designed to test the upper RFZ between the 5,000 and 4,400 feet (1,524 and 1,341 meters) amsl (Golder Associates, 2014). The program used both surface and underground drilling to delineate the zone. A total of 107 drillholes (55,180.1 feet [16,818.9 meters]) were drilled between 2005 and 2006 (Table 7-7).

Table 7-7: Atna Drilling 2005-2006

 

           
 Company    Surface RC   Surface core   UG Core    Total Holes     Total Footage 
   # Holes     Footage (feet)     # Holes     Footage (feet)     # Holes     Footage (feet) 
                 

Atna

  25   16,455.0   34   23,238.6   48   15,486.5   107   55,180.1

Source: Osgood Mining Company LLC

 

 

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Surface drilling began in May of 2005. The majority of these holes were core holes, which were pre-collared via RC drilling and completed with core drilling. Fifty-nine (59) drillholes, totaling 39,693.6 feet (12,098.6 meters) of drilling, were completed from surface.

Underground drilling began in September of 2005 after drifting was completed and underground drill rigs became available. In total, 48 holes aggregating 15,486.5 feet (4,720.3 meters) of underground drilling were completed in the Ogee, CX West, and Range Front targets.

 

  7.2.1.6

PMC (Barrick) Drilling 2007

In August of 2007, surface exploration and development drilling began using an Eklund RC drill rig and a Major Drilling core rig. Targets tested included portions of the CX and RFF, Ogee Zone, and the HPR104 area. The HPR104 area is north of the Granite Creek Mine.

Twenty-three (23) surface holes (18,916.2 feet [5,765.7 meters]) were completed during the latter part of 2007 as shown in Table 7-8. The results of the drilling were disappointing in that only thin, sub-economic zones of underground mining gold grades were intersected.

Table 7-8: PMC - Barrick Drilling 2007

 

         
Company   Surface RC   Surface Core   Total Holes   Total Footage
  # Holes   Footage (feet)   # Holes   Footage (feet)
             

PMC (Barrick)

  7   4,935.0   16   13,981.2   23   18,916.2

Source: Osgood Mining Company LLC.

 

  7.2.1.7

PMC (Barrick) 2008 Drilling

Surface drilling began in January of 2008 with three core drills and one RC drill testing areas north of the CX West pit. The core drilling was focused on completing holes pre-collared by RC drilling in 2007 and testing the deep potential of the Getchell Fault system north of the Granite Creek Mine, which had associated gravity and MT anomalies (Golder Associates, 2014). RC drilling was primarily focused on pre-collaring holes for follow-up core drilling north of the CX/CX West pits. Surface core drilling was completed in April of 2008. RC drilling continued throughout 2008, with the focus on drilling pilot holes for potential dewatering well locations.

Underground exploration began in April 2008 as discussed in Section 7.1.4. SMD was contracted to rehabilitate existing underground workings and drive exploration headings into the Ogee and CX zones. SMD supplied an underground RC drill for closely spaced definition drilling, and Connors Drilling was contracted to conduct underground core drilling. The SMD contract was terminated in May of 2008. Connors Drilling remained on site and brought in a second

 

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underground core rig in mid-July. Both core rigs continued operation through mid-December, testing the Ogee Zone and conducting widely spaced drilling within the RFZ.

In August 2008, a second surface drilling program was initiated to twin RC holes in key areas of the resource suspected of having downhole contamination. Two core rigs and one RC rig (to pre-collar holes) were used. A third surface core rig was also brought in to complete one deep hole to test the Mag fault-Delaney fault intersection south of the resource area. The drilling program was completed in mid-December and all drilling equipment removed from site.

During 2008, total surface drilling included 29 RC holes totaling 27,370 feet (8,342.4 meters) and 50 core holes totaling 48,715.6 feet (14,848.5 meters). Underground drilling included 4 RC holes for 930 feet (283.5 meters) and 56 core holes totaling 19,756 feet (6,021.6 meters) (Table 7-9).

Table 7-9: PMC – Barrick Drilling 2008

 

             
Company   Surface RC   Surface Core   UG RC   UG Core   Total Holes   Total
Footage
  # Holes  

Footage

(feet)

  # Holes   Footage
(feet)
  # Holes   Footage
(ft)
  # Holes   Footage
(feet)
                     

PMC (Barrick) 

  29   27,370.0   50   48,715.6   4   930.0   56   19,756.0   139   96,771.6

Source: Osgood Mining Company LLC

HPR104

During the 2008 drilling program, eight holes were drilled north of the Pinson deposit resource area. These holes were designed to twin earlier PMC drilling that were drilled to test the intersection of the Range Front and Linehole Faults. The results of the initial drilling could not reproduce the thick low -grade intercept identified in an earlier hole, hole HPR104. This was considered to constitute downhole contamination in hole HPR104, and the hole was removed from the database. A second round of core drilling did intersect thin, higher-grade mineralization. Hole BPIN-008 intercepted 21.5 feet grading 0.620 opt at a depth of 1,378 feet (Golder Associates, 2014). This mineralization appeared to be structurally controlled by the intersection of the Linehole Fault and the Upper/Lower Comus contact 900 feet northeast of the main portal.

Deep Exploration Targets

Two deep drillholes, BPIN-010C and BPIN-011A, were drilled in 2008. Hole BPIN-010C was drilled to a depth of 2,845.5 feet (867.3 meters) and was designed to test the Lower Comus Formation adjacent to structures identified from a 2006 gravity survey (Golder Associates, 2014). The hole bottomed in Upper Preble Formation, and assay results proved negative. Hole BPIN-011A was drilled to a depth of 2,778 feet (846.7 meters) and ended in argillite and shale of the

 

 

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Upper Comus (Golder Associates, 2014). The hole was designed to test the projected intersection of the Mag and Delaney faults. Analyses of chip samples indicated a 60-foot (18.3-meter) zone of low grade- gold (0.029 opt [0.99 g/t]) at 1,440 feet (438.9 meters) hosted in silicified Upper Comus claystone and shale (Golder Associates, 2014). Subsequent analyses of core from the entire hole indicated narrow zones of mineralization associated with decarbonatization and pyritized sediments.

 

  7.2.1.8

2012 Atna Mag Pit Core Drilling

In 2012, Atna completed four PQ-size core holes, totaling 2,086.5 feet (636 meters), to acquire samples for column leach testing from mineralized material within the Mag Pit resource area. The holes were drilled along strike of the known mineralized zone, with each hole intersecting potential high-grade material. In addition to the metallurgical holes an additional 56 underground exploration RC holes totaling 7,495 feet (2,284.5 meters) were drilled in the Ogee Zone. Table 7-10 summarizes the drilling conducted by Atna in 2012.

Table 7-10: Atna Drilling 2012

 

         
Company    Surface Core    UG RC    Total Holes    Total Footage
   # Holes    Footage (feet)    # Holes    Footage (feet)
             

Atna

   4    2,086.5    56    7,495.0    60    9,581.5

Source: Osgood Mining Company LLC

 

  7.2.1.9

2013 – 2015 Atna Underground Development RC Drilling

Between 2012 and 2015, Atna completed 120 underground RC holes totaling 24,573 feet (7,489.9 meters) (Table 7-11). These holes were designed to confirm continuity of mineralization and to delineate stope configuration within the Ogee Zone for mining.

Table 7-11: Atna Drilling 2013 – 2015

 

       
Company    UG RC    Total Holes    Total Footage
   # Holes    Footage (feet)
         

Atna

   120    24,573.0    120    24,573.0

Source: Osgood Mining Company LLC

 

7.2.2

Representative Drill Sections and Plan

Figure 7-10 shows the drill plan of the Property in the area of the 2021 Mineral Resource, shown by a red outline. The drillholes are coded by operator and significant time periods. Figure 7-11 shows a plan view with section lines of the Open Pit area. Figure 7-12 to Figure 7-15 show representative vertical sections through the four Open Pit areas. Figure 7-16 shows a vertical section through the underground resource area.

 

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All drill results presented in Section 7.2 are from previous operators. i-80 had not conducted drilling on the Property at the time of the 2021 resource model.

 

Figure 7-11: Plan View Section Lines of Granite Creek Mine Project

 

 

LOGO

Note: Red outlines show the outline of the open pits.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

 

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Figure 7-12: Vertical Section A-A1 of the Mag Pit Area

 

 

LOGO

Notes: Blue lines are faults. Black line is a topographic surface. Not all items listed in the legend are on all sections.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

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Figure 7-13: Vertical Section B-B1 of the Pit CX and C Area

 

 

LOGO

Notes: Blue lines are faults. Black line is a topographic surface. Not all items listed in the legend are on all sections.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

 

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Figure 7-14: Vertical Section C-C1 of the Pit A Area

 

 

LOGO

Notes: Blue line is a fault. Black line is a topographic surface. Not all items listed in the legend are on all sections.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

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Figure 7-15: Vertical Section D-D1 of the Pit B Area

 

 

LOGO

Notes: Blue lines are faults. Black line is a topographic surface. Not all items listed in the legend are on all sections.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

 

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Figure 7-16: Vertical Section E-E1 of the Underground Resource Area

 

 

LOGO

Notes: Blue lines are faults. Top blue line is a topographic surface. Not all items listed in the legend are on all sections.

Source: AMC Mining Consultants (Canada) Ltd 2019

 

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7.2.3

Drilling, Sampling, and Recovery Factors

There are no drilling, sampling or recovery factors that could materially impact the accuracy and reliability of the results. Drilling has been discussed in this section and sampling and recovery factors are discussed in Section 7.

7.3 Update to Drilling Statistics to Include i-80 Drilling and Land Package Expansion

The Granite Creek land position has expanded since the previous resource estimate was published in 2021. This sub-section contains a summary of all holes drilled within the current land package, and the previous sub-sections describe in greater detail holes drilled within the core land package. No discoveries have been made beyond the core land package, and all holes outside the core land package were drilled by previous operator PMC. Holes without significant mineralized intercepts serve primarily to augment geological knowledge of the property and do not contribute to the resource estimation beyond defining where no resource exists. This section also details drilling completed by i-80, which had recently acquired the property and had not yet commenced drilling when the previous resource estimate was published. Table 7-12 lists number of holes drilled within the current property boundary by type and operator (includes holes drilled from surface and underground), and Figure 7-17 through Figure 7-19 show drilling by previous operators.

Table 7-12 Drillholes Within the Current Property Boundary by Type and Operator

 

Company  

Core Holes
(includes RC pre-

collar with
Core Tail)

  Core
Footage
  RC Holes   RC Footage   Rotary
Holes
  Rotary
Footage
  Total Holes   Total
Footage

 i-80

  225   152,941   1   2,100   0   0   226   155,041

 Atna

  113   68,334   201   49,920   0   0   314   118,254

 PMC-Barrick

  123   91,006   41   33,375   0   0   164   124,381

 PMC

  46   76,297   1,369   631,061   387   124,298   1,802   831,656

 Totals

  507   388,578   1,612   716,456   387   124,298   2,506   1,229,332

 

 

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Figure 7-17 Drilling Completed by PMC

 

 

LOGO

 

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Figure 7-18 Drilling Completed by PMC with Barrick as Operator

 

LOGO

 

 

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Figure 7-19 Drilling Completed by Atna

 

LOGO

 

7.3.1

i-80 Drilling

Drilling at Granite Creek is ongoing. The holes drilled by i-80 presented in this section were drilled from April 2021 through December 2022 and had complete assay results by March 2023, which was the cutoff date for data to be included in the current underground resource estimate. i-80 primarily uses core drilling for sample collection, although one RC water well was drilled and sampled. Most surface holes are pre-collared using RC down to the water table, then completed with HQ size core. Most surface drilling has focused on the Ogee and South Pacific Zones. Underground holes were all drilled as HQ size core and focused on the Otto, Rangefront, Ogee, and South Pacific zones near the existing workings. A Cubex RC rig is used by ore control geologists to assist with short term mining decisions, but the RC holes are not merged into the resource database. Practical Mining recommends managing the underground RC drilling more

 

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attentively to ensure the results are of suitable quality for use in resource estimation. Figure 7-20 shows holes drilled by i-80.

Figure 7-20 Drilling Completed by i-80

 

LOGO

 

7.3.2

Representative Cross Sections

Example sections showing drilling in the underground resource area are shown in Figure 7-22 through Figure 7-24. Holes drilled by i-80 are labeled with hole name and shown with thicker traces. Faults and mineralized envelopes modeled at 0.1 oz Au (3-gram) cutoff grade are shown for reference. Figure 7-21 shows the section locations.

 

 

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Figure 7-21 Plan View Showing Section Locations through the Underground Resource Area

 

 

LOGO

 

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Figure 7-22 Section A-A’ Showing Drilling in the CX Zone, 100 ft thick, looking North

 

LOGO

Figure 7-23 Section B-B’ Showing Drilling in the Otto and Ogee Zones, 25 ft thick, looking North

 

LOGO

 

 

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Figure 7-24 Section C-C Showing Drilling in the South Pacific Zone, 50 ft thick, looking North

 

 

LOGO

 

7.4 Hydrogeology

 

7.4.1

 Sampling Methods and Laboratory Determinations

Hydrogeological data, including water table measurements, pore pressure distribution and direction of groundwater flow, were normally collected in conjunction with exploration and geotechnical investigations in pre-construction studies and later from hydrogeological studies for on-going programs in pit and underground mining areas.

Groundwater dewatering and monitoring wells are the primary method of collecting hydrogeological data in support of mining operations, as well as the collection of pore pressure data which can be converted to groundwater level elevations from a network of vibrating wire piezometers (VWPs). Another source of data is hydrologic testing. Most wells that are drilled undergo hydrologic testing to establish aquifer parameters. These tests range from injection (slug) tests, air-lift tests, short-term and long-term pumping tests, and spinner logging. Data obtained from testing operations are analyzed using industry standard analytical methods. Analytical and numerical groundwater flow models have been developed using hydraulic parameters using testing results, in addition to 3D geological modeling.

 

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From approximately 1980 through 2024 a total of 14 dewatering wells and 42 monitoring wells were completed in the Project area. In 2005, rapid infiltration basins (RIBs) were constructed east of the Project area to infiltrate groundwater pumped from dewatering operations into downgradient, permeable alluvial sediments. During 2022 and 2023, 19 vibrating wire piezometers were installed in the area underground mining operations (HGL, 2022 and 2024). To further assist in underground dewatering operations, one dewatering well was completed in 2023 (HGL, 2023) and another dewatering well was deepened in 2024 (LRE, in preparation) to capture additional groundwater yield. Currently, there are four active dewatering wells, 41 active monitoring wells, and 15 active vibrating wire piezometers across 5 locations (Figure 7-25). Current dewatering pumping rates range from 100 gpm to 750 gpm at the four dewatering wells. All dewatering wells are monitored, controlled and data are logged using a supervisory control data acquisition system (SCADA).

According to permitting requirements, 11 monitoring wells are sampled on a routine basis and analyses run for the State of Nevada Profile I suite at a certified analytical laboratory, currently Western Environmental Testing Laboratory (WETLAB), Reno, NV. Monitor wells and exploration drill holes that have piezometers installed are monitored for water levels and piezometric heads. Surface water is also measured and sampled on a routine basis as required by various permits.

 

7.4.2

 Hydrogeology Investigations

Throughout the span of various mine property owners and operators, the Project area has been the subject of multiple studies aimed at characterizing the hydrogeologic properties of the stratigraphy within the Project area and the surrounding region (Table 7-13). WMC (1998) established an early conceptual hydrogeological model and characterized the physical properties of major water bearing geologic units in the Mag and CX pit areas. Continuing in the early 2000s through 2018, additional hydrogeologic studies were completed by WMC, SWS, and Piteau Associates in support of groundwater monitoring, dewatering operations, water balances, and RIB design (WMC 1998, 2000, 2002, and 2005; SWS 2014; and Piteau Associates 2018).

More recently, i80 contracted HydroGeoLogica (HGL), now part of LRE Water, to conduct operations for monitoring of groundwater levels and pore pressures, plan and oversee operations of dewatering wells, and groundwater flow modeling for local-scale dewatering and regional scale permitting.

 

7.4.3

 Hydrogeologic Description

The Granite Creek Mine is in the Great Basin region of the Basin and Range Physiographic Province. Mountain ranges trending north-south with parallel intermontane basins characterize the

 

 

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Figure 7-25 Well Locations

 

 

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Note: Active dewatering wells identified with red squares. Major structural faults are shown as black lines representing intersections at an elevation of 4,500 ft amsl. Current and planned underground workings are shown in orange

 

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Table 7-13: Timeline for Hydrogeologic Characterization with Relationship to Mining

 

 

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terrain. The entire region is a closed drainage system with all the permanent streams flowing to interior “sinks” such as the Carson and Humboldt sinks, or interior lakes such as Pyramid and Walker. Elevations in the area range from about 4,000 ft amsl in the basins, to over 9,000 ft amsl in the surrounding ranges. The local terrain near the mine area is in the transition from the bedrock, mountain front zone to the alluvial, basin-fill zone.

 

7.4.3.1

   Surface Water

The Granite Creek Mine is in the Kelly Creek Hydrographic Area of the Middle Humboldt Watershed that lies within the Humboldt Basin. The Middle Humboldt has a catchment area of approximately 3,200 square miles draining to the Humboldt River to the southwest.

Granite Creek is an ephemeral stream sourced from seasonal snow melt originating in a bowl below the crest of the Osgood Mountains above Granite Creek Canyon, immediately west of the Project. Stream flow, when present, is currently diverted and routed through a series of pipes and culverts above the south part of the CX Pit and rejoins the original stream channel about 1,100 feet southeast of the CX Pit. The water in Granite Creek typically infiltrates into the permeable alluvial deposits of the middle and lower pediment slopes within 1,500 to 2,500 feet down gradient of the site. Annual stream flow generally occurs from February through July. Generally, no flow is observed in the channel west of the project by the end of June. Based on previous estimates, mean average annual flow is estimated at 0.28 cubic feet per second (cfs) (200 acre-ft/yr or 125 gpm) with springtime flows averaging about 1.9 cfs.

High flow rates in Granite Creek occur during the spring runoff period. When peak seasonal stream flow exceeds the diversion capacity, the overflow has historically been routed via a pipeline to the floor of the mined-out A and B pits. The surface water has infiltrated rapidly through the pit floors and into the CX shear zone, which is hydrologically connected to the CX Pit and the underground development. During spring 1998, a period of record precipitation in northern Nevada, and following peak runoff events, a rise in groundwater levels was noted beneath the floor of the CX Pit necessitating an increase in the dewatering rate from the CX Pit.

Permitting requires annual sampling of surface water at two stations in the Granite Creek channel, when present during the first or second quarters; one station upgradient from the mine site and the other station downgradient of the mine site at the eastern property boundary. Results of hydrochemical analysis of Granite Creek samples indicate that the average chemistry in Granite Creek is similar to that of the CX shear zone bedrock groundwater hydrologic unit (also referred to as the CX hydrologic block) as discussed further in the following section. Collected samples generally report constituent concentrations that meet NDEP RVs.

 

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7.4.3.2

   Groundwater

Groundwater generally moves from recharge areas along the range front of the Osgood Mountains to the east-southeast, towards the central part of the Kelly Creek Valley basin. Two main groundwater systems are recognized in the mine area: 1) alluvial units and 2) and sedimentary and metamorphic bedrock units:

 

 

Alluvial Groundwater System: groundwater moves south – southeast in the alluvial deposits throughout the entire Kelly Creek Valley basin towards the Humboldt River. The alluvial groundwater system in the mine area has been historically monitored by numerous monitoring wells and is well understood.

 

 

Saturated alluvium exists on the east and southeast sides of Mag Pit and its thickness increases significantly to the east of Mag Pit toward the lower elevation alluvial basin. There is no saturated alluvium to the north and west of the Mag Pit or in the immediate vicinity of the CX Pit. Groundwater flow within the alluvium is generally characterized as porous media flow. Hydraulic layering of the alluvium is known to occur, controlled in part by the presence of lower permeability horizontal lenses of fine-grained materials. There is a general increase in the proportion of fine-grained materials closer to Mag Pit. The saturated thickness of the alluvium increases to the east as the underlying bedrock surface dips beneath the valley floor. The pre-mining alluvial water table in the area of Mag Pit was 4,654 ft amsl. The lowest occurrence of alluvium in the east wall of the Mag Pit is about 4,590 ft amsl. The maximum saturated alluvial thickness in the east wall of the Mag Pit was therefore about 65 ft prior to mining Mag Pit.

 

 

Bedrock Groundwater System - groundwater flow within the bedrock units in both the Mag and CX pit areas is predominately controlled by stratigraphy and geologic structure. The bedrock hydraulic properties are therefore highly variable, and flow is dependent on the frequency and alignment of open fracture sets. Major faults, some displaying individual offsets on the order of several hundred feet, function as significant hydrologic features by: providing offsets, juxtaposing geologic units of differing hydrologic characteristics, providing preferred pathways for groundwater flow in fracture zones parallel to the fault plane, and forming gouge filled barriers to horizontal flow perpendicular to the fault plane. These characteristics favor the formation of hydraulically isolated blocks of bedrock bounded by the high angle faults in the Project mine area.

Regionally, groundwater recharge occurs to both the alluvium and the bedrock of the upper piedmont slope elevations and, during years of high run-off, to the alluvium at middle and lower piedmont slope elevations. Groundwater moves towards the center of the basin in the thickening sequences of alluvial deposits. Most natural discharge from the basin occurs through

 

 

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evapotranspiration from the alluvial deposits beneath the valley floor. Locally, historical and current dewatering of the Mag Pit and CX Pit, formation of the Mag Pit lake, as well as the underground mine workings exploration and development to facilitate mining of deeper ore reserves north of CX Pit, has influenced direction of groundwater movement in the vicinity of the mine. Local groundwater movement is also influenced by delivery of water from dewatering operations to a rapid infiltration basin (RIBs) constructed in the alluvial aquifer system for downgradient recharge to the basin.

 

7.4.4

 Mine Dewatering

 

7.4.4.1

      Mag Pit

The Mag Pit was mostly excavated in calcareous mudstones, siltstones, carbonaceous shales with interbedded limestone units exposed in the lower east and southeast walls. The rocks are mostly fine grained and belong to the Upper Comus formation, with locally silicified zones that are present along many of the cross faults. Intrusive rocks and breccias are also common in the pit area and exhibit a high degree of argillization. The main geological structures in the Mag Pit show a strong north northwest-south southeast alignment. During dewatering operations for mining of the pit, groundwater level drawdown was seen to extend along this trend to the northwest and southeast of the pit, but the extent of drawdown along strike was limited by the presence of cross faults.

The Mag Pit area is overlain by basin-fill alluvium, which increases from a thin veneer (less than 30 ft thick) above the west wall to over 200 ft thick along much of the east wall. The alluvial-bedrock contact is down-dropped to the to the east immediately behind the east pit wall. Locally, in the pit area, the alluvium was reported to be fine grained and low permeability close to the east pit wall, which was notably different to the high permeability alluvium penetrated by the mine-water supply wells located in deeper basin sediments east of Mag Pit. Reporting postulated that the alluvium close to the Mag Pit area was locally influenced by debris flow material resulting in lower permeability.

Dewatering the Mag Pit occurred during the period 1987 through April 1998. During the initial period between 1987 and 1991, only sumps were used to dewater the pit. The pumping rate was small and included only minor intermittent pumping from bedrock sumps in the pit floor. In 1991, in pit mining in the lower pit became increasingly difficult because of fracturing and water strikes in the central pit area, consequently dewatering wells #12 and #13 were installed to the north and south of the central pit area, along the main strike of the north northwest-south southeast structures. Several west-northwest and east-southeast cross faults were exposed within the pit. These faults created low-permeability barriers to groundwater flow across the main structural trend. The dewatering rate for the pit was typically within the range 400 to 700 gpm, with much of the water

 

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derived as a result of seepage faces from the alluvium. The dewatering of the Mag Pit ceased in April 1998 and water was allowed to start accumulating in the base of the pit.

In June 1998, the standing water in the base of the pit was 4,550 ft amsl, about 117 ft lower than a 2002 stabilized elevation of the lake of 4,667 ft amsl (WMC, 2002). The Mag Pit lake was rapidly filled during 2000 using a combination of water from the CX pumping system (which was operational at the time) and water from alluvial water supply wells #7 and #8. Rapid filling of the lake was initiated on February 24, 2000, and completed on August 17, 2000 (175 days). A total of 200.4 million gallons was pumped into the lake, which raised the lake elevation from 4,606 ft amsl to approximately 4,671 ft amsl. In December 2024, i80 reported a lake level of 4632.78 ft amsl, about 34- ft below the measurement reported in 2002.

Prior to the start of rapid filling activities, approximately 1,620,000 yd3 of mainly alluvial material was backfilled into the northern end of the pit. The material was placed up to an elevation of approximately 4,625-4,640 ft amsl. The goal of the backfilling was to 1) buttress the unstable northwest and north pit wall, 2) isolate part of the geochemically reactive subunits of Upper Comus in the lower part of the pit, and 3) provide a source of additional alkalinity for the juvenile lake waters. The placed material reduced the area of the lake from approximately 15.8 acres to 12.2 acres.

In summary, the Mag Pit is hydraulically connected with both the local bedrock and regional alluvial groundwater system. During the past 15 years, the Mag Pit lake level has declined consistently with the alluvial groundwater levels, potentially caused by a combination of factors: 1) increased irrigation pumping from the central part of the Kelly Creek groundwater basin to the east and southeast, 2) lower than average precipitation and reduced recharge along the east side of the Osgood Range mountain front, and 3) leakage of groundwater into the highly fractured rock of the CX faulted zones due to dewatering operations for the CX Pit through low-permeable, Mag Pit fault barriers. Based on available data and analyses, the cause of lowering of the Mag Pit lake is likely due to a combination of these factors.

 

7.4.4.2

   CX Pit and Underground Mine Workings

The structural alignment and geometry of the CX Pit differs from the Mag Pit area. The CX Pit lies along the line of the east-northeast-trending, steeply southeast dipping faulted shear zones where the Lower Comus Formation consists of variably metamorphosed, inter-bedded carbonaceous shale and limestone. These sedimentary rocks have been altered by metamorphic processes, producing calc-silicate marble zones within the limestone and argillite. A later period of alteration, associated with mineralization events, has caused localized decalcification, kaolinization and silicification, but there remains a considerable about of unaltered limestone

 

 

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within the middle and upper walls of the pit. After all alteration events, the rocks in the CX pit have been pervasively oxidized by meteoric waters.

Alteration and mineralization associated with the CX pit occurs along and adjacent to the shear zone. The shear zone is an important structural element that allows groundwater flow along strike within the CX bedrock block and contains the bulk of CX pit.

The adit and decline in the CX Pit was collared in 2005. The decline passes through steeply-dipping, tightly folded rocks of the Lower Comus formation and penetrates the faulted axis of an upright anticline along the axis of the Line Hole target area. This structure controls the location of a strongly altered and mineralized zone containing abundant iron oxide. The fault controlling the Line Hole Extension mineralization dips steeply to the northwest and has a reverse (southeast down) sense of movement. This zone is now referred to as the Ogee Zone with respect to the underground workings. Bedding on the south side of the fault typically dips steeply towards the southeast, whereas bedding on the north side of the fault dips steeply towards the northwest. Most underground development and mining in the CX pit has occurred within unmineralized rocks of the Upper and Lower Comus formation.

Early dewatering of the CX pit was also carried out using a system of sumps down to the final floor elevation of approximately 4689 ft amsl. In 2005, underground exploration activities were initiated by Atna Resources in the vicinity of the CX pit under an exploration agreement with Pinson Mining Company (PMC). An exploration incline and decline were driven from within the CX pit, with the portal collared in June 2005 on the northwest portion of the 4,760 ft amsl bench. Atna’s exploration program was suspended in April 2006. In February 2008, PMC resumed underground exploration activities from the 2005 decline.

Dewatering operations to lower the groundwater elevation below the anticipated underground exploration and development commenced in December 2005. Groundwater levels in the bedrock hydrologic block containing the CX pit were lowered to an elevation below the pit floor and the small pit lake in the lower pit was dewatered.

The CX block has subsequently been dewatered from 2007 through present to facilitate underground exploration of deeper ore zones. There are currently four operating dewatering wells APW-1, BPW-3, BPW-5 and GCW-6 (Figure 7-25) pumping from the CX block hydrogeologic unit at a combined average rate of approximately 1,450 gpm, with an additional pumping of approximately 1,000 gpm collected from sumps in the underground mine workings. This combined pumping rate has caused groundwater levels to decline to an elevation of approximately 4,310 ft amsl at the face of current decline mining operation.

 

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7.4.5

 Dewatering Treatment and Discharge

Water from the dewatering wells that is not utilized for operations is currently discharged to Rapid Infiltration Basins (RIBs) on the east side of Getchel Mine Road through HDPE pipelines. Two of the four permitted RIBs (NEV2005102) have been constructed to date, with discharge to one of the two cells at any given time. When RIB maintenance is required, discharge is routed to the dormant cell. Current dewatering efforts are well under the permitted 6,900 GPM threshold of the RIBs and the RIB infiltration is sufficiently limiting surface ponding in the active cell.

Due to arsenic levels that are above Nevada Profile 1 standards, BPW-5 and GCW-6 require treatment prior to discharge to the RIB’s. Currently, the Water Treatment Plant (WTP) processes 380 GPM from BPW-5 and 110 GPM from GCW-6.

 

7.4.6

 Groundwater Flow Model

Supporting tasks for the Project include groundwater flow model development for dewatering assessment, predictions for passive inflow related to planned underground mine development and workings (UGWs), and regional-scale permitting requirements for potential effects on Kelly Creek Basin. To conduct this work, HGL (2023) subcontracted and supervised INTERA Incorporated (INTERA) to construct the groundwater flow model, to calibrate the model using historical records, and finally to use the model to assess effectiveness of current production wells for dewatering the planned UGWs and predict passive inflow to UGWs during development.

Two endmember predictive model scenarios (Scenario 1 and Scenario 2) were developed to estimate the upper and lower limits of passive inflow to UGWs that may result from the UG mine plan and current dewatering infrastructure. The range of results reflects the current uncertainty in hydraulic parameters in the proposed mine area. Quarterly stress periods are used to represent the predictive period spanning 17 quarters (approximately 2024 through 2028), when the planned UGWs are developed in accordance with the July 2023 Granite Creek underground mine plan provided by Practical Mining LLC (2023). Figure 7-26 provides a graphical summary showing passive inflow rates during UGWs build-out on a quarterly basis, as well as total dewatering rates from the production wells and combined dewatering rates from passive inflow and production well dewatering rates.

Passive inflow for Scenario 1 was predicted at 50 gallons per minute (gpm) in early time, then increasing to 3,450 gpm through the predictive period, whereas Scenario 2 predicted approximately 50 gpm initially, and then increasing to 1,900 gpm. Differences between model scenarios were largely controlled by the conductance parameter defined in the model, which is linked to the hydraulic conductivity value that represents drain cells in the model representing

 

 

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UGWs development. Total dewatering rates from wells APW-1, BPW-3, BPW-5, and GCW-6 were predicted to decrease with the progression of the planned UGWs due to the decline hydraulic head in the underground workings. The inability of dewatering wells to sufficiently depressurize the aquifer ahead of planned underground mining results in increasing passive inflows during UGWs build-out.

The range in model results is reasonable given the current understanding of the groundwater system, observed water level data, and responses to pumping in nearby wells (both during the pumping test and operational pumping). Based on historical dewatering rates required to dewater Mag and CX Pits and likely compartmentalized hard-rock groundwater flow regime of the UGWs, the actual passive inflow rates during UGW development may trend toward the lower limit endmember of the predicted range at the end of UG mining. Results of the modeling work are preliminary and updates are underway to refine the work with updated calibration both at local and regional scales.

 

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Figure 7-26 Predictive and Passive Inflows from Scenarios One and Two

 

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8

Sample Preparation, Analysis and Security

8.1 Sampling Methods and Approach

Drilling at the Property used both surface RC and core drilling along with underground core and RC drilling. The RC drilling was used primarily to pre-collar holes to bedrock followed by core drilling. This was done to minimize costs by not core drilling through unmineralized material overlying the mineralized fault zones. Core drilling provides a higher confidence in sample quality versus RC drilling along with providing additional data for engineering studies and detailed geologic definition of structurally controlled high grade mineralized zones.

The primary objective of the drilling programs was to collect clean, uncontaminated representative samples that were correctly labeled when drilled and logged, and that could be accurately tracked from the drill rig to the assay laboratory. Atna, PMC (Barrick) Exploration, and i-80 Gold used similar sampling and analytical protocols.

 

8.1.1

Reverse Circulation Drilling

 

8.1.2

Sampling Methods

In this drilling method, cuttings produced by the bit are sent up the drill pipe into a cyclone at surface, where the sample is homogenized prior to collection. From the cyclone, the sample is processed through a rotary splitter that takes a representative split of the sample (usually a quarter split), sending a split portion to the sample port, with the remainder to the reject port. Samples are placed into 10-by-17-inch sample bags that have been clearly labeled with the drillhole number and a unique numbering sequence prepared beforehand using a spreadsheet. This spreadsheet helps in tracking bag numbers, feet drilled, and quality control samples. A representative sample of each interval drilled is also preserved in chip trays that are clearly labeled with the hole number and drill interval for future reference.

 

8.1.3

Recovery

Sample recovery for RC drilling is measured by weight of material collected, which is usually eight to ten pounds of material from the quarter split in a typical six-inch diameter hole. Historical RC sample recovery was excellent. Full five to ten-pound bags of sample were collected from every interval. The only exception were 15 samples out of 6,100 that were collected by Atna. The missing samples occurred in an isolated zone of badly broken ground.

 

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8.1.4

Sample Intervals

Typical truck-mounted RC drill rigs use 20-foot drill rods, with samples collected in five-foot intervals. Atna, PMC, and i-80 used this sampling procedure in their drilling programs.

For each RC hole drilled, the drill crew was provided with a sequentially numbered set of sample bags. The outsides of the bags were marked with the drillhole number and a sample number.

To ensure that blanks and standards were inserted into the sample stream correctly (every tenth sample), several steps were taken. First, the sampler was provided with chip trays that were labelled with both the true footage and the corresponding bag number. Second, an incompletely labeled set of sample bags was provided that did not include bags for the standards or blanks. Third, since the total depth of the hole was not known prior to drilling, bags for duplicate samples (collected every 100 feet) were labeled with the letters “A”, “B”, “C”, etc. and flagged with a tear-off paper tag. For i-80 Gold, duplicate samples were not labeled with a letter but rather were kept in the same number sequence and noted as a duplicate on the sample sheet for the driller.

Samples were allowed to drain/dry at the sample site, which was routinely visited by the geologist in charge of the drill program to ensure accurate numbering of the sample suite. Once drained and/or dried, the samples were re-located from the drill site to the shipment staging area, where personnel relabeled the bags containing the duplicate samples by assigning the correct sequential number in the case of Atna and PMC. This ensured that they were “blind” to the laboratory personnel. The samples were then loaded into 4 x 4 x 3-foot wooden or plastic crates in preparation for pickup by the assay lab.

 

8.1.5

Logging

Representative rock chips for each 5-foot run were collected in clearly labeled 20 compartment plastic chip trays. These trays were taken to the logging facility, where the geologist logged the chips with the aid of a binocular microscope. For Atna and PMC, the geologist recorded lithology, mineralization, alteration, and other pertinent features on a paper drill log. A schematic graphic log was also produced to aid in interpretation of the stratigraphic sequence. Geologists with i-80 Gold recorded geologic information directly into the acQuire database.

8.2 Diamond Drilling

 

8.2.1

Sampling Methods

At the drill site, the drill crew was responsible for obtaining a complete and representative sample of the cored interval. This interval is usually five feet in length but may be shorter depending on

 

 

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how difficult the ground conditions are. Core is recovered from the core barrel via a wire line core tube, which may be outfitted with an inner “triple-tube.”

For core obtained using a triple-tube system, the core was placed on a rack, and the drill crew recorded rock quality determination (RQD) values on a worksheet and photographed the core. For holes drilled with conventional core barrels, RQD values were recorded later by a geologist from the core in the box.

At the drill site, once the RQD values were recorded and the core photographed, the drill crew placed the core in waxed cardboard boxes that were labeled with the company name, Property, hole ID, box number, and from-to footage. Core boxes were partitioned in five, two-foot-long sections totaling 10 feet in length. As core was drilled, it was placed in the core boxes in sequential order from top of the run to bottom of the run. A wooden block was inserted at the end of each run, and at the driller’s discretion, to indicate problems with drilling, such as caving, voids, or core tube mismatches. The last block of each run was marked with the ending footage on the thin edge of the block and two numbers on the larger surface.

If the core was not photographed for RQD purposes, the drillers marked the breaks they made to fit the core into the core boxes with the letter “M or X” on each side of the break, so it was not counted in the RQD analysis. After boxing, each core box was securely closed with elastic banding and loaded into the driller’s vehicle for transport to the logging area, at which point it was unloaded and logged. i-80 Gold transports all core from a staging area at the mine site to the Lone Tree mine site logging facility with a third party contractor on a flat-bed trailer once per day.

 

8.2.2

Recovery

Core recovery is measured by the ratio of the length of drill core recovered versus the length of the drilled run and is expressed in percent. Core recovery was excellent with greater than or equal to 99% core recovered (Golder Associates, 2014). Where core loss was recorded, it amounted to less than two feet in zones where voids were present in the stratigraphy.

 

8.2.3

Sample Intervals

Once the core was logged, the geologist determined the sample intervals to be sent to the laboratory. The geologist adhered to a set of guidelines to better define boundaries between mineralized material and barren samples. Original core blocks, inserted by the driller to mark the end of a drill run, served as the primary sample boundary, subject to the rules below; where a conflict existed between the inserted core blocks and the guidelines, the guidelines prevailed, and extra blocks were inserted by the geologist to compensate:

 

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A sample must not cross a geologic contact.

 

   

A sample must not cross an obvious alteration boundary, including oxidation.

•  A sample must not exceed seven feet long for Atna/PMC and ten feet long for i-80 and only be that long if it occurred in barren material, with 5-foot (1.5-meter) samples being the optimum.

•  Any core blocks that do not mark a sample boundary, for whatever reason (such as “cave,” “loss,” “void,” etc.) must be labeled in black marker for photographic visibility.

Each block that marked a sample boundary was outlined or highlighted in red marker, and the interval boundaries were entered into a sample sequence log. Sample intervals generally ranged from one to six feet in length and averaged 4.6 feet.

During the core sampling process, the sampler was provided with the geologic core log and the sample sequence to allow the sampler to have a better understanding of why and how the sample boundaries were picked and to act as a check on the geologist’s accuracy.

For Atna and PMC, the condition of the rock and whether it was mineralized or not dictated the splitting method of the core. Unmineralized rock was split with a hydraulic splitter. Mineralized and silicified intervals were sawn with a water-cooled diamond-bladed rock saw. Mineralized un-silicified intervals were also typically sawn, but in some instances split with the hydraulic splitter. For i-80 Gold, all competent samples were sawn with a water-cooled diamond-bladed rock saw. Broken mineralized core was separated and divided into two equal portions by all companies.

To avoid sampling bias, the core was sawn or split parallel to the vertical axis of the core. The portion of the core to be saved was placed in the core box in its original position with the core blocks in place, and the box was rubber banded for additional security. The sampled half of the split core was bagged, and the bags were placed in 4- x 4- x 3-foot (1.2- x 1.2- x 0.9-meter) wooden or plastic crates for shipment to the laboratory. For Atna and PMC, the remaining core was palletized, covered with tarps, and moved to an outdoor cement pad for storage and reference. It is unknown if this storage facility was secure. i-80 Gold palletized the core, covered it with tarps, and moved it to a lay-down yard near the cutting facility at Lone Tree. The facility at Lone Tree is secured by a locked gate at all times.

 

8.2.4

Logging

Once the core was received at the logging facility, it was arranged sequentially from top of the hole to bottom of the hole.

 

 

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For Atna and PMC, data was captured on paper drill logs including footage of the core runs, lithology, alteration, major structural features, bedding dips, and fractures. A horizontal line was drawn across the log, indicating footage where core blocks were present within the drilled core. Footage of core drilled and recovery were also recorded. Intervals with no recovery were indicated on the drill log by horizontal lines crossing the entire page, with a blanked-out zone of “no information,” making it readily apparent where information was missing. For i-80 Gold all geologic data was logged directly into the acQuire database.

Any discrepancies in the footage shown on the core blocks or in core recovery were noted by the logging geologist on the log. Where there was missing core, additional core blocks were inserted by the geologist reflecting the missing interval and a cursory explanation written on the core block stating why the interval was missing.

Additionally, for Atna and PMC, graphic logs of the lithology were also produced to reflect the major rock types using conventional or agreed upon symbols. Major structural features including contact relationships, dips and fractures, bedding, and veins were plotted on the log and described as angle from core axis. Alteration and mineralization styles were also recorded along with a description of the lithology.

8.3 Sample Security

Methods for securing samples by companies conducting work at the Property prior to the formation of PMC are unknown. Between 1970 and 1996, during which time PMC was actively mining at the Property, samples were sent to the mine laboratory for analyses. It is not known what provisions PMC employed for sample security.

When Homestake operated PMC, samples were picked up and transported to the laboratory by ALS Chemex as part of the chain of custody. In 2003 and from 2007 to 2008, Barrick, as operator of PMC, conducted drilling programs. It is uncertain what protocols were employed by Barrick to ensure sample security.

Atna conducted exploration and development drilling between 2004 and 2006 and from 2012 to 2015. Once a set of samples was ready for shipment to the laboratory, the laboratory was contacted for a job number and a pickup time by the laboratory scheduler. It is unknown if samples were stored onsite or whether the sample storage area was secured. Both RC chips and core samples were placed in numbered bags and the bags placed in 4- x 4- x 3-foot wooden crates for shipping, along with a transmittal sheet indicating whether the samples were core or RC cuttings, the range of sample numbers, and the total number of samples. In some instances, an Atna geologist travelling to Reno delivered samples to the lab.

 

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During the control of the property by i-80 Gold, samples are secured at all times. At both the Granite Creek and Lone Tree mine sites all samples are kept behind a locked and controlled gate on the property until pick-up by a third party contractor or assay lab.

8.4 Sample Preparation and Analysis

 

8.4.1

PMC 1970 – 1996

Sample preparation procedures for the Granite Creek Mine were not recorded.

PMC’s standard assaying practice was to run assays using atomic absorption (AA) methods. For all assays, this was generally done on a cyanide leach to aid in identifying leachable material (Sim, 2005). At some unknown point, PMC changed this to only run fire assay with AA finish on samples over 0.01 opt (0.34 g/t). Check assays were performed on high-grade zone samples at third-party laboratories. Detection limits for the PMC samples varied from <0.003 to <0.001 opt (<0.1 to <0.03 g/t), depending on the age of the assay.

 

8.4.2

PMC - Homestake 1997 – 2000

When Homestake operated PMC, assays were analyzed by ALS Chemex in Reno, NV (ALS). Samples were prepared at ALS as follows:

 

   

Primary crush and mill to 80% passing -10 mesh

 

   

300-gram split of material for pulverization to 90% passing -150 mesh

 

   

30-gram split for digestion and assay

Samples were assayed using the Au-AA23 fire assay method with AA finish. Analyses were reported in parts per billion (ppb). Samples reporting Au values greater than 10,000 ppb were re-assayed by fire assay with a gravimetric finish.

Detection limits for gold analyses performed by ALS Chemex were 5 ppb and 0.0005 opt (0.017 g/t). For statistical purposes, most of the Homestake holes that reported “detection limit” gold were converted to 2.5 ppb and 0.0003 opt. (These values were subsequently converted back to -5 ppb and -0.0005 opt in the current database).

 

8.4.3

PMC Barrick 2000 – 2008

American Assay Laboratories (AAL) located in Sparks, Nevada was used by PMC (Barrick) to prepare and analyze samples generated from its drilling programs.

 

 

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Samples were dried, weighed, and crushed using either a roll or jaw crusher. A split of crushed material was pulverized for further analytical work. Samples were analyzed for gold using a one assay ton (29.116 gram) fire assay with AA finish (Fire AA). Samples with a fire assay greater than 0.005 opt (0.17 g/t) were subject to a cyanide soluble leach assay by AA spectroscopy to determine gold recovery and carbon and sulfur analysis for metallurgical evaluation. Samples returning an initial gold assay >5 parts per million (ppm) were subject to fire assay with a gravimetric finish.

In addition to gold, PMC (Barrick) also had the samples analyzed for an additional 69 elements using an aqua regia digestion with an Induced Coupled Plasma Atomic Emission Spectroscopy (ICP-AES finish). PMC (Barrick) employed its own internal quality assurance/quality control (QA/QC) protocols. Once the assay results were received via email, the exploration database manager loaded the assay data into AcQuire database management software (ACQ). The ACQ software evaluated the gold values of the standards and flagged any standards that performed outside of acceptable limits. Failed standards were documented and reviewed by the geologist in charge of the project. Depending on the rate of failure, a selection of samples, or possibly the entire batch, was rejected and another round of analyses requested by the geologist.

When samples needed re-assaying, the lab was notified of the failures, and a list of samples to be re-assayed were sent to the lab. Upon receipt of the results of the re-assayed samples by the database supervisor, they were loaded into ACQ, and XY-scatter plots were generated for the geologist to review for approval or rejection. Should the second round of analyses be rejected, a third round would ensue until acceptable results were achieved. Check samples were also collected and sent to a second lab to evaluate potential laboratory bias. It is unknown which laboratories were used to analyze the check samples.

 

8.4.4

Atna 2004 – 2013

Atna used Inspectorate American Laboratories (IAL), an ISO 9002-accredited facility located in Reno, Nevada, as their primary analytical lab for the Granite Creek Mine Project. Sample preparation procedures used by IAL follow.

The samples were dried and weighed prior to crushing. Crushing used a two-stage process. Once the sample was dried, it was passed through a jaw crusher to reduce it to a uniform size. It then passed through a roll mill to reduce the sample to >80% passing -10 mesh. A 300-gram split of this material was obtained using a Jones riffle splitter. The split material was further reduced to >90% passing -150 mesh using a ring and puck pulverizer.

 

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After pulverization, a 30-gram sample of pulp was taken and digested and analyzed for gold using standard fire assay with AA finish. Samples returning gold values greater than 3 g/t were subjected to gravimetric analyses.

 

8.4.5

Atna Underground 2011 – 2016

The new mine lab constructed adjacent to the administration building in 2012 was in operation from 2012 to 2016.

Underground samples were transported to the on-site laboratory by Atna personnel. Samples were logged in and checked against sample transmittal sheets. Samples were then dried and weighed before being passed through a small jaw crusher to minus 3/8-inch (0.95-centimeter) passing. Crushed material was then passed through a Jones splitter, multiple times if necessary, to produce a 200-gram to 300-gram sample split for pulverization. The pulp split was then transferred to the ring and puck pulverizer for grinding to 80% passing 150 mesh. Pulverized material was weighed out to a 30-gram fire assay sample charge (Pinson Mine, 2015).

 

8.4.6

i-80 Gold 2021 – 2025

Underground production samples are transported daily to the Lone Tree mine site assay lab by a third party contractor. Samples are logged in and checked against sample transmittal sheets. Samples are then dried in an oven at 80°C before being passed through a large jaw crusher reducing size to 2 inch and then a small jaw crusher passing 1 1⁄2 inch. A 250g split is then taken and the sample dried again. The samples are then pulverized to 85% passing 150 mesh creating a 250g pulp. Pulverized material is weighed out to a 30-gram fire assay sample charge. All samples are also assayed by a 10g gold cyanide shake method for an hour.

Exploration samples of both core and RC from underground and surface are assayed at a third party laboratory. These assay labs have included ALS Minerals, American Assay Laboratories, and Paragon Geochemical, all located in Sparks, Nevada with their respective procedures listed below.

Samples submitted to ALS Minerals (ALS) of Sparks, NV, an ISO 9001 and 17025 certified and accredited laboratory were assayed for gold and multi-element. Samples submitted through ALS are dried, crushed, and pulverized to 85% passing -200 mesh, creating a 250g pulp. Samples are then analyzed using Au-AA23 (Au; 30g fire assay) and ME-ICP41 (35 element suite; 0.5g Aqua Regia/ICP-AES). Samples containing greater than 10 g/t gold are analyzed by fire assay with a gravimetric finish (Au-GRA21). ALS also undertakes their own internal coarse and pulp duplicate analysis to ensure proper sample preparation and equipment calibration.

 

 

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Samples submitted to American Assay Laboratories (AAL) of Sparks, NV, an ISO 9001 and 17025 certified and accredited laboratory were assayed for gold and multi-element. Samples submitted through AAL are dried, crushed, and pulverized to 85% passing -200 mesh, creating a 300g pulp. Samples are then analyzed using FA-PB30-ICP (Au; 30g fire assay) and ICP-2OA36 (36 element suite; 0.5g Aqua Regia ICP-OES+MS). Samples containing greater than 10 g/t gold are analyzed by fire assay with a gravimetric finish (G-FAAU). AAL undertakes their own internal coarse and pulp duplicate analysis to ensure proper sample preparation and equipment calibration.

8.5 Data Validation

 

8.5.1

Summary

The Property database has been subjected to three major campaigns of data validation by Atna, Barrick, and most recently OMC. The details of data validation completed by Atna and Barrick are described in detail in previous Technical Reports (Sim, 2005; Atna Resources Ltd., 2007; Gustavson, 2012; Golder Associates, 2014; AMC, 2020), Atna (2007), Gustavson (2012), and Golder (2014). A summary of this work is described herein.

 

8.5.2

Atna Review of Prior Data

Atna completed a detailed review of historic data as part of due diligence studies, upon acquiring the Property. This process involved comparing data stored within a historic Microsoft Access database with digital files, databases, Vulcan files, and records stored onsite. Errors were corrected based on a “well maintained filing system containing most, if not all, drill logs, downhole surveys and Homestake assays” (Atna Resources Ltd., 2007). Validation errors such as overlapping samples and length discrepancies (i.e., surveys beyond hole depth) were investigated and corrected as appropriate.

Atna was unable to verify PMC analytical results because much of the historical analysis had been completed using the mine laboratory, and original certificates were not available. To assess historical analytical results, Atna reanalyzed 652 drill sample pulps from mineralized intercepts within the CX and Range Front target area. The pulps were sourced from the onsite pulp library maintained by PMC. Check pulp samples were submitted with Certified Reference Materials (CRMs). Atna concluded that re-assay results confirmed the accuracy of original Homestake and PMC assay results.

Atna subsequently completed two separate database audits. The first audit involved the selection of 20% of the 370 holes within the database, extracting assays greater than 0.08 opt (2.74 g/t) and checking assays. Out of 216 assays, 16 errors were noted and corrected. A second audit was

 

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completed by checking 15% of the drillholes completed by Atna in the Phase 2 program of 2006. Out of 1,653 assays, a total of 12 errors were identified.

 

8.5.3

Barrick Review of Prior Data

On exercising their earn-back option with Atna, Barrick conducted a detailed verification review of the historical drillhole database. This included reviewing the use of standards, blanks, and duplicates along with a second round of checks on the data entry and database maintenance. The results of the verification program are documented in an internal Barrick report that concluded that, “…10% of the database was checked, and it was considered adequate for use in a Scoping Level study…” (Golder Associates, 2014).

Barrick broadened the scope of their investigation of potential Mineral Resources at the Granite Creek Mine Project to include open pit potential and initiated a check of the accuracy of the historical database within an area of interest, which included checks on drillhole collars for 2,014 holes.

Barrick contracted Geostrata LLC of Bluffdale, Utah, to complete data verification checks on historical data. Collar coordinates, downhole surveys, from and to intervals, and assay values were reviewed. Six errors were identified out of 208 collars checked. Errors comprised transcription errors, where the collar coordinates or hole length was incorrect, and field errors, where data had been entered into the incorrect field. Out of a total of 18,013 assays, a total of 184 errors were identified (1%). Errors comprised:

 

   

Data in the database but not in the drill log and vice versa

 

   

Incorrect numbers in the database according to the drill log

 

   

Discrepancy transcribing nil, trace, no sample, or detection limit values

 

   

Sample type is recorded in the drill log but not in the database

 

   

No assay data is available via certificate or drill log, but there was data in the database

Table 8-1 provides a summary of the errors.

 

 

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Table 8-1 Summary of Errors Within the Granite Creek Project Database

 

Company    Total
Assays
Reviewed
   Missing
Data
   Incorrect
Numbers
   Discrepancies
Nil, Trace, No
Sample,
Detection
Limit
   Sample
Type Errors
   No
Certificate
or Drill Log
   Total
Number of
Errors

Atna

   1867    3 (0.16%)    0    0    0    0    3 (0.1%)

PMC (Barrick)

                            

Cordex

   179    0    0    7 (3.9%)    0    0    7 (3.9%)

Cordilleran

   435    4 (0.9%)    0    2 (0.4%)    0    1 (0.2%)    7 (1.6%)

PMC (Homestake)

   3319    5 (0.15%)    3 (0.09%)    11 (0.3%)    11 (0.3%)    2 (0.06%)    32 (0.9%)

Pinson Mine Co.

   12,392    71 (0.57%)    47 (0.3%)    16 (0.1%)    0    1 (0.008%)    135 (1.0%)

Total

   18,013    83 (0.46%)    50 (0.27%)    36 (0.19%)    11 (0.06%)    4 (0.02%)    184 (1.0%)

 

8.5.4

OMC Data Compilation and Validation

 

8.5.4.1

Database Compilation

In January of 2017, OMC contracted Maxwell Resources (Maxwell) to perform data migration of the drillhole database into their proprietary DataShed database software. Maxwell was supplied with collar, downhole survey, lithology, and original assay files.

While in operation, both mine labs used a digital assay file management system to keep track of assay and other data generated from drilling programs. Only raw digital assay files were located for assays generated by the new mine lab. The new mine lab used an Excel file with multiple tabs to record assay data throughout the assaying process. Only the tab marked as “final assay” was used by Maxwell and OMC for data uploads into DataShed. Assay data from the old mine lab was only available as paper copies with hand-written assays on the form. These paper copies were used to validate assay data in the DataShed database.

Maxwell supplied OMC with an SQL database in February of 2017. During the process of migrating the database into the new software, Maxwell noted that assay files were in various formats and that there were multiple errors in collar information.

All gold assays, including Cyanide Au and calculated values, were uploaded into one Au field. There were also a significant number of generic methods that had unknown (“UN_UN”) listed for the analytical method. The new data uploaded from the various labs added more analytical methods. After reviewing the database, it was determined that additional Au fields were needed to separate out the various analytical methods, i.e., Cyanide Au (Au_CN field), along with a field for calculating ounces per ton (opt) (AU_CALC field). It is important to be able to specify the

 

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analytical method used for Au analyses since DataShed automatically ranks the methods from most reliable method to least reliable method.

 

8.5.4.2

Database Corrections

In 2018, OMC corrected the errors found by Maxwell during their data migration process. Errors that were corrected included duplicate holes, core recovery issues, and interval data that went past total hole depth.

In addition, assay batches that were not uploaded correctly were flagged with a “NOCERT” or “assay method unknown” identifier.

In April of 2019, OMC contracted AMC and CSA Global to perform separate Mineral Resource updates on the Pinson underground mineralized zone. After detailed review of the drillhole database, AMC and CSA Global separately expressed concern with the number of “NO CERT” and “assay method unknown” assays. An area of interest surrounding the underground mineralized zone was subsequently defined, and original assay certificates were sourced and reloaded where possible. Analytical methods associated with assay data were updated during this process. Standards and blanks were also compiled and uploaded.

Details of assays reloaded are presented in Table 8-2 and Table 8-3.

Table 8-2 Initial Data Set and 18 April 2019 Data Subset

      Starting Database    18 April 2019 Database
Samples    77,475    77,660
Number of samples with “NOCERT”    58,740    48,498
Percentage of database with “NOCERT”    75.80%    62.40%

Table 8-3 Assay Certificates and Samples Uploaded by Laboratory

Laboratory    Number of Batches    Number of Samples
American assay lab    66    9,098
Inspectorate    164    13,626
Pinson Mine    132    2,921
Total    362    25,645

Notes: Numbers are from the defined area of interest

Certificate headers contain the certificate identification, analyte, laboratory method, and assay unit. The raw assay headers from all the labs had to be re-formatted to facilitate direct import to DataShed. All certificates, regardless of the lab of origin, had the identifier “_2019” added to the end of the certificate number to aid in separating assays from the same certificate but which had

 

 

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different loading parameters. The 18 April 2019 database described in this section was used in the Mineral Resource estimate.

Significant work has been completed on the transfer of the old database into the new DataShed database and additional clean-up work still needs to be performed on the DataShed database to ensure its completeness and increase confidence in the data.

8.6 Quality Assurance/Quality Control Overview

QA/QC data has been compiled from available databases for all drilling activities completed since 2005. No QA/QC data is available for work occurring prior to this time.

Drilling programs completed at the Property between 2005 and 2015 included QA/QC monitoring programs, which comprised the insertion of CRMs, blanks, and duplicates into the sample streams on a batch by batch basis. Table 8-4 provides a summary of QA/QC samples included during this period.

Table 8-4 QA/QC 2005 – 2015

Year   Company   Drill samples   CRMs   Blanks   Field duplicates

2005

  Atna   7,330   267   289   23

2006

  4,859   265   263   39

2007

  Barrick   3,644   123   107   2

2008

  17,661   403   265   197

2012

  1,515   0   0   0

2013

  Atna   3,360   0   0   0

2015

  1,320   23   0   0

Total

      39,689   1,081   924   261

Notes:

   

Counts of individual samples. Multiple analyses types per sample (i.e., fire assay and gravimetric).

   

Based on year drilled.

Source: AMC Mining Consultants (Canada) Ltd. using data provided by Osgood Mining Company LLC

Table 8-5 shows the insertion rates of QA/QC samples between 2005-2015.

Table 8-5 QA/QC 2005 – 2015 Insertion Rates

Year   Company   CRM’s   Blanks   Field duplicates   QA/QC 1

2005

  Atna   3.6%   3.9%   0.3%   7.9%

2006

  5.5%   5.4%   0.8%   11.7%

2007

  Barrick   3.4%   2.9%   0.1%   6.4%

2008

  2.3%   1.5%   1.1%   4.9%

 

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2012

      0.0%   0.0%   0.0%   0.0%

2013

  Atna   0.0%   0.0%   0.0%   0.0%

2015

  1.7%   0.0%   0.0%   1.7%

Total

      2.7%   2.3%   0.7%   5.7%

Notes:

   

Counts of individual samples. Multiple types of analyses per sample (i.e., fire assay and gravimetric).

   

1 Insertion rate for CRM, Blanks and Field Duplicates combined.

   

Based on year drilled.

Source: AMC Mining Consultants (Canada) Ltd. using data provided by Osgood Mining Company LLC

8.7 Certified Reference Materials

A total of 37 different CRMs were used at the Property between 2005 and 2015. CRMs were supplied by Rocklabs of New Zealand.

CRMs comprised on average 2.7% (and up to 5.5%) of samples submitted to the laboratory. CRM insertion formed part of the QA/QC program consistently in the period between 2005 and 2008. CRMs, during this time, were generally included systematically at a rate of 1 in 20 to 1 in 25 samples. CRMs do not appear to have been consistently used since 2008.

CRMs used in the 2005 and 2006 programs are discussed in the 2007 NI 43-101 Technical Report titled “Technical Report Update Pinson Gold Property, Humboldt County, Nevada, USA” effective 1 June 2007 (Atna Resources Ltd., 2007). There is no documentation available regarding CRM procedures for programs after 2006.

Rocklabs CRMs were stored in bulk in plastic bin in the logging trailer. Individual CRMs were created by measuring 100 grams of the appropriate CRM into kraft envelopes. Packaged CRMs were then stored in separate labeled bins and inserted regularly into the sample stream.

Table 8-6 and Table 8-7 summarize CRMs by year and company.

 

 

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Table 8-6: CRMs used in each year

 Period     Company      # CRMs    CRMs Used

2005

  Atna    16   OxA45, OxE21, OxH29, OxK18, OXL25, OxN33, OxP32, SF12, SG31, SI15, SJ10, SK11, SN16, SP17, SQ18, UNKNOWN

2006

   16   OxA45, OxE21, OxH29, OxI54, OxJ36, OxK18, OXL25, OxN33, OxP32, SF12, SI15, SJ10, SK11, SN16, SP17, SQ18

2007

  Barrick    15   OxA59, OxC58, OxD57, OxF53, OxG60, OxH52, OxI54, OxK48, OxN49, OxP50, SF23, SG31, SJ32, SK33, SN26

2008

   18   OxA59, OxC58, OxD57, OxF53, OxG60, OxH52, OxI54, OxJ36,OxK48, OxN49, OxP50, SF23, SG31, SI25, SJ32, SK33, SN26, UNKNOWN

2012

  Atna    0    

2013

   0    

2015

   6   OxK119, OxN117, OxP91, SK78, SN75, SP73

Table 8-7: CRMs Used by Year and Company (2005 – 2015)

CRM ID    Expected
Au Value
(ppm)
   Stand
Dev
   Number of CRMs used 1
   Atna    Barrick    Atna    Total
   2005    2006    2007    2008    2015

OxA45

   0.081    0.0069    2    13                   15

OxA59

   0.082    0.0052              3    37         40

OxC58

   0.201    0.007              7    30         37

OxD57

   0.413    0.012              13    42         55

OxE21

   0.651    0.026    30    26                   56

OxF53

   0.810    0.029              4    30         34

SF12

   0.819    0.028    36    18                   54

SF23

   0.831    0.027              7    38         45

SG31

   0.996    0.028    1         4    36         41

OxG60

   1.025    0.028              10    27         37

OxH52

   1.291    0.025              18    28         46

OxH29

   1.298    0.033    24    21                   45

SI25

   1.801    0.044                   22         22

SI15

   1.805    0.067    1    4                   5

OxI54

   1.868    0.066         1    6    33         40

OxJ36

   2.398    0.073         3         1         4

SJ10

   2.643    0.06    2    16                   18

SJ32

   2.645    0.068              5    30         35

OxK18

   3.463    0.132    21    2                   23

OxK48

   3.557    0.042              10    22         32

OxK119

   3.604    0.105                        3    3

SK33

   4.041    0.103              9    15         24

SK78

   4.134    0.138                        4    4

SK11

   4.823    0.11    21    26                   47

 

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CRM ID    Expected
Au Value
(ppm)
   Stand
Dev
   Number of CRMs used 1
   Atna    Barrick    Atna    Total
   2005    2006    2007    2008    2015

OXL25

   5.852    0.105    29    25                   54

OxN33

   7.378    0.208    33    28                   61

OxN49

   7.635    0.189              19    6         25

OxN117

   7.679    0.207                        2    2

SN16

   8.367    0.217    17    21                   38

SN26

   8.543    0.175              2    3         5

SN75

   8.671    0.199                        4    4

OxP91

   14.820    0.3                        3    3

OxP50

   14.890    0.493              6    3         9

OxP32

   14.990    0.44    3    15                   18

SP17

   18.125    0.434    25    32                   57

SP73

   18.170    0.42                        7    7

SQ18

   30.490    0.88    22    14                   36

Notes:

   

1 Counts of individual samples. Multiple analyses types per sample (i.e., fire assay and gravimetric).

 

   

Based on year drilled.

Source: AMC Mining Consultants (Canada) Ltd. using data provided by Osgood Mining Company LLC

8.8 GRE Discussion on QA/QC

The in-house QA/QC procedures for Granite Creek Mine Project (between 2005 and 2015) were reviewed. This review included:

•   a considerable quality of data analysis and validation work performed by AMC in prior technical reports (AMC, 2020), (AMC, 2019)

 

 

a review of available data, checked against the Granite Creek Mine Project database.

This review generated the following discussion and analysis.

 

8.8.1

GRE Discussion on CRMs

A total of 1,081 CRMs were inserted into the sample stream from 2005 to 2015 drilling campaigns program, including 555 CRMs by Atna in 2005, 2006, 2013, and 2015 and 526 CRMs by Barrick in 2007, 2008, and 2012. A total of 37 different CRMs were used at the property between 2005 and 2015 (Table 8-7).

Figure 8-1 shows a scatter plot of the certified value for each assay standard compared to the assay value obtained. The laboratory’s analytical results generally correlate well with the standard values

 

 

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with few outliers. A 45-degree line represents an excellent correlation between the standard assay certified value and actual assay results. This line passes through all of the sample sets, with the majority of the points directly adjacent to the line, indicating acceptable accuracy performance for the standards. Larger scatter is seen as the grade of the samples increases. The increase in scatter is within an acceptable range in the opinion of the QP.

Figure 8-1: Assay Standard Results (2005-2015)

LOGO

In addition to control charts contained in (AMC, 2020), GRE selected some additional control charts to monitor the analytical performance of an individual CRM over time and to validate prior conclusions Control lines are also plotted on the chart for the expected value of the CRM, two standard deviations above and below the expected value, and three standard deviations above and below the expected value. CRM assay results are plotted in order of analysis. Control charts at various grades for the two main campaigns of work are presented for select CRMs (outlined in Table 8-8) in Figure 8-2 to Figure 8-7.

Table 8-8: CRMs Selected by GRE for Control Charts

CRM    Au Value (ppm)    No. CRMs    Campaign

OxG60

   1.025    36    2007-2009

OxI54

   1.868    40    2007-2009

OXL25

   5.852    54    2005-2006

SG31

   0.996    41    2007-2009

SJ32

   2.645    36    2007-2009

SQ18

   30.49    36    2005-2006

 

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Figure 8-2: CRM OxG60 (2007 – 2009) FA-ICP-ES

 

LOGO

Figure 8-3: CRM OxI54 (2007 – 2009) FA-ICP-ES

 

LOGO

Figure 8-4: CRM OXL25 (2005 – 2006) FA-GRAV

 

LOGO

 

 

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Figure 8-5: CRM SG31 (2007 – 2009) FA-ICP-ES

 

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Figure 8-6: CRM SJ32 (2007 – 2009) FA-ICP-ES

 

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Figure 8-7: CRM SQ18 (2005 – 2006) FA-AAS

 

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In general, CRMs show reasonable analytical accuracy but relatively poor precision when compared against the certified standard deviation. This poor precision occurs in a number of CRMs from two laboratories over a period of four years. At this time, it is not possible to definitely determine the cause of CRM high failure rate.

 

8.8.2

GRE Discussion on Blanks

GRE reviewed and checked all blank samples in the database provided by i-80. Figure 8-8 shows the assay results of the blanks used in the QA/QC program between 2005 and 2008. A total of 1,249 blanks returned 270 excursion values, with a maximum value of 1.02 ppm Au. Apart from four samples, the remaining samples are below the probable lower limit of the cutoff grade. 78.4% of the samples are below the detection limit. GRE believes the results indicate there is no artificially introduced contamination in the sampling preparation process that would materially affect the mineral resource estimate.

Figure 8-8: Fire Assay Blank Samples (2005-2015)

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8.8.3

GRE Discussion on Duplicates

A total of 287 duplicate samples in the database provided by i-80 were checked by the QP. Figure 8-9 shows a comparison graph of the field duplicates. The scatter plots indicate some scatter in the data, with R2 values of 0.93. The scatter increases as the grade values increase but are still within acceptable ranges in the opinion of the QP.

 

 

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Figure 8-9: Laboratory Duplicate Comparison (2005-2015)

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8.9 Quality Assurance/Quality Control Overview by PM (2021-2025)

i-80 collected samples from 14 surface exploration holes drilled in 2021 and three exploration holes drilled in 2022 on the Property for the design of pen pits. Drilling programs completed at the Property in 2021 and 2022 included QA/QC monitoring programs, which involved the insertion of CRMs, blanks, and duplicates into the sample streams on a batch-by-batch basis. Table 8-9 provides a summary of QA/QC samples included during this period, while Table 8-10 shows the insertion rates of QA/QC samples for the 2021 and 2022 drilling programs.

Table 8-9: QA/QC 2021 and 2022

Year    Company    Drill samples    CRMs    Blanks   

Field

duplicates

   Preparation duplicates
2021    i80    1,395    55    51    34    38
2022    800    47    51    48    24
Total         2,195    102    102    82    62

Table 8-10: QA/QC 2021 and 2022 Insertion Rates

Year    Company    CRM’s    Blanks    Field duplicates    Preparation duplicates    QA/QC
2021    i80    3.9%    3.7%    2.4%    2.7%    12.8%
2022    5.9%    6.4%    6.0%    3.0%    21.3%
Total        

4.6%

   4.6%    3.7%    2.8%    15.9%

 

 

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8.10

PM Discussion on QA/QC 2021

i-80’s in-house QA/QC procedures in 2021 involved submitting 55 Certificate Reference Materials, 51 blank samples, 34 field duplicates, and 38 pulp duplicates to the laboratory for the 1,395 drill samples. The standards were sourced from CDN Resource Laboratories, and the blanks were purchased from Ron’s Seed and Supply in Winnemucca. The blank material consisted of 50 bags of Vigoro marble chips, which are gravel-sized. Field duplicates were 1⁄4 core samples, while RC duplicates were created by placing a splitter on the cyclone of the RC rig. As the rock emerged from the cyclone, it was evenly distributed between two sample bags.

i-80 geologists routinely reviewed their assay results. The results fall within the anticipated range of variability as described by the standards’ manufacturer, and as a result, the QP is of the opinion that there is no indication of systematic errors that might be due to sample collection or assay procedures.

8.10.1 PM Discussion on CRMs

i-80 used CRMs CDN-GS-7J, CDN-GS-8C, CDN-GS-30C, CDN-GS-P1A, and CDN-GSP6E for the 2021 drilling program. In total, CRMs for gold were inserted into the sample stream at a rate of four standards per 100 sample assays for all 1,395 core and RC samples for the 2021 drilling program.

Analysis of CRM charts for the high and lower gold grades showed no obvious errors or bias (see Figure 8-10 through Figure 8-14).

 

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Figure 8-10: CRM CDN-GS-7J for the 2021 Drilling Program

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Figure 8-11: CRM CDN-GS-8C for the 2021 Drilling Program

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Figure 8-12: CRM CDN-GS-30C for the 2021 Drilling Program

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Figure 8-13: CRM CDN-GS-P1A for the 2021 Drilling Program

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Figure 8-14: CRM CDN-GS-P6E for the 2021 Drilling Program

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8.10.2 GRE Discussion on Blanks

GRE’s QP reviewed and checked all blank samples in the database provided by i-80 for the 2021 drilling program. For all 1,395 drill samples, 51 blank samples were inserted in the sample stream at a rate of three and a half blank samples per 100 rock drill samples. Figure 8-15 shows the assay results of the blanks used in the QA/QC program in the 2021 drilling program.

The remaining samples, except four, are below the threshold, which is five times more than the detection limit. GRE believes the results indicate no artificially introduced contamination in the sampling preparation process that would materially affect the mineral resource estimate.

 

 

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Figure 8-15: Blank Results for the 2021 Drilling Program

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8.10.3 GRE Discussion on Duplicates

For the 2021 drilling program, i-80 considered 34 field duplicates and 38 preparation duplicates for all 1,395 core and RC, at a rate of 2.4 and 2.7 for field and preparation samples per 100 sample intervals. Field duplicate samples were prepared the same way as all assay samples and assayed at the laboratories.

The Q-Q plot for field duplicates shows a few scatters that are acceptable for field duplicates, confirming that high-grade mineralization zones are mainly associated with an inhomogeneous distribution of mineralization along samples (Figure 8-16).

The Q-Q plot for Preparation duplicates effectively indicates that there is no scatter, with R2 values of 0.99 (Figure 8-17).

 

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Figure 8-16: Field Duplicate Samples for the 2021 Drilling Program

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Figure 8-17: Preparation Duplicate Samples for the 2021 Drilling Program

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8.11

PM Discussion on QA/QC 2022

i-80’s in-house QA/QC procedures in 2022 were limited to submitting 47 Certificate Reference Materials, 51 blank samples, 48 field duplicates, and 24 pulp duplicates to the laboratory for all 800 drill samples. The standards were purchased from CDN Resource Laboratories, and the blanks

 

 

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were purchased from Ron’s Seed and Supply in Winnemucca the same as the material used in the 2022 drilling program.

Field duplicates are 1⁄4 core samples. RC duplicates were prepared the same as the 2022 drilling program.

The results fall within the anticipated range of variability as described by the standards’ manufacturer, and as a result, the QP is of the opinion that there is no indication of systematic errors that might be due to sample collection or assay procedures.

8.11.1 PM Discussion on CRMs

i-80 used CRMs CDN-GS-7J, CDN-GS-30C, and CDN-GSP6E for the 2022 drilling program. In total, CRMs for gold were inserted into the sample stream at a rate of 5.8 standards per 100 sample assays for all 800 core and RC samples for the 2022 drilling program.

Analysis of CRM charts for the high and lower gold grades showed no obvious errors or bias (see Figure 8-18 through Figure 8-20).

Figure 8-18: CRM CDN-GS-7J for the 2022 Drilling Program

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Figure 8-19: CRM CDN-GS-30C for the 2022 Drilling Program

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Figure 8-20: CRM CDN-GS-P6E for the 2022 Drilling Program

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8.11.2 GRE Discussion on Blanks

GRE’s QP reviewed and checked all blank samples in the database provided by i-80 for the 2022 drilling program. For all 800 drill samples, 51 blank samples were inserted in the sample stream at a rate of 6.3 blank samples per 100 drill samples. Figure 8-21 shows the assay results of the blanks used in the QA/QC program in the 2022 drilling program.

The remaining samples, apart from five, are below the threshold, which is five times more than the detection limit. GRE believes the results indicate there is no artificially introduced

 

 

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contamination in the sampling preparation process that would materially affect the mineral resource estimate.

Figure 8-21: Blank Results for the 2022 Drilling Program

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8.11.3 PM Discussion on Duplicates

For the 2022 drilling program, i-80 considered 48 field duplicates and 24 preparation duplicates for all 800 core and RC, at a rate of six and three for field and preparation samples per 100 sample intervals. Field duplicate samples were prepared the same way as all assay samples and assayed at the laboratories.

The Q-Q plot for field duplicates shows one scatter that is acceptable for field duplicates, confirming that high-grade mineralization zones are mainly associated with an inhomogeneous distribution of mineralization along the samples (Figure 8-22).

The Q-Q plot for Preparation duplicates effectively indicates that there is no scatter, with R2 values of 0.99 (Figure 8-23).

 

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Figure 8-22 Field Duplicate Samples for the 2022 Drilling Program

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Figure 8-23: Field Duplicate Samples for the 2022 Drilling Program

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8.12 Conclusions

Drilling programs completed at the Property between 2005 and 2015 have included QA/QC monitoring programs that have incorporated the insertion of CRMs, blanks, and duplicates into the sample streams.

 

 

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In 2021, GRE QP reviewed all of AMC’s work on available QA/QC data between 2005 and 2015 (AMC, 2020). In 2025, i-80 provided all QA/QC data from surface exploration holes drilled in 2021 and 2022 to GRE, and GRE reviewed all of them and found no material errors. GRE also reviewed and checked QA/QC Procedures and the database provided by i-80. GRE confirmed discussions and recommendations made in prior technical reports and noted the following:

Formal, written procedures for data collection and handling should be developed and made available to PMC field personnel. These should include procedures and protocols for fieldwork, logging, database construction, sample chain of custody, and documentation trail. These procedures should also include detailed and specific QA/QC procedures for analytical work, including acceptance/rejection criteria for batches of samples.

 

 

A detailed review of field practices and sample collection procedures should be performed on a regular basis to ensure that the correct procedures and protocols are being followed.

 

 

Review and evaluation of laboratory work should be an on-going process, including occasional visits to the laboratories involved.

In general, the QA/QC sample insertion rates used fall below general accepted industry standards. For future exploration campaigns, standards, blanks, and duplicates including one standard, one duplicate, and one blank sample should be inserted every 20 interval samples, as is common within industry standards.

CRM samples show a reasonable level of accuracy but poor to moderate precision when using standard deviations provided by the CRM supplier. A maximum of three to five different CRM samples would be adequate to monitor laboratory performance at the approximate cut-off grades, average grades, and higher grades of the deposits.

Blank sample results are considered acceptable and suggest no systematic contamination has occurred throughout the analytical process.

Duplicate sample results show suboptimal performance, which may be a result of the heterogenous nature of mineralization, uncrushed samples, and sampling variance. Overall duplicate samples appear to be positively biased, with duplicate results returning higher grade than original samples.

Previous reporting suggests that umpire sampling has been completed at the Property. The results of this sampling were not available in the drillhole database and therefore the QP was not able to assess accuracy of the primary laboratory.

 

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Although it is not possible to guarantee that there are no material impacts on the local scale, overall, based on the checking and reviewing the previous technical report dated 2020, GRE considers the assay database to be acceptable for Mineral Resource estimation.

 

 

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9

Data Verification

The following text describes the activities performed and methods employed to personally verify the data that forms the foundation of the report. In summary, these methods included an on-site inspection of the project site and chip tray storage facility, check sampling, and manual auditing of the project database. The QPs noted no limitations nor failures to verify data.

9.1 GRE Site Inspection (2021)

GRE’s QPs Dr. H. Samari, T. Lane, L. Breckenridge, and R. Mortiz conducted an on-site inspection of the project on April 20, 2021.

While on-site, Dr. Samari conducted a general geological inspection of the Pinson area, including visual inspection of key geologic formations, lithologies, structural geology, and mineralization. Dr. Samari checked all lithologies on the ground with the latest prepared geologic maps prepared by Osgood (2016).

At the time of site visit, entire core boxes of four holes, BMAG-019C, BMAG-020C, UGOG-017, and UGOG-034, were ready to be inspected by Dr. Samari. Historic RC and core samples were stored at the Pinson site in the open space with thick water-resistant covers (see Photo 9-1).

Photo 9-1 Core Boxes Are Stored at the Granite Creek Site

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9.2 Visual Sample Inspection and Check Sampling

During the site visit on April 20, 2021, about 752 RC and core sample intervals from four separate drill holes, BMAG-019C, BMAG-020C, UGOG-017, and UGOG-034, were selected for visual inspection based on a review of the drill hole logs. The samples inspected accurately reflect the

 

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lithologies and sample descriptions recorded on the associated drill hole logs and within the project database.

In 2021, to verify the assay results, GRE collected a total of seven check samples (from four separate drill holes: BMAG-019C, BMAG-020C, UGOG-017, and UGOG-034) and two surface samples. All samples were bagged and labeled by GRE. Samples were packed and delivered by GRE the QP to Hazen Research Inc. (Hazen) in Golden, Colorado, USA (Photo 9-2).

On May 03, 2021, GRE received Hazen’s analytical report on nine selected samples by fire assay method for both gold and silver. The certificate of analysis from Hazen is given in Table 9-1. Except for sample UGOC-034-528-531, with an amount of 20 ppm silver, other samples showed less than three ppm of Ag.

A comparison of the original versus check assay values for the seven core samples shows good correlation between the results, with an R2 of 0.9944 (Figure 9-1). Standard t-Test statistical analysis was completed to look for any significant difference between the original and check assay population means. The results of the t-Test showed no statistically significant difference between the means of the two trials (original versus check assay).

 

 

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Photo 9-2 Sample Intervals Selected for Check Assay

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Table 9-1: Summary Table of Hazen Results with Original Assays

No.     Sample No.    Type of Sample   

Original Au

Assay ppm

  

Hazen Au

Assay ppm

1     UGOC-034-489-491.1    Core    27.497    23.7
2     UGOC-034-498-502    Core    10.697    9.9
3     UGOC-034-528-531    Core    25.851    26.0
4     UGOC-017-414-419    Core    1.24    1.5
5     BMAG-019-727-733.5    Chip    2.8389    2.6
6     BMAG-019-829-835    Chip    0.0309    <0.2
7     BMAG-020-781.4-785    Chip    2.844    3.0
8     GRE-R.S.S.1-St.3   

Surface sample

(Chip)

   -    <0.2
9     GRE-R.S.S.2-St.6   

Surface sample

(Chip)

   -    <0.2

Figure 9-1: Sample Correlation Plot

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Two surface rock chip samples, GRE-R.S.S.1 and GRE-R.S.S.2, were taken by GRE from the upper Comus formation (gray shale) and lower Comus formation (gray limestone), respectively (Photo 12 3). The assays show that when these two formations, which are the main gold deposit targets within the property, are not affected by faults, alterations, and mineralization conditions, they are barren (Table 9-1). The result emphasizes that mineralization on the Property exhibits strong structural control.

 

 

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Photo 9-3 The Location of Two Surface Rock Chip Samples, the Upper Comus Formation (GRE-R.S.S.1-St.3, left) and the Lower Comus Formation (GRE-R.S.S.2-St.6, right)

 

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9.3 Database Audits

The manual audit of the database by GRE compared approximately 10% of the original assay certificates with the database for the 2021 and 2022 drill campaigns intended for open pit design and found no material errors. GRE recommends that i-80 establish a routine, internal mechanical audit procedure to check for overlaps, gaps, total drill hole length inconsistencies, non-numeric assay values, or any missing information in the database. After any significant update to the database, an internal mechanical audit should be conducted. The results of each audit, including any corrective actions taken, should be documented to provide a running log of the database validation.

9.4 QP Opinions on Adequacy

Based on their area of expertise, the QPs present the following opinions on data verification and adequacy.

Based on the review of the project database and all existing project documents, and GRE’s observations of the geology and mineralization at the project during the site visit, GRE considers the lithology, mineralization, and assay data contained in the project database to be reasonably accurate and suitable for use in estimating mineral resources.

 

GRE believes that the metallurgical testing was completed for the Granite Creek project by a number of well-known commercial metallurgical laboratories. GRE reviewed the sample selection

 

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and compositing used in the metallurgical test work and found that the selection of samples was representative for this type of deposit and geology. GRE performed several mathematical tests to validate the metallurgical balances presented in the test work and found the data presented in the metallurgical reports to be consistent with practices performed by reputable independent test laboratories. A complete discussion of the test work is provided in Section 10. Though much of the work is historical in nature, the work appears to be professionally completed and is well documented, is supported by production data, and is suitable for estimation of CIL recovery calculations in this IA.

Mining and processing methods, costs, and infrastructure needs were verified by comparison to other similar sized open pit gold mines operating in the western USA and experience of GRE. Cost data used in the report was sourced from the most recent Infomine cost data report. All costs used in the analysis were verified and reviewed by GRE and were assessed to be current and appropriate for use. Finally, after the economic study was performed, the overall operating costs for different aspects of the operation (mining, process, and general & admin) were benchmarked against similar sized mines and recent technical reports to determine if they were similar; the results did benchmark well to other operations and economic studies.

The taxation rates used and applied were values available from US government sources at the time of the economic analysis.

9.5 Practical Mining Drillhole Database Verification

Practical Mining performed an initial inspection of the drillhole database to identify drillholes not aligning with surface topography or underground mine workings, as well as holes with excessive downhole survey deviation. i-80 staff performed statistical analysis on assays to identify potential downhole sample contamination in RC holes. Potential issues were identified in 62 holes, which were excluded from the database pending further review. These holes are listed in Table 9-2.

 

 

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Table 9-2 Excluded Drillholes

Sample

Contamination

  Excessive Downhole
Survey Deviation
  Assay Issue   Collar elevation doesn’t
correlate with topo
   Multiple Conflicting
Collar Survey Records
AMW-002   HPC-075   PRC-13-065   APRF-241    BRFC-002
AP4665-001   HPC-127     APRF-254    BRFC-003
APRF-284   Magmet-004     APRF-263    BRFC-007
BRFC-036   OCR-29     HPR-021    HPR-005
HPR-004   PRC-13-105     HPR-053    HPR-008
HPR-050   RH-27A     HPR-058    HPR-012
HPR-070       HPR-087    HPR-013
HPR-104       HPR-122    HPR-015
RH-27B       HPR-123    HPR-038
RHA-552A       HPR-124    HPR-041
      RCH-516    HPR-047
      RCH-659    HPR-059
      RH-130A    HPR-071
      RH-130B    OG2-155-1C
      RH-145    PM9214-1
      RH-146    RCH-1366
      RH-147B    RCH-1725
      RH-199    RCH-1727
      RH-200B    RCH-1730
      RH-210    RCH-1731
      RHA-1659    RCH-1732
      ATA-40    RH-345
         RH-346

851 drillholes were flagged for use in the estimate, and 59 holes (representing about 7% of the data set) were selected for detailed review. The holes selected for review were chosen to represent the area of interest in an even spatial distribution as well as represent different operators over time (PMC, Barrick, Atna and i-80.) Table 9-3 summarizes holes drilled by type and operator.

Table 9-3 Drill Holes Selected for Review by Type and Operator

Company    Core (or RC pre-collar
with Core Tail)
   RC    Rotary    Type Requested    Unavailable

i80

   131          21 core   

Barrick

   67    7       4 core   

Atna

   73    140       5 core, 8 RC    6 RC

PMC

   18    337    78    1 core, 17 RC, 3 rotary    1 core, 9 RC, 2 rotary

Totals

   289    484    78    59    18

Practical Mining requested original hardcopy data records for the selected holes including collar location surveys, downhole deviation surveys, geology logs, and assay certificates. Records were

 

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unavailable for 18 of the historically drilled holes, leaving 41 holes to be reviewed, which represent about 5% of the drillholes used in the estimation. The detailed data review demonstrated good overall correlation between the database and the original hardcopy data.

78 rotary holes are included in the database. These were drilled historically in the area of the CX pit prior to mining and have predominantly been mined out by the CX pit. For the current analysis, they were used primarily for modeling the location of the mineralized structures, which were depleted by the pit topography in the model. Intercept locations of rotary holes were corroborated by viewing with blast hole data where available (blast holes were not used in the model.)

Collar survey records were available for 38 holes. Mismatches were identified in collar locations for two of the selected holes. It was determined that the database was exporting planned locations instead of surveyed locations for a series of nine holes drilled in 2021, and the error was corrected. This does not affect the current analysis because the holes are outside the underground resource area. Practical Mining viewed all holes in Vulcan to confirm collars coincide with topography or underground mine workings, and 22 holes were excluded. Some of the excluded holes may become acceptable for use if their locations can be confirmed. 23 holes were found to have multiple conflicting collar survey records, and further attempts should be made to identify the correct location surveys; those 23 holes were excluded from the current resource estimation (Table 9-2).

Downhole deviation survey records were available for 28 holes. Of the 21 PMC holes requested, only eight appeared to have been surveyed, and none of the records were available. Of the 13 Atna holes requested, two did not have downhole surveys and 11 had been surveyed, of which four had records available for review. All of the requested Barrick holes had been surveyed, with one lacking archived records for review. Downhole survey records were available for all of the selected i-80 holes. All of the available downhole survey records match the values in the database. All hole traces were viewed in Vulcan and six with excessive deviation were excluded from the mineral resource estimation.

Geology logs were available for 41 of the requested holes. Logs match the database quite well. i-80 logs geology data directly into acQuire which eliminates the possibility of data entry error. Barrick paper logs matched the data in the database. Atna geology logs appear to have been simplified when they were digitized into the database, particularly in the alteration fields. The lithology and formation fields match fairly closely. PMC geology logs generally matched the database with three exceptions: one hole had a geology log that had not been entered in the database, one hole had a 5-foot discrepancy in the TD, and one hole had a discrepancy in the depth of a unit contact. Practical Mining viewed all drillhole traces coded by lithology in Vulcan and observed that the drill data coincides very well with i-80’s lithological and structural models.

 

 

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Assay certificates were available for 41 of the requested holes. Certificate assay values were compared with the values in the database and only one mismatch was identified, a minor error where the preliminary value was exported instead of the final value. Practical Mining viewed all drillhole traces coded by assay grades in Vulcan and noted that grade and thickness correlate well between adjacent holes and along geological contacts. Table 9-4 summarizes the number of holes reviewed per data field.

Table 9-4 Drillhole Data Fields Reviewed

      Collar
Surveys
   Downhole
Surveys
   Geology
Logs
   Assay
Certificates

Holes Reviewed

   38    28    41    41

Percent of Population

   4.5%    3.3%    4.8%    4.8%

Practical Mining concludes the database is suitable for use in the mineral resource estimation.

 

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10 Mineral Processing and Metallurgical Testing

 

10.1

Introduction

Multiple historical metallurgical programs were completed from 1999 to 2014. Dawson Metallurgical Labs and McClelland Labs completed these programs. Both Homestake Mining and Atna commissioned the work. The test programs included cyanide solubility testing, pregnant solution robbing testing, bottle roll testing, percolation column testing, carbon-in-leach (CIL) testing and autoclave testing. A more recent program was completed in 2023 by FLS focusing on oxidative pretreatment of the underground sulfide material using pressure oxidation.

The Granite Creek Mine was an operating open pit mine, processing oxide material using heap leaching and conventional milling from 1980 to 1999. Although the majority of the current resource at Pinson is similar to the historically processed material, the deeper material is more difficult to treat than the historic oxide material. Atna mined high grade mineralized material from the Ogee underground deposit between 2012 and 2013. This material was treated at the Twin Creeks autoclave facility under a toll treatment agreement. Newmont Mining previously operated the Twin Creeks facility and is now operated by Nevada Gold Mines, a Newmont / Barrick joint venture of which Barrick is the operating partner.

The Author has reviewed the historical metallurgical testwork programs on Pinson feed material including:

 

   

Report on Heap Leach, Direct and CIL Cyanidation, and “Preg-Robbing” Tests – Various Mag Pit Samples and Composites, and CX Pit Bulk Material, MLI Job No. 2532, Addendum, and Change Orders #1, #2, and #3, March 1999. (McClelland, 1999a) (Homestake)

 

   

Report on Column Heap Leach Testing, Pinson CX Pit Material Bulk Samples, MLI Job No. 2630, June 1999 (McClelland, 1999b) (Homestake)

 

   

Summary Report on Material Variability Testing – Mag Pit Pinson Drill Core Composites, MLI Job. No. 3746, 7 February 2013. (McClelland, 2013) (Atna)

 

   

Summary Report on Heap Leach Cyanidation Testing – Mag Pit Pinson Drill Core Composites, MLI Job No. 3746, 16 January 2014. (McClelland, 2014) (Atna)

 

   

Pinson Underground Autoclave-Cyanide Leach Tests, DML P-2895A,B&C, April 14,2006. (Dawson, 2006a) (Atna)

 

   

Results of Sample Preparation and Head Analysis on Ogee Samples, DML P-2895D April 2006. (Dawson, 2006b) (Atna)

 

 

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Wilmot Metallurgical Consulting Met Test Work Results Atna-Pinson Project, (no date) (Wilmot, 2006).

 

   

Dawson autoclave leach report, DML P-2895, final date November 3, 2005. (Dawson, 2005)(Atna)

 

   

Osgood Mining Company – Granite Creek, Pressure Oxidation, Carbon-in-leach Testing, Preg-Rob Testing, FLS, March 2023. (FLS, 2023)

These reports are the basis for estimating recoveries for the various materials at the Pinson site, including the Mag Pit, the CX Pit, and the Ogee underground. Recoveries used in support of the economic evaluation are detailed within the geometallurgical subsections within this Section 13.

Tables are from various reports, and the units have been left as direct quotes.

 

10.2

Metallurgical Test Work

 

10.2.1

McClelland Laboratories, Inc. March and June 1999

During March and June of 1999, McClelland Laboratories completed a test work program on samples sourced from the Mag Pit and CX Pit on behalf of Homestake Mining. The program was described as a multiphase program testing various Mag Pit bulk high grade samples, core and cuttings composites, and a CX Pit bulk material sample. The results were reported under two different McClelland labs job numbers: #2630 and #2532 (McClelland, 1999a; McClelland, 1999b).

Bulk material samples from Mag Pit (Mag Pit I to Mag Pit VI) and CX pit were tested for the following items:

 

   

Pregnant solution robbing (preg-robbing) tests to establish preg-robbing characteristics

 

   

pH control tests to determine lime requirements for subsequent agitated cyanidation tests and column percolation leach tests

 

   

Direct and CIL Cyanidation tests on the CX Pit (CX-2) bulk material sample to confirm the non-preg-robbing character of the material

 

   

Column percolation leach tests on the Mag Pit bulk material samples at 4-inch size and the CX bulk material sample at three different feed sizes: run-of-mine (ROM), 3-inch, and 3⁄4-inch.

Samples for the program were as follows:

 

   

Six Mag Pit bulk material samples that were sampled from the pit. These samples were labeled as “Mag Pit I” through “Mag Pit VI”. The specific coordinates of these samples were not included in the document.

 

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One bulk material sample from the CX Pit labelled “C2.” The specific coordinates for this sample were not provided.

 

   

Six Mag Pit composites were made of drilled cuttings material. These were labeled as “Mag Pit Cuttings Composite 1” through “Mag Pit Cuttings Composite 6.” Drill hole and drill intervals were documented for all of these samples.

 

   

Five Mag Pit composites were made of drill core material. These were labelled as “Mag Pit Drillcore Composite 1” through “Mag Pit Drillcore Composite 5.” Drill hole and drill intervals were documented for all of these samples.

 

   

A Mag Pit master composite was made up of the Mag Pit drill core composites. All intervals and proportions of the composites were documented.

The drillhole IDs and intervals used to make up the Mag Pit cuttings and drill core composites are shown in Table 10-1.

Table 10-1: Mag Pit Composites for 1999 Test Work Program

Sample Description  

Drill Hole

Identification

   Intervals
     
Mag pit cuttings composite 1    HPC - 129    505 to 555, 600 to 615
     
Mag pit cuttings composite 2   HPC - 109    155 to 165
     
Mag pit cuttings composite 2   HPC - 129    190 to 200
     
Map pit cutting composite 3   HPC 129    455 to 465, 490 to 495, 570 to 580, 585 to 590
     
Mag pit cutting composite 4   HPC - 109    255 to 275, 280 to 290, 310 to 320
     
Mag pit cutting composite 5   HPC -129    210 to 220, 255 to 270, 275 to 285, 470 to 485
     
Mag pit cutting composite 5   HPC 109    200 to 205, 215 to 235
     
Mag pit cutting composite 6   HPC 109    275 to 280, 290 to 295, 300 to 305, 320 to 340, 345 to 350, 360 to 370

The Mag Pit master composite was created using material from Mag pit cutting composite 1 to Mag pit cutting composite 6. The material was blended using 11.6% of Mag Pit 1 sample, 21.7% of Mag Pit 2, 14.5% of Mag Pit 3, 24.6% of Mag Pit 4, and 27.6% of Mag Pit 5.

Preg-robbing tests were completed on some of the samples to determine the preg-robbing characteristics of the Mag Pit and CX Pit samples. In these tests, barren solutions were “spiked” with a diluted gold solution to create a 1 ppm gold solution. The donor solution was a barren solution from column leaching of oxide materials. A standard cyanide leach bottle roll test with a 1-kilogram (kg) rock charge was completed on the slurry with the spiked solution. The pregnant leach solution was then assayed for gold at 2, 6, 24, 48, 72, and 96 hours. The percentage of gold that was preg-robbed was determined by the following formula:

 

 

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Original gold solution concentration - Final gold  solution concentration

 
  Original gold solution concentration (%)  

Preg-robbing test work completed in this manner does not account for any gold dissolution from the 1-kg rock charge; it only considers gold loss from the spike solution. Typically, tests are conducted first (base case) without spiking to determine how much gold will be dissolved, if any. The test is then repeated with the spiked solution, and the base test gold dissolution is included within the preg-rob calculation.

Table 10-2 shows the percentages of preg-robbed gold. Negative values indicate where the final gold concentration was higher than the original gold concentration. Cyanide is destroyed in the 1 ppm gold solution before conducting the test. Some samples have negative preg-robbed values. These samples leach gold during the test. This is most likely due to the presence of Weak Acid Dissociable cyanide, which was not destroyed and is yet available for continued leaching of the material during the bottle roll test.

Table 10-2: Preg-Robbing Test Results from the 1999 Test Work Program

 

Sample    Sample Type    Feed Size    Preg-robbed Gold  (%)

Mag Pit I

  

Bulk Ore

  

P80 3⁄4-inch (19 millimeter [mm])

   87.9

Mag Pit II

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   76.6

Mag Pit III

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   -8.6

Mag Pit IV

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   76.0

Mag Pit V

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   -8.6

Mag Pit VI

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   -5.0

CX-2

  

Bulk Ore

  

P80 3/4-inch (19 mm)

   -31.0

Mag Pit cuttings composite 1

  

Drill core

  

10 Mesh (1.7 mm)

   14.1

Mag Pit cuttings composite 2

  

Drill core

  

10 Mesh (1.7 mm)

   14.9

Mag Pit cuttings composite 3

  

Drill core

  

10 Mesh (1.7 mm)

   15.5

Mag Pit cuttings composite 4

  

Drill core

  

10 Mesh (1.7 mm)

   19.1

Mag Pit cuttings composite 5

  

Drill core

  

10 Mesh (1.7 mm)

   60.5

Mag Pit cuttings composite 6

  

Drill core

  

10 Mesh (1.7 mm)

   47.4

Many samples had relatively high preg-robbing values (greater than 50%), demonstrating that preg-robbing is a potential significant issue for some Pinson material (Table 10-2).

 

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Plotting of preg-robbed percentage vs. the Total Organic Carbon (TOC) of the feed material shows a trend but the correlation is not statistically significant. However, plotting gold solubility vs. TOC shows a statistically significant trend. Solubility (%) is calculated as follows:

 

Cyanide Leach Gold Grade  

x100

Fire Assay Gold Grade 

Figure 10-1: Gold Cyanide Solubility and Total Organic Carbon Influence

 

LOGO

The TOC appears to have a strong influence on cyanide-soluble gold extraction. The organic carbon is capable of adsorbing gold during cyanide leaching, reducing the final gold recovery. This can have a major implication in the selection of the leaching and recovery process. Processes like CIL have activated carbon present during leaching that helps reduce the impact of preg-robbing carbon, while processes like heap leaching do not.

Predictors of recovery as they relate to Pinson and the currently available metallurgical database is discussed within the geometallurgical section.

Cyanide leach bottle roll tests were completed on the Mag Pit bulk material samples using caustic soda (NaOH) to adjust pH, rather than hydrated lime. The test work report postulated that NaOH passivates the preg-robbing (carbonaceous) surfaces by occupying active carbon sites with hydroxide (OH–) ions so the gold cyanide complex Au(CN)2– ions do not absorb onto the active carbon sites. Reducing the amount of Au(CN)2– ions that are absorbed onto the carbon sites would improve gold recovery. For each sample, two tests were conducted: at pH 10.5 and 12 (using NaOH to adjust pH). The results of these tests are shown in Table 10-3.

 

 

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The test data shows an increase in recovery with higher pH as well as a reduction in cyanide consumption. The reduced cyanide consumption is most likely a direct result of the increase in pH.

Generally, the largest recovery increases between the pH 12.0 tests and the pH 10.5 tests were associated with samples that showed the highest preg-robbed gold. This would indicate the OH– ions are passivating sites on the available pre-robbing carbon and preventing the uptake of gold. No tests were completed using lime. Some gold processing facilities have demonstrated success using a lime-boil for similar issues. However, CIL processing is likely the best alternative for material of this nature.

There were no baseline tests using lime on these samples, so a proper comparison between lime and NaOH cannot be completed.

Table 10-3: NaOH Bottle Roll Tests from 1999 Test Work Program

Sample   

Preg-
robbing

Factor (%) 

   NaOH Tests
   pH 10.5 Tests    pH 12.0 Tests    Difference in Gold
Recovery between pH
12.0  Tests and pH 10.5
Tests
  

Gold

Recovery (%) 

  

Cyanide

Consumption
(pound [lb]/short

ton)

  

Gold Recovery

(%)

   Cyanide Consumption
(lb/short ton)

Mag Pit I

   87.9    8.5    1.1    32.0    1.0    23.5

Mag Pit II

   76.6    13.2    3.0    24.7    1.5    11.5

Mag Pit III

   -8.6    74.2    1.4    83.6    0.7    9.4

Mag Pit IV

   76.0    26.4    1.6    40.8    0.5    14.4

Mag Pit V

   -8.6    50.0    1.8    53.8    0.4    3.8

Mag Pit VI

   -5.0    62.2    1.1    65.0    0.3    2.8

CIL tests were completed on the CX-2 bulk material and Mag Pit cuttings samples. The objective of these tests was to test the applicability of CIL processes to Pinson open pit material. The CIL process is used to help overcome the impact of the organic pre-robbing carbon that naturally occurs in some of the Pinson material. The conditions of these tests were:

 

   

Tests were conducted in agitated bench-scale beakers

 

   

Samples were ground to a P80 of 200 mesh (75 microns [µm])

 

   

72 hours residence time

 

   

Kinetic samples taken at 6 hours, 12 hours, 24 hours, 36 hours, and 48 hours

 

   

Hydrated lime was added to raise the pH to 10.5

 

   

A sodium cyanide (NaCN) concentration of 1 gram per liter (g/L)

 

   

Pulp density of 40% solids weight for weight (w/w)

 

   

Activated carbon was added to absorb the gold in solution onto the carbon.

 

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The results of these tests are shown in Table 10-4.

Table 10-4: CIL Tests from 1999 Test Work Program

 

Sample   

CIL Gold

Recovery

(%)

  

Cyanide Consumption

(lb/short ton)

   TOC (%)    S= (%)    Solubility (%)

CX-2 Bulk ore

   88.2    2.4    0.35    <0.01    91.9

Mag Pit cuttings I

   94.4    3.3    4.55    0.05    19.5

Mag Pit cuttings II

   75.3    2.3    4.25    0.12    19.6

Mag Pit cuttings III

   59.7    3.0    0.15    <0.01    84.5

Mag Pit cuttings IV

   82.9    4.8    4.00    0.83    22.8

Mag Pit cuttings V

   55.0    3.0    0.40    <0.01    50.0

Mag Pit cuttings VI

   87.5    3.9    0.45    <0.01    61.8

With the exception of Mag Pit 3 and Mag Pit 5 samples, these tests generally achieved high gold recoveries. However, these lower recoveries appear to be an anomaly when compared to column tests on the same material. When comparing the solubility value to the CIL gold recovery, it can be seen there is not a strong correlation between the two values. There is also no relationship between the CIL recovery and the sulfide sulfur grade (S=). It appears that the CIL process can overcome the presence of organic carbon in this material. There is a relationship between gold feed grade and CIL recovery as shown in Figure 10-2. Based on this test work, a CIL process would be applicable to Pinson material.

Figure 10-2: CIL Recovery and Head Grade Influence

 

LOGO

 

 

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Column leach tests were conducted on some of the samples from the 1999 program. Sixteen columns were completed representing 11 different samples. The samples were from the Mag Pit, bulk samples, core, and bulk samples from the CX Pit. The conditions of these tests were:

 

   

Leach time of between 50 and 90 days. If the kinetic leach curve demonstrated that a test was approaching terminal gold recovery, the test was stopped.

 

   

Varying crush sizes.

 

   

Hydrated lime was added to agglomerate the material in the column.

 

   

Lime was added to most tests to raise the pH to 10.5.

 

   

NaOH was added to the Mag Pit I and Mag Pit II samples, given the successful NaOH bottle roll tests on these samples. The pH was initially 10.5 but was increased to 12.0 later in the test to ascertain the impact on leaching.

 

   

NaCN was added at an initial concentration of 1 g/L and was pumped into the columns at a rate of 0.005 gpm per square foot (/ft2) of cross-sectional area.

 

   

Three tests with varying particle sizes were conducted on the CX Pit sample to ascertain the impact of crush size on gold recovery.

Three tests were conducted on the Mag Pit master composite where pH and alkali were varied:

 

   

Test 1: pH 10.5 (lime)

 

   

Test 2: pH 11.8 (lime.

 

   

Test 3: pH 11.8 (NaOH)

Table 10-5 shows the results from the 1999 column leach tests.

Table 10-5: Column Leach Tests from 1999 Test Work Program

 

Sample    Sample Type     Feed Size
(inches)
   Gold Recovery 
(%)
   Cyanide
Consumption 
(lb/short ton) 
   Lime
Consumption (lb/ 
short ton) 

Mag Pit I

   Bulk ore    -4    18.8    9.9    5.2

Mag Pit II

   Bulk ore    -4    35.3    9.0    10.2

Mag Pit III

   Bulk ore    -4    93.1    4.6    5.2

Mag Pit IV

   Bulk ore    -4    49.5    5.3    12.0

Mag Pit V

   Bulk ore    -4    51.7    3.9    2.5

Mag Pit VI

   Bulk ore    -4    60.7    3.7    4.0

Mag Pit 2

   Drill core    -1    69.0    4.0    11.0

Mag Pit 3

   Drill core    -1    62.0    1.6    9.6

Mag Pit 4

   Drill core    -1    47.9    1.5    8.1

Mag Pit 5

   Drill core    -1    61.7    2.1    10.0

Mag Pit master (pH 10.5, Lime)

   Drill core    -1    65.0    6.3    8.4

Mag Pit master (pH 11.8, Lime)

   Drill core    -1    70.7    4.2    19.3

 

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Sample    Sample Type     Feed Size
(inches)
   Gold Recovery 
(%)
   Cyanide
Consumption 
(lb/short ton) 
   Lime
Consumption (lb/ 
short ton) 

Mag Pit master (pH 11.8, NaOH)

   Drill core    -1    69.0    3.5    n/a

CX Pit, CX-2

   Bulk ore    -6    77.7    5.1    3.0

CX Pit, CX-2

   Bulk ore    P80 3    81.7    4.8    3.0

CX Pit, CX-2

   Bulk ore    P80 3⁄4    82.2    5.4    3.0

This test work program had the following findings:

 

   

There was a wide range of gold recoveries, varying from 19% to 93%

 

   

Gold recovery closely followed the gold cyanide solubility percentage

 

   

Sulfide grade was generally too low to show any impact on gold recovery

 

   

The bulk sample tested from the CX Pit showed near 5% improvement in recovery when crushing from 6 inches to 3⁄4 inch. There was very little difference in recovery between the 3 inch and 3⁄4 inch size, less than 1%.

The tests on the Mag Pit master composite sample had the following conclusions:

 

   

Increasing pH demonstrated an increase in gold recovery

 

   

The NaOH and the lime test (pH 11.8) had a slightly less than 2% recovery difference.

Figure 10-3: Column Recovery and Solubility Influence

 

LOGO

Figure 10-3 shows that the gold cyanide solubility influences the column gold recovery (all particle sizes and parameters shown). This is expected as the degree of gold solubility is highly dependent on the presence of organic carbon as shown in Table 10-4. Unmitigated organic carbon will

 

 

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adsorb gold from solution negatively impacting the solubility percentage and the gold recovery from a column test.

McClelland reported on an additional column leach test program in June 1999. The materials were all from the CX Pit area. Coordinates within the pit were not provided, although a bench elevation within the pit was provided for each sample. Three bulk samples, 4840-Silty, 4840-Clay Ore and 4680–Typical CX-4, were tested at a 100% passing 1-inch crush size in percolation columns. The materials leach very well and rapidly. The majority of the leaching was complete in 30 days. The samples were run from 78 to 95 days, including leaching and washing.

Table 10-6 Column Leach Tests from June 1999 Test Work Program

 

Sample    Sample
Type
  

Column
Feed

Size

   Gold
Recovery
(%)
   Cyanide
Consumption
(lb/short ton
   Lime
Consumption
(lb/short
ton)
  

Final

pH

  

Calculated
Head

(opt)

4840 Silty   

Bulk

Material

  

P100

1-inch

   91.9    3.16    3    10.1    0.111

4840

Clay Ore

  

Bulk

Material

  

P100

1-inch

   95.6    2.36    3    9.4    0.110

4680

Typical

CX-4

  

Bulk

Material

  

P100

1-inch

   94.2    2.73    3    10.9    0.119

The grade of these samples was high, above what the typical heap leach feed would be. McClelland noted in the final report that the Silty and Clay Ore materials demonstrated moderate to severe percolation issues. Agglomeration was recommended for commercial heap leaching operations. During the column testing, the pH levels were lower than desired. Additional lime, above 3 pounds per short ton, will be necessary to keep cyanide consumption to a minimum. For typical column leaching, the cyanide consumption is high. The CX materials are very amenable to heap leaching.

 

10.2.2

McClelland Laboratories Inc 2013 & 2014

McClelland Laboratories completed a metallurgical test work program on Mag Pit samples in 2013 and 2014 on behalf of Atna Resources Ltd. (Atna).

The 2013 program used 32 drill core composite samples. The samples were well identified by the drill hole and down-hole depth. The 32 samples were then subjected to detailed head analysis, ICP scan, carbon and sulfur speciation analysis, and preg-robbing tests.

Bottle roll tests were completed in pairs, with one being 80% passing 1⁄4-inch and the second as a 150 Mesh sample (P100 100 um). The program intended to complete an evaluation of the impact

 

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of feed size, potential for heap leaching, and especially testing for preg-robbing problems associated with the Pinson materials.

A summary of the drillholes and intervals used to make up the samples for this program are shown in Table 10-7. This is sourced from the appendix within the McClelland 2013 report (McClelland, 2013).

Table 10-7 Sample Composite List from 2013 Test Work Program

 

Drillhole    Sample    Interval
   To (ft)    From (ft)
Magmet-001    Magmet-001-01    0.0    27.5
   Magmet-001-02    27.5    99.5
   Magmet-001-03    99.5    157.0
   Magmet-001-04    170.5    228.5
   Magmet-001-05    228.5    251.5
   Magmet-001-06    251.5    302.5
   Magmet-001-07    302.5    364.5
   Magmet-001-08    364.5    415.5
Magmet-002    Magmet-002-01    211.5    254.5
   Magmet-002-02    254.5    292.0
   Magmet-002-03    292.0    337.0
   Magmet-002-04    337.0    397.0
   Magmet-002-05    397.0    446.0
   Magmet-002-06    450.0    497.0
   Magmet-002-07    497.0    567.5
   Magmet-002-08    567.5    599.8
Magmet-003    Magmet-003-01    179.0    225.0
   Magmet-003-02    225.0    283.0
   Magmet-003-03    283.0    304.5
   Magmet-003-04    304.5    361.5
   Magmet-003-05    361.5    409.0
   Magmet-003-06    409.0    459.5
   Magmet-003-07    459.5    514.0
Magmet-004    Magmet-004-01    125.0    148.0
   Magmet-004-02    148.0    220.5
   Magmet-004-03    220.5    270.0
   Magmet-004-04    270.0    330.0
   Magmet-004-05    330.0    372.0
   Magmet-004-06    372.0    415.0
   Magmet-004-07    415.0    439.0

 

 

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Drillhole    Sample    Interval
   To (ft)    From (ft)
     Magmet-004-08    439.0    496.0
     Magmet-004-09    496.0    551.5

Cyanide leach bottle roll tests were conducted on Mag Pit samples. The objective of these tests was to identify the impact of particle size on gold recovery. The conditions for the bottle roll tests were:

 

   

A pulp density of 40% solids w/w

 

   

pH was maintained between 10.8 and 11.2 using lime

 

   

NaCN was added at a concentration of 1 g/L

 

   

The bottle rolls were sampled at 6, 12, 24, 36, and 48 hours

 

   

The tests were terminated at 48 hours and final analysis was completed

Preg-rob factors were measured for each sample. The preg-robbing factor methodology was not presented within the laboratory report.

A summary of the results from the bottle roll tests is shown in Table 10-11.

Table 10-8 Bottle Roll Tests Results from 2013 Test Work Program

 

Sample   

Au head

grade
(oz/ton)

   S= (%)    TOC (%)    Solubility
(%)
   Preg-Rob
Factor
   BRT Gold recovery (%)
   P80 1⁄4-
inch
   150
Mesh
   Difference 150#  and
P80 1⁄4-inch
Magmet-001-01    0.031    0.03    0.24    84.5    0    77.4    81.3    3.9
Magmet-001-02    0.089    0.04    3.49    20.7    93    50.0    47.8    -2.2
Magmet-001-03    0.032    0.01    0.46    77.3    0    72.0    82.1    10.1
Magmet-001-04    0.030    0.02    1.78    66.0    36    57.1    69.2    12.1
Magmet-001-05    0.057    0.20    3.47    1.8    97    17.4    6.0    -11.4
Magmet-001-06    0.050    0.53    3.67    12.4    92    24.4    27.8    3.4
Magmet-001-07    0.058    1.24    4.15    1.7    100    4.3    5.8    1.5
Magmet-001-08    0.018    1.14    3.81    16.1    87    20.0    25.0    5.0
Magmet-002-02    0.030    0.94    2.75    6.6    92    8.3    9.5    1.2
Magmet-002-03    0.005    0.57    1.73    N/A    84    50.0    40.0    -10.0
Magmet-002-04    0.128    0.83    2.48    50.8    76    78.4    86.4    8.0
Magmet-002-05    0.114    0.69    4.57    34.7    78    50.0    70.7    20.7
Magmet-002-06    0.043    1.57    5.07    7.3    95    15.4    17.1    1.7
Magmet-003-04    0.014    1.12    3.39    N/A    0    53.8    71.4    17.6
Magmet-003-05    0.059    2.55    0.40    20.2    10    76.8    87.0    10.2
Magmet-003-06    0.022    1.80    0.18    51.4    11    75.0    80.0    5.0
Magmet-003-07    0.044    1.05    3.34    67.9    48    48.7    81.0    32.3

 

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Sample   

Au head

grade
(oz/ton)

   S= (%)    TOC (%)    Solubility
(%)
   Preg-Rob
Factor
   BRT Gold recovery (%)
   P80 1⁄4-
inch
   150
Mesh
   Difference 150#  and
P80 1⁄4-inch
Magmet-004-04    0.018    0.59    3.58    57.7    42    42.9    55.6    12.7
Magmet-004-05    0.006    1.15    3.66    35.8    15    50.0    71.4    21.4
Magmet-004-06    0.018    1.64    2.94    32.0    19    47.1    75.0    27.9
Magmet-004-08    0.023    0.98    1.90    41.8    10    42.9    54.5    11.6

Some conclusions can be drawn from the results of the tests. The 1⁄4-inch material had a recovery range from 4.3% to 78.4%, with an average of 45.8%. Gold recoveries for the 150 Mesh material ranged from 5.8% to 87.0%, averaging 52.9%. The recovery was not very sensitive to feed size considering the size difference.

Figure 10-4: Bottle Roll Recovery and Solubility Influence

 

LOGO

McClelland stated the refractory nature of the Pinson material is poorly understood. Preg-robbing (PR factor) assay results indicate 11 of the 23 samples would be expected to exhibit moderate to severe preg-robbing character, PR factor >50%. There is a strong correlation between the calculated solubility and the gold extraction from the bottle roll tests, and a related correlation between the TOC grade and the solubility, as would be expected when activated carbon is not employed in the leaching (CIL).

The sulfide grade did not show a strong correlation to the BRT gold extraction but in most cases the sulfide grade was low.

 

 

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Cyanide leach bottle roll tests were completed on a select set of the Mag Pit samples to determine the impact of using NaOH rather than lime to abate the preg-robbing impact. For each sample type, tests were completed at pH 10.5 and 12.0. The conditions of these tests were:

 

   

NaOH was added to raise the pH

 

   

Samples were crushed to a P80 of 1⁄4-inch

 

   

Residence time of 48 hours

Table 10-12 shows that increasing the pH using NaOH increased gold recovery for all tests.

Table 10-9 Results from Bottle Roll Tests Using NaOH from 2013 Testwork Program

 

Sample   

Au Head Grade

(opt)

   Gold Recovery (%)
   pH 10.5 Test    pH 12.0 Test    Difference Between pH 12.0 Test and
pH 10.5 Test

Magmet-001-02

   0.089    53.2    60.8    7.6

Magmet-001-05

   0.057    14.0    43.1    29.1

Magmet-002-02

   0.030    12.5    20.8    8.3

Magmet-002-06

   0.043    20.5    35.1    14.6

Magmet-003-05

   0.059    75.0    78.2    3.2

Column leach tests were conducted on composites of the Mag Pit samples. The samples used to make up these composites are shown in Table 10-7. The rationale behind the compositing methodology was not clear from the report.

Column leach tests and bottle roll tests were conducted on each composite to determine gold recovery kinetics and reagent addition rates. The conditions of the bottle roll tests were:

 

   

48 hours residence time

 

   

P80 of 1⁄4-inch

 

   

Hydrated lime was added to raise the pH to 12.0

 

   

Residence time for the column leach tests varied between 72 and 76 days

Generally, the tests were conducted on samples that had been crushed to 2 inches. The Mag Column 2 sample had an additional test on a sample crushed to 1⁄2-inch to ascertain the impact of size on gold recovery. The 1⁄2-inch column recovered essentially the same as the 2-inch column of the same material and grade.

Lime was added to agglomerate the single 1⁄2-inch column. The lime was cured in the column for 72 hours prior to applying the leach solution. Agglomeration is required for any material that might exhibit percolation issues. Materials that exhibit mild percolation issues in the lab may exhibit

 

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more severe percolation issues during commercial operations. Larger and taller lab column tests can help predict the potential for percolation issues.

Lime additions were based on the bottle roll test lime requirements. This is typical for column tests and one of the intents of the bottle roll tests.

The results of the McClelland column tests are shown in Table 10-13.

Table 10-10 Sample Composition for Column Leach Tests from 2013 Test Work Program

 

Sample    % of
Composite
Mag Column 1
Magmet-001-01    14.3
Magmet-001-03    15.5
Magmet-001-04    33.2
Magmet-003-05    17.1
Magmet-004-04    19.9
Mag Column 2
Magmet-002-04    18.5
Magmet-002-05    8.8
Magmet-003-04    28.4
Magmet-003-06    14.5
Magmet-003-07    13.8
Magmet-004-06    13.2
Magmet-004-08    12.8
Mag Column 3
Magmet-001-02    26.2
Magmet-001-05    6.4
Magmet-001-06    10.1
Magmet-001-07    17.7
Magmet-001-08    13.9
Magmet-002-02    12.2
Magmet-002-06    13.5
Mag Column 4
Mag Column 1    13.8
Mag Column 2    43.1
Mag Column 3    43.1

 

 

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Table 10-11: Bottle Roll and Column Test Results from 2013 Test Work Program

 

Sample   Calc Au Grade
(opt)
  Test Type   Feed Size
(inches)
  Leach Time
(days)
  Gold
Recovery (%)
  Cyanide
Consumption
(lb/short ton)
  Lime
Consumption
(lb/short ton)
Mag Column 1    0.042   Bottle roll   1⁄4   2   52.4   0.6   14.6
  0.034   Column   2   73   61.8   2.5   14.6
Mag Column 2    0.053   Bottle roll   1⁄4   2   71.7   1.1   21.4
  0.051   Column   2   76   82.4   3.3   21.4
  0.050   Column   1⁄2   72   82.0   3.8   21.4
Mag Column 3    0.040   Bottle roll   1⁄4   2   32.5   0.7   16.1
  0.057   Column   2   72   50.9   2.6   16.2
Mag Column 4    0.043   Bottle roll   1⁄4   2   51.2   0.8   17.3
  0.049   Column   2   76   65.3   3.0   17.4

Table 10-11 shows that the gold recoveries in the column tests varied from 51% to 82%. The Mag Column 2 tests showed no benefit to gold recovery by crushing finer. The 1⁄2 inch and 2-inch columns had similar recovery for the materials tested. The gold solubility percentage correlated to the column and bottle roll gold extractions.

Figure 10-5: Column and Bottle Roll Recovery and Solubility Influence

 

LOGO

 

10.2.3

Dawson Metallurgical Program 2005 and 2006

Dawson Metallurgical Laboratories completed autoclave metallurgical test work programs on samples from underground (Ogee samples) on behalf of Atna (Dawson, 2005; Dawson, 2006a; Dawson, 2006b). The 2005 report reported on samples LR&RR and 33941, 33942, 34259. These samples were treated to determine applicability of the material for autoclave treatment. This initial program was followed by the 2006 test work and report, which appears to be the final report.

This test work program was completed on the following samples:

 

   

A composite from the Ogee underground deposit labelled “Right Rib and Left Rib”

 

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Composites from the RFZ labelled as:

 

   

LR&RR

 

   

RF_Met-1 (33941)

 

   

RF_Met-2 (33942)

 

   

RF_Met-4 (34259)

 

   

Composites from the CX Zone labelled as (drill footages identified):

 

   

APCX-204

 

   

APCX-211

 

   

APCX-219

 

   

APCX-226

 

   

Undefined samples:

 

   

AMW-002

 

   

Met1 and Met 2 from the 2005 program (re-tested)

 

   

Met1 and Met 2 from the 2006 program (re-tested)

The objective of these programs was to determine whether the underground materials identified as refractory could be treated using autoclave pre-treatment. Atna was considering contracting with a third party for autoclave treatment and downstream processing of the underground material.

The scope of these test work programs included:

 

   

Head assays including gold, sulfur speciation, and carbon speciation

 

   

Baseline cyanide leach shake-out tests on ground feed samples

 

   

Pressure oxidation test work: grinding of samples to either a P80 of 75 µm or 45 µm.

 

   

Acidulation, where sulfuric acid was added to achieve a pH of 1.8 to 2.0 and processed for one hour. The purpose of this stage was to digest carbonate minerals ahead of the autoclave stage, which is a standard methodology for whole material autoclave treatment in Nevada.

The acid leach residue was then processed in an autoclave with the following conditions:

 

   

Temperature of 225 °C

 

   

Residence time of one hour

 

   

Pulp density of 35% solids w/w

 

   

Oxygen overpressure of 460 pounds per square inch (psi)

 

 

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Lime was added to the autoclave residue to raise the pH to a range of 10.0 to 10.5 prior to leaching with cyanide. The autoclave residue underwent a cyanide leaching test simulating CIL (addition of carbon) processing to determine gold recovery.

The cyanide leach gold recoveries from the baseline tests and the autoclave tests are shown in Table 10-11. The refractory material responded well to the autoclave pre-treatment, and, for those tests with a baseline cyanide leach (no pre-treatment) compared to the autoclave pre-treatment, there was an increase from an average recovery of 52% to 92%.

An attempt was made to fit an equation to demonstrate the relationship between sulfide sulfur analysis and baseline gold recovery. The Coefficient of Determination (R2) was very low for the equation and for this reason it is not presented. There was an inverse relationship between sulfide sulfur content and cyanide solubility (no autoclave treatment). The solubility in this case showed a reasonable correlation to the sulfide grade.

Figure 10-6: Solubility and Sulfide Influence – Ogee Samples

 

LOGO

There is a range of recoveries from 11% to 86% for those samples with measured sulfide sulfur. The previous test work completed by McClelland labs noted that the sulfide sulfur content did not correlate well with cyanide solubility. Other potential issues impact cyanide solubility, such as the presence of organic carbon, and not all of the gold is directly associated with pyrite, as demonstrated by the wide range of baseline cyanide solubility tests. Table 10-12 Samples MET1 and MET2 did not undergo baseline cyanide leach tests. These samples were from prior autoclave tests and were submitted for re-testing.

 

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Table 10-12: Autoclave Pre-treatment Tests from Dawson Test Work Program

 

Sample    Year of Test
work Program
   Grind P80
(µm)
   Gold Head
Assay (opt)
   Total Sulfide
Sulfur (%)
   Total Carbon
(CO2) Head
Assay (%)
   Cyanide Leach Gold
Recovery (%)
   Baseline
Tests
   Tests on
Autoclave
Residue
Ogee Samples
Ogee (Right Rib + Left Rib)    2005    75    0.40    0.0    0.82    86    93
RF Zone Samples
RF_Met-1 (33941)    2005    75    0.24    1.21    2.27    52    93
RF_Met-2 (33942)    2005    75    0.43    2.61    1.76    61    95
RF_Met-4 (34259)    2005    75    0.43    2.32    2.43    11    89
CX Zone Samples
APCX-204    2006    75    0.27    0.00    5.34    94    N/A
APCX-211    2006    75    0.33    0.00    4.29    85    N/A
APCX-219    2006    75    0.33    0.84    0.77    60    91
APCX-226    2006    45    0.56    1.53    2.70    42    94
Undefined Samples
AMW-002    2006    75    0.33    0.03    0.35    77    N/A
MET 1    2005    75    0.51    1.21    2.27    N/A    93
MET 2    2005    75    0.32    2.61    1.76    N/A    95

 

10.2.4

 FLS Metallurgical Program 2023

In early 2023 FLS was contracted to undertake a series of tests related to the potential underground material. The objective of this report was to provide the following: ore characterization, mineralogical testing, comminution testing, acid-alkaline batch pressure oxidation (POX), followed by batch cyanidations, and preg robbing tests. This program also included a continuous POX test followed by batch neutralization testing and subsequent cyanidation, and cyanide detoxification testing. A batch of POX-CIL tests for 3 different composite blends was also undertaken.

 

10.2.4.1

  Samples Characterizations

The samples received were assayed with the results shown in Table 10-13.

 

 

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Table 10-13: Underground Samples Head Assays from FLS Program

 

Sample
Number
   Description    ID    Au opt    Ag opt    S= %    Corg %
1   

OG Zone Upper (Sulfide)

   OGU    0.424    0.135    1.96    0.44
2   

OG Zone Lower (Sulfide)

   OGL    0.830    0.274    2.98    0.12
    

OG Zone Oxide

   OGOX    0.364    0.337    0.43    0.34
3   

OG Zone High Grade (Sulphide) Variability Sample

   OGHG    1.10    0.216    3.89    0.10
4   

OG Zone Low Grade (Sulphide) Variability Sample

   OGLG    0.246    0.65    2.73    0.36
    

OG Zone Comminution Sample No. 1

   OGCOM1    0.372    4.47    2.33    0.33
    

OG Zone Comminution Sample No. 2

   OGCOM2    1.15    0.285    0.04    0.03
    

OG Zone Comminution Sample No. 3

   OGCOM3    0.303    0.207    0.15    0.02
    

OG Zone Comminution Sample No. 4

   OGCOM4    0.290    0.123    0.00    0.19
    

OG Zone Comminution Sample No. 5

   OGCOM5    0.765    0.108    0.00    0.03
    

OG Zone Comminution Sample No. 6

   OGCOM6    0.376    0.056    1.41    1.12
5   

Otto Zone Upper (Sulfide)

   OTU    0.487    0.117    1.14    0.18
6   

Otto Zone Lower (Sulfide)

   OTL    0.437    0.319    1.67    0.38
7   

Otto Zone High Grade (Sulfide) Variability Sample

   OTHG    0.595    0.376    0.79    0.19
8   

Otto Zone Low Grade (Sulfide) Variability Sample

   OTLG    0.182    0.032    1.48    0.22
    

Otto Zone Comminution Sample No. 1

   OTCOM1    0.245    0.00    3.01    0.11
    

Otto Zone Comminution Sample No. 2

   OTCOM2    0.406    0.029    1.81    0.47
    

Otto Zone Comminution Sample No. 3

   OTCOM3    0.277    0.031    2.84    0.27
9   

Adams Peak Zone Upper

   APU    0.223    0.027    5.84    0.33
10   

Adams Peak Zone Lower

   APL    0.133    0.143    4.29    0.25
11   

Adams Peak Zone High Grade Variability Sample

   APHG    0.596    0.029    2.86    0.20
12   

Adams Peak Zone Low Grade Variability Sample

   APLG    0.188    0.240    3.20    0.13
    

Adams Peak Zone Comminution Sample No. 1

   APCOM1    0.322    0.091    3.12    0.37
    

Adams Peak Zone Comminution Sample No. 2

   APCOM2    0.343    0.439    3.85    0.17
13   

Otto/Adams Peak Zone Composite

   OAPC    0.331    0.107    2.25    0.24
14   

Deep Range Front Zone Variability Sample

   DRFV    0.184    0.081    1.73    0.26
15   

Range Front Zone Variability Sample

   RFV    0.213    0.083    2.33    0.17
16   

South Pacific Zone Variability Sample

   SPZV    0.588    0.036    2.79    0.75

Mineralogy (XRD) and swelling clay analysis was conducted on the samples. The gangue mineralogy consists of quartz, k-feldspar, muscovite, clays (kaolinite and swelling clay) and calcite. Pyrite and marcasite were present in all samples as the primary sulfides. Although most samples had minor amounts of swelling clay, two samples showed higher percentages that may require attention during POX treatment (APL and APLG).

 

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10.2.4.2  Comminution Testing

Comminution testing (Ball Mill Bond Work Index, BBWi) was conducted on selected samples.

Comminution testing was limited to Bond ball mill work index testing in consideration of the fact that no comminution design is required for this study. Production will be initially toll milled at a Nevada Gold Mines process facility or eventually through the refurbished Lone Tree process facility. Bond ball mill work index tests were conducted using a 140 mesh (106) µm closing screen size, aiming to achieve a grind size k80 of 200 mesh (75 µm).

 

   

The overall average is 16.4 (imperial), classified as hard while the 75th percentile value (normal design value) is 19.0 (imperial), classified as very hard.

 

   

For the OG Zone samples the average value is 18.9 (imperial), classified as very hard while the 75th percentile value is 20.3 (imperial), classified as very hard. The OG Zone oxide samples have similar hardness values compared to the sulphide samples.

 

   

For the Otto Zone samples the average value is 16.2 (imperial), classified as hard while the 75th percentile value is 17.4 (imperial), classified as hard.

 

   

For the Adams Peak Zone samples the average value is 12.8 (imperial), classified as medium while the 75th percentile value is 13.2 (imperial), also classified as medium.

 

   

The Deep Range Front and Range Front Zone samples are classified as medium hardness while the South Pacific Zone sample is classified as very hard.

 

10.2.4.3

Cyanide Shake Tests and Preg-Robbing Testing

Analytical Direct Cyanide Leach Shake Tests were conducted to provide a baseline recovery on representative samples from each of the 17 selected samples. Tests were done on pulverized samples and conducted in centrifuge tubes. The tubes were placed on a shaker for 60 minutes before centrifuging and before the collection of the pregnant solution for gold analysis by AA. Sample pH was adjusted using lime slurry to a pH of 10.5.

Preg-robbing leach test samples were pulverized and placed on shakers for 60 minutes after being spiked with a stock solution containing a known amount of gold. This test was performed to measure the preg-rob index (PRI) of each sample by comparing the amount of gold adsorbed onto the solids to the leached gold.

Cyanide shake gold extractions ranged from 4.8% to 54.7%, with the OG oxide sample having the highest extraction. The overall average gold extraction was 26.2%, reflecting the refractory sulphide nature of the majority of the samples. The average baseline CIL gold recovery was 31%. All samples demonstrated some degree of preg robbing. Historical testing showed that Mag Pit

 

 

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samples also demonstrated preg robbing, so the current results are not unexpected. The average preg-robbing index was 17.9%, ranging from 4.4% to 54.1% (OAPC sample). If the shake tests are used as a cyanide leach test proxy, CIL improved recoveries by an average of 4.1% (Otto Zone upper and lower samples were well above this average). This is not a significant increase showing CIL does counteract the preg robbing to a small extent but doesn’t overcome the refractory nature of the samples. The only oxide sample (OGOX from the OG Zone) showed the highest extraction at 54.7% Au but only 31% Au CIL recovery. This is well below expected recovery for an oxide sample. OGOX showed low preg robbing at 7.23%, suggesting the sample has some refractory qualities.

In some deposits, organic carbon can be corelated to pre-robbing index No relationship is apparent between organic carbon content and preg-robbing or CIL gold recoveries.

 

10.2.4.4

Initial BTAC Tests and CIL Testing

All metallurgical samples (except for OGOX) were run through a series of BTAC tests under a series of tests with six different conditions shown in Table 10-14 . Continuous POX testing was employed to confirm the batch testing results. Test conditions B, C, D and F replicate Lone Tree autoclave operating conditions. Trona addition can counteract the problems caused by swelling clays in POx and was evaluated in test conditions C and D. Test conditions A and E include acid conditions with pre-acidulation at two temperatures. Conditions A and E were acidulated for an hour using 98% concentrated sulfuric acid. For Condition A, sulphuric acid was added in a stoichiometric ratio to enable carbonate destruction. Condition E included sulphuric acid addition to target a CO3/S2- weight ratio of 1.

Six different Batch Autoclave Tests (BTAC) conditions were run on each of the selected zones as shown. As a result of the sulfide oxidation and gold recovery determined during the batch testing for the material sample OAPC, the three conditions selected for the continuous POX run were conditions A, B, and E. Condition A had the highest average gold recovery of 89% with 96% sulfide oxidation, condition B with an average of 69% gold recovery and 60% sulfide oxidation—expected to improve at a larger scale and thus chosen over condition F, and finally condition E with the second best average gold recovery of 79% and 72% sulfide oxidation.

Table 10-14: Underground Samples Batch Autoclave Conditions from FLS Program

POX Condition    A    B    C    D    E    F

Acidulation

   Yes    No    No    No    Yes    No

POX

   Acid    Alkaline    Alkaline    Alkaline    Acid    Alkaline

Trona Dosage (lb./ton)

   None    None    20    10    None    None

Temperature (oF)

   437    390    390    390    390    390

O2 Overpressure (psig)

   100    100    100    100    100    100

Target Gauge Pressure (psig)

   455    305    305    305    305    305

Pulp Density (% solids)

   30    30    30    30    30    30

Particle Size, k80 (mesh)

   200    200    200    200    200    270

Retention Time (min)

   60    45    45    45    45    45

 

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Carbon-in-Leach Bottle Roll Tests were conducted each sample and on the discharge from each BTAC test. These tests aimed to evaluate the effects of the different batch POX conditions on gold and silver recoveries. The conditions for these CIL tests were the same as the baseline CIL tests.

CIL Operating Conditions

 

   

Temperature (°F): 70

   

Pulp Density (% solids): 35

 

   

Carbon (lb./gal.): 0.17 (20 g/L)

 

   

pH Control: 10.5-11

 

   

Initial Cyanide Addition (lb./ton NaCN): 5.0

 

   

Residence Time (hours): 24

Averaged results by zone are shown in Table 10 -15.

Table 10-15: Underground Samples Baseline and Batch Pressure Oxidation CIL Results from FLS Program

BTAC
 Conditions 
   Parameter    OG Zone    Otto Zone    Adam’s Peak   

Otto/

Adam’s Peak

   Deep Range
Front
   Range Front     South Pacific  
                 
N/A   

Baseline Recovery

(% Au)

   41.50    46.75    9.25    24.00    10.00    34.00   43.00
                 
A    S Oxidation (%)    86.26    99.49    98.87    99.33    95.87    96.62   99.48
   Recovery (% Au)    84.12    88.77    90.65    83.18    89.05    94.96   95.65
  

NaCN Consumption

(lb./ton)

   1.32    1.5    1.32    1.00    1.22    0.62   1.60
                 
B    S Oxidation (%)    71.11    56.10    57.71    44.00    32.95    42.06   98.68
   Recovery (% Au)    85.17    67.12    61.33    56.88    42.79    59.32   94.19
  

NaCN Consumption

(lb./ton)

   2.82    2.14    1.46    3.62    3.24    2.62   3.64
                 
C    S Oxidation (%)    66.56    38.66    32.35    38.22    35.26    38.63   43.01
   Recovery (% Au)    75.75    66.71    45.43    52.17    43.33    57.64   79.09
  

NaCN Consumption

(lb./ton)

   2.22    2.14    3.20    3.58    3.42    3.82   2.34
                 
D    S Oxidation (%)    60.60    37.91    42.81    35.56    31.21    34.76   80.47
   Recovery (% Au)    80.40    66.93    52.71    50.99    40.10    57.79   84.98
  

NaCN Consumption

(lb./ton)

   1.50    2.22    3.30    1.84    3.30    3.80   3.30

 

 

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BTAC
 Conditions 
   Parameter    OG Zone    Otto Zone    Adam’s Peak   

Otto/

Adam’s Peak

   Deep Range
Front
   Range Front     South Pacific  
                 
E    S Oxidation (%)    65.20    92.98    67.32    68.67    36.42    42.92   98.84
   Recovery (% Au)    86.59    89.31    70.02    81.70    41.47    63.04   93.95
  

NaCN Consumption

(lb./ton)

   2.62    1.98    2.84    0.76    0.94    3.60   3.70
                 
F    S Oxidation (%)    67.68    60.17    58.13    44.89    32.95    45.06   98.80
   Recovery (% Au)    86.40    72.76    65.74    61.13    45.18    60.99   94.83
  

NaCN Consumption

(lb./ton)

   1.52    2.66    1.42    2.14    3.02    2.78   4.16

Observations on the results:

 

   

Baseline CIL tests confirmed the refractory nature of the samples with an average recovery of 31.3% Au.

 

   

Acidic BTAC conditions produced the highest gold recoveries:

 

   

BTAC A conditions produced the highest S= oxidation, averaging 95.6% and highest average recovery averaging 88.6% Au. Gold recoveries ranged from 70.2% (99.5% S= oxidation) with sample OGLG to 95.7% (99.6% S= oxidation) with sample SPZV.

 

   

BTAC E conditions produced the next highest S= oxidation, averaging 71.8% and next highest average recovery averaging 79.0% Au. Gold recoveries ranged from 42.2% (31.8% S= oxidation) with sample APHG to 94.5% (61.7% S= oxidation) with sample OGL.

 

   

Alkaline BTAC conditions produced lower S= oxidations and lower gold recoveries.

 

   

BTAC F conditions produced the best alkaline results, likely from the finer grind compared to the baseline alkaline conditions (B). The average S= oxidation was 60.3% with an average gold recovery of 72.6%.

 

   

BTAC B conditions had an average S= oxidation of 59.8%, with an average gold recovery of 69.2%.

 

   

The alkaline condition tests with trona addition produced the lowest S= oxidations and the lowest gold recoveries. BTAC C conditions had an average S= oxidation of 44.1%, with an average gold recovery of 61.5%. BTAC D conditions had an average S= oxidation of 46.7%, with an average gold recovery of 64.6%.

In terms of the how samples from the various zones responded:

 

   

Adam’s Peak and Deep Range Front are the most refractory zones based on their baseline CIL recoveries. The South Pacific Zone variability sample responded well to all BTAC conditions.

 

   

OG Zone samples demonstrate a trend between increasing gold recovery with increasing gold head grade. In terms of the primary zones, OG Zone samples had the highest recoveries, regardless of the POx conditions. The highest average OG Zone gold

 

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  recoveries was with conditions F, (alkaline with finer grind), followed by BTAC A conditions.

 

   

Otto Zone and Adam’s Peak Zone samples showed no clear relationship between gold head grade and gold recovery.

 

   

The highest average Otto Zone gold recoveries with acidic BTAC conditions (E followed by A).

 

   

The highest average Adams Peak Zone gold recoveries was with BTAC A conditions. The average gold recoveries for the other three conditions were significantly lower.

 

   

Composite OAPC best responded to acidic BTAC conditions A and E and responded poorly to alkaline conditions with respect to gold recovery.

 

   

SPZV responded well to all BTAC conditions, with an average gold recovery of 90.5%. Sulphide oxidations ranged from 43.0% with BTAC C conditions (79.0% Au recovery) to 99.5% with BTAC A conditions (95.7% Au recovery).

RFV and DRFV only responded well to BTAC A conditions, with gold recoveries of 95.0% and 89.1% respectively. Gold recoveries for the other conditions were all below 60%.

Overall, there is a positive trend between S= oxidation and gold recovery as shown in Figure 10-7.

Figure 10-7: CIL Gold Recovery as a Function of Sulfide Sulfur Oxidation – Underground Samples

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10.2.4.5

Follow Up BTAC Tests and Roasting Testing

Follow tests aimed at optimizing alkaline POx conditions and to compare against bench top roasting. The blends are shown in Table 10-16. The objective was primarily to evaluate longer retention time under alkaline conditions (Lone Tree autoclave) and to confirm the results from adding trona to POx. Predicted sulphide oxidations and up tests were completed on three sample blends. The test conditions are shown in Table 10-17.

Table 10-16: Granite Creek Underground Metallurgical Sample Blends for Additional Testing

Sample    Composition

Blend 1

   65% OTHG, 34% APHG

Blend 2

   50% OTLG, 34% OUT, 16% APL

Blend 3

   60% OUT, 27% APHG, 13% OTLG

Table 10-17: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Test Conditions

Test Condition    I    II    III    IV    V

Acidulation

   No    No    No    No    No

POx Condition

   Alkaline    Alkaline    Alkaline    Alkaline    Alkaline

Trona Dosage (lb./ton)

   None    None    40    20    None

Temperature (°F)

   390    390    390    390    390

O2 Overpressure (psig)

   100    100    100    100    100

Pulp Density (% solids)

   30    30    30    30    30

Particle Size, k80 (mesh)

   200    200    200    200    200

Retention Time (min)

   60    45    45    45    75

Benchtop roasting (BTR) was performed in a Carbolite HTR rotary reactor tube furnace. A dry ground sample was weighed into a tared baffles borosilicate glass reactor and subjected to a two-

 

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stage roast with oxygen (@99.9% O2 purity) gas applied across the roaster bed. BTR test conditions included:

 

   

Temperature 1 (°C): 986

 

   

Residence 1 Time (min): 30

 

   

Temperature 2 (°C): 1,058

 

   

Residence 2 Time (min): 15

 

   

Ramp Rate (°F/min): 9

 

   

Oxygen Rate (cc/min): 2

 

   

Sample Feed Weight (lb.): 1.1-1.2 (500-550 g)

 

   

Results for sulfide oxidation are shown in and results for gold recovery are shown in Table 10-18 and Table 10 -19 respectively. Overall, the actual sulphide oxidations exceeded predicted values and gold recoveries slightly exceeded predicted values. Extended retention times did not make a significant impact on gold recoveries. Trona additions again did not prove to be beneficial. BTR provided superior sulfur oxidation and gold recoveries for all blend samples.

Table 10-18: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Sulfide Oxidation Results Compared to Predicted Results

     
Test
Condition 
  Actual Results   Predicted Results
  Blend 1   Blend 2   Blend 3   Average   Blend 1   Blend 2   Blend 3   Average

I

  48   57   49   51   -   -   -   -

II

  43   52   42   46   43   60   50   51

III

  39   49   39   42   39   26   33   33

IV

  37   35   41   38   36   33   37   35

V

  55   52   45   51   -   -   -   -

BTR

  99.0   93.9   98.2   97.0                

 

   

 

 

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Table 10-19: Granite Creek Underground Metallurgical Testing Program Follow Up BTAC Gold Recovery Results Compared to Predicted Results

     
Test
Condition 
  Actual Results   Predicted Results
  Blend 1   Blend 2   Blend 3   Average   Blend 1   Blend 2   Blend 3   Average

I

  46   75   71   64   -   -   -   -

II

  65   74   70   70   63   69   66   66

III

  67   73   68   69   63   64   64   64

IV

  52   73   69   65   62   68   65   65

V

  69   77   66   71   -   -   -   -

BTR

  84.0   85.0   80.0   83.0                

 

10.2.4.6

Continuous Pressure Oxidation Testing

 

10.2.4.6.1.

Continuous Pressure Oxidation Operation

A continuous POx test was completed using sample OAPC for three conditions shown in Table 10-20. These conditions were based on the batch BTAC results for the OAPC sample with highest sulfide oxidation and gold recovery. Run 1-Condition B was an alkaline test with no reagents added with six 45-minute turnover intervals, Run 2-Condition E was a partially acidulated test (only partial carbonate destruction) with six 45-minute turnover intervals, and Run 3–Condition A was fully acidulated (total carbonate destruction) with six 60-minute turnover intervals and 437°F.

The three test runs operated back-to-back for approximately 16 hours excluding autoclave heating and cooling time. The first set of POx profile samples were collected once steady state was achieved and another set was collected after three volume changeovers. In addition to sampling, 4.22 gal. (16 L) of POx discharge were collected and weighed at each changeover for downstream testing. Continuous autoclave feed rate for Runs 1 and 2 was approximately 29.3 lb./hour (13.31 kg/hour) of solids to provide 45-minute residence time runs, and for Run 3 the feed rate was approximately 22.0 lb./hour (9.98 kg/hour) to provide the 60-minute residence time.

 

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Table 10-20: Granite Creek Underground Metallurgical Testing Program Continuous POx Run Test Conditions

  Test Condition   B (Run 1)   E (Run 2)   A (Run 3)    
 

POx Condition

  Alkaline   Acid   Acid  
 

Acidulation

  No   Partial   Complete  
 

Trona Dosage (lb./ton)

  0   0   0  
 

Temperature (°F)

  390   390   437  
 

O2 Overpressure (psig)

  100   100   100  
 

Total Pressure (psig)

  304   304   454  
 

Pulp Density (% solids)

  30   30   30  
 

Particle Size, k80 (mesh)

  200   200   200  
 

Retention Time (min)

  45   45   60  

 

10.2.4.6.2.

Continuous Pressure Oxidation Results

The primary objective of the POX process is to oxidize sulfide sulfur to liberate contained refractory gold. Figure 10-8 shows sulfide oxidation by autoclave compartment, starting with the autoclave feed and ending with autoclave discharge. Sulfide oxidation starts quickly in compartments 1⁄2 and reached near full oxidation by compartment 3 for Run 2 and Run 3. As expected, Run 1 did not reach the same level of oxidation as the next two runs under alkaline conditions.

 

 

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Figure 10-8 Granite Creek POx Pilot Plant Sulfide Oxidation Profile

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Source: TR Raponi Consulting Ltd. (2023)

Continuous POX discharge leached in a hot stir tank showed no significant gold recovery difference compared to POX discharge leached in a bottle roll at ambient temperature. The effect temperature had on the discharge was increased lime and cyanide consumption.

Table 10 -21 compares the continuous conditions. Conditions A and E had much higher gold recovery and sulfide oxidation when compared to Condition B but also consumed significantly more lime and recovered much less silver.

 

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Table 10-21: Underground Samples (OAPC) Continuous Autoclave Tests from FLS Program

           
         
Parameter   

Baseline

 

Results

   Run 1 (B) – Alkaline    Run 2 (E) – Partial Acidulation    Run 3 (A) – Full Acidulation
Bottle Roll Conditions    Hot Stir    Bottle Roll    BTAC    Hot Stir    Bottle Roll    BTAC    Hot Stir    Bottle Roll    BTAC
Sulphide
Oxidation (%)
        49    44    97    69    96    99
CIL Recovery
(% Au)
   24    63    63    57    89    91    81.7    91    91    83
CIL Recovery
(% Ag)
   16    50    42    -    3    3    -    2    1    -
Lime
Consumption
(lb./ton)
   2.60    10.5    6.0    -    86.6    78.38    -    111.76    105.50    -
Cyanide
Consumption
(lb./ton)
   4.90    5.90    3.58    3.62    5.12    3.66    0.76    6.60    3.28    1.00

The acid POX conditions (A and E) had much higher sulfide oxidation when compared to alkaline POX (B). This resulted in higher gold recoveries but also required significantly more lime for neutralization. As a result, the increased lime consumption likely caused much of the silver to be locked within jarosite, which did not appear to form in alkaline POX.

 

10.2.4.7

Cyanide Destruction Testing

Cyanide destruction tests were performed using the SO2/air process. The SO2/air process was originally developed and patented by Inco Ltd. (now Vale). SO2 plus air oxidizes cyanide into cyanate, catalyzed by the addition of copper ions. Typical retention times to achieve <5 ppm weak acid dissociable cyanide (CNWAD) are 1 to 2 hours, at pH levels of 7.5 to 9.5.

SO2 is now usually provided in the form of sodium metabisulphite solution dissolved on site or elemental Sulphur combusted to generate SO2 on site for larger users.

This process is capable of achieving discharge concentrations of < 1 mg/L CNWAD. The addition rate of SO2 to cyanide is optimized for consumption and desired discharge cyanide concentration. The process is not suited to directly reducing total cyanide (CNT).

A total of 8.45 gal. (32) L of transitional continuous POX material was collected as discharge from the beginning of Run 2. This material was used to conduct a bulk Carbon-in-Leach test for 24 hours using the autoclave feed reactor and the standard CIL conditions.

After the 24-hour leach, the sample underwent cyanide detox testing with continuous sodium metabisulfate (SMBS) and copper sulfate addition. Conditions of the tests are shown Table 10-23.

 

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Table 10-22: Underground Blend Samples Cyanide Detox Conditions from FLS Program

     
Test Number   1   2
Feed Slurry   Bulk CIL
Oxidizing Agent   SMBS
Feed CNWAD (ppm)   177   182
SO2:CNWAD (mass basis)   8   5
Retention Time (h)   1   1
Copper (ppm)   25   25
pH range   8.0 - 9.0   8.0 - 9.0
Number of Turnovers   6   6
Total Continuous Time (h)   4.5   4.5
Temp °C   77°F   77°F

The CIL slurry was screened for carbon and prepared for the cyanide detox set up. Two of the 32 L of carbon-free slurry were then used to prime the cyanide detox set up, and detox testing was started. Samples were taken from the feed every two hours, and from reactors 1, 2, and the discharge every hour. Cyanide speciation assays for each of these solutions were determined.

Reagents consumptions in both cyanide detox tests can be found in Table 10-24. Lime was used during bulk CIL testing to create the feed for both detox tests.

Table 10-23: Underground Cyanide Detox Reagent Consumption from FLS Program

     
Reagent   Detox 1   Detox 2

Lime Consumption (lb./ton)

  2.88   2.88

SO2 Consumption (lb./ton)

  6.90   6.84

Lime:SO2 (wt/wt)

  0.42   0.60

CuSO4 Maintained (ppm)

  29.6   32.1

CuSO4 Consumption (lb./ton)

  0.14   0.18

Table 10-24 shows the final cyanide results from both cyanide detox tests. The additional SMBS added in Test 1 is shown to result in significantly less CNWAD when compared to Test 2.

Table 10-24: Underground Cyanide Detox WAD from FLS Program

     

Time

(hours)

  Detox 1   Detox 2
  CNWAD (ppm)

1

  22.5   51.1

2

  25.9   42.8

3

  23.2   46.1

4

  33.2   50.5

 

 

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10.2.4.8

Solids Liquids Separation Testing

 

10.2.4.8.1.

Thickener Testing

POX discharge slurry samples were retained from the POX Runs 1, 2, and 3 plus product slurry from Detox 1 (POx Run 2 sample) for solids-liquids separation testing for optimum sizing and selection of thickener operating parameters. Testing was for pre-leach thickener duty for the POx discharge samples and tailings thickener on the cyanide detox sample.

The Run 1 sample was used for the flocculant screening testing. The flocculants tested were 905VHM, 910, VHM, 913VHM, 923VHM and 934VHM. Flocculant 913VHM was found to provide the best overflow clarity and settling velocities when compared to four other flocculants. The findings were consistent for all samples and 913 VHM flocculant was selected for subsequent testing on all samples. This flocculant can be substituted with any comparable anionic polyacrylamide flocculant with a medium molecular weight and low charge density.

Flux testing showed the optimum feedwell suspended solids concentration for flocculation to be approximately 4% by weight for all samples except for the Run 3 sample which is approximately 2% by weight. These diluted densities are much lower than seen for most applications that are in the 10% to 15% solids by weight. Testing shows all samples produced acceptable overflow clarity.

The Run 1 sample (alkaline POX) and the Detox Tailings were the only samples that produced acceptable underflow densities at 51% and 48% solids by weight. Run 2 and Run 3 samples produced thickener underflow densities that were no better than POx discharge densities. Underflow slurry yield stresses are above what is considered the maximum yield stress for centrifugal pumping of 25 Pa.

Thickening results indicate that use of a thickener downstream of POx is not recommended.

 

10.2.4.8.2.

Filtration Testing

FLSmidth conducted pressure filtration tests using a bench-scale filtration testing unit. The bench-scale testing unit can simulate FLSmidth’s recessed chamber and membrane squeeze chamber configurations allowing for various feed solids concentrations, pressure profiles, and cake thicknesses. Filter testing was completed on the Detox Tailings sample to assess the potential for filtered tailings disposal. The pressure filtration test was fed at 48% solids by weight to simulate feeding from a thickener underflow.

While the final filter cake moisture and density are suitable for disposal in a typical disposal area, the filtration rate is about an order of magnitude below what is typical for this duty. The filtration

 

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equipment requirements to implement filtered tailings would be inordinately high but would provide suitable tailings for dry disposal.

10.3  Sample Representativity

Many of the samples used for the metallurgical test work were bulk samples collected either underground or on the benches of the open pits. The precise location of these samples was not fully documented, although the samples were identified as to which pit they originated from. Mag pit and some C and CX samples did include some core drill samples that were well documented as to the precise drill hole location and interval.

10.3.1 Overview

Samples used for the metallurgical test work have been sourced from the open pits (Mag Pit, CX Pit, and underground). Within each zone, drilling has been localized to relatively small portions of the deposit. The metallurgical response of the samples is likely to represent the general behavior of the zone, but sampling different areas of each zone to confirm the metallurgical response will reduce uncertainty. Additional targeted drilling in different zones is recommended to mitigate the risk. Figure 10-9 and Figure 10-10 show the recommended targeted drill hole locations.

 

 

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Figure 10-9 Plan View Showing Metallurgical Sample Locations

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Figure 10-10 Isometric View Showing Metallurgical Sample Locations

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10.3.2 Bulk Samples

Bulk samples were sourced from the Mag Pit and CX Pit. Six bulk samples of approximately 1,000 lbs each were sourced from the Mag Pit and designated Mag Pit I through Mag Pit VI. One bulk sample of approximately 4,300 lbs was sourced from the CX Pit. The locations of the samples were not reported, so it is not possible to assess whether they are representative of the eventual Mineral Resource volume.

10.3.3 Drillhole Samples

The samples selected from drilling on the Project over its life are listed in Table 10-25.

 

 

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Table 10-25: Drillhole Sample Selection and Testing Matrix

     
     
Sample ID    Location    Testing
     
HPR109    Mag Pit    Preg-robbing, bottle roll, column percolation leach tests
HPR129    Mag Pit
HPC142    Mag Pit
HPC143    Mag Pit
     
Magmet-001    Mag Pit    Bottle roll, column percolation leach tests
Magmet-002    Mag Pit
Magmet-003    Mag Pit
Magmet-004    Mag Pit
     
APCX-204    CX Zone    Bottle roll, column percolation leach tests
APCX-211    CX Zone
APCX-219    CX Zone
APCX-226    CX Zone
     
AMW-002    CX Zone    Bottle roll, column percolation leach tests
     
UGOG-004    Underground resource    Head assays and CN soluble Au only.
UGOG-010    Underground
UGOG-013    Underground
UGOG-015    Underground
UGOG-017    Underground
UGOG-018    Underground
UGOG-019    Underground
UGOG-021    Underground
UGOG-022    Underground

Mag Pit drillholes intersect only one end of the mineralized domain. The sampling of the CX Pit is heavily clustered, and much of the mineralized domain has not been assessed metallurgically. Ogee (from the old underground developments) metallurgical test drilling intersects a restricted portion of the mineralized domain. The lens parallel to the existing workings is not intersected by any drilling. No metallurgical test work is available for the A Pit and B Pit.

Generally, drilling intersects only limited areas of the mineralized domains, and testing of additional areas is recommended. Selection of further drilling and sampling for metallurgical testing should be guided by a future mine plan. Metallurgical testing that spatially represents all zones of the project is recommended.

 

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10.3.4 Metallurgical Composite Assembly

For all test work conducted, composites for metallurgical test work were prepared by combining drillhole intervals. From the information available, it is not apparent if the composites were prepared in a way that represents the grade and mineralogical variability within the deposits.

The samples provided by Homestake in 1999 were composited in a manner that tended to reduce the variability in the provided samples. Table 10-26 displays the composite assay and the highest and lowest assays of the intervals in the composite.

Table 10-26: Composite Assays

   
     
Composite   Composite Assay (opt)   High/Low Interval (opt)
     
Mag Pit Cuttings composite 1   0.094   0.027 – 0.266
     
Mag Pit Cuttings composite 2   0.070   0.028 – 0.103
     
Mag Pit Cuttings composite 3   0.068   0.036 – 0.125
     
Mag Pit Cuttings composite 4   0.074   0.035 – 0.102
     
Mag Pit Cuttings composite 5   0.059   0.014 - 0.142
     
Mag Pit Cuttings composite 6   0.076   0.041 – 0.163
     
Mag Pit Drill Core Composite 1   0.032   0.014 – 0.058
     
Mag Pit Drill Core Composite 2   0.077   0.003 – 0.162
     
Mag Pit Drill Core Composite 3   0.098   0.021 – 0.191
     
Mag Pit Drill Core Composite 4   0.058   0.016 – 0.093
     
Mag Pit Drill Core Composite 5   0.146   0.015 – 0.272

The samples do not represent the full variability of the mineralization, and test work should be undertaken on samples that represent different grade variations of the mineralization.

10.4  Deleterious Elements

Both arsenic (As) and mercury (Hg) are present in the mineralization, which is very common in Nevada gold deposits. The arsenic is not cyanide-soluble. However, the mercury is cyanide-soluble and must be collected using appropriate technology at any thermal device processing, stripping, or regenerating carbon. Much like precious metals, mercury will report to the carbon.

Naturally occurring pregnant solution robbing organic carbon is also present within some of the materials at Granite Creek. This limits the applicable processing methods for these materials. High preg-robbing materials are unsuitable for heap leaching and should be treated by CIL methods.

Any autoclave or roasting treatment for the underground refractory material will mobilize the arsenic. If adequate iron is present within the autoclave discharge, the arsenic can be fixed as ferric- arsenate. Any material treated by third-party toll treatment will potentially be subject to additional charges for mercury, arsenic, sulfides, and organic carbon.

 

 

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10.4.1 Homestake Mining 1999

Test work conducted for Homestake did not report the presence or deportment of arsenic.

Mercury was assayed in Mag Pit and CX Pit bulk material samples. Assays ranged from 2.53 ppm to 43.12 ppm Hg; levels high enough to require consideration of mercury capture during refining.

TOC was present at greater than 4% in Mag Pit samples I, II, and IV, all of which displayed significant preg-robbing characteristics. TOC was less than 0.4% in the other samples, and gold recoveries were high. Based on the currently available metallurgical test work, there is not a well-defined relationship between TOC and the level of preg-robbing, but there does appear to be a relationship with the cyanide-soluble gold. A relationship between TOC, sulfide sulfur, and recovery is most likely.

10.4.2 Atna Resources 2005

Drillhole samples from the underground extension of the CX Pit were received for autoclave and cyanide leach testing. These were assayed for arsenic, and measured values between 0.054% and 1.65% were reported and are tabulated with gold by fire assay in Table 10 -27.

Table 10 -27: Gold and Arsenic Assays CX Pit

        
     
Sample    Au (ppm) by Fire Assay    As (%) by AA   
     
APCX-204    8.16    0.066   
     
APCX-211    8.33    0.130   
     
APCX-219    10.25    0.180   
     
APCX-226    17.5    1.650   
     
AMW-002    10.29    0.054   

A further four samples designated R Rib & L Rib, 33941, 33942, and 34259 were also received for autoclave and cyanide leach testing. These tests were completed on three composite samples from the RFZ and one composite from the Ogee Zone mineralization. These samples are understood to be samples collected from the previously mined underground mineralized body. Assays for As and Hg were not reported and assumed not measured.

Approximately 200 samples from nine drillholes in the Ogee Underground resource area were submitted for sample preparation and assaying in March 2006 (Dawson, 2006a). The individual samples were composited into 21 samples for further work. Gold assays ranged from 7.1 ppm to 54.7 ppm and arsenic from 0.09% to 0.46%.

 

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10.4.3 Atna Resources 2013

Thirty-two (32) drill core composites from the Mag Pit area were submitted for heap leach amenability testing. A full elemental analysis was done on each sample, including As, Hg, copper (Cu), and organic carbon. The range of assays is shown in Table 10 -28.

Table 10 -28: Mag Pit Drill Core Composite Assays

   
Element    Assay Range
   
As    77 ppm to 671 ppm
   
Hg    2.1 ppm to 30 ppm
   
Cu    8.7 ppm to 154 ppm
   
C    0.14% to 5.1%

Twenty-three (23) of these composites were tested for heap leach amenability.

10.4.4 Osgood 2023

All the material for the testing campaign was received. Twenty-eight (28) samples were obtained for the various tests outlined in the campaign. A total of 621.6 kg of material was received. Drill hole identifiers were supplied with the samples and included intervals from 57 holes. Mineralogy detected arsenopyrite in all of the sample composites. Realgar was identified in low quantities in several of the samples. It does not appear that As or Hg was assayed for directly in the samples.

10.5  Geometallurgical Modeling

The test programs included cyanide solubility testing, pregnant solution robbing testing, bottle roll testing, percolation column testing, CIL testing, and autoclave testing. The objective of the study was to determine the factors that impact cyanide solubility. Since the cyanide solubility information is unavailable for all the drill holes, a geometallurgical model was created to predict the cyanide solubility based on the available information in the drill hole data. A trend between the cyanide solubility and the column test was determined to predict the heap leach recovery.

The study is divided into 5 zones: Mag pit area, C and CX pit area, A pit area, and B pit area (associated with the open pit zones), and a single underground zone. In all the zones, the gold grade, alterations, and cyanide solubility in some intervals are available. A model for cyanide solubility was created in all zones based on the available cyanide solubility information, see Table .

 

 

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Table 10-29: Available Cyanide Solubility Data in Different Zones

Zone   

Number of Fire

Assays Available

  

Number of CN

Solubility Data

Available

  

Percent of Data with

CN Solubility

Information

Mag Pit    26,589    20,722    77.9%
C and CX Pit    42,732    23,849    55.8%
A Pit    4,234    1,211    28.6%
B Pit    5,109    1,994    39.0%
Underground    139,191    56,887    40.9%

The deposit is mainly associated with silicification, iron oxide, argillization, pyrite, decarbonatization, carbonate, carbon, hydrogen chloride, bleaching, propylitic, and realgar orpiment alterations. All the alterations were logged as shown in Table 10 -30.

Table 10-30: Numerical Equivalent Alteration Codes.

  
   
Alteration Code    Description
0    None
1    Trace (Incipient)
2    Weak (Patchy, Poorly developed)
3    Moderate (Occurs through most rock)
4    Strong (Occurs throughout, textures remain)
5    Complete (Destruction of lithologic texture)

To perform the geometallurgical modeling, different analyses were performed in all the zones:

 

  1.

Principal Component Analysis: A principal component analysis was performed to better understand the relationship between the elements in the ICP geochemical data and its variability. Principal component analysis reduces the variables in a data set to components that attempt to describe the greatest amount of variance in the data set. The first component describes the greatest amount of variance in the data set, the second component is orthogonal to the first and describes the second greatest amount of variance, and so on.

 

  a.

Scree Plots: A scree plot shows the variance of different principal components (variables). On a scree plot, dimensions indicate the amount of variance in the data set described by each component.

 

  b.

Biplots: A biplot allows information on the variables of a data matrix to be displayed graphically, where variables are displayed as vectors. The further away these vectors are from a Principal Component origin, the more influence they have on that Principal Component. Biplots also hint at how variables correlate with the principal components and one another: a small angle implies positive correlation, a large one suggests a negative correlation, and a 90° angle indicates no correlation between two variables.

 

  2.

Regression Tree: Regression techniques contain a single output (cyanide solubility) variable and one or more input variables (alteration mineral species, elevation, and assay grades). The output variable is numerical, and the general regression tree building

 

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methodology allows input variables to be a mixture of continuous and categorical variables. A regression tree is generated when each decision node in the tree contains a test on some input variable’s value. The terminal nodes of the tree contain the predicted output variable values. A regression tree is built through a process known as binary recursive partitioning, which is an iterative process that splits the data into partitions or branches and then continues splitting each partition into smaller groups as the method moves up each branch. The prediction provided by the regression tree model is the mean cyanide solubility for the groups created.

 

  3.

Multivariate Adaptive Regression Splines (MARS): MARS are a form of a non-parametric regression technique and can be seen as an extension of linear models that automatically model nonlinearities and interactions between variables. MARS can handle both continuous and categorical data. Building MARS models often requires little or no data preparation. The hinge functions automatically partition the input data, so the effect of outliers is contained. In this respect, MARS is similar to recursive partitioning which also partitions the data into disjointed regions, although using a different method. The result of the GRE’s non-parametric geometallurgical modeling effort was a formula (the MARS model formula) to determine the cyanide solubility for gold.

10.5.1 Cyanide Solubility for Different Zones

For open pits material only the samples with Au >0.1 ppm (0.003 oz/ton) were considered in all zones. Assays less than this value were considered below cut-off grade and were removed from the study. The material below 0.1 ppm was removed to reduce the impact of waste material on the model.

 

10.5.1.1

Mag Pit Zone

There were 11,812 samples with cyanide solubility information and Au grade above 0.1 ppm. Figure 10 -11 shows the Scree plot, where the first component comprises of the maximum variance. Figure 10 -12 shows the biplot, where gold, pyrite, and iron oxides have strong positive correlation and strong negative correlation to elevation. Figure 10 -13 shows the regression tree model, which shows that gold cyanide solubility is highly dependent on gold grade and elevation.

The MARS model for the gold cyanide solubility in the Mag pit area is shown by two different models, depending on the gold grade. The first model was created where the gold grade is less than 15 ppm, and the second model was created when the gold grade is above 15 ppm.

 

1.

Au >=0.1 ppm and Au <15 ppm

The gold cyanide solubility when gold grade is greater than 0.1 ppm and less than 15 ppm is given by:

 

 

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AuCN ppm = 0.5888 – 0.6124 * max(0, 1.33 – Au ppm)

+ 0.7781 * max(0,Au ppm – 1.33) + 0.26 * max(0, 2 – Pyrite)

– 0.4271 * max(0,Pyrite – 2) – 0.001047 * max(0, 4920 – Elevation(ft))

– 0.0009474 * max(0,Elevation(ft) – 4920)

The cyanide solubility equation uses gold grade, pyrite alteration, and elevation in the model. The graph showing the observed and predicted gold cyanide solubility is given in Figure 10 -14. The model has an R2 of 0.82, implying that 82% of the variations are explained by the model.

Figure 10 -11 PCA- Scree Plot for Mag Pit

LOGO

 

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Figure 10-12 PCA – Biplot for Mag Pit

     LOGO

Figure 10-13 Regression Tree Model for Mag Pit

LOGO

 

 

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Figure 10-14 Observed and Predicted Cyanide Solubility for Gold (ppm)

     LOGO

 

2.

Au >=15 ppm

The gold cyanide solubility when gold grade is greater than 15 ppm is given by:

AuCN ppm = 1.463 + 0.1723 * max(0,Au ppm – 20.74)

+ 0.01517 * max(0,Elevation(ft) – 3934)

The cyanide solubility equation uses gold grade and elevation in the model. The graph showing the observed and predicted gold cyanide solubility is given in Figure 10 -15. The model has an R2 of 0.84, implying that 84% of the variations are explained by the model.

 

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Figure 10-15 Observed and Predicted Cyanide Solubility for Gold (ppm)

    LOGO

 

10.5.1.2

C and CX Pit Zone

There were 5,656 samples with cyanide solubility information and Au grade above 0.1 ppm. Figure 10 -16 shows the scree plot, where the first component comprises of the maximum variance and second component has relatively higher variance. Figure 10 -17 shows the biplot, where gold, pyrite, and iron oxides have a weak positive correlation and weak negative correlation to elevation. Figure shows the regression tree model, which shows that gold cyanide solubility is highly dependent on only gold grade.

Figure 10-16 PCA- Scree Plot for C and CX Pit

    LOGO

 

 

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Figure 10-17: PCA – Biplot for C and CX Pit

    LOGO

 

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Figure 10-18: Regression Tree Model for C and CX Pit

LOGO

The MARS model for the gold cyanide solubility in C and Cx pit area is shown as:

AuCN ppm = 23.87 – 0.8948 * max(0, 26.91– Au ppm)

+ 0.6038 * max(0, Au ppm – 26.91) + 0.2289 * max(0, 1 – Pyrite)

– 1.748 1 * max(0, Pyrite – 1) – 0.00116 * max(0, 4308 – Elevation(ft))

The cyanide solubility equation uses gold grade, pyrite alteration, and elevation in the model. The graph showing the observed and predicted gold cyanide solubility is given in Figure . The model has an R2 of 0.92, implying that the 92% of the variations are explained by the model.

 

 

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Figure 10-19: Observed and Predicted Cyanide Solubility for Gold (ppm)

   LOGO

 

10.5.1.3

A Pit Zone

There were 487 samples with cyanide solubility information and Au grade above 0.1 ppm. Figure 10 -20 shows the scree plot, where the first component comprises the maximum variance. Figure 10 -21: PCA – Biplot for A Pit shows the biplot, where gold and elevation have a positive correlation. Figure the regression tree model shows that gold cyanide solubility is highly dependent on only gold grade.

The MARS model for the gold cyanide solubility in A pit area is shown as:

AuCN ppm = 1.365 – 0.9794 * max(0, 1.44 – Au ppm)

+ 1.005 * max(0, Au ppm – 1.44)

The cyanide solubility equation uses only gold grade in the model. The graph showing the observed and predicted gold cyanide solubility is given in Figure 10 -23 The model has an R2 of 0.999, implying that the 99.9% of the variations are explained by the model.

 

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Figure 10-20: PCA- Scree Plot for A Pit

    LOGO

Figure 10-21: PCA – Biplot for A Pit

     LOGO

 

 

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Figure 10-22: Regression Tree Model for A Pit

   LOGO

Figure 10-23 Observed and Predicted Cyanide Solubility for Gold (ppm)

    LOGO

 

10.5.1.4

B Pit Zone

There were 617 samples with cyanide solubility information and Au grade above 0.1 ppm Figure 10 -24 shows the scree plot, where the first component comprises of the maximum variance. Figure 10 -25 shows the biplot, where gold and elevation have a positive correlation. Figure 10 -26 shows the regression tree model, which shows that gold cyanide solubility is highly dependent on gold grade and pyrite alteration.

 

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The MARS model for the gold cyanide solubility in B pit area is shown as:

AuCN ppm = 3.002 – 0.6082 * max(0, 4.121 – Au ppm)

+ 1.071 * max(0, Au ppm – 4.121) – 0.4479 * max(0, 1– FeOx)

– 0.4769 * max(0, Pyrite – 1)

Figure 10-24 PCA- Scree Plot for B Pit

    LOGO

 

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Figure 10-25: PCA – Biplot for B Pit

LOGO

Figure 10 -26: Regression Tree Model for B Pit

 

 

LOGO

 

 

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The cyanide solubility equation uses gold grade, pyrite alteration, and iron oxide alteration in the model. The graph showing the observed and predicted gold cyanide solubility is given in Figure 10 -27. The model has an R2 of 0.91, implying that 91% of the variations are explained by the model.

Figure 10 -27: Observed and Predicted Cyanide Solubility for Gold (ppm)

 

LOGO

 

10.5.2

Cyanide Solubility Estimation in the Block Model

Using the drill hole information in each pit and the underground zone, the gold grades were estimated, and alteration information was coded for all the blocks. All the blocks are associated with the elevation, and the elevation for the block center is used for cyanide solubility estimation. Using the MARS model for each domain, the cyanide solubility information was calculated. This information was later used to calculate the recovery for each block.

 

10.5.3

Metallurgical Test and Recovery

Column test and the CIL test information are available in the McClelland April 1999 report (McClelland, 1999a). The test samples are primarily located in the Mag pit area and a single sample in the C and CX pit area. Autoclave results are available in the DML Wilmot 2005 -2006 Memo (Wilmot, 2006).

There are 16 Column leach tests available and seven CIL tests. The number of samples and the locations of the samples are not spatially representative of the deposit.

 

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Table 10 -31: Column Test and CIL Test (McClelland April 1999 Report)

Sample Composite  

  Sample  

Type

 

Column

 Feed Size 

 

Column

Au

 Recovery 

Percent

 

 NaCN 

lb/ton

 

Lime

 lb/ton 

   NaOH   

C

 organic 

Percent

 

CIL Au

 Percent 

 

Calc

 Head 

opt

 

CN

 Solubility

%

Mag Pit I

  Bulk High-
Grade
Material
  -4 inch   18.8%   9.94   5.2   2.8   4.55%   94.4%   0.138   19.5%

Mag Pit II

  -4 inch   35.3%   8.96   10.2   2.34   4.25%   75.3%   0.085   19.6%

Mag Pit III

  -4 inch   93.1%   4.55   5.2       0.15%   59.7%   0.058   84.5%

Mag Pit IV

  -4 inch   49.5%   5.30   12.0       4.00%   82.9%   0.105   22.8%

Mag Pit V

  -4 inch   51.7%   3.85   2.5       0.40%   55.0%   0.029   50.0%

Mag Pit VI

  -4 inch   60.7%   3.66   4.0       0.45%   87.5%   0.028   61.8%

CX Pit, CX-2

  -6 inch   77.7%   5.11   3.0       0.35%   88.2%   0.091   91.9%

CX Pit, CX-2

  P80 3
inch
  81.7%   4.80   3.0               0.089   91.9%

CX Pit, CX-2

  P80 3/4
inch
  82.2%   5.39   3.0               0.090   91.9%

Mag Pit 2

  Core Comp   Nom 1
inch
  69.0%   3.98   11.0               0.058   77.1%

Mag Pit 3

  Core Comp   Nom 1
inch
  62.0%   1.60   9.6               0.079   44.2%

Mag Pit 4

  Core Comp   Nom 1
inch
  47.9%   1.51   8.1               0.048   72.0%

Mag Pit 5

  Core Comp   Nom 1
inch
  61.7%   2.08   10.0               0.141   22.6%
Mag Pit Master (pH 10.5)   Core Comp   Nom 1
inch
  65.0%   6.26   8.4       2.25%       0.080   49.4%

 

 

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Sample Composite  

  Sample  

Type

 

Column

 Feed Size 

 

Column

Au

 Recovery 

Percent

 

 NaCN 

lb/ton

 

Lime

 lb/ton 

   NaOH   

C

 organic 

Percent

 

CIL Au

 Percent 

 

Calc

 Head 

opt

 

CN

 Solubility

%

Mag Pit Master (pH 11.8, Lime)   Core Comp   Nom 1
inch
  70.7%   4.15   19.3       2.25%       0.082   49.4%
Mag Pit Master (pH 12.0, NaOH)   Core Comp   Nom 1
inch
  69.0%   3.48   N/A   10.7   2.25%       0.084   49.4%

Assumption: The recovery properties observed in the Mag pit are similar to the entire area (Mag Pit, C and Cx pit, A Pit, B Pit, and Underground Zone).

 

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10.5.3.1

Recovery Model

Recovery models are based on Table 10 -31. The recovery models are based on the actual metallurgical tests. The Heap Leach Recovery model is used for both the open pit areas (Mag pit, C and Cx pit, A pit, and B pit) and the Underground area. CIL recovery model is used only in the open-pit area. The autoclave is used only for the underground area, except for any oxide material encountered above the 4670-foot level. The material above the 4670-foot level in the underground area can be treated in the heap leach or autoclave depending on the revenue generated by the blocks in the different processing options.

 

10.5.3.1.1.

Heap Leach Recovery

Heap Leach recovery is determined by plotting the cyanide solubility with the column recovery. Figure 10 -28 shows that a few samples have lower cyanide solubility but have variable column recovery, ranging from 20% to 70%.

A few samples show that the model has 20% column recovery at 20% cyanide solubility, 50% Column Recovery at 50% cyanide solubility, and 60% Column Recovery at 60% cyanide solubility. Looking at the trend, a conservative model is created, where up to 60% cyanide solubility, the heap leach recovery is 60%, and above 60% cyanide solubility, the model follows

LOGO

 

 

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Figure 10 -28: Cyanide Solubility vs Column Recovery

LOGO

 

10.5.3.1.2.

Carbon in Leach Recovery

Within the Mag Pit, the samples were looked at, and the trend between the calculated head grade and CIL Recovery was plotted (Figure 10 -29).

Figure 10 -29: Calculated Head Grade vs Carbon in Leach Recovery

LOGO

Comparing the CIL results to column leach tests of the same samples revealed that several outliers needed to be addressed. In one case, a column test showed a gold extraction of 93% while the CIL test provided only 55% gold recovery, two additional samples were removed for similar reasons. To create a more realistic model, the outliers were removed as shown in Figure 10 -30. Multiple

 

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regression analysis did not find a statistical relationship between CIL recovery and organic carbon or sulfide sulfur. In most cases, the sulfide sulfur grade of the samples was low.

The number of CIL tests available was low, but significant direct leach bottle roll tests (BRT) informed the selection of the CIL results. Gold extraction from BRT showed a strong relationship to the organic carbon concentrations and a weaker relationship to sulfide sulfur. The sulfide sulfur relationship was likely impacted by the low sulfide grade variability examined. CIL testing indicated that the preg-robbing impact of the TOC was largely overcome in most cases.

Figure 10-30: Solubility vs BRT Recovery

 

LOGO

By parsing the data Figure 10 -31 a reasonable trend between the head grade and recovery can be observed. This trend is used to predict CIL recovery, which is capped at 95%.

 

LOGO

 

 

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Figure 10-31 Calculated Head Grade vs Carbon in Leach Recovery (Outlier Removed)

LOGO

It is important to note that no limit on sulfide sulfur has been included in this recovery formula but it likely would play a role. Additional test work is required to better define this relationship and careful ore control measures will need to be adopted to ensure that refractory material is not directed to the CIL.

 

10.5.3.1.3.

Autoclave Recovery

For the autoclave recovery tests, 25 samples were available (Table 10 -32, Table and Table 10 -34). The majority of the tests had a high recovery, ranging from 81% to 97%. One sample had a low recovery (OGLG) and the reason for this was not determined.

 

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Table 10-32: DML Wilmot 2005 -2006, Memo Autoclave Test Results (Samples from 2005)

October 2005  

One hour 225 C, 460 psi, Acidulate to 1.8 to 2.0 pH Autoclave Discharge, 2

gm/l NaCN 24 hr

              
Description   Test#  

 Sample 

ID

 

Grind

P80

 (microns) 

 

 Head 

Assay

(ppm)

 

 Head 

Calc

(ppm)

 

Leach

 Residue 

(ppm)

 

Gold

 Extracted 

(ppm)

 

Gold

 Recovery 

(%)

  Consumption  

CN

 Shake 

Test

 

CaO

 (Kg/Mt) 

 

NaCN

 (Kg/Mt) 

Ogee channel sample, oxide   1   LL&RR   75   13.71   12.22   0.81   11.4   93.2%   4.2   1.32   86%
Range Front Sample, surface drilling   4   34259   75   14.74   14.8   1.49   13.1   88.5%   32.1   4.72   11%
Range Front Sample, surface drilling   2   33941   75   8.23   9.01   0.58   8.4   93.0%   19.6   1.56   53%
Range Front Sample, surface drilling   3   33942   75   14.74   15.63   0.67   14.9   95.2%   47.2   2.37   61%
Range Front Sample, surface drilling   5   33942   45   14.74   15.57   0.41   15.1   97.0%   47.7   2.44   61%

Table 10 -33: DML Wilmot 2005 -2006, Memo Autoclave Test Results (Samples from 2006)

April 14, 2006   One hour 225 C, 460 psi, Acidulate to 1.8 to 2.0 pH  Autoclave Discharge, 2
gm/l NaCN 24 hr
              
Description   Test#  

 Sample 

ID

 

Grind

P80

 (microns) 

 

 Head 

Assay

(ppm)

 

 Head 

Calc

(ppm)

 

Leach

 Residue 

(ppm)

 

Gold

 Extracted 

(ppm)

 

Gold

 Recovery 

(%)

  Consumption  

CN

 Shake 

Test

 

CaO

 (Kg/Mt) 

 

NaCN

 (Kg/Mt) 

CX Sample composites   8   APCX-219   75   10.25   8.35   0.72   7.6   91.0%   17.5   1.2   60%

 

 

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CX Sample composites    9    APCX-226     75      17.5     15.81     0.89      14.84      93.8%      21.2       2.38       42% 
From Barrick pre-ground   6   MET 1   88   n/a   15.96   0.96   14.8   92.7%   71.4   1.36   n/a
From Barrick pre-ground   7   MET 2   88   n/a   10.06   2.06   7.81   77.6%   23.4   0.88   n/a
Coarse sample -3/8   2   MET 1   75   8.23   9.01   0.58   8.38   93.0%   19.6   1.56   n/a
Coarse sample -3/8   3   MET 2   75   14.74   15.63   0.67   14.88   95.2%   47.2   2.37   n/a

Table 10-34: Autoclave Test Results (Samples from 2023)

 

      Condition A - Acid    Condition E - Partial  Acid

Sample

ID

  

S=

Oxidation

(%)

   CIL Recovery   

S=

Oxidation

(%)

   CIL Recovery
  

Au

(%)

  

Ag

(%)

  

Au

(%)

  

Ag

(%)

OGU

   97    91    13    47    82    84

OGL

   57    94    0    62    94    0

OGHG

   91    81    15    91    94    1

OGLG

   100    70    0    61    76    53

OTU

   100    88    42    99    90    0

OTL

   99    87    27    98    90    77

OTHG

   99    92    45    74    87    62

OTLG

   100    88    87    100    90    0

APU

   98    95    1    95    95    0

APL

   98    87    3    97    88    20

APHG

   100    91    0    32    42    0

APLG

   100    90    2    45    55    0

OAPC

   99    83    19    69    82    16

DRFV

   96    89    0    36    41    30

RFV

   97    95    0    43    63    0

SPZV

   99    96    0    99    94    0

 

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10.5.4 Recovery in the Block Model

The recovery in the block model uses the estimated cyanide solubility information (Section 10.5.2). For each block, the recovery equations from Section 10.5.3 are used to calculate the recovery in different processing destinations.

 

10.6

Conclusions

10.6.1 Sample Representativity

Within each zone, drilling has been localized to relatively small portions of the mineralized domains, as seen in Figure 10-9 and Figure 10-10. The samples’ metallurgical response is likely to represent the zone’s general behavior, but additional sampling of each zone to confirm the metallurgical response will reduce uncertainty. The lack of this metallurgical drilling remains a risk to the project.

10.6.2 Test Work on Open Pit Samples

Cyanide leach bottle roll tests and column leach tests were completed on samples from both the Mag and CX open pits. Both Homestead and Atna commissioned these tests.

The test work demonstrated that many of the Mag Pit samples had high preg-robbing factors due to carbonaceous material in the feed. Due to the variable preg-robbing characteristics of the feed material, a higher degree of representativity of the Mag Pit should be evaluated.

Bottle roll tests were conducted on Mag Pit samples using NaOH as an alternative to hydrated lime, as a method of treating material with preg-robbing characteristics. These tests demonstrated that raising the pH improved gold recovery and decreased cyanide consumption.

A column leach test on a Mag Pit sample showed that there was no gold recovery benefit in using NaOH rather than lime (at the equivalent pH).

Test work on ground materials showed that Mag Pit materials were amenable to CIL methods. CIL treatment showed low impact from the TOC. Gold recoveries ranged from 83% to 94%.

Column leach tests on the Mag Pit samples achieved gold recoveries in the range of 19% to 82%.

Column leach tests on the CX Pit samples achieved gold recoveries of 82%.

 

10.7

Recommendations

The following recommendations have been put forward:

 

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10.7.1 Test Work Recommendations

A metallurgical drilling program should be undertaken to collect samples within the various zones representing the spatial, mineralogical, and grade differences. The collected samples should be tested for the following:

 

   

Paired fire assays and cyanide soluble assays to define cyanide solubility.

 

   

Bottle roll tests with and without carbon to predict reagent consumption as well as amenability to CIL treatment and to evaluate the impact of sulfide sulfur on the CIL performance.

 

   

Column leach tests at various sizes to predict field recovery for material to be heap leached. This should be performed on those materials with a cyanide solubility of greater than 50%. Recovery by size fraction should be completed as part of the testing program.

 

   

Conduct SAG and ball mill testing to determine the work index.

 

   

Additional autoclave pretreatment of underground materials should be completed, especially for those materials that showed lower gold extraction.

 

   

Infill the drill hole database with TOC and S= assays.

 

   

Conduct arsenic and mercury assays on all samples employed for metallurgical testing.

10.7.2 Geometallurgy Recommendations

The geometallurgical work completed as part of this technical report should be expanded using the planned metallurgical test program results. The intent will be to confidently define those materials that can be treated by heap leaching or CIL methods and those that require autoclave treatment.

 

 

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11 Mineral Resource Estimates

 

11.1

Introduction

Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305 provides the following definitions for mineral resources:

Mineral resource is a concentration or occurrence of material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. A mineral resource is a reasonable estimate of mineralization, taking into account relevant factors such as cut-off grade, likely mining dimensions, location or continuity, that, with the assumed and justifiable technical and economic conditions, is likely to, in whole or in part, become economically extractable. It is not merely an inventory of all mineralization drilled or sampled.

Inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. The level of geological uncertainty associated with an inferred mineral resource is too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability. Because an inferred mineral resource has the lowest level of geological confidence of all mineral resources, which prevents the application of the modifying factors in a manner useful for evaluation of economic viability, an inferred mineral resource may not be considered when assessing the economic viability of a mining project, and may not be converted to a mineral reserve.

Indicated mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of adequate geological evidence and sampling. The level of geological certainty associated with an indicated mineral resource is sufficient to allow a qualified person to apply modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Because an indicated mineral resource has a lower level of confidence than the level of confidence of a measured mineral resource, an indicated mineral resource may only be converted to a probable mineral reserve.

Measured mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of conclusive geological evidence and sampling. The level of geological certainty associated with a measured mineral resource is sufficient to allow a qualified person to apply modifying factors, as defined in this section, in sufficient detail to support detailed mine planning and final evaluation of the economic viability of the deposit. Because a measured mineral resource has a higher level of confidence than the level of confidence of either an indicated

 

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mineral resource or an inferred mineral resource, a measured mineral resource may be converted to a proven mineral reserve or to a probable mineral reserve.

Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no guarantee that all or any part of the Mineral Resource will be converted into Mineral Reserve. Confidence in the estimate of Inferred Mineral Resources is insufficient to allow

 

 

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the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure.

The Mineral Resource Statement presented herein represents the updated Mineral Resource Estimate for the Granite Creek deposit located in Humboldt County, Nevada, USA. This Mineral Resource Estimate was prepared by GRE for i-80 to complete a S-K 1300 Technical Report Summary. The most recent previous Mineral Resource Estimates were contained in the reports titled:

 

   

“Preliminary Economic Assessment NI 43-101 Technical Report, Granite Creek Mine Project, Humbolt County, Nevada, USA” (GRE, 2021)

 

   

“Technical Report, Osgood Pinson Deposit NI 43-101 Technical Report, Osgood Mining Company, LLC, Humboldt County, Nevada, USA” produced by Osgood Mining Company, LLC (AMC, 2019).

The mineral resources were estimated in conformity with and are reported in accordance with the S-K 1300 requirements. This mineral resource estimate includes inferred mineral resources, which are defined as “that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. The level of geological uncertainty associated with an inferred mineral resource is too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability…” (CFR, 2018). There is no certainty that the inferred mineral resources will be converted to the measured or indicated categories through further drilling or into mineral reserves, once economic considerations are applied. Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the mineral resource will be converted into mineral reserves. The project presently has no mineral reserves. Whittle Pit optimization was applied to the open pit mineral resource estimate to assess the reasonable prospects for economic extraction for the resource.

The open pit Mineral Resource Estimate for the Granite Creek Mine Project was completed by GRE. The effective date of the resource statement is December 31, 2024. In the opinion of GRE, the Mineral Resource Estimate reported here is a reasonable representation of the mineral resources found in the open pit portion of the Granite Creek Mine Project at the current level of sampling.

 

11.2

Drill Hole Database

GRE performed a data validation of the drill hole database prepared by i-80 for the Granite Creek deposit and determined it to be of suitable accuracy to perform a mineral resource estimate for the

 

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property. More detail regarding the validation of the drill hole database can be found in Section 9. The drill hole data for the Granite Creek Mine Project was delivered as a separate .csv file that contained exploration and production collar locations, drill hole survey orientations, sample intervals with gold assays in ppm, geologic intervals with lithology, alteration type, and alteration strength. The collar locations are projected in a local grid system, with planar and elevation units in feet. All downhole intervals are captured in feet.

The complete data set contained assays, collar, and survey data for a total of 2,855 exploration holes (surface, underground, and trench samples) and 695 production holes (surface and underground). Drilling is a mix of RC drilling, diamond drilling, and RC pre-collar with diamond drilling to final depth. The exploration assay file contains 212,839 gold assays. The production data assay file contains 1,477 gold assays. The drill hole collar locations are shown in Figure 11-1.

Figure 11-1: Drill Holes Used Plan View on Topography

 

 

LOGO

Note: This figure is intended to show the relative distribution of surface drill hole collars on topography around the areas of interest. This figure does not show all collars that have been drilled.

A number of negative, missing, and blank assay values exist in the drill hole data files provided to GRE by i-80. Missing intervals and values were assumed to be non-mineralized and therefore

 

 

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assigned a value of half of the most common detection limit used to assay the samples. Negative assay values were replaced according to Table 11-1.

Table 11-1: Negative Values in Drill Hole Database

 

       
Non-Positive
Value
  Interpretation   

Count In 

Data Set

  Action Taken
     

-0.005 

  below detection limit of 0.005 opt    6417   replaced with 0.0857 ppm
     

-0.003 

  below detection limit of 0.003 ppm    2723   replaced with 0.0015 ppm
     

-0.9943 

  below detection limit of 0.029 opt    429   replaced with 0.0857 ppm
     

-5557 

  Sample Not Received    3   Omit
     

-5556 

  Sample Not Received    80   Omit
     

-0.0343 

  Half the detection limit of 0.002opt    926   replaced with 0.0343 ppm
     

-0.1714 

  below detection limit for 0.005 opt    52   replaced with 0.0857 ppm
     

-3394.2842 

  conversion of -99 opt to ppm    13   replaced with 0.0857 ppm

 

11.3

Topography

Topography was provided by i-80 as dxf files with triangulated surfaces. The files included both as built surfaces showing dimensions of previously mined pits at their maximum depths, and present topography which includes backfill, pits, dumps, and surrounding topography. The current topographic data was loaded into Leapfrog Geo and used to constrain the block model. The topographic data provided by i-80 was not rectangular, which is required within Leapfrog to generate models; therefore, GRE extrapolated topographic data around the edges to form a rectangular surface (Figure 11-2). The extrapolated area, however, is not part of the resource estimate.

 

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Figure 11-2: Current Topography Used for Resource Estimation

 

LOGO

 

11.4

Geologic Model

The geologic model used to complete the mineral resource estimate was developed by GRE using grouped majority composites for lithology based on data provided to GRE as part of the drill hole database. Material below the current topography and above the as-built surface was classified as backfill and assigned an Au ppm grade of zero. See Figure 11-3 for illustration of the geologic model used in the resource estimation. The model was validated for geologic accuracy and found to be suitable for the purpose of Mineral Resource Estimation by GRE.

 

 

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Figure 11-3: Geologic Model Oblique View

 

LOGO

 

11.5

Open Pit Estimation

 

11.5.1

Estimation Domains

Estimation zones were recreated by bounding the assays that surround the pit areas (Figure 11-4). The underground zone to the North of Zone 3/CX Pit was not considered since this area was estimated in the underground resource section of this report. Vertical extents of the estimation zones range from 5500 feet amsl to 2500 feet amsl.

 

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Figure 11-4: Open Pit Estimation Zones

 

LOGO

Table 11-2 summarizes the estimation zone numbers along with the corresponding pits.

Table 11-2: Open Pit Estimation Zone and Pit Name

 

   

Estimation

Zone

  Pit Name

Zone 1

  A Pit

Zone 2

  MAG Pit

Zone 3

  CX Pit

Zone 4

  B Pit

Numeric indicator models were constructed to better define the high-grade mineralized domains contained within the generalized estimation domains fit around the existing pits. The parameters used to define the indicator models are shown in Table 11-3.

Table 11-3: Open Pit Numeric Indicator Model Parameters

         
Estimation Zone    Indicator
Model
Cutoff (ppm)
  

ISO

Value

   Search Distance
(feet)
   Dynamic
Anisotropy

Zone 1

   1.0    0.4    200    CX Fault

Zone 2

   1.0    0.3    250    Mag Fault

 

 

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Zone 3 CX

   0.1    0.4    180    CX Fault

Zone 3 SOS DIKE

   0.1    0.4    100    SOS Dike

Zone 3 SOS XSECTION 

   0.1    0.3    150    SOS X Section Fault

Zone 4

   1.0    0.3    200    NA*

*Because Zone 4 did not use dynamic anisotropy, a global trend set to the following parameters was used: dip 90, dip Azimuth 100, pitch 75, and ellipse ratios max. 200, int. 200, min. 100.

Initial search ellipse orientations were set from examining the spatial orientation of the composites greater than 1 g/t (see Figure 11-5). After initial construction of the high grade solids using indicator models, it was noted that some of the high-grade numeric models had voids or otherwise poor geometry. When examining the fault structures, it was noted that the high grade domain corresponded well with the location and orientation of several fault structures. GRE then attempted to add a structural trend using dynamic anisotropy to the numeric model by constructing a structural trend from the fault meshes. This improved the continuity of the indicator models and helped eliminate voids that were previously present in the indicator models. ISO factors, which are defined as the probability that the enclosing volume encloses the values above the cutoff, for interpolants were based on visually examining the mineralized body and using an iterative process to select a value that produced a reasonable geologic shape. If small islands of volume were created off the major trend, they were clipped out.

 

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Figure 11-5: Example of Numeric Indicator High Grade Trend Analysis Mag Pit

 

LOGO

Note: The opaque red solid represents the high grade domain. The transparent red solid is the low grade domain. Composite grades are shown at a cutoff of 1ppm.

One resource estimation zone, Zone 3, was broken into several sub domains after failing to produce a reasonable high grade numeric indicator model using a single global trend within the Zone 3 domain. It was noted that during the initial attempt, branching solids formed displaying three distinct trends. Upon further investigation it was found that these trends corresponded to fault structures that crosscut the Zone 3 domain. The identified structural trends are the CX fault, the SOS Dike, and the SOS X-Section.

To construct estimation sub-domains, GRE offset the fault meshes both forwards and backwards to a thickness that contained the majority of the high-grade intercept (see Figure 11-6). Separate numeric estimators were constructed within these domains, and high grade and low-grade zones were defined within the zones that contained the mineralized trends (see Figure 11-7). To later avoid estimation boundary issues during resource estimation, the volumes on either side of the mineralized domains were separated into sub domains. These include the HW 1, HW 2, HW 3, and FW zones.

 

 

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Figure 11-6: Open Pit Zone 3 Sub-Domains

 

LOGO

Figure 11-7: High Grade and Low Grade Open Pit Domains in CX Fault

 

LOGO

 

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Note: The high grade zone is shown in green, the low grade domain is shown in purple.

 

11.5.1.1

Domain Analysis

To check the validity of the high grade and low-grade estimation domains, box and whisker plots were constructed, as shown in Figure 11-8. Generally, a good correlation was observed between the high grade and low-grade solids and the distribution of the grades contained within them.

Figure 11-8: Box and Whisker Plot of Open Pit Estimation Domains

LOGO

 

 

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Figure 11-9: HG and LG Distributions in Zones 1 and 2

 

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Figure 11-10: HG and LG Distributions in Zones 3 and 4

 

LOGO

 

11.5.2

Assay Compositing

Sample data was composited to intervals of equal length to ensure that the samples used in statistical analysis and estimations were equally weighted. To accomplish compositing, GRE first examined the interval histogram to determine the most common assay length (see Figure 11-11).

 

 

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Figure 11-11: Open Pit Interval Length Statistics of Au ppm Assays

 

LOGO

Once it was determined that five feet was the primary assay length, GRE evaluated various compositing lengths using 5-foot intervals to avoid splitting assays. It was decided that compositing on a 20-foot interval represented a significant decrease in the variance of the data while not adversely decreasing the mean of the data set, as shown in Table 11-4. Therefore, GRE selected a 20-foot composite interval.

Table 11-4: Open Pit Compositing Interval Statistics

Statistic    AU_ppm_Assays    Composite Interval
   5-foot    10-foot    15-foot    20-foot    25-foot

Count

   221,062    264,712    133,121    88,217    66,103    52,752

Length

   1,323,435.3    1,323,396.9    1,323,367.4    1,322,037.1    1,321,859.4    1,321,327.9

Mean

   0.362    0.362    0.361    0.358    0.357    0.354

 

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Statistic    AU_ppm_Assays    Composite Interval
   5-foot    10-foot    15-foot    20-foot    25-foot

SD

     2.58        2.49        2.31        2.08        1.95        1.87  

CV

   7.1    6.9    6.4    5.8    5.5    5.3

Variance

   6.6    6.2    5.4    4.3    3.8    3.5

Minimum

   0    0    0    0    0    0

Q1

   0.0170    0.0171    0.0171    0.0171    0.0171    0.0171

Q2

   0.0170    0.0171    0.0274    0.0343    0.0343    0.0343

Q3

   0.0860    0.0857    0.0857    0.0857    0.0857    0.0857

Maximum

   290.06    290.06    290.06    119.89    94.33    85.08

A box plot comparison of the 20-foot composited and the uncomposited assays is shown in Figure 11-12. This comparison shows that compositing, while not changing the mean or quartiles, does drastically reduce the maximum value of grades.

Figure 11-12: Open Pit Compositing Comparison 20 Foot Intervals

 

LOGO

Table 11-5: Open Pit Compositing Comparison 20 Foot Intervals

     
Statistic    Composited      Uncomposited  
     
Count      50,633        221,675  

 

 

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Statistic    Composited      Uncomposited  
     
Length      1,011,524        1,327,165  
     
Mean      0.45        0.36  
     
SD      2.22        2.57  
     
CV      4.93        7.13  
     
Variance      4.93        6.63  
     
Minimum      0        0  
     
Q1      0.02        0.02  
     
Q2      0.04        0.02  
     
Q3      0.09        0.09  
     
Maximum      94.33        290.06  

 

11.5.3

Evaluation of Outliers

Cumulative probability plots for gold were completed for the composites within each estimation domain (see Figure 11-13 for an example). A break in the population was identified and marked with the clipping line. Based on this analysis, GRE applied a maximum allowable value for the gold grade within each separate domain, as shown in Table 11-6.

 

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Figure 11-13: Example of Open Pit Cumulative Log Probability Plot Zone 1 HG

 

LOGO

Table 11-6: Open Pit Upper Clipping Au ppm Values by Domain

 

     
Zone    Sub-Domain    Clipping Value
     

Zone 1

   HG    35
   LG    NA
     

Zone 2

   HG    10
   LG    5
     

Zone 3 

   CX HG    20
   CX LG    10
   SOS DIKE HG    NA
   SOS DIKE LG    NA
   SOS XSECTION HG    NA
   SOS XSECTION LG    NA
   HW 1    7

 

 

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Zone    Sub-Domain    Clipping Value
   
   HW 2    3
   
   HW 3    7
   
   FW    12
     

Zone 4

   HG    NA
   LG    3

 

11.5.4

Density

Density was assigned to each domain in the block model based on a combination of rock type and grade, as shown in Table 11-7. The bulk densities are the same as those used in the 2020 Getchell Project Technical Report (AMC, 2020) and were originally supplied by OMC. The results for each domain fit well with GRE’s experience with similar rock types.

Table 11-7: Open Pit Domain Density Summary

 

     
Unit  

Au>=0.008
opt

(tonne/m3)

   

Au<0.008
opt

(tonne/m3)

 
     

Backfill

    1.85       1.85  
     

Alluvium

    1.85       1.85  
     

Granodiorite

    2.7       2.7  
     

Upper Comus

    2.5       2.7  
     

Lower Comus

    2.51       2.64  
     

Preble

    2.42       2.6  

 

11.5.5

Variography

After iterative analysis, a good fit for the gold grade variography was found using pairwise relative variograms. The pairwise relative variogram helps to smooth the variogram by scaling g(h) using the square of the mean of each sample pair of the data from calculating g(h). This makes the interpretation of the variogram model easier, and all variances calculated this way are relative to the mean of the sample pairs within the distribution.

Variogram analysis was completed on the samples within each of the high grade and low grade estimation domains to establish the direction of maximum continuity between sample pairs. The range for each variogram was found using a global variogram. The nugget was determined by examining the downhole variograms and determining where the short-range trend crossed the y-axis. Variograms were orientated along the strike and dip of the visually observed high grade trend of the composites, with the major axis oriented along the direction of maximum continuity.

 

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Table 11-8: Open Pit Variogram Parameters

Zone    Sub-Domain    Dip   

Dip

 Azimuth 

   Pitch   

 Major 

Axis

 

Semi-

 Major 

Axis

 

 Minor 

Axis

   

Zone 1

   Overall   38   134   75   200   200   100  
   HG   38   134   160   160   160   75  
   LG   38   134   75   175   175   60  

 Zone 2 

   Overall   50   70   105   250   250   100  
   HG   50   70   75   180   125   75  
   LG   50   70   105   300   200   200  

Zone 3

   Overall*   56   135   75   160   160   125  
   CX   56   135   75   180   160   70  
   CX HG   56   135   75   80   80   50  
   CX LG   56   135   75   50   50   50  
   SOS DIKE   65   170   45   100   80   50  
   SOS DIKE HG   65   170   45   100   80   75  
   SOS DIKE LG   65   170   45   100   50   25  
   SOS XSECTION   80   170   80   150   125   50  
   SOS XSECTION HG   80   170   80   80   80   80  
   SOS XSECTION LG   80   170   105   150   125   50  
   HW 1   56   135   150   160   100   125  
   HW 2   56   135   100   160   160   125  
   HW 3   56   135   75   160   160   125  
   FW*   56   135   70   140   125   125  

Zone 4

   Overall   90   100   75   200   200   100  
   HG   90   100   90   60   60   25  
   LG   90   100   75   300   225   75  

11.5.6 Block Model Parameters

A 3D block model was developed to represent the deposit using a block size of 25 feet x 25 feet x 20 feet. The block model dimensions and model limits are shown in Table 11-9. The coordinate system used for the 3D modelling was based on the local grid system using imperial units of feet. The block model is un-rotated and contains no sub blocking.

Table 11-9: Block Model Parameters Open Pit

  Parameter    Value   
  Base point    6000,7000,5800 (X,Y,Z)   
  Parent block size    25x25x20 (X,Y,Z)   
  Azimuth    0   

 

 

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  Boundary size    9000,7000,3000 (X,Y,Z)   
  Size in Blocks    360x280x150 (X,Y,Z)   
  Sub-blocking    None   

11.5.7 Estimation Domains

The estimation domains used to constrain the mineral resource estimate resulted from the numeric indicator models developed as part of the geologic model as discussed in Section 11.5.1. Figure 11-14 shows an overview of the estimation domains that were used to constrain the mineral resource estimation for the open pit mineral resource estimate.

Figure 11-14: Open Pit Numeric Indicator Models

LOGO

11.5.8 Estimation Parameters

Estimation within mineralized domain boundaries was performed using an inverse distance squared method with a minimum of 4 samples, a maximum of 20 samples, and a drill hole limit of 2. Declustering objects were applied to all high-grade estimation domains. Dynamic anisotropy was applied where it was applicable based on faults that structurally control mineralization. The exception to this was Zone 4, which has no apparent structural control that has yet been mapped. In this case, search ellipse orientation was determined from examining the spatial orientation of the composites greater than 1 g/t.

 

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Search distances of the domained estimators were based on the variography for each sub-domain, as discussed in Section 11.5.4. Search distances for domains that showed poor variography were replaced by the overall sub-domain or overall domain search distances. This method was used for the Zone 4 HG and LG estimations and the Zone 3 sub-domains. All estimations used hard boundaries with the exception of the boundaries between Zone 1 HG and Zone 3 HG and between Zone 1 LG and Zone 4 LG. Soft 40-foot boundaries were set up with filters between these boundaries since they are immediately adjacent to each other and could potentially have continuity in grade estimation across these boundaries. The inverse distance estimation parameters for each domain are given in Table 11-10.

Table 11-10: Open Pit ID2 Estimation Parameters

Zone   Sub-Domain  

Dynamic

 Anisotropy 

   Trend     Major   

Semi-

 Major 

   Minor 

Zone 1

  HG   Yes   CX Fault   160   160   75
  LG   Yes   CX Fault   175   175   60

 Zone 2 

  HG   Yes   MAG Fault   180   125   75
  LG   Yes   MAG Fault   300   200   200

Zone 3

  CX HG   Yes   CX Fault   180   160   70
  CX LG   Yes   CX Fault   180   160   70
  SOS DIKE HG   Yes   SOS Dike   100   80   50
  SOS DIKE LG   Yes   SOS Dike   100   80   50
  SOS XSECTION HG   Yes   SOS Xsection   150   125   50
  SOS XSECTION LG   Yes   SOS Xsection   150   125   50
  HW 1   Yes   CX Fault   160   100   125
  HW 2   Yes   CX Fault   160   160   125
  HW 3   Yes   CX Fault   160   160   125
  FW   Yes   CX Fault   160   160   125

Zone 4

  HG   No   90,100,90   200   200   100
  LG   No   90,100,75   200   200   100

After each domained estimator was constructed, a combined estimator was used to assign a hierarchical value to each domained estimation to produce a single gold grade value. The combined estimator hierarchy is shown in Table 11-11.

Table 11-11: Open Pit Combined Estimator Hierarchy

     Priority   

Domained

Estimation

   Priority   

Domained

Estimation

   
  1   Zone 1 HG   11  

Zone 3 HW 3

 
  2   Zone 1 LG   12  

Zone 3 HW 1

 
  3   Zone 2 HG   13  

Zone 3 HW2

 
  4   Zone 2 LG   14  

Zone 3 FW

 

 

 

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     Priority   

Domained

Estimation

   Priority   

Domained

Estimation

   
  5   Zone 3 CX HG   15  

Zone 4 HG

 
  6   Zone 3 CX LG   16  

Zone 4 LG

 
  7   SOS Dike HG   17  

Zone 1

 
  8   SOS Dike LG   18  

Zone 2

 
  9   SOS Xsection HG   19  

Zone 3

 
  10   SOS Xsection LG   20  

Zone 4

 

The overall domain grade estimations were assigned to priority 17 to 20 so that they would fill in areas of the numeric estimator domains that lacked the required number of samples to be able to estimate grade due to the small volume being estimated that excluded drill holes.

11.5.9 Geometallurgical Modeling

Section 10.510.5 goes into detail of how the gold recovery model was estimated and implemented in the block model.

Cyanide solubility was compared to all available interval information from the drilling data: gold assay, alteration, lithology, depth, etc. From this available data, a principal component analysis, regression tree, and multivariate adaptive regression spline analysis were performed. Using multivariate adaptive regression spline analysis model was created to predict cyanide solubility in different zones using the available drilling data. Heap Leach (HLCH) recovery is determined by plotting the cyanide solubility with the column recovery. The carbon in leach (CIL) recovery equation was determined by plotting the trend with the calculated head grade and CIL recovery.

The input fields required in the recovery equations were added to the block estimations. Then, the recovery equations were applied to the block model for HLCH and CIL recoveries. These recoveries, along with the Whittle inputs from Table 11-15, were used to determine which of the two processes would be applied to each block.

11.6 Open Pit Resource

11.6.1 Block Model Validation

Validation of the estimated block grades for the Granite Creek deposit was completed for each of the estimation domains. The resource block model estimate was validated by:

 

   

Completing a series of visual inspections by comparisons of gold assay and composite grades to estimated block values across the deposit in both horizontal and vertical sections.

 

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Statistical comparison of parameters such as means, quantiles, and variance between 20-foot composites, Nearest Neighbor (NN), Inverse Distance squared (ID2), and Ordinary Kriged (OK) estimators to ensure that the grade estimations are representative of the composites they are based on.

 

   

Comparing average composite sample values with average estimated block grades along east, north, and elevation orientations using swath grade trend plots.

11.6.1.1 Visual Inspection

The model was examined in plan and section views to compare to drill hole locations and grades. Plan views and section views for each of the estimation areas are shown in Figure 11-15 though Figure 11-22. Comparison of the model grade from the assays did not reveal any major discrepancies.

Figure 11-15: Open Pit Zone 1 Visual Comparison Composite to Block Model Grade Plan View

LOGO

 

 

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Figure 11-16: Open Pit Zone 2 Visual Comparison Composite to Block Model Grade Plan View

LOGO

 

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Figure 11-17: Open Pit Zone 3 Visual Comparison Composite to Block Model Grade Plan View

LOGO

 

 

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Figure 11-18: Open Pit Zone 4 Visual Comparison Composite to Block Model Grade Plan View

LOGO

 

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Figure 11-19: Open Pit Zone 1 Section Composites and Block Model Cross Section

LOGO

Figure 11-20: Open Pit Zone 2 Section Composites and Block Model Cross Section

LOGO

 

 

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Figure 11-21: Open Pit Zone 3 Section Composites and Block Model Cross Section

LOGO

 

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Figure 11-22: Open Pit Zone 4 Section Composites and Block Model Cross Section

LOGO

 

11.6.1.2

Statistical Comparison

To ensure that the grade estimations are representative of the composites they are based on and validate the resource estimation results, the block model grade estimation statistics were analyzed. GRE compared the means, quantiles, and variance between 20-foot composites, Nearest Neighbor (NN), Inverse Distance squared (ID2), and Ordinary Kriged (OK) estimators, as shown in Table 11-12. Blocks are confined to the 2000 $/tr oz Whittle pit.

Table 11-12: Open Pit Comparison of Composite Values to Grade Estimation Methods

Parameter    Composites     Parameter     NN    ID2    OK

Count

   66,300     Block Count     216,544     253,284     216,544 

Mean

   0.36     Mean    0.35     0.31     0.36 

SD

   1.95     SD    0.81     1.07     0.78 

CV

   5.47     CV    2.31     3.49     2.19 

Variance

   3.80     Variance    0.66     1.14     0.62 

Minimum

   0.00     Minimum    0.00     0.00     -0.35 

Q1

   0.02     Q1    0.03     0.02     0.03 

Q2

   0.03     Q2    0.08     0.05     0.08 

Q3

   0.09     Q3    0.27     0.09     0.28 

 

 

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Parameter    Composites     Parameter     NN    ID2    OK

Maximum

   94.33     Maximum    32.11     35.00     27.39 

As expected, the NN estimator generates more blocks due to the lack of restrictions of having to use multiple samples and multiple drillholes to estimate a block grade. Both ID2 and OK produced similar quantiles, means, variance, etc. The one marked difference seen between ID2 and OK is that it is possible to have a negative value in Kriging as seen in the minimum value of the OK estimator.

Figure 11-23: Cumulative Frequency of Composite and Block Data

LOGO

 

11.6.1.3

Swath Plots

Swath plots of the various estimation methods (NN, ID2, and OK) were used to compare the results from each estimation method to the composite values and examine which method smoothed the estimated grades. As an example, swath plots are provided below for the Zone 1 high grade and low -grade domains. The swatch plots show a general trend that the ID2 estimator smoothed out drastic swings in grade while not over smoothing local variability.

 

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Figure 11-24: Open Pit Swath Plot X axis, Zone 1 High Grade Domain

LOGO

Figure 11-25: Open Pit Swath Plot Y axis, Zone 1 High Grade Domain

LOGO

 

 

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Figure 11-26: Open Pit Swath Plot Z axis, Zone 1 High Grade Domain

LOGO

11.6.2 Mineral Resource Classification

Block model quantities and grade estimates for the Granite Creek deposit were classified according to the S-K 1300 definitions for Mineral Resources and Mineral Reserves (CFR, 2018).

Mineral resource classification involved a two-step process using minimum distances and minimum numbers of samples to define resource classification initially before applying numeric indicator model to define a more continuous and reasonable resource classification. The criteria used in the first step of the resource classification are listed in Table 11-13. Parameters used for the numeric indicator models for the second step of the resource classifications are listed Table 11-14.

Table 11-13: Open Pit Mineral Resource Classification Parameters

Resource

Class

  

Minimum

Distance

  

Minimum

Number of

samples

Measured    50    7
Indicated    100    5
Inferred    150    NA

 

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Table 11-14: Open Pit Parameters for Resource Class Numeric Indicator Model

 Resource 

Class

   Shape   

 Interpolant 

Distance

  ISO
Measured   Isotropic   250   0.4
Indicated   Isotropic   250   0.4
Inferred   NA   NA   NA

Because the classification was performed across all resource estimation domains, an isotropic search was used along with an interpolant distance of 250, which was based on the average continuity of grade seen in the deposit. No numeric indicator model was constructed for the inferred resource class, rather it was defined as any block with a calculated gold grade that did not fall within the measured or indicated numeric indicator domains that had a calculated gold grade. A plan view of the estimated resource classes is shown in Figure 11-27.

Figure 11-27: Open Pit Constrained Resource Class All Areas Plan View

LOGO

11.6.3 Mineral Resource Statement

S-K 1300 (CFR, 2018) defines a mineral resource as: “a concentration or occurrence of material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. A mineral resource is a reasonable estimate of mineralization, taking into account relevant factors such as cut-off grade, likely mining

 

 

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dimensions, location or continuity, that, with the assumed and justifiable technical and economic conditions, is like to, in whole or in part, become economically extractable…”. The mineral resources may be impacted by further infill and exploration drilling that may result in increase or decrease in future resource evaluations. The mineral resources may also be affected by subsequent assessment of mining, environmental, processing, permitting, taxation, socio-economic, and other factors. Mineral resources are not mineral reserves and do not have demonstrated economic viability. As a result, no mineral reserves have been estimated as part of this study. There is no certainty that all or any part of the mineral resources will be converted into a mineral reserve.

The requirement, “reasonable prospects for economic extraction,” generally implies that the quantity and grade estimates meet certain economic thresholds and that the mineral resources are reported at a cutoff grade considering appropriate extraction scenarios and processing recoveries. To meet this requirement, GRE considered that major portions of the Granite Creek deposit are amenable for open pit extraction.

To determine the quantities of material offering “reasonable prospects for economic extraction” by an open pit, GRE constructed open pit scenarios developed from the resource block model estimate using Whittle’s Lerchs-Grossman miner “Pit Optimizer” software. Reasonable mining assumptions were applied to evaluate the portions of the block model (Measured, Indicated, and Inferred blocks) that could be “reasonably expected” to be mined from an open pit. The optimization parameters presented in Table 11-15 were selected based on experience and benchmarking against similar projects. The results are used as a guide to assist in the preparation of a mineral resource statement and to select an appropriate resource reporting cutoff grade. GRE considers that the blocks located within the resulting conceptual pit envelope show “reasonable prospects for economic extraction” and can be reported as a mineral resource.

Table 11-15: Granite Creek Resource Parameters for Open Pit Optimization

Parameter    Items    Unit    Value
Costs   

Mining Cost

(waste/mineralized

material)

   $/tonne mined    2.46
   Heap Leach1   

$/tonne

mineralized

material treated

   9.04
   Carbon in Leach2   

$/tonne

mineralized

material treated

   17.22

 

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Parameter    Items    Unit    Value
Recovery   

Heap Leach (HLCH)

Recovery with CN

Solubility <60

   %    CN Solubility*100
  

HLCH Recovery

with CN Solubility

>= 60

   %   

((0.1225 * [Au_ppm]) +

0.4164)*100

   CIL Recovery    %   

((0.5388 * CN Solubility) +

0.3201)*100

Net revenue

gold

   Gold price3    $/oz    2,040
  

Selling costs and

penalties4

   $/oz    114
Royalty   

Total royalty

(simplified)

   %    6.00%
Slope angles    Slope Angle    degrees    41
Limits    HLCH    tonnes per year    2,975,000
   CIL    tonnes per year    1,050,000

HLCH and CIL costs include $1.56/tonne milled for admin costs.

Various royalties are applicable at various points throughout the mine life, however for the scope of this iA, GRE has used a single 6% royalty for the open pit mineral resource.

The gold price used for this analysis is the 36-month trailing average gold price as of December 31, 2024.

This selling cost is used to apply the 6% royalty

Due to the large ratio of deposit size to block size and method of grade estimation, the grade model is fully diluted, and the resource is 100% recoverable as estimated.

The Granite Creek open pit mineral resource constrained by a Whittle pit shell that corresponds to a gold price of $2,040 per troy ounce is shown in Table 11-16. The reader is cautioned that the results from the pit optimization are used solely for testing the “reasonable prospects for economic extraction” by an open pit and do not represent an attempt to estimate mineral reserves. There are presently no mineral reserves for the project.

Table 11-16: Granite Creek Open Pit Mineral Resource

Class    Zone   

Total

Process

Material

(1000s

Tonnes)

  

Total

Process
Material

(1000s

Tons)

  

Au Grade

(g/t)

  

Au Grade

(opt)

  

Total
Contained
Au (10000s

t. oz)

Measured

   Pit B    2,910    3,207    1.32    0.042    123.41
   Pit A    563    620    1.07    0.034    19.30
   CX    10,889    12,003    1.30    0.042    455.27

 

 

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Class    Zone   

Total

Process

Material

(1000s

Tonnes)

  

Total

Process
Material

(1000s

Tons)

  

Au Grade

(g/t)

  

Au Grade

(opt)

  

Total
Contained
Au (10000s

t. oz)

     MAG    12,000    13,228    1.21    0.039    467.97
   Total    26,362    29,059    1.26    0.040    1,065.95

Indicated

   Pit B    360    397    1.10    0.035    12.73
   Pit A    689    760    0.80    0.026    17.78
   CX    2,973    3,277    1.25    0.040    119.62
   MAG    7,317    8,066    0.93    0.030    219.16
   Total    11,339    12,499    1.01    0.033    369.29

Measured +

Indicated

   Pit B    3,270    3,604    1.29    0.042    136.14
   Pit A    1,252    1,380    0.92    0.030    37.08
   CX    13,862    15,280    1.29    0.041    574.89
   MAG    19,317    21,293    1.11    0.036    687.13
   Total    37,701    41,558    1.18    0.038    1,435.24

Inferred

   Pit B    32    36    0.64    0.021    0.67
   Pit A    205    226    0.59    0.019    3.88
   CX    1,347    1,485    1.16    0.037    50.24
   MAG    563    620    1.11    0.036    20.17
   Total    2,148    2,367    1.09    0.035    74.95

1) The effective date of the Mineral Resources Estimate is December 31, 2024.

2) The Qualified Person for the estimate is GRE.

3) Mineral resources are not ore reserves and are not demonstrably economically recoverable.

4) Mineral resources are reported at a 0.30 g/t cutoff, an assumed gold price of 2,040 $/tr. oz, using variable recovery, a slope angle of 41 degrees, 6% royalty, heap leach processing cost $9.04 per tonne (includes admin), CIL processing cost of $17.22 per tonne (includes admin).

11.6.4 Calculation of Cutoff Grade

The cutoff grade of 0.30 ppm used for the Mineral Resource Statement was calculated as the maximum of the following:

 

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Heap leach cutoff grade:

 

Process & G&A Cost:

  

$9.04

Cost @ 70% Recovery (average heap leach recovery): $12.91/tonne

Au Price:

  

$2,040/Au oz

Economic Cutoff grade

  

$12.91/$2040 = $0.00633 oz/tonne

  

= 0.20 g/tonne or ppm

CIL cutoff grade:

 

Process & G&A Cost:

  

$17.22

Cost @ 85% Recovery (average CIL recovery):    $20.85/tonne
Au Price:    $2,040/Au oz
Economic Cutoff grade    $20.85/$2040 = $0.010 oz/tonne
   = 0.30 g/tonne or ppm

These calculated cutoff grades represent marginal cutoff grades. Mining costs are not included because they are applied during pit optimization.

11.6.5 Sources of Uncertainty

Sources of uncertainty in the mineral resource estimate are described below.

 

  a)

Data: As noted in Section 11.2, a number of negative, missing, and blank assay values exist in the drill hole data files. These values were replaced as shown in Table 11-1. GRE followed standard practice in this regard; however, these intervals represent some uncertainty but are believed to not have a significant impact on the resource estimate.

 

  b)

Geologic Model: The geologic model was prepared by a QP geologist and represents GRE’s understanding of the project geology as of the effective date of this report; however, the geologic model may evolve as more data becomes available. This may have an impact on the resource estimation in the future.

 

  c)

Classification Criteria: GRE followed standard practice in determining Measured, Indicated, and Inferred resources; however, the interpretations may evolve as more data becomes available.

 

  d)

Grade Interpolation: GRE interpolated grades following completion of a variography analysis of the data. The variogram ellipsoid directions and ranges could evolve as more data becomes available, which could impact the resource estimation in the future.

 

  e)

Parameters for Open Pit Optimization: GRE used mining, processing, G&A, royalty, and selling costs from the 2021 Technical Report (GRE, 2021) and the 36-month trailing average gold price as of December 31, 2024. Changes in any of the values, for example from potential tariffs and trade wars increasing operating costs, would alter the resulting Lerchs-Grosman pit shell and the resource reporting within that pit shell.

 

 

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  f)

Cutoff Grade Calculation: GRE used processing and G&A costs, average royalty, and average recoveries from the 2021 Technical Report (GRE, 2021) to calculate economic cutoff grades. Changes in any of the values would alter the cutoff grade.

 

  g)

Metal Price: Use of the 36-month trailing average gold price is deemed reasonable. It is likely that acceptable gold prices will change over time and could be influenced by shifts in the economy.

11.6.6 Mineral Resource Sensitivity

Table 11-17 shows the sensitivity of the mineral resource to cutoff grade in each domain.

Table 11-17: Granite Creek Mineral Resource Sensitivity to Cutoff Grade

 

Deposit  

 Cutoff 

Grade

(ppm)

   

Mass

 (1000sn 

tonnes)

   

Mass

 (1000s tons) 

   

Au

 Grade 

(g/t)

   

Au

 Grade 

(opt)

   

Au

 Contained 

(million

tr oz)

 
   

             

Pit B

    0.1       4,655       5,132       0.89       0.029       0.134  
    0.15       3,994       4,402       1.02       0.033       0.131  
    0.2       3,543       3,905       1.13       0.036       0.128  
    0.25       3,201       3,529       1.22       0.039       0.126  
    0.3       2,910       3,207       1.32       0.042       0.123  
    0.35       2,662       2,935       1.41       0.045       0.121  
    0.4       2,476       2,729       1.49       0.048       0.119  
    0.45       2,283       2,517       1.58       0.051       0.116  
    0.5       2,098       2,312       1.68       0.054       0.113  

Pit A

    0.1       1,052       1,160       0.66       0.021       0.022  
    0.15       890       981       0.75       0.024       0.022  
    0.2       762       839       0.85       0.027       0.021  
    0.25       646       712       0.96       0.031       0.020  
    0.3       563       620       1.07       0.034       0.019  
    0.35       486       536       1.18       0.038       0.019  
    0.4       424       467       1.30       0.042       0.018  
    0.45       370       407       1.43       0.046       0.017  
    0.5       320       353       1.58       0.051       0.016  

 

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Deposit  

 Cutoff 

Grade

(ppm)

   

Mass

 (1000sn 

tonnes)

 

Mass

 (1000s tons) 

 

Au

 Grade 

(g/t)

 

Au

 Grade 

(opt)

 

Au

 Contained 

(million

tr oz)

CX

    0.1     17,873   19,702   0.86   0.028   0.495
    0.15     14,942   16,470   1.01   0.032   0.483
    0.2     13,162   14,508   1.12   0.036   0.473
    0.25     11,837   13,048   1.22   0.039   0.464
    0.3     10,889   12,003   1.30   0.042   0.455
    0.35     10,083   11,114   1.38   0.044   0.447
    0.4     9,401   10,363   1.45   0.047   0.439
    0.45     8,787   9,686   1.52   0.049   0.430
    0.5     8,208   9,048   1.60   0.051   0.421

Mag

    0.1     16,755   18,469   0.92   0.030   0.497
    0.15     15,088   16,631   1.01   0.032   0.490
    0.2     14,006   15,439   1.07   0.035   0.484
    0.25     12,970   14,297   1.14   0.037   0.477
    0.3     12,000   13,228   1.21   0.039   0.468
    0.35     11,133   12,272   1.28   0.041   0.459
    0.4     10,263   11,313   1.36   0.044   0.448
    0.45     9,472   10,442   1.44   0.046   0.438
    0.5     8,807   9,708   1.51   0.049   0.427

Pit B

    0.1     1,291   1,423   0.44   0.014   0.018
    0.15     931   1,027   0.56   0.018   0.017
    0.2     700   772   0.68   0.022   0.015
    0.25     499   550   0.87   0.028   0.014
    0.3     360   397   1.10   0.035   0.013
    0.35     284   313   1.31   0.042   0.012
    0.4     229   252   1.53   0.049   0.011
    0.45     208   229   1.64   0.053   0.011
    0.5     183   201   1.81   0.058   0.011

Pit A 

    0.1     1,076   1,186   0.58   0.019   0.020
    0.15     936   1,032   0.65   0.021   0.019
    0.2     821   905   0.71   0.023   0.019

 

 

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Deposit  

 Cutoff 

Grade

(ppm)

   

Mass

 (1000sn 

tonnes)

 

Mass

 (1000s tons) 

 

Au

 Grade 

(g/t)

 

Au

 Grade 

(opt)

 

Au

 Contained 

(million

tr oz)

      0.25     754   831   0.76   0.024   0.018
    0.3     689   760   0.80   0.026   0.018
    0.35     618   681   0.86   0.028   0.017
    0.4     535   590   0.93   0.030   0.016
    0.45     467   514   1.01   0.032   0.015
    0.5     411   453   1.08   0.035   0.014

CX

    0.1     6,222   6,858   0.69   0.022   0.137
    0.15     4,800   5,291   0.85   0.027   0.132
    0.2     3,898   4,297   1.01   0.033   0.127
    0.25     3,301   3,639   1.15   0.037   0.123
    0.3     2,973   3,277   1.25   0.040   0.120
    0.35     2,740   3,020   1.33   0.043   0.117
    0.4     2,532   2,791   1.41   0.045   0.115
    0.45     2,329   2,567   1.49   0.048   0.112
    0.5     2,188   2,412   1.56   0.050   0.110

Mag

    0.1     11,982   13,208   0.64   0.021   0.247
    0.15     10,312   11,367   0.72   0.023   0.240
    0.2     9,142   10,078   0.79   0.026   0.234
    0.25     8,161   8,996   0.86   0.028   0.227
    0.3     7,317   8,066   0.93   0.030   0.219
    0.35     6,584   7,257   1.00   0.032   0.212
    0.4     5,939   6,546   1.07   0.034   0.204
    0.45     5,260   5,799   1.15   0.037   0.194
    0.5     4,616   5,089   1.24   0.040   0.185

Pit B 

    0.1     61   68   0.41   0.013   0.001
    0.15     44   48   0.54   0.017   0.001
    0.2     42   46   0.55   0.018   0.001
    0.25     38   42   0.58   0.019   0.001
    0.3     32   36   0.64   0.021   0.001
    0.35     25   27   0.74   0.024   0.001
    0.4     23   25   0.77   0.025   0.001

 

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Deposit  

 Cutoff 

Grade

(ppm)

   

Mass

 (1000sn 

tonnes)

 

Mass

 (1000s tons) 

 

Au

 Grade 

(g/t)

 

Au

 Grade 

(opt)

 

Au

 Contained 

(million

tr oz)

      0.45     23   25   0.77   0.025   0.001
      0.5     21   23   0.80   0.026   0.001

Pit A

    0.1     457   503   0.37   0.012   0.005
    0.15     376   414   0.42   0.014   0.005
    0.2     327   360   0.46   0.015   0.005
    0.25     263   290   0.52   0.017   0.004
    0.3     205   226   0.59   0.019   0.004
    0.35     175   193   0.63   0.020   0.004
    0.4     145   160   0.69   0.022   0.003
    0.45     123   135   0.74   0.024   0.003
    0.5     105   115   0.78   0.025   0.003

CX

    0.1     2,694   2,969   0.66   0.021   0.058
    0.15     2,093   2,307   0.82   0.026   0.055
    0.2     1,701   1,875   0.97   0.031   0.053
    0.25     1,491   1,643   1.07   0.035   0.051
    0.3     1,347   1,485   1.16   0.037   0.050
    0.35     1,221   1,346   1.25   0.040   0.049
    0.4     1,127   1,243   1.32   0.042   0.048
    0.45     1,055   1,163   1.38   0.044   0.047
    0.5     984   1,085   1.44   0.046   0.046

Mag

    0.1     1,486   1,638   0.54   0.018   0.026
    0.15     1,243   1,370   0.63   0.020   0.025
    0.2     1,031   1,137   0.72   0.023   0.024
    0.25     781   861   0.88   0.028   0.022
    0.3     563   620   1.11   0.036   0.020
    0.35     488   538   1.24   0.040   0.019
    0.4     444   489   1.32   0.042   0.019
    0.45     430   474   1.35   0.043   0.019
    0.5     425   469   1.36   0.044   0.019

 

11.7 Underground Mineral Resources

Practical Mining LLC estimated the Granite Creek underground Mineral Resource using all drilling and geological data available through March 29, 2023.

 

 

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11.7.1 Structural and Mineralized Grade Shell Modelling

The Granite Creek structural model includes 11 major and 46 minor faults. Underground mineralization is controlled by seven of the major faults. The range front fault separates the cretaceous granodiorite to the northwest from the Ordovician Upper and Lower Comus formation. The underground mineralization is hosted almost entirely within the Lower Comus.

Underground mineralization is contained within the fault zones and strikes north easterly with sub vertical dip to the southeast. It is subdivided into the CX, Otto, Ogee and South Pacific Zones. The zones are defined by 0.10 opt grade shells trending parallel to the fault orientations (Figure 11-28).

 

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Figure 11-28 Major Faulting and Underground 0.10 opt Grade Shells

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11.7.1.1 Drill Data and Compositing

 

  1.1.6.1.1.

Drill Data Set

The drilling data set within the bock model boundary consists of 2,346 drillholes. Of these 93 were excluded due to collar location discrepancies, downhole survey errors or suspected downhole contamination. Table 11-18 summarizes the drilling within the block model extents by operator and hole type.

Table 11-18 Summary of Drilling Within Block Model Extents

COMPANY

   Count    Hole Type    Length (ft)

Atna

   55    Core    19,730

Atna

   197    RC    47,890

Atna

   51    RC/Core    44,886

Barrick

   1    Rotary    1,000

Barrick

   1    RC/Rotary    1,940

Barrick

   44    RC/Core    49,532

Barrick

   71    Core    30,843

Barrick

   34    RC    24,455

Barrick

   1    Monitor    1,340

PMC

   316    Rotary    99,315

PMC

   15    Monitor    8,635

PMC

   1,151    RC    505,759

PMC

   28    RC/Core    45,748

PMC

   4    Unknown    1,797

PMC

   5    Core    5,205

i-80

   50    RC/Core    70,028

i-80

   184    Core    87,159

i-80

   16    RC    2,595

Unknown

   15    Monitor    7,415

Unknown

   2    RC    465

Unknown

   4    Pump    3,740

Unknown

   5    Core    0

Unknown

   3    Met    1,535
    

         

Total

   2,253         1,061,012

 

 

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11.7.1.2

Compositing

Drill holes were composited into ten-foot lengths starting and ending with the 0.004 or 0.10 opt grade shells. The grade shells act as hard boundaries and only composites within the shell are used to estimate grades within a particular shell.

 

11.7.1.3

Statistics

Univariate statistics were calculated for the 0.004 and 0.20 Au opt grade shell. The results are shown in Table 11-19 and Figure 11-29 and Figure 11-30.

Table 11-19 Composite Statistics

Shell    # Comps    Min    Max    Mean    Std Dev    CV
au004    16452    0.0001    19.83    0.030    0.208    6.861
au1    2631    0.0001    3.780    0.317    0.334    1.054

 

 

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Figure 11-29 Histogram of 0.004 Au opt Composites

 

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Figure 11-30 Histogram of 0.10 Au opt Composites

 

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11.7.1.4

Density

The drill database contains 419 density measurements collected by past operators of the Granite Creek Project. These values are summarized by lithology in Table 11-20.

Table 11-20 Density Values Used in the Underground Model

Lithologic Unit

  

Density (tons/ft3)

Qal

  

0.0578

Cp

  

0.0826

Ocl

  

0.0820

Ocu

  

0.0814

Kgd

  

0.0819

 

11.7.1.5

Block Model

Block model blocks are 20x20x20 feet with sub-blocking to 5x5x5 feet at the 0.004 and 0.10 Au opt grade shells. The model extends across the Ogee, Otto, South Pacific and CX zones.

 

11.7.1.6

Grade Capping

Grade capping as applied to the 0.10 and 0.004 Au opt composites. Composite grades exceeding the cap value within the 0.004 Au opt grade shells are restricted for use only within the 20x20x20 foot block that contains the capped composite and it is not used to estimate any grades beyond the containing block. For the 0,10 opt grade shells the capped composite is constrained for use within the 10x10x10 foot block. Cap Grades are listed in Table 11-21.

Table 11-21 Underground Grade Capping Values

Grade

Shell

  

Cap Grade

Au opt

  

# Comps

Affected

0.004    0.1    248
0.10    1.32    48

 

11.7.1.7

Grade Estimation and Resource Classification

Block grades were estimated using the Inverse Distance Cubed (ID3) method. Anisotropic search parameters were set to the general orientation of the grade shells as shown in Table 11-22.

 

 

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The required number of composites to classify a block as mineral resource must come from at least two drill holes. A block cannot be classified on the basis of only one drill hole. The mineral resource classification parameters are shown in Table 11-23.

Table 11-22 Ellipsoid Search Parameters

Shell    Bearing         Plunge    Dip
ogee004    67         0    -90
otto004    40         0    -63
spz004    33         0    -56
cx004_01    52         0    -40
cx004_02    0         0    0
cx004_03    0         0    0
cx004_04    30         0    0
cx004_05    90         0    0
cx004_06    90         0    -25
cx004_07    90         0    -28
cx004_08    90         0    -15
cx004_09    90         0    -50
cx004_10    90         0    -55
au1_02    0         0    0
cx1_01    52         0    -42
cx1_02    58         -22    -21
cx1_03    90         -52    -35
cx1_04    90         0    -52
cx1_05    90         0    -70
cx1_06    90         0    -53
cx1_07    90         0    0
cx1_08    0         0    0
cx1_09    90         0    -55
cx1_10    0         0    -13
cx1_11    90         0    -57
cx1_12    90         0    -68
cx1_13    90         0    0
cx1_14    90         0    -15
cx1_15    90         0    -27
cx1_16    90         0    -37
ogee1    67         0    90
otto1    40         0    -63
spz1    33         0    -56

Table 11-23 Resource Classification Parameters

Class    Major (ft)    Semi (ft)    Minor (ft)   

Min

Samp

  

Max

Samp

  

Min.

DH

meas    75    75    37.5    8    16    2
ind    150    150    75    6    12    2
inf    300    300    150    4    12    2

 

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11.7.1.8

Mined Depletion and Sterilization

The fraction of Blocks that intersect the mined volume survey or surface topography is calculated and the block tonnage adjusted accordingly.

11.7.2 Model Validation

 

11.7.2.1

Visual Comparison

Cross sections showing modelled block grades and drill hole composites provide a visual comparison in a localized area. Comparative cross sections showing blocks greater than 0.10 Au opt are shown in Figure 11-31 through Figure 11-33.

Figure 11-31 Comparative Cross Section Through Otto and Ogee Zones

 

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Figure 11-32 Comparative Cross Section Through the CX Zone

 

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Figure 11-33 Comparative Cross Section through the South Pacific Zone

 

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11.7.2.2

Drift Analysis

Drift analysis or swath plots graphically compare drill hole composite grades to model grades in a specified slice direction and thickness across the modelled extents. Modelled grades should closely follow drilling composite grades. Drift analysis for 100-foot slices are presented in Figure 11-34 and Figure 11-35.

 

 

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Figure 11-34 Easterly Drift Analysis

 

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Figure 11-35 Elevation Drift Analysis

 

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11.7.2.3

Reconciliation

i-80 splits the heading survey at the recorded face distance for each round mined. These then correspond to muck assays and grade control tracking. All mineralization determined to be refractory with muck samples above 0.058 opt and oxide material above 0.075 opt is sent to NGM for processing. Low grade oxide material between 0.020 and 0.075 is sent to the Lone Tree heap leach facility. (See Section 13.2.4).

Table 11-24 show the material sent to each process by month in 2024 and Table 11-25 shows the corresponding modeled high grade mineralization contained within the mined volumes and the monthly variance. The monthly ounce variance is shown in Figure 11-36 and the cumulative percentage variance is shown in Figure 11-37.

High grade mill to high grade model reconciliation for all of 2024 shows a 55% increase in tons with a 46% decrease in grade and 16% decrease in ounces. The underperformance of the process with respect to grade and ounces is due to both planned dilution and overbreak. Some high grade mineralization was mixed with low grade mineralization.

Table 11-24 Mineralization Processed in 2024

     Refractory Autoclave    Oxide Carbon in Leach    High Grade Mill    Lone Tree Heap Leach
   Tons   

Au

opt

   Au oz    Tons   

Au

opt

   Au oz    Tons   

Au

opt

   Au oz    Tons   

Au

opt

   Au oz

Jan

   6,812    0.298    2,027    3,778    0.255    963    10,590    0.282    2,990    9,804    0.109    1,069

Feb

   11,135    0.238    2,655    3,294    0.195    642    14,429    0.228    3,297    5,663    0.110    623

Mar

   4,633    0.185    856    3,062    0.310    948    7,695    0.234    1,804    3,783    0.097    365

Apr

   2,291    0.154    352    2,040    0.278    567    4,331    0.212    919    7,592    0.100    757

May

   726    0.130    94    2,067    0.226    467    2,793    0.201    561    11,190    0.102    1,140

Jun

   939    0.235    221    5,019    0.251    1,260    5,958    0.249    1,481    18,201    0.115    2,084

Jul

   2,641    0.183    483    5,978    0.243    1,453    8,619    0.225    1,935    24,069    0.084    2,027

Aug

   1,928    0.126    244    3,803    0.251    955    5,731    0.209    1,198    25,791    0.109    2,810

Sep

   3,815    0.261    994    5,025    0.211    1,060    8,840    0.232    2,054               

Oct

   4,338    0.214    928    5,036    0.230    1,158    9,374    0.223    2,086               

Nov

                  3,342    0.276    922    3,342    0.276    922               

Dec

                  4,853    0.243    1,179    4,853    0.243    1,179               

Total

   39,257    0.225    8,852    47,297    0.245    11,575    86,553    0.236    20,427    106,093    0.103    10,875

 

 

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Table 11-25 2024 High Grade Block Model Predicted and High Grade Mill - Model Variance

      Block Model    High Grade Mill – Model
Variance
  Percentage Variance
   Tons    Au opt    Au oz    Tons   Au opt   Au oz   Tons   Au opt   Au oz

Jan

   6,512    0.500    3,253    4,078   (0.217)   (263)   63%   -43%   -8%

Feb

   2,676    0.312    834    11,753   (0.083)   2,462   439%   -27%   295%

Mar

   6,214    0.346    2,148    1,480   (0.111)   (343)   24%   -32%   -16%

Apr

   1,848    0.329    607    2,483   (0.116)   312   134%   -35%   51%

May

   3,220    0.369    1,188    (427)   (0.168)   (626)   -13%   -46%   -53%

Jun

   5,525    0.509    2,810    432   (0.260)   (1,329)   8%   -51%   -47%

Jul

   3,717    0.668    2,482    4,901   (0.443)   (547)   132%   -66%   -22%

Aug

   5,002    0.458    2,291    729   (0.249)   (1,092)   15%   -54%   -48%

Sep

   5,686    0.330    1,878    3,153   (0.098)   176   55%   -30%   9%

Oct

   3,177    0.526    1,672    6,198   (0.304)   415   195%   -58%   25%

Nov

   5,658    0.410    2,319    (2,316)   (0.134)   (1,397)   -41%   -33%   -60%

Dec

   6,554    0.430    2,815    (1,701)   (0.187)   (1,636)   -26%   -43%   -58%

Total

   55,789    0.435    24,296    30,764   (0.200)   (3,870)   55%   -46%   -16%

Figure 11-36 Monthly High Grade Mill to Model Au Ounce Variance in 2024

 

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Figure 11-37: 2024 Cumulative High Grade Mill to Model Variance

 

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Reconciliation of all processed material to the model is shown in Table 11-26 and graphically in Figure 11-38 and Figure 11-39. Compared to the high grade reconciliation the model performed much better, processed tons were 5% higher than modeled and processed grade was 15% higher. The model underestimated ounces by 19%.

These results indicate that:

 

   

High grade mineralization is mixed with low grade mineralization, and;

 

   

A significant amount of low grade mineralization is being mined beyond the model limits.

Table 11-26 All Block Model Predicted and Mill - Model Variance in 2024

      Block Model    All Mill –Model Variance   Percentage Variance
   Tons    Au opt    Au oz    Tons   Au opt   Au oz   Tons   Au opt   Au oz

Jan

   19,687    0.174    3,417    707   0.025   641   4%   15%   19%

Feb

   9,791    0.097    949    10,302   0.098   2,970   105%   101%   313%

Mar

   13,733    0.166    2,278    (2,256)   0.023   (109)   -16%   14%   -5%

Apr

   10,304    0.071    728    1,619   0.070   948   16%   99%   130%

May

   11,772    0.133    1,562    2,210   (0.011)   139   19%   -8%   9%

Jun

   17,233    0.179    3,091    6,925   (0.032)   474   40%   -18%   15%

 

 

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      Block Model    All Mill –Model Variance   Percentage Variance
   Tons    Au opt    Au oz    Tons   Au opt   Au oz   Tons   Au opt   Au oz

Jul

   15,558    0.174    2,705    17,130   (0.053)   1,257   110%   -30%   46%

Aug

   16,503    0.154    2,538    15,019   (0.027)   1,470   91%   -17%   58%

Sep

   14,063    0.154    2,172    (5,223)   0.078   (118)   -37%   50%   -5%

Oct

   9,958    0.178    1,772    (583)   0.045   314   -6%   25%   18%

Nov

   15,094    0.168    2,533    (11,752)   0.108   (1,610)   -78%   64%   -64%

Dec

   15,339    0.195    2,989    (10,486)   0.048   (1,810)   -68%   25%   -61%

Total

   169,035    0.158    26,734    23,611   0.004   4,568   4%   15%   19%

Figure 11-38 2024 Monthly All Processed to Model Au Ounce Variance

 

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Figure 11-39 2024 Cumulative All Processed to Model Variance

 

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11.7.2.4

Factors that May Affect Mineral Resource

Areas of uncertainty that may materially impact the Mineral Resource Estimates include:

• Changes to long term metal price assumptions.

• Changes to the input values for mining, processing, and G&A costs to constrain the estimate.

• Changes to local interpretations of mineralization geometry and continuity of mineralized domains.

• Changes to the density values applied to the mineralized zones.

• Changes to metallurgical recovery assumptions.

• Variations in geotechnical, hydrogeological and mining assumptions.

• Changes to assumptions with an existing agreement or new agreements.

• Changes to environmental, permitting, and social license assumptions.

• Logistics of securing and moving adequate services, labor, and supplies could be affected by epidemics, pandemics and other public health crises.

 

 

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11.7.3 Reasonable Prospects for Economic Extraction

SK 1300 requires mineral resources demonstrate “Reasonable Prospects for Economic Extraction” (RPEE). Stope optimizer software is well suited to meet this requirement. The software will produce stope designs that meet minimum minable geometric shapes that exceed the cutoff grade. These shapes will include necessary low grade or waste dilution required to produce a minable geometry.

Granite Creek mineral resources are defined by a mining geometry consistent with the drift and fill mining method. The dimensions of a minimum minable stope cross section are 15 feet wide x 15 feet high. Individual stope lengths vary from a minimum of 20 feet to a maximum of 100 feet.

11.7.4 QP Opinion

Practical Mining is not aware of any environmental, legal, title, taxation, socioeconomic, marketing, political, or other relevant factors that would materially affect the estimation of Mineral Resources that are not discussed in this Technical Report Summary.

Practical Mining is of the opinion that the Mineral Resources for the Project, which were estimated using industry accepted practices, have been prepared and reported using S-K 1300 definitions.

Technical and economic parameters and assumptions applied to the Mineral Resource Estimate are based on parameters received from i-80 and reviewed by Practical Mining to determine if their appropriateness.

The QP considers that all issues relating to all relevant technical and economic factors likely to influence the prospect of economic extraction can be resolved with further work.

11.7.5 Underground Mineral Resources

Mineral Resources for the Granite Creek underground mine are summarized in Table 11-27.

Table 11-27 Summary of Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

Zone      ktons      ktonnes      Au opt      Au g/t      Au koz
Measured

Ogee

     88      80      0.244      8.4      22

Otto

     59      53      0.256      8.8      15

Meas Total

     147      133      0.249      8.5      37
Indicated

CX

     8      7      0.391      13.4      3

Ogee

     181      164      0.352      12.1      64

 

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Zone      ktons      ktonnes      Au opt      Au g/t      Au koz

Otto

     295      268      0.316      10.8      93

South Pacific

     223      203      0.286      9.8      64

Ind Total

     707      641      0.317      10.9      224
Measured and Indicated

CX

     8      7      0.391      13.4      3

Ogee

     269      244      0.317      10.9      85

Otto

     354      321      0.306      10.5      108

South Pacific

     223      203      0.286      9.8      64

M&I Total

     854      775      0.305      10.5      261
Inferred

CX

     97      88      0.351      12.0      34

Ogee

     42      38      0.563      19.3      24

Otto

     187      170      0.401      13.7      75

South Pacific

     536      486      0.361      12.4      194

Inf Total

     862      782      0.378      13.0      326

Notes Pertaining to Underground Mineral Resources:

 

  1.

Mineral Resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of $27.25 per ounce. Refer to Section 16.1 for price selection details.

 

  2.

Mineral Resources have been estimated using gold metallurgical recoveries of 85.2% to 94.2% for pressure oxidation. Payment for refractory mineralization sold to a third party is 58%. Oxide CIL mineralization payments vary from 40% to 70% based upon the grade of the mineralization.

 

  3.

The cutoff grade for refractory Mineral Resources varies from 0.151 to 0.184 opt. for acidic conditions. The cutoff grade for oxide mineral resources is 0.075 opt.

 

  4.

The contained gold estimates in the Mineral Resource table have not been adjusted for metallurgical recoveries.

 

  5.

Numbers have been rounded as required by reporting guidelines and may result in apparent summation differences.

 

  6.

A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

 

  7.

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

 

  8.

Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors.

 

  9.

Mineral Resources have an effective date of December 31, 2024, and;

 

  10.

The reference point for mineral resources is in situ.

 

 

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12 Mineral Reserve Estimates

The Granite Creek Project does not have any mineral reserves.

 

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13 Mining Methods

 

13.1

Open Pit

 

13.1.1

Introduction

The Granite Creek Mine Project will employ conventional open pit mining techniques using front end loaders and rear dump rigid frame haul trucks. As discussed in Section 17, open pit material will be treated using CIL circuit. The mine plan is designed to deliver an average of 10,000 tonnes of potentially economically viable material per day from the open pit to the crusher which will then be run through the CIL mill. The average daily waste production rate over the life of the mine is 84,750 tonnes per day. Waste material would be either placed on waste rock storage facilities (WRSF) or as backfill in previously mined open pits.

There are three distinct open pit production areas on the project: B pit, CX-A pit, and Mag pit. The CX and Mag pits were each designed with three phases, for a total of seven mining phases for the project.

13.1.2 Whittle Pit Shell Analysis

Whittle pit shell analysis was used to provide a basis for creating the pit designs. The objective of the Whittle pit optimization was to maximize the economic extraction of the mineral resources contained in the block model. The inputs used to develop the Whittle pit shell analysis are listed in Table 13-1.

Table 13-1: Granite Creek Open Pit Mine Project Whittle Pitshell Analysis Parameters

Parameter    Items    Unit    Value
Costs   

Mining Cost

(waste/mineralized material)

   $/tonne mined    2.46
   Heap Leach*    $/tonne mineralized material treated     9.04
   Carbon in Leach**    $/tonne mineralized material treated     17.22
Recovery   

HLCH Recovery

CN Solubility <60

   %    CN Solubility*100
  

HLCH Recovery

CN Solubility >= 60

   %   

((0.1225 * [Au_ppm]) +

0.4164)*100

 

 

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Parameter    Items    Unit    Value
    

CIL Recovery

   %   

((0.5388 * CN Solubility) +

0.3201)*100

Net Revenue Gold    Gold price    $/oz    2,040
   Selling costs and penalties***    $/oz    114
Royalty    Total royalty (simplified)    %    6.00%
Slope angles    Slope Angle    degrees    41
Limits    HLCH    tonnes per year    2,975,000
   CIL    tonnes per year    1,050,000

* CIL costs include $1.1/tonne milled for admin costs

** Various royalties are applicable at various points throughout the mine life, however for the scope of this IA GRE has used a single 6% royalty for the open pit mineral resource.

*** This selling cost is used to apply the 6% royalty

Due to the large ratio of deposit size to block size and method of grade estimation, the grade model is fully diluted, and the resource is 100% recoverable as estimated.

Revenue factors from 0.245 to 1.47 in 0.049 increments were applied to the base gold price of $2,040 per troy ounce to examine gold prices from $500 to $3,000 in $100 increments. After Whittle pit shells were run, GRE analyzed each resource area by examining the marginal impact on undiscounted cashflow. This analysis examines the impact that each incremental increase in the pit shell has on the undiscounted cashflow divided by the number of tonnes that are processed. GRE examined each pit area and selected a case that gave a local spike in the marginal impact on undiscounted cashflow at a revenue factor equal to or less than the base price of $2,040 as shown in Figure 13-1 through Figure 13-4. Pit shells were not adjusted for overlap of backslopes. Visually the overlap only occurs at the top of the pits. However, the effect of this overlap is relatively small and the results of this analysis can still guide the Whittle pit shell selection.

 

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Figure 13-1 Marginal Impact Undiscounted Cashflow Mag Pit

 

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Figure 13-2 Marginal Impact Undiscounted Cashflow CX Pit

 

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Figure 13-3 Marginal Impact Undiscounted Cashflow Pit B

 

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Figure 13-4 Marginal Impact Undiscounted Cashflow Pit A

 

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A summary of the selected Whittle pit shells used to constrain the resource is listed in Table 13-2: Selected Whittle Pit Shells for Resource Areas.

 

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Table 13-2: Selected Whittle Pit Shells for Resource Areas

Area     

   Whittle Pit  

Shell

   

Mag

  17

CX

  16

Pit A

  18

Pit B

  14

13.1.3 Pit Design

Based on previous engineering analysis performed by Golder (Golder Associates, 2014), GRE used a triple bench format consisting of triple 20-foot vertical benches with a horizontal 44-foot catch bench every three vertical benches. The resulting open pit parameters are listed in Table 13-3. In less competent zones of safety, benches will be wider or placed at more frequent intervals to reduce the slope angle.

Table 13-3: Pit Parameters

Pit Design Parameters  

Value

 (degrees) 

    

Max Inter-ramp Angle Hard Rock

  41   

Max Bench Face Angle

  68   

 

 

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Figure 13-5 Cross-Section of Typical Pit Slope

 

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13.1.4 Block Model Coding

 

13.1.4.1

Processing Method

GRE coded each block in the block model with a processing method applicable for the block, either heap leach or CIL, or coded the block as waste based on the economic value of each block. The economic value of the block is calculated as:

 

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The selling price, royalty, selling cost, mining cost, processing cost, and admin cost used in the equation were the preliminary values input into the Whittle analysis, as summarized in Table 13-1. For each block, the economic value for the heap leach and CIL were evaluated using the above equation. Blocks were coded to be processed through CIL operation if the economic value for the block was positive and higher than the heap leach operation. Blocks were coded to be processed through heap leach operation if the economic value for the block was positive and higher than the CIL operation. Remaining blocks were coded as waste material.

 

13.1.4.2

Recovery

GRE coded each block in the block model with three potential recoveries:

 

   

Multi-process recovery: each block coded with a CIL process method was coded with its calculated CIL recovery and each block coded with a heap leach process method was coded with its calculated heap leach recovery

 

   

Heap Leach only recovery: each block coded with a CIL processing method that was also heap leachable was coded with its calculated heap leach recovery and each block coded with a heap leach processing method was coded with its calculated heap leach recovery

 

   

CIL only recovery: each block coded with a CIL processing method was coded with its calculated CIL recovery and each block coded with a heap leach processing method was coded with its calculated CIL recovery.

13.1.5 Mining Sequence

The proposed mining sequence is based on known engineering information, economic factors, and environmental considerations. The production pits would be sequentially mined with minor overlap of simultaneous production dependent on short term scheduling needs. The proposed mining sequence begins with Pit B and is shown in Table 13-4.

Table 13-4: Summary of Pit Phases

Pit    Start Day        End Day       
Pit B    -129    208   
CX 1    118    618   
CX 2    423    1,077   
CX 3    997    1,969   
MAG A    1,969    2,137   
MAG B    1,918    2,656   
MAG C    2,388    3,103   

 

 

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13.1.6 Base Case

GRE selected the CIL processing at a cutoff grade of 0.85 g/t for high grade material for the base case with a low grade to high grade cutover grade of 0.25 g/t. The resources within the base case pits and phases are shown in Table 13-5.

Table 13-5: Base Case Pit Resource

 Pit/Phase     

High

Grade CIL

Tonnes

(1000s)

    

Low

Grade

CIL

Tonnes

(1000s)

      

Total Tonnes

(1000s)

      

Au Troy

Ounces

Contained

(1000s)

      

Au Grade

(g/t)

       Stripping
Ratio
 

B

       1,740.7          1,883.6          26,846.3          132.0          1.13          6.41  

CX Phase 1

       1,981.1          2,315.5          47,821.5          172.1          1.25          10.13  

CX Phase 2

       2,137.3          2,707.6          60,396.5          169.3          1.09          11.47  

CX Phase 3

       3,465.8          4,037.3          90,929.1          313.1          1.30          11.12  

Mag Phase 1

       489.4          1,498.3          18,423.9          48.3          0.76          8.27  

Mag Phase 2

       4,546.1          1,695.5          51,912.1          291.7          1.45          7.32  

Mag Phase 3

       4,549.7          1,806.7          47,847.8          270.8          1.33          6.53  

Total

       18,910.1          15,944.4          344,177.3          1,397.2          1.25          8.87  

13.1.7 Mine Scheduling

A preliminary mining schedule was generated from the base case pit resource estimate. GRE used the following assumptions to generate the schedule:

 

   

Mining Production Rate: 10,000 tonnes per day

 

   

Mine Operating Days per Week: 7

 

   

Mine Operating Weeks per Year: 52

 

   

Mine Operating Shifts per Day: 2

 

   

Mine Operating Hours per Shift: 12

Pre-stripping of waste was included if waste occurred on a bench that had no corresponding processable material or if the tonnage of waste on a bench exceeded ten times the tonnage of processable material on that bench. The production rate for pre-strip benches was set to 20 times the leach material production rate, or 100,000 tpd. Processable material mined along with pre-stripped waste was placed into stockpiles for later processing.

For all other benches, all waste on a bench was scheduled to be mined over the same duration as the processable material on that bench. This scheduling method resulted in some years with high waste quantities relative to the processable material quantity mined. GRE used pre-stripping and

 

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phasing, as described above, as much as possible to smooth out the production, but the limitations of the scheduling program resulted in some inefficiencies.

For the economic model, the project was scheduled by quarter for any pre-production years and for the first two production years, then by year for the remainder of the mine life. The mining schedule is summarized in Table 13-6 and illustrated in Figure 13-6.

 

 

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Table 13-6: Granite Creek Mine Project Open Pit Base Case Mine Schedule Summary

Pit/Phase    Year -1   Year 1   Year 2   Year 3   Year 4   Year 5   Year 6   Year 7   Year 8   Year 9   Total
High Grade CIL Tonnes (1000s)  
B      103.7        1,637.0        0.0        0.0        0.0        0.0        0.0        0.0        0.0        0.0        1,740.7   
CX Phase 1      0.0       370.1       1,611.0       0.0       0.0       0.0       0.0       0.0       0.0       0.0       1,981.1  
CX Phase 2      0.0       0.0       188.1       1,949.1       0.0       0.0       0.0       0.0       0.0       0.0       2,137.3  
CX Phase 3      0.0       0.0       0.0       96.0       1,228.1       884.1       1,257.6       0.0       0.0       0.0       3,465.8  
Mag Phase A      0.0       0.0       0.0       0.0       0.0       0.0       489.4       0.0       0.0       0.0       489.4  
Mag Phase B      0.0       0.0       0.0       0.0       0.0       0.0       202.1       3,374.2       969.8       0.0       4,546.1  
Mag Phase C      0.0       0.0       0.0       0.0       0.0       0.0       0.0       16.9       2,777.1       1,755.7       4,549.7  
Total      103.7       2,007.0       1,799.1       2,045.2       1,228.1       884.1       1,949.2       3,391.1       3,746.9       1,755.7       18,910.1  
Low Grade CIL Tonnes (1000s)  
B      187.7       1,695.9       0.0       0.0       0.0       0.0       0.0       0.0       0.0       0.0       1,883.6  
CX Phase 1      0.0       808.4       1,507.1       0.0       0.0       0.0       0.0       0.0       0.0       0.0       2,315.5  
CX Phase 2      0.0       0.0       803.4       1,904.2       0.0       0.0       0.0       0.0       0.0       0.0       2,707.6  
CX Phase 3      0.0       0.0       0.0       141.4       2,510.8       647.1       738.0       0.0       0.0       0.0       4,037.3  
Mag Phase A      0.0       0.0       0.0       0.0       0.0       0.0       1,498.3       0.0       0.0       0.0       1,498.3  
Mag Phase B      0.0       0.0       0.0       0.0       0.0       0.0       768.2       894.4       33.0       0.0       1,695.5  
Mag Phase C      0.0       0.0       0.0       0.0       0.0       0.0       0.0       47.5       1,703.3       55.9       1,806.7  
Total      187.7       2,504.3       2,310.5       2,045.5       2,510.8        647.1       3,004.5        941.9       1,736.2       55.9        15,944.4  
Waste Tonnes (1000s)  
B      12,936.7       10,285.4       0.0       0.0       0.0       0.0       0.0       0.0       0.0       0.0       23,222.0  

 

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Pit/Phase    Year -1   Year 1   Year 2   Year 3   Year 4   Year 5   Year 6   Year 7   Year 8   Year 9   Total
CX Phase 1      0.0        24,683.9        18,841.0        0.0        0.0        0.0        0.0        0.0        0.0        0.0        43,524.9   
CX Phase 2      0.0       0.0       29,115.8       26,435.8       0.0       0.0       0.0       0.0       0.0       0.0       55,551.6  
CX Phase 3      0.0       0.0       0.0       9,878.6       35,287.1       31,496.9       6,763.5       0.0       0.0       0.0       83,426.0  
Mag Phase A      0.0       0.0       0.0       0.0       0.0       0.0       16,436.2       0.0       0.0       0.0       16,436.2  
Mag Phase B      0.0       0.0       0.0       0.0       0.0       0.0       27,215.2       16,465.9       1,989.5       0.0       45,670.5  
Mag Phase C      0.0       0.0       0.0       0.0       0.0       0.0       0.0       16,680.8       22,749.2       2,061.4       41,491.4  
Total      12,936.7       34,969.3       47,956.8       36,314.4       35,287.1       31,496.9       50,414.9       33,146.7       24,738.8       2,061.4       309,322.8  
Au Troy Ounces (1000s)  
B      12.3       119.7       0.0       0.0       0.0       0.0       0.0       0.0       0.0       0.0       132.0  
CX Phase 1      0.0       37.1       135.1       0.0       0.0       0.0       0.0       0.0       0.0       0.0       172.1  
CX Phase 2      0.0       0.0       21.1       148.2       0.0       0.0       0.0       0.0       0.0       0.0       169.3  
CX Phase 3      0.0       0.0       0.0       5.6       118.9       68.7       119.9       0.0       0.0       0.0       313.1  
Mag Phase A      0.0       0.0       0.0       0.0       0.0       0.0       48.3       0.0       0.0       0.0       48.3  
Mag Phase B      0.0       0.0       0.0       0.0       0.0       0.0       20.0       212.0       59.6       0.0       291.7  
Mag Phase C      0.0       0.0       0.0       0.0       0.0       0.0       0.0       1.5       170.4       99.0       270.8  
Total      12.3       156.8       156.1       153.8       118.9       68.7       188.2       213.5       230.0       99.0       1,397.2  

 

 

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Figure 13-6 Granite Creek Mine Project Base Case Mine Schedule

 

LOGO

13.1.8 Mine Operation and Layout

Limited facilities for administrative offices, warehouse, and other facilities are present at the site. Other facilities, such as crushing, and CIL plants will need to be constructed.

GRE developed conceptual layouts for the project, including waste dump locations and sizes, leach pad location and size, tailings storage facility, and stockpile locations and sizes. Figure 13-7 illustrates the conceptual project layout with pits, pads, and dumps. Phased site layout plans for the duration of open pit mining are shown in Figure 13-8 through Figure 13-14.

 

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Figure 13-7 Conceptual Project Layout

 

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Figure 13-8 Phased Pit and Site Plan Layout B Pit

 

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Figure 13-9 Phased Pit and Site Plan Layout B Pit & CX Phase 1

 

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Figure 13-10 Phased Pit and Site Plan Layout B Pit & CX Phases 1 & 2

 

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Figure 13-11 Phased Pit and Site Plan Layout B Pit & CX Phases 1, 2, & 3

 

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Figure 13-12 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase1

 

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Figure 13-13 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase 1 & 2

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Figure 13-14 Phased Pit and Site Plan Layout B Pit; CX Phases 1, 2, & 3; and Mag Phase 1, 2, & 3

 

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13.1.8.1

Waste Rock Pile

To store the waste material generated during mining activities, two waste rock piles are proposed. The waste piles would be located south of the CX pit and east of the Mag pit. Additionally, as mining progresses, waste rock would be backfilled in portions of the mined-out B and CX pits. These locations are selected to minimize hauling distances and disturbed acreage. Up to approximately 100 million loose cy (approximately 153 million tonnes) of waste rock would be mined and placed into the waste rock piles and approximately 111 million cy (approximately 157 million tonnes) would be backfilled into mined out pits. Waste rock piles would be engineered to have overall final 3H:1V ultimate slopes.

 

13.1.9

Drilling and Blasting

Fresh mineralized material and waste rock is comprised of a mix of shale, limestone, dolomite, conglomerates, and granodiorite. All of this material would require drilling and blasting prior to excavation. Some areas within the pits to be excavated consist of alluvium or previous backfill; those areas would not require drilling and blasting, except to the extent drill holes are needed for grade control.

Drilling and blasting would employ conventional techniques, which would entail drilling 7-inch diameter blastholes spaced on 18-foot centers. The rock would be blasted with ammonium nitrate fuel oil (ANFO) blasting agent initiated with shock tube, boosters, and nonel blasting caps. Potential noise and dust from blasting is not anticipated to impact the surrounding community due to the project’s remote location far away from residential or commercial structures.

 

13.1.10

Loading and Hauling

The blasted rock or backfill would be loaded with a 17-cy capacity front end loader into 133-tonne capacity haul trucks.

 

13.1.11

Haul Roads

Haulage pit ramps were designed with a minimum width of 90 feet and a maximum gradient of 10 percent. Haul ramps and roads have been designed to accommodate two-way traffic using 133-tonne haul trucks, water diversion ditches, and safety berms. Minor sections of temporary ramping for development purposes may be steeper and narrower. Haulage roads outside of the pit areas would typically be 90 feet wide, and in some areas would be up to 150 feet wide to allow for turning lanes, surface drainage, and separate lanes for auxiliary vehicle traffic. A minimum cross slope of 2% on haul roads will accommodate water drainage.

 

 

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13.1.12

Mining Mobile Equipment

A variety of mobile equipment likely to be used in conducting mining operations is presented in Table 13-7.

Table 13-7: Granite Creek Mine Project Open Pit Mobile Equipment Sizes and Quantities

 

Major Equipment

  

Max Quantity

Loader CAT 993K

  

6

Haul Truck CAT 785D

  

20

Bulldozer CAT D10

  

3

Drill

  

6

Support Equipment

    

Wheel Dozer

  

1

Wheel Loader

  

1

Water Truck

  

2

ANFO Truck

  

1

Lube Truck

  

2

Mechanics Truck

  

2

Grader

  

1

Minor Equipment

    

Small Excavator

  

1

Backhoe

  

1

Small Crane

  

1

Light Plant

  

6

4x4 Pickup

  

10

Equipment sizes and quantities may vary slightly over the life of the mine in response to changes in stripping ratios, haul distances, or other factors.

13.2   Underground

The Granite Creek Mine is operated by a local contractor. Table 13-8 and Table 13-9 list personnel levels and underground equipment provided. The contractor has operated a number of mines in northern Nevada over the past thirty years.

Table 13-8: Contractors Personnel

Supervision / Overhead

  

Count  

  

Superintendent

   1        

 

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Assistant Superintendent

  

2

  

Safety Superintendent

   2   

Master Mechanic

  

1

  

Electrician

   1   

Engineer (Billed on a Day Rate)

  

1     

  

Total Supervision / Overhead

   8   

Rotating Crew Manpower (Per Shift)

       

Shifter

  

1

  

Miners

   4   

Operators - Mucker / Truck / Jam

  

6

  

Shift Mechanic

   2   

Luber/Nipper

  

1

  

Shotcrete

   2   

Batch Plant

  

1

  

Total Crew Manpower 17

   17   

Rotating Crew Total (4 Crews) 68

  

68

  

Day Shift (7 days)

       

Mechanics

   2   

Total Day Shift Crew

  

4

  

Total

   80   

Table 13-9: Contractors Underground Equipment

Granite Creek Equipment List Quantity

ANFO Truck with boom, Eimco 975

  

1

Man basket and ANFO pot

  

1

Single Boom Jumbo, Sandvik D05, backup unit

  

1

Two Boom Jumbo, Epiroc 282

  

1

Two Boom Jumbo, Epiroc M2C, Tunnel Manager

  

1

Mechanized Bolters with Screen Handler, SandvikDS310, Robolt 5, and D05.

  

3

LHD, 4 cubic yard, Sandvik T6

  

1

LHD, 6 cubic yard, Cat R1600G/H

  

3

U/G Articulated Truck, Cat AD30, w/ ejector bodies

  

4

Shotcrete Spray Truck, YMCO 462, SMD design

  

1

Shotcrete Remix Truck, Elmac open top and Normet Utimec 1500

  

1

U/G Fuel/Lube Truck, Getman 644

  

2

i-80 personnel are responsible for:

 

   

Site security;

 

 

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Designation of the proper destination for mined mineralized material;

 

   

Material movement from the portal to appropriate stockpile;

 

   

Screening of mineralized material;

 

   

Operation of the water treatment plant and;

 

   

Contractor oversight.

i-80 fills some of these positions with contract labor as necessary. Table 13-10: i-80 Personnel list i-80 personnel including positions filled with contract labor.

Table 13-10: i-80 Personnel

Description

 

Count

Manager

 

1

Safety/Security

 

6

Geology

 

2

Ore Control Techs

 

4

Engineer

 

1

Surveyor

 

1

Assay

 

3

Equipment Operators

 

4

Labor

 

2

Clerk

 

1

Janitor

 

2

Total

 

27

 

13.2.1 Development

Development drifting is excavated 15-feet wide by 17-feet high to allow room for a large diameter ventilation duct and 30-ton truck. Decline gradient cannot exceed +/- 13%. Ventilation is provided through a series of raises and crosscuts located inside the spiral. Two bored raises totaling 909 ft and 1,382 ft are planned to supplement ventilation in the lower levels of the mine. The life-of-mine development plan is shown in Figure 13-15.

 

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Figure 13-15 Existing (Shaded Blue) and Planned Mine Development

 

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13.2.2 Production

The initial attack ramp to a set of stopes is driven at a +15% grade. Once the initial level is mined and filled the sill of the ramp is excavated to reach the next level. Three levels or 45 vertical feet of mineralization are typically excavated from a single attack ramp. Where sufficient setback distance is available an attack ramp could access four levels.

Underhand drift and fill mining is well suited to the mineralization geometry and ground conditions at Granite Creek. The cross section of individual stope cuts measures 15-feet high by 15-feet wide, large enough for six yd3 Load Haul Dump (LHD. Mining parallel to strike is preferred. Stope cuts are mined sequentially across the mineralized zone until all mineralization above the cutoff grade is extracted. Once a level is mined and backfilled mining initiates on the next level down.

Figure 13-16 Typical 3-Cut Stope and Hanging Wall Attack Ramp

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13.2.3 Ground Support

Primary ground support consists of welded wire mesh and eight-foot Swellex rock bolts with four-foot by four-foot foot spacing. Additional bolts are added to pull the wire tight to the back. Primary support is installed to completely cover the back and to within five feet of the sill. When necessary

 

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additional support may consist of two to three inches of shotcrete, 12-foot super Swellex bolts, or grouted cable bolts up to sixteen feet in length.

Figure 13-17 Primary Ground Support Installation

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13.2.3.1 Backfill

All stopes where mining is planned alongside or below are backfilled with Cemented Rock Fill (CRF). Backfill aggregate is sourced from waste rock mined in the CX-West open pit. Suitable waste rock is any clean rock free of clay. It is crushed to a nominal three-inch maximum size. The aggregate is mixed with cement slurry to produce a mixture containing six to eight percent cement. Backfill is then loaded in a haul truck where it is transported to the stope. There, a modified LHD with an extended boom and push plate affixed to the end will work the CRF until all the void spaces are filled (Figure 13-18). If no future mining is planned alongside or below the stope cut it can be left open or filled with waste rock from development headings.

 

 

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Figure 13-18 Cemented Rock Fill in Adjacent Cut

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13.2.4 Granite Creek Mineralization Control Procedures

The ore control geologist on shift makes an effort to view each active heading. The geologist takes a photo of the face and makes notes.

Muck samples are collected by haul truck drivers at the windrow. Development headings receive one sample per round and are automatically shipped as waste unless otherwise directed by the geologist. For ore headings, a sample is collected at the rate of roughly one sample per two trucks. The driver uses a hand-held sample scoop to fill a sample bag half full (five to ten pounds), walking along the windrow and taking a scoop every five feet, making sure to collect both coarse and fine material. Sample bags have a tag with duplicate bar codes separated by a perforation. The perforated portion of the tag includes space to hand write sample source information including mine level, heading ID and distance, date and shift. The haul truck driver places completed sample bags in a designated location near the mine office trailer. The ore control geologist collects the samples accumulated from the previous day and night shifts and uses the sample tag information to generate a sample submittal for the laboratory and fill the information to the ore control database. The ore control geologist inserts QAQC samples into the sample stream. A contract driver transports muck samples, Over the Road (OTR) truck samples as well as any drill samples to the Lone Tree laboratory once per day.

 

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The Lone Tree lab analyzes samples for Au grade by fire assay and cyanide absorption, sulfide %, TOC, CO3, and preg-rob potential. The results are used to characterize each round as oxide or autoclave refractory, high grade or low grade, or waste (Table 13-11). Assays must be approved by the database administrator before the ore control geologist can enter assay results in the ore control database and flag windrows for routing. The geologist ties color coded flagging associated with the assessed ore type to a lath at the end of the windrow. The windrow can then be moved to the stockpile corresponding to its ore type. The process typically takes three days from mine face mucking to ore type determination and flagging.

High grade oxide and sulfide ores are screened to 3 inches at the stockpile. The minus 3-inch portion is loaded for shipment to the appropriate ore processing location (autoclave or oxide mill). Screened oversize oxide material is placed in the low grade oxide stockpile, which is shipped to the heap leach facility on a low priority basis. Oversized sulfide material is transferred to long term on-site low grade sulfide stockpiles and is not shipped.

Ore is shipped to processing facilities using contract OTR trucks. The truck driver receives a ticket number at the security gate and gives the ticket number to the loader operator at the stockpile. The loader operator loads the first bucket, then spills a small portion of every other bucket into a small sample pile on the ground during the loading process. Once the loaded truck departs, the loader operator collects a sample from the sample pile and labels the bag with an ID associated with the truck ticket number. Sample bags are waterproof to preserve moisture content. The loader operator places the samples at the designated sample location at the end of the shift, where they are collected by the ore control geologist who prepares a laboratory sample submittal and enters the sample information into acQuire. Trucks are weighed near the security gate when departing the mine, and security personnel email a report of truck tons and ticket numbers at the end of the shift.

Table 13-11 Mineralization Routing Criteria

Criteria   

3rd Party Refractory

Autoclave

  

3rd Party Carbon in

Leach

   Lone Tree Heap Leach

Au opt

   0.058    0.075    0.020 <= Au opt <= 0.075

Sulfide

   >= 1%    <= 0.6%    <= 0.6%

Total Organic Content (TOC)

   <= 0.5%    <= 0.50%    <= 0.50%

Carbonate (CO3)

   <= 15%    N/A    N/A

Preg Rob

   <= 40%    <= 40%    <= 40%

Cyanide Solubility

   N/A    >= 50%    >= 50%

 

 

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13.2.5 Mine Production Plan

Individual drift advance rates are estimated to be between six and eight feet per day. The production plan presented in Table 13-12 represents the material mined from the collection of drifts available at any given time and subject to the constraints of people and equipment availability. At least eight drifts mining mineralized material are required to achieve 500 tons per day.

Table 13-12: Annual Production and Development Schedule (Including Inferred Mineral Resources)

 

                     
Calendar Year    2025     2026     2027     2028     2029     2030     2031     2032     2033     Total 
 
Mineralized Material Mined
               

   Total Mineralization Mined (000’s Tons)   212.4   206.7   221.9   242.2   274.4   205.5   167.7   58.5   -   1,589.4
               
     Gold Grade (Ounce/Ton)   0.328   0.394   0.341   0.346   0.324   0.316   0.316   0.354   -   0.339
               
     Contained Gold (000’s Ounces)   69.7   81.4   75.8   83.9   88.9   65.0   53.0   20.7   -   538.4
 

Production Mining

               
     Stope Development and Drift and Fill Mining (000’s Tons)   212.4   206.7   221.9   242.2   274.4   205.5   167.7   58.5   -   1,589.4
               
     Mineralization Production Rate (tpd)   582   566   608   662   752   563   460   160   -   435
 

Backfill

               
     Total CRF Backfill (000’s Tons)   212.4   206.7   221.9   242.2   274.4   205.5   167.7   58.5   -   1,589.4
 

Waste Mining

               
     Expensed Waste (000’s Tons)   115.0   99.9   111.3   110.8   124.3   89.5   75.6   28.0   -   754.4
               
     Primary Capital Drifting (Feet)   6,675   7,913   5,233   1,694   -   -   -   -   -   21,515
               
     Capital Raising (Feet)   1,050   640   180   180   150   150   -   -   -   2,350
               
     Capitalized Mining (000’s Tons)   180   201.6   116.4   38.5   1.1   1.1   -   -   -   539
             

Total Tons Mined (000’s Tons)

  507.6   508.2   449.6   391.6   399.7   296.1   243.4   86.5   -   2,882.6
             

Mining Rate (tpd)

  1,391   1,392   1,232   1,070   1,095   811   667   236   -   658

 

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Fifty percent of the mineral resource tons are inferred and 56% of the contained gold is inferred. The mine production plan without inferred mineral resources presented in Table 13-13 is a gross factorization of the mine plan containing inferred mineral resources. No adjustments have been made to capital development, productivities, or unit costs.

Table 13-13: Annual Production and Development Schedule (Excluding Inferred Mineral Resources)

 

                     
Calendar Year    2025     2026     2027     2028     2029     2030     2031     2032     2033     Total 
 
Mineralized Material Mined
               

   Total Mineralization Mined (000’s Tons)   105.7   102.9   110.5   120.5   136.5   102.3   83.5   58.5   -   820.3
               
     Gold Grade (Ounce/Ton)   31.0   36.2   33.7   37.3   39.5   28.9   23.6   9.2   -   0.292
               
     Contained Gold (000’s Ounces)   31.0   36.2   33.7   37.3   39.5   28.9   23.6   9.2   -   239.4
 

Production Mining

               
     Stope Development and Drift and Fill Mining (000’s Tons)   105.7   102.9   110.5   120.5   136.5   102.3   83.5   29.1   -   791.0
               
     Mineralization Production Rate (tpd)   290   282   303   329   374   280   229   160   -   225
 

Backfill

               
     Total CRF Backfill (000’s Tons)   105.7   102.9   110.5   120.5   136.5   102.3   83.5   29.1   -   791.0
 

Waste Mining

               
     Expensed Waste (000’s Tons)   57.2   49.7   55.4   55.2   61.8   44.6   37.6   14.0   -   375.5
               
     Primary Capital Drifting (Feet)   6,675   7,913   5,233   1,694   -   -   -   -   -   21,515
               
     Capital Raising (Feet)   1,050   640   180   180   150   150   -   -   -   2,350
               
     Capitalized Mining (000’s Tons)   180   201.6   116.4   38.5   1.1   1.1   -   -   -   539
             

Total Tons Mined (000’s Tons)

  343.1   354.2   282.2   214.2   199.4   147.9   121.1   43.0   -   1,705.2
             

Mining Rate (tpd)

  940   970   773   585   546   405   332   118   -   389

 

 

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14 Recovery Methods

 

14.1

Introduction

The Granite Creek processing facility was selected based on the metallurgical performance of the material. The material generally responds well to cyanide leaching but the presence of organic carbon (TOC) in some of the material hinders the gold recovery. To overcome the impacts of TOC, a carbon-in-leach (CIL) process was selected. The test work has shown that the CIL process largely overcomes the negative impacts of the organic carbon. Further, the CIL test work indicates a substantial gold recovery benefit compared to heap leaching. Heap leaching may still be a future processing option for lower grade materials but the capacity is largely driven by the gold cutover grade. At current gold prices, the use of heap leaching is not justified.

Oxide production from the Granite Creek underground operation will be processed through the planned Granite Creek CIL plant on site.

Refractory production from the Granite Creek underground operation will be initially processed via milling, pressure oxidation followed by carbon in leach (CIL) or roasting followed by CIL at a third party processing facility. The most recent metallurgical testing is described in Section 9, Mineral Processing and Metallurgical Testing, and supports processing parameters at these facilities at NGM.

Granite Creek underground production will be classified based on gold grade, level of oxidation and refractory characteristics (e.g. presence of preg-robbing components in ore, refractory sulfide components) that contribute to recovery at processing facilities and will be routed based on an integrated process production plan that is devised for maximum economic returns. Generally, materials grading >0.080 oz/ton will be routed to a third party processing facility.

Once operational in 2028, production will be processed at i-80’s Lone Tree pressure oxidation (POX) operation.

 

14.2

Process Description

The Granite Creek project would employ open pit mining with a CIL system on a 365 days per year 24 hour per day basis with 91% availability. Run-of mine (ROM) material at a nominal size of approximately 200 mm (8 inches) will be fed into a large jaw crusher. The jaw crusher would be equipped with a dump pocket capable of allowing direct dump from haul trucks or loaders. The dump pocket will have a static grizzly and hydraulic rock breaker to handle oversize material.

The jaw crusher is also equipped with a vibrating grizzly feeder that allows undersize to pass and oversize is fed to the crusher. The crusher will operate to produce a nominal P80 of 100 mm (4

 

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inch). Crushed material will advance to a live stockpile equipped with vibrating feeders to feed the downstream semi-autogenous (SAG) mill. Powdered lime will be added to the SAG mill feed belt via a silo and screw feeder.

The grinding circuit consists of a single SAG and ball mill. The SAG mill grinds the material to a nominal 19 mm (3⁄4 inch) and discharges across a trommel screen to reject oversize scats. The slurry reports to the mill discharge sump that feeds the cyclone pack. The cyclone underflow discharges to a ball mill that discharges back into the common mill sump. The ball mill is also fitted with a trommel to remove oversize scats. The cyclone overflow has a P80 target of 75um at 35% solids.

A pre-leach thickener is then used to thicken the ground material and flocculant is added to improve settling rates. The thickener underflow is then pumped to a series of CIL tanks, where the slurry flows countercurrent to the activated carbon. Cyanide and lime are added in the CIL circuit as required. The thickener overflow reports to the process water tank.

Gold is extracted via cyanidation and quickly adsorbed onto the active carbon. Carbon flows counter-current to the slurry and is recovered from the first tank. Inter-tank screens prevent the carbon from leaving the CIL tanks and recessed impeller pumps advance the carbon.

The loaded carbon extracted from tank 1 is transferred to a large column vessel. This carbon is then treated with a dilute acid wash to remove any calcium deposits. After water rinsing, the carbon is then eluted with cyanide and caustic to remove the adsorbed gold and silver. The resulting pregnant solution is then pumped to the electrowinning process and the precious metal cathodes are smelted into doré bars.

The eluted carbon is transferred to a regeneration kiln. Thermal regeneration is used to reactivate the barren carbon from the elution column. This reactivated carbon is pumped back into the last CIL tank. Fresh carbon may be added to this stream as well. Tailings from the CIL tanks are pumped to a tails thickener with the underflow reporting to the tailings storage facility (TSF). The overflow reports to the process water tank. Figure 14-1 shows the conceptual flowsheet.

 

 

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Figure 14-1: Conceptual Flowsheet

 

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14.2.1 Crusher Circuit

The crusher is designed to process approximately 555 metric tonnes per hour on an 18-hour basis. The crusher capacity is designed for 10,000 metric tonnes per day (tpd) at 75% availability. The jaw crusher is 1.22x1.52 meters (48x60 inch) with a 220 kW (300 HP) motor.

The run of mine feed passes over the static grizzly to reject oversize. The crusher is equipped with a hydraulic rock breaker. The feed then passes to a vibrating grizzly with a 150-mm (6 inch) opening. The undersize reports directly to the jaw crusher discharge conveyor (CV-110) while the oversize feeds the jaw crusher. The jaw crusher would crush to a nominal 100-mm (4 inch), with the crushed product reporting to a live stockpile.

The crusher discharge conveyor has a metal detector, magnet, and weigh scale. The stockpile has a live capacity of approximately 10,000 tonnes and has two vibrating feeders that feed the mill feed conveyor.

 

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14.2.2 Grinding Circuit

The mill feed conveyor (CV-150) feeds the SAG mill. Lime is added to the belt from a silo with a screw feeder. The conveyor has a magnet, metal detector, and weigh scale. The conveyor feeds the primary semi-autogenous mill (SAG). Process water is added to the feed to achieve the desired percent solids.

The selected SAG mill is 9.75m diameter by 3.35 long (32’ diameter by 11’ long) with a 5,000 kW (6800 HP) motor. The SAG mill has a trommel to discharge oversize material (scats) out of the circuit. No provision has been made for a pebble crusher to reprocess the scats in this design.

The SAG mill discharge reports to a common mill sump. The mill sump material is pumped to a cyclone pack (10 x 0.66-meter diameter cyclones) with the cyclone underflow reporting to the ball mill and the overflow reporting to the grind thickener. No gravity circuit has been included in the design.

The ball mill selected for this circuit is 7.31-meter diameter by 12.19-meter long (24-foot diameter by 40-foot long) equipped with a 7,134 kW (9700 HP) motor. The target P80 of the cyclone overflow is 75 um.

The cyclone overflow reports to the stock tank which is then pumped to a 63-meter diameter grind thickener. The grind thickener increases the solids density to reduce the size of the CIL tanks and provide improved density control.

14.2.3 Carbon in Leach (CIL) Circuit

Grind thickener underflow is pumped to the CIL circuit consisting of six mechanically agitated tanks operating in series designed to provide 48 hours of retention time. Each tank has a live volume of approximately 5,920 m3 with an 85% volume utilization. Slurry flows sequentially through tanks with the activated carbon retained in each tank by inter-tank screens. Carbon is advanced from the end of the train to the front of the train sequentially using recessed impeller pumps. The carbon flows countercurrent to the slurry. As gold is extracted from the ore, it is adsorbed onto activated carbon. The loaded carbon is extracted from tank 1 via pumping slurry across an external vibrating screen. The slurry returns to the tank and the carbon is then transported to the acid wash column.

To replace the loaded carbon removed in tank 1, regenerated or fresh carbon will be pumped to CIL tank 6. Slurry discharging from tank 6 gravitates to a carbon safety screen to recover any carbon leaking from worn screens or overflowing tanks. The slurry then proceeds to the tailings thickener.

 

 

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14.2.4 CIL Strip Circuit

The screened loaded carbon from the first CIL tank is pumped to the acid wash column. The loaded carbon is acid washed with dilute hydrochloric to remove calcium and adsorbed metals. The spent acid is neutralized and disposed of. After acid washing, the carbon is rinsed with water before gold and silver elution.

Elution is conducted by the modified ZADRA system at a rate of 18 tonnes/day. A solution of caustic and cyanide is passed through the elution column to remove the adsorbed gold. The rich electrolyte is pumped to electrowinning cells, where the gold and silver are recovered on the cathodes. The cathodes are washed, and the recovered sludge is refined in a conventional induction furnace after drying. The circuit is designed to conduct two daily strip cycles with a total metal recovery of approximately 740 oz per day (90% gold). The doré produced is assayed and stored in a vault before being shipped off-site for payment.

Barren carbon from the elution column is returned to the CIL circuit after passing across a carbon sizing screen. Fine carbon from the screen underflow is stockpiled and sent for separate off-site recovery. Approximately 50% of the barren carbon reports to an indirect fired kiln for thermal regeneration. The regenerated carbon reports to a quench tank before being pumped to the carbon sizing screen. Fresh makeup carbon is first sent to an attrition tank for fines removal before being pumped to the carbon sizing screen. The fine carbon from the screen underflow is captured in a plate and frame filter.

Tailings from the CIL circuit pass to a 63-meter diameter thickener and are thickened to 60% solids density. The thickener underflow reports to the tailings storage facility (TSF) and the overflow reports to the process water tank. Cyanide destruction may be required on this solution before it is circulated back to the circuit. The comminution circuit does not employ cyanide because of variable TOC grades.

 

14.3

Refractory Processing

14.3.1 Third Party Processing

The third party autoclave circuit processes 4 - 5 million tons per year and consists of primary crushing, two parallel semi-autogenous grinding (SAG) Mill-Ball Mill grinding circuits with pebble crushing, five parallel autoclaves capable of acid pressure oxidation (POX) and three of which are capable of alkaline POX, two parallel calcium thiosulphate (CaTS) leaching circuits with resin-in-leach (RIL), electrowinning for gold recovery, and a refinery producing doré

 

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bullion from both autoclave and roaster circuits.

Gold recovery estimates are based on both testwork and operational history at both facilities with curves utilized for both depending on operating strategy and ore characteristics.

The current autoclave LOM has an average recovery of 50% when running solely alkaline ore (one SAG-Ball Mill circuit and 3 autoclaves to process 4.0 million tons per year) and an average of 73% after converting the RIL circuit to CIL and running single refractory ore. The average LOM gold recovery is 65%. The simplified POX flowsheet is shown in Figure 14-2.

Figure 14-2 Third Party POX Simplified Flowsheet

 

LOGO

14.3.2 Lone Tree Pressure Oxidation Facility

i-80 Gold plans to process single refractory mineralization from Granite Creek at their Lone Tree Mill in a hub and spoke arrangement.

 

 

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14.3.2.1

Lone Tree Mill Historic Processing

The Lone Tree Mine is located immediately adjacent to I-80, approximately 12 miles west of Battle Mountain, 50 miles east of Winnemucca, and 120 miles west of Elko. Mining commenced at Lone Tree in April 1991 with the first gold pour in August of 1991. In 1993, a POX circuit was added to the facility, which included a SAG / ball mill circuit, followed by a thickening circuit, the POX process for refractory gold ores, and finally CIL, carbon stripping, and refining.

In 1997, a 4,500 tpd flotation plant was constructed to make concentrate to supplement the feed to the POX circuit, as well as to ship excess concentrate to Newmont’s Twin Creeks POX plant or to its Carlin roaster. The Lone Tree processing facilities were shut down at the end of 2007. Since that time, the mills have been rotated on a regular basis to lubricate the bearings. In general, the facility is still in place with most of the equipment sitting idle.

i-80 Gold Corp’s objective is to refurbish and restart the POX circuit and associated unit operations, including the existing oxygen plant, as it was operating before the shut-down, while meeting all new regulatory requirements. The flotation circuit is not being considered for restart. The POX circuit will have the capability to operate under either acidic or basic conditions.

In order to restart the process plant, new environmental regulations in relation to allowable mercury emissions must be met. In February 2011, the NDEP and the EPA brought about new standards to limit mercury emissions to 127 lb of mercury for every million tons of ore processed. In order to meet this requirement, the Lone Tree facility will require several environmental upgrades prior to restarting.

 

14.3.2.2

Lone Tree Facility Block Flow Diagram

A block flow diagram for the Lone Tree Mill facility is included in Figure 14-3. The block flow diagram contains the following major processing areas:

Ore Reclaim, Grinding and Thickening and Acidulation

Pressure Oxidation

POX Off-gas Treatment and Quench Water Loop

Neutralization, Carbon-in-Leach, and Cyanide Destruction

Tailings Thickening and Filtration

Acid Wash, Carbon Stripping, and Carbon Regeneration

Electrowinning and Refinery

Plant and Instrument Air

Oxygen Plant

 

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Reagent Preparation and Storage

Process and Plant Service Cooling Towers

Water Distributions

Steam Generating Plant and Propane Storage.

 

 

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Figure 14-3 Lone Tree Facility Block Flow Diagram

 

LOGO

Source: i80Gold (2025)

 

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14.3.3 Key Design Criteria

The Lone Tree Pressure Oxidation (POX) Facility restart will have minimal changes made from the 1993 Process Design Criteria (PDC). A new PDC was developed based on the expected production sources as defined by i-80.

Key process design criteria are summarized in Table 14-1.

Table 14-1: Summary of Key Process Statistics

 

Criteria     Units       Value  
     

Annual Mill Throughput

  tons   912,500
     

Daily Throughput (per calendar day)

  tons   2,500
     

Operating Throughput of Ore to Autoclave Circuit (LTH feed)

  tph   122.5
     

Operating Time / Availability

  %   85
     

Design Sulfur Treatment Rate

  tph S   2.7
     

Gold Recovery

  %   Varies
     

Silver Recovery

  %   Varies

14.3.4 Lone Tree Facility Description

 

14.3.4.1

Mill Feed Reclaim

The purpose of the Mill feed reclaim area is to store and reclaim material for processing, which has been shipped to the lone tree processing facility via highway ore trucks.

Run of mine (ROM) crushed material is delivered to the stockpile area. Material from various mining locations, namely Granite Creek, Cove, and Archimedes, is dumped at designated locations within the storage area and blended into facility feed stockpiles.

The stockpile area will have the capacity to store multiple days of mined and crushed material to accommodate the production shipment schedule to site. Additionally, the reclaim area is utilized for feed blending for the POX circuit. This blending will be used to manage the sulfide sulfur concentrations, gold grades, and carbonate grades through the autoclave to ensure stable circuit operation within the design window for the plant.

 

 

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14.3.4.2

Comminution

The purpose of comminution area is to reduce the particle size of the feed ore to the target autoclave circuit feed size for sufficient sulfide oxidation kinetics and gold recovery within the autoclave. The comminution area contains an SABC circuit with a dedicated SAG (semi-autogenous grinding mill) and ball mill to reduce the feed particle size to the target grind size. The SAG mill is fed via a conveyor from the dump hopper. The ball mill cyclone overflow is directed to the POX feed thickening conveyor.

 

14.3.4.3

Thickening and Acidulation

The purpose of the thickening area is to prepare the slurry for autoclave process by densifying the product of the grinding circuit to improve storage capacity of the downstream slurry storage tanks, improve the autoclave heat balance by reducing the water transferred to the autoclave and improving the possible solids flow through the autoclave feed pumps. The dense slurry is stored in two acidulation tanks that provide a combined storage / acidulation retention time of 12 hours. The acidulation tanks ensure a continuous feed to the autoclave plant, unaffected by upstream throughput variations.

 

14.3.4.4

Pressure Oxidation

The POX autoclave circuit includes the slurry pre-heaters, autoclave feed, autoclave, and the POX ancillary services: autoclave agitator seal system, oxygen supply, high pressure cooling water, and high-pressure steam. The Lone Tree Facility restart includes provisions to operate the circuit in alkaline or acidic modes depending on the feed carbonate concentration among other factors.

 

14.3.5

Slurry Heaters

The purpose of the slurry heaters is to capture excess energy discharged from the autoclave and pre-heat the feed slurry prior to the autoclave process reducing the total energy input required to operate the autoclave. The heating is achieved in two stages consisting of a series of two refractory lined counter-current splash slurry heater vessels. The heat source is flashed steam released from the autoclave discharge slurry during the pressure letdown process. The splash slurry heaters are direct contact heat exchanger and provide a means of heat recovery via steam condensation. This reduces the off-gas load on the downstream off-gas equipment and reduces the required input steam.

 

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14.3.6 Autoclave Feed

The purpose of the autoclave feed area is to increase the pressure of the pre-heated slurry to above the autoclave operating pressure to facilitate transfer into the autoclave at the required pressure using the autoclave feed pumps.

14.3.7 Autoclave

The purpose of the autoclave is to oxidize the refractory sulfide minerals under acidic or alkaline conditions to liberate the gold trapped in the sulfide sulfur minerals. The autoclave at Lone Tree is designed to operate at 389 °F and 297 PSI(g) with a slurry residence time of 40 - 50 minutes and consists of 4 compartments. The design expects a 78% - 97% cumulative sulfide sulfur oxidation through the autoclave depending on operating conditions. In either operating condition high purity oxygen is introduced to all four compartments of the autoclave at controlled rates to oxidize the fed sulfide minerals. Due to the low sulfur grades steam is required to be continuously fed to the autoclave to maintain the kinetically required oxidation rates to achieve the sulfide sulfur oxidation extent. The autoclave slurry is discharged through a level control choke valve and is fed to the high pressure flash vessel.

14.3.8 Flash System

The purpose of the flash system is to reduce the pressure and temperature of the autoclave discharge, making it suitable for subsequent unit operations downstream. The oxidized slurry undergoes a controlled pressure and temperature reduction process as it passes through two stages of flashing vessels located downstream of the last autoclave compartment.

 

14.3.8.1

POX Off-gas Treatment

The purpose of the POX off-gas treatment area is to effectively eliminate particulate matter present in the POX vent stream, while simultaneously reducing the temperature and volume of the vent gas through direct contact condensation. This process serves to alleviate the burden imposed on downstream equipment, ensuring their optimal performance, and mitigates the environmental impact by minimizing emissions. The off-gas treatment circuit also includes a mercury removal step to minimize autoclave mercury emissions to the environment.

 

14.3.8.2

Slurry Coolers

The purpose of slurry coolers is to reduce the temperature of the incoming slurry from the lowpressure flash vessel to prepare it for the downstream neutralization and CIL circuits through a series of water cooled shell and tube heat exchangers.

 

 

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14.3.8.3

Neutralization

The purpose of neutralization circuit is to neutralize all free acid in the slurry, precipitate the heavy metals as their hydroxides and raise the pH to approximately 10 to ensure cyanide stability in the CIL circuit for personnel safety and process optimization. The neutralization circuit is dosed with lime slurry to raise the pH of the autoclave discharge slurry. The neutralized slurry from this circuit is then fed to the CIL circuit for gold recovery.

 

14.3.8.4

Carbon-in-Leach

The purpose of CIL circuit is to leach and extract gold and silver from the oxidized slurry from neutralization using cyanidation and carbon adsorption. The CIL circuit provides retention time of 24 to 28 hours. The CIL circuit consists of 6 mechanically agitated tanks arranged in a series. The agitators prevent solid settlement and maximize contact time to improve gold and silver recovery. The carbon flows counter current to the slurry flows and the loaded carbon is sent to an elution circuit for carbon stripping and regeneration. Unloaded carbon is fed into the last tank of the CIL circuit. The leached slurry is transferred from to the cyanide destruction circuit.

 

14.3.8.5

Elution

The purpose of the elution circuit is to elute precious metals from the loaded carbon and transfer the resulting loaded solution of high gold concentration (pregnant eluate) to the refinery to generate doré.

 

14.3.9

Carbon Acid Wash

The purpose of acid wash is to rinse the loaded carbon form CIL with dilute nitric acid solution prior to the carbon stripping process. Carbonate scale builds up on the activated carbon during the CIL process and fouls the carbon’s adsorption properties by depositing a layer of scale. If left intact, over time the scale will limit the adsorption capacity of the carbon and will cause softening of the carbon in the regeneration kiln. The loaded carbon from CIL is first treated within the carbon acid wash vessel prior to treatment within the carbon stripping vessel.

 

14.3.10

Carbon Stripping

The purpose of the carbon strip circuit is to strip the cleaned loaded carbon from the acid wash vessel of the adsorbed gold using a Pressure ZADRA Strip scheme. The ZADRA strip uses several bed volumes of a recirculated solution to strip the precious metals off the loaded carbon. The cyanide solution is buffered by caustic to assist with gold elution. The stripped carbon is then sent to carbon regeneration circuits. The loaded solution is next processed in the electrowinning circuit.

 

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14.3.11

Elution Mercury Abatement

The purpose of elution mercury abatement system is to condition the off gas leaving the pregnant and barren solution tank to remove fine particulate, solution aerosols and condensed and gas phase mercury.

 

14.3.11.1

Carbon Regeneration

The purpose of the carbon regeneration circuit is to restore the activated carbon’s ability to recover gold from the cyanidation circuit solutions. The circuit also permits the introduction of new carbon to the process and removes carbon fines from the process.

 

14.3.12

Carbon Regeneration Kiln

As carbon is used in the CIL and elution circuits, the surface and internal pore structure becomes contaminated with organic species. The organics foul the carbon, slow the gold adsorption rate, and decrease the gold loading capacity of the carbon. The carbon reactivation electric kiln is a horizontal rotary kiln that is specifically designed for this purpose.

 

14.3.13

Carbon Fines Handling

Carbon fines are transferred by gravity from the reactivated carbon vibrating screen, carbon reactivation feed vibrating screen, kiln feed hopper, and carbon reactivation electric kiln. The carbon fines are dewatered in a filter press and discharged into supersacks for external sale.

 

14.3.13.1

Refinery

The purpose of the refinery circuit is to recover gold cyanide solutions via electrowinning and produce doré bullion bars.

 

14.3.14

Electrowinning

The purpose of the electrowinning (EW) circuit is to recover gold from the pregnant solution by applying a voltage across electrodes immersed in the pregnant solution. Rich solution from the pregnant solution tank is transferred through the EW cells to electrowin the gold.

 

14.3.15

Refining

The purpose of the refining process is to produce doré bars void of other contaminants including but not limited to mercury.

The sludge from the EW cells is first processed in a mercury retort oven to remove the co-captured mercury from the precious metals recovery steps. The retorted gold sludge is then processed in a melt furnace to produce the final mine grade doré bars.

 

 

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14.3.15.1

Cyanide Destruction

The purpose of the cyanide destruction circuit is to effectively reduce the concentration of cyanide in the final tail discharge and the recycled process water, ensuring compliance with predefined environmental standards and regulations and improving the safety of the operation by reducing cyanide concentrations outside of the CIL and elution circuits. The circuit targets a specific concentration limit of 2.5 mg/L of residual weakly acid-dissociable cyanide (CNWAD). This reduction is accomplished through the application of the SO2/air cyanide destruction process, which oxidizes the cyanide to meet the required concentration level. The cyanide destruction circuit is fed directly from the slurry discharge from the CIL circuit.

 

14.3.15.2

Tailings Preparation

The purpose of the tailings circuit is to increase the density of the detoxified tailings to aid with dry stacking of tailings residue. Additionally, this circuit produces process water for internal use within the facility. The tailings preparation circuit consists of a thickener as a first stage of solids densification. The thickener underflow is then fed to a tailings filtration circuit which dewaters the tailings sufficiently to support tailings dry stacking. The de-watered tailings from the filter presses are then dry stacked at the tailings storage facility.

The water removed from the tailings slurry is used as process water within the facility to offset water requirements. Excess process water is processed via a reverse osmosis circuit to provide supplemental permeate water to offset fresh water requirements.

 

14.3.15.3

Water Distributions

There are eight types of defined water services at Lone Tree:

Fresh water – generally used for reagent make-up and water washing streams.

Gland water – used to supply gland water to slurry pumps.

Mill water – used to provide dilution water within the milling circuit.

Potable water – used for safety showers and sanitary uses.

Demineralized water – primarily used to supply the steam generating plant.

Process water – used for washing and slurry dilutions. Additionally, generally feeds the reverse osmosis circuit to generate permeate water.

Quench water – used within the POX off-gas circuit as the source of direct cooling water.

Excess water – discharged from the main processing facility to the existing heap leach facility for treatment.

 

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14.3.15.4

Solution Cooling

The purpose of the cooling area is to reject heat absorbed within the process to atmosphere. The solution cooling area includes the process service cooling circuit and the plant service cooling circuit. The process cooling circuit rejects the heat from the autoclave cooling circuit and the elution circuit heat exchangers. The plant service cooling circuit provides trim heat rejection from various equipment support systems throughout the design.

 

14.3.15.5

Reagents

Each set of compatible reagent preparation and storage systems is located within dedicated containment areas to prevent erroneous mixing of reagents. Storage tanks are equipped with level indicators, instrumentation, and alarms to reduce the risk of spills during normal operation. Appropriate ventilation, fire and safety protection, safety shower stations and Safety Data Sheet stations are located throughout the facility.

 

14.3.16

Oxygen Plant

High purity oxygen is primarily used for oxidation of sulfide during the POX process, of iron conversion from ferrous to ferric in the neutralization circuit, and of cyanide to cyanate in cyanide destruction. Furthermore, during cyanidation, the addition of oxygen maximizes the rate of gold dissolution. At Lone Tree, a cryogenic ASU produces high purity oxygen. The unit uses pressure swing adsorption technology for front end purification and production of high-pressure oxygen at 95% purity.

 

14.3.16.1

Instrument and Plant Air

The Lone Tree facility includes separate instrument and plant air systems to support the facilities air requirements.

 

14.3.17

Utilities Consumption

The plant consumptions for water and power are provided for the average processing case below and consider the design blend of material to be processed within the Lone Tree Facility for the design life of operation.

 

14.3.17.1

Water Consumption

Table 14 -2 provides a summary of the water consumption by type for the Lone Tree processing facility.

 

 

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Table 14 -2: Lone Tree Facility Water Consumption by Type

 

   
Type   Consumption (gpm)
   

Mill Water

  1 550
   

Fresh Water

  570
   

Permeate Water

  195
   

Low Pressure Gland Water

  105
   

High Pressure Gland Water

  170
   

Demineralized Water

  110
   

Potable Water

  15

 

14.3.17.2

Electrical Power Requirements

The estimated annual electrical energy requirements for the Lone Tree processing facility are summarized by area in Table 14 -3.

Table 14 -3: Lone Tree Facility Energy Usage by Area

 

   
Area  

Annual Energy

 Consumption (MWh/y) 

   

000 – General Plant Wide

  2 250
   

180 – Water System

  930
   

181 – Potable Water

  240
   

182 – Process Water (RO and Process Water Tank)

  4 900
   

210 – Ore Reclaim

  770
   

240 – Refinery

  2 310
   

241 – POX Grinding

  26 920
   

242 – POX Grinding Thickening and Acidulation

  1 890
   

244 – Neutralization and CIL and Acid Storage

  6 540
   

245 – Carbon Stripping

  4 090
   

247 – CND

  690
   

248 – Reagents

  2 640
   

249 – Plant Air and Propane

  3 310
   

250 – Pressure Oxidation (POX) and POX Utilities

  15 540
   

251 – POX Demineralized Water System

  2 660
   

275 – Tailings Filtration

  13 690
   

300 – Plant Wide Electrical and Instrumentation

  4 000
   

305 – ABS and CN Storage

  160

 

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Area  

Annual Energy

 Consumption (MWh/y) 

   

320 – POX Mercury Abatement

  900
   

340 – Quench Water Treatment

  4 020
   

255 – Oxygen Plant

  40 090
   

099 – Existing Plant Areas

  3 570
   

Total

  142 090

 

 

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15

Infrastructure

 

15.1

Operations Dewatering

There are four dewatering wells operating (APW-1, BPW-3, BPW-5 and GCW-6), three in place at the time of purchase. one of which was deepened, and one new drilled by well i-80 in 2022. A fifth well exists but does not benefit the current mining and is not in operation. The wells are pumping from the CX block hydrogeologic unit at a combined average rate of approximately 1,410 gpm with an additional pumping of approximately 1,000 gpm collected from sumps in the underground mine workings (Figure 7 -25). Water is discharged to the rapid infiltration basins (RIBs) for infiltration back into the downgradient alluvial basin aquifer. A pipeline has been constructed to connect the dewatering circuit to the treatment circuit so the dewatering water can be treated if required, as well operations requirements for dust suppression and backfill material. The dewatering well pump parameters are listed in Table 15 -1.

 

15.2

Operations Monitoring Wells and VWPs

Monitoring wells and VWPs are used to collect hydrogeological data in support of mining operations. Currently, there are 41 active monitoring wells and 15 active vibrating wire piezometers across 5 locations (Figure 7 -25). Construction and recent water level data are provided in Table 15 -2..

 

15.3

Operations RIBs

Water from the dewatering wells that is not utilized for operations is currently discharged to Rapid Infiltration Basins (RIBs) on the east side of Getchel Mine Road through HDPE pipelines. Two of the four permitted RIBs (NEV2005102) have been constructed to date, with discharge to one of the two cells at any given time (Figure 7 -25). When RIB maintenance is required, discharge is routed to the dormant cell. Current dewatering efforts are well under the permitted 6,900 gpm threshold of the RIBs and the RIB infiltration is sufficiently limiting surface ponding in the active cell.

A portable rental water treatment plant capable of 800gpm is in use currently to treat water from the two deepest wells (BPW-5 and GCW-6) which do not meet minimum requirement quality for direct discharge to the RIBS. The mag pit serves as a diversion route for plant upsets, but is limited by a total number of gallons discharged.

 

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Figure 15 -1 Water Treatment Plant

LOGO

 

15.4

Operations Water Supply

Well WW-8, east of the Getchell road supplies potable water well to the Project (Figure 7 -25). The well is completed in basin alluvial deposits to a depth of 580 ft and equipped with a pump capable of supplying 60 gpm.

 

15.5

Underground Development

The mine is accessed through either of two portals, and dual egress has been established for most areas of the mine. Over 9,000 feet (2,743 meters) of underground workings have been completed Where dual egress is not possible, rescue chambers have been installed. Equipment is repaired in an underground mine shop. Air doors and a ventilation fan provide required air supply to the workings in compliance with Mine Safety and Health Administration (MSHA) standards.

 

15.6

Other Infrastructure

Existing infrastructure at the Project includes an office building, dry and warehouse facilities, and a lined stockpile area on the surface. Landline telephone and digital subscriber line service are available at the Project site. Cellular phone service is also available, but is dependent on the strength of receiving antennas, topography, and lines of sight. A fiber optic line provides wifi throughout surface infrastructure and key areas of the underground to support phone, radio, and process control instrumentation.

 

 

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Table 15 -1 Dewatering Well Completion Details

   
      Mine Coordinates   Elevation   Casing
Diameter
  Well
Depth
  Static
Water
Level
  Screened
Interval
  Average
Pumping
Rate
  Pump
Power
 

Pump
Set-

Depth

Well Identifier

   Easting
(ft)
  Northing
(ft)
  (ft amsl)   (inches)   (ft

bgs)

  (ft
bgs)
  (ft bgs)   (gpm)   (hp)   (ft bgs)

APW-1

   9890.2   10154.3   4722.3   18   620   380   120-140
160-180
200-600
  210   100   574

BPW-3

   10188.9   9474.8   5057.1   18   1391   780   500-540
580-620
660-700
740-1380
  750   400   1307

BPW-5

   10387.1   11126.9   5093.9   1812   2222   708   679 -1380
1400 - 2222
  380   200   1290

GCW-6

   10310.9   11742.9   5153.2   14   2093   742   803-2083   110   150   1950

 

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Table 15-2: Summary of Locations, Construction Information and Water Levels for Dewatering Wells, VWPs, Monitoring Wells and Piezometers

                 
Identifier   Local Mine
Coordinates
 

Elevation 

of 

Land 

Surface 

(ft  amsl)a 

            Open Interval of Well 
or VWP Setting 
       Static Water Level     
  Easting 
(ft) 
  Northing 
(ft) 
  Year 
Completed 
  Inclination
(degrees)b
 

Depth 

(ft bls)c 

 

Elevation 

(ft amsl) 

 

Geologic 

Unit(s) 

 

Depth 

(ft 

bls)d 

 

Elevation 

(ft amsl)e 

 

Date 

(dd/mm/yyyy) 

  Comments
   

Dewatering Wells

                                               
APW-1   9889.60   10152.60   4722.28   2005   -90   120 to 600   4602 to 4122   Ocl   308.1   4414.18   27/08/2024   CX Pit (active)
BPW-2   9804.81   10554.13   4762.98   2008   -90   200 to 920   4563 to 3843   Ocl   385.4   4377.58   23/11/2024   CX Pit (inactive)
BPW-3   10188.94   9474.81   5057.14   2008   -90   500 to 1380   4557 to 3677   Ocl   662.53   4394.61   10/07/2024   South of CX Pit (active)
BPW-4   10806.54   9132.50   5011.94   2008   -90   540 to 1380   4472 to 3632   Ocl   627.5   4384.44   07/10/2024   South of CX Pit (inactive)
BPW-5   10387.06   11126.94   5093.90   2008 / 2024   -90   679 to 2222   4415 to 2872   Ocu; Ocl   703.06   4390.84   17/12/2024   Between Mag and CX West Pits (active)
GCW-06   10310.90   11742.90   5153.20   2022   -90   803 to 2083   4350 to 3070   Ocu; Ocl   1852.3   3300.90   27/12/2024   Between Mag and CX Pits (active); deepened 2025
 

Water Supply Wells

WW-8   15141.92   8899.17   4756.11   1987   -90   210 to 560   4546 to 4196   Qal   190   4566.11   31/08/2000   East side of county road (inactive); no sounding port for manual DTW
   

VWPsa

                                               
iGS22-17   11081.00   11264.70   4830.70   2022   ---   ---   ---   Ocu               Mag Pit (active)
iGS22-17D_4188   10801.60   11478.90   ---   2022   -62   736   4188   Ocu   ---   ---   ---    
iGS22-17C_3735   10588.50   11598.00   ---   2022   -62   1250   3735   Ocl   ---   ---   ---    
iGS22-17B_3391   10428.30   11688.90   ---   2022   -62   1640   3391   Ocl   ---   ---   ---    
iGS22-17A_3233   10354.60   11732.70   ---   2022   -62   1820   3233   Ocl   ---   ---   ---    
iGS22-25   10411.80   11448.50   5104.00   2022   ---   ---   ---   Ocu               Between Mag and CX Pit (active)
iGS22-25D_4281   10081.30   11747.90   ---   2022   -63   937   4281   Ocl   ---   ---   ---    
iGS22-25C_4190   10049.50   11779.40   ---   2022   -64   1038   4190   Ocl   ---   ---   ---    
iGS22-25B_4077   10010.20   11819.40   ---   2022   -63   1164   4077   Ocl   ---   ---   ---    
iGS22-25A_3978   9975.60   11855.50   ---   2022   -63   1275   3978   Ocl   ---   ---   ---    
iGS22-26   11247.10   12508.00   5092.00   2022   ---   ---   ---   Qal               North of Mag Pit (active)
iGS22-26D_4199   11147.30   12590.70   ---   2022   -83   902   4199   Ocu   ---   ---   ---    
iGS22-26C_3983   11124.00   12609.50   ---   2022   -82   1120   3983   Ocu   ---   ---   ---    
iGS22-26B_3635   11087.10   12638.80   ---   2022   -82   1472   3635   Ocu   ---   ---   ---    
iGS22-26A_3384   11061.10   12659.10   ---   2022   -80   1725   3384   Ocl   ---   ---   ---    
iGS23-10A   11005.50   12407.30   5111.00   2023   ---   ---   ---   Ocu               North of Mag Pit (active)
iGS23-10A_4412   10935.20   12481.40   ---   2023   -81   707   4412   Ocu   ---   ---   ---    
iGS23-10A_3861   10857.40   12517.30   ---   2023   -81   1264   3861   Ocu   ---   ---   ---    

 

 Practical Mining LLC   March 26, 2025 


             

 

iGS23-10A_3700   10830.50   12526.90   ---   2023   -81   1428   3700   Ocu   ---   ---   ---    
iGS23-10A_3579   10811.70   12533.50   ---   2023   -81   1552   3579   Ocl   ---   ---   ---    
iGS23-02A   11091.40   11430.30   4826.60   2023   ---   ---   ---   Ocu               Mag Pit (active)
iGS23-02A_4128   10792.40   11906.90   ---   2023   -52   898   4128   Ocu   ---   ---   ---    
iGS23-02A_3747   10634.20   12132.80   ---   2023   -52   1371   3747   Ocu   ---   ---   ---    
iGS23-02A_3641   10592.10   12195.10   ---   2023   -52   1500   3641   Ocl   ---   ---   ---    
   

Monitor Wells

                                               
GMWCX-1   9844.60   12588.10   5258.10   1997   -90   505 to 545   4753 to 4713   Ocl   382.8   4875.30   11/12/2024    
GMWCX-2   8409.50   11701.20   5580.20   1997   -90   465 to 505   5115.2 to 5075.2   Kgd   278.92   5301.28   05/11/2024    
GMWCX-3   8235.70   10619.10   5321.30   1997   -90   278 to 318   5043.3 to 5003.3   Cpy   274.45   5046.85   05/11/2024    
GMWCX-4   7785.30   7991.50   5335.50   1997   -90   610 to 670   4725.5 to 4665.5   Cpy   506   4829.50   05/11/2024    
GMWCX-5   10341.30   8970.70   5073.00   1997   -90   490 to 523   4583 to 4550   Ocl   DRY   DRY   11/05/2024    
GMWCX-5D   10008.58   8670.70   5109.77   2008   -90   1000 to 1080   4109.77 to 1029.77   Ocl   711.68   4398.09   23/12/2024   Replacement for GMWCX-5
AMW-1   9857.09   10652.34   4762.31   2005   -90   900 to 960   3862.31 to 3802.31   Ocl   383.55   4378.76   23/12/2024    
BPZ0802   11070.54   9224.37   4995.80   2008   -90   ---   ---   Qal   DRY   DRY   ---    
BPZ0803   9329.24   11014.22   5194.53   2008   -90   ---   ---   Ocl   678.8   4515.73   16/05/2024    
GMW-HLMW-1   14221.21   4917.50   4692.27   1989   -90   254   ---   Qal   264.06   4428.21   27/09/2023    
GMW-RCH-588   15614.11   10449.75   4788.60   1995   -90   505   ---   Qal   193.94   4594.66   28/09/2023    
GMW-W7A   20480.58   10155.92   4657.70   1995   -90   110   ---   Qal   125.94   4531.76   28/09/2023    
MW 8   15140.35   8999.35   4757.01   1998   -90   212   ---   Qal   154.87   4602.14   02/10/2023    
RCH-1101   11440.58   12178.09   5060.51   1992   -90   264   465   Ocu   270.71   4789.80   05/06/2024    
RCH-1280   13419.41   9893.00   4850.00   1991   -90   ---   ---   Qal   211.25   4638.75   23/12/2025    
RCH-1305   10859.00   10333.45   5012.79   1992   -90   300   ---   Ocu   DRY   DRY   ---   Records indicate well has collapsed
RCH-1308   10981.17   9783.77   4996.49   1992   -90   285   ---   Ocu   DRY   DRY   ---   Records indicate well has collapsed
RCH-1309   10642.73   10744.50   5048.88   1991   -90   337   ---   Ocu   DRY   DRY   ---   Records indicate well has collapsed
RCH-1515   12602.88   9750.12   4895.68   1993   -90   219 to 465   4676.68 to 4430.68   Qal   258.1   4637.58   05/06/2024    
RCH-1516   12395.45   10333.65   4904.70   1993   -90   229   ---   Qal   DRY   DRY   ---   Records indicate well has collapsed
RCH-1517   12458.51   10705.76   4915.72   1993   -90   229   ---   Qal   DRY   DRY   ---   Records indicate well has collapsed
RMW2NE   15915.24   7792.24   4718.68   2005   -90   178 to 218   4540.68 to 4500.68   Qal   104.16   4614.52   15/02/2024    
RMW2SE   15902.63   7292.08   4711.48   2005   -90   158 to 198   4553.48 to 4513.48   Qal   91.35   4620.13   15/02/2024    
RMW2W/GMWMW2A   13432.00   9194.40   4838.20   1992   -90   140   500   Qal   203.91   4634.29   20/03/2024    
RMW3NE   15621.68   6584.88   4710.37   2005   -90   158 to 198   4552.37 to 4512.37   Qal   90.52   4619.85   15/02/2024    
RMW3SE   15488.44   6250.25   4706.13   2005   -90   158 to 198   4548.13 to 4508.13   Qal   87.59   4618.54   15/02/2024    

 

 

 Practical Mining LLC   March 26, 2025 


 Page 349  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

RMW3W   13681.94   7080.52   4804.22   2005   -90   278 to 318   4526.22 to 4486.22   Qal   179.19   4625.03   15/02/2024    
WELL 10   10454.99   11200.32   5084.57   1992   -90   242   542   Ocu   DRY   DRY   16/05/2024   Records indicate well has collapsed
WELL 2A   13452.52   9148.66   4839.80   1992   -90   144 to 450   4695.8 to 4389.8   Ocu   204.15   4635.65   05/06/2024    
WELL 6   8381.24   9858.41   5168.70   1998   -90   274   ---   Cpy   248.75   4919.95   16/05/2024    
WSW-W#11   10907.57   12300.59   5117.49   1998   -90   105 to 505   5012.49 to 4612.49   Ocu   DRY   DRY   ---    
WSW-W#2   13024.00   9420.00   4861.90   1992   -90   251 to 555   4610.9 to 4306.9   Ocu   228.83   4633.07   05/06/2024    
WSW-W#9B   10323.65   11966.86   5173.91   1998   -90   264 to 617   4909.91 to 4556.91   Ocl   DRY   DRY   ---    
 

Rib Piezometers

RPZ2E   14833.50   8020.25   4768.41   2005   -90   58 to 138   4710.41 to 4630.41   Qal   90.03   4678.38   15/02/2024    
RPZ2N   14732.84   8348.93   4771.00   2005   -90   58 to 138   4713 to 4633   Qal   128.5   4642.50   15/02/2024    
RPZ2S   14595.45   7762.05   4773.27   2005   -90   58 to 138   4715.27 to 4635.27   Qal   81.05   4692.22   15/02/2024    
RPZ2W   14496.42   8090.13   4785.56   2005   -90   58 to 138   4727.56 to 4647.56   Qal   88.58   4696.98   15/02/2024    
RPZ3E   14388.32   6615.05   4763.06   2005   -90   58 to 138   4705.06 to 4625.06   Qal   DRY   DRY   15/02/2024    
RPZ3N   14377.55   6939.79   4772.25   2005   -90   58 to 138   4714.25 to 4634.25   Qal   DRY   DRY   15/02/2024    
RPZ3S   14140.12   6414.30   4766.55   2005   -90   58 to 138   4708.55 to 4628.55   Qal   DRY   DRY   15/02/2024    
RPZ3W   14124.26   6741.97   4778.56   2005   -90   58 to 138   4720.56 to 4640.56   Qal   DRY   DRY   15/02/2024    

 

a)

feet above mean sea level; for wells, elevation of land surface at surface casing; for VWPs elevation of surface casing at land surface is provided;  

 

b)

degrees from horizontal at bottom of well or depth of VWP along inclined borehole using IDS survey 

 

c)

feet below land surface for wells; feet along inclined borehole for VWPs based on IDS inclination survey and Leapfrog Geologic Model positioning 

 

d)

feet below land surface for wells 

 

e)

feet below land surface for wells; depth to water subtracted from collar elevation 

 

 Practical Mining LLC   March 26, 2025 


 Page 350  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

16

Market Studies and Contracts

 

16.1

Precious Metal Markets

Gold and silver are fungible commodities with reputable smelters and refiners located throughout the world. The price of gold has reached all-time highs in 2024 with December’s price averaging 2,644 per ounce. As of December, 2024 the three-year trailing average gold price was $2,044 per ounce and the two-year trailing average price was $2,166 per ounce. The three -year and two-year trailing average prices for silver in December 2024 were $24.50 and $25.88 per ounce respectively. Historical plots for both are shown in in Figure 16 -1.

Figure 16 -1 Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average

 

LOGO

Issuers may also rely on published forecasts from reputable financial institutions. The current long term price forecast by CIBC is $2,169 and per ounce and $27.61 per ounce for gold and silver respectively (CIBC., 2025).

 

 Practical Mining LLC   March 26, 2025 


             

 

Commodity prices for Mineral Reserves are chosen not to exceed financial institution forecasts or the three-year trailing average price. Commodity pricing for the estimation of Mineral Resources can be 10% to 20% higher than that used for Mineral Reserves. The gold price selected for estimating Mineral Resources disclosed in this technical report is $2,175. The silver price selected is $27.25 per ounce.

 

16.2

Contracts

 

16.2.1

Orion and Sprott Financing Package

The Company entered into a financing package with OMF Fund III (F) Ltd. an affiliate of Orion Mine Finance (collectively “Orion”) on December 31, 2021, and a fund managed by Sprott Asset Management USA, Inc. and a fund managed by CNL Strategic Asset Management, LLC (“Sprott”) on December 9, 2021 (together the “Finance Package”).

The Financing Package in its aggregate consists of:

a.  $50 million convertible loan (the “Orion Convertible Loan”)

b.  $10 million convertible loan (the “Sprott Convertible Loan” and together with the Orion Convertible Loan, the “Convertible Loans”)

c.  $45 million gold prepay purchase and sale agreement entered into with affiliates of Orion (the “Gold Prepay Agreement”), including an accordion feature potentially to access up to an additional $50 million at i-80 Gold’s option

d.  $30 million silver purchase and sale agreement entered into with affiliates of Orion (the “Silver Purchase Agreement”), including an accordion feature to potentially access an additional $50 million at i-80 Gold’s option and an amended and restated offtake agreement entered into with affiliates of Orion (the “A&R Offtake Agreement”)

e.  5,500,000 warrants of the Company issued to Orion (the “Orion Warrants” and together with the Orion Convertible Loan, Gold Prepay Agreement, Silver Purchase Agreement and the A&R Offtake Agreement, the “Orion Finance Package”).

Under the Gold Prepay Agreement, i-80 Gold was due to deliver to Orion 3,000 troy ounces of gold for each of the quarters ending March 31, 2022 and June 30, 2022, and thereafter, 2,000 troy ounces of gold per calendar quarter until September 30, 2025 in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold. i-80 Gold may request an increase in the $45 million prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the Gold Prepay Agreement.

 

 

 Practical Mining LLC   March 26, 2025 


 Page 352  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

The final Gold Prepay Agreement includes an amendment to adjust the quantity of the quarterly deliveries of gold, but not the aggregate amount of gold, to be delivered by the Company to Orion over the term of the Gold Prepay Agreement. Under the amended Gold Prepay Agreement, commencing on the date of funding, the Company is required to deliver to Orion 1,600 troy ounces of gold for the quarter ending March 31, 2022, 3,100 troy ounces of gold for the quarter ending June 30, 2022, and thereafter 2,100 troy ounces of gold per calendar quarter until September 30, 2025, in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold, subject to adjustment as contemplated by the terms of the Gold Prepay Agreement. As the funding from Orion did not occur until April 2022, payment for the delivery of 1,600 ounces for the quarter ending March 31, 2022 was offset against the $45 million of proceeds received from Orion.

Under the Silver Purchase Agreement, commencing April 30, 2022, i-80 Gold will deliver to Orion 100% of the silver production from the Granite Creek and Ruby Hill projects until the delivery of 1.2 million ounces of silver, after which the delivery will be reduced to 50% until the delivery of an aggregate of 2.5 million ounces of silver, after which the delivery will be reduced to 10% of the silver production solely from the Ruby Hill Project. Orion will pay i-80 Gold an ongoing cash purchase price equal to 20% of the prevailing silver price. Until the delivery of an aggregate of 1.2 million ounces of silver, i-80 Gold is required to deliver the following minimum amounts of silver (the “Annual Minimum Delivery Amount”) in each calendar year: (i) in 2022, 300,000 ounces, (ii) in 2023, 400,000 ounces, (iii) in 2024, 400,000 ounces, and (iv) in 2025, 100,000 ounces. Upon a construction decision for the Ruby Hill project, comprised of one or both of the Ruby Deep or Blackjack Deposits, which construction decision is based on a feasibility study in form and substance satisfactory to Orion, acting reasonably, i-80 Gold will have the right to request an additional deposit from Orion in the amount of $50 million in aggregate in accordance with the terms of the Silver Purchase Agreement.

Both the Gold Prepay Agreement and the Silver Purchase Agreement were funded on April 12, 2022 with i-80 Gold receiving net proceeds of $71.6 million after netting the aforementioned March 31, 2022 gold delivery and closing costs as further described in Note 10 and Note 24 in the Company’s Financial Statements.

The main amendments reflected in the A&R Offtake Agreement include the increase in the term of the agreement to December 31, 2028, the inclusion of the Granite Creek and Ruby Hill projects, and the increase of the annual gold quantity to up to an aggregate of 37,500 ounces in respect of the 2022 and 2023 calendar years and up to an aggregate of 40,000 ounces in any calendar year after 2023. During the year ended December 31, 2022, Orion assigned all of its rights, title and interest under the A&R Offtake Agreement to TRR Offtakes LLC, now Deterra Royalties Limited.

 

 Practical Mining LLC   March 26, 2025 


             

 

On September 20, 2023, the Company entered into an Amended and Restated (“A&R”) Gold Prepay Agreement with Orion, pursuant to which the Company received aggregate gross proceeds of $20 million (the “2023 Gold Prepay Accordion”) structured as an additional accordion under the existing Gold Prepay Agreement.

The 2023 Gold Prepay Accordion will be repaid through the delivery by the Company to Orion of 13,333 troy ounces of gold over a period of 12 quarters, being 1,110 troy ounces of gold per quarter over the delivery period with the first delivery being 1,123 troy ounces of gold. The first delivery will occur on March 31, 2024, and the last delivery will occur on December 31, 2026. Obligations under the A&R Gold Prepay Agreement, including the 2023 Gold Prepay Accordion, will continue to be senior secured obligations of the Company and its wholly-owned subsidiaries Ruby Hill Mining Company, LLC and Osgood Mining Company, LLC and secured against the Ruby Hill project in Eureka County, Nevada and the Granite Creek project in Humboldt County, Nevada.

The remaining terms of the A&R Gold Prepay Agreement remain substantially the same as the existing Gold Prepay Agreement. The Company may request an increase in the prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the A&R Gold Prepay Agreement.

In connection with the 2023 Gold Prepay Accordion, the Company issued to Orion warrants to purchase up to 3.8 million common shares of the Company at an exercise price of C$3.17 per common share until September 20, 2026, and extended the expiry date of 5.5 million existing warrants by an additional 12 months to December 13, 2025.

 

16.2.2 

Orion Offtake

In February of 2025, i-80 Gold and Orion entered into an offtake agreement (the “Orion Offtake Agreement”). The Orion Offtake Agreement has similar terms to the current A&R Offtake Agreement with Deterra Royalties Limited and will commences upon its expiry. The Orion Offtake Agreement expires on December 31, 2034

Under the terms of the Offtake Agreement, the Company agreed to sell, and Orion agreed to purchase (i) an aggregate of 29,750 ounces of refined gold for 2021, and (ii) up to an aggregate of 31,500 ounces of refined gold annually (the “Annual Gold Quantity”) from the Company’s Eligible Projects until March 1, 2027. The Company’s Eligible Projects include the South Arturo Project, the McCoy-Cove Project, and the Granite Creek Project. The final purchase price to be paid by Orion will be, at Orion’s option, a market-referenced gold price in U.S. dollars per ounce during a defined pricing period before and after the date of each sale. In the event that the Company does not produce the Annual Gold Quantity in any given year, the obligation is limited to those ounces actually produced.

 

 

 Practical Mining LLC   March 26, 2025 


 Page 354  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

16.3

Previous Financing Agreements

 

16.3.1

South Arturo Purchase and Sale Agreement (Silver)

The Company entered into a Purchase and Sale Agreement (Silver) (the “Stream Agreement”) with Nomad, which was connected to South Arturo, whereby the Company will deliver to Nomad (i) 100% of the refined silver from minerals from the main stream area, and (ii) 50% of the refined silver from the exploration stream area. Nomad will pay an ongoing cash purchase price equal to 20% of the silver market price on the day immediately preceding the date of delivery and will credit the remaining 80% against the liability. Following the delivery of an aggregate amount of refined silver equal to $1.0 million to Nomad under the Stream Agreement, Nomad would continue to purchase the refined silver at an ongoing cash purchase price equal to 20% of the prevailing silver price. The liability for the Stream Agreement was included in the net asset value in connection with the asset exchange with Nevada Gold Mines LLC (“NGM”) discussed in the “Lone Tree and Ruby Hill Acquisition”, and therefore, is no longer impacting the Financial Statements as of December 31, 2021.

 

16.3.2

Autoclave Mineralized Material Purchase Agreement

The company has negotiated an agreement with a third party to sell up to 1,000 tons per day mineralized material for a fixed recovery of 58% of the contained gold at the average gold price during the month. In exchange there are no processing costs, refining or sales costs deducted from the purchase price. Transportation of the material to the processing site remains the responsibility of i-80. This agreement will apply to all refractory material mined from Granite Creek prior to the restart of the Lone Tree facility in 2028.

 

16.3.3

Contract Mining

Granite Creek mining is performed by a qualified contractor. The contract is structured to pay on footage advance with no allowance or additional payment for overbreak. Additional items are ground support required in addition to primary ground support and hourly rates for labor or equipment when work outside the scope is requested. There are no monthly fixed administration costs added.

 

16.3.4

Other Contracts

The company also intends to negotiate contracts for underground mine development, production mining, and over-the-road haulage with reputable contractors doing business in northeast Nevada. At the time of this report these negotiations have not been initiated. From time to time the company enters into other contracts for goods and services as a routine course of business.

 

 Practical Mining LLC   March 26, 2025 


             

 

17 Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups

 

17.1

Environmental Setting

The site is a producing underground operation built on a historic mine site that has been impacted by operations and exploration since the 1940s. The majority of disturbances have been reclaimed. In the valley to the east of the mine, there are several center-pivot irrigation systems raising hay. These adjacent water users may be beneficially impacted by the mine contributing groundwater to the RIBs. They are far enough away (and in the opposite direction of the prevailing wind) thus making them unlikely to be impacted by noise or dust from operations.

 

17.2

Geochemistry

The site has had limited geochemical characterization throughout history. Most of the geochemical test work was performed by WMCI in 1998. This study involved acid-base accounting, metals enrichment by acid-digestion and ICP-MS, and kinetic tests. 51 rock samples were tested statically, and 15 of those samples were selected for kinetic cell testing. Samples were selected from a variety of lithologies and locations but were designed to primarily focus on the Mag and CX future pit wall material (WMC Consultants, 1998).

Results from the ICP-MS analysis showed that major element abundance was controlled by rock type, with calcium abundant in carbonate-bearing rocks, and silica and aluminum concentrations abundant in silici-clastic rocks. Arsenic was found to be elevated in three samples and associated with hydrothermal deposits. ABA results indicated that the rock had low acid-generating potential (AP). Tested rocks had low sulfur, neutral paste pH, and abundant neutralization potential (NP) that resulted in 49 out of 51 samples being classified as non acid-generating based on the CANMET standards of NP/AP >3 (WMC Consultants, 1998).

Kinetic cells were run for a minimum of 20 weeks, with some running for a total of 28 weeks. Rates of ARD generation were tracked weekly, and the change in acidity and alkalinity was used to provide a quantitative estimate of whether the retained alkalinity would outlast the acid generation. Only 1 cell showed potential for acid generation: an argillite with 0.47% sulfur. This sample had consistently acidic pHs (4 – 2.2) with sulfate in the hundreds to thousands of mg/L range. All other kinetic cells had neutral to basic pH (7-9), alkalinity between 20-40 mg/L, and no quantifiable acidity. Monthly leachate analyses largely confirmed the weekly results. Metals in leachate were generally within the reference values, but some samples showed elevated levels of antimony and arsenic multiple times after the initial stabilization period. Additionally, the 2 cells that were uncertain under the acid-generation calculation showed levels of aluminum, antimony,

 

 

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 Page 356  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

arsenic, copper, iron, lead, and thallium after the initial stabilization period that were above Nevada Profile I reference values at the time. While the remaining cells were not over the standard at the time, the arsenic and antimony reference values were lowered in 2006 to 0.001 and 0.01 mg/L, which makes all the cells retrospectively over the standard. However, there was not a particular rock type that was consistently exhibiting acid generating or metal leaching potential (WMC Consultants, 1998).

Since 1998, the site performed periodic sampling characterizing waste rock authorized for disposal by backfill to the bottom of the CX pit. In the sample set taken from 2005 to 2022, the NP/AP ratio of this rock has varied from 2.5 to 568, confirming the presence of limestone layers within the Comus Sediments (see Section 7.0), that will readily neutralize any acid generated from the dissolution of sulfide minerals (Osgood Mining Company LLC, 2023). In addition to ABA tests, Meteoric Water Mobility Procedure (MWMP) tests have been performed on the waste rock deposited in the bottom of the CX pit. The weighted average of MWMP values based on volume of rock placed, the minimum, maximum, and geometric mean of constituents in exceedance of Nevada Profile I reference values is tabulated in Table 17 -1 (Stantec Consulting Services, 2023).

Table 17 -1 Weighted Average Concentrations of MWMP Results of Rock Placed in CX Pit 2005 - 2022

 

Analyte   

NDEP

Profile I

Reference

Value

  

Weighted

Average

   Maximum    Minimum    Mean
Antimony (mg/L)    0.006    0.190    0.070    0.001    0.184
Arsenic (mg/L)    0.010    0.510    2.200    0.015    0.423
Nitrate (mg/L)    10    51.3    150.0    0.6    46.7
Sulfate (mg/L)    500    151    800    3    184
TDS (mg/L)    1000    590    1,900    27    601

It is important to note that all the rock deposited in the CX pit will be covered with an engineered evapotranspiration (ET) cover (see Section 17.3). This will be protective of water quality impacts from rock leachate.

 

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The project, with the current available geochemical data, does not appear to pose ARD risk and only appears to pose minimal metal leaching (ML) risk in regards to antimony and arsenic release. To mitigate this ML risk, the mine operates a water treatment plant (see Section 17.1.4). GRE suggests that the project expand upon the current geochemical characterization to include a thorough characterization of future waste rock and tailings, as this will be required for future permitting efforts (See Section 23).

17.2.1 Onsite Water Quality

Water is sampled from many sources:

 

   

Underground dewatering wells (APW1, BPW3, BPW5, and GCW-06)

 

   

Background groundwater wells,

 

   

Underground mine sumps,

 

   

Surface water,

 

   

Mag pit lake,

 

   

RIBs, and

 

   

Influent and effluent from the Water Treatment Plant (WTP).

Bedrock groundwater from the underground dewatering wells has variable chemistry, which is reflective of the variable groundwater flow between different mineralized and unmineralized geologic units. In general, the groundwater from unmineralized blocks has lower arsenic and antimony than from mineralized blocks. The pH of bedrock groundwater ranges between 6.8 to 8.4, with total dissolved solids from 180 to 1,500 mg/L and alkalinity between 74 to 134 mg/L. Some bedrock groundwater also exceeds the Nevada Profile I reference values for metals, particularly for arsenic, cadmium, iron, manganese, nickel, and zinc (Osgood Mining Company LLC, 2023).

The site has maintained continuous bedrock groundwater quality monitoring for the purposes of compliance with Water Pollution Control permits (WPCP). A summary of the most recent water quality data collected is compiled in Table 17 -2. The Nevada Profile I reference values for antimony and arsenic are 0.006 mg/L and 0.01 mg/L respectively.

 

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Table 17 -2 Water Quality April 2023-Jan 2025

 Location ID

  

Antimony mg/L Range

  

Arsenic mg/L Range

APW1

  

0.0025-0.0041

  

0.029 - 0.059

BPW3

  

0.0025

  

0.02 – 0.027

BPW5

  

0.0025

  

0.012 – 0.037

GCW-06

  

0.0035 – 0.0085

  

0.33 – 1

RIB Distribution Pipeline

  

0.0025

  

0.023 – 0.028

WTP Effluent

  

0.0025 – 0.003

  

0.005 – 0.015

Most natural water meets Nevada reference values for antimony. The majority of bedrock groundwater onsite exceeds Nevade Profile I reference values for arsenic, however, the WTP effluent water quality demonstrates the WTP’s ability to meet arsenic reference values. An arsenic attenuation study is ongoing to address the high arsenic concentrations found in most groundwater.

Underground mine water is most impacted by metal leaching with analyzed samples collected from April to August 2024 exhibiting Nevada Profile I reference value exceedances for arsenic (ranging from 0.03 mg/l – 0.375 mg/l), antimony (ranging from 0.015 mg/l – 0.115 mg/l) and thallium (ranging from 0.003 mg/l – 0.014 mg/l). Range-front background alluvial groundwater quality has historically been relatively consistent over the period of record, with few exceedances of Nevada Profile I reference values, and some indication that natural chemical attenuation is occurring (Enviroscientists, Inc. Water Management Consultants, 2005). Alluvial groundwater on the project is monitored by numerous wells and is well understood. Onsite alluvial groundwater generally meets Nevada Profile I water quality reference values, with most trace metals at or below analytical laboratory standards (Osgood Mining Company LLC, 2023).

Granite Creek is located adjacent to the site and flows ephemerally during the spring and summer in response to snow melt and precipitation events. It is currently diverted through a series of pipes and culverts around the southern rim of the CX Pit to the original stream channel location downgradient of the pit. The water quality is consistently good, with all constituent concentrations below the Nevada reference values for surface water (Osgood Mining Company, LLC, 2020).

The Mag pit lake water quality has been monitored consistently since 2015. Samples have been taken from the top, middle, and bottom of the water column to establish any chemical differences in water quality with depth. Over 10 years of sampling, the average arsenic surface concentration of the Mag pit is 0.029 mg/L, the middle of the water column in the Mag pit has 0.029 mg/L arsenic, and the bottom of the Mag pit has 0.032mg/L arsenic. All layers show consistently high total dissolved solids (around 1000 mg/L) and sulfate (490-580 mg/L). The bottom of the Mag pit also appears to be elevated in manganese up to maximum of 0.53 mg/L in 2021 (LRE Water, 2024).

 

 

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17.2.2 Pit Lake Future Water Quality

The mine plan assumes that the CX pit will be backfilled and will not create a pit lake; however, a pit lake study exists. GRE believes that this study retains its relevance because it is the best-available study at present to discuss the water quality of the future MAG pit lake which will form after mining ceases. Similar to the CX pit study, ongoing work from LRE (unpublished as of the effective date) also shows that the Mag pit will be a hydrologic sink.

A thorough investigation of future CX pit wall material and backfill was conducted by WMC in 1998. This was followed up by a pit lake model to predict future pit lake water quality. WMC found that the majority of future pit wall rock were acid-neutralizing, with paste pH >7 and average sulfur content of 0.053% weight. Most rock had significant neutralization potential, with an average of 115 tons CaCO3/1000 tons rock.

The lake will behave as a hydrologic sink with no discharge of impounded waters to the surface or groundwater. The waste rock backfilled to the bottom of the CX pit will be inundated by rising post-mining pit lake waters to an estimated minimum depth of approximately 30 feet. Inundation of the backfill will cut off the oxygen supply and reduce or eliminate the potential for the backfill to generate acidic conditions. Pit lake predictive modeling reports indicate that long-term post-mining lake water will be in compliance with WPCP NEV2005103, with many metals concentrations less than the analytical laboratory reporting limit, significantly below Nevada reference values with the exception of arsenic (0.019 mg/l). In the event that post mining CX pit lake arsenic concentrations rise above acceptable levels, the modeling predicts that the addition of ferric sulfate solution (Fe2(SO4)3) at a rate of 0.23 grams per gallon would reduce the concentrations of arsenic to below the analytical laboratory reporting limit in the short- and long-term (Osgood Mining Company LLC, 2023). The model result is well-supported by the ferric sulfate dosing program that was tested at the CX pit lake in 2001 to help reduce arsenic concentrations reported in the pit lake at that time (Beale & Feehan, 2005).

17.2.3 Water Treatment Plant

The project currently has an 850 gpm capacity Veolia mobile water treatment plant (WTP). All the WTP storage and treatment components are contained within HDPE lined containment areas and consist of three areas: Surge Tank, Water Treatment Plant, and Sludge Dewatering System. The surge tank has a 10,000-gallon capacity and is used to blend water from mine contact water (200-250 gpm), BPW5 (400 gpm), and GCW-06 (120 gpm) to obtain adequate flow.

Water is treated for arsenic and antimony. Treated water is then discharged to the RIBs or diverted to the MAG pit which are regulated under WPCP NEV2005102.

 

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The water treatment plant utilizes ferric iron to precipitate antimony and arsenic from solution. It is made up of the following components:

 

   

Metal Precipitation Reactor

 

   

Addition of ferric sulfate or ferric chloride and sodium hydroxide during aeration with blowers

 

   

Actiflo Clarifier

 

   

Solids settle out of solution and are sent to sludge dewatering system

The Sludge Dewatering System consists of a splitter box that splits the sludge from the water, pumps that pump water back to the Actiflo clarifier, and a 1,500-gallon HDPE sludge tank. Sludge is then deposited in geobags, and the supernatant is recirculated back to the metals precipitation reactor. The geobag containment system consists of a two-layer HDPE liner system, with the bottom liner used for leak detection. Material from the geobags is mixed with waste rock and placed in the CX pit (approved by NDEP-BMRR).

An additional water treatment plant (a modular twin of the existing system) will be designed and built to accommodate the greater water disposal needs of the project.

Water Volumes

Water treatment needs are variable over time. At its peak, the underground mine is expected to produce 2900 gpm at its maximum dewatering extent. 80% (2300 gpm) is expected to come from dewatering wells, and 20% (600 gpm) is expected to come from contact water from underground sumps. The current IA considers that 2900 gpm of underground water is treated. However, there is an opportunity to create an improved water balance and to greatly decrease the treatment requirements (see Section 27).

Furthermore, the MAG pit will require dewatering. 2 years prior to open pit operations, the mine must commence dewatering the MAG pit at a rate of 450 gpm. This dewatering is anticipated to take 4 months to evacuate the volume of 69M gallons. MAG pit water has ~0.035 mg/L arsenic, which will be allocated to the TSF pond for forced evaporation.

 

17.3

Environmental Studies and Issues

The estimated cost to close and reclaim the Granite Creek Project is approximately $3 million (Osgood, 2023). This amount includes closure of all permitted surface and underground mining and exploration related disturbance at the Project and is calculated using standardized reclamation cost estimators that assess the following:

 

   

Exploration drill hole abandonment

 

   

Exploration roads and pads

 

   

Waste rock dumps

 

   

Heap leach pads

 

   

Roads

 

   

Pits

 

   

Foundations and buildings

 

 

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Other demolition and equipment removal

 

   

Sediment and drainage control

 

   

Process ponds

 

   

Landfill

 

   

Yards

 

   

Waste disposal

 

   

Well abandonment

 

   

Underground related infrastructure and portals closure

 

   

Miscellaneous costs

 

   

Monitoring

 

   

Construction management

 

   

Mobilization and demobilization.

Bonds in the amount of approximately $3 million are held by the BLM to address surface and underground reclamation and closure related activities (Osgood, 2023). There are no other known environmental liabilities associated with pre-Project operations (Osgood, 2024).

17.4 Social or Community Impacts

The following information on community relations and stakeholder consultation has been provided by Osgood Mining personnel (2024).

Historical mining activities date back to the 1940’s at the Granite Creek property (a.k.a., Pinson Mine), with intermittent periods of operation continuing to the present day. Over its history, the operation has contributed significantly to the economic development of Humboldt County through job creation and both direct and indirect economic benefits.

Osgood Mining Company periodically hosts Town Hall meetings in Winnemucca, Nevada, to offer operational updates to local stakeholders. The company also collaborates with the Nevada 95-80 Regional Development Authority during its annual economic development conference and works closely with the Humboldt County School District and Great Basin College, supporting various educational initiatives that benefit local students and the community at large.

Beyond these partnerships, Osgood Mining places a high priority on maintaining positive, long-term relationships with local government officials, ranchers, and neighboring landowners, ensuring that all parties are heard, respected, and included in the development process. Through these efforts, Osgood Mining strives to be a responsible and engaged community partner, prioritizing the well-being of the area and its residents throughout the duration of the Granite Creek Project.

 

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17.5 Permits

Osgood Mining Company is currently permitted to carry out mining operations and reclamation activities at the Project site. This permitting allows it to carry out the exploration, geotechnical and metallurgical field work recommended in this Report. Specific permits related to site activities are presented in Table 17 -3.

Table 17 -3 Granite Creek Mine Project Significant Permits

 Permit Name    Agency    Permit Number
Plan of Operations Granite Creek Mine Project   

BLM

   NVN-064101
Class II Air Quality Operating Permit   

NDEP-BAPC

   AP1041-3086.02
Mercury Operating Permit to Construct   

NDEP-BAPC

   MOPTC AP1041-3089 (De Minimis)
Water Pollution Control Permit - Rapid Infiltration Basins   

NDEP-BMRR

   NEV2005102
Water Pollution Control Permit - Granite Creek Mine   

NDEP-BMRR

   NEV2005103
Mine Reclamation Permit   

NDEP-BMRR

   0047
Granite Creek UG Mine Reclamation Permit   

NDEP-BMRR

   0242
   Mining Stormwater General Permit   

NDEP-BWPC

   NVR300000: MSW-42365
   Onsite Sewage Disposal System   

NDEP-BWPC

   GNEVOSDS09S0177
Hazardous Materials Storage Permit   

Nevada State Fire Marshal

   12441012106
  Waters of the United States Jurisdictional Determination   

USACE

  

Request for Approved Jurisdictional Determination (AJD) submitted to USACE November 2022

 

 

17.6 Water Use Permits

Water rights at the Granite Creek Mine have a total combined duty of 9,853 acre-feet annually (AFA), of which 1,149 AFA is for consumptive use (Osgood, 2024).

Water usage for the Project is managed via three certified water rights and ten permits, three of which are block permits. All water rights are subject to State Engineer’s Order 1087 (Block Order).

All use from the mine, including consumptive and non-consumptive use, is reported monthly on a site pumpage report and the specific meter readings are recorded and subsequently uploaded monthly to NDWR’s online meter database.

17.7 Future Permitting Requirements

The permits discussed above allow for the current ongoing underground operation at Granite Creek. Major permit revisions, as well as additional permits, will be required for the proposed plan of operations in this IA. The following sections details the anticipated new permits, permit revisions, and permitting efforts that Granite Creek will need to face for the mining plan described in this IA.

 

 

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17.7.1 National Environmental Policy Act (NEPA)

The National Environmental Policy Act (NEPA) is likely the largest single permitting hurdle that the project will face. The NEPA process is required when disturbances are anticipated to take place on federal lands and non-patented mining claims. It is reasonable to expect that this NEPA permitting effort will require the completion of an Environmental Impact Statement (EIS) through the BLM. An EIS is usually a lengthy process involving:

 

   

A large database of baseline data (prior to the anticipated mining impact)

 

   

A Plan of Operations (PoO) amendment describing the mining plan in detail

 

   

An assessment of the environmental impacts as a result of operations

 

   

A discussion of mitigation measures

 

   

An evaluation of the effectiveness of mitigation measures

 

   

A wide variety of supporting and supplemental environmental reports

The EIS is prepared by a third party hired by the BLM (not the mining company, and not the consultants who prepare the PoO amendment and baselines studies) but paid for with mining-company funds. The EIS is submitted to the BLM, where it is given a public comment period. After a process that often takes years from the commencement of baseline data collection, the BLM provides a Record of Decision (ROD), which acts as the permit.

At minimum, because the site has never had a full EIS, the following supplementary reports will be required:

 

   

Geochemistry of tailings and waste rock

 

   

MAG pit lake study

 

   

Backfill study for the mine waste below the water table in the CX pits

 

   

Groundwater and surface water resource studies and water quality studies

 

   

Seepage and groundwater quality studies for the TSF

 

   

Noise and vibration

 

   

Air quality

 

   

Wildlife and impacted biology

 

   

Archeology and cultural resources

 

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17.7.2 State Permits

State permits are required for air quality protection, groundwater protection, surface water protection, and water rights. All of the permits presented in Table 17 -3 will require revision with the amended PoO. Several key state permits are described below

17.7.2.1 Water Pollution Control Permit

The WPCPs are granted by the state of Nevada (NDEP-BMRR) and cover any potential discharge of water to surface or groundwater. This permit will require revision to be consistent with the amended PoO. An arsenic attenuation study will likely need to be conducted as part of the efforts to mitigate arsenic metal leaching risk and to establish an alternative discharge standard for the RIBs (see Section 20.1.4).

17.7.2.2 Reclamation Permits

Reclamation permits are overseen by NDEP-BMRR with the BLM providing supporting input. Existing reclamation permits for surface mining and underground mining will have to be revised. A new closure bond must be calculated and provided in anticipation of the new mining impacts in accordance with the PoO proposed in the IA. Section 20.3 discusses the closure plan and the reclamation cost estimate.

17.7.2.3 Other State Permits

Sewage disposal permits, stormwater permits, and air quality permits must be updated to be consistent with the PoO amendment specifics. It is important to note that the MAG pit dewatering is currently permitted under the exploration permit, and it can commence prior to acquiring other permits and prior to the ROD on the EIS.

17.7.3 Monitoring Requirements

Currently the site has been executing all environmental monitoring requirements required to maintain the WPCPs, air quality permits, reclamation permits, and other state permits associated with the small-scale underground operation. Samples of surface and groundwater are required quarterly, and quarterly and annual reports are provided to the regulators. Reclamation permits require annual disturbance reports and bond updates every three years. The stormwater permit requires quarterly inspections, an annual report, and an annual fee. All permit monitoring requirements are up-to-date. Widespread additional monitoring will be required in support of the NEPA permitting process, as well as after the ROD and well into closure.

17.8 Mine Closure

The mine closure cost estimate was derived using the Standardized Reclamation Cost Estimator Version 1.4.1, developed by the Nevada Division of Environmental Protection. This software is used by the Nevada regulators to calculate closure bond and closure cost requirements, and as a result, it is the only tool that could be reasonably applied for this estimate.

 

 

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GRE has expanded, modified and updated prior closure bond calculations provided by i-80 Gold. These included the version 1.4.1, 2025 update of the costs to close the current configuration of the Granite Creek Mine. These prior bond calculations were augmented and expanded to include the elements that would be constructed in the mine plan evaluated by this IA.

17.8.1 Mine Closure Design Criteria

The following design criteria and assumptions were applied to closure:

 

   

All key regulatory closure requirements will be met.

 

   

All regrading of mine waste structures is performed during operations. This includes creating tailings and waste rock facilities at a closure-ready 3:1 slope.

 

   

Some regrading of the waste rock dump and TSF will be required to recontour benches and to create a more-effective surface water drainage pattern.

 

   

Excess water on the TSF must be evaporated using an enhanced evaporation method. The surface will be allowed to sit dry for one year to consolidate and stabilize to a “trafficable” surface condition. It will take approximately one year to evaporate the tailings pond using the forced evaporator systems which were purchased for the MAG pit dewatering (see Section 17.1.4 above).

 

   

Waste rock and pit backfill is to be covered by 12 inches of cover and 5 inches of growth media. This is consistent with the prior closure cover.

 

   

An approved rangeland grass mix will be used. Grassland and wildlife habitat with grazing is the anticipated post-mining land use.

 

   

Because of the ML risk, the TSF, pit backfill, and WRSF will be covered with an Evapotranspiration Cover (ET Cover). In arid climates, ET covers have been shown to perform as well as HDPE covers in preventing mine waste leachate. For the PEA, the ET cover is assumed to be 18 inches thick.

 

   

All buildings, power lines, and other infrastructure will be removed.

 

   

The RIBs will be filled-in and reclaimed.

 

   

Four stream channels will be reclaimed, allowing precipitation in the hills to have a reclaimed and restored streambed in which it can flow through the reclaimed mine down into the valley. Granite Creek will be restored across the pit backfill and adjacent to the waste dumps to a channel similar to the pre-mining flowpath.

 

   

The water quality in the MAG pit lake will unlikely meet Nevada water quality standards due to elevated arsenic standards (see Section 17.1.1). However, because the lake will be a terminal sink for water, this IA does not consider it a long-term risk to groundwater quality. GRE assumes that an ecological risk assessment will likely conclude that the marginal arsenic concentrations (estimated at 35 ppb) will not be a significant risk to migratory waterfowl.

 

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GRE assumes that seepage from the CX pit backfill will not have a significant impact on groundwater quality and be contained within the evaporative cone-of-depression for the MAG pit lake and/or will be protected by the ET cover.

 

   

With an ET cover, GRE assumes that the WRDs will not create a surface or groundwater quality impact.

 

   

After dewatering and consolidation and with an ET cover, GRE assumes that the TSF will not cause a long-term surface water or groundwater seepage water quality issue.

 

   

Other than the issues discussed above, there are no other potential water quality or water quantity impacts at Granite Creek upon closure.

17.8.2 Closure Costs

The closure costs are summarized in Table 17 -.5

Table 17 -4 Mine Closure Cost Summary

 

Category   

Cost (Millions ,

USD)

   Notes

Earthwork Recontouring

   8.68   

Minimal recontouring is required because waste facilities are constructed at final closure grade.

Revegetation/Stabilization

   0.706   

Follows example of previous successful revegetation on site

Detoxification/Water

Treatment/Disposal of Waste

   0.430   

Includes anticipated waste disposal and evaporation of water in TSF

Structure, Equipment and

Facility Removal, and Misc.

   1.57   

Includes plant removal, new fence around MAG pit lake upon closure.

Monitoring

   1.63   

Assumes more monitoring wells due to larger footprint.

Construction Management

and Support

   0.28   

Calculated by the Reclamation Spreadsheet based on Nevada-based experience.

Contingency and Indirect

Costs

   4.68   

Recommended indirect costs with contingency as set in the SRCE model.

Total

   17.98   

This value is entered into the cost model.

 

 

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17.8.3 Closure Cost Limitations

The closure calculation is preliminary. Additional studies are required to confirm the design criteria are correct. Nearly all these studies are ultimately part of the EIS and permitting effort (see Section 17.2).

17.9 Local Procurement and Hiring

i80 gold has a specific corporate policy for local procurement and hiring. This program will be described in greater detail in subsequent SK-1300 reports.

 

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18 Capital and Operating Costs

18.1 Open Pit Capital Cost Estimate

The open pit capital cost estimate has been prepared for the IA under the assumption of processing of open pit mined material at 10,000 tpd through a CIL. Project costs were estimated using first principles, cost data from Infomine (Infomine, 2024), and the experience of senior staff. The estimate assumes that the project will be operated by a contractor; therefore, no mining equipment capital costs were included as this equipment would be provided by the contractor.

The initial capital costs are incurred in the years prior to production. GRE’s QP expects there will be three to five years of continued exploration, engineering, and permitting prior to a production decision.

Initial capital costs are defined as all costs until a sustained positive cash flow is reached. This includes labor and development costs in the pre-production years. Sustaining capital is defined as the capital costs incurred in the periods after a sustained positive cash flow is achieved through the end of the mine life.

All capital cost estimates cited in this Report are referenced in US dollars with an effective date of December 31, 2024.

Table 18 -1: Granite Creek Open Pit Mine Project Capital Costs

Item   Year
-2
  Year
-1
  Year
1
  Year
2
  Year
3
  Year
4
  Year
5
  Year
6
  Year
7
  Year
8
  Year
9
  Year
10
  Year
11
   Year
12
   Capital
Cost
(millions)

Open Pit Mining

Equipment

  $0.0   $0.0   $0.0   $0.0   $0.9   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    $0.0    $0.9

Capitalized Waste

  $0.0   $23.3   $6.8   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    $0.0    $30.2

CIL Process

  $73.3   $73.3   $0.0   $0.0   $0.0   $4.6   $4.6   $4.6   $4.6   $4.6   $0.0   $0.0   $0.0    $0.0    $169.6

Infrastructure

  $4.7   $4.7   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    $0.0    $9.4

G&A

  $0.0   $4.8   $0.0   $0.0   $0.1   $0.0   $0.0   $0.1   $0.0   $0.0   $0.1   $0.0   $0.0    $0.0    $5.1

Sustaining

  $0.0   $0.1   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    $0.0    $0.1

Working

  $0.0   $0.0   $10.6   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    ($10.6)    $0.0

Reclamation

  $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $9.0    $9.0    $18.0

Permitting

  $5.0   $5.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0   $0.0    $0.0    $10.0

Contingency

  $19.5   $21.9   $1.7   $0.0   $0.2   $1.2   $1.2   $1.2   $1.2   $1.2   $0.0   $0.0   $0.0    $0.0    $49.2

Total

  $102.5   $133.1   $8.5   $0.0   $1.2   $5.8   $5.8   $5.8   $5.8   $5.8   $0.1   $0.0   $9.0    $9.0    $292.4

A contingency of 25% was applied to all capital costs.

 

 

 Practical Mining LLC   March 26, 2025 


 Page  369  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

Initial capital costs total $235.6 million, as detailed in Table 18 -2.

Table 18 -2: Granite Creek Open Pit Mine Project Initial Capital Costs

Item

   Initial Capital
 Cost (millions) 

Open Pit Mine

   $0.0

Capitalized Waste

   $23.3

Plant

   $146.6

G&A

   $4.8

Infrastructure

   $9.4

Working

   $0.0

Sustaining

   $0.1

Reclamation

   $0.0

Permitting

   $10.0

Contingency

   $41.4

Total

   $235.6

18.1.1 Sustaining

Sustaining capital costs are set at 10% of the average yearly owner’s mobile equipment operating costs, or $0.1 million, and are incurred in year -1.

18.1.2 Facilities

All buildings and associated infrastructure installed on the property on a permanent or semi-permanent basis are considered facilities. They include material and installation costs. These costs are incurred in years -2 and -1.

Each item’s capital cost was estimated based on knowledge of nearby mine operations or senior engineers’ experience. Table 18 -3 shows total cost for each facility item.

Table 18 -3: Granite Creek Open Pit Mine Project Facilities Capital Cost

Item

    Capital Cost 
(millions)

Haul Roads

   $0.5

Office

   Existing

Warehouse

   $1.2

Mine Shop

   $4.1

Fuel Bay

   $0.1

Wash Bay

   $0.2

4x4 Pickup

   $0.3

 

 Practical Mining LLC   March 26, 2025 


             

 

Item    Capital Cost
(millions)
Security and Fencing    Existing
Surface Water Management    $0.6
Water Well with Pump    Existing
New Well Pump    Existing
Back Up Gen Set    $0.4
Sub-Station    Existing
Power Line 33KV    $2.1

Total

   $9.4

18.1.3 Process Plant

Costs for the CIL plant are incurred in years -2 and -1. Costs for tailings dam expansions are incurred in years 4 through 8. The plant capital costs are summarized in Table 18 -4.

Table 18 -4: Granite Creek Open Pit Mine Project Plant Capital Costs

Item    Capital
Cost
(millions)
CIL    $146.6
Tailings Expansion    $23.0
Total    $169.6

18.1.4 Mine Equipment

Because the project was assumed to be contractor-operated, no mine equipment capital costs were included, with the exception of pit dewatering pumps and evaporators, which are incurred in year 3 and total $0.9 m.

18.1.5 G&A Capital

General and administrative (G&A) capital costs include computers, software, technical support equipment, and office equipment. Initial capital costs for computers are $50k, occurring in year -1, with replacement costs occurring every three years for the life of the project. Initial capital costs for software are estimated at $150k, occurring in year -1, with supplemental costs of $15,000 every year for the life of the project. The total G&A capital costs are summarized in Table 18 -5.

Table 18 -5: Granite Creek Open Pit Mine Project G&A Capital Costs

Item    Capital Cost
(millions)
Computers    $0.2
Software    $0.3

 

 

 Practical Mining LLC   March 26, 2025 


 Page  371  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

Item    Capital Cost
(millions)

Tech Equipment

   $0.1

Office Equipment

   $0.3

Metallurgical/Geotechnical Drilling and Assaying

   $0.8

Total

   $1.7

18.1.6 Working Capital

Working capital is the necessary cash on hand for the next period’s operating cost. The estimated total is $10.6 million. This cost is recovered at the end of production.

18.1.7 Closure

Closure costs are estimated over two years at the end of production for closure and covering of waste storage facilities and the TSF. Total cost for site closure is $18.0 million. Additional details on closure costs are presented in Section 17.8.

18.2 Open Pit Operating Cost Estimate

The operating costs assume contractor operation. A 10% contractor’s premium was applied to all operating unit costs, labor unit rates, and supplies. Operating costs are summarized in Table 18 -6.

Table 18 -6: Granite Creek Open Pit Mine Project Operating Cost Summary

Item    Total Operating
Cost (millions)
   Unit Operating
Cost
   Unit  

Mining

   $637.0    $1.93      $/tonne mined  

Processing

   $343.6    $9.86      $/tonne processed  

G&A

   $53.3    $1.58      $/tonne processed  

Contingency

   $206.8              
Total    $1,240.7              

18.2.1 Labor

Hourly labor for the project is based on the number of people needed to operate and support equipment for each shift in a day plus additional crew to fill in for absences. Salaried labor in the project is based on job positions filled regardless of production changes or equipment units needed. Table 18 -7 through Table 18 -10 show the required labor, and Table 18 -11 shows the estimated mining and G&A labor costs by year. Processing labor costs are built into the processing unit costs of $9.86/tonne.

 

 Practical Mining LLC   March 26, 2025 


             

 

Table 18-7: Granite Creek Open Pit Mine Project Hourly Laborers by Year

Position    Year
-1
     Year 1      Year 2      Year 3      Year 4      Year 5      Year 6      Year 7      Year 8      Year 9  

Drill Operator

     12        24        28        24        24        20        24        20        16        8  

Blaster

     6        12        14        12        12        10        12        10        8        4  

Blaster Helper

     6        12        14        12        12        10        12        10        8        4  

Haul Truck Driver

     4        36        56        68        68        80        64        36        36        12  

Loader/Shovel Operator

     4        20        24        24        28        24        24        16        16        8  

Dozer Operator

     8        16        16        16        16        16        16        16        16        12  

Loader Operator

     4        4        4        4        4        4        4        4        4        4  

General Equipment Operator

     13        13        13        13        13        13        13        13        13        13  

Water Truck Driver

     8        8        8        8        8        8        8        8        8        8  

Lube Truck Driver

     8        8        8        8        8        8        8        8        8        8  

Laborer

     8        8        8        8        8        8        8        8        8        8  

Heavy Duty Mechanic

     24        48        59        62        63        65        60        45        44        29  

Light Duty Mechanic

     4        4        4        4        4        4        4        4        4        4  

Tire Man

     4        4        4        4        4        4        4        4        4        4  
Total      113        217        260        267        272        274        261        202        193        126  

Table 18-8: Granite Creek Mine Open Pit Mine Project Salaried Workers, Mine Management

Position    Number Each
Year
 

Mine Superintendent

     1  

Mine Engineer

     1  

Geologist

     1  

Surveyor/Tech

     1  

General Foreman

     1  

Shift Supervisor

     4  

Total

     9  

 

 

 Practical Mining LLC   March 26, 2025 


 Page  373  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

Table 18 -9: Granite Creek Mine Open Pit Mine Project General and Administrative Positions by Year

Position    Number
Each Year of
Active Open
Pit Mining
     Number Each
Year of
Stockpile
Processing
 

General Manager

     1        1  

Purchasing Manager

     0        0  

Purchaser

     1        1  

Chief Accountant

     1        1  

Accounting Clerk

     2        1  

Human Resources/Relations Manager

     1        1  

Human Resources/Payroll Clerk

     1        1  

Security/Safety/Training Manager

     1        1  

Safety Officer

     2        1  

Environmental Supervisor

     0        1  

Environmental Technicians

     2        1  

Logistics Administrator

     0        0  

IT Manager

     0        0  

Warehouseman ON SITE

     2        2  

Accounts Payable Clerk

     1        0  

Receptionist/Secretary

     1        0  

Guards

     4        4  

Drivers

     1        0  

Laborers / Janitorial On Site

     2        1  
Total      23        17  

Table 18 -10: Granite Creek Mine Open Pit Mine Project Processing Positions by Year

Position    Number Each Year
of Processing - CIL
 
Metallurgical Staff  

Superintendent

     1  

General Foreman

     1  

Maintenance Foreman

     1  

Shift Foreman

     4  

Chief Assay Chemist

     1  

Sr Metallurgist

     1  

Metallurgist

     1  

Process Technician

     0  

Instrument Technician

     0  

Subtotal

     10  

 

 Practical Mining LLC   March 26, 2025 


 Page  374  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Position    Number Each Year
of Processing - CIL
Laboratory

Sample Prep

   4

Assayers

   2

Analytical

   0

Subtotal

   6
Crusher

Operator

   4

FEL Operator

   4

Maintenance

   1

Electrical

   1

Subtotal

   10

Grinding

    

Operator

   8

Maintenance

   2

Electrical

   2

Subtotal

   12
CIL

Operator

   8

Maintenance

   2

Electrical

   1

Subtotal

   11
Strip Circuit

Operator

   24

Maintenance

   2

Electrical

   2

Subtotal

   6
Tailings

Operator

   0

Maintenance

   0

Electrical

   0

Subtotal

   0

Total

   39

Table 18 -11: Granite Creek Open Pit Mine Project Labor Costs by Year (millions)

Item    Year
-1
     Year
1
     Year
2
     Year
3
     Year
4
     Year
5
     Year
6
     Year
7
     Year
8
     Year
9
     Total  

Open Pit Hourly Labor

   $ 3.5      $ 19.7      $ 23.7      $ 24.3      $ 24.7      $ 24.0      $ 23.7      $ 18.2      $ 17.4      $ 5.6      $ 184.7  

 

 

 Practical Mining LLC

  March 26, 2025 


             

 

Item    Year
-1
     Year
1
     Year
2
     Year
3
     Year
4
     Year
5
     Year
6
     Year
7
     Year
8
     Year
9
     Total  

Open Pit Salaried Labor

   $ 0.5      $ 1.4      $ 1.4      $ 1.4      $ 1.4      $ 1.3      $ 1.4      $ 1.4      $ 1.4      $ 0.7      $ 12.3  

G&A Labor

   $ 1.0      $ 2.8      $ 2.8      $ 2.8      $ 2.8      $ 2.7      $ 2.8      $ 2.8      $ 2.8      $ 2.3      $ 25.2  

Total

   $ 5.0      $ 23.8      $ 27.8      $ 28.4      $ 28.9      $ 28.0      $ 27.9      $ 22.4      $ 21.5      $ 8.6      $ 222.3  

18.2.2 Mining Equipment and Consumables

Mining equipment includes production equipment and support equipment. Mining production equipment hours are calculated using the equipment productivity estimates and the number of tonnes required to be moved. It was assumed that all mining will be contractor-operated. GRE included a 20% surcharge on the estimated operating costs for account for contractor markup.

Mining support equipment hours are calculated using the number of shifts that the equipment is operated per day, the number of pieces of equipment, and the operating hours per day. The operating hours per day are calculated assuming utilization of 85%, availability of 90%, and two twelve-hour shifts per day. Table 18 -12 and Table 18 -13 summarize the mining costs by year.

Table 18 -12: Granite Creek Open Pit Mine Project Mining Equipment Costs by Year (millions)

Item    Year
-1
     Year
1
     Year
2
     Year
3
     Year
4
     Year
5
     Year
6
     Year
7
     Year
8
     Year
9
     Total  

Open Pit Production Equipment

   $ 7.0      $ 25.2      $ 38.3      $ 34.9      $ 31.7      $ 33.4      $ 45.5      $ 27.2      $ 25.5      $ 4.9      $ 273.6  

Open Pit Support Equipment

   $ 2.1      $ 6.0      $ 6.0      $ 6.0      $ 6.0      $ 5.8      $ 6.0      $ 6.0      $ 6.0      $ 3.0      $ 52.6  

Total

   $ 9.1      $ 31.2      $ 44.3      $ 40.8      $ 37.7      $ 39.1      $ 51.5      $ 33.1      $ 31.5      $ 7.8      $ 326.2  

Table 18 -13: Granite Creek Open Pit Mine Project Blasting Costs by Year (millions)

 

Item    Year
-1
     Year
1
     Year
2
     Year
3
     Year
4
     Year
5
     Year
6
     Year
7
     Year
8
     Year
9
     Total  

Explosives

   $ 2.3      $ 2.3      $ 6.3      $ 8.5      $ 6.6      $ 6.3      $ 5.7      $ 8.4      $ 6.2      $ 5.0      $ 0.7  

Primers

   $ 0.3      $ 0.3      $ 0.8      $ 1.0      $ 0.8      $ 0.8      $ 0.7      $ 1.0      $ 0.7      $ 0.6      $ 0.1  

Material Control/ Sample Testing

   $ 0.6      $ 0.6      $ 1.8      $ 2.4      $ 1.9      $ 1.8      $ 1.6      $ 2.4      $ 1.8      $ 1.4      $ 0.2  

Misc

   $ 0.2      $ 0.2      $ 0.5      $ 0.5      $ 0.5      $ 0.5      $ 0.5      $ 0.5      $ 0.5      $ 0.5      $ 0.3  

Total

   $ 3.4      $ 3.4      $ 9.4      $ 12.5      $ 9.7      $ 9.3      $ 8.4      $ 12.4      $ 9.2      $ 7.5      $ 1.2  

 

 Practical Mining LLC   March 26, 2025 


 Page  376  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

18.2.3 Process Plant

The processing operating costs include labor, reagents and consumables, and power. The unit rate for processing is $9.86/tonne of material processed. In addition, $0.83/tonne was included for rehandling of material from stockpiles. A summary of the process operating costs is provided in Table 18 -14.

Table 18-14: Granite Creek Mine Project Processing Costs by Year (1000s)

Item    Year 1      Year 2      Year 3      Year 4      Year 5      Year 6      Year 7      Year 8      Year 9     

Year

10

     Year 11      Total  

CIL Processing

     $20.7        $33.2        $33.2        $33.2        $33.2        $33.2        $33.2        $33.2        $33.2        $33.2        $11.1        $330.2  

Rehandle

     $0.1        $1.4        $1.2        $1.9        $2.2        $1.3        $0.1        $0.0        $1.4        $2.9        $1.0        $13.4  

Total

     $20.9        $34.6        $34.4        $35.0        $35.3        $34.4        $33.2        $33.2        $34.6        $36.0        $12.0        $343.6  

18.2.4 Taxes and Royalties

GRE prepared a generalized tax computation for the Granite Creek Mine Project. The following is a summary of tax elections incorporated into this tax computation:

 

   

The Granite Creek Open Pit Mine Project consists of a single mine and property

 

   

The Granite Creek Open Pit Mine Project will elect to treat mine development costs as incurred as deferred expenses

 

   

The Granite Creek Open Pit Mine Project will elect out of bonus depreciation.

 

   

The Granite Creek Open Pit Mine Project will elect to depreciate long-lived assets under the unit of production basis and all other assets will be depreciated using either 7-year or 15-year straight line depreciation

 

   

The Granite Creek Open Pit Mine Project will elect to deduct reclamation costs under Section 468.

Royalties were included in the cost estimation on a block by block basis. The total royalty applied totaled approximately 5.7% of the gross revenue.

18.2.5 General and Administrative

General and administrative costs were estimated for four phases of the mine plan: open pit production operating and stockpile operating. The G&A costs include both salaried and hourly labor, supplies, office equipment, and anticipated regular expenses. Open pit production years have a G&A cost of $5.2 million per year; stockpile operations years have a G&A cost of $0.9 million per year.

 

 Practical Mining LLC   March 26, 2025 


             

 

18.3

Granite Creek Underground

18.3.1 Capital Costs

Capital cost estimates are based on past actual, supplier costs, and internal estimates. Capital contingencies are calculated at 15% of capital development and delineation drilling and 25% on everything else. Capital costs estimates are within a range of accuracy of +/- 50% and are suitable for an Initial Assessment evaluation.

Table 18 -155 Capital Cost Estimates ($000’s)

      2025   2026   2027   2028   2029   2030   Total

Capital Development

   $19,553   $20,760   $12,755   $4,617   $600   $600   $58,885

Delineation

Drilling

   $6,000   $2,000   $2,000   $2,000   $2,000   $2,000   $16,000

Water Treatment Plant

   ($3,300.0)   ($3,500.0)   ($2,646.0)               $9,446

Feasibility Study

   $1,000                        

Underground Electrical

   ($500.0)   ($1,000.0)   ($500.0)   ($500.0)   ($500.0)   ($500.0)   $3,500

Fans/Ventilation

   ($100.0)   ($500.0)   ($100.0)   ($100.0)   ($100.0)   ($100.0)   $1,000

Contingency

   ($5,058)   ($4,664)   ($3,025)   ($1,142)   ($540)   ($540)   $14,969

Total

   ($35,511)   ($32,424)   ($21,026)   ($8,359)   ($3,740)   ($3,740)   $104,800

Unit mine development costs are derived from actual expenditures and work done over the period January through August 2024.

Table 18 -16 Mine Development Unit Costs

Description    $/ft  

Primary Drifting (15 ft x 17 ft)

   $ 2,300  

Secondary Horizontal Access (15 ft x 15 ft)

   $ 2,300  

Raise Bore (10 ft dia.)

   $ 4,000  

Excludes contingency

18.3.2 Closure and Reclamation

Total reclamation costs are estimated at $7.4M or $17.69/ounce produced. Reclamation costs are only for the underground mine-related disturbance. Legacy reclamation costs are included in the open pit estimates. Table 18 -1 show reclamation cost detail.

 

 

 Practical Mining LLC   March 26, 2025 


 Page  378  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Table 18 -17 Annual Bonding, Reclamation and Closure Costs ($000s)

        2025 -
2032
     2033–
2037
     2038-
2042
     Total 

Reclamation Bonding

     515      -      -      4,120 

Reclamation

     -      403      -      2,015 

Closure and Monitoring

     -      -      250      1,250 

Total

     515      403      250      7,385 

18.3.3 Underground Mine Operating Costs

Operating unit costs are summarized in Table 18 -1. The unit cost accuracy is within a +/- 50% range and are suitable for an Initial Assessment evaluation. The mining and trucking costs are derived from analysis of actual cost and production data over the period January through August 2024. They include all contractor charges, owner supplied materials and services.

Table 18 -18 Unit Operating Cost Estimates

Item      Unit Cost      Units

Stope Attack Ramps

     $110.59       $/ ton  

Drift and Fill

     $110.59       $ /ton  

Cemented Backfill

     $37.93       $/fill ton  

Gob Fill

     $13.00       $ /fill ton  

Expensed Waste

     $110.59       $/waste ton  

Lone Tree Acid POX Processing

     $106.00       $/ton  

Lone Tree Alkaline POX Processing

     $70.81       $/ton  

Trucking to Twin Creeks

     $8.16       $ /ton  

Trucking to Lone Tree

     $17.08       $ /ton  

Total cost, cost per ton and cost per ounce are shown in Table 18 - and Table 18 - for the with and without inferred mineral resource cases respectively.

Table 18 -19 Total and Unit Operating Costs (With Inferred Mineral Resources)

     Total
Costs
     Unit
Cost
     Cost per
Ounce
 
   ($M)      ($/ton
milled)
     ($/oz Au)  

Mining

   $ 331.7      $ 208.7      $ 794  

 

 Practical Mining LLC   March 26, 2025 


             

 

    Total
Costs
  Unit
Cost
  Cost per
Ounce
     ($M)   ($/ton
milled)
  ($/oz
Au)

Transportation & Processing

  $98.8   $62.1   $237

G&A, Royalties & Net Proceeds Tax

  $140.0   $88.1   $335

By-Product Credits

           

Total Operating Cost/Cash Costs

   $570.5     $358.9    $1,366

Closure & Reclamation

  $7.4   $4.6   $18

Sustaining Capital

  $88.8   $55.9   $213

All-in Sustaining Costs (1)

  $666.6   $419.4   $1,597

Excludes Resource Conversion Drilling

Table 18 -20 Total and Unit Operating Costs (Without Inferred Mineral Resources)

 

    Total
Costs
  Unit
Cost
  Cost per
Ounce
     ($M)   ($/ton
milled)
  ($/oz
Au)

Mining

  $171.2   $208.7   $923

Transportation & Processing

  $51.7   $63.1   $279

G&A, Royalties & Net Proceeds Tax

  $92.4   $112.6   $498

By-Product Credits

           

Total Operating Cost/Cash Costs

   $315.3     $384.3    $1,699

Closure & Reclamation

  $7.4   $9.0   $40

Sustaining Capital

  $88.8   $108.2   $479

All-in Sustaining Costs (1)

  $411.5   $501.6    $2,217 

Excludes Resource Conversion Drilling

18.3.4 Cutoff Grade

Cutoff grades vary depending upon process location and recovery. Cutoff grades for both refractory process locations are shown in Table 18 -.

 

 

 Practical Mining LLC   March 26, 2025 


 Page  380  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Table 18 -21 Cutoff Grades for Lone Tree POX and Mineralization Sales Agreement

 

     Acid POX   Alkaline POX   NGM

Gold Price ($/oz)

  $2,175

Nevada Commerce and Excise Tax

  1.151%

Refining and Sales ($/oz)

  $1.85

Royalty

  6%

Recovery1

  90%   82.5 – 94.2%   58%

Process Capacity (tpd)

  2500   2,500   1000

Mine Capacity (tpd)

  1000

Mining Costs ($/ton)

  $208.70

Haulage Cost

  $17.08   $17.08    $8.16 

Process Cost

   $106.00    $70.81   -

Incremental Cutoff Grade (opt)

  0.067   0.053 - 0.052   0.050

Mine Limited Cutoff Grade (opt)

  0.182   0.177 - 0.136   0.185

Fixed Costs ($ /ton)$

  $39.49

Process Limited Cutoff Grade (opt)

  0.203   0.201 – 0.159   0.218

 

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19

Economic Analysis

Readers are advised that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability under S-K 1300. This IA is preliminary in nature and includes inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. Readers are advised that there is no certainty that the results projected in this IA will be realized.

 

19.1

Taxes

19.1.1 Federal

The United States Government tax rate on corporations is 21% of taxable income. Taxable income is determined by offsetting revenue with depreciation, amortization, and depletion deductions. Unused depreciation and amortization deductions can be carried forward to the following year. The carryforward balance for the Granite Creek project at the beginning of 2025 is $91.9M.

19.1.2 Nevada

 

19.1.2.1

Net Proceeds Tax

Net mining proceeds are taxed at a rate of up to 5%. Net proceeds are generally defined as revenue less the costs of production. Capital investments are deductible using straight line depreciation over a 20-year period.

 

19.1.2.2

Excise Tax

The State legislature enacted an excise tax that went into effect in 2022. The tax applies to gross revenue from the extraction of gold and silver. The tax is two-tiered. Revenues greater than $20,000,000 and less than $150,000,000 are taxed at 0.75% while revenues above $150,000,000 are taxed at 1.1%.

 

19.1.2.3

Sales and Use Tax

Equipment and supplies for use in mining are subject to the sales and use tax. The tax rate for Eureka County is 6.85%.

 

19.1.2.4

Commerce Tax

The commerce tax is imposed on businesses with annual revenue exceeding $4,000,000. The commerce tax rate for mining companies is 0.051% of revenue above $4,000,000.

 

 

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Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

19.1.2.5

Modified Business Tax

All employers subject to Nevada Unemployment Compensation are also subject to the Modified Business Tax (MBT) on total gross wages less employee healthcare benefits paid. The MBT rate is 1.378%. The first $50,000 of gross wages is exempt from MBT.

19.1.3 Property Taxes

Property or ad valorem taxes are based on the value of the property, both real and personal. The Nevada constitution caps the property tax rate at five dollars for every $1,000 of assessed value. It is also capped by statute at $3.64 per $100 of assessed value. The assessed value in Nevada is 35% of the taxable value. Real and personal property taxes attributable to Real and personal property taxes attributable to Osgood Mining Company are summarized in Table 19 1 below. The total tax due for the 2023-2024 tax year is $100,996.60 (Humboldt County Assessor 2024).

Table 19 -1 Real and Personal Property Taxes

 

 Location   APN   Taxable
Value
    Assessed 
Value
    Tax
Rate  
    Annual Tax 

Osgood 38N 42E 21

  07-0121-05     $27,000                      $210.79

Murphy/Osgood 38N 42E 28

  07-0121-07     $11,000                     $85.88

Osgood 38N 42E 29

  07-0121-06     $27,000                     $161.36

Premier 38N 42E 21

  07-0121-08     $27,000                     161.36

Osgood 38N 42E 33

  07-0121-33     $27,000                     $6,615.54

Real Property Total

        $7,234.63

Ruby Hill Mine

  Kelly Creek Groundwater                           $6,404.45

Personal Property Mining Equipment

  MM000015                           $94,594.15

Total

                              $100,996.60

 

19.2

Granite Creek Underground

19.2.1 Cash Flow

Granite Creek underground is an operating mine and is in the initial production stage and ramping up to full production in 2025. Positive cash flow occurs early in the economic model with corresponding high IRR and short payback times.

A constant dollar cash flow analysis combining the mine production schedule including inferred mineral resources presented in Section 13.2 combined with the commodity pricing of Section 16.1 and the capital and operating costs of Section 18.3 is presented in Table 19 -2 and Table 19 -3 and graphically in Figure 19 -1 and Figure 19 -2. Results for the production plan excluding inferred mineral resources is presented in Table 19 -4 and Table 19 -5 and graphically in Figure 19 -3 and Figure 19 -4. Financial statistics from both cases are presented side by side in Table 19 -6.

 

 Practical Mining LLC   March 26, 2025 


             

 

Table 19 -2 Income Statement (Includes Inferred Mineral Resources)

 

   
     Production
       2025       2026       2027       2028       2029       2030       2031       2032       2033       Total  
     

Total Revenue

  $88   $103   $96   $165   $176   $130   $107   $42   $ -   $908
     

Mining Cost

  ($44.3)   ($41.8)   ($45.3)   ($48.2)   ($54.5)   ($40.4)   ($33.3)   ($11.8)   $ -   ($319)
     

Haulage and Processing

  ($1.7)   ($1.7)   ($1.8)   ($24.3)   ($27.7)   ($20.3)   ($16.1)   ($5.1)   $ -   ($99)
     

Electrical Power

  ($1.3)   ($1.4)   ($1.6)   ($1.7)   ($1.7)   ($1.7)   ($1.5)   ($1.2)   $ -   ($12)
     

Site Administration

  ($8.6)   ($8.6)   ($7.7}   ($7.7)   ($7.7)   ($7.7)   ($7.7)   ($7.7)   $ -   ($64)
     

Refining and Sales

  ($0.1)   ($0.1)   ($0.1)   ($0.1)   ($0.1)   ($0.1)   ($0.1)   $0.0   $ -   ($1)
     

Royalties & NV Taxes

  ($6.8)   ($8.6)   ($7.5)   ($14.8)   ($15.5)   ($10.7)   ($8.7)   ($2.9)   $ -   ($76)
     

Total Cash Cost

  ($63)   ($62)   ($64)   ($97)   ($107)   ($81)   ($67)   ($29)   $ -   ($570)
     

Cash Cost per Ounce1 ($/oz)

  $1,551   $1,307   $1,455   $1,275   $1,328   $1,357   $1,363   $1,479   $ -   $1,366
     

EBITDA

  $25   $41   $32   $68   $68   $49   $40   $14   $ -   $338
     

Reclamation Accrual

  ($1)   ($1)   ($1)   ($1)   ($1)   ($1)   ($1)   ($0)   $0   ($7)
     

Depreciation

  ($19)   ($26)   ($27)   ($49)   ($53)   ($41)   ($34)   ($13)   $ -   ($262)
     

Total Cost

  ($82)   ($89)   ($92)   ($147)   ($162)   ($123)   ($102)   ($43)   $ -   ($840)
     

Income Tax

  ($3)   ($5)   ($2)   ($7)   ($6)   ($3)   ($3)   ($1)   $ -   ($29)
     

Net Income

  $3   $10   $2   $12   $8   $3   $2   ($1)   $ -   $39

 

 

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Osgood Mining 

Company LLC. 

 

Table 19 -3 Cash Flow Statement (Includes Inferred Mineral Resources)

 

   
     Production
       2025       2026       2027       2028       2029       2030       2031       2032    

  2033-

  2042

    Total  
     

Net Income

  $3   $10   $2   $12   $8   $3   $2   -$1    $0   $39
     

Depreciation

  $19   $26   $27   $49   $53   $41   $34   $13   $0   $262
     

Reclamation

  $0   $0   $0   $1   $1   $1   $0   $0   $0   $0
     

Working Capital

  -$7    $0   $0   -$4    -$1    $3   $2   $4   $3   $0
     

Operating Cash Flow

  $15   $36   $29   $58   $61   $48   $38   $17   $3   $301
     

Total Capital

  -$36    -$32    -$21    -$8    -$4    -$4    $0   $0   $0   -$105 
     

After Tax Cash Flow

  -$21    $4   $8   $49   $57   $44   $38   $17   $3   $197
     

Cumulative Cash Flow

  -$21    -$17    -$9    $40   $97   $141   $180   $197   $200    

 

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Table 19-4 Income Statement (without Inferred Mineral Resources)

 

   
     Production
       2025       2026       2027       2028       2029       2030      2031      2032      2033       Total  
     

Total Revenue

  $39   $46   $43   $74   $78   $58   $48   $19   $-   $404
     

Mining Cost

  ($22)   ($21)   ($23)   ($24)   ($27)   ($20)   ($17)   ($6)   $-   ($159)
     

Haulage and Processing

  $0   $0   $0   ($10)   ($11)   ($8)   ($7)   ($4)   $-   ($52)
     

Electrical Power

  $0   $0   $0   ($10)   ($11)   ($8)   ($7)   ($4)   $-   ($12)
     

Site Administration

  ($9)   ($9)   ($8)   ($8)   ($8)   ($8)   ($8)   ($8)   $-   ($64)
     

Refining and Sales

  ($0)   ($0)   ($0)   ($0)   ($0)   ($0)   ($0)   ($0)   $-   ($0)
     

Royalties & NV Taxes

  ($3)   ($3)   ($3)   ($6)   ($6)   ($4)   ($3)   ($1)   $-   ($28)
     

Total Cash Cost

  ($35)   ($35)   ($36)   ($51)   ($56)   ($44)   ($37)   ($21)   $-   ($315)
     

Cash Cost per Ounce1 ($/oz)

  $1,971   $1,648   $1,824   $1,514   $1,566   $1,648   $1,687   $2,438   $-   $1,699
     

EBITDA

  $4   $11   $7   $22   $22   $14   $11   ($2)   $-   $88
     

Reclamation Accrual

  ($1)   ($1)   ($1)   ($1)   ($1)   ($1)   ($1)   ($0)   $-   ($7)
     

Depreciation

  ($19)   ($26)   ($27)   ($49)   ($53)   ($41)   ($34)   ($13)   $-   ($262)
     

Total Cost

  ($55)   ($62)   ($63)   ($101)   ($111)   ($86)   ($72)   ($35)   $-   ($585)
     

Income Tax

  $-   $-   $-   $-   $-   $-   $-   $-   $-   $-
     

Net Income

  ($16)   ($16)   ($21)   ($28)   ($33)   ($28)   ($24)   ($16)   $-   ($181)

 

 

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Osgood Mining 

Company LLC 

 

Table 19-5 Cash Flow Statement (Without Inferred Mineral Resources)

 

   
     Production
       2025       2026       2027       2028      2029       2030       2031       2032       2033  -  
2042 
    Total  
     

Net Income

  ($16)   ($16)   ($21)   ($28)   ($33)   ($28)   ($24)   ($16)   $0   ($181)
     

Depreciation

  $19   $26   $27   $49   $53   $41   $34   $13   $0   $262
     

Reclamation

  $0   $0   $0   $1   $1   $1   $0   ($0)   ($3)   ($0)
     

Working Capital

  ($4)   $0   ($0)   ($2)   ($1)   $1   $1   $2   $3   $0
     

Operating Cash Flow

  ($1)   $11   $6   $20   $21   $15   $11   ($1)   ($3)   $81
     

Total Capital

  ($30)   ($30)   ($19)   ($6)   ($2)   ($2)   $0   $0   $0   ($89)
     

After Tax Cash Flow

  ($36)   ($22)   ($15)   $12   $17   $11   $11   ($1)   ($36)   ($24)
     

Cumulative Cash Flow

  ($36)   ($58)   ($73)   ($61)   ($44)   ($33)   ($22)   ($23)   ($24)    

 

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Table 19 -6 Financial Statistics

 

     
        With Inferred         Without  Inferred   
   

Gold price (US$/oz)

  $2,175
   

Silver price (US$/oz)

  $27.25
   

Mine life (years)

  8
     

Average mineralized mining rate (tons/day)

  435   225
     

Average grade (oz/t Au)

  0.339   0.292
     

Average gold recovery (autoclave %)

  78%   78%
     

Average annual gold production (koz)

  52   23
     

Total recovered gold (koz)

  418   186
     

Sustaining capital (M$)

  $88.8   $88.8
     

Cash cost (US$/oz) 1

  $1,366   $1,699
     

All-in sustaining cost (US$/oz) 1,2

  $1,597   $2,217
     

Project after-tax NPV5% (M$)

  $155   ($30)
     

Project after-tax NPV8% (M$)

  $135   ($33)
     

Project after-tax IRR

  84%   -12.7%
     

Payback Period

  3.2 Years   NA
     

Profitability Index 5%3

  12.6   0.7

Notes:

  7.

Net of byproduct sales;

  8.

Excluding income taxes, resource conversion drilling, corporate G&A, corporate taxes and interest on debt;

  9.

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates breakeven;

  10.

This IA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the IA will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability;

  11.

Inferred mineral resources constitute 50% of mass and 56% of gold ounces of all mineral resources. The “Without Inferred” statistics presented are a gross factorization of the mine plan without any redesign of mine excavations or recalculation of productivities and costs. Capital costs are the same for the “With Inferred” and “Without Inferred” scenarios. The “Without Inferred” scenario is presented solely to illustrate the project’s dependence on inferred mineral resources;

  12.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable Canadian and United States securities regulations. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and

 

 

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Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

 

statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law.

Figure 19 -1 Gold Production and Cost per Ounce (With Inferred)

 

LOGO

 

 Practical Mining LLC   March 26, 2025 


             

 

Figure 19 -2 Cash Flow Waterfall Chart (With Inferred)

 

LOGO

Figure 19 -3 Gold Production and Cost per Ounce (Without Inferred)

 

LOGO

 

 

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Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Figure 19 -4 Cash Flow Waterfall Chart (Without Inferred)

 

LOGO

19.2.2 Sensitivity Analysis

The sensitivity of NPV, IRR and profitability index are shown in Figure 19 -5 through Figure 19 -6. The Granite Creek underground mine’s transition from development to production stage along with sustaining capital comprising 100% of all remaining capital expenditures provides the mine resilience to negative variance in gold price, operating costs and capital costs. The mine is most sensitive to gold price fluctuations. The gold price can decline to $1,565 per ounce or 28% before the after tax cash flow turns negative.

 

 Practical Mining LLC   March 26, 2025 


             

 

Figure 19 -5 NPV 5% Sensitivity

 

LOGO

Figure 19 -6 NPV 8% Sensitivity

 

LOGO

 

 

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 Page  392  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC 

 

Figure 19 -7 IRR Sensitivity

 

LOGO

Figure 19 -8 Profitability Index Sensitivity

 

LOGO

 

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19.3

Open Pit

19.3.1 Model Cases

A multi-scenario analysis method was used to analyze the economic performance of the project by varying the cutoff grade used, processing method or combination of methods, owner vs. contractor operations, tailings type, and equipment sizes.

GRE evaluated the following options:

 

   

High grade cutoff grades of 0.35, 0.4, 0.45, 0.5, 0.55, 0.6, 0.65, 0.7, 0.75, 0.8, 0.85, 0.9, 0.95, and 1.0 g/t. All cases also included low-grade material between 0.25 g/t and the high grade cutoff grade.

 

   

Three processing options: both CIL and heap leach, heap leach only, and CIL only.

 

   

Two tailings types: conventional and dry stack

 

   

Multiple sets of equipment sizes.

After analyzing the economic results of all cases considered, GRE selected the CIL only case with 0.85 g/t high grade cutoff, contractor operation, conventional tailings, and 133-tonne haul trucks and 21.9-tonne loaders as the base case as it results in the best overall economic results.

 

19.3.1.1

Economic Analysis

GRE performed an economic analysis of the project by building an economic model based on the following assumptions:

 

   

Federal corporate income tax rate of 21%

 

   

Nevada taxes:

 

  o

Proceeds of Minerals Tax – variable, with a maximum of 5% of Net Proceeds

 

  o

Property tax – 2.5605%

 

  o

Nevada gold and silver mine royalty – variable, with a maximum of 1.1% of gross revenue

 

   

Sales and use taxes are not included in the model

 

   

Equipment depreciated over a straight 7 or 15 years and has no salvage value at the end of mine life

 

   

Loss carried forward

 

   

Depletion allowance, lesser of 15% of net revenue or 50% of operating costs

 

   

Gold price of $2,175 per troy ounce, selected based on the CIBC January 2025 Consensus Gold Price Estimate (see Section 19 for more information)

 

   

Gold recovery calculated as detailed in Section 13

 

 

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Company LLC 

 

   

Block by block royalties totaling 5.7% of gross revenue, and a 10% net profit royalty

19.3.2 Results

GRE considered the following key economic parameters to determine the best scenario: NPV@5%, IRR, payback period, mine life, and initial capital cost. Table 19 - summarizes the results of the economic model.

 

 Practical Mining LLC   March 26, 2025 


             

 

Table 19 -7: Granite Creek Open Pit Mine Project Summary of Economic Model

 

                               
Item   Year -2     Year -1     Year 1     Year 2     Year 3     Year 4     Year 5     Year 6     Year 7     Year 8     Year 9     Year 10     Year 11     Year 12     Total  

Production

                                                                                                                       

CIL Tonnes Processed (‘000s)

    0.0       0.0       2,187.5       3,500.0       3,500.0       3,500.0       3,500.0       3,500.0       3,500.0       3,500.0       3,500.0       3,500.0       1,167.0       0.0       34,854.5  

Contained CIL Au Oz Processed (‘000s)

    0.0       0.0       131.7       144.5       142.8       115.6       99.0       166.6       196.9       184.3       136.2       59.8       19.9       0.0       1,397.2  

Recovered CIL Au Oz (‘000s)

    0.0       0.0       113.8       125.8       123.8       100.2       85.6       145.4       170.7       159.5       117.6       51.0       17.0       0.0       1,210.5  

Revenue

                                                                                                                       

Gross Revenue (M$)

  $ 0.0     $ 0.0     $ 247.6     $ 273.6     $ 269.3     $ 217.9     $ 186.2     $ 316.3     $ 371.3     $ 346.8     $ 255.8     $ 110.9     $ 37.0     $ 0.0     $ 2,632.7  

Refining/Selling Cost (M$)

  $ 0.0     $ 0.0     $ 0.6     $ 0.6     $ 0.6     $ 0.5     $ 0.4     $ 0.7     $ 0.9     $ 0.8     $ 0.6     $ 0.3     $ 0.1     $ 0.0     $ 6.1  

Transportation Charges (M$)

  $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0     $ 0.0  

Royalty (M$)

  $ 0.0     $ 0.0     $ 10.0     $ 13.1     $ 12.6     $ 10.4     $ 8.2     $ 16.4     $ 26.1     $ 26.2     $ 19.0     $ 6.2     $ 2.1     $ 0.0     $ 150.3  

Net Smelter Revenue (M$)

  $ 0.0     $ 0.0     $ 237.0     $ 259.9     $ 256.1     $ 207.0     $ 177.6     $ 299.2     $ 344.4     $ 319.8     $ 236.1     $ 104.5     $ 34.8     $ 0.0     $ 2,476.4  

Total Operating Costs (M$)

  $ 0.0     $ 0.0     $ 101.2     $ 152.5     $ 142.5     $ 140.4     $ 144.4     $ 165.9     $ 137.0     $ 131.9     $ 64.6     $ 44.8     $ 15.5     $ 0.0     $ 1,240.7  

Owners Royalty Add Back Pre-Tax

  $ 0.0     $ 0.0     $ 0.0     $ 0.1     $ 0.8     $ 1.2     $ 0.2     $ 3.1     $ 13.6     $ 15.2     $ 10.6     $ 1.4     $ 0.5     $ 0.0     $ 46.7  

Pre-Tax Operating Cash Flow

  $ 0.0     $ 0.0     $ 135.8     $ 107.5     $ 114.5     $ 67.8     $ 33.4     $ 136.3     $ 221.0     $ 203.1     $ 182.2     $ 61.0     $ 19.7     $ 0.0     $ 1,282.4  

Taxes

                                                                                                                       

Federal Taxes (M$)

  $ 0.0     $ 0.0     $ 15.9     $ 11.6     $ 11.4     $ 5.4     $ 2.4     $ 15.0     $ 28.7     $ 24.5     $ 23.5     $ 5.7     $ 1.5     $ 0.0     $ 145.5  

State Taxes (M$)

  $ 0.0     $ 0.0     $ 7.9     $ 6.5     $ 6.3     $ 4.3     $ 3.1     $ 7.5     $ 11.6     $ 10.5     $ 9.2     $ 2.7     $ 0.8     $ 0.0     $ 70.4  

After-Tax Operating Cash Flow

  $ 0.0     $ 0.0     $ 112.1     $ 89.3     $ 96.7     $ 58.2     $ 28.0     $ 113.8     $ 180.7     $ 168.1     $ 149.5     $ 52.7     $ 17.5     $ 0.0     $ 1,066.6  

Nevada Property Taxes (M$)

  $ 0.0     $ 0.0     $ 3.0     $ 2.4     $ 2.0     $ 1.6     $ 1.2     $ 0.9     $ 0.8     $ 0.8     $ 0.7     $ 0.0     $ 0.0     $ 0.0     $ 13.3  

Total Capital Costs (M$)

  $ 102.5     $ 133.1     $ 8.5     $ 0.0     $ 1.2     $ 5.8     $ 5.8     $ 5.8     $ 5.8     $ 5.8     $ 0.1     $ 0.0     $ 0.0     $ 0.0     $ 274.4  

Net Cash Flow After Tax (M$)

  ($ 102.5   ($ 133.1   $ 90.0     $ 86.1     $ 82.2     $ 44.2     $ 17.7     $ 93.8     $ 153.4     $ 142.7     $ 131.5     $ 46.7     $ 6.5     $ 1.6     $ 660.9  

Cumulative Net Cash Flow After Tax (M$)

  ($ 102.5   ($ 235.6   ($ 145.6   ($ 59.5   $ 22.7     $ 66.9     $ 84.6     $ 178.4     $ 331.8     $ 474.5     $ 606.0     $ 652.7     $ 659.3     $ 660.9          

 

 

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Table 19 - summarizes the key economic results for the base case.

Table 19 -8: Granite Creek Open Pit Mine Project Key Economic Results

 

After Tax Economic Measure    Value  
After Tax NPV@5% (millions)      $417.2  
After Tax IRR      28.7%  
Initial Capital (millions)      $254.7  
Payback Period (years)      3.72  
All-in Sustaining Cost ($/oz Au Produced)      $1,227.4  
Cash Cost ($/oz Au Produced)      $1,180.5  

19.3.3 Sensitivity Analyses

GRE evaluated the after-tax NPV@5% and IRR sensitivity to changes in gold price, capital costs, and operating costs. For this analysis, GRE used a base case gold price of $2,175/oz. The results indicate that the after-tax NPV@5% and IRR are most sensitive to gold price and grade, moderately sensitive to operating cost, and least sensitive to capital costs (Figure 19 - and Table 19 - for NPV@5%, and Figure 19 - and Table 19 - for IRR).

Figure 19 -9: After Tax NPV@5% Sensitivity to Varying Gold Price, Capital Costs, and Operating Costs

LOGO

Table 19 -9: After Tax NPV@5% (1000s) Sensitivity to Gold Price, Capital Costs, and Operating Costs

 

Variable    % of Base Case  
     60%        80%        100%        120%        140%  

Capital Cost

     $527.5        $474.0        $417.2        $406.1        $339.4  

 

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Variable    % of Base Case  

Operating Cost

     $703.2        $566.2        $417.2        $290.6        $94.4  

Gold Price

     ($128.8)        $174.4        $417.2        $681.5        $943.6  

Gold Grade

     ($136.6)        $169.1        $417.2        $688.7        $960.4  

Figure 19 -10: IRR Sensitivity to Varying Gold Price, Capital Costs, and Operating Costs

LOGO

Table 19 -10: IRR Sensitivity to Varying Gold Price, Capital Costs, and Operating Costs

 

Variable    % of Base Case  
     60%        80%        100%        120%        140%  

Capital Cost

     48.7%        36.8%        28.7%        24.1%        19.2%  

Operating Cost

     43.4%        36.5%        28.7%        21.1%        10.3%  

Gold Price

     -4.0%        15.5%        28.7%        40.9%        51.8%  

Gold Grade

     -4.6%        15.2%        28.7%        41.2%        52.6%  

19.3.4 Inferred Mineral Resource Impacts on Economics

This economic evaluation includes inferred mineral resources, which are considered to have a level of geological uncertainty too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability. The quantity of inferred mineralized tonnes included in the evaluation is 1,891.5 ktonnes, and the quantity of inferred Au ounces is 63.5 koz, representing 5.4% of the mineralized tonnes and 4.5% of the Au ounces.

As the percentages of inferred mineralized tonnes and Au ounces are quite low, the impact of including them in the economic evaluation is low. Removing the inferred tonnes and ounces would

 

 

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result in an after-tax cash flow of $600 million, an after-tax NPV@5% of $378 million, and an IRR of 27.4%.

19.3.5 Conclusions of Economic Model

The project economics shown in the IA are favorable, providing positive NPV values at varying gold prices, capital costs, and operating costs. The IA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves under CIM Definition Standards. Readers are advised that there is no certainty that the results projected in this preliminary economic assessment will be realized.

 

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20 Adjacent Properties

Most of the mineral rights surrounding Granite Creek are owned or controlled by Nevada Gold Mines. The only reported mining activities adjacent to Granite Creek were conducted by Nevada Gold Mines at the Turquoise Ridge Complex (TRC) where approximately 28 Moz Au have been produced since 1938 (Table x-x). The TRC is in Humboldt County, approximately 64 km northeast of Winnemucca and approximately 10 km north of the Granite Creek Mine. Historic and current production at TRC includes open-pit and underground mining.

Deposits that compose the TRC are Carlin-type, structurally and stratigraphically controlled and sediment-hosted. They containing disseminated micrometer-sized gold occurring on arsenic-rich pyrite rims, primarily within decalcified, carbonaceous rocks. Preferred host lithologies for gold mineralization are the Comus Formation, followed by the Valmy and Etchart Formations. Sub-microscopic gold mineralization is associated with arsenian pyrite, quartz, calcite, realgar, and orpiment. Gold mineralization is likely Eocene in age, and it is overprinted in some areas by a late stage of realgar, orpiment, and calcite. Gold-bearing zones can be located close to granodiorite and dacite dikes and beneath basaltic sills, evidencing the importance of rheologic contacts to mineralization (Barrick., 2024).

Practical Mining is unable to verify the information on production, mineral resources and mineral reserves included for the adjacent properties. This type of adjacent property information is not necessarily indicative of the mineralization at Granite Creek. In addition, the QP is not aware of any declared mineral resource that might have an impact on Granite Creek’s Mineral Resources, Mineral Reserves or mining operations.

 

 

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21 Other Relevant Data and Information

The authors are not aware of any other relevant data or information.

 

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22 Interpretation and Conclusions

 

22.1

Drilling

Drilling programs completed at the Property between 2005 and 2015 have included QA/QC monitoring programs that have incorporated the insertion of CRMs, blanks, and duplicates into the sample streams.

In 2021, QP GRE reviewed all of AMC’s work on available QA/QC data between 2005 and 2015 (AMC, 2020). In 2025, i-80 provided all QA/QC data from surface exploration holes drilled in 2021 and 2022 to GRE, and GRE reviewed all of them and found no material errors. GRE also reviewed and checked QA/QC Procedures and the database provided by i-80. GRE confirmed discussions and recommendations made in prior technical reports and noted the following:

 

   

Formal, written procedures for data collection and handling should be developed and made available to PMC field personnel. These should include procedures and protocols for fieldwork, logging, database construction, sample chain of custody, and documentation trail. These procedures should also include detailed and specific QA/QC procedures for analytical work, including acceptance/rejection criteria for batches of samples.

   

A detailed review of field practices and sample collection procedures should be performed on a regular basis to ensure that the correct procedures and protocols are being followed.

   

Review and evaluation of laboratory work should be an on-going process, including occasional visits to the laboratories involved.

In general, the QA/QC sample insertion rates used fall below general accepted industry standards. For future exploration campaigns, standards, blanks, and duplicates including one standard, one duplicate, and one blank sample should be inserted every 20 interval samples, as is common within industry standards.

CRM samples show a reasonable level of accuracy but poor to moderate precision when using standard deviations provided by the CRM supplier. A maximum of three to five different CRM samples would be adequate to monitor laboratory performance at the approximate cut-off grades, average grades, and higher grades of the deposits.

Blank sample results are considered acceptable and suggest no systematic contamination has occurred throughout the analytical process.

Duplicate sample results show suboptimal performance, which may be a result of the heterogenous nature of mineralization, uncrushed samples, and sampling variance. Overall duplicate samples appear to be positively biased, with duplicate results returning higher grade than original samples.

 

 

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Previous reporting suggests that umpire sampling has been completed at the Property. The results of this sampling were not available in the drillhole database and therefore the QP was not able to assess accuracy of the primary laboratory.

Although it is not possible to guarantee that there are no material impacts on the local scale, overall, based on the checking and reviewing the previous technical report dated 2020, GRE considers the assay database to be acceptable for Mineral Resource estimation.

 

22.2

Environmental

The environmental, permitting and social conditions are favorable for the project. The site is a producing underground operation built on a historic mine site that has been impacted by operations and exploration since the 1940s.

The project, with the current available geochemical data, does not appear to pose ARD risk and only appears to pose minimal metal leaching (ML) risk with respect to antimony and arsenic release. To mitigate this ML risk, the mine operates a water treatment plant (see Section 17.2). GRE suggests that the project expand upon the current geochemical characterization to include a thorough characterization of future waste rock and tailings, as this will be required for future permitting efforts (See Section 23).

The site has all relevant State and Federal permits to allow for the current ongoing underground operation at Granite Creek (see Section 17.3). Major permit revisions, as well as additional permits, will be required for the proposed plan of operations in this IA. Section 17.3 details the anticipated new permits, permit revisions, and permitting efforts that Granite Creek will need to face for the mining plan described in this IA. Due to the favorable jurisdiction, and despite the challenges of the National Environmental Policy Act (NEPA) permitting process, it is anticipated that the permits can be acquired in ~3 years.

The project has sufficient water rights to operate and will be self-sufficient for water due to the underground and open pit dewatering requirements.

 

22.3

Open Pit Conclusions

22.3.1 Metallurgical Conclusions

 

22.3.1.1

Sample Representativity

Within each zone, drilling has been localized to relatively small portions of the mineralized domains, as seen in Figure 10 1 and Figure 10 2The samples’ metallurgical response is likely to represent the zone’s general behavior, but additional sampling of each zone to confirm the

 

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metallurgical response will reduce uncertainty. The lack of this metallurgical drilling remains a risk to the project.

 

22.3.1.2

Test Work on Open Pit Samples

Cyanide leach bottle roll tests and column leach tests were completed on samples from both the Mag and CX open pits. Both Homestead and Atna commissioned these tests.

The test work demonstrated that many of the Mag Pit samples had high preg-robbing factors due to carbonaceous material in the feed. Due to the variable preg-robbing characteristics of the feed material, a higher degree of representativity of the Mag Pit should be evaluated.

Bottle roll tests were conducted on Mag Pit samples using NaOH as an alternative to hydrated lime, as a method of treating material with preg-robbing characteristics. These tests demonstrated that raising the pH improved gold recovery and decreased cyanide consumption.

A column leach test on a Mag Pit sample showed that there was no gold recovery benefit in using NaOH rather than lime (at the equivalent pH).

Test work on ground materials showed that Mag Pit materials were amenable to CIL methods. CIL treatment showed low impact from the TOC. Gold recoveries ranged from 83% to 94%.

Column leach tests on the Mag Pit samples achieved gold recoveries in the range of 19% to 82%.

Column leach tests on the CX Pit samples achieved gold recoveries of 82%.

22.3.2 Mineral Resource Conclusions

Mineral Resources for the Granite Creek open pit mine project have been prepared to industry best practices and conform to the resource categories defined by the SEC in S-K 1300. In the option of the QP, the resource evaluations reported herein are a reasonable representation of the Granite Creek open pit deposits.

In the opinion of the QP, the Mineral Resource model presented in this report is representative of the informing data, which is of sufficient quality and quantity to support the Mineral Resource estimate to the classifications applied.

22.3.3 Mining

The open pit mine plan for the is based on conventional mining techniques, reasonable production assumptions, and consideration of risks to achieving the mine plan.

 

 

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22.3.4 Economics

The open pit project economics shown in the IA are favorable, with an after-tax NPV@5% of $417.2 million and IRR of 28.7%. The sensitivity analysis shows that the economics are most sensitive to gold price and grade, then operating costs, and then capital costs.

The IA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves under CIM Definition Standards. Readers are advised that there is no certainty that the results projected in this preliminary economic assessment will be realized.

 

22.4

Underground Conclusions

22.4.1 Metallurgy

 

  1.

Granite Creek underground samples were refractory with baseline CIL gold recoveries ranging from 9% to 46%, averaging 31%;

  2.

Shake flask tests with gold cyanide spikes were used to determine preg robbing index. The average preg-robbing index was 17.9%, ranging from 4.4% to 54.1%.;

  3.

Bench top autoclave batch pressure oxidation tests were completed on all samples with 2 sets of acid conditions and four sets of alkaline conditions. Acid conditions resulted in resulted in higher sulfur oxidations and higher gold recoveries;

  4.

Three continuous pressure oxidation runs were completed with two acid and one alkaline sets of conditions based on the batch results. The continuous results followed the results of the batch tests with the acid conditions producing the higher sulfur oxidations and gold recoveries;

  5.

Overall gold recoveries increased with increasing sulfur oxidation;

  6.

Cyanide destruction tests on CIL tailings using the SO2/air process reduced weak acid dissociable cyanide concentrations to below 50 ppm using established reagent addition rates and retention time;

  7.

Thickening and filtration tests on CIL tailings showed unacceptable thickening properties and filtration rates. Thickening and filtration of pressure oxidation streams is not recommended.

  8.

Arsenic concentrations in the samples averaged 0.29%, largely occurring as arsenian pyrite with only trace amounts of arsenopyrite.

  9.

Sulfide minerals were predominantly pyrite with some marcasite.

 

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  10.

Mercury concentrations ranged from 31 ppm to 138 ppm, averaging 81 ppm. These concentrations will require mercury capture and abatement equipment in the process flowsheet.

22.4.2 Mining and Infrastructure

 

  1.

Mine infrastructure has been completed.

  2.

Production ramp up has reached approximately 400 tons per day.

  3.

The mining contractor is in place with the full complement of equipment and personnel.

  4.

Decline development has accessed 700 vertical feet of mineralization of the Otto and Ogee zones. Development has reached the top of the South Pacific zone allowing additional active production stopes.

  5.

The drill lateral drift over the South Pacific zone has been completed.

  6.

Reconciliation of the model to mill indicates process head ounces exceed model by 19%. This appears to be from mining in a larger low grade halo around high grade core.

  7.

Processed grade is lower than the life-of-mine planned grade due to extensive mining of marginal mineralization below the economic cutoff grade.

22.4.3 Economics

At $2,175 per ounce gold price the Granite Creek underground mine provides an 84% IRR due to the advanced stage of the underground mining with infrastructure in place. The projects NPV5% is $155M.

 

 

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23

Recommendations

 

23.1

Granite Creek Open Pit

The QP’s recommend the following items and budget to advance the Granite Creek project towards production (Table 23 -1).

Table 23 -1 Estimated Costs to Complete the 2-Year Program

Exploration Cost Area    Total  

Drilling (Exploration QA/QC, Metallurgy, and Geotechnical)

   $ 5,000,000  

Metallurgical Testing

   $ 400,000  

Permitting, including all baseline studies

   $ 9,000,000  

Engineering

   $ 750,000  

Total

   $ 15,150,000  

23.1.1 Metallurgical Recommendations

The following recommendations have been put forward:

 

23.1.1.1

Test Work Recommendations

A metallurgical drilling program should be undertaken to collect samples within the various zones representing the spatial, mineralogical, and grade difference. The collected samples should be tested for the following:

 

   

Paired fire assays and cyanide soluble assays to define cyanide solubility.

 

   

Bottle roll tests with and without carbon to predict reagent consumption as well as amenability to CIL treatment and to evaluate the impact of sulfide sulfur on the CIL performance.

 

   

Column leach tests at various sizes to predict field recovery for material to be heap leached. This should be performed on those materials with a cyanide solubility of greater than 50%. Recovery by size fraction should be completed as part of the testing program.

 

   

Conduct SAG and ball mill testing to determine the work index.

 

   

Additional autoclave pretreatment of underground materials should be completed, especially for those materials that showed lower gold extraction.

 

   

Infill the drill hole database with TOC and S= assays.

 

   

Conduct arsenic and mercury assays on all samples employed for metallurgical testing.

 

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23.1.1.2

Geometallurgy Recommendations

The geometallurgical work completed as part of this technical report should be expanded using the planned metallurgical test program results. The intent will be to confidently define those materials that can be treated by heap leaching or CIL methods and those that require autoclave treatment.

23.1.2 Environmental Recommendations

At the IA stage, the environmental recommendations focus on the efforts required to get state and federal permits. These recommendations are summarized below.

 

23.1.2.1

Requirements for the EIS

As mentioned in Section 17.7, the NEPA permitting process is expected to result in the need to complete an EIS. This process can take many years (even in a favorable jurisdiction like Nevada). As a result, the site will need baseline studies and supplemental environmental reports to prepare itself for the permit process.

To begin, the site needs several baseline reports. These will likely be:

 

   

Air Quality

 

   

Biological Resources

 

   

Surface Water

 

   

Groundwater

 

   

Geochemistry (including mine waste rock leachate, TSF leachate, TSF supernatant, and pit lakes)

 

   

Archeological and cultural resources

Because the kinetic geochemical tests have a year-long duration, the geochemistry study may be the critical path for permitting (and possibly production).

SERs are part of the NEPA process. Mr. Breckenridge of GRE recommends the commencement of the following

 

   

Geochemistry study

 

   

Pit lake study for the MAG pit

 

   

Backfill study for the mine waste below the water table in the CX pit

 

   

RIB water quality impact study

Because the site has never had a full EIS, it may require the following additional SERs:

 

 

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Noise and vibration

 

   

Visual impacts

 

   

Air quality

 

   

Biology

 

   

Archeology

These additional reports should be started as soon as possible so they do not slow the critical path to production.

Some of these studies, such as the hydrogeologic study predicting the underground mine dewatering and the potential impact of underground dewatering on water resources, were currently underway as of the effective date of this study.

 

23.2

Granite Creek Underground

23.2.1 Recommendations

 

1.

Metallurgical Testing

  a)

Establish sampling using the most recent mine plan to select samples to evaluate pressure oxidation with CIL cyanidation under Lone Tree conditions. Testing should also include baseline CIL tests and roasting testing as a comparison.

  b)

Testing should attempt to establish head grade and extraction relationships for use in more detailed resource modelling;

  c)

Mineralogy impacts need to be established and geologic domains within each resource need to be determined;

  d)

Additional comminution data should be collected to assess hardness variability within the zones and any potential impacts on throughput in the Lone Tree process plant..

  e)

The resource model should be advanced to include arsenic, TCM, TOC, mercury, as these will be important for predicting grades if toll process offsite is used and potentially for estimating extractions within the resources;

  f)

The estimated cost for the suggested next phase metallurgical program is to $350,000 based on current market pricing.

2.

Feasibility Study

  a)

Complete the resource conversion drilling program in the South Pacific zone;

  b)

Update the mineralization models, and;

  c)

Complete the feasibility study.

3.

Dilution Control

  a)

Align stope drift excavation direction parallel to strike wherever possible;

 

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  b)

Minimize mining outside the high grade 0.10 opt grade shells, and;

  c)

Reduce stope drift widths when the mineralized high grade zone is less than 15 feet.

 

23.2.2

Underground Feasibility Study Work Program

The recommended work program for the underground mine is listed in Table 23 -1

Table 23 - Feasibility Study Work Program

 

 Task

   Cost $M

 South Pacific Infill Drilling

   $ 6.0

 Metallurgical Testing

   $ 0.5

 Feasibility Study -

   $ 0.5

 Contingency

   $ 1.15

 Total -

   $ 8.15

 

 

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24 References

AMC. 2020. Technical Report Getchell Project NI 43-101 Technical Report, Premier Gold Mines Limited/Premier Gold Mines, USA, Inc., Humboldt county, Nevada, USA. 2020.

—. 2019. Technical Report, Osgood Pinson Deposit NI 43-101 Technical Report, Osgood Mining Company, LLC, Humboldt County, Nevada, USA. 2019.

Atna Resources Ltd. 2007. Technical Report Update Pinson Gold Property, Humboldt County, Nevada, USA. 2007.

Barrick. 2008. Pinson Geological/Geophysical Targeting. 2008.

Beale, Geoff and Feehan, Tom. 2005. Hydrology related to the proposed Pinson declines. Reno, NV: Water Management Consultants, 2005.

BLM. 2001. Granite Creek Administrative Draft Environmental Assessment N63-EA01-xx, prepared for Pinson Mining Company. 2001.

Chadwick, T. H. 2002. Barrick Gold Exploration Inc. Pinson Mine, Humboldt Co., NV. Plate 1: Structure and Lithology of the CS, C and Portions of the A Pits, April. 2002.

Chevillon, V. E., et al. 2000. Geologic Overview of the Getchell Gold Mine Geology, Exploration and Ore Deposits, Humboldt County, Nevada. [ed.] A. E.J. Crafford. Geology and Ore Deposits 2000: The Great Basin and Beyond, Geological Society of Nevada Symposium Proceedings. 2000, pp. 113-121.

CIM. 2014. Definition Standards for Mineral Resources and Mineral Reserves. s.l. : CIM Standing Committee on Reserve Definitions, 2014.

Dawson. 2005. Dawson Autoclave Leach Report, 28 October, updated 3 November. 2005.

—. 2006a. Pinson Underground Project, Results of Head Analyses and Autoclave/Cyanide Leach Tests on Samples from the Atna-Pinson Project, Project No. P-2895A, B & C, April 14. 2006a.

—. 2006b. Results of Sample Preparation and Head Analysis on Ogee Samples from the Atna-Pinson Project, Project No. P-2895D, April 24. 2006b.

Edmondo, G., McDonald, D. and Stanley, W. R. 2007. Technical Report Update Pinson Gold Property, Humboldt County, Nevada USA, June. 2007.

 

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 i-80 Gold Corp   References   Page  411  

 

Enviroscientists, Inc. Water Management Consultants, 2005. Water Pollution Control Permit Application – Pinson Exploration Infiltration Project Atna Resources. 2005.

Foster, J. M. and Kretschmer, E. L. 1991. Geology of the Mag Deposit, Pinson Mine, Humboldt County, Nevada. [ed.] G. L. Raines, et al. Geology and Ore Deposits of the Great Basin: Geological Society of Nevada 1990 Symposium Proceedings. 1991, pp. 845-856.

Foster, J. M. 1994. Gold in Arsenian Framboidal Pyrite in Deep CX Core Hole DDH-1541: Unpublished Pinson Mining Company Report. 1994.

Fritz Geophysics. 2007. Data Processing and Interpretation of Gravity Data, Pinson Property, Elko County, Nevada. 2007.

Golder Associates. 2014. NI 43-101 Technical Report Pinson Project Preliminary Feasibility Study, Humboldt County, Nevada, Atna Resources Ltd. 2014.

Gustavson. 2012. NI 43-101 Technical Report on the Mineral Resources of the Pinson Mine, Humboldt County, Nevada. 2012.

HGL, 2022b. i80 Gold Granite Creek Mine – Summary of 2022 VWP Installations and Preliminary Monitoring Results. Prepared for i80 Gold Corp, November 1, 2022.

HGL, 2023. Dewatering Well GCW-06, Granite Creek Mine, Humboldt County, Nevada. Prepared for i80 Gold Corp, June, 7 2023.

HGL, 2023. Groundwater Flow Model Results for Assessing Dewatering and Passive Inflow of Planned Underground Mine Workings, i80 Gold Granite Creek Mine, Humboldt County, NV, August 14, 2023.

Hofstra, A. H. and Cline, J. S. 2000. Characteristics and Models for Carlin-Type Gold Deposits. Society of Economic Geologists Reviews. 2000, Vol. 13, pp. 163-220.

Hotz, P. E. and Willden, R. 1964. Geology and Mineral Deposits of the Osgood Mountains Quadrangle, Humboldt County, Nevada: U.S. Geological Survey Professional Paper 431. 1964.

2020. Infomine. costs.infomine.com. [Online] December 2020.

https://costs.infomine.com/.

Jones, A. E. 1991. Tectonic Significance of Paleozoic and Early Mesozoic Terrane

 

 

 Practical Mining LLC   March 26, 2025 


 Page  412  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

Accretion in Northern Nevada: University of California Berkeley, Ph.D. dissertation. 1991.

Kretschmer, E. L. 1985. Geology of the Mag Deposit, Pinson Mining Company Internal Report. 1985.

—. 1984. Geology of the Pinson and Preble Gold Deposits, Humboldt County, Nevada. Arizona Geological Society Digest. 1984, Vol. 15, pp. 59-66.

Leonardson, R. W. 2016. Pinson Narrative Nov 2016: Unpublished Osgood Mining LLC Company Report. 2016.

LRE Water. 2024. i-80 Gold Corp Granite Creek Mine Dewatering Updated. PowerPoint Presentation, 2004.

LRE Water, 2025. Results of Deepening and Testing Dewatering Well BPW-5, Granite Creek Mine, Humboldt County, Nevada. Prepared for i80 Gold Corp., in preparation.

Lupo, J.F. 2005. Heap Leach Facility Liner Design. s.l. : Golder Associates, Inc., Lakewood, Colorado, 2005.

Magee Geophysical Services. 2006. Detailed Gravity Survey over the Pinson Property. 2006.

McClelland. 1999b. Report on Heap Leach Cyanidation, Column Leach Tests—Pinson Mine CX Pit Bulk Ore Samples, MLI Job No. 2630, June 16. 1999b.

—. 1999a. Report on Heap Leach, Direct and CIL Cyanidation and “Preg-Robbing” Tests—Various Mag Pit Samples and Composites, and CX Pit ‘bulk Ore, MLI Job No. 2532, Addendum, and C.O. #1, 2 and 3, March. 1999a.

—. 2014. Summary Report on Heap Leach Cyanidation Testing—Mag Pit Pinson Drill Core composites, MLI Job No. 3746, January 16. 2014.

—. 2013. Summary Report on Ore Variability Testing—Mag Pit Pinson Drill Core Composites, MLI Job No. 3746, February 7. 2013.

McLachlan, C. D., Struhsacker, E. M. and Thompson, W. F. 2000. The Gold Deposits of Pinson Mining Company: A Review of the Geology and Mining History through 1999, Humboldt County, Nevada. [ed.] A. E.J. Crafford. Geology and Ore Deposits 2000: The Great Basin and Beyond: Geological Society of Nevada Symposium Proceedings.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   References   Page  413  

 

2000, pp. 123-153.

Muntean, J. L., et al. 2011. Magnetic-Hydrothermal Origin of Nevada’s Carlin-type Gold Deposits. Nature. 2011, Vol. 4, pp. 122-127.

Nevada Gold Mines LLC. 2020. Technical Report on the Turquoise Ridge Complex, State of Nevada, USA, NI 43-101 Report. 2020.

OMC. 2020. Fact Sheet, Osgood Mining Company, Pinson Mining Project, Permit No. Nev2005103 (Renewed 2018). 2020.

—. 2019. OMC Pinson Metallurgy Summary April 2019. 2019.

Osgood Mining Company, LLC. 2023. Application to Renew Water Pollution Control Permit NEV2005103. s.l.: Osgood Mining Company, 2003.

Osgood Mining Company, LLC. 2020. Fact Sheet for Pinson Rapid Infiltration Basins WPCP. 2020.

Pinson Mine. 2015. Internal Ore and Ore Stockpile Tracking and Control, 5 Oct. 2015.

Piteau Associates, 2018. Mag and CX Pits Updated Pit Lake Models. Prepared for Osgood Mining Company LLC, Reno, NV, September 10, 2018.

Practical Mining LLC, 2023. Granite Creek LOM Development and Stopes. Email communication and files received July 6, 2023, from Dagny Odell (Practical Mining) to Daniel Weber (HGL). Drawing files (.dwg): GraniteCreekLOMDev.dwg and GraniteCreekLOMStopes.dwg.

Radtke, A. S. 1985. Geology of the Carlin Gold Deposit, Nevada, 1985: USGS Professional Paper 1267. 1985.

Ridgley, V., et al. 2005. Project Update Pinson Gold Project, Humboldt County, Nevada, USA. [ed.] H. N. Rhoden, R. C. Steinenger and P. G. Vikre. Symposium 2005 Window to the World: Geological Society of Nevada 2005 Symposium Proceedings. 2005, p. 103.

Schlumberger Water Services (SWS), 2014. CX Pit Lake Modeling Update. Prepared for Atna Resources Inc., Golconda, Nevada. February 19, 2014, 5298/R1.

Silberman, M. L., Berger, B. R. and Koski, R. A. 1974. K-Ar Age Relations of Granodiorite

 

 

 Practical Mining LLC   March 26, 2025 


 Page  414  

Initial Assessment of the Granite Creek Mine,

Humboldt County, NV

 

Osgood Mining 

Company LLC. 

 

Emplacement and Tungsten and Gold Mineralization Near the Getchell Mine, Humboldt County, Nevada. Economic Geology. 1974, Vol. 69, pp. 646-656.

Sim, R. 2005. Revised Technical Report on the Pinson Gold Property, Humboldt County, Nevada. 2005.

Stantec Consulting Services. 2023. Waste Rock Management Plan – Granite Creek Mining Project. s.l. : Osgood Mining LLC, 2023.

Stenger, D. P., et al. 1998. Deposition of Gold in Carlin-type Deposits: The Role of Sulphidation and Decarbonation at Twin Creeks, Nevada. Economic Geology. 1998, Vol. 93, pp. 201-215.

Stoker, P. T. 2006. Newmont Australia Technical Services Sampling Notes, AMC Report to Australia Technical Services, January. 2006.

Thompson, W. F. 2003. [Geologic] Stoping Report for the Pinson Properties: Internal Company Report, 2 March. 2003.

Wallace, A. and Wittkopp, R. W. 1983. The Mode of Occurrence of Gold at the Pinson Mine as Determined by Microprobe Analyses, and Metallurgical Implications: Unpublished Pinson Mining Company Report. 1983.

Water Management Consultants, Inc. (WMC), 1998. Results of Hydrogeologic and Geochemical Studies at Pinson and Provisional Closure Plan for the CX and Mag Pits. Prepared for Pinson Mining Company, Winnemucca, Nevada. July 1998, 2163/R2.

Water Management Consultants (WMC), Inc. 2000. Amended Closure Plan for the CX Pit. Prepared for Pinson Mining Company, September 19, 2000.

Wilmot. 2006. Wilmot Metallurgical Consulting Met Testwork Results Atna-Pinson Project. 2006.

WMC Consultants. 1998. Results of Hydrogeologic and Geochemical Studies at Pinson and Provisional Closure Plan for the CX Pt, Chapter 4: Geochemical Test Work. 1998.

WMC, 2002. Pinson Mining Company – CX Pit Status Report on Closure and Updated Pit Lake Modeling Predictions. Prepared for Pinson Mining Company, Winnemucca, Nevada. January 2002, 2163/R9.

WMC, 2005. Hydrology Related to the Proposed Pinson Declines. Technical

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   References   Page  415  

 

Memorandum, prepared for Enviroscientists, Inc.

 

 

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Company LLC. 

 

25 Reliance on Information Provided by the Registrant

i-80 provided information regarding the following

 

   

Status of surface and mineral estate, Section 3;

 

   

Geologic setting and mineral deposits, Section 6;

 

   

Hydrogeologic study and reports prepared by LRE Water, Section 7.4, 15.1 through 15.3;

 

   

2024 Processing statistics. Section 11.7.2;

 

   

Lone Tree POX facility description and flow sheet. Section 14.3.

 

   

CIBC market study. Section 16.1;

 

   

Summary of financing arrangements. Section 16.2;

 

   

Status of permitting, environmental liabilities and closure estimates. Section 17.

 

   

Underground mine cost accounting reports for 2024, Section 18.3;

 

   

Underground mine production data for 2024, Section 11.72 and 18.3;

The Registrant provided information appears reasonable and is within industry norms or information is from reputable third party organizations.

 

 Practical Mining LLC   March 26, 2025 
EX-99.4 5 d913666dex994.htm EX-99.4 EX-99.4

Exhibit 99.4

 

LOGO

S-K1300 Initial Assessment & Technical Report Summary for the Cove Project, Lander County, Nevada Effective Date: December 31, 2024 Report Date: March 26, 2025 Prepared for: i-80 Gold Corp. 5190 Neil Road, Suite 460 Reno, Nevada 89502 Prepared By: Practical Mining LLC. TR Raponi Consulting Ltd. Montgomery & Associates


 Page ii .  

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  i-80 Gold Corp. 

 

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Date and Signature Page

The undersigned prepared this Technical Report Summary (TRS) report, titled: S-K1300 Initial Assessment & Technical Report Summary for the Cove Project, Lander County, Nevada, dated the 26th day of March 2025, with an effective date of December 31, 2024, in support of i-80 Gold Corporations initial S-K 1300 disclosure of Mineral Resource estimates for the Cove Project.

Dated this March 26, 2025

/s/ Practical Mining LLC

Practical Mining LLC

495 Idaho Street, Suite 205

Elko, Nevada 89815, USA

/s/ T.R. Raponi Consulting Ltd.

T.R. Raponi Consulting Ltd.

15-223 Rebecca Street

Oakville, ON Canada L6K 3Y2

/s/ Montgomery & Associates

Montgomery & Associates

132 S. State Street

Salt Lake City, Utah 84111

 

 

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  i-80 Gold Corp. 

 

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S-K1300 Initial Assessment & Technical Report

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  Page v . 

 

Table of Contents

 

Date and Signature Page

     iii  

Table of Contents

     v  

List of Tables

     xi  

List of Figures

     xiv  

List of Abbreviations

     xvii  

1.  Summary

     18  

1.1.   Introduction

     18  

1.2.   Property Description

     19  

1.3.   Geology and Mineral Resource

     19  

1.4.   Metallurgical Testing and Processing

     21  

1.5.   Mining, Infrastructure, and Project Schedule

     22  

1.6.   Economic Analysis

     23  

1.7.   Conclusions

     25  

1.8.   Recommendations

     27  

2.  Introduction

     29  

2.1.   Registrant for Whom the Technical Report Summary was Prepared

     29  

2.2.   Terms of Reference and Purpose of this Technical Report

     29  

2.3.   Sources of Information

     29  

2.4.   Details of Inspection

     29  

2.5.   Report Version

     30  

2.6.   Units of Measure

     30  

2.7.   Coordinate Datum

     31  

2.8.   Mineral Resource and Mineral Reserve Definitions

     31  

2.9.   Qualified Person

     34  

3.  Property Description and Location

     35  

3.1.   Property Description

     35  

3.2.   Status of Mineral Titles

     36  

3.3.   Property Holding Costs

     39  

 

 

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3.4.   Environmental Liabilities

     40  

3.5.   Permits/Licenses

     40  

4.  Accessibility, Climate, Local Resources, Infrastructure, and Physiography

     41  

4.1.   Accessibility

     41  

4.2.   Climate

     41  

4.3.   Local Resources

     41  

4.4.   Infrastructure

     41  

4.5.   Physiography

     42  

5.  History

     44  

5.1.   Previous Owners

     44  

5.2.   Historic Exploration

     44  

5.3.   Historic Resource Estimates

     45  

5.4.   Historic Mining

     45  

6.  Geologic Setting, Mineralization and Deposit

     48  

6.1.   Regional Geology

     48  

6.2.   Local Geology

     49  

6.3.   Structural Geology

     55  

6.4.   Mineralization Controls

     56  

6.5.   Post Mineral Faulting

     57  

6.6.   Mineralization

     57  

6.7.   Deposit Types

     60  

7.  Exploration

     61  

7.1.    Geologic

     61  

7.1.1.  Resource Conversion Drilling

     65  

7.1.2.  Drilling

     66  

7.1.3.  Historic Drilling Methodology

     72  

7.1.4.  Current Drilling Methodology

     73  

7.1.5.  Sampling Methodology

     74  

7.1.6.  Core Recovery

     75  

 

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  Page vii . 

 

7.2.   Hydrogeology

     75  

7.2.1.  Sampling Methods and Laboratory Determinations

     75  

7.2.2.  Hydrogeology Investigations

     78  

7.2.3.  Hydrogeologic Description

     79  

7.2.4.  Mine Dewatering

     82  

7.2.5.  Dewatering Discharge

     83  

7.2.6.  Groundwater Flow Model

     84  

8.  Sample Preparation, Analysis and Security

     94  

8.1.   Pre-2012

     94  

8.1.1.  Sample Preparation Procedures

     94  

8.1.2.  Laboratory Analysis Procedures

     95  

8.1.3.  Security

     96  

8.2.   Premier 2012-2018

     96  

8.3.   i-80 2023-Present Underground Resource Conversion Drilling

     97  

8.4.   Quality Assurance and Quality Control

     97  

8.4.1.  Standards and Blanks

     97  

8.4.2.  Duplicate Assays

     101  

9.  Data Verification

     104  

10.  Mineral Processing and Metallurgical Testing

     106  

10.1.  Historical Metallurgical Test Work

     106  

10.1.1.  2008 KCA Program

     106  

10.1.2.  2009 KCA Program

     108  

10.1.3.  2017 SGS Programs

     108  

10.1.4.  2021 KCA Program

     113  

10.1.5.  2022 FLSmidth Program

     116  

10.2   Mineral Processing and Metallurgical Discussion

     117  

10.2.  QP Opinion

     117  

10.3.  Conclusions and Recommendations:

     117  

10.3.1.  Conclusions:

     118  

 

 

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10.3.2.  Recommendations

     118  

11.  Mineral Resource Estimates

     120  

11.1.  Introduction

     120  

11.2.  Modeling of Lithology and Mineralization

     125  

11.3.  Drill Data and Compositing

     129  

11.3.1.  Drill Data Set

     129  

11.3.2.  Compositing

     130  

11.4.  Density

     131  

11.5.  Statistics and Variography

     131  

11.6.  Grade Capping

     133  

11.7.  Block Model

     134  

11.8.  Grade Estimation and Resource Classification

     136  

11.9.  Mined Depletion and Sterilization

     138  

11.10. Model Validation

     139  

11.10.1.  Estimation Comparison

     139  

11.10.2.  Visual Comparison

     143  

11.10.3.  Swath Plots

     146  

11.10.4.  Model Smoothing Checks – Grade Tonnage Curves

     151  

11.11. Factors That May Affect Mineral Resources

     154  

11.12. Reasonable Prospects for Economic Extraction

     154  

11.12.1.  QP Opinion

     154  

11.13. Mineral Resources

     155  

12.  Mineral Reserve Estimates

     156  

13.  Mining Methods

     157  

13.1.  Mine Development

     157  

13.1.1.  Access Development

     157  

13.1.2.  Ground Support

     158  

13.1.1.  Ventilation and Secondary Egress

     160  

13.1.2.  Dewatering

     161  

 

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13.2.  Mining Methods

     161  

13.2.1.  Drift and Fill

     161  

13.3.  Underground Labor

     162  

13.4.  Mobile Equipment Fleet

     163  

13.5.  Mine Plan

     164  

14.  Recovery Methods

     169  

14.1.  INTRODUCTION

     169  

14.2.  Gold Quarry Roaster

     169  

14.3.  Lone Tree Pressure Oxidation Facility

     171  

14.3.1.  Lone Tree Mill Historic Processing

     171  

14.3.2.  Lone Tree Facility Block Flow Diagram

     171  

14.3.3.  Key Design Criteria

     172  

14.3.4.  Lone Tree Facility Description

     174  

14.3.5.  Utilities Consumption

     181  

15.  Infrastructure

     183  

15.1.  Dewatering

     183  

15.1.1.  History

     183  

15.1.2.  Infrastructure

     183  

15.2.  Electrical Power

     184  

15.3.  Mine Facilities

     186  

15.4.  Backfill

     187  

16.  Market Studies and Contracts

     189  

16.1.  Precious Metal Markets

     189  

16.2.  Contracts

     190  

16.2.1.  Private Placement Offering

     190  

16.2.2.  Orion and Sprott Financing Package

     192  

16.2.3.  Equinox Investment

     192  

16.2.4.  Private Placement

     192  

16.3.  Previous Financing Agreements

     192  

 

 

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16.3.1.  Offtake Agreement

     192  

16.3.2.  South Arturo Purchase and Sale Agreement (Silver)

     192  

16.4.  Roaster Toll Milling Agreement

     193  

16.5.  Other Contracts

     193  

17.  Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups

     194  

17.1.  Social or Community Impacts

     195  

17.2.  Permitting

     196  

17.3.  Closure and Reclamation Requirements

     197  

17.4.  Closure and Reclamation

     199  

17.5.  QP Statement

     200  

18.  Capital and Operating Costs

     201  

18.1.  Capital Costs

     201  

18.2.  Closure and Reclamation

     203  

18.3.  Operating Costs

     203  

18.4.  Cutoff Grade

     204  

19.  Economic Analysis

     206  

20.  Adjacent Properties

     215  

21.  Other Relevant Data and Information

     216  

22.  Interpretation and Conclusions

     217  

22.1.  Risks and Opportunities

     219  

22.2.  Work Program

     219  

23.  References

     223  

24.  Reliance on Information Provided by the Registrant

     226  

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  Page xi . 

 

List of Tables

 

Table 1-1 Cove Mineral Resources

     20  

Table 1-2 Financial Statistics

     24  

Table 2-1 Personal Inspections by Qualified Professionals

     30  

Table 2-2 Units of Measure

     31  

Table 2-3 QP Responsibility for Sections of This Report

     33  

Table 3-1 Property Holding Costs

     40  

Table 5-1 Historic Resource and Reserve Estimates

     46  

Table 5-2 Historic Cove and McCoy Mine Production 1986 through 2006

     46  

Table 7-1 List of Drilling by Operator

     67  

Table 7-2 Type of Drilling by Zone

     71  

Table 7-3 Summary of Hydrogeological Studies

     78  

Table 7-4 Calculated Model Calibration Statistics

     89  

Table 8-1 Pre-2012 ICP Analysis

     94  

Table 8-2 ICP Analysis 2012 - 2018

     95  

Table 8-3 Gold Blank and Standard Summary Statistics

     98  

Table 9-1 Data Review Summary

     105  

Table 10-1 Metallurgical Testing Programs

     106  

Table 10-3 Roasting Test Conditions

     109  

Table 10-4 Chlorination Program Test Results

     115  

Table 10-52022 FLSmidth Program Test Results

     116  

Table 11-1 Geology Codes

     126  

Table 11-2 Identification Codes for 3 g/t Grade Lenses

     126  

Table 11-3 Identification Codes for 0.2 g.t Grade Lenses

     127  

Table 11-4 Drill Hole Summary

     129  

Table 11-5 Composite Summary

     130  

Table 11-6 Density

     131  

Table 11-7 Gold Composite Statistics

     132  

Table 11-8 Silver Composite Statistics

     133  

Table 11-9 Composite Grade Capping

     133  

Table 11-10 Block Model Variables

     134  

Table 11-11 Estimation Parameters

     137  

Table 11-12 Classification Conditions

     137  

Table 11-13 Estimate Comparison for Gold versus a Nearest Neighbor at 0 Cutoff

     139  

Table 11-14 Estimate Comparison for Silver versus a Nearest Neighbor at 0 Cutoff

     141  

Table 11-15 Cove Mineral Resources

     155  

Table 13-1 Underground Workforce 2028 through 2029

     162  

 

 

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Table 13-2 Peak Underground Workforce beginning 2030

     163  
Table 13-3 Underground Mobile Equipment and Support Equipment for Exploration Development Phase      163  

Table 13-4 Underground Mobile Equipment and Support Equipment for Peak Production Mining

     163  

Table 13-5 Heading Productivity

     164  

Table 13-6 Annual Production and Development Schedule (Including Inferred Mineral Resource)

     164  

Table 13-7 Annual Production and Development Schedule (Excluding Inferred Mineral Resource)

     166  

Table 14-1 Summary of Key Process Statistics

     172  

Table 14-2 Lone Tree Facility Water Consumption by Type

     182  

Table 14-3 Lone Tree Facility Energy Usage by Area

     182  

Table 15-1 Backfill Scoping Tests 28-Day Unconfined Compressive Strength (psi)

     188  

Table 17-1 Cove Project Existing Permits

     196  

Table 17-2 McCoy Cove Reclamation Bonds

     199  

Table 17-3 Annual Closure and Reclamation Costs ($M)

     199  

Table 18-1 Project Capital Costs ($M)

     201  

Table 18-2 Mine Development Unit Costs

     201  

Table 18-3 Mine Development Capital ($M)

     202  

Table 18-4 Dewatering Capital ($M)

     202  

Table 18-5 Facilities and Site General ($M)

     202  

Table 18-6 Closure and Reclamation Costs ($M)

     203  

Table 18-7 Unit Operating Costs

     203  

Table 18-8 One Way Trucking Distance to Nevada Metallurgical Plants

     203  

Table 18-9 Operating and Capital Costs

     204  

Table 18-10 Cutoff Grades for Alkaline POX and Roaster

     204  

Table 19-1 Income Statement Includes Inferred) (Millions $US except Unit Cost per Ounce)

     207  

Table 19-2 Cash Flow Statement (Includes Inferred)

     207  

Table 19-3 Income Statement (Excludes Inferred) (Millions $US except Unit Cost per Ounce)

     208  

Table 19-4 Table 19-5 Cash Flow Statement (Excludes Inferred)

     208  

Table 19-6 Financial Statistics1

     209  

Table 22-1 Project Risks

     219  

Table 22-2 Opportunities

     219  

Table 22-3 Work Program Estimated Costs (US$M)

     220  

Table 24-1 Reliance on Information Provided by the Registrant

     226  

 

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Summary for the Cove Project, Lander County,

Nevada

  Page xiii . 

 

 

 

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  i-80 Gold Corp. 

 

List of Figures

 

Figure 1-1 Section View of Cove Mineralization looking NE

     20  

Figure 1-2 Project Timeline

     23  

Figure 3-1 Cove Location Map

     36  

Figure 3-2 Mining Claim Map

     38  

Figure 3-3 Summa and Chiara Royalties and Resource Location

     39  

Figure 4-1 McCoy-Cove Project Area Looking Southeast

     43  

Figure 6-1 Regional Geology

     51  

Figure 6-2 Triassic Stratigraphy and Mineralization

     52  

Figure 7-1 Exploration Targets

     62  

Figure 7-2 Barrick and Premier Exploration Drilling

     64  

Figure 7-3 Underground Exploration Potential

     65  

Figure 7-4 Resource Conversion Drill Program

     66  

Figure 7-5 Plan View of Drill Holes Used for the Current Analysis

     67  

Figure 7-6 Sample Section of CSD-Gap and Gap Hybrid Drilling

     68  

Figure 7-7 Sample Section of Helen Zone Drilling

     69  

Figure 7-8 Sample Section of CSD Zone Drilling

     70  

Figure 7-9 Sample Section of 2201 Zone Drilling

     71  

Figure 7-10. Vibrating Wire Piezometer and Groundwater Monitor Well Locations

     77  

Figure 7-11 Hydrogeologic Study Area/Model Area

     81  

Figure 7-12 Historical Cove Dewatering Rates

     83  

Figure 7-13 Historical Discharges to RIBs

     84  

Figure 7-14 Model Grid Extent

     86  

Figure 7-15 Representative Simulated and Observed Hydrographs around the Cove Pit Lake

     88  

Figure 7-16 Simulated and Observed Cove Pit Lake Stage

     89  

Figure 7-17 Projected Mine Elevation for Dewatering During Life of Mine

     90  

Figure 7-18 Simulated Water Levels During the Life of Mine

     91  

Figure 7-19 Simulated Dewatering Rates

     92  

Figure 7-20 Dewatering Well Field Locations Used For Predictive Modeling

     93  

Figure 8-1 Blank Assay Results

     99  

Figure 8-2 SP 37 Standard Reference Material Results

     99  

Figure 8-3 CDN-GS-22 Certified Reference Material Results

     100  

Figure 8-4 CDN-GS-5H Certified Reference Material Results

     100  

Figure 8-5 GS912-1 Certified Reference Material Results

     101  

Figure 8-6 Prep Duplicates - ALS Reno

     102  

Figure 8-7 Lab Check Duplicates

     103  

Figure 11-1 Plan View of Cove Mineralized Zones

     122  

 

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Nevada

  Page xv . 

 

Figure 11-2 Section View of Cove Mineralized Zones looking NE

     123  

Figure 11-3 Section View of the Helen Zone looking NW

     124  

Figure 11-4 Section View of the 2201 Zone looking NW

     125  

Figure 11-5 Low Grade Envelope

     127  

Figure 11-6 Section Looking AZ 315 Showing Mineralized Lenses in the GAP Zone

     128  

Figure 11-7 Sterilization Surfaces

     139  

Figure 11-8 Comparison of Composite and Estimated Block Gold Grades, Helen Zone

     143  

Figure 11-9 Comparison of Composite and Estimated Block Gold Grades, Gap Zone

     144  

Figure 11-10 Comparison of Composite and Estimated Block Gold Grades, CSD Zone

     145  

Figure 11-11 Comparison of Composite and Estimated Block Gold Grades, 2201 Zone

     145  

Figure 11-12 Gold Swath Plots of Helen Zone 3103

     146  

Figure 11-13 Silver Swath Plots of Helen Zone 3103

     147  

Figure 11-14 Gold Swath Plots of Gap Zone 2208

     148  

Figure 11-15 Silver Swath Plots of Gap Zone 2208

     149  

Figure 11-16 Gold Swath Plots of CSD Zone 1106

     149  

Figure 11-17 Silver Swath Plots of CSD Zone 1106

     150  

Figure 11-18 Gold Swath Plots of 2201 Zone 1302

     150  

Figure 11-19 Silver Swath Plots of 2201 Zone 1302

     151  

Figure 11-20 Helen Zone Grade Tonnage Plots

     152  

Figure 11-21 Gap Zone Grade Tonnage Plots

     152  

Figure 11-22 CSD Zone Grade Tonnage Plots

     153  

Figure 11-23 2201 Zone Grade Tonnage Plots

     153  

Figure 13-1 Exploration Development

     157  

Figure 13-2 Plan view showing portal, main haulages, and two raises to surface

     158  

Figure 13-3 Formation RQD

     159  

Figure 13-4 Ventilation Schematic

     160  

Figure 13-5 Depiction of Drift and Fill method

     162  

Figure 13-6 Production Profile (Including Inferred Mineral Resources)

     166  

Figure 13-7 Production Profile without Inferred Mineral Resources

     168  

Figure 14-1 Nevada Gold Mines Gold Quarry Roaster Simplified Flowsheet

     170  

Figure 14-2 Lone Tree Facility Block Flow Diagram

     173  

Figure 15-1 Proposed Dewatering Well Location

     184  

Figure 15-2 Electrical Demand

     185  

Figure 15-3 Electrical Site Plan

     186  

Figure 15-4 Mine Facilities Layout

     187  

Figure 16-1 Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average

     189  

Figure 19-1 Project Timeline

     206  

 

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp. 

 

Figure 19-2 Gold Production and Unit Costs (With Inferred)

     211  

Figure 19-3 Cash Flow Waterfall Chart Including Pre-Construction Costs (With Inferred)

     211  

Figure 19-4 Gold Production and Unit Costs (Without Inferred)

     212  

Figure 19-5 Cash Flow Waterfall Chart Including Pre-Construction Costs (Without Inferred)

     212  

Figure 19-6 NPV 5% Sensitivity

     213  

Figure 19-7 Profitability Index 5% Sensitivity

     213  

Figure 19-8 IRR Sensitivity

     214  

 

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 i-80 Gold Corp  

S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  Page xvii . 

 

List of Abbreviations

 

       

A

  

Ampere

  

kA

  

kiloamperes

   

AA

  

atomic absorption

  

kCFM

  

thousand cubic feet per minute

   

AGP

  

Acid Generation Potential

  

Kg

  

Kilograms

   

Ag

  

Silver

  

km

  

kilometer

   

ANFO

  

ammonium nitrate fuel oil

  

km2

  

square kilometer

   

ANP

  

Acid Neutralization Potential

  

kWh/t

  

kilowatt-hour per ton

   

Au

  

Gold

  

LoM

  

Life-of-Mine

   

AuEq

  

gold equivalent

  

m

  

meter

   

btu

  

British Thermal Unit

  

m2

  

square meter

   

°C

  

degrees Celsius

  

m3

  

cubic meter

   

CCD

  

counter-current decantation

  

masl

  

meters above sea level

   

CIL

  

carbon-in-leach

  

mg/L

  

milligrams/liter

   

CoG

  

Cut off grade

  

mm3

  

cubic millimeter

   

cm

  

centimeter

  

MME

  

Mine & Mill Engineering

   

cm2

  

square centimeter

  

Moz

  

million troy ounces

   

cm3

  

cubic centimeter

  

Mt

  

million tonnes

   

cfm

  

cubic feet per minute

  

MTW

  

measured true width

   

CRec

  

core recovery

  

MW

  

million watts

   

CSS

  

closed-side setting

  

m.y.

  

million years

   

CTW

  

calculated true width

  

NGO

  

non-governmental organization

   

°

  

degree (degrees)

  

NI 43-101

  

Canadian National Instrument 43-101

   

dia.

  

diameter

  

oz

  

Troy Ounce

   

EA

  

Environmental Assesment

  

opt

  

Troy Ounce per short ton

   

EIS

  

Environmental Impact Statement

  

oz/ton

  

Troy Ounce per short ton

   

EMP

  

Environmental Management Plan

  

%

  

percent

   

FA

  

fire assay

  

PLC

  

Programmable Logic Controller

   

Ft

  

Foot

  

PLS

  

Pregnant Leach Solution

   

Ft2

  

Square foot

  

PMF

  

probable maximum flood

   

Ft3

  

Cubic foot

  

POO

  

Plan of Operations

   

g

  

Gram

  

ppb

  

parts per billion

   

g/L

  

gram per liter

  

ppm

  

parts per million

   

g-mol

  

gram-mole

  

QAQC

  

Quality Assurance/Quality Control

   

g/t

  

grams per metric tonne

  

RC

  

reverse circulation drilling

   

ha

  

hectares

  

ROM

  

Run-of-Mine

   

HDPE

  

Height Density Polyethylene

  

RQD

  

Rock Quality Description

   

HTW

  

horizontal true width

  

SEC

  

U.S. Securities & Exchange Commission

   

ICP

  

induced couple plasma

  

Sec

  

second

   

ID2

  

inverse-distance squared

  

SG

  

specific gravity

   

ID3

  

inverse-distance cubed

  

SPT

  

Standard penetration test

   

mm

  

millimeter

  

ton

  

US Short Ton

   

mm2

  

square millimeter

  

Tonne

  

Metric Tonne

 

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp 

 

  1.

Summary

 

  1.1.

Introduction

Practical Mining LLC (Practical or PM) was engaged by i-80 Gold Corp., and Premier Gold Mines USA, Inc. (together i-80, Premier or the Company) to prepare an Initial Assessment (IA) Technical Report Summary (TRS) on the McCoy Cove Project (Cove or The Project) in Lander County, Nevada. This Technical Report (TRS) has been prepared in accordance with Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items and 1300 through 1305).

This TRS dated the 26th day of March 2025 with an effective date of December 31, 2024 is the initial statement of mineral resources for i-80 Gold Corporation (i-80) under S-K regulations. The Company has previously reported mineral resources under Canadian NI 43-101 regulations. This TRS presents an underground mine plan, metallurgical testing, hydrogeologic summary, and financial analysis for the proposed Cove Underground Mine.

The mineral resource estimation generated for the 2021 Preliminary Economic Assessment (PEA) included only drillholes completed prior to January 1, 2021. Since then, i-80 has completed roughly 5,700 feet of underground drift development. A resource conversion drill program consisting of 125 drill holes totaling approximately 140,000 feet is ongoing and expected to be completed in Q3 2025. The resource estimation will be updated following the completion of the drill program. This TRS updates the mineral resource estimate with current metal pricing and cost estimates.

Cautionary Notes:

 

  1.

The financial analysis contains certain information that may constitute “forward-looking information” and “forward-looking statements” (together, “forward-looking information”) under applicable Canadian and United States securities legislation. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors

 

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 i-80 Gold Corp  

Summary

  Page 19 

 

 

which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in i-80’s management information circular under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this report is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. i-80 disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law; and

  2.

This IA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the IA will be realized.

 

  1.2.

Property Description

The Cove Project covers 32,110 acres and is located 32 miles south of the Town of Battle Mountain, in the Fish Creek Mountains of Lander County, Nevada. It is centered approximately at 40°22’ N and 117°13’ W and lies within the McCoy Mining District. The project area contains 1,728 unpatented and nine patented mining claims owned 100% by i-80, through its wholly-owned subsidiaries Premier Gold Mines USA, Inc. and Au-Reka Gold Corp. The unpatented claims are on land administered by the Bureau of Land Management (BLM).

The Cove deposit consists of the Helen, Gap, CSD, and 2201 zones. They are located beneath the historically mined Cove open pit and extend approximately 2,000 feet northwest from the pit. The Cove deposit was mined by Echo Bay Mines Ltd. (Echo Bay) between 1987 and 2001. During this period the Cove deposit produced 2.6 million ounces of gold and 100 million ounces of silver. Gold and silver production from heap leach pads continued until 2006.

 

  1.3.

Geology and Mineral Resource

The Cove Project contains four structurally controlled mineralized zones within the Triassic sedimentary package. The Helen and Gap zones are Carlin Style disseminated refractory gold deposits. The Cove South Deep (CSD) gold and silver mineralization is associated with disseminated sulfides and is characterized by Ag:Au ratios of 50:1 to over 100:1. The 2201 zone is comprised of sulfide mineralization within sheeted stockwork veins with high concentrations of lead and zinc (Figure 1-1).

 

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp 

 

Figure 1-1 Section View of Cove Mineralization looking NE

LOGO

Mineral Resources are constrained to high-grade wireframe models constructed at a nominal 0.09 opt (3 g/t) grade shell. The mineral resource estimate relies on data from 387 core drill holes totaling 548,038 feet and 1,010 reverse circulation (RC) drill holes totaling 579,443 feet. From these drill holes, 3,146 samples were flagged within the high-grade wireframes to be used in grade estimation.

Parent block dimensions are 100 ft x 100 ft x 100 ft with sub-block dimensions as small as 1 ft x 1 ft x 1 ft. Block grades were estimated using Inverse Distance Cubed (ID3) methods.

A block is classified as Indicated if there are at least two composites within an average distance of 100 feet or less and at least one of the samples is within fifty feet. A block is classified Inferred if there are at least two composites within 300 feet but more than 100 feet. Cove Mineral Resources as of December 31, 2024 are presented in Table 1-1.

Table 1-1 Summary of Cove Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

                 
      Tons
(000)
  

 Tonnes 

 (000) 

  

  Au    

(opt) 

  

  Au  

  g/t  

  

  Ag  

  (opt)  

  

  Ag  

  (g/t)  

  

 Au ozs 

 (000) 

  

 Ag ozs 

 (000) 

   
      Indicated Mineral Resource
   

Helen

   743    674    0.271    9.3    0.074    2.6    201    55
   

Gap

   280    254    0.219    7.5    0.239    8.9    61    72
   

CSD

   275    249    0.175    6.0    1.603    55.0    48    441

 

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Summary

  Page 21 

 

                 
      Tons
(000)
  

 Tonnes 

 (000) 

  

  Au    

 (opt)  

  

  Au  

  g/t  

  

  Ag  

  (opt)  

  

  Ag  

  (g/t)  

  

 Au ozs 

 (000) 

  

 Ag ozs 

 (000) 

   

Total Indicated

   1,298    1,177    0.239    8.2    0.438    15.0    310    568
   
      Inferred Mineral Resource
   

Helen

   1,743    1,582    0.245    8.4    0.083    2.9    427    146
   

Gap

   2,229    2,022    0.244    8.4    0.262    9.0    543    585
   

CSD

   319    290    0.173    5.9    1.685    57.8    55    538
   

2201

   168    153    0.780    26.7    1.016    34.8    131    171
   

Total Inferred

   4,459    4,047    0.259    8.9    0.323    11.1    1,156    1,439

Notes:

 

  1.

Mineral resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of 27.25 per troy ounce. Refer to Section 16.1for a discussion on metal pricing;

 

  2.

Mineral resources have been estimated using gold metallurgical recoveries ranging from 73.2% to 93.3% for roasting and 78.5% to 95.1 % for pressure oxidation;

 

  3.

Roaster cutoff grades range from 4.15 to 5.29 Au g/t (0.121 to 0.154 opt) and pressure oxidation cutoff grades range from 3.83 to 4.64 Au g/t (0.112 to 0.135 opt);

 

  4.

The effective date of the mineral resource estimate is December 31, 2024;

 

  5.

Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors;

 

  6.

An inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An inferred mineral resource has a lower level of confidence than that applying to an indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration; and

 

  7.

The reference point for mineral resources is in situ.

 

  1.4.

Metallurgical Testing and Processing

Metallurgical testing of Cove samples dates back to the 1960’s. The most recent relevant testing are programs in 2008 and 2009 at Kappes Cassiday Associates and 2017 at SGS Lakefield Research.

The testing has generally shown that Helen and Gap resources based on the composites tested appear to be generally refractory to conventional whole ore cyanidation and will need some type of oxidation process to significantly increase gold extractions over whole ore cyanidation. Most of the samples were also double refractory where gold is both bound within sulfides and active carbonaceous matter which adsorbs dissolved gold (known as “preg-robbing”) faster than activated carbon. Both the sulfides and the preg-robbing carbonaceous matter requires higher temperature oxidation which can be achieved through roasting followed by carbon-in-leach. The testing showed that Helen Zone samples are generally more amenable to roasting and carbon-in-

 

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp 

 

leach (CIL) processing. Gap Zone samples were more amenable to treatment using pressure oxidation and residue CIL processing.

The 2017 data set was too small to establish any clear relationships between mineralogy, head grades, and recoveries for either Helen or Gap samples. although it is clear that mineralogy factors such as arsenic content and total carbonaceous matter or total organic carbon influence recoveries. Improved recovered were achieved using either roasting and calcine leaching/CIL or pressure oxidation followed by leaching or CIL compared to direct leaching or CIL alone.

In 2020, testing was commenced on Helen and Gap samples to evaluate chlorination as a lower cost method for sulfide oxidation and passivation of active carbonaceous matter. This method was previously employed at several Nevada operations. This method was not pursued due to higher than anticipated operating cost estimates.

 

  1.5.

Mining, Infrastructure, and Project Schedule

Access to the mineralized zones will be through a portal located just north of the Cove Pit. Primary development totals 23,048 feet with gradients up to +/- 15%. Ventilation and secondary egress will be gained through a ventilation intake portal in the southwest pit wall and an exhaust raise equipped with a personal hoist for evacuation.

Drift and fill mining with mining heights of fifteen feet will be the primary method for extraction of the Helen and Gap mineral resource. Where the mineralized lenses thicken, breasting the sill or back can recover additional mineralization. Waste rock from development and waste reclaimed from historic dumps will be used for Cemented Rock Fill (CRF) or unconsolidated (GOB) fill as appropriate to achieve high levels of extraction. Development and production mining will be performed by a qualified mining contractor thus reducing the capital requirements for the Project.

A trucking contractor will transport mineralization mined over local, state, and federal roads for processing at a third party roaster or at i-80’s Lone Tree pressure oxidation (POX) facility.

Dewatering will be accomplished using 15 surface wells and a pit lake barge producing up to 43,000 gallons per minute (gpm) Dewatering water will be piped to several Rapid Infiltration Basins (RIBs) constructed at the northern project boundary. The RIB locations have been selected to prevent recharge into the Cove hydrogeologic system. During the summer irrigation season water will be provided to local ranchers for irrigation of alfalfa crops.

Long term electrical power demands up to 11.5 MW will be supplied by NV Energy via an existing 120kV transmission line which connects the Project site to NV Energy’s Bannock substation. A new substation and 13.8kV distribution system will be constructed at Cove.

 

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 i-80 Gold Corp  

Summary

  Page 23 

 

Power for initial mine development and underground delineation drilling will be provided from an existing 24.9kV distribution line that also terminates on the property. A substation for the 24.9kV line and a distribution line to the portal site was constructed in 2019. Power will be stepped down from the 120kV substation and use the same distribution line.

Baseline environmental studies have been undertaken to facilitate the National Environmental Protection Act (NEPA) administered by federal land management agencies. This, along with dewatering, constitutes the critical path to production. The overall project timeline is shown in Figure 1-2.

Figure 1-2 Project Timeline

LOGO

 

  1.6.

Economic Analysis

Capital spending over the life of the project is subdivided into three categories. Pre-development spending of $17.3M encompasses definition drilling, baseline data collection, engineering, hydrogeologic investigations, metallurgical testing and permitting. Construction capital is required for Helen and Gap dewatering, infrastructure and mine development and is projected at $157.4 M over a two-year period commencing in 2028. Sustaining capital includes mine development and facilities. Sustaining capital totals $49.1M commencing in 2030.

Gold recovery will total 740,000 ounces over the eight-year mine production life. Material mined for processing averages 0.305 Au opt. Full production is reached after two years of ramp up in 2031 and averages 1,190 tpd from 2031 through 2036.

 

 

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S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp 

 

The Helen and Gap zones contain 70% and 89% inferred mineral resources respectively. The results without inferred are the result of a gross factorization of the production stream. There has been no adjustment to capital development, dewatering capital or mine facilities capital. Furthermore, there has not been any recalculation of productivities or operating costs due to the lower production rates.

Table 1-2 Financial Statistics

         With Inferred         Without  Inferred  

Gold price (US$/oz)

   $2,175

Silver price (US$/oz)

   $27.25

Mine life (years)

   8

Average mineralized mining rate (tons/day)

   1,010    200

Average grade (oz/t Au)

   0.305    0.313

Average gold recovery (roaster %)

   79%    79%

Average gold recovery (autoclave %)

   86%    86%

Average annual gold production (koz)

   92    19

Total recovered gold (koz)

   740    148

Pre-development capital ($M)

   $17.2    $17.2

Mine construction capital ($M)

   $157.4    $157.4

Sustaining capital (M$)

   $49.1    $49.1

Construction Start Date

   1/1/2028
Economic Indicators Post Construction Decision 3

Cash cost (US$/oz) 1

   $1,201    $1,697

All-in sustaining cost (US$/oz)2

   $1,310    $2,240

All in cost (US$/oz) 5

   $1,635    $3,302

Project after-tax NPV5% (M$)

   $274    ($160)

Project after-tax NPV8% (M$)

   $216    ($159)

Project after-tax IRR

   30%    NA

Payback Period

   5.5 Years    NA

Profitability Index 5%4

   2.4    0.2
Financial Statistics Including Pre-Construction Period (1/1/2025 – 12/31/2027) 6

All in cost (US$/oz) 5

   $1,658    $3,419

 

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 i-80 Gold Corp   Summary   Page 25 

 

         With Inferred         Without  Inferred  

Project after-tax NPV5% (M$)

   $256    ($178)

Project after-tax NPV8% (M$)

   $198    ($177)

Project after-tax IRR

   25%    NA

Payback Period

   6.8 years    NA

Profitability Index 5%4

   1.9    0.2

Notes:

  1.

Net of byproduct sales;

 

  2.

Excluding income taxes, pre-construction capital, construction capital, corporate G&A, corporate taxes and interest on debt;

 

  3.

Discounted to 2028, Construction Start;

 

  4.

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates breakeven;

 

  5.

Excluding corporate G&A, corporate taxes and interest on debt;

 

  6.

Discounted to 2025;

 

  7.

This IA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the IA will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability;

 

  8.

Inferred mineral resources constitute 70 % of the Helen zone and 89% of the Gap zone. The “Without Inferred” statistics presented are a gross factorization of the mine plan without any redesign of mine excavations or recalculation of productivities and costs. Capital costs are the same for the “With Inferred” and “Without Inferred” scenarios. The “Without Inferred” scenario is presented solely to illustrate the projects dependence on inferred mineral resources.

 

  9.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable Canadian and United States securities regulations. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law;

 

  1.7.

Conclusions

Metallurgical Testing

 

 

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  1.

Head assaying for both the Helen Zone and Gap indicated that the gold in the two resources will likely be finely disseminated and not amenable to gravity gold recovery;

 

 

  2.

The mineralogy of the Helen and Gap resources differ in two significant areas, the first being that the Helen appears to be lower in arsenic content than the Gap resource and that the Gap resource appears to be lower on average in TCM and TOC than the Helen resource;

 

 

  3.

The Helen composite arsenic assays indicate the mineral resources in Helen are lower in arsenic content than those in the Gap;

 

 

  4.

Based on the composites tested the Helen Zone appears to generally be more amenable to roasting and CIL processing;

 

 

  5.

Based on the composites tested, the Gap resource appears to generally be more amenable to pressure oxidation and CIL processing; and,

 

 

  6.

The data set was too small to establish any clear relations between mineralogy, metal head grade, and extractions for either resource although it is clear that mineralogy factors such as arsenic content and TCM or TOC are influencing extractions.

Toll Processing

 

  1.

The feed specifications appear to be somewhat rigid and could preclude some material being sent to the toll processor or result in penalties. Blending may allow shipment of some off-specification material provided appropriate material is available for onsite blending prior to shipping to the toll processor;

 

  2.

The terms appear to be consistent and typical with those encountered in the industry; and,

 

  3.

The recovery terms appear to be the result of analyzing the metallurgical data provided by i-80 Gold.

Mining and Infrastructure

 

  1.

Mining conditions typical for sedimentary deposits in the northeastern Nevada extensional tectonic environments are anticipated;

  2.

Dewatering the Helen and Gap zones will require up to fifteen wells and a pit lake barge pumping up to 43,000 gpm.

Financials

 

  1.

Capital requirements total $206.5M excluding $17.3M in pre-construction capital;

  2.

The project achieves NPV 5% of $274M and NPV 8% of $216M (excluding preconstruction capital);

 

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 i-80 Gold Corp  

Summary

  Page 27 

 

  3.

When including the pre-construction capital, the NPV 5% reduces to $256M and the NPV 8% is $198M; and,

  4.

The estimated payback period is 5.6 years with an IRR of 30%.

 

  1.8.

Recommendations

Resource Delineation and Exploration

 

  1.

Resource delineation drilling will be completed in 2025 along with an updated resource model and feasibility study.

  2.

The existing Cove Pit prohibits drilling the Gap extension area and portions of the Gap deposit. These are the most prospective nearby areas for adding significant Mineral Resources; and,

  3.

Expansion of the 2201 Zone could add high grade mineralization to the project which could be accessed through the Helen and Gap infrastructure.

Dewatering

 

  1.

A detailed hydrogeologic study has been completed along with estimates of pumping rates, water table drawdown estimates and construction cost estimates. This information will be included in the upcoming feasibility study.

Mining

 

  1.

A geotechnical data collection program is ongoing. Characterization of geotechnical parameters should be included in the upcoming feasibility study:

 

  a.

The objectives of the program are to characterize the mining horizons using the Rock Mass Rating (RMR) system;

 

  b.

Collect downhole Acoustic Tele Viewer (ATV) drill logs to collect joint orientation data for mine designs and accurately estimate ground support requirements; and,

 

  c.

Collect full core samples for physical rock property testing.

 

  2.

Complete additional testing of potential back fill sources to optimize the Cemented Rock Fill (CRF) mix design; and,

 

  3.

Complete a ventilation simulation to predict Diesel Particulate Matter (DPM), carbon monoxide, and other contaminant concentrations.

Metallurgical Testing

 

  1.

Additional metallurgical testing will be needed to thoroughly investigate the variability and viability of Helen and Gap resources to evaluate pressure oxidation with CIL cyanidation

 

 

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  i-80 Gold Corp 

 

 

under Lone Tree conditions. Testing should also include baseline CIL tests and roasting testing as a comparison. Sampling objectives will include:

 

   

Samples from GAP and Helen zones and their major lithological units; Favret, Panther Dolomite. The samples should also address spatial variability within each zone.

 

   

Sample intrusive formations in each zone.

 

   

Assess variability of the responses to roasting and calcine cyanidation across the resources;

 

   

Assess variability of the responses to pressure oxidation and residue cyanidation across the resources;

 

   

Testing should attempt to establish head grade and extraction relations for use in more detailed resource modelling;

 

   

Mineralogy impacts need to be established and geologic domains within each resource need to be determined, and;

 

   

Additional comminution data should be collected to assess hardness variability within the zones and any potential impacts on throughput in the Lone Tree process plant.

 

  2.

The resource model should be advanced to include arsenic, TCM, TOC, mercury, lead, zinc, total copper selenium, barium, cobalt, nickel, and cadmium as these will be important for predicting grades if toll process offsite is used and potentially for estimating extractions within the resources;

 

  3.

The estimated cost for the suggested next phase metallurgical program is to $850,000 based on current market pricing.

Permitting and Development Decision

 

  1.

Baseline data collection in support of the Environmental Impact Statement should be done simultaneously to reduce the Project’s critical path and bring forward production; and

  2.

The project should proceed directly with a feasibility study to support a development decision.

 

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 i-80 Gold Corp  

Introduction

  Page 29 

 

  2.

Introduction

 

  2.1.

Registrant for Whom the Technical Report Summary was Prepared

This TRS was prepared for i-80 Gold Corporation and its subsidiaries Premier Gold Mines USA, Inc. and AuReka Gold Corporation (together i-80) in accordance with the requirements of the Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305) for i-80.

 

  2.2.

Terms of Reference and Purpose of this Technical Report

This Technical Report Summary is i-80’s initial statement of mineral resources for the Cove Project under regulation S-K 1300 and presents an Initial Assessment (IA) based on indicated and inferred mineral resources.

The quality of information, conclusions, and estimates contained herein are based on: i) information available at the time of preparation and ii) the assumptions, conditions, and qualifications set forth in this report. This IA is a preliminary technical and economic study of the economic potential of all or parts of mineralization to support the disclosure of mineral resources. This IA is preliminary in nature. It includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the Initial Assessment will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

 

  2.3.

Sources of Information

This report is based in part on internal Company technical reports, previous studies, maps, published government reports, Company letters and memoranda, and public information as cited throughout this report and listed in Section 0 References. Reliance on Information Provided by the Registrant is listed in the Section 24 when applicable.

2.4. Details of Inspection

Table 2-1 summarizes the details of the personal inspections on the property by each qualified person or, if applicable, the reason why a personal inspection has not been completed

 

 

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Table 2-1 Personal Inspections by Qualified Professionals

Company    Discipline    Dates of Personal
Inspection
   Details of Inspection
Practical Mining    Mining, Mineral Resources and Mineral Reserves    October 16, 2024    Site specific hazard training, examined core and core logging procedures, examined underground exploration decline, observed core drilling operations
Raponi Engineering    Metallurgical Testing and Mineral Processing    None    The Cove Project does not have facilities for mineral processing.
Montgomery and Associates    Hydrogeology          

 

  2.5.

Report Version

This TRS presents the inaugural statement of the Cove Project mineral resources by i-80 under 17 CFR § 229.1300. The Company has most recently disclosed mineral resources for the project under Canadian Securities NI 43-101 regulations with the report Titled “Preliminary Economic Assessment for the Cove Project, Lander County, Nevada” dated January 25, 2021.

 

  2.6.

Units of Measure

U.S. Imperial units of measure are used throughout this document unless otherwise noted. The units of measure used in this report are shown in Table 2-2. Currency is expressed as United States Dollars unless otherwise noted.

 

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Introduction

  Page 31 

 

Table 2-2 Units of Measure

US Imperial to Metric conversions

Linear Measure

1 inch = 2.54 cm

1 foot = 0.3048 m

1 yard = 0.9144 m

1 mile = 1.6 km

Area Measure

1 acre = 0.4047 ha

1 square mile = 640 acres = 259 ha

Weight

1 short ton (st) = 2,000 lbs = 0.9071 metric tons

1 lb = 0.454 kg = 14.5833 troy oz

Assay Values

1 oz per short ton = 34.2857 g/t

1 troy oz = 31.1036 g

1 part per billion = 0.0000292 oz/ton

1 part per million = 0.0292 oz/ton = 1g/t

 

  2.7.

Coordinate Datum

Spatial data utilized in analysis presented in this TR are projected to UTM Zone 11 North American Datum 1983 feet. All spatial measurements are in international survey feet.

 

  2.8.

Mineral Resource and Mineral Reserve Definitions

The terms “mineral resource” and “mineral reserves” as used in this Technical Report Summary have the following definitions.

Mineral Resources

17 CFR § 229.1300 defines a “mineral resource” as a concentration or occurrence of material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. A mineral resource is a reasonable estimate of mineralization, taking into account relevant factors such as cut-off grade, likely mining dimensions, location or continuity, that, with the assumed and justifiable technical and economic conditions, is likely to, in whole or in part, become economically extractable. It is not merely an inventory of all mineralization drilled or sampled.

A “measured mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of conclusive geological evidence and sampling. The level of

 

 

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geological certainty associated with a measured mineral resource is sufficient to allow a qualified person to apply modifying factors, as defined in this section, in sufficient detail to support detailed mine planning and final evaluation of the economic viability of the deposit. Because a measured mineral resource has a higher level of confidence than the level of confidence of either an indicated mineral resource or an inferred mineral resource, a measured mineral resource may be converted to a proven mineral reserve or to a probable mineral reserve.

An “indicated mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of adequate geological evidence and sampling. The level of geological certainty associated with an indicated mineral resource is sufficient to allow a qualified person to apply modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Because an indicated mineral resource has a lower level of confidence than the level of confidence of a measured mineral resource, an indicated mineral resource may only be converted to a probable mineral reserve.

An “inferred mineral resource” is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. The level of geological uncertainty associated with an inferred mineral resource is too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability. Because an inferred mineral resource has the lowest level of geological confidence of all mineral resources, which prevents the application of the modifying factors in a manner useful for evaluation of economic viability, an inferred mineral resource may not be considered when assessing the economic viability of a mining project and may not be converted to a mineral reserve.

Mineral Reserves

17 CFR § 229.1300 defines a “mineral reserve” as an estimate of tonnage and grade or quality of indicated and measured mineral resources that, in the opinion of the qualified person, can be the basis of an economically viable project. More specifically, it is the economically mineable part of a measured or indicated mineral resource, which includes diluting materials and allowances for losses that may occur when the material is mined or extracted.

This Initial Assessment was compiled by Practical Mining LLC (Practical or PM), Raponi Engineering (Raponi) and Montgomery and Associates (M&A). All three firms are third-party firms comprising experts in their respective fields in accordance with 17 CFR § 229.1302(b)(1). i-80 has determined that all three firms meet the qualifications specified under the definition of qualified person in 17 CFR § 229.1300

 

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 i-80 Gold Corp  

Introduction

  Page 33 

 

Table 2-3 QP Responsibility for Sections of This Report

Section     Title    QP Firm
1.1.   

Introduction

  

Practical, Raponi

1.2.   

Property Description

  

Practical

1.3.   

Geology and Mineral Resource

  

Practical

1.4.   

Metallurgical Testing and Processing

  

Raponi

1.5.   

Mining, Infrastructure, and Project Schedule

  

Practical

1.6.   

Economic Analysis

  

Practical

1.7.   

Conclusions

  

Practical, Raponi, M&A

1.8.   

Recommendations

  

Practical, Raponi M&A

2   

Introduction

  

Practical

3   

Property Description and Location

  

Practical

4   

Accessibility, Climate, Local Resources, Infrastructure, and Physiography

  

Practical

5   

History

  

Practical

6   

Geologic Setting, Mineralization and Deposit

  

Practical

7   

Exploration

  

Practical

7.1   

Geologic

  

Practical

7.2   

Hydrogeologic

  

M&A

8   

Sample Preparation, Analysis and Security

  

Practical

8.1.   

Pre-2012

  

Practical

9   

Data Verification

  

Practical

10   

Mineral Processing and Metallurgical Testing

  

Raponi

11   

Mineral Resource Estimates

  

Practical

12   

Mineral Reserve Estimates

  

Practical

13   

Mining Methods

  

Practical

14   

Recovery Methods

  

Raponi

15   

Infrastructure

  

Practical

16   

Market Studies and Contracts

  

Practical

17    Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups   

Practical

18.1.   

Capital Costs

  

Practical, Raponi

18.2.   

Closure and Reclamation

  

Practical

18.3.   

Operating Costs

  

Practical, Raponi

18.4.   

Cutoff Grade

  

Practical

19   

Economic Analysis

  

Practical

20   

Adjacent Properties

  

Practical

21   

Other Relevant Data and Information

  

Practical

22   

Interpretation and Conclusions

  

Practical, Raponi, M&A

23   

Recommendations

  

Practical, Raponi, M&A

 

 

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Section     Title    QP Firm
24   

References

  

Practical, Raponi, M&A

25   

Reliance on Information Provided by the Registrant

  

Practical, Raponi, M&A

 

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 i-80 Gold Corp  

Property Description and Location

  Page 35 

 

  3.

Property Description and Location

 

  3.1.

Property Description

The McCoy-Cove Project covers 32,110 acres and is located 32 miles south of the Town of Battle Mountain, in the Fish Creek Mountains of Lander County, Nevada. It is centered approximately at 40°22’ N and 117°13’ W and lies within the McCoy Mining District (Figure 3-1).

The Cove Mineral Resource consists of the Helen, Gap, CSD, and 2201 deposits. They are located beneath the historically mined Cove open pit and extend approximately 2,000 feet northwest from the pit. The historic McCoy open pit is located approximately 0.6 mi to the southwest. The Cove deposit was mined by Echo Bay Mines Ltd. (Echo Bay) between 1987 and 2001 and produced 2.6 million ounces of gold and 100 million ounces of silver. McCoy was mined between 1986 and 2001 and produced approximately 0.88 million ounces of gold and 3.0 million ounces of silver. Gold and silver production from heap leach pads continued until 2006. The McCoy-Cove project area also includes several exploration targets.

The Project is located on federal land administered by the US Department of Interior - Bureau of Land Management (BLM) and patented mining claims.

 

 

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  i-80 Gold Corp 

 

Figure 3-1 Cove Location Map

LOGO

 

  3.2.

Status of Mineral Titles

The McCoy-Cove Project consists of 1,727 100%-owned unpatented claims and nine owned patented claims. The claim map provided by i-80 is shown in Figure 3-2.

On June 15, 2006, Victoria Gold Corporation (Victoria) entered into a “Minerals Lease and Agreement” to lease a portion of the Project from Newmont. Under the terms of the Minerals Lease and Agreement, Victoria was subject to escalating yearly work commitments in the aggregate amount of $8.5 million over a period of seven years (consisting of $0.3 million, $0.7 million, $1.0 million, $1.25 million, $1.5 million, $1.75 million, and $2.0 million, respectively, in each year of the first seven years of the agreement dated June 15, 2006), of which $1.0 million was a firm obligation and was to be expended by June 15, 2008 (completed). Excess expenditures were allowed to be carried forward. Newmont acknowledged that Victoria spent over $9.1 million in exploration at the Project between June 15, 2006 and March 16, 2009 and satisfied the work commitment of Section 2(a) of the Minerals and Lease Agreement.

 

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Property Description and Location

  Page 37 

 

On June 14, 2012, Premier, through its wholly owned subsidiary, Au-Reka Gold Corporation (Au-Reka Gold), acquired a 100% interest in the Cove portion of the Project from Victoria pursuant to an asset purchase agreement dated June 4, 2012 (Cove Purchase Agreement). In connection with the acquisition, Premier paid an aggregate of C$8,000,000 on closing, C$4,000,000 of which was paid in cash and the balance of which was satisfied by the issuance of 892,857 common shares of Premier. In addition, Premier issued a promissory note (Cove Acquisition Promissory Note) in the amount of C$20,000,000 payable in C$10,000,000 allotments on the first and second anniversary dates of the closing date of the acquisition. The Cove Promissory Note was repaid in full in June 2014. The Company also reimbursed Victoria in the amount of $1,206,277 in respect of exploration and related activities conducted on the Cove portion of the Project between March 15, 2012 and the closing of the transaction.

Pursuant to the Cove Purchase Agreement in the event of production from the Cove portion of the Project, Premier will make additional payments to Victoria in the aggregate amount of C$20,000,000 (consisting of cash and/or the equivalent value of Premier common shares, at Premier’s option), payable in four installments of $5,000,000 each upon the cumulative production, to Premier’s account, of 250,000, 500,000, 750,000 and 1,000,000 troy ounces of gold from the Cove potion of the Project (Deferred Bullet Payment Consideration).

In September 2014, Premier entered into an agreement with Newmont to acquire a 100% interest in the property. Upon closing of the transaction, Premier paid Newmont $15 million, replaced bonding of approximately $4 million via a surety policy, and transferred to Newmont all land sections that comprised the South Carlin Property. In addition, Premier made staged payments to Newmont over 18 months equal to $6 million. Additional details of the transaction included the elimination of Newmont’s previous “back-in” rights to the Project, a 10-year good faith milling agreement for ores mined at McCoy-Cove and retention of a 1.5% NSR in the property.

Premier entered into an earn-in agreement (Barrick Earn-In Agreement) dated December 11, 2017, but effective January 8, 2018, with certain subsidiaries of Barrick Gold Corporation (Barrick).

Pursuant to the Barrick Earn-In Agreement, Barrick had the option to earn a 60% interest in the exploration portion of the Project (McCoy Joint Venture Property) by spending $22.5 million in exploration before June 30, 2022 (Barrick Earn-in). The McCoy Joint Venture Property excludes the “Cove Deposits” (being the claims within the “Carveout”) portion of the Project which was retained solely by Premier.

On December 16, 2020 Premier Gold Mines announced that it entered into an agreement with Equinox Gold whereby Equinox Gold would acquire all the outstanding shares of Premier. At Closing, Premier would spin-out i-80 Gold Corporation. i-80 was to include 100% of Premier’s interest in the McCoy/Cove Project.

 

 

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  i-80 Gold Corp 

 

Figure 3-2 Mining Claim Map

LOGO

Cove - mccoy Land Position map

 

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 i-80 Gold Corp  

Property Description and Location

  Page 39 

 

On February 6, 2020 Barrick formally terminated the Earn-In Agreement and all land subject to it is once again 100% held by Premier.

On May 17, 1977 Houston Oil and Minerals granted a 2% net Smelter return royalty to the Summa Corporation. The royalty applies to the portion of the claims in existence on that date.

A 2% net smelter return royalty was collectively reserved by Robert Chiara, et. al., on the Lone Star 1 – 4 claims. These claims do not contain any of the Cove Mineral Resource. Figure 3-3 shows the area covered by the Summa and Chiara Royalties.

Figure 3-3 Summa and Chiara Royalties and Resource Location

LOGO

 

  3.3.

Property Holding Costs

Unpatented claims have annual maintenance fees of $200 per claim payable to the Bureau of Land Management and a notice of intent to hold (NIH) in the amount of $12 per claim payable to Lander

 

 

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  i-80 Gold Corp 

 

County. The BLM MLRS mining claim database shows all claim fees paid through September 2025. County NIH fee payments are current.

Patented claims are subject to property taxes, which were current at the time of this report. There was also a special assessment for Reese River Basin water rights in 2024, which was paid. Annual property holding costs are shown in Table 3-1.

Table 3-1 Property Holding Costs

       
Description    Payee    Quantity     Amount  
       
Unpatented Claim Maintenance Fee    BLM    1,728      $345,600.00  
       
Notice of Intent to Hold    Lander County       1,728      $20,748.00  
       
Patented Claim Property Taxes    Lander County    9      $281.20  
       
2024 Reese River Basin water right special assessment    Lander County         $10,578.73  
                    
       
Total                $377,179.81  

 

  3.4.

Environmental Liabilities

The Project was under active reclamation by Newmont from 2003 to 2014. Activities include re-contouring and seeding of the dumps, leach pads, and tailings facility. All surface infrastructure outside of the maintenance shop and guard shack has been removed.

Au-Reka is responsible for all environmental liabilities related to the closure of the McCoy-Cove Project as well as final clean-up of surface drill pads and minor drill roads. All closure activities other than evaporation of the tailings facility, reclamation of water treatment and storage ponds, reclamation of exploration drill pads, and water quality testing have been temporarily put on hold pending the potential for future production out of the Cove-Helen underground.

The authors are not aware of any additional environmental liabilities on the property. The authors are not aware of any other significant factors and risks that may affect access, title, or the right or ability to perform the proposed work program on the property.

 

  3.5.

Permits/Licenses

Currently, Au-Reka is working under the Cove-Helen Underground Mine Project Plan of Operations (POO No. NVN-088795) approved 2018. The POO authorizes Au-Reka to complete up to 100 acres of surface exploration disturbance as well as an underground exploration decline and subsequent bulk sample of up to 120,000 tons.

 

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 i-80 Gold Corp  

Accessibility, Climate, Local Resources, Infrastructure,

and Physiography

  Page 41 

 

4.  Accessibility, Climate, Local Resources, Infrastructure, and Physiography

 

  4.1.

Accessibility

Access to the Project area is via State Highway 305, 30 miles south from the town of Battle Mountain, and then west approximately seven miles along the paved McCoy Mine Road. Battle Mountain is located on Highway 80, approximately 70 miles west of Elko, Nevada.

Topographic elevations vary across the project site from 4,600 to 7,200 feet amsl.

 

  4.2.

Climate

The climate in Lander County is typical of the high-desert environment. Average July temperatures range between 65°F and 75°F in the lower valleys and cooler in the higher elevations. Summer highs in the valleys are approximately the mid-90°F, with temperatures in the range of 50°F or 60°F at night. Winter temperatures average between 20°F and 30°F in the valleys with the possibility of frost from early September through June.

Average rainfall is 10 in to 15 in, with less than 10 in. of rain in the lowest areas and up to 20 in. occurring in the mountains. The majority of precipitation falls between November and May, with the possibility of summer thunderstorms.

All of the mining operations in the area operate 365 days per year.

 

  4.3.

Local Resources

The McCoy Mining District has a long history of mining activity, and mining suppliers and contractors are locally available. Both experienced and general labor is readily available from the towns of Elko in Elko County (100 miles north and east of the Project) and Winnemucca in Humboldt County (83 miles north and west of the Project). Some services are also available in Battle Mountain (30 miles north of the Project). There are a number of mining operations in the area and as such there is always competition for employees.

 

  4.4.

Infrastructure

Dirt track access roads are located throughout the property for exploration access. The Project exploration facilities consist of a guard shack, mechanic shop and numerous shipping containers used as storage sheds in the laydown and core storage yards.

 

 

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Nevada Energy (formerly Sierra Pacific) power lines run to the property at the McCoy-Cove Project. Power is available at the site from a 120 kV transmission line and a 24.9 kV distribution line.

Previous mining within the Project has left a legacy of:

 

   

Cove open pit and pit lake;

 

   

Reclaimed leach pads;

 

   

Tailings dam (partially reclaimed);

 

   

Reclaimed dumps; and

 

   

Reclaimed infiltration basins.

All aforementioned facilities except for the tailings dam have been released by state and federal agencies and are considered reclaimed.

 

  4.5.

Physiography

The Project lies in the Basin and Range Province, a structural and physiographic province comprised of generally north to north-northeast trending, fault bounded mountain ranges separated by alluvial filled valleys.

The property is located on the northeastern side of the Fish Creek Mountains. Elevation in the McCoy Mining District ranges from about 4,800 feet to 6,900 feet above sea level. The valley in the Helen deposit area is at approximately the 4,800 feet elevation and the area overlying the deposit has an elevation of approximately 5,500 feet.

Vegetation is typical of the high desert; greasewood characterizes the salt flats, sagebrush dominates the alluvial fans, and piñon and juniper are found on the mountain slopes. Rabbit brush, white sage, and mountain mahogany are also present (Figure 4-1).

 

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and Physiography

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Figure 4-1 McCoy-Cove Project Area Looking Southeast

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5.  History

Gold was first discovered in the McCoy Mining District in 1914 by Joseph H. McCoy. Production through 1977 included approximately 10,000 ounces of gold plus minor amounts of silver, lead, and copper. Production in these early years came from placers and from gold-quartz veins that occurred in northeast striking faults and in intersections of northeast and northwest striking faults. Most of the non-placer production, however, came from argillized and oxidized skarn at what became the McCoy open pit mine.

 

  5.1.

Previous Owners

Summa Corporation (Summa), a Howard Hughes company, acquired most of the mining claims in the McCoy Mining District in the 1950s and 1960s. In 1977, Houston Oil and Minerals Corporation (Houston) purchased the McCoy-Cove Project. Gold Fields Mining Corporation (Gold Fields) leased the property in 1981 until September 1984, whereupon the property was returned to Tenneco Minerals Company (Tenneco), which had acquired Houston. Echo Bay Mines Ltd. (Echo Bay) purchased the precious metal holdings of Tenneco in October 1986. Newmont took ownership of the Cove and McCoy properties in February 2003 following the merger between TVX Gold Inc. (TVX), Echo Bay, and Kinross Gold Corporation (Kinross).

Victoria Gold Corp (Victoria) leased the property in June 2006 as previously described in Section 3. In June 2012, Premier entered into an agreement to acquire the lease of the McCoy-Cove Project from Victoria and subsequently acquired a 100% interest in the land package from Newmont in September 2014.

 

  5.2.

Historic Exploration

Modern exploration for copper and gold in the McCoy Mining District started in the 1960s by Bear Creek Mining Company and Pilot Exploration drilling in 1967. Summa conducted extensive exploration on the McCoy skarn deposit from 1969 to 1977. Summa also undertook regional geologic mapping of 55 square miles (including the McCoy-Cove Project area) and extensive rock chip surveys.

Houston explored the property in 1980, including geologic mapping, soil geochemical surveys, ground magnetic surveys, and drilling.

Gold Fields conducted an extensive induced polarization (IP) program, airborne magnetic surveys, detailed rock chip sampling, as well as limited geologic mapping and drilling between 1981 and 1984.

 

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In 1985, Tenneco undertook drilling, metallurgical testing, engineering and feasibility studies and began mining the McCoy deposit in February 1986. Tenneco also began systematic district-wide exploration in 1985 with the collection of 500 stream sediment samples from an eight-square mile area around the McCoy deposit. Evidence of what would become the Cove deposit was found in early 1986, when seven samples yielded gold values of between 15 ppb and 72 ppb with associated anomalous Ag, As, Hg, Sb, and Ti. Subsequent detailed geologic mapping identified jasperoid, manganiferous limestone, and outcrops of altered felsic dikes in the area of the anomalous samples. Surface rock chip samples of these rocks all contained significant gold mineralization. Tenneco’s detailed mapping covered a large area that included both McCoy and Cove and extended to the north, west, and south. In September and October 1986, a total of 147 soil samples were collected from the B and C soil horizons over the altered area at Cove on a 100-foot by 200-foot grid.

Echo Bay continued the systematic district exploration program initiated by Tenneco that included stream sediment, soil, and rock chip sampling plus geologic mapping, exploration trenching using a bulldozer and drilling. Later soil sampling at Cove defined a gold anomaly measuring 2,800 feet long by 100 feet to 600 feet wide, with gold values ranging from 100 ppb to 2,600 ppb. Bulldozer trenching exposed ore grade rock over the entire length of this soil anomaly. Echo Bay discovered the Cove deposit with drilling in January 1987. By March 1987, Echo Bay had drilled 42 shallow exploration holes and development drilling began in late March. Echo Bay drilled 458 reverse circulation (RC) holes totaling 315,000 feet from January 1987 through June 1988 and 51 core holes totaling approximately 65,800 feet through 1989 (Briggs, 2001).

In 1999, Echo Bay drilled eight surface drill holes totaling 6,700 feet on the Cove South Deep deposit. This drilling, combined with bulk sampling from an underground exploration drift, confirmed the presence of a high-grade zone (0.25 opt Au) that could be mined by underground methods (Briggs, 2001). Detailed underground drilling of this deposit continued during 2000 as mining proceeded.

Newmont drilled 15 vertical holes on the property from 2004 to 2005. Victoria began exploring the property in 2006 resulting in the discovery of the Carlin-style Helen Zone immediately northwest of the Cove pit.

 

  5.3.

Historic Mining

The earliest known significant mining was in the early 1930s at the Gold Dome mine, previously located on northeast side of the present McCoy open pit mine. This operation included a 250-foot shaft and five levels of workings at 50-foot intervals producing gold grades ranging between 0.25 opt and 2.0 opt.

 

 

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Table 6-2 summarizes the annual production between 1986 and 2006 at the McCoy and Cove mines. Tenneco commenced mining at the McCoy open pit mine in 1986 and Echo Bay began open pit mining of the Cove deposit in 1988, accompanied by three phases of underground mining.

Underground access at the Cove Mine was via a decline with rubber-tire machines using a room and pillar mining method. From 1988 to 1993, underground mining was used to recover high grade ore ahead of the pit. In 1999, additional underground mining at Cove South Deep (CSD) recovered approximately 300,000 tons of mineralization beyond the ultimate pit limits. The mineralization was relatively flat lying from 10 feet to 80 feet thick. Longhole stoping and drift and fill methods were used with cemented rock fill (CRF).

Conventional open pit mining methods were utilized at Cove open pit, with drilling and blasting of ore on 20-foot benches (double benched to 40 feet) and waste on 30-foot benches (double benched to 60 feet). The lower sulfide orebody was reached in late 1991.

Processing of low grade, run-of-mine heap leach ores from Cove began in 1992 and mining of high-grade ores was completed in 1995. Open pit mining ended at Cove in October 2000.

In 1996, the mill facility was expanded from 7,500 stpd to 10,000 stpd, with milling of stockpiled ores from the Cove open pit beginning in the second half of 1997. Mill recoveries declined during the remaining life of the mine as lower grade, more refractory ores were processed. By October 2000, the mill was processing 11,369 stpd. As of that date, the gold grade was 0.055 opt Au and plant gold recovery was 51.8%; silver grade was 4.00 opt Ag and plant silver recovery was 71.5%.

The mill contained gravity, flotation, and cyanide leach circuits. Through 2006, a total of 3.41 million ounces of gold and 110.2 million ounces of silver were produced from Cove and McCoy, with the vast majority of both metals reportedly coming from the Cove deposit. Approximately 2.6 million ounces of gold were produced from the Cove open pit.

Table 5-2 Historic Cove and McCoy Mine Production 1986 through 2006

Mineralized Material Processed   Oxide   Sulfide   Heap Leach          
Year  

Milled

Oxide

Tons

(000)

 

Milled

Sulfide

Tons

(000)

 

Heap

Leach

Tons

(000)

 

Au

(opt)

 

Ag

(opt)

 

Au

(opt)

 

Ag

(opt)

 

Au

(opt)

 

Ag

(opt)

 

Au

Ounces

  Ag Ounces
   

1986

  -   -   1,851   -   -   -   -       34,035   na
   

1987

  -   -   4,292   -   -   -   -   0.04   -   90,788   56,800
   

1988

  -   -   2,994   -   -   -   -   0.053   1.14   104,009   764,116
   

1989

  1,358   -   5,696   0.107   3.21       0.02   0.44   214,566   2,259,653
   

1990

  2,004   201   5,709   0.084   0.82   0.227   6.17   0.021   0.2   255,044   1,982,455
   

1991

  2,094   364   5,174   0.077   1.7   0.194   8.42   0.02   0.69   284,327   5,619,007

 

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Mineralized Material Processed   Oxide   Sulfide   Heap Leach          
Year  

Milled

Oxide

Tons

(000)

 

Milled

Sulfide

Tons

(000)

 

Heap

Leach

Tons

(000)

 

Au

(opt)

 

Ag

(opt)

 

Au

(opt)

 

Ag

(opt)

 

Au

(opt)

 

Ag

(opt)

 

Au

Ounces

  Ag Ounces
   

1992

  1,483   990   9,029   0.075   2.54   0.163   7.57   0.014   0.6   301,512   7,921,496
   

1993

  2,308   552   8,938   0.107   4.61   0.136   4.65   0.017   0.88   395,608   12,454,338
   

1994

  506   2,304   7,892   0.126   6.71   0.143   4.91   0.013   0.48   359,360   10,443,151
   

1995

  497   2,151   4,355   0.15   5.42   0.104   5.23   0.018   0.49   310,016   11,905,806
   

1996

  -   3,287   6,068   -   -   0.086   3.14   0.018   0.27   271,731   7,102,348
   

1997

  -   3,391   6,494   -   -   0.061   4.54   0.018   0.29   187,034   11,021,708
   

1998

  -   4,306   4,112   -   -   0.046   2.95   0.021   0.26   167,494   9,412,823
   

1999

  -   4,452   4,178   -   -   0.038   3.02   0.022   0.37   124,536   8,430,072
   

2000

  -   4,172   1,809   -   -   0.053   3.71   0.024   0.93   162,784   12,328,297
   

2001

  -   -   -   -   -           94,633   6,451,425
   

2002

  -   -   -   -   -           33,142   1,987,421
   

2003

  -   -   -   -   -           4,699   706
   

2004

  -   -   -   -   -           8,454   64,335
   

2005

  -   -   -   -   -           2,740   776
   

2006

  -   -   -   -   -           2,939   596
                                             
   

Total

  10,250   26,170   78,591   0.10   2.93   0.08   3.98   0.02   0.48   3,409,451   110,207,329

 

 

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6.  Geologic Setting, Mineralization and Deposit

 

  6.1.

Regional Geology

The McCoy-Cove Project is located in the central Nevada portion of the Basin and Range Province, which underwent regional extension during the Tertiary that created the present pattern of alternating largely fault bounded ranges separated by alluvial filled valleys (Figure 6-1). Prior to this extension, central Nevada had been the site of numerous tectonic events, including at least two periods of regional compression. The property lies west of the central part of the Battle Mountain-Eureka Trend.

During the Paleozoic, central Nevada was the site of the generally north-northeast trending continental margin of North America, along which pre-orogenic rocks of Cambrian to Early Mississippian age were deposited. A carbonate platform sequence was deposited to the east along the continental margin, with siliceous and volcanic rocks deposited to the west. In Late Devonian to Early Mississippian time during the Antler Orogeny, rocks of the western assemblage moved eastward along the Roberts Mountains thrust, perhaps as much as 90 miles over the eastern assemblage carbonate rocks. A post-orogenic assemblage of coarse clastic sedimentary rocks of Mississippian to Permian age was shed eastward from an emerging highland to the west, overlapping the two earlier facies.

During Pennsylvanian and Permian time, chert, pyroclastic rocks, shale, sandstone, conglomerate, and limestone of the Havallah sequence were deposited in a deep eugeosynclinal trough to the west of the Antler orogenic belt. These rocks were thrust eastward along the Golconda thrust over the Antler overlap assemblage in Late Permian and Early Triassic time during the Sonoma Orogeny. The Golconda thrust is exposed to the west of the Roberts Mountains thrust.

Mesozoic rocks, primarily shallow water siliciclastic and carbonate units with minor volcanic and volcaniclastic rocks, are found in this part of Nevada. At least three additional tectonic events are recorded in late Paleozoic and Mesozoic time, including the formation of the late Jurassic Luning-Fencemaker fold and thrust belt in western and central Nevada. The most recent events in the Great Basin are widespread Cenozoic volcanism and extensional faulting. Late Jurassic (168-143 Ma), Cretaceous (128-90 Ma), and Eocene to Oligocene (43-30 Ma) intrusions have been reported from this part of Nevada.

 

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  6.2.

Local Geology

The stratigraphy of the McCoy Mining District is well documented and has been described in detail by Emmons and Eng (1995) and Johnston (2003). Generalized Triassic stratigraphy of the local area is presented in Figure 6-2 and the major lithological units are described below.

HAVALLAH FORMATION

The Permian Havallah Formation is the deepest drilled unit on the property and is composed of reddish-brown to green argillite and chert. Where it hosts veins, the Havallah displays alteration envelopes containing fine-grained quartz-illite/sericite. The total thickness of the Havallah across the property is unknown. Its contact with the overlying Dixie Valley Formation is sometimes demarcated by the presence of an unconformable rhyodacite tuff (assumed to be Koipato Formation), while in other areas of the property, it is simply defined by the change from coarse-grained clastic conglomerates and sedimentary breccias to argillite.

KOIPATO FORMATION

Locally, at the contact between the Dixie Valley Formation and the Havallah, there is a maroon rhyodacite tuff assumed to be part of the Permo-Triassic Koipato sequence described by Silberling and Roberts (1962). The upper and lower contacts of this rhyodacite tuff are unconformities.

DIXIE VALLEY FORMATION

The early Middle Triassic Dixie Valley Formation consists primarily of coarse-grained conglomerates and intercalated dolomitic sandstones, as well as lesser fossiliferous limestone units generally restricted to the upper portion of the formation.

FAVRET FORMATION

The late Middle Triassic Favret Formation, approximately 750 feet thick, consists of an upper fossiliferous limestone unit containing ammonites and pelecypods, a middle unit of finely interbedded silty limestones and limestones (principal Carlin-style ore host), and a basal unit of debris flow fossil hash containing ammonites, pelecypods, and star-shaped crinoids.

AUGUSTA MOUNTAIN FORMATION – HOME STATION MEMBER

The late Middle Triassic Home Station Member is 100 feet to 150 feet thick and was previously described as a thicker unit consisting of massive calcareous and dolomitic limestone with lenses or beds of sandstone and conglomerate (Kuyper et al., 1991). Johnston (2003) however, classified this unit as silty dolostones based on exposures in the Cove open pit which displayed medium to

 

 

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dark grey, very thickly bedded (greater than 3 feet) dolostone consisting of three to 25 volume percent quartz grains (averaging 0.0016 in. diameter) in a recrystallized dolomite matrix. The clastic components of Kuyper et al.’s (1991) Home Station are now classified as Panther Canyon and the lower limestone is now considered the upper part of the Favret Formation. Although the contact between the Home Station Member and the overlying Panther Canyon Member was described as gradational by Kupyer et al. (1991), Johnston (2003) mapped the contact in the Cove open pit as sharp, and Premier and i-80 geologists use a prominent lag gravel deposit (generally less than 15 feet to 20 feet thick) to mark this contact.

 

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Figure 6-1 Regional Geology

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Figure 6-2 Triassic Stratigraphy and Mineralization

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AUGUSTA MOUNTAIN FORMATION – PANTHER CANYON MEMBER

The Panther Canyon Member at Cove is divided into two informal units, the lower Dolostone Sub member and upper Transitional Sub member. The lower Dolostone Sub member unit is generally 140 feet to 165 feet thick and consists of a 20-30 foot thick gray chert pebble conglomerate at the base that transitions to a dark gray dolostone with local thin bedding and then to a well bedded, pink, white, and beige silty dolostone with common stromatolitic textures. Individual beds are typically less than three feet in thickness. The contact with the overlying Transitional Sub member is very gradational over a distance of approximately 10 feet, but is picked in geochemistry where magnesium percentage becomes less than 4%. The upper Transitional Sub member is a 500 feet thick unit which coarsens upward, from a basal primary maroon and dark gray sandstone with common liesegang banding, through middle carbonate cemented silt- and red bed sandstones, to conglomerate near the top. The general transition is not smooth, however, as contrasting lithologies are interspersed throughout the unit at all levels, typically as lensoid bodies. This Transitional Sub member can be further separated into a lower carbonate rich and an upper clastic section as follows:

 

   

Lithologies in the 140-165 feet thick lower carbonate rich section are highly variable. Although the strata are primarily made up of dolostone, lenses, and beds of carbonate cemented siltstone and very fine-grained sandstone, coarser sandstone and conglomerate are abundant. The lower 80 feet of this section consists principally of massive dolostone. Typical strata in the upper 80 feet of this section consist of 0.001 in. to 0.003 in. diameter, subrounded, moderately sorted quartz grains. Individual beds are typically less than 3.3 feet in thickness. The diagenetic cement is calcite, but it has been dissolved and/or replaced by illite-sericite where hydrothermally altered.

   

The 300 feet thick upper clastic section in the Transitional Submember generally consists of fine-grained sandstone to cobble conglomerate. The thickness of bedding is highly variable, but the conglomerate beds are generally thicker (up to 16 feet thick) than the sandstone beds (up to 3.3 feet thick). Crossbedding is common, and conglomeratic strata typically grade upwards from relatively coarse to relatively fine grained sediments. Detrital grains and cobbles consist of chert, quartzite, and quartz. These grains are rounded to subrounded and moderately sorted. Primary porosity, which was originally high, ranges up to 20 volume percent as observed by Johnston (2003). The contact with the overlying Smelser Pass Member is sharp to gradational over several tens of feet.

AUGUSTA MOUNTAIN FORMATION – SMELSER PASS MEMBER

 

 

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The Smelser Pass Member unit is volumetrically the largest at Cove with a maximum thickness of just over 900 feet. The unit is predominantly a microcrystalline limestone with abundant recrystallized bioclasts, however, the upper 500 feet contain very minor thin interlaminated calcareous shale beds. The limestone is thick bedded to massive, with individual beds ranging from three feet to 16 feet in thickness. Macro allochemical remains consist of partial to complete brachiopods, pelecypods, gastropods, crinoids, corals, sponges, and ammonites, in decreasing order of abundance. The lowermost beds contain up to 15 volume percent of 0.0006 in. diameter quartz grains.

The Smelser Pass Member is separated from the overlying Oligocene tuffaceous sediments and Tuff of Cove Mine by an angular unconformity. Kuyper et al. (1991) determined that the upper 575 feet of the Smelser Pass were removed by erosion prior to deposition of the Oligocene units. More than 2,100 feet of the Triassic Cane Spring and Osobb Formations, which overlie the Smelser Pass Member elsewhere in the McCoy Mining District, are also missing at Cove. Much of the Smelser Pass Member has been subjected to supergene oxidation, giving the originally medium grey limestone an orange to brown appearance.

TUFF OF COVE MINE

The tuff of Cove Mine, previously thought to be the 33.8 Ma Caetano Tuff, has a maximum thickness of approximately 1,500 feet in the deepest parts of the paleovalley it filled. It consists of 0.016 in. to 0.276 in. long fragments of plagioclase, biotite, potassium-feldspar, and resorbed quartz phenocrysts in a glassy to devitrified matrix. Phenocrysts comprise 40 volume percent and matrix 60 volume percent of the rock. John et al. (2008) reported a 40Ar/39Ar age of approximately 34.2 Ma on a set of samples including some collected in the northern Fish Creek Mountains.

INTRUSIVE IGNEOUS ROCKS

Abundant dikes and sills are encountered in drilling at Cove, and historic convention at the property has been to classify them as either “felsic” or “mafic.” The majority are “felsic” and can be mapped at surface associated with and occupying the main faults extending from the Eocene Brown stock at McCoy. Though commonly altered, their textural similarities to the unaltered granodioritic feldspar porphyry of the Brown Stock suggest that they were of similar composition. These dikes are light grey to white in colour due to sericitic or argillic alteration. Their porphyritic texture is preserved. They may be observed over drill hole intercepts ranging in length from less than 0.5 to 215 vertical feet and are usually steeply dipping. Less altered samples collected from the Cove open pit retain evidence for secondary biotite replacing hornblende suggesting a weak potassic alteration event that has been overprinted by lower temperature alteration events at depth.

 

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The Gold Dome is the most prominent “felsic” dike at the deposit and is cross-cut by both polymetallic veins and pervasively altered by weak Carlin-style mineralization.

As a result of the intense alteration, many occurrences of rocks of different composition have been incorrectly logged as “felsic.” Multi-element geochemistry was used in 2016 to reclassify all igneous rocks by filtering for high occurrences of Cr, Ni, and V. When the reclassified lithologies were remodelled in 3D it became apparent that the mafic intrusive rocks are present as thin, laterally extensive, stacked sills that terminate down the northeast limb of the Cove anticline. As a result of that exercise, two distinct trends were discovered in the Ni and V concentrations of these mafic dikes and sills. Whole rock geochemistry and subsequent remodelling confirms the presence of three distinct mafic compositions and this detailed discrimination has been continued with more recent drilling. The mafic intrusions are classified as “Type 1” characterized by high V and lower Ni, “Type 2” characterized by low V and higher Ni, and “Type 3” characterized by strongly elevated Ce, La, and Th with elevated Ni and V. Type 1 in drill core is typically dark green in color, contains abundant calcite filling vesicles, and may be magnetic. Though the Type 1 sills have a strong spatial association to Carlin-style mineralization across the deposit, they are rarely mineralized and can be devoid of As, Au, and Ag in direct contact with mineralized limestone. Typically, the Type 1 intrusions show pervasive propylitic alteration with very occasional weak argillic alteration. The Type 2 and 3 intrusions are generally light green to white in color and can be difficult to distinguish from similarly altered “felsic” dikes. They appear to have been hornblende-biotite porphyries prior to alteration and commonly contain magnetite. They also share a spatial association to Carlin-style mineralization but, unlike Type 1 sills, are very commonly mineralized (up to 20 ppm Au, 20 ppm Ag) and strongly argillically altered. Type 2 and 3 intrusions are less prevalent overall than Type 1 sills. Type 2 and 3 intrusions are generally restricted to filling a northwest striking fault corridor that marks the southern boundary of both the Gap and Helen zones.

QUATERNARY ALLUVIUM

Emmons and Eng (1995) divided the Quaternary surficial units in the McCoy Mining District into alluvium, talus, and colluvium. Quaternary sediments exposed in the Cove open pit were not differentiated in this study. These sediments include unconsolidated sand and gravel, and are less than 215 feet thick.

 

  6.3.

Structural Geology

Deposits on the McCoy-Cove Project are related to specific structural features.

MAJOR DEFINING STRUCTURES

 

 

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The major structure and control on fluid movement is the broad northwest-striking, gently southeast-plunging Cove anticline interpreted as a fault propagation fold over a deep northwest striking reverse fault identified in deep drill holes under the Cove pit. While the reverse fault can be identified in the 2201 zone, its presence at the Gap and Helen Zones is uncertain due to limited drilling in areas that would confirm its continuation. A northwest striking vertical dike called the Northwester Dike (classified as “type 2 and 3”) extends from the Bay fault through the Gap and into the Helen. It appears to prohibit the flow of mineralizing fluids to the southwest in areas between the major northeast striking faults. Though there is no discernible separation on the dike, it may be related to a near vertical to steeply southwest dipping fault mapped in the pit by Echo Bay geologists called the Northwester fault.

The other major structures for fluid movement and mineralization are a number of northeast striking normal faults (Cay, Blasthole, Bay, 110, and Gold Dome). The northeast striking faults commonly host altered granodioritic dikes, the largest of which is the Gold Dome. The north-south striking Lighthouse fault also contains altered granodioritic dikes and is believed to have had both pre- and post-mineralization movement.

These faults and structures were defined and confirmed by:

 

   

Surficial and open pit geologic mapping by Echo Bay, Victoria, and Premier;

 

   

Offset observed during detailed cross section work by Premier in 2016; and

 

   

Oriented core measurements by Victoria and Premier, especially in the Helen and Gap.

 

  6.4.

Mineralization Controls

Carlin-style mineralization appears to be controlled by a combination of the axis of the Cove anticline, normal faults that cut the anticline, mafic sills and dikes throughout the property, and contacts between different sedimentary units. Generally, the highest grades are found where the rhythmically bedded unit of the Favret Limestone is cut by mafic dikes and sills along the axis of the anticline, and especially where this area is cut by apparent small-scale, unmapped faults. Lower-grade (0.05 opt to 0.25 opt Au) Carlin-style mineralization in the Helen and Gap zones is typically found along the Favret-Home Station contact and the contact between the Panther Canyon’s upper conglomerate unit and lower dolomite unit.

The northeast striking faults commonly contain quartz-sericite-pyrite and argillic altered granodioritic dikes that carry low to anomalous values of Au and Ag. Carlin-style mineralization in the Favret and other units is typically bounded by these northeast structures with higher grades focused in the axis of the anticline and lower grades with associated pathfinder elements (As, Sb,

 

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Tl, Hg, etc.) typically along the margins of the anticline as well as immediately adjacent to these major structures.

In the 2201 zone, structural controls are poorly defined, however, vein-bearing Au occurrences do trend northwest and may be related to structures formed in the hanging wall of the deep-seated reverse fault or to the near vertical to steeply southwest dipping Northwester fault.

 

  6.5.

Post Mineral Faulting

There is at least one instance of significant post-mineral faulting. The Striper Splay is believed to be a splay off of the Lighthouse fault which is known to have both pre- and post-mineralization movement. It dips steeply northeast and strikes approximately 320° along the northeast limb of the Cove anticline causing significant post-mineral normal displacement before terminating against the Bay/110 fault complex. The overlying volcanics are not significantly faulted, as defined by holes NW-1, NW-2 & 2A, and NW-3. It is likely there is minor post-mineral movement on all northeast and north striking faults as a result of Basin and Range extension beginning the Miocene and continuing through present day.

 

  6.6.

Mineralization

There are four distinct mineralization types known on the property: Carlin-style, polymetallic sheeted veins, carbonate replacement (Manto) and skarn. The Helen, Gap and CSD deposits are Carlin-style deposits while the 2201 zone is comprised of steeply dipping polymetallic sheeted veins.

CARLIN-STYLE (AU-AG)

The gold in Carlin-style deposits is usually sub-micron in size and generally occurs in pyrite and arsenical pyrite. An envelope characterized by decalcification, silicification, and argillization accompanied by anomalous amounts of silver, arsenic, antimony, thallium, and mercury often accompanies mineralization. The Carlin-style mineralization at Cove is relatively rich in silver compared to similar deposits elsewhere in northern Nevada (Johnston, 2003). When Carlin-style mineralization occurs in the silty limestones and packstones of the Favret Formation and Home Station Dolomite, decarbonatization replaces fine-grained calcite and/or dolomite with quartz and forms very fine-grained illite and pyrite. Diagenetic pyrite was probably present in the Helen Zone before Carlin-style mineralization based on the abundant presence of subhedral pyrite grains that bear no arsenian rims. The arsenic-bearing pyrite precipitated as a product of Carlin-style mineralization in the Helen are fine-grained (~10 microns) patchy, anhedral “fuzzy” pyrite generally smaller than the diagenetic pyrite grains. In the CSD zone, most pyrite grains in

 

 

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high-grade samples are larger (~20 microns), display spectacular, sharp geochemical zonations, and are rimmed with arsenian pyrite or stoichiometric arsenopyrite. The few samples studied from the Gap under the SEM suggest it shares more in common with the CSD zone though its silver content is lower overall.

POLYMETALLIC SHEETED VEINS (AU-AG±PB-ZN)

The polymetallic veins in the 2201 zone are enveloped by a zone of illitization of the conglomerate matrix detected by sodium cobaltinitrite staining and confirmed by scanning electron microscope (SEM) analysis. Minor silicification is relatively common, especially in the conglomerate, however, it is not present everywhere and not always directly associated with mineralization.

CARBONATE REPLACEMENT (AG-PB-ZN±AU)

Carbonate replacement mineralization occurs as local pods of manto-style mineralization characterized by massive sulfide (pyrite-sphalerite-galena) replacing basal limestone at the Dixie Valley/Favret contact. Mineralization is discontinuous and generally defined by high-grade Ag-Zn-Pb±Au.

SKARN (AU-AG±CU)

Skarn mineralization at the historic McCoy pit occurs as both endoskarn and exoskarn mineralization characterized by a predominantly garnet-diopside-magnetite mineral assemblage.

The Carlin-style mineralization across the deposit appears to represent an evolving system from a “primary” endmember represented by the CSD zone with higher Ag/Au, coarser-grained pyrite, and a close proximal relationship to Ag-Pb-Zn-(Au) mineralization to the “evolved” endmember represented by the Helen Zone with lower Ag/Au, very fine-grained pyrite, and weak spatial association with any other styles of mineralization. The Gap can be considered a “transition” zone between the two endmembers until more petrography is conducted on the recently discovered Gap to test this hypothesis. Helen Zone geochemistry is distinct from the CSD zone in many ways. For samples greater than 1 ppm Au, less than or equal to 100 ppm Ag, and confirmed to be Carlin-style mineralization by core photo review, the Helen Zone has an average Ag/Au ratio of approximately 0.85 whereas the CSD zone is 2.25. Gold in both the Helen and CSD zones correlates with As, Sb, and Hg, however, Au correlates moderately (0.52 correlation coefficient) with Ag in the CSD zone but more weakly (0.3652 correlation coefficient) in the Helen Zone.

 

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Like the geochemistry, the mineralization in the Helen and CSD is also distinct. The As-bearing (assumed to also be Au-bearing) pyrite in the Helen are generally finer-grained, less euhedral, and more poorly zoned than the As-bearing CSD zone pyrite. Helen pyrite overall have lower As content – ranging from just at detection limit (~0.3 wt% to 0.5 wt%) to 2.1 wt% – than the CSD zone which contains pyrite with arsenic contents ranging from detection limit to 6 wt%. The SEM-EDS system first detected trace elements such as Te, Tl, Hg, Sb, and even Au and Ag in CSD zone pyrite, while electron microprobe analysis confirmed the presence of Au, Ag, As, Tl, Hg, Sb, and Pb in CSD mineralization. Other pyrite in the CSD zone contain fewer trace elements but still display complex elemental zoning and growth patterns visible only in backscatter electron imaging. The complicated nature of the mineralized pyrite at the CSD zone is suggestive of a more complex and long-lasting mineralizing event in comparison to the seemingly simple Helen mineralization.

In the 2201 zone, Au correlates with Ag, As, Cu, Fe, Pb, Sb, and Zn – a distinctly different grouping of elements from the CSD, Gap, and Helen Zones. The 2201 zone veins typically occur as sheeted veins and range in thickness from 0.1 cm to 6.5 cm and contain both quartz and carbonate minerals as gangue. Generally, the calcite and dolomite-dominant veins are shallower and thinner whereas the quartz (-carbonate)-bearing veins are deeper and can reach widths of 15 cm. The sulfides are mostly pyrite, sphalerite, and galena with arsenopyrite, chalcopyrite, and pyrrhotite also locally present. Visible gold is mostly limited to the thicker veins and is always observed along the margins with coarse-grained quartz. When microscopic, the gold is present as electrum with approximately 15 wt% Ag (measured on SEM-EDS) and hosted within sulfides such as chalcopyrite or arsenopyrite. Galena may also carry up to 10 wt% Ag. An oriented hole drilled in 2014 (PG14-23) provided some structural data for the vein-type mineralization. There were no trends for veins grouped by gangue or thickness, however, when grouped by depth, the data show that veins shallower than 1,750 feet generally strike northeast-southwest with varying dips and veins deeper than 1,900 feet generally strike northwest-southeast and dip steeply in both directions.

 

 

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  6.7.

Deposit Types

Mineralization at McCoy-Cove consists of two mineralization styles, Carlin-style and polymetallic sheeted veins, as outlined in Section 6 of the report. The Carlin-style mineralization within the Helen, Gap, and CSD zones comprises approximately 85% of the existing resource with high gold and silver grades occurring as both stratabound and structurally controlled mineralization at the intersection of the Cove anticline and favorable lithologic beds, structures, intrusive dikes and sills.

The polymetallic 2201 zone is a separate deposit from the shallower Carlin-style mineralization and is believed to be a structurally controlled sheeted vein system. Veining is oriented northwest, with vein geometry being controlled by a deeper northwest striking reverse fault. Due to its depth, the 2201 zone has seen limited drilling since its original discovery in late 2013, however, additional infill and step-out drilling in the future will help to better define deposit potential and mineralization controls.

 

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7. Exploration

 

  7.1.

Geologic

McCoy-Cove is a large property with advanced-stage deposits as well as numerous sparsely tested prospective areas. Historical exploration from the 1960’s to 2012 included stream sediment (silt) sampling, soil sampling, rock chip sampling, geophysical surveys, and geologic mapping. Since acquiring the property in 2012 through mid-2018 when the mineral resource estimate was completed, Premier carried out soil sampling, field mapping, geophysics and drilling projects. Highlights of Premier’s exploration through mid-2018 included the discovery of the 2201 and CSD Gap zones as well as the re-interpretation of the litho-structural model, which resulted in expansion and improved continuity throughout the Cove-Helen zone. The updated litho-structural model has helped guide property-wide target generation.

Numerous exploration targets have been identified within the McCoy-Cove land package. All targets are thought to be Carlin-style and/or polymetallic 2201-style mineralization. Since mid-2018, exploration efforts have focused on eight areas: Windy Point, Antenna, Alpha, Davenport, Lakeside, Saddle, Reflection, and Hidden Valley (Figure 7-1).

 

 

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Figure 7-1 Exploration Targets

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The Windy Point area contains significant surface alteration that includes decarbonatization and silicification. Surface samples have returned >1.2 opt Au, with silver and base metals largely absent. The Cove anticline continues through Windy Point. Here, the anticline has an east-west orientation rather than northwest-southeast as in the Cove pit. Granodiorite dikes and mafic sills are present both at the surface and in drill core. Several drill holes have intersected >300 ft intervals of <0.06 opt Au with isolated intervals of >0.2opt Au.

The area between the McCoy and Cove pits is referred to as the Antenna target. Eocene granodiorite dikes extend from the Brown stock northward through Antenna and into the Cove pit. Geochemically anomalous surface rock-chip samples of sedimentary rocks generally coincide spatially with granodiorite dikes at Antenna. Drilling has intersected multiple >0.15 opt Au intervals displaying both Carlin-style and polymetallic mineralization.

The Alpha target is located 1 km south of Windy Point. Surface rock-chip samples have returned >3 opt Au from jasperoid. Intrusive sills outcrop at surface with jasperoid occurring in proximity to the sills. In 2019, one deep hole was drilled into Alpha, intersecting weakly anomalous Carlin-style geochemistry and alteration.

Davenport and Lakeside are both pediment targets that lack substantial drilling. Most drilling in these areas consist of shallow condemnation holes drilled by Echo Bay. A large magnetic high is present at Davenport and was drill tested in 2020. Results indicate the magnetic high is a large granodiorite sill. Mesozoic sedimentary rocks are present beneath the granodiorite. Two deep holes drilled at Davenport in 2020 intersected long intervals of geochemically anomalous Carlin-style altered rock. In addition, one hole intersected polymetallic 2201-style mineralization in the Dixie Valley formation. The pediment remains a large, underexplored area on the property capable of containing a large Carlin-style ore body.

The Saddle target is located 0.3 mi. south of the western margin of Windy Point. This target lies along the north striking Saddle fault at the intersection with northwest striking faults. A historic drill hole intersected 60 feet of 0.2 opt Au in the Panther Canyon formation. The drill hole was ended before reaching the favorable Home Station dolomite and Favret limestone units. Future exploration work should include twinning the historic hole and drilling deeper to intersect the underlying favorable stratigraphic units.

The Reflection target is located southwest of the McCoy pit. At Reflection, northeast striking granodiorite dikes extend from the Brown stock where they intersect the margin of the Jurassic McCoy Pluton, a north striking anticline, and the large north striking Saddle fault. In addition, a string of jasperoid outcrops extend northwest along the McCoy Pluton margin in the target area. A single hole tested favorable stratigraphic horizons in the target area in 2022 without any significant intercepts of mineralization or alteration. Metasomatic iron skarns are present along the

 

 

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margins of the McCoy Pluton and have been mined historically. Iron-bearing minerals may be present in the sedimentary rocks near the margin of the McCoy Pluton representing a compelling target in non-traditional host rocks given sulfidation is the primary mode of gold deposition for Carlin-style fluids.

The Hidden Valley target is located north of the Helen and Gap zones where a northwest striking anticline is mapped at surface. The anticline appears to be a much tighter fold than the Cove anticline and may be a parasitic fold to the Cove anticline. Drilling at Hidden Valley should target the axis of the anticline in the Favret limestone near the intersection with northeast striking granodiorite dikes.

In January 2018, Premier and Barrick entered into the Barrick Earn-In Agreement which included a significant exploration budget commitment from Barrick to be spent on the McCoy-Cove property. Exploration on the Joint Venture Property began in mid-2018 and continued until Barrick exercised its right to terminate the agreement on February 6, 2020. Work completed by Barrick included detailed surface mapping, soil sampling, gravity survey, and drilling. Barrick drilled 30 holes and Premier has drilled 16 holes since mid-2018. None of the new holes intersect the modelled resource area. Eight of Premier’s holes were drilled for piezometer installations. (Figure 7-2)

Figure 7-2 Barrick and Premier Exploration Drilling

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Underground mine development commenced in 2022 and roughly 4,900’ of development was completed down to the 4640 elevation to provide platforms for infill drilling. A roughly 140,000 foot resource conversion drill program was ongoing at the time of this report. Further mine development will provide platforms for exploration drilling, including access to the difficult-to-reach prospective Gap Extension target under the Cove pit (Figure 7-3).

Figure 7-3 Underground Exploration Potential

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  7.1.1.

Resource Conversion Drilling

All drilling since January 2023 has been conducted from recently mined underground drill platforms for resource conversion of the Helen and Gap zones. The drill project is ongoing and is not included in the resource. The resource will be updated to include the resource conversion holes when the drill project is complete, projected for Q1 2025. The project includes about 125 drillholes totaling 140,000 feet of drilling.

 

 

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Figure 7-4 Resource Conversion Drill Program

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  7.1.2.

Drilling

The McCoy-Cove drill hole database is large, containing many holes drilled across the large land package. For the current resource estimate, the drill data was filtered to contain only holes within and near the Helen, CSD, CSD-Gap, Gap Hybrid and 2201 Zones. A total of 1,397 holes totaling 1,127,481 feet of drilling were included in the current estimate. Holes were drilled using both core and reverse circulation (RC) methods. Premier drilled 123 of the holes, and the remainder were drilled by Victoria, Newmont and Echo Bay. Figure 7-5 shows a plan view of the drill holes, and Table 7-1 lists the type and extent of drilling completed by each operator.

 

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Figure 7-5 Plan View of Drill Holes Used for the Current Analysis

 

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Table 7-1 List of Drilling by Operator

         
Year    Drill Hole Type    Operator    Number
of Holes
  

 Length 

Drilled

(ft)

   
1985-2000    Reverse Circulation    Echo Bay    788    520,194
   
1999-2000    Cubex (RC)    Echo Bay    201    22,829
   
1987-2000    Diamond Drill    Echo Bay    251    216,059
   
2004-2005    Reverse Circulation    Newmont    13    22,080
   
2006-2009    Diamond Drill    Victoria    21    47,118

 

 

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Year    Drill Hole Type    Operator    Number
of Holes
  

 Length 

Drilled

(ft)

   
2013-2017    Reverse Circulation    Premier    8    14,340
   
2012-2018    Diamond Drill    Premier    115    284,862
   
Total              1,397    1,127,481

Figure 7-6 through Figure 7-9 show 100-foot thick sample sections of drilling in the CSD-Gap, Helen, CSD and 2201 zones. Holes drilled by Premier are labeled and shown with thicker traces. Models of lithologic surfaces and 3-gram grade polygons are shown for reference.

Figure 7-6 Sample Section of CSD-Gap and Gap Hybrid Drilling

 

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Figure 7-7 Sample Section of Helen Zone Drilling

 

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Figure 7-8 Sample Section of CSD Zone Drilling

 

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Figure 7-9 Sample Section of 2201 Zone Drilling

 

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Recent drill projects have predominantly been completed by coring, while RC drilling was used extensively to delineate historic pit and underground resources. Accordingly, the recently discovered Helen, 2201 and CSD-Gap zones were modeled almost exclusively using core holes, while the pit-proximal CSD Zone and low-grade lenses were modeled using a mix of RC and core. Table 7-2 details the proportion of core drilling used to model each zone. The authors have carefully reviewed the data and consider both core and RC data to be reliable.

Table 7-2 Type of Drilling by Zone

             
Zone       

 Mineral Lens 

Codes

  

 Number 

of Holes

    

RC

 Composites 

    

Core

 Composites 

    

Total

 Composites 

     % Core  
   

CSD Gap

   220X      27        0        327        327        100  
   

GAP Hybrid

   500X      27        1        132        133        99  
   

CSD

   110X      269        1,276        699        1,975        35  
   

Helen

   310X, 320X,
330X, 340X
     65        23        871        894        97  
   

2201

   130X, 140X      8        0        53        53         100  

 

 

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Zone       

 Mineral Lens 

Codes

  

 Number 

of Holes

    

RC

 Composites 

    

Core

 Composites 

    

Total

 Composites 

     % Core  
   

Low Grade CSD, Gap and Gap Hybrid

   22000      157        3,897        6,153        10,050        61  

 

  7.1.3.

Historic Drilling Methodology

Evan et al., (2011) described drilling protocols for Victoria:

“Victoria diamond drill holes NW-01 to NW-09, inclusive, were spotted by hand-held GPS. This included collar, foresight and backsight. Drill holes NW-10 to NW-15, inclusive, were surveyed by All Points North, registered Nevada Land Surveyors. A Brunton compass was used to set the drill head.

“All diamond drill holes were proposed and collared based on the property grid, which was referenced in a historical digital terrain map (DTM) created prior to full scale mining and reclamation.” (page 74)

“All Victoria diamond drilling was completed from surface retrieving whole core. The holes were collared HQ size and reduced down to NQ size dependent upon ground or drilling conditions. Drill muds were utilized to ensure consistent core recovery.” (pg. 71)

“Victoria downhole surveys were completed using a North Seeking Gyro (NSG) by Major Technical Services and International Directional Services. NSGs eliminated the need for sighting on surface (gyro-compass alignment) and offered high accuracy. Generally, NSG surveys were performed twice, once at mid-hole and again at hole completion. Readings of dip and azimuth were taken at nominal 50 ft intervals.

“RPA notes that no directional tests were taken during regular drilling operations. Holes NW-02 and NW-09A were unable to be downhole surveyed as the holes had to be abandoned due to poor ground conditions.” (Page 74)

Formal records of Newmont and Echo Bay drill procedures have not been located, but methods are assumed to have followed industry standard.

 

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  7.1.4.

Current Drilling Methodology

 

  7.1.4.1.

Drill Hole Placement

Initial surface collar locations are based on drill plan targeting – collar locations are marked in the field by a geologist using a handheld global positioning system (GPS) device loaded with coordinates from drill plans in either Gemcom or MapInfo project files. A wooden collar picket is marked with both the azimuth and dip designations. The azimuth is also painted in a line on the ground directly in-line with the collar picket allowing the drill rig to line up on the correct bearing from the collar location. The geologist re-confirms both azimuth and dip once the rig is lined up on the drill pad using a Brunton compass. After drilling is complete, holes are abandoned and marked with a metal tag cemented into the collar. A final collar location survey is performed by a professional contract surveyor. The project uses UTM NAD83 Zone 11N international feet coordinate system.

 

  7.1.4.2.

Downhole Surveys

International Directional Services (IDS) of Elko performs downhole surveys on all drill holes. Holes are surveyed on 50-foot intervals using a north-seeking gyroscopic downhole survey tool.

 

  7.1.4.3.

RC Drilling Procedures

Holes are drilled using industry standard RC drilling equipment. Samples are collected on five-foot intervals using a cyclone sample collector. The sample interval is written on the sample bag using permanent marker. Drilling advances are paused at the end of each sample run to ensure the complete sample has been collected and avoid contamination of the following sample. The optimum sample size collected is approximately one quarter to one half of a 17-inch by 22-inch sample bag (about 4.5 to 9 kg or 10-20 pounds.)

 

  7.1.4.4.

Core Drilling Procedures

Core holes are drilled using HQ (about 3-inch diameter) core. Holes may be reduced to NQ (about 2.4-inch diameter) to permit continuation of a hole in difficult drill conditions. Premier has used both standard and triple-tube tooling. Triple-tube is preferable in broken ground because it facilitates placement of core into the core box, allowing the sample to remain more intact. Drilled material is placed in wax-impregnated core boxes. Drillers label the end of the core run to the nearest half of a foot, and measure and record the recovery in feet on wooden blocks, which are placed in the core box at the end of each drilled interval. Core boxes are labeled with company

 

 

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name, property, bore hole identifying number (BHID), box number and drilled interval. The authors believe the drilling procedures are adequate.

 

  7.1.5.

Sampling Methodology

Boxed core is transported by i-80 personnel from Cove to the Lone Tree core storage yard adjacent to the logging facility. When a geologist is ready to log the hole, a geotechnician moves the core inside the logging facility, places it on logging tables and washes off any drill mud. A geologist records detailed geology data directly into acQuire software. Data fields include geotechnical, sampling intervals, lithology, alteration, oxide, sulfide, structure, point data, veins and density.

Sample intervals are chosen by the geologist based on detailed geologic observations. Sample intervals may range from ten feet to a minimum of one foot, with a maximum of five feet in areas of interest. The geologist marks sample intervals on the core and staples a sample ticket double-stub in the core box at the end of the sample interval. Sample IDs are automatically generated in acQuire, with a prefix that designates the project. Sample tickets are then printed out with sample IDs. Logged core boxes are photographed with a high-resolution camera while wet and then stacked on a wooden pallet prior to being moved to the sample cutting area.

The geologist prints a cut-sheet from acQuire software with the sample numbers and intervals and gives the cut-sheet to the geotechnician. The geotechnician places a sample bag in a five-gallon plastic bucket on the floor next to the core saw. The core is sawed in half, the left piece is placed into the sample bag, and the right piece goes back into the core box. In the case of broken core, the sampler does their best to divide the sample equally. Once the interval is split, the geotechnician takes one part of the double sample stub from the core box and staples it to the sample bag. The remaining sample stub remains in the core box for future reference. The geotechnician then ties the sample bag shut and marks the sample off the cut-sheet. The tied sample bags are stored in a sample bin for the lab driver to pick up.

The geologist assigns QAQC samples while logging targeting 5% blanks, 5% standards, and 2.5% field duplicates. The geologist attempts to place blanks after high-grade samples where available. The geologist also attempts to place standards proximal to mineralized zones with standard gold values approximately that of the mineralized zone gold values. However, since the gold value of the rock cannot be known prior to assay, the standard value may not always compare well to the mineralized zone. The geotechnician places the blanks and duplicates with their sample tags in the sample bin with the regular core samples. The standards are placed in a small sample bag with the corresponding sample ID. The standards corresponding to a single hole are then placed in a larger bag prior to shipment to the assay lab.

 

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The geologist completes a sample submittal sheet. The assay lab driver picks up the samples from the Lone Tree core shed and is given a chain of custody form with sample ID’s for the shipment. An electronic copy of the sample submittal form is emailed to the assay lab.

Drill hole status, such as splitting, sample dispatch date, batch ID, and dates of both preliminary and final results, are tracked in acQuire as well as on ALS Mineral’s online portal.

The authors believe the sampling procedures are adequate.

 

  7.1.6.

Core Recovery

Historic core recovery was described by (Evan et al., 2011):

“Overall core recovery for Victoria’s diamond drilling at the Cove Project is estimated at 90%.

“In RPA’s opinion, these values are likely to be overestimated based on the broken nature of the samples retrieved.” (Page 86)

The average recovery for core drilled by Premier is about 90%, which is consistent with historic recovery measurements. Recovery is calculated by measuring the length of material between blocks in the core box and dividing that length by the drilled interval length. It is difficult to measure length accurately for a broken interval of core, and the tendency is to over-estimate recovery in broken intervals. This is a typical problem for drilling in Northern Nevada, and the authors believe that 90% is a reasonable estimate of recovery. Although any sample with less than 100% recovery is sub-optimal, the authors believe the samples provide a reasonable representation of the rock package.

7.2. Hydrogeology

 

  7.2.1.

Sampling Methods and Laboratory Determinations

Hydrogeological data, including groundwater elevation measurements were collected in conjunction with exploration in pre-construction studies and later from hydrogeological studies in the pit and planned underground mining areas. Groundwater elevations were used to develop a piezometric surface and determine the direction of groundwater flow (hydraulic gradient),

A network of vibrating wire piezometers (VWPs) was the primary method of determining water levels to support of mine site characterization, see Figure 7 1 for collar locations of the VWPs. Some water levels were collected from wells. Most wells that were drilled underwent hydrologic testing to estimate aquifer parameters. These tests included short-term and long-term pumping tests. Data obtained from well testing were analyzed using industry standard analytical methods.

 

 

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Analytical and numerical groundwater flow models were developed based on 3D geological modeling and supported by the site-specific aquifer test analysis results.

The McCoy-Cove mine operated from 1985 to 2001. Dewatering occurred from 1988 until mid-2001 and utilized up to 24 dewatering wells, 13 sumps, and numerous horizontal drain holes. Dewatering water, apart from water that was consumed by the mining and milling processes, was placed in a series of rapid infiltration basins (RIB), located approximately one mile north of the pit, where the water infiltrated into the alluvium. In 2001, after mining ceased in the Cove pit and Cove underground, the project went into closure. Groundwater levels have been recovering since that time with the Cove pit forming a pit lake. All dewatering wells from the 1985 to 2001 program have been abandoned in accordance with Nevada Division of Water Resources (NDWR) regulations. Many of the historical monitoring locations from this era have also been abandoned in accordance with NDWR regulations. Existing alluvial monitoring wells, located in the vicinity of the McCoy-Cove mine, consist of heap leach pad monitoring wells, tailings dam monitoring wells, and RIB monitoring wells. Additionally, irrigation wells are monitored as part of the site’s Nevada Division of Environmental Protection Bureau of Mining Regulation and Reclamation (NDEP BMRR) Water Pollution Control Permit (WPCP).

According to permitting requirements, seven monitoring wells are sampled on a routine basis and analyses run for the State of Nevada Profile I suite at a certified analytical laboratory. Monitoring wells and exploration drill holes that have piezometers installed (VWPs) are monitored for piezometric heads. Surface water related to historical mine features is also monitored on a routine basis as required by various permits. Along with collar locations of the VWPs, the monitor well and test well locations are shown in Figure 7-10.

 

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Figure 7-10. Vibrating Wire Piezometer and Groundwater Monitor Well Locations

 

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  7.2.2.

Hydrogeology Investigations

Throughout the span of various mine property owners and operators, the Project area has been the subject of multiple studies aimed at characterizing the hydrogeologic properties of the stratigraphy within the Project area and the surrounding region. Hydrologic Consultants, Inc. (HCI) conducted various hydrogeologic conceptualization studies and groundwater flow modeling studies in the 1990s and early 2000s. In the 2010s, Itasca performed an additional groundwater flow model update. The studies by HCI focused primarily on dewatering of the Cove Pit while the studies in the 2010s and 2020s focused on the hydrogeologic conceptualization of the Gap and Helen deposits and dewatering of these deposits along with dewatering of the Cove pit lake (Table 7-3).

Table 7-3 Summary of Hydrogeological Studies

     
Year    Title    Author
     

1990

   Conceptual hydrogeologic model of Cove pit area as of June, 1990    Hydrologic Consultants, Inc.
     

1992

   Updated conceptual hydrogeologic model and estimates of future dewatering costs as of September 1992    Hydrologic Consultants, Inc.
     

1993

   Status report Cove pit dewatering program    Hydrologic Consultants, Inc.
     

1994

   Updated conceptual hydrogeologic model and estimate of future dewatering costs for Cove pit    Hydrologic Consultants, Inc.
     

1995

   Updated conceptual hydrogeologic model and estimate of future dewatering costs for Cove pit as of April 1995    Hydrologic Consultants, Inc.
     

1997

   Hydrogeologic framework and numerical ground-water flow modeling of McCoy/Cove Mine, Lander county, Nevada    Hydrologic Consultants, Inc.
     

1997

   Updated conceptual hydrogeologic model and predicted dewatering requirements for remaining life of McCoy/Cove mine    Hydrologic Consultants, Inc.
     

1999

   1999 update of hydrogeologic framework and numerical ground-water flow modeling of McCoy/Cove Mine, Lander County, Nevada    Hydrologic Consultants, Inc.
     

2001

   2001 update of numerical ground-water flow modeling for McCoy/Cove mine, Lander county, Nevada    Hydrologic Consultants, Inc.
     

2016

   Numerical groundwater model and predictions of Cove pit-lake infilling , McCoy Cove Mine    Itasca

 

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2018

   Cove Helen hydrogeologic characterization, numerical model update and preliminary dewatering plan    Piteau
     

2021

   Groundwater flow modeling conducted for simulation of Cove Pit-Lake recover, Lander County, Nevada    Montgomery & Associates
     

2023

   Flow modeling for the McCoy-Cove Project    Montgomery & Associates
     

2024

   Flow modeling for the Cove Gap/Helen Project    Montgomery & Associates

In 2017 Au-Reka contracted with Montgomery & Associates (M&A) to provide hydrogeologic services. M&A installed VWPs and test wells, performed pumping tests, performed surface water monitoring, and developed and updated the numerical groundwater flow model of the project area.

 

  7.2.3.

Hydrogeologic Description

The Hydrogeologic Study Area (HSA) lies within the Basin and Range province which stretches from the Sierra Nevada to the Wasatch Front. The HSA is shown in (Figure 7-11). The Project is located on the northeastern flank of the Fish Creek Mountains within the Lower Reese River hydrographic basin. The HSA comprises most of this basin along with portions of Middle Reese River, Antelope Valley, Jersey Valley, and Buffalo Valley hydrographic basins.

Areal recharge and groundwater underflow are the sources of groundwater in the HSA. Areal recharge occurs from infiltration of precipitation in the HSA either as direct recharge or as infiltration of runoff along streams. Underflow enters the HSA through Buffalo Valley, Middle Reese River Valley, and Antelope Valley. Groundwater leaves the HSA as underflow through Lower Reese River Valley and Jersey Valley, groundwater evapotranspiration through naturally occurring processes or from agriculture, and from mine use.

 

  7.2.3.1.

Surface Water

Surface water occurs within the HSA as intermittent, ephemeral, and perennial streams, and as seeps and springs – see Figure 7 3. Most drainages and streams, including the Reese River, are intermittent or ephemeral and only flow during particularly wet years or large storm events. The Reese River, within the HSA is characterized as a series of muddy pools and only flows during wet seasons. Beginning in 2022, M&A began monitoring seeps and springs (Montgomery and Associates 2023). In 2023, additional seeps and springs were added to the group being monitored along with flows and wet/dry reaches of select drainages (Montgomery and Associates 2024a).

 

 

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M&A has continued to conduct these surface water surveys each year, with one sampling event during high spring runoff and a second in late fall during low flow (M&A 2023a, 2024a, and 2025).

 

  7.2.3.2.

Groundwater

Geologically, the HSA is comprised of three general geologic assemblages: Paleozoic, Mesozoic, and Cenozoic. The Paleozoic rock assemblage accounts for the basement rock within the HSA. The Mesozoic assemblage overlays the Paleozoic and hosts the Project. The Cenozoic assemblage contains the many volcanic units within the HSA as well as the Quaternary alluvial sediments that comprise the basin fill.

Groundwater is generally believed to follow topography meaning that groundwater within the mountains and highlands flow down the topographic gradient toward the basin fill sediments. Groundwater flows within the basin fill sediments along or parallel to the valley axis towards discharge points. Groundwater discharge is to evapotranspiration zones such as playas, or out of the HSA by underflow. Historical mining and agricultural pumping have affected groundwater flow locally within the HSA.

Groundwater underflow in the alluvial sediments occurs along the HSA boundary in Buffalo Valley, Antelope Valley, and in the sediments beneath the Reese River at the southeastern corner of the HSA.

Historically groundwater pumpage has occurred due to agriculture in Buffalo Valley, Antelope Valley, and Lower and Middle Reese River Valleys. Due primarily to agricultural pumping, water levels are generally declining. Groundwater pumpage for mining has occurred during the historic McCoy/Cove mining and currently at the NGM Phoenix project located on the southern flanks of Battle Mountain at the northern edge of the HSA. At the McCoy/Cove mine, excess dewatering water was discharged to, and infiltrated into the basin fill sediments in Lower Reese River Valley.

Locally at the project, groundwater generally flows from the west or the upland areas towards the east and out into Lower Reese River Valley. This general flow is observed with the exception of the area around the Cove Pit lake which acts as a local evaporative sink.

 

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Figure 7-11 Hydrogeologic Study Area/Model Area

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  7.2.4.

Mine Dewatering

Beginning in 1988 wells were installed to dewater the Cove pit. As mining progressed additional wells were installed to augment the dewatering of the pit. Dewatering was accomplished through 24 dewatering wells, 13 sumps, and numerous horizontal drain holes.

Starting in mid-1989, dewatering wells pumped between 2,000 and 3,000 gpm in addition to the water pumped from underground sumps. As mining continued, additional dewatering wells were installed. By 1992, dewatering operations were producing over 10,000 gpm, reaching a maximum dewatering rate of nearly 23,000 gpm in late 1994, primarily from dewatering wells. The maximum annual dewatering rate in 1993 and 1994 was just over 19,000 gpm. Horizontal drains ranging from 300 to 600 feet long were installed to depressurize portions of the open pit highwall, primarily east of the Lighthouse Fault where water levels were higher. Average annual dewatering rates from 1988 through 2000 are shown in Figure 7-12.

Underground mining resumed beneath the pit in 1999 and continued until July 2001 with the bulk of the mining focused on the southeast corner below the Cove Pit. The final Cove Pit floor elevation of 3,895 feet was reached in October 2000 and dewatering achieved 675 feet of drawdown. After underground mining ceased in July 2001, the pit was partially backfilled with waste rock along the south and southwest walls and dewatering activities were scaled back. Several of the dewatering wells were kept in operation for an additional 6 months and their production was pumped into the pit lake to quickly submerge sulfide-bearing waste rock that was placed near the bottom of the pit. The pit lake has reached an elevation of 4,628.1 ft amsl as of June 2023.

 

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Figure 7-12 Historical Cove Dewatering Rates

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  7.2.5.

Dewatering Discharge

Historic dewatering operations at McCoy-Cove mine occurred between 1988 and 2001, see Figure 7 12 showing the annual dewatering rates. Most pumped groundwater was discharged into RIBs located northeast of the Cove Pit. Annual average infiltration rates at the historic RIBs ranged between approximately 3,400 gpm and 17,500 gpm. The yearly average RIB infiltration rates are shown on Figure 7-13.

 

 

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Figure 7-13 Historical Discharges to RIBs

 

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  7.2.6.

Groundwater Flow Model

The first groundwater flow model at Cove was prepared by HCI, in the early 1990s with continued updates through 2001. Modeling was restarted in the mid-2010s by Itasca and Piteau to estimate dewatering needs for the Gap and Helen deposits. In 2021 M&A updated the groundwater flow model (Montgomery and Associates 2021). Over the past five years M&A has updated and recalibrated the groundwater flow model two additional times; with the most recent model prepared for NEPA baseline characterization (Montgomery and Associates 2023b, 2024b).

 

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  7.2.6.1.

Model Overview

The numerical groundwater flow model code used for this study is MODFLOW-USG (Panday, et al., 2017). This model code is an unstructured grid version of the 3D finite-difference code MODFLOW (McDonald and Harbaugh, 1988). MODFLOW-USG allows for horizontal and vertical grid refinement in areas of interest while allowing for coarser gridding on the regional scale.

As shown on Figure 7-11 and Figure 7-14, the numerical groundwater flow model area is approximately 1,200 square miles. The model grid’s horizontal discretization uses quadtree refinement which subdivides each “parent” cell into 4 “children” cells. The model grid has 5 levels of quadtree refinement. The largest parent cell has an equal x-y length of 3,200 feet which is sequentially divided into x-y lengths of 1,600 feet, 800 feet, 400 feet, and 200 feet. The 3,200 feet by 3,200 feet grid cells are used to simulate the regional areas outside of the McCoy-Cove and Phoenix mine sites. The level of refinement increases as the model cells approach the mine sites. The smallest refinement cell, 200 feet by 200 feet, is located within the immediate area of the mine sites.

The top of the model is defined by the land surface. The base of the model is set uniformly to be at zero ft amsl to reduce the potential interactions between the model bottom boundary and mine dewatering stresses. The geologic units within the Model Area are grouped into hydrogeologic units (HGU). The model’s vertical discretization is based on the geologic model of the hydrostratigraphy. Each HGU is assigned a single or multiple layers depending on its total thickness.

The historical simulation begins on December 31, 1978, and ends on January 1, 2024. The first stress period lasts one day and simulates steady-state conditions representative of pre-development conditions. Stress periods two through 10 last one year and simulate transient conditions through 1987. Stress periods 11 and 12 simulate 1988 with stress period 12 representing the beginning of the McCoy-Cove mining. Annual stress periods are resumed until stress period 25, which ends on July 31, 2001, when McCoy-Cove mining operations stopped, and the Cove Pit Lake filling began. Stress period 26 simulates the time it takes from August 1, 2001, through January 31, 2002, to fill the pit lake by pumping from several McCoy-Cove dewatering wells. The pit lake was intentionally quick filled with water to stabilize the geochemistry. Annual stress periods continue through 2018. 2019 is divided into many stress periods to simulate the aquifer testing conducted in 2019. Annual stress periods then resume for 2020 through 2022. 2023 is also divided into many stress periods to simulate the aquifer testing conducted in 2023.

 

 

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Figure 7-14 Model Grid Extent

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  7.2.6.2.

Model Calibration

Numerical model calibration is the process used to simulate hydrogeologic conditions measured in the past and adjust model parameters so that simulated historical conditions can be approximated. This process is often called history matching because the numerical model is used to match historical conditions. The hydrogeological conditions used for history matching include historical groundwater levels and changes in groundwater levels, surface water flows, and groundwater flows. Calibration is performed on historical water levels, estimated groundwater budgets, and the aquifer tests performed in 2019 and 2023 using pumping from WE01 and WE02 test wells.

From the late 1980s through the middle of 2001, McCoy-Cove dewatering operations lowered water levels in the vicinity of the Cove Pit. After dewatering ceased in 2001, water levels began to recover, and the Cove Pit Lake formed. In the pit, and on the northern and eastern side of the Cove Pit, the simulated and observed water levels closely match. This match between observed and simulated water levels indicates the model bulk hydraulic properties in the pit and to the north and east are representative of actual hydraulic properties. Hydrographs from the western and southern sides of the Cove Pit tend to under-simulate the observed drawdown (Figure 7-15). The underprediction of the simulated drawdowns likely means that the simulated hydraulic conductivities to the west and south of the pit are slightly higher than actual. The simulated and observed Cove Pit Lake filling curves are tightly correlated (Figure 7-16). The correlation between simulated and observed pit lake stages is an indicator that the bulk hydraulic properties surrounding the Cove Pit are reasonably represented within the model.

The steady state residual mean is 19 feet, which means that on average the simulated steady state, pre-development water levels are 19 feet lower than observed. The transient residual mean is -27 feet, indicating the transient simulated water levels are on average 27 feet higher than observed. The absolute residual mean is a measurement of the average magnitude of the residual, with 38 feet and 63 feet the absolute residual mean for the steady state and transient periods, respectively. When evaluating model calibration, a scaled statistic of less than 10% is evidence of a reasonable model calibration and is calculated by dividing the statistic by the observation range. All scaled statistics calculated for the model are below 10% and indicate an adequately calibrated model.

 

 

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Figure 7-15 Representative Simulated and Observed Hydrographs around the Cove Pit Lake

 

 

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Figure 7-16 Simulated and Observed Cove Pit Lake Stage

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Table 7-4 Calculated Model Calibration Statistics

 

       
Group   

 Steady 

State

   Transient     Total 
   

Number of Locations

   62   144   252
   

Number of Observations

   63   3,429   3,538
   

Observation Range

   2,661   1,011   3,344
   

Residual mean (feet)

   19   -27   -25
   

Absolute residual mean (feet)

   38   63   63
   

Root mean square error (feet)

   69   94   95
   

Scaled root mean square error (feet)

   2.6%   9.3%   2.8%
   

Scaled absolute mean residual (feet)

   1.4%   6.2%   1.9%

 

  7.2.6.3.

Predictive Groundwater Flow Modeling of Dewatering for Gap and Helen Mining

The proposed Project will mine the Gap and Helen deposits beneath the bottom of the westernmost portion of the Cove Pit and extend to the west-northwest of the Cove Pit. The underground mine begins with the decline at the surface along the northern side of the Cove Pit and then extends

 

 

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down to approximately 3,400 ft amsl. With the decline primarily constructed down to the approximate water table of 4,630 ft amsl, mining down to the 3,400 ft level would occur over 9 years from 2029 through the end of 2037 (Figure 7-17).

Figure 7-17 Projected Mine Elevation for Dewatering During Life of Mine

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Dewatering will be required because the ore body is hosted entirely below the current water table. The hydrogeological impacts from the proposed Project are largely related to the required dewatering of approximately 1,000 feet during the first couple of years of mining ending in

 

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2030. To simulate Helen and Gap mining, the groundwater model is temporally extended through the end of mining, 2037, and then for an additional 500 years to simulate post-mining hydrogeologic recovery.

Predictive modeling is undertaken to simulate the dewatering efforts needed for the Project in addition to its related simulating hydrogeological impacts. Figure 7-18 shows the simulated water levels of the Cove Pit Lake and of 2 representative locations within the Project underground workings: one in the Gap deposit and a second in the Helen Deposit. As evidenced by the dashed lines, simulated water levels are just above or below the workings during the life of mine. Figure 7 10 also shows that the simulated stage from the pit lake drops below the bottom of the pit lake by the end of 2032, resulting in the pit lake being completely dewatered.

Figure 7-18 Simulated Water Levels During the Life of Mine

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Dewatering will occur through direct pumping from the Cove Pit Lake and from dewatering wells. Dewatering rates through 2037 are shown on Figure 7-19. The water to be managed peaks in 2027 at approximately 55,000 gallons per minute (gpm) and declines to approximately 24,000 gpm at the end of 2037. Pit lake pumping also peaks in 2027 at approximately 27,500 gpm and declines to nothing when the pit lake becomes dewatered at the end of 2032. Well dewatering peaks in 2028 at approximately 33,000 gpm. Dewatering was simulated using 15 wells screened through the vertical extent of the underground workings and extending down approximately 100 ft below the workings. Dewatering well locations are shown in Figure 7-20. Of these 15 dewatering wells, 2

 

 

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are the existing test wells (WE-01 and WE-02). The remaining 13 dewatering wells will need to be constructed.

Figure 7-19 Simulated Dewatering Rates

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Residual passive inflow (RPI) represents seepage water within the underground workings that would need to be managed and sumped. RPI is limited to approximately 100 gpm throughout the predictive simulation. All managed water is discharged to RIBs. The peak annual discharge to RIBs is approximately 43,000 gpm in 2028. The annualized average discharge rate to the RIBs declines from the peak value in 2028 to approximately 24,000 gpm in 2037.

 

  7.2.6.4.

Groundwater Model Summary

A numerical groundwater flow model was developed to estimate dewatering rates required for the Project. Existing hydrologic datasets, including measured water levels and results from pumping tests, were used to calibrate the numerical flow model. Following model calibration, the model was used to simulate predictive dewatering rates.

 

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Figure 7-20 Dewatering Well Field Locations Used For Predictive Modeling

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  8.

Sample Preparation, Analysis and Security

 

  8.1.

Pre-2012

Of the 21 Echo Bay RC holes, only seven were presented with assay results. RPA was unable to determine the sample preparation laboratory or procedures for the Echo Bay and Newmont RC holes. RPA assumes that they were prepared to industry standards at the time either in-house or at a commercial facility. The Echo Bay samples were analyzed by Rocky Mountain Geochemical Corp. in West Jordan, Utah. The Newmont samples were analyzed by ALS Chemex in Sparks, Nevada. As per the ALS Chemex certificates, pulp samples were received and a 50-element aqua regia inductively coupled plasma (ICP) analytical package (ME-MS41) was run. The ICP elements, and their ranges in ppm or percent, are listed below:

Table 8-1 Pre-2012 ICP Analysis

 

                     

Ag

   0.01-100      Cu     0.2-10,000      Na     0.01%-10%      Ta     0.01-500
   

Al

   0.01%-25%      Fe    0.01%-50%      Nb    0.05-500      Te    0.01-500
   

As

   0.1-10,000          Ga    0.05-10,000          Ni    0.2-10,000          Th    0.2-10,000
   

B

   10-10,000      Ge    0.05-500      P    10-10,000      Ti    0.005%-10% 
   

Ba

   10-10,000      Hf    0.02-500      Pb    0.2-10,000      TI    0.02-10,000
   

Be

   0.05-1,000      Hg    0.01-10,000      Rb    0.1-10,000      U    0.05-10,000
   

Bi

   0.01-10,000      In    0.005-500      Re    0.001-50      V    1-10,000
   

Ca

   0.01%-25%      K    0.01%-10%      S    0.01%-10%      W    0.05-10,000
   

Cd

   0.01-1,000      La    0.2-10,000      Sb    0.05-10,000      Y    0.05-500
   

Ce

   0.02-500      Li    0.1-10,000      Sc    0.1-10,000      Zn    2-10,000
   

Co

   0.1-10,000      Mg    0.01%-25%      Se    0.2-1,000      Zr    0.5-500
   

Cr

   1-10,000      Mn    5-50,000      Sn    0.2-500          
   

Cs

   0.05-500        Mo    0.05-10,000        Sr    0.2-10,000              

Note: ppm unless otherwise indicated

Fire Assay (FA) with an atomic absorption (AA) finish was utilized for gold assays (Au-AA23 package). Any gold FA values over 3 ppm were rerun by gravimetric methods (Au-GRA21). The detection limit for both gold assaying methods was 0.005 ppm. (Roscoe Postle Associates Inc., 2017)

Victoria’s Cove samples were all prepared and analyzed by the Inspectorate assay laboratory located in Sparks, Nevada. The following discussion relates specifically to Victoria’s samples.

 

  8.1.1.

Sample Preparation Procedures

Upon receipt by Inspectorate the core samples were reviewed, sorted, and oven dried (230oF). The samples were crushed to +80% passing 10 mesh by jaw crusher and pulverized to +90% passing

 

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150 mesh by ring and puck. The samples were then split by a splitter; one half of the sample was set aside as the “reject” and the remaining half sample split again. This process was continued until the sample equalled three-fourths of the volume of a pulp envelope. The total rejects were tied, tagged, and palletized.

 

  8.1.2.

Laboratory Analysis Procedures

Gold assays were first run by FA with an AA finish with a detection limit of 5 ppb. Any gold FA values over 3 ppm were rerun by gravimetric methods. Silver assays were also run by FA/AA with a detection limit of 0.1 ppm.

A summary of Inspectorate’s FA method is described below:

 

   

Samples are received from weigh-room in labelled envelopes;

 

   

Crucibles are set up in trays of twenty by numbers assigned from Laboratory Information Management System (LIMS);

 

   

Crucibles are charged with the appropriate type and amount of flux;

 

   

Samples are transferred from the envelopes to the appropriately labelled crucible, copper spikes are inserted, and inquarting is conducted;

 

   

Additional reagents are added to the crucible if needed and sample and flux is mixed with cover flux added on to the top of charge;

 

   

Crucibles in sets of 80 charges are then loaded into pre-heated gas fusion furnace and fusion is conducted for one hour at 2,100°F;

 

   

Upon completion of fusion, molten lead-slag is poured into numbered conical moulds. Unsatisfactory fusions are submitted back to the weighing room for reweigh;

 

   

Fusions are allowed to cool and the moulds are transferred in order to the slagging station. Slag is removed with hammer, and lead buttons are cubed and placed in numbered trays;

 

   

MgO cupels are heat treated in the cupel furnace at 1,800°F for a minimum of five minutes to drive off moisture. Cupels are then carefully evaluated for cracks or erosion and are discarded accordingly;

 

   

Lead buttons are loaded into cupels in order and the set is then loaded with a fork into an electric oven set at 1,800°F;

 

   

Upon full cupellation (lead adsorption), the cupels are allowed to cool and the resulting Ag ± Au prills are placed into numbered trays;

 

   

For AA finish, the prills are dissolved in aqua regia and analyzed on the ICP; and

 

   

For gravimetric finish, the prills are placed in parting cups, approximately two-thirds full with 20% Nitric Acid to dissolve the silver, and then heated on a hotplate at 125°F until parted. The gold bead is then allowed to cool, transferred to cups, rinsed with cold de-

 

 

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ionized water, and allowed to dry. The cups are fired at 1,560oF for approximately five minutes, and then allowed to cool. The resulting doré bead is weighed on a microbalance.

A multi-element ICP analytical package was also run for most samples. The ICP elements determined including their detection limits in ppm are presented below:

Table 8-2 ICP Analysis 2012 - 2018

               

Ag

   0.1-100    Co    1-10,000    Mg    100-100,000    Sc    1-10,000
   

Al

   100-100,000    Cr    1-10,000    Mn    5-10,000    Se    0.2-1,000
   

As

   5-10,000    Cu    1-10,000    Mo    1-10,000    Sr    0.2-10,000
   

B

   10-10,000    Fe    100-100,000    Na    100-100,000    Ti    100-100,000
   

Ba

   10-10,000    Ga    0.05-10,000    Ni    1 -10,000    TI    10-100,000
   

Bi

   2-10,000    Hg    3-100,000    P    10-50,000    V    1-10,000
   

Ca

   100-100,000    K    100-100,000    Pb    2-10,000    W    10-5,000
   

Cd

   0.5-1,000    La    2 -10,000    Sb    2-10,000    Zn    2-10,000

 

  8.1.3.

Security

Security measures taken to ensure the validity and integrity of the samples collected included:

 

   

Chain of custody of drill core from the drill site to the core logging area;

 

   

Buildings were kept locked when not in use;

 

   

Core sampling was undertaken by technicians under the supervision of Victoria geologists;

 

   

All intersections were kept in the Reno office; and

 

   

Inspectorate was storing all the rejects and pulps indefinitely.

 

  8.2.

Premier 2012-2018

Drill hole samples collected by Premier were sent for assay analyses to three independent laboratories:

 

   

American Assay laboratory, Sparks, Nevada, accredited ISO/IEC 17025:2005;

 

   

Inspectorate America Corporation, Sparks, Nevada, accredited ISO 9001:2008 and ISO/IEC 17025:2005; and

 

   

ALS Minerals, Vancouver, British Columbia, accredited ISO/IEC 17025:2005.

From 2012 until end of 2014, samples were sent for analyses to Inspectorate laboratories. Starting with 2015, samples were sent to ALS. The pulp sample checks were sent to the American Assay laboratory.

 

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The sample preparation and gold FA procedures for the Premier 2012-2016 drilling programs at all the laboratories are essentially the same as described above except that gold FA results greater than 10 g/t Au are re-assayed by FA/gravimetric.

In addition to the fire assay analysis, the 2016-2022 program included analysis of gold and silver by screen metallic methods when visible gold was noted in the polymetallic sheeted veins intercepted in the 2201 zone. The program also incorporated a 42-element, four-acid, ICP-mass spectrometry, ultra-trace level analysis.

The sample preparation, analysis, and security procedures at the Project are adequate for use in the estimation of Mineral Resources.

 

  8.3.

i-80 2023-Present Underground Resource Conversion Drilling

When sampling is complete for a hole, the sample bin is picked up by a driver for ALS Minerals and delivered to the assay lab. ALS is independent of i-80. ALS Minerals is ISO 9001 and 17025:2017 certified. Samples are dried, weighed, screened, crushed to 70% passing 10 mesh, split to 250g with a riffle splitter, then pulverized to 85% passing 200 mesh. Samples submitted through ALS Minerals (4977 Energy Way, Reno, NV 89502 or 1345 Water St. Elko, NV 89801) were analysed for Au using 30g fire assay, aqua regia digestion with AAS finish (code Au-AA23), with detection range 0.005 to 10 ppm Au. Samples with an Au result greater than 10 ppm Au were analysed using 30g fire assay with gravimetric finish (code Au-GRAV21), detection range 0.05 to 10,000 ppm Au. Samples are also assayed with a 48 element suite (code ME-MS61) using 0.5g 4-acid digestion with ICP-AES finish. For samples containing >2 ppm Au an additional assay is performed for Hg by ICP-MS (Hg-MS42). The ALS ICP-AES facility is located at 2103 Dollarton Hwy, North Vancouver, BC, Canada. This drill program is ongoing and has not been included in the current resource estimate.

Cut core is stored on pallets at the Lone Tree core yard. Coarse rejects are returned from the assay lab and stored in 55-gallon drums at the Lone Tree core yard. Leftover sample pulp material is returned from the assay lab and stored in the pulp warehouse at Lone Tree.

 

  8.4.

Quality Assurance and Quality Control

 

  8.4.1.

Standards and Blanks

A total of 69 different blank and gold standard reference materials have been used at Cove. Table 8-3 presents the results of the most frequently assayed materials. The null hypothesis test compares the calculated t-statistic to the t value for a 95% confidence level. Acceptance of the test indicates that the lab mean is within the 95% confidence limit of the standard value. A rejection result from the test does not necessarily mean the data is not representative of the expected value but rather

 

 

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that the test was inconclusive. Groups which have a high out limit frequency are not necessarily reject by the t-test if the standard deviation for the group is not excessively high.

Table 8-3 Gold Blank and Standard Summary Statistics

                 
ID     Count    

  Lab Mean  

PPM

  

  Lab Std  

Dev

  

Out of

  Limit  

Rate

 

  Std. Value  

PPM

     t-statistic       ta /2     

Null

 Hypothesis 

Test

   

Blank

   1880    0.054    0.626    14%   0.005    3.380   1.646    Reject
   

CDN-GS-P8C

   166    0.777    0.108    3%   0.784    (0.843)   1.974    Accept
   

CDN-GS-P4E

   392    0.519    0.412    3%   0.493    1.256   1.966    Accept
   

SP37

   283    17.892    2.560    0%   18.140    (1.631)   1.968    Accept
   

CDN-ME-1301

   150    0.550    0.434    20%   0.437    3.180   1.976    Reject
   

CDN-GS-22

   144    22.600    2.523    14%   22.940    (1.618)   1.977    Accept
   

CDN-GS-5L

   193    4.747    0.551    10%   4.740    0.180   1.972    Accept
   

OREAS 503b

   116    0.695    0.013    0%   0.695    (0.086)   1.981    Accept
   

G912-1

   109    7.356    0.112    0%   7.290    6.132   1.982    Reject
   

OxJ120

   107    2.365    0.044    45%   2.365    0.033   1.983    Accept
   

CDN-GS-5H

   99    4.004    1.930    48%   3.840    0.847   1.984    Accept
   

CDN-GS-2M

   85    2.917    3.865    22%   2.210    1.686   1.989    Accept
   

CDN-GS-12

   77    9.423    1.770    32%   9.980    (2.760)   1.992    Reject
   

CDN-GS-11

   63    3.398    0.877    21%   3.400    (0.020)   1.999    Accept
   

CDN-GS-4D

   47    3.839    0.408    19%   3.810    0.483   2.013    Accept
   

G307-7

   32    7.964    0.102    0%   7.750    11.837   2.040    Reject
   

SQ48

   45    30.229    0.327    22%   30.250    (0.433)   2.015    Accept
   

OXI81

   43    1.815    0.126    28%   1.807    0.418   2.018    Accept
   

OXD87

   34    0.412    0.024    18%   0.417    (1.162)   2.035    Accept
   

CDN-GS-30

   33    33.553    0.786    3%   33.500    0.390   2.037    Accept
   

G909-4

   33    7.496    0.176    0%   7.520    (0.770)   2.037    Accept
   

CDN-GS-15B

   31    15.619    2.179    13%   15.980    (0.924)   2.042    Accept

 

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Figure 8-1 Blank Assay Results

LOGO

Figure 8-2 SP 37 Standard Reference Material Results

LOGO

 

 

 

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Figure 8-3 CDN-GS-22 Certified Reference Material Results

LOGO

Figure 8-4 CDN-GS-5H Certified Reference Material Results

LOGO

 

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Figure 8-5 GS912-1 Certified Reference Material Results

LOGO

 

  8.4.2.

Duplicate Assays

Duplicate assays are performed under two scenarios. The geologist can instruct the lab to duplicate the pulp of a specified sample (Figure 8-6) or the pulp can be sent to another lab for check assay (Figure 8-7). Both types of duplicates show good replication of assay values.

 

 

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Figure 8-6 Prep Duplicates - ALS Reno

LOGO

 

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Figure 8-7 Lab Check Duplicates

LOGO

It is the authors opinion that the sample preparation, security, and analytical procedures are adequate for the estimation of Mineral Resources.

 

 

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  9.

Data Verification

Practical Mining received the McCoy Cove drill hole database from Premier’s Geology team. Premier exported the data as csv files for Practical Mining from Maxwell Geoservices software. The authors imported the data into Maptek Vulcan software and identified holes within the resource area. The authors selected 5 percent of holes from the resource dataset for detailed review. The selected holes are a spatial and temporal sampling of the data, the majority consisting of holes drilled by Victoria and Premier because most older holes are in the mined area and supported by past production. Premier supplied copies of the raw data for the selected holes to the authors.

The authors compared the raw data with the corresponding records in the database. Records reviewed include assay values for gold and silver, collar location surveys, and downhole deviation surveys. The authors observed no significant problems with the data, and conclude the data is suitable for use in the resource estimation.

The authors did not observe any mismatches between assay certificates and the database. Minor inconsistencies in the handling of missing data were noted. Sampled intervals which lack assay data typically have a blank cell in the assay column of the csv, but holes AX-10 and AX-22 contained negative values. Those holes were subsequently corrected by re-importing into Maxwell Geoservices software. All missing data cells were assigned -99 for use in Vulcan software, including holes AX-10 and AX-22, so the database inconsistency did not affect the estimation.

Collar surveys are collected by professional land surveyors and reported to Premier in Excel spreadsheets. Collar surveys are occasionally duplicated on subsequent surveyor visits, and surveys will vary slightly due to limits in precision. The authors noted one collar with a slight mismatch between the surveyor’s spreadsheet and the database, however the small difference in distance has an insignificant effect on hole placement and may be attributed to multiple surveys of the same collar.

The authors did not observe any mismatches between downhole survey reports and the database. Table 9-1 summarizes the scope of the detailed drill data review.

 

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Table 9-1 Data Review Summary

 

           
      Holes in
Data Set
     Holes
Audited
     Collar Survey
Coordinates
Reviewed
     Downhole
Surveys
Reviewed
     Assay
Certificates
Reviewed
 
   

Number of Drill Holes

     1,397        70        70        70        88  
   

Percent of Population Reviewed

              5%        5%        5%        6%  

All holes were checked for overlapping intervals using Vulcan, and there were none. Hole traces were viewed in Vulcan to confirm there were no extreme survey deviations. Lithology was viewed in Vulcan to confirm that the geology model conforms to the geology data.

In summary, the authors observed no significant problems with the data, and conclude the data is suitable for use in the resource estimation.

 

 

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10. Mineral Processing and Metallurgical Testing

 

  10.1.

Historical Metallurgical Test Work

Metallurgical testing reviewed in regard to the Cove Project including date, laboratory, and listed accreditation are shown in Table 10-1.

Table 10-1 Metallurgical Testing Programs

 

       
Date    Laboratory    Accreditation    Program
2008   

Kappes Cassiday &

Associates (KCA)

7950 Security Circle

Reno, NV 89506

   No certifications listed on website. Independent of i-80.    Whole Ore Leaching and Flotation Tests;
2009   

Kappes Cassiday &

Associates (KCA)

7950 Security Circle

Reno, NV 89506

   No certifications listed on website. Independent of i-80.    Roasting and Cyanidation of Calcine, Hot Lime Treatment, and Flotation Tests
2017   

SGS Lakefield Research Ltd. (SGS)

185 Concession St,

Lakefield, ON K0L 2H0,

Canada

   Conforms to the requirements of the ISO/IEC 17025 standard for specific registered tests. Independent of i-80.    Whole ore leaching, roasting, pressure oxidation and cyanidation of oxidation products
2021   

Kappes Cassiday &

Associates (KCA)

7950 Security Circle

Reno, NV 89506

   No certifications listed on website. Independent of i-80.    Whole ore chlorination oxidation and leaching
2022   

FLSmidth Minerals Testing and Research Center

7068 S. FLSmidth Dr

Midvale, UT 84047

 

   No certifications listed on website. Independent of i-80.    Whole ore pressure oxidation

 

  10.1.1.

2008 KCA Program

The 2008 KCA test program was conducted on nine (9) composites from the Helen Zone. The testing included:

 

   

Bottle Roll direct cyanidation of each composite;

 

   

Bottle roll Carbon-In-Leach (CIL) cyanidation of each composite; and

 

   

Rougher and Scavenger Flotation on each composite.

The whole ore cyanidation tests gave generally poor gold extractions ranging from 1% to 23%.

 

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The CIL cyanidation tests gave higher gold extractions ranging from 49% to 82%.

The difference between the whole ore cyanidation and the CIL cyanidation indicates a pregnant solution robbing (“preg robbing”) component in the composites tested.

The flotation tests gave gold recoveries into a concentrate ranging from 24% to 59%. The corresponding concentrate weight recoveries ranged from 9 to 13%.

The flotation tests gold recoveries were low and did not demonstrate a strong amenability towards flotation.

Based on the suspicion that the relatively poor cyanidation results from the 2008 testing were due to the preg-robbing carbonaceous content of the materials tested the 2009 KCA test program investigated three types of processes to mitigate the effects of carbonaceous matter. Testing was conducted using a composite constructed from the composite remains from the 2008 program. The testing included:

 

   

Head Characterization for Carbonaceous and Sulfide Material in 2008 Drill hole interval samples used in 2008 composite construction;

 

   

Roasting followed by cyanidation of calcine using both direct cyanidation and CIL cyanidation of the calcines;

 

   

Hot Lime treatment of the Composite; and

 

   

Flotation.

The head analyses indicated organic carbon contents ranging from 0.03% to 0.96% with an average of 0.44%. The sulfide sulfur content ranged from 0.15% to 1.79% with an average of 1.02%. The assays confirmed the presence of carbonaceous material as well as potential refractory aspects related to sulfide sulfur content.

Roasting tests were conducted using a 650°C for two hours. The gold extraction for the direct cyanidation of the calcine was 87% while the extraction using CIL cyanidation of the calcine was 90% which indicates that after calcination there are still active preg-robbing factors.

The hot lime treatment was conducted on a sample of the composite ground to 80% passing 74 µm to which a lime addition of 100 lbs/ton was made. Hot lime treatment may be effective in recovering previously preg-robbed gold. The slurry was heated to 100°C and agitated for 8 hours. The slurry was then leached with cyanide. The gold extraction for the hot lime test was 40%.

Two flotation tests were conducted, the first using a rougher, scavenger, cleaner simulation, the second simulating four stages of rougher flotation. The gold recovery from the first test was 54%

 

 

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into a concentrate with a 17.6% weight recovery. The second test gave a gold recovery of 31% into a concentrate with a weight recovery of 20.7%.

 

  10.1.2.

2009 KCA Program

The 2009 tests confirmed the presence of carbonaceous material, the likely cause of preg-robbing observed in the whole ore cyanidation tests.

The tests indicated that roasting and calcine cyanidation may be an effective treatment for the material tested. The hot lime treatment and flotation tests did not match the roast and calcine cyanidation gold recoveries.

 

  10.1.3.

2017 SGS Programs

Eleven composites from the Helen Zone and ten from the Gap were sent to SGS Canada Inc., Lakefield, Ontario, Canada in 2017. The samples were selected from drill holes and drill hole intervals to provide spatial representation of the deposits both vertically and horizontally. The primary objectives of the test program included:

 

   

Head grade characterizations of the physical and metallurgical properties of each resource to meet requirements for third-party processing;

 

   

Provide preliminary metallurgical data for the resource targets for potential metallurgical third-party processing;

 

   

Determine precious metal extractions, and deportments, reagent consumptions, with the following process routes:

 

   

Whole ore leaching;

 

   

Roasting followed by calcine leaching; and

 

   

Pressure oxidation followed by leaching of neutralized slurry.

 

   

Roasting and pressure oxidation test conditions used were based on those provided by a potential third-party toll processor.

The SGS program scope did not include testing to optimize roasting or pressure oxidation conditions nor to develop process design data a purpose-built processing facility for Cove.

Gold head grades for the samples were:

 

   

Helen samples averaged 11.7 g/t Au, ranging from 4.50 g/t Au to 32.6 g/t Au;

 

   

Gap samples averaged 17.1 g/t Au, ranging from 4.38 g/t Au to 37.6 g/t Au.

Baseline leach tests showed the extreme refractory nature of the Cove deposit. Results from 24 hours leaching showed:

 

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Helen samples averaged 10.8% Au recovery, skewed by one test with a recovery of 90.8% Au. Without this result, the average is 2.5% Au.

 

   

Gap samples averaged 1.2% Au recovery.

Roasting conditions used are shown in Table 10-2.

Table 10-2 Roasting Test Conditions

 

   
Conditions    Value
 

Stage 1:

   

Temperature:

   530°C
   

CO2 flow:

   0.8 L/min
   

O2 flow:

   1.2 L/min
   

Time:

   30 min
 

Stage 2:

   

Temperature:

   570°C
   

O2 flow:

   2 L/min
   

Time:

   15 min

The roasting and calcine leach tests showed the following:

 

 

Sulfide oxidation with the Helen composites ranged from 85.9% to 97.0% while for the Gap composites ranged from 87.9% to 98.1%;

 

 

Carbonate oxidation was inconsistent, but the Helen composites was generally low but with the Gap composites was somewhat higher. Low carbonate concentrates often resulted in negative oxidations;

 

 

Gold leach recoveries from the Helen composite calcines was variable ranging from 63.5% to 90.8%;

 

 

Silver leach recoveries from the Helen composite calcines was variable ranging from 9.6% to 56.5%;

 

 

Gold leach recoveries from the Gap composite calcines was variable ranging from 54.4% to 89.4%; and

 

 

Silver leach recoveries from the of Gap composite calcines was also variable ranging from 23.1% to 77.0%.

Pressure oxidation (POX) tests were conducted under fixed conditions and not for optimization of oxidation conditions. The tests were conducted under the following conditions:

 

 

225°C;

 

 

60 minutes retention time;

 

 

700 kPa oxygen overpressure.

 

 

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The pressure oxidation and oxidation product leach tests showed:

 

 

Pressure oxidation resulted in high sulfide oxidation of both Helen and Gap samples;

 

 

Carbonate removal in the Helen composites averaged 97.2% while carbonate removal in the Gap composites averaged 82.5% but this average was skewed lower from two Gap composites having very low head carbonate contents;

 

 

Gold leach recoveries of Helen composite POX products ranged from 0.3% to 96.6%. This indicates that even with oxidation of sulfide and to a lesser extent carbonate removal, pressure oxidation is less effective than roasting for improving gold recoveries in Helen samples;

 

 

Silver leach recoveries of Helen composite POX products ranged was highly variable ranging from 6.7% to 69.6%;

 

 

Gold leach recoveries of Gap composite POX products was variable ranging from 5.7% to 73.6%. This indicates that even with oxidation of sulfide and to a lesser extent carbonate removal, pressure oxidation is less effective than roasting for improving gold recoveries in Gap samples;

 

 

Silver leach recoveries of Gap composite POX products was also variable ranging from 52.5% to 81.7%; and

 

 

The data set was too small to establish any clear relationship between mineralogy, head grade and leach recovery although it is clear that mineralogical factors such as arsenic grade and total carbonaceous matter or total organic content impact leach recoveries with pressure oxidation and POX product leaching.

A second phase of testing was completed to investigate the reasons for the low recoveries with both roasting and pressure oxidation that occurred in the Phase 1 tests. Calcine and POX products were generated on selected composites from the Helen and Gap, using the same conditions from phase 1. The calcines and POX products were split in two. One half of each spilt was leached under the same conditions used in phase 1. The other half of each split was processed using carbon-in-leach (CIL). CIL was used to determine if preg robbing was still occurring from active organic carbon species still present in either the calcine or POX product.

Leach and CIL tests on the POX product showed that:

 

With the Helen samples:

 

   

Gold leach (48 hours) recoveries ranged from 0.6% to 5.1% while CIL tests (also 48 hours) showed recoveries from 62.3% to 81.9%, significantly higher than the direct leach;

 

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Silver leach (48 hours) recoveries ranged from 36.2% to 86.9% while with CIL tests ranged from 76.8% to 86.9% significantly higher than the direct leaching;

 

 

With the Gap samples:

 

   

Gold leach (48 hours) recoveries ranged from 1.6% to 77.8% while with CIL tests (also 48 hours) ranged from 70.5% to 95.9%, significantly higher than the direct leach;

 

   

Silver leach (48 hours) recoveries ranged from 19.9% to 84.1% while with CIL tests ranged from 71.6% to 87.0%, significantly higher than the direct leach;

The phase 2 POX tests on the selected composites confirmed that preg robbing occurred in direct leaching of the POX products and that CIL can increase gold and silver recoveries significantly for both the Helen and Gap samples.

The phase 2 roasting tests showed that CIL cyanidation of the calcine could increase gold and silver recoveries but were lower than expected. Diagnostic leach tests of the phase 2 calcine leach residues were completed to investigate gold deportment. The diagnostic leach test results showed:

 

 

The estimated amount of gold associated with iron oxides, ferrites or calcite in the Helen composites calcine leach residues ranged from 8.2% to 17.9% and averaged 11.0%, with the remaining gold estimated to be in siliceous gangue which ranged from 9.2% to 18.0% and averaged 12.8%;

 

 

The estimated amount of gold associated with iron oxides, ferrites or calcite in the Gap composites calcine leach residues ranged from 11.7% to 35.9%, with the remaining gold estimated to be in siliceous gangue which ranged from 2.6% to 14.0%;

 

 

The gold deportment tests show that the Helen has more gold associated with siliceous material than the Gap samples which showed more gold associated with the iron oxides, ferrites, or calcite following roasting; and

 

 

The data also suggests that the specified roasting conditions from a potential toll roasting operation may not be optimal for the Helen or Gap material.

One possible development path for the Cove project is third party processing of production through either existing roasting and calcine cyanidation or existing pressure oxidation and residue cyanidation facilities.

Premier Gold solicited two items from a prospective toll operator with both roasting and pressure oxidation (POX) processes and their associated cyanidation processes for the respective calcines or POX residues.

 

 

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The first item included the test protocols and test conditions for laboratory bench scale batch roasting and pressure oxidation test conditions for the 2017 metallurgical testing. The conditions provided approximate the expected operating conditions in the prospective toll operator’s roasting and POX facilities.

The second item Premier Gold solicited was terms and conditions for toll milling and treating Helen resource material. Premier Gold provided a package of Helen metallurgical data for the roasting and POX tests from the 2017 test program to the prospective toll process operator for their consideration and as the basis for toll processing resource material through either the toll operator’s roasting or POX facilities.

The test data indicates that the Helen composites were generally more amenable to roasting and calcine CIL cyanidation than POX and residue CIL cyanidation. The assay data for the Helen composites indicates that there may be some problems from some areas to meet roaster feed specifications. Onsite blending of Helen resource material to meet specifications prior to shipping to the toll processor provided that resource material is available for blending will likely be required.

Conversely, the Gap composite test data were generally more amenable to POX and residue CIL cyanidation. Again, blending would likely have to be used prior to shipping offsite to provide on specification material to the toll processor.

 

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  10.1.4.

2021 KCA Program

 

 

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In May 2020, a metallurgical test program was initiated to evaluate a lower capital cost oxidation method of Cove samples. The Cove resource does not support the capital required for a whole ore roasting operation, which includes a dry crushing and grinding circuit. Chlorination offers a lower capital cost approach but typically has higher operating costs. The purpose of the tests is to determine if chlorination provides a viable process route. Optimization testing to support process design will be required if the method is found to be a viable option.

Chlorination is a proven process that was used to oxidize sulfide minerals and deactivate active carbonaceous mineralization. This process was used in the 1980’s and 1990’s at two Nevada operations (Newmont Carlin and Jerritt Canyon, John O. Marsden, 2006). Chlorination provides a lower capital cost approach for treatment of refractory sulfide ore requiring a whole ore treatment approach (rather than concentrate). Chlorine is highly soluble in water and dissolves to form hydrochloric and hypochlorous acids. Both are strong oxidizing agents and will oxidize sulfide minerals. The mechanism for deactivating of carbonaceous mineralization is not well understood but from operational experience has been shown to be effective with organic carbon concentrations in the range of 0.5% to 1.0%.

The scope of work for the program evolved to a two-phase program:

Baseline leach and CIL tests followed by oxidation with calcium hypochlorite at ambient and elevated temperature. Oxidation products were then subjected to CIL tests.

Oxidation using chlorine gas followed by CIL tests.

The program includes testing of two Helen composite samples and two Gap composite samples:

•   G1: CSD Gap Upper Zone in Favret lithology formation with an average head grade of 11.55 g/t Au;

•   G2: CSD Gap Lower Zone in Favret lithology formation with an average head grade of 13.65 g/t Au;

•   H1: Helen upper zone with Panther Canyon and Home Station lithology formation, with generally less organic carbon than other zones with an average head grade of 9.84 g/t Au;

•   H2: Helen lower zone with Favret lithology formation, which includes previous defined metallurgical zone H2 to H5 with an average head grade of 16.13 g/t Au.

The Gap samples contained an average 0.23% arsenic and 4.4 g/t mercury while the Helen samples contained an average0.15% arsenic and 12 g/t mercury. The mercury concentrations are elevated and will require capture and abatement in the process.

The results from both phases of this program are shown in Table 10 3. Phase 1 results are based on an equivalent chlorine gas addition of 280 kg/t (140 kg/t available chlorine) and 120 kg/t added NaOH. Phase 2 results are based on 70 kg/t added chlorine gas (35 kg/t available chlorine) and 80 kg/t added NaOH. While Phase 1 was exploratory in nature, Phase 2 showed that the chlorine oxidation process for the Cove Project is feasible.

 

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Table 10-3 Chlorination Program Test Results

     
     Phase 1   Phase 2

KCA

Sample

No.

   Description   

Direct

Leach

 Extraction1 

(% Au)

 

CIL

  Recovery  

(% Au)

 

Ca(OCl)2

  Oxidation2   

Recovery,

(% Au)

 

Hot

Ca(OCl)2

  Oxidation +  
Abundant

NaOH3

Recovery

(% Au)

 

Hot

Ca(OCl)2

Oxidation

+
  Abundant  

NaOH2

(% Au)

 

Hot

Chlorine

Gas

  Oxidation1  

Recovery

(% Au)

88501

A

 

G1

Composite

  0%   18%   51%   84%   73%   90%

88502

A

 

G2

Composite

  0%   10%   41%   91%   67%   77%

88503

A

 

H1

Composite

  25%   51%   89%   86%   80%   84%

88504

A

 

H2

Composite

  0%   19%   72%   92%   72%   90%
                                     
Projected  Recovery   6%   25%   63%   88%   73%   85%

Notes:

  1.

Target grind k80 = 75 µm

  2.

Target grind k80 38 µm

  3.

Target grind k80 = 20 µm

The results demonstrate the refractory nature of the samples with direct gold leach extractions of 0% in three of the four samples. CIL gold recoveries ranged from 10% to 51%, averaging 25%.

 

  10.1.4.1.

Phase 1 Results

Ambient chlorination with calcium hyprochlorite with 140 kg/t NaOH showed CIL gold recoveries from 41% to 89%, averaging 63%. Hot (70oC) chlorination with calcium hypochlorite and excess NaOH showed CIL gold recoveries from 84% to 92%, averaging 88%.

 

  10.1.4.2.

Phase 2 Results

The initial testing included calcium hypochlorite at 75oC plus excess NaOH at a slightly coarser grind the phase 2 hot tests (k80 = 38 µm instead of 20 µm). The results were inferior with CIL gold extractions ranging from 67% to 80%, averaging 73%. The next set of tests were conducted with at a coarser grind (k80 = 75 µm), with chlorine gas and lime for pH control. The results were an improvement with CIL gold extractions ranging from 77% to 90%, averaging 85%. Although not as high as the hot phase 1 tests, the coarser grind and the use of lime instead of NaOH results in a more economic process.

 

 

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  10.1.5.

2022 FLSmidth Program

Remaining samples from the 2021 KCA program were sent to the FLSmidth Minerals Testing and Research Center for pressure oxidation testing with bench top autoclaves. Pressure oxidation conditions included:

 

   

Grind to k80 = 70 µm.

   

Slurry density = 35% solids.

   

Trona addition = 10 kg/t.

   

Temperature = 195oC

   

Retention time = 45 minutes

   

Oxygen overpressure = 760 kPa

CIL tests were performed on the oxidized samples under the following conditions:

 

   

Temperature = 60oC;

   

Slurry density = 35% solids;

   

Cyanide addition = 2.5 kg/t NaCN.

   

Carbon concentration = 20 g/L slurry.

   

Retention time = 24 hours.

Table 10-42022 FLSmidth Program Test Results

The results of the testing are shown in Table 10-4.

 

           

Sample

No.

   Description   

  Calc..Head  

Grade (g/t

Au)

 

CIL

Residue

Grade (g/t

Au)

  

CIL Recovery

(% Au)

  

S= Oxidation2

(%)

88501

A

  

G1

Composite

   12.28   4.79    59.6    22.9

88502

A

  

G2

Composite

   14.03   6.30    57.7    64.2

88503

A

  

H1

Composite

   9.54   1.35    90.2    0.0

88504

A

  

H2

Composite

   16.03   5.76    63.3    48.2

 

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The results are anomalous with the highest Au recovery coming from the sample with no sulphide oxidation. This sample had a baseline CIL gold extraction of 51% from the 2021 KCA program. Pressure oxidation increased baseline CIL recoveries for all samples. The Gap sample CIL recoveries are below those obtained in the 2017 SGS program while the Helen samples are roughly in line with the 2017 SGS results.

 

  10.2

Mineral Processing and Metallurgical Discussion

Testing programs completed demonstrate that samples from Gap and Helen zones are refractory to standard cyanide leaching. The samples showed significant levels of preg robbing from naturally occurring carbonaceous matter and refractory gold in sulfide minerals. Some samples showed better amenability to roasting while others showed better amenability to pressure oxidation. Processing of Cove production will require a facility with refractory treatment either in a new process plant or through an existing operation that can provide third party processing.

 

  10.2.

QP Opinion

In the opinion of the QP, the data are adequate for the purposes used in this Technical Report and the analytical procedures used in the analyses are of conventional industry practice. More work will be performed in the future to characterize the Cove deposit.

The main deleterious elements in the Cove samples are arsenic and mercury. The arsenic is encountered in the sulfide plant from arsenopyrite, which would be oxidized either by roasting or pressure oxidation. Both refractory processes are able to fix arsenic for safe disposal within tailings streams. Mercury is encountered in the cyanide leaching circuits and is recovered with specialized capture and abatement equipment within carbon elution and regeneration and gold refining areas of a process plant. The QP considers these methods of mitigation to be consistent with best industry practices.

 

  10.3.

Conclusions and Recommendations:

The following are the major conclusions and recommendations from the historical metallurgical test programs:

 

 

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  10.3.1.

Conclusions:

 

  1.

Head assaying for both the Helen Zone and Gap indicated that the gold in the two resources will likely be finely disseminated and not amenable to gravity gold recovery;

The mineralogy of the Helen and Gap resources differ in two significant areas, the first being that the Helen appears to be lower in arsenic content than the Gap resource and that the Gap resource appears to be lower on average in TCM and TOC than the Helen resource;

The Helen composite arsenic assays indicate that the Helen mineral resource is lower in arsenic content that the Gap resource;

The Helen and Gap resources based on the composites tested appear to be doubly refractory to conventional cyanidation and require both sulfide oxidation and passivation of active carbonaceous mineralization to significantly increase gold extractions;

Based on the composites tested the Helen Zone appear to generally be more amenable to roasting and CIL processing;

Based on the composites tested, the Gap resource appears to generally be more amenable to pressure oxidation and CIL processing;

The data set was too small to establish any clear relations between mineralogy and metal head grade and extractions for either resource although it is clear that mineralogy factors such as arsenic content and TCM or TOC are influencing extractions using either roasting and calcine cyanidation or pressure oxidation and residue cyanidation.

 

  10.3.2.

Recommendations

 

  1.

Additional metallurgical testing will be needed to thoroughly investigate the variability and viability of Helen and Gap resources to roasting and pressure oxidation with CIL cyanidation for which a program evaluating thirty to forty composites from each resource is suggested with objectives as follows:

 

 

Determine the location and number of samples required to represent the resources through geo-metallurgical analysis

 

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Assess variability of the responses to roasting and calcine cyanidation across the resources;

   

Assess variability of the responses to pressure oxidation and residue cyanidation across the resources;

   

Testing should attempt to establish head grade and extraction relations for use in more detailed resource modelling;

   

Mineralogy impacts need to be established and geologic domains within each resource need to be determined, and;

   

Additional comminution data should be collected to assess variability within the resources.

The resource model should be advanced to include arsenic, TCM, TOC, mercury, lead, zinc, total copper selenium, barium, cobalt, nickel, and cadmium as these will be important for predicting grades if toll process offsite is used and potentially for estimating extractions within the resources;

The estimated cost for the suggested next phase metallurgical program is to $850,000 based on current market pricing.

 

 

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  11. 

Mineral Resource Estimates

 

  11.1.

Definitions

Securities and Exchange Commission (SEC) S-K regulations (Title 17, Part 229, Items 601 and 1300 through 1305 provides the following definitions for mineral resources:

Mineral resource is a concentration or occurrence of material of economic interest in or on the Earth’s crust in such form, grade or quality, and quantity that there are reasonable prospects for economic extraction. A mineral resource is a reasonable estimate of mineralization, taking into account relevant factors such as cut-off grade, likely mining dimensions, location or continuity, that, with the assumed and justifiable technical and economic conditions, is likely to, in whole or in part, become economically extractable. It is not merely an inventory of all mineralization drilled or sampled.

Inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. The level of geological uncertainty associated with an inferred mineral resource is too high to apply relevant technical and economic factors likely to influence the prospects of economic extraction in a manner useful for evaluation of economic viability. Because an inferred mineral resource has the lowest level of geological confidence of all mineral resources, which prevents the application of the modifying factors in a manner useful for evaluation of economic viability, an inferred mineral resource may not be considered when assessing the economic viability of a mining project, and may not be converted to a mineral reserve.

Indicated mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of adequate geological evidence and sampling. The level of geological certainty associated with an indicated mineral resource is sufficient to allow a qualified person to apply modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Because an indicated mineral resource has a lower level of confidence than the level of confidence of a measured mineral resource, an indicated mineral resource may only be converted to a probable mineral reserve.

Measured mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of conclusive geological evidence and sampling. The level of geological certainty associated with a measured mineral resource is sufficient to allow a qualified person to apply modifying factors, as defined in this section, in sufficient detail to support detailed mine planning and final evaluation of the economic viability of the deposit. Because a measured mineral resource has a higher level of confidence than the level of confidence of either an indicated

 

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mineral resource or an inferred mineral resource, a measured mineral resource may be converted to a proven mineral reserve or to a probable mineral reserve.

The mineral resource estimate presented herein is an update of the previous technical report, there were no material changes to the methodologies or assumptions within the estimation process from the previous Technical Report and no new drill data has been incorporated in this estimate.

The effective date of this mineral resource estimate is December 31, 2024. All data coordinates are referenced to UTM Zone 11N NAD83 international survey feet and quantities are given in imperial units unless indicated otherwise.

The gold and silver mineralization at the Project was estimated using Vulcan versions 9.1.8 and 11.1.0 modeling software using the Inverse Distance Cubed (ID3) estimation method. A Nearest Neighbor method was run for comparison. The estimate was performed by Practical Mining LLC.

The Cove area includes four distinct mineralized zones: CSD, GAP, Helen, and 2201. The mineralized zones follow a southeast to northwest trend beginning below the historic Cove pit and extending over 6,000 feet to the northwest. Figure 11-1 shows the location of the zones.

 

 

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Figure 11-1 Plan View of Cove Mineralized Zones

 

LOGO

Zones are bounded by fault blocks. The Helen Zone lies north of the Gold Dome fault. The GAP zone lies between the Gold Dome and 110 Faults. A prospective, unmodeled zone lies between the 110 and Cay faults. The CSD and 2201 Zones lie south of the Cay fault. All zones are bounded by the CR fault to the northeast. Figure 11-2 shows the faults bounding the mineralized zones.

 

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Figure 11-2 Section View of Cove Mineralized Zones looking NE

 

LOGO

The Helen Zone is divided into four sub-zones: Upper Helen, Lower Helen, Upper Helen Wedge, and Lower Helen Wedge. The Upper Helen Zone is situated in the Home Station and Favret Formations, while the Lower Helen Zone is in the Dixie Valley Formation. The JE Fault cuts through the northern one-third of the Helen Zone striking east-west and dipping 68° N. The offset forms a wedge of mineralized material between the JE and CR faults. The Helen sub-zones are shown in Figure 11-3.

 

 

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Figure 11-3 Section View of the Helen Zone looking NW

 

LOGO

Mineralization included in the current estimate is characterized predominantly as disseminated Carlin style, except for the 2201 zone, which is polymetallic. Some polymetallic mineralization has been observed in the Gap Hybrid zone. Mineralization is controlled both lithologically and structurally. Disseminated mineralization tends to occur in lenticular geometries following both favorable bedding and T1-type sills, which are generally low angle. The sills are depicted in Figure 11-2 and Figure 11-3. Polymetallic vein mineralization is also lithologically and structurally controlled, generally with higher grades occupying narrow high angle structures with adjacent moderate grades lying along favorable low angle bedding. Figure 11-4 shows bedding parallel mineralization with high-angle mineralized structures in the 2201 zone.

 

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Figure 11-4 Section View of the 2201 Zone looking NW

 

LOGO

 

  11.2.

Modeling of Lithology and Mineralization

Premier geologists prepared preliminary geologic and grade shell models for the Cove area based on drill hole logging, assay data and geologic mapping. The Premier models served as the basis for resource modelling performed by Laura Symmes, Senior Geologist Practical Mining LLC.

Lithologic contacts were modeled by connecting corresponding logged drill hole intercepts in adjacent holes to form a surface representing each geologic formation. Surfaces were also created for significant lithologies associated with mineralization, including mafic sills. Faults were modeled using drill hole intercepts and by interpreting offset of lithologic surfaces. While structural interpretation is ongoing, the authors find the current lithology and structure models to be reasonable and applied no significant edits to Premier’s work. Table 11-1 lists the database codes for the modeled lithologic surfaces.

 

 

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Table 11-1 Geology Codes

 

   
Formation Name    Abbreviation
   

Tuff of Cove Mine

   Tc
   

Augusta- Smelser Pass member

   Tras
   

Augusta- Panther Canyon member

   Trap
   

Augusta- Home Station member

   Trah
   

Favret

   Trfv
   

Dixie Valley

   Trdv
   

T1 Mafic Sill

   T1
   

T2 Mafic Sill

   T2

Gold mineralization was modelled on 100-foot vertical sections facing azimuth 315. Polygons were digitized around drill hole intercepts with significant gold assay values. Intercepts in adjacent holes were connected within a polygon so that the polygons lie generally parallel with bedding and sills. Using the lithology model as a guide, polygons on adjacent sections at similar stratigraphic depths were connected to create mineral lenses. The mineral lenses were then trimmed against ore controlling faults. To model the higher grades in the 2201 zone, very high-grade intercepts were connected with high angle polygons oriented sub-parallel to the CR fault. The remaining moderate grade intercepts were digitized parallel to bedding.

i-80’s grade model conformed to a strict 3 g/t cutoff. PM modified this to allow lower grades locally in order to maintain continuity so long as the composite grade of the interval remained above 3 g/t. Each mineral lens was assigned a unique numerical code as listed in Table 11-2.

Table 11-2 Identification Codes for 3 g/t Grade Lenses

 

   
Zone    Mineral Lens Codes
   

CSD_GAP

   2203, 2204, 2205, 2206, 2207, 2208, 2209
   

GAP Hybrid

   5001, 5002, 5003, 5004, 5005, 5006, 5007, 5008, 5009, 5010, 5011
   

CSD

   1101, 1104, 1105, 1106, 1107, 1108, 1109, 1112, 1113, 1114, 1115, 1116, 1117, 1118, 1119, 1120
   

Upper Helen

   3101, 3102, 3103, 3104, 3105, 3106, 3107
   

Upper Helen Wedge

   3301, 3302, 3303, 3304, 3305, 3306
   

Lower Helen

   3202, 3203, 3204, 3205, 3206, 3207, 3208, 3209, 3210, 3211
   

Lower Helen Wedge

   3400, 3401, 3402, 3403, 3404, 3405, 3406, 3407, 3408, 3409
   

2201 high grade

   1301, 1302, 1303

 

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Zone    Mineral Lens Codes
   

2201

   1400, 1401, 1403, 1404, 1405, 1406

Practical Mining also digitized a low grade mineral envelope at an approximate 0.2 g/t cutoff which surrounds all the zones except 2201. The later does not have sufficient data to construct a low-grade envelope. The low-grade envelope was divided by zone and assigned codes. In Figure 11-5, the low-grade halo is translucent with the 3 g/t lenses visible inside.

Figure 11-5 Low Grade Envelope

 

LOGO

Several areas of low grade mineralization outside the low grade mineral envelope were identified and modelled and assigned codes including suffix X. Low grade codes are listed in Table 11-3.

Table 11-3 Identification Codes for 0.2 g.t Grade Lenses

 

   
Zone    Mineral Lens Codes   
   

Low_CSD_GAP and Gap Hybrid

   22000
   

Low_CSD

   11000
   

Low_Upper Helen

   31000
   

Low_Upper Helen Wedge

   33000
   

Low_Lower Helen

   32000
   

Low_Lower Helen Wedge

   34000
   

Low_NE of CR Fault

   10000

 

 

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Zone    Mineral Lens Codes     
   

Low_Other

   1100X, 2200X, 3300X, 3400X

Figure 11-6 shows the modeled grade lenses and low-grade halo on a section in the GAP zone.

Figure 11-6 Section Looking AZ 315 Showing Mineralized Lenses in the GAP Zone

 

LOGO

 

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  11.3.

Drill Data and Compositing

 

  11.3.1.

Drill Data Set

The drill data set used for the resource estimation contains 1,397 drill holes totaling 1,127,481 feet of drilling, of which 579,443 feet is RC and 548,038 feet is core. Premier has identified a subset of RC holes drilled prior to 2012 which may be affected by grade contamination, and those holes were excluded from the data set used for the estimation. The remaining RC holes correlate well with the surrounding core holes. One hole, NW-9A, was excluded due to lack of survey data.

Premier provided the drill data to Practical Mining as csv files. Gold and silver assays were converted from g/t to opt by dividing by 34.2857 in Excel, and blank values were assigned the value -99. The CSV files were then imported into a Vulcan ISIS database. A flag field was added to the ISIS database to contain numerical code of modeled lenses. Samples within the grade model polygons were flagged with the corresponding mineral lens code using the Vulcan flagging utility. The 3 g/t polygons take precedence over the low-grade polygons for lens code flagging. Of the 1,397 holes analyzed, 1,204 intersect at least one modeled grade polygon. Of these, 370 were flagged by the 3 g/t polygons and 1,195 were flagged by the low-grade polygons. An overview of drill hole and sample statistics is shown in Table 11-4.

Table 11-4 Drill Hole Summary

 

Data

Population

          Core        RC        Total  
     

All Holes

   No. Holes      387        1,010        1,397  
   Length Drilled (ft)      547,787.0        579,694.0        1,127,481.0  
   No. Samples      74,913        101,637        176,550  
   Length Sampled (ft)      430,357.5        572,744.0        1,003,101.5  
     

Holes with

Flags for

3g Lenses 

   No. Holes      195        175        370  
   Length Drilled (ft)      329,645.2        49,160.0        378,805.2  
   No. Flagged Samples      1,957        1,189        3,146  
   Length Flagged Samples (ft)      8,941.2        5,950.0        14,891.2  
     

Holes with

Flags for

Low Grade 

Lenses

   No. Holes      370        825        1,195  
   Length Drilled (ft)      513,909.9        463,031.0        976,940.9  
   No. Flagged Samples      32,094        36,946        69,040  
   Length Flagged Samples (ft)      161,085.2        187,141.5        348,226.7  

 

 

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  11.3.2.

Compositing

Gold and silver assay values were composited into 5-foot lengths beginning at the drill hole collar. Compositing intervals were truncated and a new compositing interval was begun where the drill hole intersects a modeled grade shell. Only samples with like flags may be composited together. If the intercept length within the grade shell is less than 5 feet, the composite consists of only that length contained within the polygon. The numerical lens flag was recorded with each composite in the composite database for use in the mineral resource estimation.

The total flagged composite length is 361,800 feet from 1,204 drill holes. Composite statistics by zone are shown in Table 11-5.

Table 11-5 Composite Summary

 

Zone

 

  

Mineral Lens
Codes

 

  

Number of
Holes

 

  

Number of
Composites

 

    

Length of
Composites
(ft)

 

 
   
CSD_GAP    220X    27      327        1,429.4  
   
GAP Hybrid    500X    27      133        500.0  
   
CSD    110X    269      1,975        8,820.3  
   
Upper Helen    310X    31      191        824.2  
   
Upper Helen Wedge    330X    26      141        615.9  
   
Lower Helen    320X    22      168        736.0  
   
Lower Helen Wedge    340X    30      394        1,784.0  
   
2201 high angle    130X    7      28        100.2  
   
2201 low angle    140X    6      25        90.0  
   
Low_CSD_GAP and Gap Hybrid    22000    157      10,050        48,039.6  
   
Low_CSD    11000    596      33,785        166,141.7  
   
Low_Upper Helen    31000    50      1,917        9,251.4  
   
Low_Upper Helen Wedge    33000    43      2,278        11,052.7  
   
Low_Lower Helen    32000    26      728        3,350.5  
   
Low_Lower Helen Wedge    34000    25      802        3,732.5  
   
Low_NE of CR fault    10000    295      8,518        41,468.7  
   
Low_Other    1100X, 2200X,
3300X, 3400X
   619      13,057        63,863.5  

 

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  11.4.

Density

Premier augmented their density data set in 2017 by submitting 29 samples from modeled 3 g/t zones to ALS for analysis. The new data includes 23 samples from the Helen and Gap zones, five samples from the 2201 zone and one sample from the CSD zone. The density data set was filtered by analysis method, and only samples analyzed using the water displacement method with wax coating were used. The data were then sorted by zone and grade, and results were averaged by zone. Results indicate that densities are similar for samples from 3 g/t grade shells and samples from low grade shells. The densities calculated for each zone are listed in Table 11-6.

Table 11-6 Density

 

     
Zone   

 Density 

(ton/ft3)

    

 Number 

Samples

   

Helen

   0.0691      29
   

GAP and GAP

Hybrid

   0.0708      17
   

CSD

   0.0772      25
   

2201 Veins

   0.0826      7
   

2201 Replacement

   0.0984      13
   

East of CR Fault

   0.0677      Default
value

 

  11.5.

Statistics and Variography

Univariate statistics for gold and silver composites within the 3 g/t grade shells and low-grade shells are presented in Table 11-7 and Table 11-8 below. The composite data are not closely spaced enough to permit construction of valid variograms.

 

 

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Table 11-7 Gold Composite Statistics

 

Zone   

No.

 

Samples

     Median      Max      Min      Mean     

Std.

 

Dev.

     Variance     

Upper

 

95% CI

    

Lower

 

95% CI

 
   
CSD      1444        0.089        20.602        0.000        0.188        0.739        0.035        0.226        0.149  
   
GAP      342        0.180        1.616        0.007        0.288        0.276        0.079        0.318        0.259  
   
Gap Hybrid      154        0.111        1.182        0.002        0.158        0.173        0.064        0.186        0.131  
   
Upper Helen      225        0.133        1.570        0.002        0.205        0.228        0.053        0.235        0.175  
   
Upper Helen Wedge      190        0.112        0.752        0.001        0.131        0.104        0.012        0.145        0.116  
   
Lower Helen      180        0.196        3.943        0.002        0.338        0.445        0.142        0.403        0.273  
   
Lower Helen Wedge      375        0.201        2.330        0.000        0.315        0.325        0.108        0.348        0.282  
   
2201 Vein      103        0.119        5.320        0.006        0.315        0.597        1.041        0.430        0.200  
   
2201 Replacement      25        0.163        1.224        0.08        0.306        0.312        0.097        0.428        0.184  
   
Low_CSD      17010        0.009        20.602        0        0.026        0.21        0.001        0.029        0.023  
   
Low_GAP and GAP Hybrid      6260        0.011        0.593        0        0.019        0.025        0.001        0.019        0.018  
   
Low_Upper Helen      1822        0.012        0.835        0        0.021        0.032        0.001        0.022        0.019  
   
Low_Upper Helen Wedge      1661        0.015        0.378        0        0.022        0.024        0.001        0.024        0.021  
   
Low_Lower Helen      828        0.007        0.285        0        0.016        0.025        0.001        0.018        0.014  
   
Low_Lower Helen Wedge      1280        0.011        1.347        0        0.026        0.06        0.001        0.029        0.022  

 

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Table 11-8 Silver Composite Statistics

 

Zone   

No.

 

Samples

     Median      Max      Min      Mean     

Std.

 

Dev.

     Variance     

Upper

 

95% CI

    

Lower

 

95% CI

 
   
CSD      521        0.852        48.673        0.000        2.396        4.391        19.281        2.773        2.019  
   
GAP      304        0.131        5.163        0.004        0.261        0.483        0.233        0.315        0.207  
   
Gap Hybrid      119        0.184        42.351        0.007        1.972        5.813        33.788        3.016        0.927  
   
Upper Helen      189        0.050        4.887        0.007        0.173        0.503        0.253        0.245        0.102  
   
Upper Helen Wedge      126        0.029        3.702        0.000        0.110        0.369        0.136        0.175        0.046  
   
Lower Helen      523        0.058        1.397        0.000        0.091        0.115        0.013        0.101        0.081  
   
Lower Helen Wedge      353        0.044        0.549        0.001        0.075        0.088        0.008        0.084        0.066  
   
2201 Vein      26        1.027        2.920        0.099        1.122        0.848        0.719        1.448        0.796  
   
2201 Replacement      25        0.653        3.987        0.102        1.043        1.142        1.304        1.49        0.595  
   
Low_CSD      17012        0.19        71.667        0        0.64        1.891        0.002        0.669        0.612  
   
Low_GAP and GAP Hybrid      6262        0.069        20.816        0        0.397        0.983        0.002        0.421        0.372  
   
Low_Upper Helen      1749        0.01        4.947        0        0.046        0.177        0.002        0.054        0.038  
   
Low_Upper Helen Wedge      1433        0.008        4.44        0        0.046        0.242        0.002        0.058        0.033  
   
Low_Lower Helen      803        0.007        0.802        0        0.02        0.04        0.002        0.023        0.017  
   
Low_Lower Helen Wedge      1233        0.005        1.031        0.001        0.016        0.057        0.002        0.019        0.012  

 

  11.6.

Grade Capping

Cap grades were applied to composites with values above a statistically determined threshold. Cap grade values were determined individually for each zone, set according to the upper 95% Cl listed in Table 11-7 and Table 11-8. For the estimation, composite values exceeding the cap grade were set to the cap grade. Of the composites within the 3 g/t grade shells 6.3% were capped. Grade capping details are listed in Table 11-9.

Table 11-9 Composite Grade Capping

 

                                                                                                                                                                                                              
Zone    Grade Cap    Composites Above
Cap
  

Number

 

of Comps.

   Capped %    Average Grade  Before
Capping
  

 

Au opt

  

 

Ag opt

  

 

Au

  

 

Ag

   Au    Ag    Au opt    Ag opt
                   
Lower Helen    1.07    0.29    14    15    168    8.3    8.9    1.61    0.49
                   
Lower Helen Wedge    0.96    0.26    18    23    394    4.6    5.8    1.41    0.46
                   
Upper Helen    0.65    0.49    11    10    191    5.8    5.2    0.98    1.48
                   
Upper Helen Wedge    0.36    0.38    5    7    141    3.5    5.0    0.52    1.20
                   
GAP Hybrid    0.48    14.93    6    7    133    4.5    5.3    1.05    23.23
                   
GAP    0.84    0.82    16    16    327    4.9    4.9    1.20    1.86
                   
CSD    0.43    8.84    138    199    1,975    7.0    10.1    1.61    7.30

 

 

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Zone    Grade Cap    Composites Above
Cap
  

Number

 

of Comps.

   Capped %    Average Grade  Before
Capping
  

 

Au opt

  

 

Ag opt

  

 

Au

  

 

Ag

   Au    Ag    Au opt    Ag opt
                   
2201 Vein    1.56    2.70    2    2    28    7.1    7.1    3.46    2.86
                   
2201 Replacement    0.87    3.61    2    2    25    8.0    8.0    1.05    3.82
                   
Low_GAP    0.09    1.86    305    1,251    10,050    3.0    12.4    0.25    5.98
                   
Low_Upper Helen    0.09    0.16    24    94    1,917    1.3    4.9    0.16    1.48
                   
Low_Lower Helen    0.09    0.06    8    26    728    1.1    3.6    0.15    0.13
                   
Low_Upper Helen Wedge    0.09    0.10    27    96    2,278    1.2    4.2    0.13    0.67
                   
Low_Lower Helen Wedge    0.09    0.05    29    25    802    3.6    3.1    0.22    0.17
                   
Low_CSD    0.09    2.56    1,333    3,745    33,785    3.9    11.1    0.37    7.30

 

  11.7.

Block Model

The block model origin was set at coordinate 1584315.0, 14647350.0, 3300.0 with bearing 45° to match the northwest trend of the deposit. The plunge and dip were both set to zero. The model extends 7,100 feet to the northwest, 2,400 feet to the northeast, and is 2,400 feet thick. The parent block size is 100 ft x 100 ft x 100 ft with a sub block size of 5 ft x 5 ft x 2.5 ft. The 2201 zone was assigned a sub block size of 1 ft x 1 ft x 1ft.

Variables were assigned to the model to contain gold and silver estimation values and other assigned values. The block model variables are listed in Table 11-10.

Table 11-10 Block Model Variables

 

         
Variables     Default           Type       Description
   
density       0    float    density
   
au_opt       -99    float    Gold - Grade Estimate (Ounces per Ton)
   
au_flag       0    byte    Gold - Estimation Flag
   
au_ndh       0    byte    Gold - Number Drill Holes
   
au_dist       0    float    Gold - Average Distance to Samples
   
au_ns       0    byte    Gold - Number of Samples
   
au_opt_nn       -99    float    Gold - Nearest Neighbor (Ounces per Ton)
   
au_nn_dist       0    float    Distance to nearest sample
   
ag_opt       -99    float    Silver - Grade Estimate (Ounces per Ton)
   
ag_flag       0    byte    Silver - Estimation Flag
   
ag_ndh       0    byte    Silver - Number Drill Holes
   
ag_dist         0    float    Silver - Average Distance to Samples

 

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Variables     Default           Type       Description
   
ag_ns       0    byte    Silver - Number of Samples
   
ag_opt_nn       -99    float    Silver - Nearest Neighbor (Ounces per Ton)
   
ag_nn_dist       0    float    Distance to nearest sample
   
aueq       -99    double    Gold Equivalence (Ounces per Ton)
   
agau       -99    double    Silver:Gold Ratio
   
mined     In situ       name    Block Status (in situ, sterile, mined)
   
classname     none       name    Classification (meas, ind, inf)
   
mii       0    byte    1 eq meas, 2 eq ind, 3 eq inf
   
aueng       0    float    Au Engineering
   
ageng       0    float    Ag Engineering
   
aueqeng       0    float    AuEq Engineering
   
zone     none       name    Zone
   
volume    -       predefined     
   
xlength    -       predefined     
   
ylength    -       predefined     
   
zlength    -       predefined     
   
xcentre    -       predefined     
   
ycentre    -       predefined     
   
zcentre    -       predefined     
   
xworld    -       predefined     
   
yworld    -       predefined     
   
zworld    -         predefined     

 

 

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  11.8.

Grade Estimation and Resource Classification

The gold and silver variables in the block model were estimated using inverse distance cubed (ID3) and nearest neighbor methods. The estimations were completed with one pass.

Anisotropic search parameters were set to the average orientation of each zone. Average orientation was determined by loading the modeled 3 g/t lenses by zone in Vulcan and visually estimating average dip and dip direction. Distances were selected based on the drill spacing of samples intercepting the lenses and on the general orientation and shape of the interpreted solids. Blocks inside of the modelled 3 g/t lenses were estimated using only composites flagged with the corresponding lens code. Blocks outside of the 3 g/t lenses were estimated using composites with the corresponding low-grade flag. Blocks lying outside the low-grade halo were not estimated. The estimation search parameters are listed in Table 11-11.

 

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Table 11-11 Estimation Parameters

 

Zone   

Mineral

Lens

Code

   Bearing    Plunge      Dip     

Major 

Search 

Axis (ft) 

    

Semi- 

major 

Search 

Axis (ft) 

    

Minor 

Search 

Axis (ft) 

     Min
Samp
     Max
Samp
   
CSD_GAP    220X    315      4.8        10.3        300        300        100        1      3
   
GAP Hybrid    500X    315      5.0        0.0        300        300        100        1      3
   
CSD    110X    315      6.3        7.1        300        300        100        1      3
   
Upper Helen    310X    315      10.7        7.0        300        300        100        1      3
   
Upper Helen Wedge    330X    315      6.2        -7.4        300        300        100        1      3
   
Lower Helen    320X    315      0.0        0.0        300        300        100        1      3
   
Lower Helen Wedge    340X    315      -1.5        0.0        300        300        100        1      3
   
2201 high angle vein 1    1301    342      0.0        73.6        300        300        300        1      3
   
2201 high angle vein 2    1302    342      0.0        73.6        300        300        300        1      3
   
2201 high angle vein 3    1303    322      0.0        73.6        300        300        300        1      3
   
2201 low angle    140X    315      7.6        15.7        300        300        100        1      3
   
Low_CSD_GAP and Gap Hybrid    22000    315      4.8        10.3        300        300        100        1      3
   
Low_CSD    11000    315      6.3        7.1        300        300        100        1      3
   
Low_Upper Helen    31000    315      10.7        7.0        300        300        100        1      3
   
Low_Upper Helen Wedge    33000    315      6.2        -7.4        300        300        100        1      3
   
Low_Lower Helen    32000    315      0.0        0.0        300        300        100        1      3
   
Low_Lower Helen Wedge    34000    315      -1.5        0.0        300        300        100        1      3
   
Low_NE of CR fault    10000    315      0.0        0.0        300        300        100        1      3
   
Low_Other_1100X    1100X    315      6.3        7.1        300        300        100        1      3
   
Low_Other_2200X    2200X    315      4.8        10.3        300        300        100        1      3
   
Low_Other_3300X    3300X    315      6.2        -7.4        300        300        100        1      3
   
Low_Other_3400X    3400X    315      -1.5        0        300        300        100        1      3

A script was run on the estimated block model to populate the classification variable. The classification categories are indicated, inferred and none. Classification was determined based on three block model variables: au_dist, au_ndh and au_nn_dist. These three variables represent, respectively, the average distance to the composites used to estimate the grade of the block, the number of drill holes contributing to the grade of the block, and the distance to the nearest composite. The default value was defined as ‘none’, which was over-written by indicated or inferred where the required conditions were satisfied. The conditions of the classification script are listed in Table 11-12.

Table 11-12 Classification Conditions

 

         
Class   

Script

Condition

  

au_dist

(ft)

   au_ndh   

au_nn_dist

(ft)

   

Indicated

  

if

  

<100

   at least 2    50 or less
   

Inferred

  

elseif

  

<=300

   at least 2     

 

 

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Class   

Script

Condition

  

au_dist

(ft)

   au_ndh   

au_nn_dist

(ft)

   

None

  

default

              

Significant parameters used in the gold and silver estimations included:

 

  1.

Only composites with a value of greater than or equal to zero were used;

  2.

A minimum of one and maximum of three composites were used;

  3.

Only one composite per drill hole was allowed;

  4.

Composites were selected using anisotropic distances oriented to the local dip and dip direction of the zone;

  5.

Only composites within a lens were used to estimate blocks within the lens;

  6.

Grades were capped using a top cut method; and

  7.

Gold and silver for blocks outside modelled 3 g/t and low-grade shells were not estimated.

 

  11.9.

Mined Depletion and Sterilization

The CSD zone lies adjacent to the historic Cove pit and was historically exploited using underground cut-and-fill and stoping methods. Part of the modelled CSD zone has been mined, and areas of in situ material near historically mined areas are rendered inaccessible. The block model includes a ‘mined’ variable which stores information identifying each block as in situ, mined or sterile. The default value is in situ. Blocks lying above the ultimate pit topography or inside the underground mine as built are defined as mined. Sterile blocks were defined using two shapes modelled in Vulcan. The first is a surface digitized 50-feet below deepest mined topography, and the second is a solid shape digitized around the underground mine as-built representing a 30-foot buffer zone. Blocks lying between the 50-foot surface and the ultimate pit topographic surface are sterile, and blocks within 30 feet of the historic underground mine are sterile. Figure 11-7 shows the sterilization surfaces in a section view of the CSD zone.

 

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Mineral Resource Estimates

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Figure 11-7 Sterilization Surfaces

 

LOGO

 

   11.10.

Model Validation

 

  11.10.1.

Estimation Comparison

The mean gold grades for each lens were compared against a nearest neighbor (representing declustered composites) in Table 11-13. Individual lens comparisons vary depending on sample support and grade variability. Overall, the ID3 estimate is slightly lower than the nearest neighbor estimate. Table 11-14 represents the same data for silver which shows the same general relationships.

Table 11-13 Estimate Comparison for Gold versus a Nearest Neighbor at 0 Cutoff

 

       
     ID3 Estimate     Nearest Neighbor     Mean  
Vein   Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Diff  
     
1101     0.186       0.056       0.284       0.245       0.136       0.089       0.193       0.077       0.27       0.27       0.102       0.051       -3.6
     
1104     0.143       0.076       0.416       0.188       0.089       0.01       0.148       0.102       0.416       0.216       0.081       0.007       -3.4
     
1105     0.172       0.079       0.43       0.218       0.12       0.002       0.172       0.105       0.43       0.245       0.105       0.002       0.0
     
1106     0.147       0.085       0.43       0.188       0.092       0.003       0.15       0.111       0.43       0.194       0.09       0.002       -2.0
     
1107     0.184       0.09       0.43       0.235       0.119       0.01       0.186       0.117       0.43       0.274       0.096       0.01       -1.1
     
1108     0.147       0.05       0.43       0.151       0.116       0.028       0.15       0.061       0.43       0.16       0.114       0.022       -2.0
     
1109     0.177       0.057       0.428       0.222       0.129       0.027       0.181       0.071       0.43       0.237       0.126       0.026       -2.2
     
1112     0.109       0.031       0.429       0.11       0.09       0.089       0.108       0.041       0.43       0.1       0.089       0.089       0.9

 

 

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     ID3 Estimate     Nearest Neighbor     Mean  
Vein   Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Diff  
     
1113     0.136       0.092       0.39       0.186       0.06       0.011       0.149       0.136       0.391       0.34       0.031       0       -8.7
     
1115     0.115       0.101       0.43       0.166       0.037       0       0.116       0.126       0.43       0.164       0.018       0       -0.9
     
1116     0.134       0.068       0.43       0.164       0.092       0.006       0.129       0.088       0.43       0.161       0.073       0.006       3.9
     
1117     0.174       0.095       0.43       0.266       0.1       0.009       0.148       0.116       0.43       0.17       0.074       0.009       17.6
     
1118     0.22       0.099       0.424       0.281       0.14       0.035       0.225       0.138       0.424       0.35       0.113       0.034       -2.2
     
1120     0.106       0.094       0.43       0.14       0.034       0.002       0.108       0.106       0.43       0.118       0.037       0       -1.9
     
2203     0.146       0.076       0.347       0.151       0.099       0.097       0.145       0.09       0.347       0.138       0.097       0.097       0.7
     
2204     0.263       0.166       0.773       0.305       0.172       0.068       0.259       0.204       0.773       0.211       0.146       0.068       1.5
     
2204     0.304       0.157       0.84       0.361       0.2       0.022       0.29       0.205       0.84       0.376       0.139       0.012       4.8
     
2205     0.244       0.12       0.609       0.338       0.147       0.101       0.24       0.139       0.609       0.376       0.13       0.101       1.7
     
2206     0.263       0.192       0.84       0.267       0.156       0.039       0.272       0.221       0.84       0.273       0.145       0.039       -3.3
     
2207     0.342       0.215       0.84       0.479       0.16       0.031       0.339       0.266       0.84       0.538       0.127       0.031       0.9
     
2208     0.282       0.173       0.84       0.375       0.147       0.006       0.286       0.216       0.84       0.423       0.115       0.006       -1.4
     
2209     0.199       0.073       0.363       0.244       0.142       0.119       0.199       0.089       0.363       0.25       0.137       0.119       0.0
     
3101     0.257       0.089       0.648       0.278       0.227       0.094       0.253       0.113       0.65       0.281       0.197       0.094       1.6
     
3102     0.177       0.073       0.318       0.261       0.125       0.09       0.181       0.081       0.318       0.258       0.125       0.09       -2.2
     
3103     0.192       0.135       0.65       0.216       0.109       0.002       0.192       0.169       0.65       0.201       0.097       0.002       0.0
     
3104     0.167       0.077       0.449       0.178       0.13       0.003       0.168       0.098       0.449       0.195       0.104       0.003       -0.6
     
3105     0.407       0.172       0.65       0.561       0.284       0.106       0.406       0.236       0.65       0.65       0.222       0.106       0.2
     
3106     0.159       0.061       0.374       0.192       0.115       0.046       0.158       0.072       0.374       0.188       0.112       0.045       0.6
     
3107     0.128       0.069       0.319       0.136       0.09       0.023       0.132       0.103       0.319       0.124       0.088       0.023       -3.0
     
3202     0.254       0.124       0.482       0.365       0.133       0.083       0.266       0.158       0.484       0.399       0.082       0.082       -4.5
     
3202     0.256       0.133       0.734       0.329       0.159       0.082       0.243       0.155       0.735       0.248       0.114       0.082       5.3
     
3203     0.405       0.151       1.069       0.464       0.329       0.11       0.416       0.223       1.07       0.458       0.299       0.11       -2.6
     
3204     0.256       0.033       0.336       0.273       0.229       0.197       0.249       0.05       0.336       0.272       0.222       0.197       2.8
     
3205     0.6       0.344       1.07       0.954       0.257       0.118       0.625       0.445       1.07       1.07       0.138       0.118       -4.0
     
3206     0.418       0.28       1.07       0.608       0.168       0.004       0.413       0.358       1.07       0.649       0.116       0.004       1.2
     
3207     0.271       0.177       0.612       0.405       0.128       0.114       0.285       0.204       0.612       0.612       0.116       0.114       -4.9
     
3208     0.159       0.038       0.256       0.17       0.139       0.1       0.163       0.054       0.256       0.177       0.101       0.1       -2.5
     
3210     0.157       0.043       0.209       0.203       0.109       0.1       0.156       0.055       0.209       0.209       0.1       0.1       0.6
     
3211     0.16       0.065       0.5       0.178       0.123       0.073       0.157       0.077       0.5       0.197       0.112       0.073       1.9
     
3301     0.14       0.057       0.36       0.17       0.104       0.03       0.134       0.079       0.36       0.155       0.093       0.03       4.5
     
3302     0.102       0.069       0.298       0.167       0.018       0.006       0.111       0.091       0.305       0.138       0.018       0.006       -8.1
     
3303     0.125       0.054       0.34       0.126       0.099       0.031       0.121       0.064       0.34       0.133       0.087       0.025       3.3
     
3304     0.169       0.061       0.36       0.216       0.127       0.022       0.169       0.089       0.36       0.236       0.114       0.001       0.0
     
3305     0.142       0.066       0.36       0.179       0.096       0.024       0.143       0.081       0.36       0.202       0.091       0.024       -0.7
     
3306     0.129       0.022       0.174       0.146       0.118       0.044       0.13       0.028       0.177       0.158       0.114       0.043       -0.8
     
3400     0.224       0.171       0.727       0.283       0.112       0.017       0.239       0.22       0.729       0.322       0.109       0.017       -6.3
     
3401     0.214       0.108       0.658       0.273       0.143       0.018       0.21       0.13       0.659       0.272       0.118       0.018       1.9

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp  

Mineral Resource Estimates

  Page 141 

 

       
     ID3 Estimate     Nearest Neighbor     Mean  
Vein   Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Diff  
     
3402     0.313       0.236       0.96       0.512       0.134       0       0.322       0.272       0.96       0.574       0.106       0       -2.8
     
3403     0.277       0.173       0.96       0.329       0.162       0.001       0.28       0.217       0.96       0.412       0.149       0.001       -1.1
     
3404     0.365       0.237       0.96       0.461       0.204       0.098       0.364       0.291       0.96       0.416       0.143       0.092       0.3
     
3405     0.182       0.082       0.367       0.25       0.115       0.114       0.184       0.084       0.367       0.255       0.115       0.114       -1.1
     
3406     0.36       0.205       0.96       0.482       0.222       0.116       0.363       0.246       0.96       0.495       0.22       0.116       -0.8
     
3407     0.267       0.17       0.96       0.378       0.136       0.056       0.273       0.226       0.96       0.42       0.104       0.056       -2.2
     
3408     0.146       0.048       0.224       0.177       0.113       0.008       0.141       0.052       0.224       0.167       0.099       0.008       3.5
     
3409     0.112       0.021       0.158       0.125       0.107       0.057       0.112       0.028       0.166       0.129       0.114       0.057       0.0
     
5001     0.3       0.068       0.407       0.322       0.239       0.202       0.302       0.078       0.407       0.322       0.202       0.202       -0.7
     
5002     0.165       0.092       0.48       0.193       0.109       0.055       0.167       0.115       0.48       0.209       0.108       0.055       -1.2
     
5003     0.179       0.04       0.413       0.191       0.153       0.114       0.177       0.055       0.415       0.179       0.153       0.104       1.1
     
5004     0.14       0.067       0.4       0.179       0.097       0.041       0.145       0.089       0.4       0.181       0.097       0.041       -3.4
     
5005     0.202       0.144       0.48       0.202       0.107       0.093       0.195       0.154       0.48       0.217       0.098       0.092       3.6
     
5006     0.186       0.096       0.48       0.201       0.119       0.091       0.18       0.128       0.48       0.159       0.094       0.091       3.3
     
5007     0.165       0.095       0.465       0.207       0.104       0.022       0.166       0.118       0.466       0.206       0.099       0.02       -0.6
     
5008     0.18       0.085       0.479       0.21       0.113       0.078       0.184       0.112       0.48       0.201       0.111       0.077       -2.2
     
5009     0.108       0.01       0.133       0.114       0.101       0.095       0.107       0.013       0.133       0.103       0.1       0.095       0.9
     
5010     0.17       0.079       0.447       0.196       0.123       0.068       0.167       0.107       0.448       0.153       0.11       0.068       1.8
     
5011     0.299       0.035       0.348       0.321       0.286       0.148       0.293       0.073       0.382       0.382       0.245       0.075       2.0

Table 11-14 Estimate Comparison for Silver versus a Nearest Neighbor at 0 Cutoff

 

       
     ID3 Estimate     Nearest Neighbor     Mean  
Vein   Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Mean    

Std.

Dev.

    Max     Q3     Q1     Min     Diff  
     
1101     1.759       1.219       5.8       2.449       0.854       0.07       1.422       1.628       8.165       2.416       0.07       0.07       23.7
     
1104     3.145       2.067       8.84       4.673       1.413       0       2.988       2.595       8.84       4.832       0.832       0       5.3
     
1105     1.21       1.643       8.84       1.676       0.05       0       1.213       2.028       8.84       1.521       0.04       0       -0.2
     
1106     2.329       2.498       8.84       4.026       0.216       0       2.42       2.891       8.84       3.653       0.245       0       -3.8
     
1107     0.33       0.289       1.437       0.537       0.056       0.004       0.335       0.34       1.437       0.5       0.047       0.004       -1.5
     
1108     2.817       1.191       6.683       3.566       1.991       0.261       2.968       1.558       6.683       3.652       1.657       0.214       -5.1
     
1109     2.871       1.698       8.653       4.016       1.282       0.065       2.831       1.922       8.653       3.907       1.15       0.06       1.4
     
1112     1.627       2.105       8.659       1.799       0.445       0.12       1.74       3.026       8.662       0.455       0.455       0.12       -6.5
     
1113     2.098       1.383       5.621       2.726       0.968       0.203       1.807       1.874       5.63       1.801       0.65       0.2       16.1
     
1115     1.882       2.733       8.84       1.782       0.252       0       1.852       2.921       8.84       1.651       0.166       0       1.6
     
1116     0.905       1.61       8.834       0.742       0.094       0       0.887       1.806       8.84       0.39       0.071       0       2.0
     
1117     0.949       0.975       8.84       0.824       0.484       0.152       0.807       1.075       8.84       0.961       0.311       0.151       17.6
     
1118     1.731       1.98       8.453       1.635       0.638       0.098       1.744       2.754       8.453       1.1       0.3       0       -0.7
     
1120     1.732       1.696       8.681       2.328       0.625       0.01       1.705       2.401       8.84       2.395       0.14       0.01       1.6
     
2203     0.051       0.022       0.087       0.07       0.041       0.007       0.053       0.025       0.087       0.087       0.044       0.007       -3.8

 

 

 Practical Mining LLC   March 26, 2025 


 Page 142  

S-K1300 Initial Assessment & Technical Report

Summary for the Cove Project, Lander County,

Nevada

  i-80 Gold Corp 

 

       
      ID3 Estimate      Nearest Neighbor      Mean
Vein    Mean      Std.
Dev.
     Max      Q3      Q1      Min      Mean      Std.
Dev.
     Max      Q3      Q1      Min      Diff
       
2204       0.264        0.194        0.805        0.336        0.114        0.012        0.262        0.215        0.805        0.332        0.107        0.012      0.8%
       
2204      0.108        0.082        0.369        0.12        0.056        0.009        0.104        0.108        0.369        0.106        0.04        0.007      3.8%
       
2205      0.236        0.217        0.82        0.347        0.067        0.016        0.245        0.254        0.82        0.305        0.04        0.016      -3.7%
       
2206      0.126        0.089        0.506        0.152        0.069        0.005        0.128        0.101        0.506        0.168        0.059        0.005      -1.6%
       
2207      0.236        0.186        0.82        0.294        0.107        0.016        0.237        0.222        0.82        0.317        0.076        0.016      -0.4%
       
2208      0.214        0.171        0.82        0.263        0.094        0.007        0.217        0.211        0.82        0.298        0.062        0.007      -1.4%
       
2209      0.49        0.2        0.82        0.688        0.365        0.075        0.503        0.269        0.82        0.82        0.356        0.075      -2.6%
       
3101      0.062        0.031        0.197        0.094        0.044        0.02        0.061        0.038        0.207        0.097        0.04        0.02      1.6%
       
3102      0.08        0.035        0.176        0.113        0.046        0.02        0.078        0.044        0.176        0.115        0.04        0.02      2.6%
       
3103      0.111        0.107        0.49        0.159        0.029        0.007        0.112        0.139        0.49        0.159        0.02        0.007      -0.9%
       
3104      0.08        0.089        0.49        0.086        0.03        0.007        0.081        0.108        0.49        0.096        0.026        0.007      -1.2%
       
3105      0.22        0.115        0.417        0.316        0.111        0.047        0.222        0.158        0.417        0.361        0.055        0.047      -0.9%
       
3106      0.157        0.111        0.49        0.211        0.072        0.007        0.15        0.14        0.49        0.222        0.035        0.007      4.7%
       
3107      0.216        0.195        0.49        0.439        0.016        0.007        0.205        0.234        0.49        0.49        0.009        0.007      5.4%
       
3202      0.017        0.02        0.143        0.021        0.007        0.001        0.006        0.034        0.265        0.001        0        0      183.3%
       
3202      0.028        0.023        0.106        0.039        0.012        0        0.028        0.028        0.106        0.058        0.008        0      0.0%
       
3203      0.094        0.07        0.29        0.138        0.033        0.012        0.091        0.093        0.29        0.1        0.012        0.012      3.3%
       
3204      0.088        0.027        0.142        0.112        0.063        0.047        0.084        0.043        0.16        0.125        0.047        0.047      4.8%
       
3205      0.126        0.071        0.29        0.153        0.071        0.044        0.128        0.091        0.29        0.123        0.05        0.044      -1.6%
       
3206      0.144        0.068        0.29        0.192        0.09        0.007        0.143        0.088        0.29        0.23        0.071        0.007      0.7%
       
3207      0.121        0.068        0.251        0.166        0.076        0.015        0.129        0.081        0.251        0.251        0.093        0.015      -6.2%
       
3208      0.085        0.05        0.213        0.094        0.055        0.035        0.086        0.061        0.213        0.09        0.058        0.035      -1.2%
       
3210      0.02        0.016        0.041        0.038        0.002        0        0.02        0.021        0.041        0.041        0        0      0.0%
       
3211      0.046        0.013        0.086        0.052        0.038        0.025        0.046        0.015        0.086        0.05        0.041        0.025      0.0%
       
3301      0.046        0.021        0.122        0.063        0.026        0        0.046        0.028        0.122        0.063        0.017        0      0.0%
       
3302      0.025        0.015        0.048        0.041        0.01        0        0.025        0.018        0.049        0.043        0.007        0      0.0%
       
3303      0.096        0.109        0.38        0.104        0.025        0.006        0.094        0.12        0.38        0.093        0.014        0.006      2.1%
       
3304      0.072        0.078        0.38        0.128        0.007        0        0.076        0.099        0.38        0.152        0.001        0      -5.3%
       
3305      0.053        0.05        0.38        0.069        0.018        0        0.054        0.065        0.38        0.077        0.009        0      -1.9%
       
3306      0.108        0.141        0.38        0.251        0.005        0.001        0.111        0.152        0.38        0.306        0.001        0.001      -2.7%
       
3400      0.023        0.024        0.129        0.037        0.003        0        0.025        0.03        0.129        0.044        0.002        0      -8.0%
       
3401      0.052        0.049        0.245        0.07        0.012        0.001        0.05        0.056        0.245        0.064        0.007        0.001      4.0%
       
3402      0.07        0.069        0.26        0.134        0.009        0        0.075        0.081        0.26        0.152        0.007        0      -6.7%
       
3403      0.056        0.036        0.26        0.061        0.034        0.001        0.054        0.044        0.26        0.061        0.023        0.001      3.7%
       
3404      0.109        0.063        0.26        0.135        0.071        0.001        0.109        0.08        0.26        0.136        0.052        0.001      0.0%
       
3405      0.046        0.005        0.057        0.047        0.04        0.04        0.046        0.005        0.057        0.05        0.04        0.04      0.0%
       
3406      0.074        0.051        0.26        0.087        0.045        0.003        0.074        0.058        0.26        0.082        0.053        0.003      0.0%
       
3407      0.084        0.063        0.26        0.125        0.035        0.009        0.087        0.083        0.26        0.143        0.02        0.009      -3.4%
       
3408      0.044        0.024        0.102        0.053        0.021        0.007        0.042        0.028        0.102        0.044        0.017        0.007      4.8%

 

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      ID3 Estimate      Nearest Neighbor      Mean
Vein    Mean      Std.
Dev.
     Max      Q3      Q1      Min      Mean      Std.
Dev.
     Max      Q3      Q1      Min      Diff
       
3409       0.039        0.037        0.22        0.037        0.021        0.009        0.04        0.053        0.26        0.035        0.015        0.009      -2.5%
       
5001      0.335        0.204        0.639        0.543        0.133        0.077        0.339        0.267        0.639        0.639        0.121        0.077      -1.2%
       
5002      0.671        0.776        3.091        0.732        0.167        0.064        0.67        0.96        3.091        0.612        0.153        0.059      0.1%
       
5003      1.721        2.195        7.508        2.337        0.091        0.055        1.812        2.654        7.508        1.491        0.088        0.038      -5.0%
       
5004      1.671        3.361        14.597        1.464        0.153        0.044        1.73        4.021        14.93        0.449        0.125        0.043      -3.4%
       
5005      1.595        3.085        14.929        0.842        0.076        0.016        1.256        3.774        14.93        0.292        0.029        0.016      27.0%
       
5006      0.425        0.738        4.802        0.427        0.07        0.038        0.377        0.916        4.87        0.56        0.062        0.038      12.7%
       
5007      1.023        0.586        3.419        1.465        0.544        0.081        0.976        0.962        3.439        1.123        0.181        0.08      4.8%
       
5008      1.594        3.097        14.93        0.838        0.109        0.053        1.462        3.343        14.93        0.72        0.108        0.053      9.0%
       
5009      0.088        0.065        0.237        0.144        0.024        0.008        0.093        0.081        0.237        0.174        0.013        0.007      -5.4%
       
5010      0.318        0.322        1.576        0.329        0.137        0.031        0.317        0.462        1.578        0.226        0.12        0.031      0.3%
       
5011      1.979        0.473        2.945        2.369        1.596        1.003        2.092        1.646        3.501        3.501        0.174        0.164      -5.4%

 

  11.10.2.

Visual Comparison

On a local scale, model validation can be confirmed by the visual comparison of block grades to composite grades. Figure 11-8 through Figure 11-11 show typical cross sections where the block and drill color schemes are identical.

Figure 11-8 Comparison of Composite and Estimated Block Gold Grades, Helen Zone

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Figure 11-9 Comparison of Composite and Estimated Block Gold Grades, Gap Zone

 

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Figure 11-10 Comparison of Composite and Estimated Block Gold Grades, CSD Zone

 

LOGO

Figure 11-11 Comparison of Composite and Estimated Block Gold Grades, 2201 Zone

 

LOGO

 

 

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  11.10.3.

Swath Plots

Further spatial model validation is provided by swath plots of individual lenses. Swath plots for a typical lens from each zone are presented in Figure 11-12 through Figure 11-19. These plots compare the average grade from the estimation to the NN from within regularly spaced swaths or slices through the lens in three dimensions (along strike, along width and vertically). Examination of the swath plots shows good agreement among the gold and silver estimation values.

Figure 11-12 Gold Swath Plots of Helen Zone 3103

 

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Figure 11-13 Silver Swath Plots of Helen Zone 3103

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Figure 11-14 Gold Swath Plots of Gap Zone 2208

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Figure 11-15 Silver Swath Plots of Gap Zone 2208

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Figure 11-16 Gold Swath Plots of CSD Zone 1106

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Figure 11-17 Silver Swath Plots of CSD Zone 1106

LOGO

Figure 11-18 Gold Swath Plots of 2201 Zone 1302

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Figure 11-19 Silver Swath Plots of 2201 Zone 1302

LOGO

 

  11.10.4.

Model Smoothing Checks – Grade Tonnage Curves

A final model validation check can be made by examining the grade tonnage distribution for the estimation, which is illustrated in Figure 11-20 through Figure 11-23. The grade tonnage curve is used to describe the tons and grade that may be present above a cutoff for mining. Smoothing in the estimate, the spacing of the informing samples, and the continuity of grades within the vein all affect the shape of the estimated grade tonnage curve. Above a 0.1 opt gold cut-off grade the curve shows gradually increasing grade and decreasing tonnage as the cut-off grade is increased.

 

 

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Figure 11-20 Helen Zone Grade Tonnage Plots

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Figure 11-21 Gap Zone Grade Tonnage Plots

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Figure 11-22 CSD Zone Grade Tonnage Plots

LOGO

Figure 11-23 2201 Zone Grade Tonnage Plots

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  11.11.

Factors That May Affect Mineral Resources

Areas of uncertainty that may materially impact the Mineral Resource Estimates include:

 

 

Changes to long term metal price assumptions.

 

 

Changes to the input values for mining, processing, and G&A costs to constrain the estimate.

 

 

Changes to local interpretations of mineralization geometry and continuity of mineralized Domains.

 

 

Changes to the density values applied to the mineralized zones.

 

 

Changes to metallurgical recovery assumptions.

 

 

Variations in geotechnical, hydrogeological and mining assumptions.

 

 

Changes to assumptions with an existing agreement or new agreements.

 

 

Changes to environmental, permitting, and social license assumptions.

 

 

Logistics of securing and moving adequate services, labor, and supplies could be affected by epidemics, pandemics and other public health crises.

 

  11.12.

Reasonable Prospects for Economic Extraction

S-K 1300 requires mineral resources demonstrate “Reasonable Prospects for Economic Extraction” (RPEE). Stope optimizer software is well suited to meet this requirement. The software will produce stope designs that meet minimum minable geometric shapes that exceed the cutoff grade. These shapes will include necessary low grade or waste dilution included with the stope design.

Cove mineral resources are defined by a mining geometry consistent with the drift and fill or drift and bench mining methods chosen. The dimensions of a minimum minable stope cross section are 20 feet wide x 15 feet high. Individual stope lengths can vary from a minimum of 20 feet to a maximum of 100 feet.

 

  11.12.1.

QP Opinion

Practical Mining is not aware of any environmental, legal, title, taxation, socioeconomic, marketing, political, or other relevant factors that would materially affect the estimation of Mineral Resources that are not discussed in this Technical Report.

Practical Mining is of the opinion that the Mineral Resources for the Project, which were estimated using industry accepted practices, have been prepared and reported using S-K 1300 definitions.

 

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Technical and economic parameters and assumptions applied to the Mineral Resource Estimate are based on parameters received from i-80 and reviewed by Practical Mining to determine if they were appropriate.

The QP considers that all issues relating to all relevant technical and economic factors likely to influence the prospect of economic extraction can be resolved with further work.

 

  11.13.

Mineral Resources

Cove mineral resources in each zone are tabulated in Table 11-15. The cutoff grade for mineral resources varies by process type and estimated recovery. It is anticipated that total gold production will be limited by mine capacity and cut off grades will range from a low of 0.112 Au opt to a high of 0.155 Au opt. Details of Cutoff grade estimation are given in section 18.4.

Table 11-15 Summary of Cove Mineral Resources at the End of the Fiscal Year Ended December 31, 2024

      Tons
(000)
    

Tonnes

(000)

    

Au

(opt)

    

Au

g/t

    

Ag

(opt)

    

Ag

(g/t)

    

Au ozs

(000)

    

Ag ozs

(000)

 
Indicated Mineral Resource
   

Helen

   743      674      0.271      9.3      0.074      2.6      201      55
   

Gap

   280      254      0.219      7.5      0.239      8.9      61      72
   

CSD

   275      249      0.175      6.0      1.603      55.0      48      441
   
Total Indicated    1,298      1,177      0.239      8.2      0.438      15.0      310      568
 
Inferred Mineral Resource
   

Helen

   1,743      1,582      0.245      8.4      0.083      2.9      427      146
   

Gap

   2,229      2,022      0.244      8.4      0.262      9.0      543      585
   

CSD

   319      290      0.173      5.9      1.685      57.8      55      538
   

2201

   168      153      0.780      26.7      1.016      34.8      131      171
   
Total Inferred    4,459      4,047      0.259      8.9      0.323      11.1      1,156      1,439

Notes:

  1.

Mineral resources have been estimated at a gold price of $2,175 per troy ounce and a silver price of 27.25 per troy ounce. Refer to Section 16.1 for a discussion on metal pricing;

  2.

Mineral resources have been estimated using gold metallurgical recoveries ranging from 73.2% to 93.3% for roasting and 78.5% to 95.1 % for pressure oxidation;

  3.

Roaster cutoff grades range from 4.15 to 5.29 Au g/t (0.121 to 0.154 opt) and pressure oxidation cutoff grades range from 3.83 to 4.64 Au g/t (0.112 to 0.135 opt);

  4.

The effective date of the mineral resource estimate is December 31, 2024;

  5.

Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant factors;

  6.

An inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity. An inferred mineral resource has a

 

 

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lower level of confidence than that applying to an indicated mineral resource and must not be converted to a mineral reserve. It is reasonably expected that the majority of inferred mineral resources could be upgraded to indicated mineral resources with continued exploration; and

  7.

The reference point for mineral resources is in situ.

12. Mineral Reserve Estimates

The Cove Project does not have any Mineral Reserves.

 

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13. Mining Methods

 

  13.1.

Mine Development

 

  13.1.1.

Access Development

Underground access to the mining areas will begin with a portal on the North side of the existing pit and ramp down. Initial work will consist of 4,557 feet of decline from the portal down to approximately the 4600-foot elevation and 889 feet of drill laterals. The drill laterals are located directly above the Helen and Gap deposits. (Figure 13-1)

Figure 13-1 Exploration Development

LOGO

The decline will serve as a starting point for subsequent development and a portion of the drill cross cuts will later serve as the part of the main ventilation intake. Primary access drifts are designed 15 feet wide and 17.5 feet high to permit 30-ton haulage trucks and provide a large cross section for ventilation. Drift gradients will vary from – 15% to + 15% to reach the desired elevation. Secondary drifts, spiral ramps and vertical raises will connect the haulage drifts to provide a pathway for ventilation to the surface and serve as a secondary escape way (Figure 13-2).

 

 

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Figure 13-2 Plan view showing portal, main haulages, and two raises to surface

LOGO

 

  13.1.2.

Ground Support

The ground conditions at the Project are typical of the northern Nevada extensional tectonic environment. Joint spacing varies from a few inches to a foot or more. It is expected that Swellex rock bolts along with welded wire mesh will be able to control all conditions encountered during decline development and stoping. Shotcrete will also be liberally applied as needed to prevent long-term deterioration of the rock mass. Under more extreme conditions, resin anchor bolts, or cable bolts can be used to supplement the primary support. Steel sets and spiling may also be used to support areas with the most severe ground conditions.

Project geologists have recorded core recovery and Rock Quality Designation (RQD) as part of their normal core logging process. Figure 13-3 summarizes RQD for each formation in the mining horizon. RQD values from 30% to the low 40% range are typical for mines in the area. RQD values are also dependent on drill orientation relative to the major joint sets and can vary widely.

The Modified Rock Mass Rating system proposed by Jakubec and Laubscher (2000) provides for additional characteristics to be considered in addition to RQD. These include filling material, joint waviness, alteration, weathering and the presence of water. A selection of core holes in the resource delineation program should be logged with the MRMR system to allow comprehensive classification of the rock mass.

 

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Figure 13-3 Formation RQD

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Joint set orientation relative to the mine opening geometry is the most significant factor in opening stability in north-east Nevada. In conjunction with the resource delineation program, Acoustic Tele Viewer logging should be obtained to determine joint orientation for each domain to optimize mine opening orientation and estimate support requirements.

 

  13.1.1.

Ventilation and Secondary Egress

Mechanized underground mining relies heavily on diesel equipment to extract the mineralized material and waste rock and to transport backfill to the stopes. Diesel combustion emissions will require substantial amounts of fresh ventilation air to remove the diesel exhaust and maintain a healthy working environment. A combination of the main access drifts and vertical raises to the surface are arranged in a manner to provide a complete ventilation circuit capable of supplying the mine with 500,000 cubic feet per minute (CFM) of fresh air. Air movement is facilitated by primary ventilation fans placed at the surface and underground in strategic locations. Small auxiliary fans and ducting will draw primary ventilation air directly into the working faces.

Figure 13-4 Ventilation Schematic

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Secondary egress will be provided by installing a personnel hoist with a capsule capable of holding up to four people. The hoist will be located at the surface of the exhasut ventilation raise.

 

  13.1.2.

Dewatering

The dewatering wells will provide the majority of mine dewatering. Small localized inflows will be captured at sumps located strategically throughout the mine and pumped to the surface where it will be commingled with the water from the dewatering wells.

 

  13.2.

Mining Methods

Due to the mostly flat geometry of the ore lenses, all planned production mining will be completed using drift and fill mining. The final choice of mining method will depend upon the geometry of the stope block, proximity to main access ramps, ventilation and escape routes, the relative strength or weakness of the mineralized material and adjacent wall rock, and finally the value or grade of the mineralized material. The choice of mining method will not be finalized until after the stope delineation and definition drilling is completed. The drift and fill method is discussed briefly in the following paragraphs.

 

  13.2.1.

Drift and Fill

Drift and Fill is a very selective mining method. A drift and fill stope is initiated by driving a waste crosscut from the access ramp to the ore. The initial ore drift is driven at planned 13-feet wide by 13-feet high dimensions, with gradient varying between +/-20% to follow the geometry of the mineralization. The minimum cut and fill drift height is eight feet to minimize dilution on the thinner mineralized lenses. Once the initial drift is driven, floor may be pulled and/or back may be breasted down to capture the full thickness of the lens. Where mining is planned adjacent to the drift, it will be backfilled with CRF prior to mining the subsequent drifts. (Figure 13-5)

 

 

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Figure 13-5 Depiction of Drift and Fill method

LOGO

 

  13.3.

Underground Labor

Approximately 5,270 feet of development will be undertaken in 2022 and 2023 to provide access for underground delineation and exploration drilling. Underground workforce requirements for this early development phase of the Project are estimated in Table 13-1. Following a positive production decision in 2024, production will increase and peak underground workforce requirements for the Project are presented in Table 13-2. This estimate was prepared using productivity rates typical for large-scale mechanized mining in North America. The Project will operate 24 hours per day seven days per week. Project operations workforce will be divided into four crews scheduled to work 14 out of every 28 days.

Table 13-1 Underground Workforce 2028 through 2029

   
Job Classification     Count 
   

Miners

   8
   

Mechanics

   4
   

Supervision

   2
   

Technical Staff

   8
   

Manager

   1
   

Total

   23

 

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Table 13-2 Peak Underground Workforce beginning 2030

   
Job Classification     Count 
   

Miners

   80
   

Mechanics/Electricians

   20
   

Supervision

   8
   

Technical Staff

   16
   

Manager

   1
   

Total

   125

 

  13.4.

Mobile Equipment Fleet

During the early exploration phase, capital development drifting will average 10-15 feet per day from 2022 into 2023. Following a positive production decision, ore production will begin in 2024 and ramp up to the steady state rate of 1,250 tpd. Mine development will follow the water level drawdown opening new production areas to sustain production. Table 13-3 lists the mining fleet necessary to achieve the development goals during the delineation drilling. Table 13-4 lists the mining fleet necessary to achieve the development and production goals for peak mining levels.

Table 13-3 Underground Mobile Equipment and Support Equipment for Exploration Development Phase

   
Description      Quantity  
   

6-Yd LHD

   1
   

30-T Haul Truck

   1
   

Jumbo Drill

   1
   

Bolter

   1
   

Fork Lift

   1
   

Lube Truck

   1
   

Grader

   1
   

Emergency Rescue

   1
   

Tractor

   2
   

UTV

   1

Table 13-4 Underground Mobile Equipment and Support Equipment for Peak Production Mining

   
Description      Quantity  
   

6-Yd LHD

   6
   

30-T Haul Truck

   8
   

Jumbo Drill

   4
   

Bolter

   4

 

 

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Description      Quantity  
   

Remix Truck

   2
   

Cement Pump

   2
   

Fork Lift

   2
   

Lube Truck

   1
   

Grader

   1
   

Emergency Rescue

   1
   

Heavy Duty Pickup

   1
   

Tractor

   3
   

UTV

   4

 

  13.5.

Mine Plan

The productivities of Table 13-5 were used to develop the production plan. The production plan is limited by overall production rates. Assuming a positive production decision in 2027, development and production rates will increase as headings become available, eventually reaching a maximum rate of 100 total feet per day and 1,300 tons of mineralized material production per day. At these rates, the mine plan is exhausted in 2036. The mine plan is summarized in Table 13-6. The production profile over the life of mine is shown in Figure 13-6.

Table 13-5 Heading Productivity

     

Heading Type

   Units    Daily Rate   
   

Primary Capital Development Drift

   Feet/Day      12  
   

Secondary Capital Development Drift

   Feet/Day      10  
   

Raise Bore

   Feet/Day      10  
   

Drop Raise

   Feet/Day      15  
   

Ore Drift Development

   Feet/Day      10  
   

Floor Pulls

   Ton/Day      300  
   

Breast Downs

   Ton/Day      100  
   

Long Hole Stoping

   Ton/Day      500  
   

Backfill

   Ton/Day      500  

Table 13-6 Annual Production and Development Schedule (Including Inferred Mineral Resource)

     Calendar Year     2028      2029      2030      2031      2032      2033      2034      2035      2036      Total 
           Mineralized Material Mined                        
    Mineralization Mined (000’s Tons)    -    155.7    170.1    471.7    445.1    486.3    410.8    430.2    270.8    2840.7

  

  Gold Grade (Ounce/Ton)    -    0.268    0.297    0.331    0.339    0.333    0.283    0.284    0.327    0.313
    Silver Grade (Ounce/Ton)    -    0.106    0.113    0.080    0.111    0.318    0.261    0.201    0.194    0.184
    Contained Gold (000’s Ounces)    -    41.8    50.5    156.0    150.8    162.1    116.3    122.1    88.4    888.1

 

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     Calendar Year     2028      2029      2030      2031      2032      2033      2034      2035      2036      Total 

  

  Contained Silver (000’s Ounces)    -    16.5    19.2    38.0    49.5    154.7    107.1    86.5    52.6    524
                                                       
    Marginal Mineralization Mined (000’s Tons)    -    10.0    2.8    18.0    28.8    12.4    12.2    7.6    16.9    108.6
    Gold Grade (Ounce/Ton)    -    0.092    0.089    0.108    0.105    0.103    0.118    0.124    0.119    0.107
    Silver Grade (Ounce/Ton)    -    0.029    0.042    0.028    0.039    0.188    0.155    0.110    0.076    0.077
    Contained Gold (000’s Ounces)    -    0.9    0.2    1.9    3.0    1.3    1.4    0.9    2.0    11.8
    Contained Silver (000’s Ounces)    -    0.3    0.1    0.5    1.1    2.3    1.9    0.8    1.3    8.4
                                                       
    Total Mineralization Mined (000’s Tons)    -    165.7    172.9    489.7    473.9    498.6    423.0    437.8    287.7    2949.3
    Gold Grade (Ounce/Ton)    -    0.257    0.294    0.323    0.325    0.328    0.278    0.281    0.314    0.305
    Silver Grade (Ounce/Ton)    -    0.101    0.112    0.079    0.107    0.315    0.258    0.199    0.187    0.180
    Contained Gold (000’s Ounces)    -    42.7    50.7    157.9    153.9    163.3    117.8    123.1    90.4    899.8
    Contained Silver (000’s Ounces)    -    16.8    19.3    38.5    50.7    157.0    108.9    87.3    53.9    532.4
         Production Mining                    
    Stope Development and Drift and Fill Mining (000’s Tons)    -    139.5    140.9    360.9    398.0    387.5    337.2    365.0    196.6    2,325.6
    Bench Mining (000’s Tons)         26.2    32.0    128.8    75.9    111.1    85.8    72.8    91.0    623.7
    Mineralization Production Rate (tpd)       454    474    1,342    1,295    1,366    1,159    1,200    786    897
         Backfill                    
    Total CRF Backfill (000’s Tons)    -    108.7    113.5    321.4    311.0    327.2    277.6    287.3    188.8    1,935.5
         Waste Mining                    
    Expensed Waste (000’s Tons)    -    37.2    13.5    46.1    45.2    27.0    32.1    34.4    13.7    249.3
    Primary Capital Drifting (Feet)    -    9,352    4,036    1,859    4,799    265    1,386              21,698
    Secondary Capital Drifting (Feet)    200    1,194    855    574    829    39    396              4,086
    Capital Raising (Feet)    -    934    578    241    264    -    1,727              3,744
    Capitalized Mining (000’s Tons)    3.5    210.3    95.1    43.1    107.7    4.9    49.8              514.4

Total Tons Mined (000’s Tons)

   3.5    413.2    281.5    579.0    626.8    530.5    504.9    472.2    301.4    3,713.0

Mining Rate (tpd)

   10    1,132    771    1,586    1,713    1,453    1,383    1,294    823    1,129

 

 

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Figure 13-6 Production Profile (Including Inferred Mineral Resources)

 

LOGO

The Helen deposit mineral resource contains 70% inferred mineral resources and the Gap deposit contains 89% inferred mineral resources. The mining schedule presented in Table 13-7 is derived by factoring the full mine plan containing both inferred and indicated mineral resources. Adjustments to productivity or recalculation of costs have not been performed. Capital development mining is identical in both cases. The without inferred mineral resource mine plan is depicted in Table 13-7 and Figure 13-7.

Table 13-7 Annual Production and Development Schedule (Excluding Inferred Mineral Resource)

     Calendar Year     2028      2029      2030      2031      2032      2033      2034      2035      2036      Total 
   
     Mineralized Material Mined 46.5    50.8    141.0    104.0    77.9    54.5    57.4    30.2      

  

  Mineralization Mined (000’s Tons)    -    46.5    50.8    141.0    104.0    77.9    54.5    57.4    30.2    562.4
    Gold Grade (Ounce/Ton)    -    0.268    0.297    0.331    0.345    0.346    0.295    0.307    0.327    0.321
    Silver Grade (Ounce/Ton)    -    0.106    0.113    0.080    0.106    0.254    0.237    0.190    0.194    0.147
    Contained Gold (000’s Ounces)    -    12.5    15.1    46.6    35.9    27.0    16.1    17.6    9.9    180.6
    Contained Silver (000’s Ounces)    -    4.9    5.7    11.3    11.0    19.8    12.9    10.9    5.9    82.5
                                                     

 

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     Calendar Year     2028      2029      2030      2031      2032      2033      2034      2035      2036      Total 
    Marginal Mineralization Mined (000’s Tons)    -    3.0    0.8    5.4    7.5    1.9    1.4    0.9    1.9    22.7
    Gold Grade (Ounce/Ton)    -    0.092    0.089    0.108    0.108    0.110    0.118    0.120    0.119    0.107
    Silver Grade (Ounce/Ton)    -    0.029    0.042    0.028    0.043    0.155    0.155    0.103    0.076    0.058
    Contained Gold (000’s Ounces)    -    0.3    0.1    0.6    0.8    0.2    0.2    0.1    0.2    2.4
    Contained Silver (000’s Ounces)    -    0.1    0.0    0.1    0.3    0.3    0.2    0.1    0.1    1.3
                                                     

  

  Total Mineralization Mined (000’s Tons)    -    49.5    51.7    146.4    111.5    79.8    55.9    58.3    32.1    585.1
    Gold Grade (Ounce/Ton)    -    0.257    0.294    0.323    0.329    0.341    0.291    0.304    0.314    0.313
    Silver Grade (Ounce/Ton)    -    0.101    0.112    0.079    0.102    0.252    0.235    0.188    0.187    0.143
    Contained Gold (000’s Ounces)    -    12.8    15.2    47.2    36.7    27.2    16.3    17.7    10.1    183.1
    Contained Silver (000’s Ounces)    -    5.0    5.8    11.5    11.4    20.1    13.2    11.0    6.0    83.8
         Production Mining                    
    Stope Development and Drift and Fill Mining (000’s Tons)    -    41.7    42.1    107.9    79.9    33.0    14.3    14.9    2.4    336.2
    Bench Mining (000’s Tons)         7.8    9.6    38.5    15.6    10.1    4.3    5.0    1.1    92.1
    Mineralization Production Rate (tpd)       136    142    401    305    219    153    160    88    178
         Backfill                    
    Total CRF Backfill (000’s Tons)    -    9.7    10.1    28.7    18.5    7.9    3.1    3.3    0.3    81.6
         Waste Mining                    
    Expensed Waste (000’s Tons)    -    11.1    4.0    13.3    11.0    2.6    4.1    4.0    0.0    50.2
    Primary Capital Drifting (Feet)    -    9,352    4,036    1,859    4,799    265    1,386              21,698
    Secondary Capital Drifting (Feet)    200    1,194    855    574    829    39    396              4,086
    Capital Raising (Feet)    -    934    578    241    264    -    1,727              3,744
    Capitalized Mining (000’s Tons)    3.5    210.3    95.1    43.1    107.7    4.9    49.8              514.4

Total Tons Mined (000’s Tons)

   3.5    271.0    150.8    203.3    230.3    87.3    109.8    62.4    32.1    3,713.0

Mining Rate (tpd)

   10    742    413    557    629    239    301    171    88    350

 

 

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Figure 13-7 Production Profile without Inferred Mineral Resources

LOGO

 

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14. Recovery Methods

 

  14.1.

INTRODUCTION

Mineralization from Cove operation which is amenable to roasting will be processed via dry grind, 0roasting followed by carbon in leach (CIL) at Nevada Gold Mines Gold Quarry. The most recent metallurgical testing is described in Section 10 Mineral Processing and Metallurgical Testing supports processing parameters at this facility at NGM. Up to 750 tons per day of Cove production will be processed at the Gold Quarry Roaster.

Beginning 2028, amenable production will also be processed through the Lone Tree pressure oxidation facility. Operating conditions will be determined. Production will likely be blended with feed from other i-80 operations.

 

  14.2.

Gold Quarry Roaster

The Gold Quarry roaster processes 3.4 to 3.6 million tons per year and consists of primary and secondary crushing, dry grinding, dual stage fluidized bed roasting, off-gas handling, mercury recovery systems, a slurry neutralization circuit, a carbon-in-leach (CIL) circuit with carbon stripping, cyanide detoxification circuit, and electrowinning for gold recovery. Roaster off-gas is processed to recovery sulfuric acid for sale or internal use.

Gold recovery estimates are based on both test work and operational history at both facilities with curves utilized for both depending on operating strategy and ore characteristics. The current roaster LOM has an average recovery of 88%.

The simplified Gold Quarry Roaster process flowsheet is shown in Figure 14-1.

 

 

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Figure 14-1 Nevada Gold Mines Gold Quarry Roaster Simplified Flowsheet

 

LOGO

 

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  14.3.

 Lone Tree Pressure Oxidation Facility

i-80 Gold plans to process single refractory ore from their Nevada mines at their Lone Tree Mill in a hub and spoke arrangement.

 

  14.3.1.  

Lone Tree Mill Historic Processing

The Lone Tree Mine is located immediately adjacent to I-80, approximately 12 miles west of Battle Mountain, 50 miles east of Winnemucca, and 120 miles west of Elko. Mining commenced at Lone Tree in April 1991 with the first gold pour in August of 1991. In 1993, a POX circuit was added to the facility, which included a SAG / ball mill circuit, followed by a thickening circuit, the POX process for refractory gold ores, and finally CIL, carbon stripping, and refining.

In 1997, a 4,500 tpd flotation plant was constructed to make concentrate to supplement the feed to the POX circuit, as well as to ship excess concentrate to Newmont’s Twin Creeks POX plant or to its Carlin roaster. The Lone Tree processing facilities were shut down at the end of 2007. Since that time, the mills have been rotated on a regular basis to lubricate the bearings. In general, the facility is still in place with most of the equipment sitting idle.

i-80 Gold Corp’s objective is to refurbish and restart the POX circuit and associated unit operations, including the existing oxygen plant, as it was operating before the shut-down, while meeting all new regulatory requirements. The flotation circuit is not being considered for restart. The POX circuit will have capability to operate under either acidic or basic conditions.

In order to restart the process plant, new environmental regulations in relation to allowable mercury emissions must be met. In February 2011, the NDEP and the EPA brought about new standards to limit mercury emissions to 127 lb of mercury for every million tons of ore processed. In order to meet this requirement, the Lone Tree facility will require several environmental upgrades prior to restart.

 

  14.3.2.  

Lone Tree Facility Block Flow Diagram

A block flow diagram for the Lone Tree Mill facility is included in Figure 14-2. The block flow diagram contains the follow major processing areas:

 

   

Ore Reclaim, Grinding and Thickening and Acidulation

 

   

Pressure Oxidation

 

 

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POX Off-gas Treatment and Quench Water Loop

   

Neutralization, Carbon-in-Leach, and Cyanide Destruction

   

Tailings Thickening and Filtration

   

Acid Wash, Carbon Stripping, and Carbon Regeneration

   

Electrowinning and Refinery

   

Plant and Instrument Air

   

Oxygen Plant

   

Reagent Preparation and Storage

   

Process and Plant Service Cooling Towers

   

Water Distributions

   

Steam Generating Plant and Propane Storage.

 

  14.3.3.  

Key Design Criteria

The Lone Tree Pressure Oxidation (POX) Facility restart will have minimal changes made from the 1993 PDC. A new PDC was developed based on the expected production sources as defined by i-80.

Key process design criteria are summarized in Table 14-1.

Table 14-1 Summary of Key Process Statistics

           
Criteria    Units              Value      
   

Annual Mill Throughput

   tons                  912,500         
   

Daily Throughput (per calendar day)

   tons              2,500     
   

Operating Throughput of Ore to Autoclave Circuit (LTH feed)

   tph              122.5     
   

Operating Time / Availability

   %              85     
   

Design Sulfur Treatment Rate

   tph S              2.7     
   

Gold Recovery

   %              Varies     
   

Silver Recovery

   %              Varies     

 

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Figure 14-2 Lone Tree Facility Block Flow Diagram

LOGO

 

 

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  14.3.4.

Lone Tree Facility Description

 

  14.3.4.1.

Mill Feed Reclaim

The purpose of the Mill feed reclaim area is to store and reclaim material for processing, which has been shipped to the lone tree processing facility via highway ore trucks.

Run of mine (ROM) crushed material is delivered to the stockpile area. Material from various mining locations – namely Granite Creek, Cove, and Archimedes – is dumped at designated locations within the storage area and blended into facility feed stockpiles.

The stockpile area will have the capacity to store multiple days output of mined and crushed material to accommodate the production shipment schedule to site. Additionally, the reclaim area is utilized for feed blending for the POX circuit. This blending will be used to manage the sulfide sulfur concentrations, gold grades, and carbonate grades through the autoclave to ensure stable circuit operation within the design window for the plant.

 

  14.3.4.2.

Comminution

The purpose of comminution area is to reduce the particle size of the feed ore to the target autoclave circuit feed size for sufficient sulfide oxidation kinetics and gold recovery within the autoclave. The comminution area contains an SABC circuit with a dedicated SAG (semi-autogenous grinding mill) and ball mill to reduce the feed particle size to the target grind size. The SAG mill is fed via a conveyor from the dump hopper. The ball mill cyclone overflow is directed to the POX feed thickening conveyor.

 

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  14.3.4.3.

Thickening and Acidulation

The purpose of the thickening area is to prepare the slurry for autoclave process by densifying the product of the grinding circuit to improve storage capacity of the downstream slurry storage tanks, improve the autoclave heat balance by reducing the water transferred to the autoclave and improving the possible solids flow through the autoclave feed pumps. The dense slurry is stored in two acidulation tanks that provide a combined storage / acidulation retention time of 12 hours. The acidulation tanks ensure continuous feed to the autoclave plant, unaffected by upstream throughput variations.

 

  14.3.4.4.

Pressure Oxidation

The POX autoclave circuit includes the slurry pre-heaters, autoclave feed, autoclave, and the POX ancillary services: autoclave agitator seal system, oxygen supply, high pressure cooling water, and high-pressure steam. The Lone Tree Facility restart includes provisions to operate the circuit in alkaline or acidic modes depending on the feed carbonate concentration among other factors.

 

  14.3.4.4.1.

Slurry Heaters

The purpose of the slurry heaters is to capture excess energy discharged from the autoclave and pre-heat the feed slurry prior to the autoclave process reducing the total energy input required to operate the autoclave. The heating is achieved in two stages consisting of a series of two refractory lined counter-current splash slurry heater vessels. The heat source is flashed steam released from the autoclave discharge slurry during the pressure letdown process. The splash slurry heaters are direct contact heat exchanger and provide a means of heat recovery via steam condensation. This reduces the off-gas load on the downstream off-gas equipment and reduces the required input steam.

 

  14.3.4.4.2.

Autoclave Feed

The purpose of the autoclave feed area is to increase the pressure of the pre heated slurry to above the autoclave operating pressure to facilitate transfer into the autoclave at the required pressure using the autoclave feed pumps.

 

 

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  14.3.4.4.3.

Autoclave

The purpose of the autoclave is to oxidize the refractory sulfide minerals under acidic or alkaline conditions to liberate the gold trapped in the sulfide sulfur minerals. The autoclave at Lone Tree is designed to operate at 389 °F and 297 PSI(g) with a slurry residence time of 40 - 50 minutes and consists of 4 compartments. The design expects a 78% - 97% cumulative sulfide sulfur oxidation through the autoclave depending on operating conditions. In either operating condition high purity oxygen is introduced to all four compartments of the autoclave at controlled rates to oxidize the fed sulfide minerals. Due to the low sulfur grades steam is required to be continuously fed to the autoclave to maintain the kinetically required oxidation rates to achieve the sulfide sulfur oxidation extent. The autoclave slurry is discharged through a level control choke valve and is fed to the high pressure flash vessel.

 

  14.3.4.4.4.

Flash System

The purpose of the flash system is to reduce the pressure and temperature of the autoclave discharge, making it suitable for subsequent unit operations downstream. The oxidized slurry undergoes a controlled pressure and temperature reduction process as it passes through two stages of flashing vessels located downstream of the last autoclave compartment.

 

  14.3.4.5.

POX Off-gas Treatment

The purpose of the POX off-gas treatment area is to effectively eliminate particulate matter present in the POX vent stream, while simultaneously reducing the temperature and volume of the vent gas through direct contact condensation. This process serves to alleviate the burden imposed on downstream equipment, ensuring their optimal performance, and mitigates the environmental impact by minimizing emissions. The off-gas treatment circuit also includes a mercury removal step to minimize autoclave mercury emissions to the environment.

 

  14.3.4.6.

Slurry Coolers

The purpose of slurry coolers is to reduce the temperature of the incoming slurry from the low-pressure flash vessel to prepare it for the downstream neutralization and CIL circuits through a series of water cooled shell and tube heat exchangers.

 

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  14.3.4.7.

Neutralization

The purpose of neutralization circuit is to neutralize all free acid in the slurry, precipitate the heavy metals as their hydroxides and raise the pH to approximately 10 to ensure cyanide stability in the CIL circuit for personnel safety and process optimization. The neutralization circuit is dosed with lime slurry to raise the pH of the autoclave discharge slurry. The neutralized slurry from this circuit is then fed to the CIL circuit for gold recovery.

 

  14.3.4.8.

Carbon-in-Leach

The purpose of CIL circuit is to leach and extract gold and silver from the oxidized slurry from neutralization using cyanidation and carbon adsorption. The CIL circuit provides retention time of 24 to 28 hours. The CIL circuit consists of 6 mechanically agitated tanks arranged in a series. The agitators prevent solid settlement and maximize contact time to improve gold and silver recovery. The carbon flows counter current to the slurry flows and the loaded carbon is sent to an elution circuit for carbon stripping and regeneration. Unloaded carbon is fed the last tank of the CIL circuit. The leached slurry is transferred to the cyanide destruction circuit.

 

  14.3.4.9.

Elution

The purpose of the elution circuit is to elute precious metals from the loaded carbon and transfer the resulting loaded solution of high gold concentration (pregnant eluate) to the refinery to generate doré.

 

  14.3.4.9.1.

Carbon Acid Wash

The purpose of acid wash is to rinse the loaded carbon form CIL with dilute nitric acid solution prior to the carbon stripping process. Carbonate scale builds up on the activated carbon during the CIL process and fouls the carbon’s adsorption properties by depositing a layer of scale. If left intact, over time the scale will limit the adsorption capacity of the carbon and will cause softening of the carbon in the regeneration kiln. The loaded carbon from CIL is first treated within the carbon acid wash vessel prior to treatment within the carbon stripping vessel.

 

 

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  14.3.4.9.2.

Carbon Stripping

The purpose of the carbon strip circuit is to strip the cleaned loaded carbon from the acid wash vessel of the adsorbed gold using a Pressure ZADRA Strip scheme. The ZADRA strip uses several bed volumes of a recirculated solution to strip the precious metals off the loaded carbon. The cyanide solution is buffered by caustic to assist with gold elution. The stripped carbon is then sent to carbon regeneration circuits. The loaded solution is next processed in the electrowinning circuit.

 

  14.3.4.9.3.

Elution Mercury Abatement

The purpose of elution mercury abatement system is to condition the off gas leaving the pregnant and barren solution tank to remove fine particulate, solution aerosols and condensed and gas phase mercury.

 

  14.3.4.10.

Carbon Regeneration

The purpose of the carbon regeneration circuit is to restore the activated carbon’s ability to recover gold from the cyanidation circuit solutions. The circuit also permits the introduction of new carbon to the process and removes carbon fines from the process.

 

  14.3.4.10.1.

  Carbon Regeneration Kiln

As carbon is used in the CIL and elution circuits, the surface and internal pore structure becomes contaminated with organic species. The organics foul the carbon, slow the gold adsorption rate, and decrease the gold loading capacity of the carbon. The carbon reactivation electric kiln is a horizontal rotary kiln that is specifically designed for this purpose.

 

  14.3.4.10.2.

  Carbon Fines Handling

Carbon fines are transferred by gravity from the reactivated carbon vibrating screen, carbon reactivation feed vibrating screen, kiln feed hopper, and carbon reactivation electric kiln. The carbon fines are dewatered in a filter press and discharged into supersacks for external sale.

 

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  14.3.4.11.

Refinery

The purpose of the refinery circuit is to recover gold cyanide solutions via electrowinning and produce doré bullion bars.

 

  14.3.4.11.1.

  Electrowinning

The purpose of the electrowinning (EW) circuit is to recover gold from the pregnant solution by applying a voltage across electrodes immersed in the pregnant solution. Rich solution from the pregnant solution tank is transferred through the EW cells to electrowin the gold.

 

  14.3.4.11.2.

  Refining

The purpose of the refining process is to produce doré bars void of other contaminants including but not limited to mercury.

The sludge from the EW cells is first processed in a mercury retort oven to remove the co-captured mercury from the precious metals recovery steps. The retorted gold sludge is then processed in a melt furnace to produce the final mine grade doré bars.

 

  14.3.4.12.

Cyanide Destruction

The purpose of the cyanide destruction circuit is to effectively reduce the concentration of cyanide in the final tail discharge and the recycled process water, ensuring compliance with predefined environmental standards and regulations and improving the safety of the operation by reducing cyanide concentrations outside of the CIL and elution circuits. The circuit targets a specific concentration limit of 2.5 mg/L of residual weakly acid-dissociable cyanide (CNWAD). This reduction is accomplished through the application of the SO2/air cyanide destruction process, which oxidizes the cyanide to meet the required concentration level. The cyanide destruction circuit is fed directly from the slurry discharge from the CIL circuit.

 

 

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  14.3.4.13.

Tailings Preparation

The purpose of the tailings circuit is to increase the density of the detoxified tailings to aid with dry stacking of tailings residue. Additionally, this circuit produces process water for internal use within the facility. The tailings preparation circuit consists of a thickener as a first stage of solids densification. The thickener underflow is then fed to a tailings filtration circuit which dewaters the tailings sufficiently to support tailings dry stacking. The de-watered tailings from the filter presses are then dry stacked at the tailings storage facility.

The water removed from the tailings slurry is used as process water within the facility to offset water requirements. Excess process water is processed via a reverse osmosis circuit to provide supplemental permeate water to offset fresh water requirements.

 

  14.3.4.14.

Water Distributions

There are eight types of defined water services at Lone Tree:

 

   

Fresh water – Is generally used for reagent make-up and water washing streams.

   

Gland water – Is used to supply gland water to slurry pumps.

   

Mill water – Is used to provide dilution water within the milling circuit.

   

Potable water – Is used for safety showers and sanitary uses.

   

Demineralized water – Is primarily used to supply the steam generating plant.

   

Process water – Is used for washing and slurry dilutions. Additionally, generally feeds the reverse osmosis circuit to generate permeate water.

   

Quench water – Is used within the POX off-gas circuit as the source of direct cooling water.

   

Excess water – Is discharged from the main processing facility to the existing heap leach facility for treatment.

 

  14.3.4.15.

Solution Cooling

The purpose of the cooling area is to reject heat absorbed within the process to atmosphere. The solution cooling area includes the process service cooling circuit and the plant service cooling circuit. The process cooling circuit rejects the heat from the autoclave cooling circuit and the elution circuit heat exchangers. The plant service cooling circuit provides trim heat rejection from various equipment support systems throughout the design.

 

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  14.3.4.16.

Reagents

Each set of compatible reagent preparation and storage systems is located within dedicated containment areas to prevent erroneous mixing of reagents. Storage tanks are equipped with level indicators, instrumentation, and alarms to reduce the risk of spills during normal operation. Appropriate ventilation, fire and safety protection, safety shower stations and Safety Data Sheet stations are located throughout the facility.

 

  14.3.4.16.1.

  Oxygen Plant

High purity oxygen is primarily used for oxidation of sulfide during the POX process, of iron conversion from ferrous to ferric in the neutralization circuit, and of cyanide to cyanate in cyanide destruction. Furthermore, during cyanidation, the addition of oxygen maximizes the rate of gold dissolution. At Lone Tree, a cryogenic ASU produces high purity oxygen. The unit uses pressure swing adsorption technology for front end purification and production of high-pressure oxygen at 95% purity.

 

  14.3.4.17.

Instrument and Plant Air

The Lone Tree facility includes separate instrument and plant air systems to support the facilities air requirements.

 

  14.3.5.

Utilities Consumption

The plant consumptions for water and power are provided for the average processing case below and consider the design blend of material to be processed within the Lone Tree Facility for the design life of operation.

 

  14.3.5.1.

Water Consumption

Table 14-2 provides a summary of the water consumption by type for the Lone Tree processing facility.

 

 

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Table 14-2 Lone Tree Facility Water Consumption by Type

   
Type     Consumption (gpm) 
   

Mill Water

   1 550
   

Fresh Water

   570
   

Permeate Water

   195
   

Low Pressure Gland Water 

   105
   

High Pressure Gland Water 

   170
   

Demineralized Water

   110
   

Potable Water

   15

 

  14.3.5.2.

Electrical Power Requirements

The estimated annual electrical energy requirements for the Lone Tree processing facility are summarized by area in Table 14 3.

Table 14-3 Lone Tree Facility Energy Usage by Area

   
Area  

Annual Energy

 Consumption (MWh/y) 

   

000 – General Plant Wide

  2 250
   

180 – Water System

  930
   

181 – Potable Water

  240
   

182 – Process Water (RO and Process Water Tank)

  4 900
   

210 – Ore Reclaim

  770
   

240 – Refinery

  2 310
   

241 – POX Grinding

  26 920
   

242 – POX Grinding Thickening and Acidulation

  1 890
   

244 – Neutralization and CIL and Acid Storage

  6 540
   

245 – Carbon Stripping

  4 090
   

247 – CND

  690
   

248 – Reagents

  2 640
   

249 – Plant Air and Propane

  3 310
   

250 – Pressure Oxidation (POX) and POX Utilities

  15 540
   

251 – POX Demineralized Water System

  2 660
   

275 – Tailings Filtration

  13 690
   

300 – Plant Wide Electrical and Instrumentation

  4 000
   

305 – ABS and CN Storage

  160
   

320 – POX Mercury Abatement

  900

 

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Area   

Annual Energy

 Consumption (MWh/y) 

   

340 – Quench Water Treatment

   4 020
   

255 – Oxygen Plant

   40 090
   

099 – Existing Plant Areas

   3 570
   

Total

   142 090
  15. 

Infrastructure

 

  15.1.

Dewatering

 

  15.1.1.

History

Dewatering of the Cove Pit occurred from 1988 until mid-2001 utilizing surface dewatering wells, sumps, and horizontal drains. Water pumped from the dewatering wells was piped to a series of rapid infiltration basins (RIBs) located north of the pit, where the water was infiltrated into the alluvium of the Reese River Valley. All wells constructed for dewatering purposes have been abandoned in accordance with Nevada Division of Water Resources regulations as part of the mine’s closure plan. Following cessation of dewatering activities, a pit lake began forming in 2001 and has reached an elevation of approximately 4,626 ft. (Piteau Associates USA Ltd., 2018)

The pit reached the ground water level in 1991. The pumping rate peaked at 19.000 gpm in 1994 and 1995. By the year 2000, the last full year of mining it had declined to 13,400 gpm. The infrastructure required to move this volume of water included 23 pumping wells and two in pit pumping stations. (Echo Bay Minerals Company, 2002)

 

  15.1.2.

Infrastructure

All of the historical dewatering infrastructure has been removed except the water monitoring wells and piezometers. i-80 will be required to construct the following:

 

   

Dewatering wells with production string;

 

   

Pit lake dewatering barge;

 

   

Electrical distribution lines and transformers to the dewatering wells and pit lake barge;

 

   

Collection pipelines to each well and pit lake barge’

 

   

Rapid Infiltration Basins, and;

 

   

Pipeline from the collection piping system to the RIBs.

 

 

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Tentative well locations are shown in Figure 15-1. Ribs will be located several miles to the North of the proposed underground mining area.

Figure 15-1 Proposed Dewatering Well Location

LOGO

 

  15.2.

Electrical Power

Dewatering constitutes 90% of electrical power demand over the Project’s duration. Demand for dewatering was estimated from projected water elevations and pumping rates and peak demand of 11.5 megawatts (MW) occurs in 2028 (Figure 15-2).

 

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Figure 15-2 Electrical Demand

LOGO

An existing NV Energy 24.9 kilovolt (kV) distribution line and meter will provide one megawatt (MW) to the Cove Project during the initial decline development and underground drilling program. Permanent power for the project will be supplied by an existing 120 kV transmission line. This line previously powered the Cove Project and extends approximately 9 1⁄2 miles from NV Energy’s Bannock substation to and terminates at the Cove Project. The line is in good condition and will not require any repairs.

The Bannock substation serves the Phoenix Mine and a geothermal power plant located in Jersey Valley. The substation has ample capacity to provide the estimated 11.5 MW of power required by the Cove Project. Prior to reconnecting the line to the grid NV Energy requires updating the switchgear at the substation to a ring configuration as a result of new standards implemented since the line was taken out of service after the cessation of activities at Cove by Echo Bay. The full cost of these upgrades will be borne by the Cove Project.

Where the lines cross the project access road a new substation will be constructed. It currently contains a 24.9/13.8 kV, 1,500 kilovolt-ampere pad mounted transformer and related equipment. Approximately 7,500 feet of distribution line connects the substation to the portal site and related surface facilities (Figure 15-3).

 

 

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The substation will be upgraded with a 120/13.8 kV transformer when permanent power is being connected that will feed the distribution line to the portal. As the dewatering wells are completed additional distribution lines will be added to connect the wells.

Figure 15-3 Electrical Site Plan

LOGO

(Quantum Electric 2017)

 

  15.3.

Mine Facilities

The proposed location of mine facilities is shown in Figure 15-4. The laydown area will contain the mine office, maintenance shop, equipment wash down bay, fuel and oil storage, employee dry facilities and warehouse.

 

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Figure 15-4 Mine Facilities Layout

LOGO

 

  15.4.

Backfill

Backfill material for unconsolidated waste fill (GOB) can be obtained from any suitable source such as development waste, open pit waste dumps, or leach pads.

Backfill material for Cemented Rock Fill (CRF) will need to meet specifications designed to achieve minimum Uniaxial Compressive Strength (UCS) specifications. This specification is designed to provide the pillar strength needed to maintain stability of adjacent underground excavations and may require screening and/or crushing. The results of backfill testing for six types of material available at Cove are shown in

Table 15-1. CRF material will be mixed at a backfill plant located near the portal and transported underground using the same truck fleet used to remove mineralized material and waste from the mine.

 

 

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Table 15-1 Backfill Scoping Tests 28-Day Unconfined Compressive Strength (psi)

       
Cement Content    4%    6%    8%
     
Aggregate Source                  
     

Waste

   440             510             830         
     

Tails

   60    90    120
     

Tuff

   90    140    240
     

Pad 3

   360    590    500
     

Pad 2

   190    200    260
     

Mill Rejects

   210    510    810

 

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Market Studies and Contracts

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  16. 

Market Studies and Contracts

 

  16.1.

Precious Metal Markets

Gold and silver are fungible commodities with reputable smelters and refiners located throughout the world. The price of gold has reached all-time highs in 2024 with December’s price averaging 2,644 per ounce. As of December, 2024 the three-year trailing average gold price was $2,044 per ounce and the two-year trailing average price was $2,166 per ounce. The three -year and two-year trailing average prices for silver in December 2024 were $24.50 and $25.88 per ounce respectively. Historical plots for both are shown in in Figure 16-1.

Figure 16-1 Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average

LOGO

Issuers may also rely on published forecasts from reputable financial institutions. The current long term price forecast by CIBC is $2,169 and per ounce and $27.61 per ounce for gold and silver respectively. (CIBC., 2025)

 

 

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Commodity prices for Mineral Reserves are chosen not to exceed financial institution forecasts or the three-year trailing average price. Commodity pricing for the estimation of Mineral Resources can be 10% to 20% higher than that used for Mineral Reserves. The gold price selected for estimating Mineral Resources disclosed in this technical report is $2,175. The silver price selected is $27.25 per ounce.

 

   16.2.

Contracts

 

  16.2.1.

Financing Agreements

 

  16.2.1.1.

Orion and Sprott Financing Package

The Company entered into a financing package with OMF Fund III (F) Ltd. an affiliate of Orion Mine Finance (collectively “Orion”) on December 31, 2021, and a fund managed by Sprott Asset Management USA, Inc. and a fund managed by CNL Strategic Asset Management, LLC (“Sprott”) on December 9, 2021 (together the “Finance Package”).

The Financing Package in its aggregate consists of:

a.  $50 million convertible loan (the “Orion Convertible Loan”)

b.  $10 million convertible loan (the “Sprott Convertible Loan” and together with the Orion Convertible Loan, the “Convertible Loans”)

c.  $45 million gold prepay purchase and sale agreement entered into with affiliates of Orion (the “Gold Prepay Agreement”), including an accordion feature potentially to access up to an additional $50 million at i-80 Gold’s option

d.  $30 million silver purchase and sale agreement entered into with affiliates of Orion (the “Silver Purchase Agreement”), including an accordion feature to potentially access an additional $50 million at i-80 Gold’s option and an amended and restated offtake agreement entered into with affiliates of Orion (the “A&R Offtake Agreement”)

e.  5,500,000 warrants of the Company issued to Orion (the “Orion Warrants” and together with the Orion Convertible Loan, Gold Prepay Agreement, Silver Purchase Agreement and the A&R Offtake Agreement, the “Orion Finance Package”).

Under the Gold Prepay Agreement, i-80 Gold was due to deliver to Orion 3,000 troy ounces of gold for each of the quarters ending March 31, 2022 and June 30, 2022, and thereafter, 2,000 troy

 

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ounces of gold per calendar quarter until September 30, 2025 in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold. i-80 Gold may request an increase in the $45 million prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the Gold Prepay Agreement.

The final Gold Prepay Agreement includes an amendment to adjust the quantity of the quarterly deliveries of gold, but not the aggregate amount of gold, to be delivered by the Company to Orion over the term of the Gold Prepay Agreement. Under the amended Gold Prepay Agreement, commencing on the date of funding, the Company is required to deliver to Orion 1,600 troy ounces of gold for the quarter ending March 31, 2022, 3,100 troy ounces of gold for the quarter ending June 30, 2022, and thereafter 2,100 troy ounces of gold per calendar quarter until September 30, 2025, in satisfaction of the $45 million prepayment, for aggregate deliveries of 32,000 troy ounces of gold, subject to adjustment as contemplated by the terms of the Gold Prepay Agreement. As the funding from Orion did not occur until April 2022, payment for the delivery of 1,600 ounces for the quarter ending March 31, 2022 was offset against the $45 million of proceeds received from Orion.

Under the Silver Purchase Agreement, commencing April 30, 2022, i-80 Gold will deliver to Orion 100% of the silver production from the Granite Creek and Ruby Hill projects until the delivery of 1.2 million ounces of silver, after which the delivery will be reduced to 50% until the delivery of an aggregate of 2.5 million ounces of silver, after which the delivery will be reduced to 10% of the silver production solely from the Ruby Hill Project. Orion will pay i-80 Gold an ongoing cash purchase price equal to 20% of the prevailing silver price. Until the delivery of an aggregate of 1.2 million ounces of silver, i-80 Gold is required to deliver the following minimum amounts of silver (the “Annual Minimum Delivery Amount”) in each calendar year: (i) in 2022, 300,000 ounces, (ii) in 2023, 400,000 ounces, (iii) in 2024, 400,000 ounces, and (iv) in 2025, 100,000 ounces. Upon a construction decision for the Ruby Hill project, comprised of one or both of the Ruby Deep or Blackjack Deposits, which construction decision is based on a feasibility study in form and substance satisfactory to Orion, acting reasonably, i-80 Gold will have the right to request an additional deposit from Orion in the amount of $50 million in aggregate in accordance with the terms of the Silver Purchase Agreement.

Both the Gold Prepay Agreement and the Silver Purchase Agreement were funded on April 12, 2022 with i-80 Gold receiving net proceeds of $71.6 million after netting the aforementioned March 31, 2022 gold delivery and closing costs as further described in Note 10 and Note 24 in the Company’s Financial Statements.

The main amendments reflected in the A&R Offtake Agreement include the increase in the term of the agreement to December 31, 2028, the inclusion of the Granite Creek and Ruby Hill projects, and the increase of the annual gold quantity to up to an aggregate of 37,500 ounces in respect of

 

 

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the 2022 and 2023 calendar years and up to an aggregate of 40,000 ounces in any calendar year after 2023. During the year ended December 31, 2022, Orion assigned all of its rights, title and interest under the A&R Offtake Agreement to TRR Offtakes LLC, now Deterra Royalties Limited.

On September 20, 2023, the Company entered into an Amended and Restated (“A&R”) Gold Prepay Agreement with Orion, pursuant to which the Company received aggregate gross proceeds of $20 million (the “2023 Gold Prepay Accordion”) structured as an additional accordion under the existing Gold Prepay Agreement.

The 2023 Gold Prepay Accordion will be repaid through the delivery by the Company to Orion of 13,333 troy ounces of gold over a period of 12 quarters, being 1,110 troy ounces of gold per quarter over the delivery period with the first delivery being 1,123 troy ounces of gold. The first delivery will occur on March 31, 2024, and the last delivery will occur on December 31, 2026. Obligations under the A&R Gold Prepay Agreement, including the 2023 Gold Prepay Accordion, will continue to be senior secured obligations of the Company and its wholly-owned subsidiaries Ruby Hill Mining Company, LLC and Osgood Mining Company, LLC and secured against the Ruby Hill project in Eureka County, Nevada and the Granite Creek project in Humboldt County, Nevada.

The remaining terms of the A&R Gold Prepay Agreement remain substantially the same as the existing Gold Prepay Agreement. The Company may request an increase in the prepayment by an additional amount not exceeding $50 million in aggregate in accordance with the terms of the A&R Gold Prepay Agreement.

In connection with the 2023 Gold Prepay Accordion, the Company issued to Orion warrants to purchase up to 3.8 million common shares of the Company at an exercise price of C$3.17 per common share until September 20, 2026, and extended the expiry date of 5.5 million existing warrants by an additional 12 months to December 13, 2025.

 

  16.3.

Roaster Toll Milling Agreement

A roaster toll milling agreement was negotiated coincident with the purchase of 100% interest in the McCoy/Cove property from Newmont, now Nevada Gold Mines. The agreement allows for processing up to 750 tpd of mineralized material at the Gold Quary roaster. Final tolling charges will be determined at the time of processing.

 

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  16.4.

Refractory Mineralized Material Sale Agreement

Refractory mineralization mined prior to 2028 will be sold to a third party for processing under an existing agreement. Payment will be made for 58% of the contained gold at the average gold price realized during the month the material was processed. The processing agreement applies to all i-80 projects and allows a maximum purchase rate of 1,000 tons per day from all i-80 operations. The QP’s have reviewed this agreement and find the terms and conditions are in accordance with industry standard practice.

 

  16.5.

Other Contracts

The company also intends to negotiate contracts for underground mine development, production mining, and over-the-road haulage with reputable contractors doing business in northeast Nevada. At the time of this report these negotiations have not been initiated. From time to time the company enters into other contracts for goods and services as a routine course of business.

 

 

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  17.

Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups

Au-Reka is the owner and operator of the Project and is responsible for all permitting requirements associated with the site including ensuring that mineral exploration activities are conducted in compliance with all applicable environmental protection legislation. The Project is primarily located on public lands administered by the Bureau of Land Management (BLM) and is subject to both Federal and State permitting requirements. Au-Reka is unaware of any existing environmental issues or compliance problems that have the potential to impede production at the Project. Au-Reka is working closely with both State and Federal regulators to ensure that the permitting and compliance strategies are acceptable and will not cause delays in production or mine development. At this time, there are no community or social impact issues regarding work being completed at the Project. Au-Reka continues to engage with the surrounding community and other external stakeholders.

The Project site is located within a previously mined area and most activities are currently being conducted or are planned on existing previously disturbed or mined areas, thereby limiting the potential environmental impacts to the site. All necessary studies and permits are in place to support the permitted exploration and test mining activities at the site.

The underground exploration decline was advanced in 2022-2023 followed by infill drilling of the ore body that continued in 2025.

Au-Reka is currently in the process of advancing the full-scale underground mining phase of the Project and, during 2023-2024, baseline studies were conducted to support the federal permitting action which is expected to be an Environmental Impact Statement (EIS) through the BLM. In addition, an associated POO Amendment will also be completed and submitted to address the Project specifics for full-scale underground mining. Nevada Division of Environmental Protection (NDEP) permits will also be secured as the Project permitting progresses.

 

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Groups

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  17.1.

Social or Community Impacts

The following information on community relations and stakeholder consultation was provided by Au-Reka personnel in 2024.

Au-Reka is deeply committed to ensuring that local ranchers and Tribal officials are actively involved in the progress of the Project, particularly in understanding both the potential benefits and the possible environmental or social impacts that may arise as the project advances. The company recognizes the importance of maintaining open and transparent communication with these key stakeholders and values their input in decision-making processes.

In addition to this, Au-Reka works closely with the Northeastern Nevada Regional Development Authority and provides project updates as it relates to regional economic development and growth. Furthermore, the company has established a strong partnership with the Lander County School District and Great Basin College, supporting various educational initiatives that benefit local students and the community at large. These initiatives include programs aimed at enhancing educational resources, offering career training opportunities, and providing support for workforce development in the region.

Beyond these partnerships, Au-Reka places a high priority on maintaining positive, long-term relationships with local government officials, tribal leaders, ranchers, and neighboring landowners, ensuring that all parties are heard, respected, and included in the development process. Through these efforts, Au-Reka strives to be a responsible and engaged community partner, prioritizing the well-being of the area and its residents throughout the duration of the Project.

 

 

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  17.2.

Permitting

Au-Reka currently holds three separate POO and associated State of Nevada Reclamation Permits in relation to the larger McCoy Cove land package. Of these, one pertains to the legacy facilities, including tailings dam and leach pads, that were shut down in 2001 and have been largely reclaimed. A second POO and Reclamation Permit pertains to exploration on the property that is not proximal to the current resource area. The third POO and Reclamation Permit pertains to the current resource area encompassing surface exploration, portal construction, initial underground development, underground delineation and exploration drilling, hydrological testing and baseline data collection. Additionally, there are three Water Pollution Control Permits associated with the site that will require modifications as part of the full-scale mining permitting actions. A list of currently held permits relevant to the exploration and development of the current resource are listed in Table 17-1.

Table 17-1 Cove Project Existing Permits

     
Permit Name    Agency    Permit Number
     
Plan of Operations - Cove Helen UG    BLM    NVN-088795
     
Plan of Operations - McCoy Cove Mine    BLM    NVN-067086
     
Plan of Operations - McCoy Cove Exploration    BLM    NVN-067716
     
Mine Reclamation Permit - Cove Helen UG    NDEP-BMRR    0342
     
Mine Reclamation Permit - McCoy Cove Mine    NDEP-BMRR    0147
     
Mine Reclamation Permit - McCoy Cove Exploration    NDEP-BMRR    0062
     
Surface Area Disturbance Permit - Cove Helen UG    NDEP-BAPC    AP1041-2192.03
     
Surface Area Disturbance Permit - McCoy Cove Exploration    NDEP-BAPC    AP1041-3762
     
Water Pollution Control Permit - Cove Helen UG    NDEP-BMRR    NEV2010102
     
Water Pollution Control Permit - Cove Helen RIBs    NDEP-BMRR    NEV2010107
     
Water Pollution Control Permit - McCoy Cove Mine    NDEP-BMRR    NEV0088009
     
Mining Stormwater General Permit    NDEP-BWPC    NVR300000: MSW-678
     
Onsite Sewage Disposal System General Permit    NDEP-BWPC    GNEVOSDS09L0112
     
Industrial Artificial Pond Permit    NDOW    S400418
Dam Safety Permit   

Nevada State

Engineer/NDWR

   J-495
     
Class III Waivered Landfill    NDEP-BSMM    SW335

 

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  17.3.

Closure and Reclamation Requirements

Reclamation of disturbed areas resulting from activities associated with the facilities included in the Project will be completed in accordance with NDEP and BLM regulations. The objectives of the reclamation work include the following:

 

   

Ensure public safety;

 

   

Reduce or eliminate potential environmental impacts;

 

   

Return the site to a condition that will support future use;

 

   

Control infiltration, erosion, sedimentation, and related degradation of existing drainages in an effort to minimize off-site impacts; and

 

   

Employ reclamation practices using proven methods that do not require ongoing maintenance.

Principal land uses in the Project area include mineral exploration and development, livestock grazing, wildlife habitat, and dispersed recreation. Following closure and final reclamation, the Project area will support the multiple land uses of livestock grazing, wildlife habitat, and recreation. Au-Reka will work with agencies and local governments and tribes to evaluate alternative land uses that could provide long-term socioeconomic benefits from the mine infrastructure. Post-closure land uses will be in conformance with the BLM Battle Mountain Resource Management Plan, Eureka-Shoshone Resource Management Plan (BLM, 1987), and Lander County zoning ordinances.

The goal of the reclamation program is to provide a safe and stable post-mining landform that supports defined land uses. To achieve this goal, the following objectives will be accomplished:

 

   

Minimize erosion and protect water resources through control of water runoff and stabilization of mine facilities;

 

   

Establish post-reclamation surface soil conditions conducive to the regeneration of a stable plant community through stripping, stockpiling, and reapplication of growth media;

 

   

Revegetate disturbed areas with a diversity of plant species in order to establish productive long-term plant communities compatible with post-mining land uses; and

 

   

Maintain public safety by stabilizing or limiting access to landforms that could constitute a public hazard.

With these objectives in mind, reclamation activities are designed to do the following:

 

 

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Stabilize the disturbed areas to a safe condition; and

 

   

Protect both disturbed and undisturbed areas from unnecessary and undue degradation.

Surface management regulations within 43 CFR 3809.420 establish performance standards that apply to this Plan. Measures to be taken to prevent unnecessary or undue degradation will be implemented during design, construction, operation, and closure of the Project and include, but are not limited to:

 

   

Regulated components of the facility will be designed and constructed to meet or exceed the BLM / NDEP / NDOW / Nevada Division of Water Resources (NDWR) design criteria. Waste rock facilities and stockpiles, which do not require engineered containment, will be evaluated for potential to release constituents and will be monitored routinely or in accordance with an approved waste rock monitoring plan;

 

   

Surface disturbance will be limited to that which is reasonably incidental to exploration, mining, and mineral processing operations;

 

   

In-pit benches, highwalls, and haul roads will be left in place;

 

   

At mine closure, a four-strand wire fence or other appropriate barrier will be placed around pit perimeters where practical and safe to do so;

 

   

Pit ramps will be barricaded in a similar manner to prevent entrance;

 

   

Mineral exploration and development drillholes, monitoring and observation wells, and production dewatering wells subject to Nevada regulations will be properly abandoned to prevent potential contamination of water resources;

 

   

Regulated wastes will be managed according to relevant regulations;

 

   

Fugitive dust emissions from disturbed and exposed surfaces will be controlled in accordance with NDEP regulations and permits;

 

   

Surface water drainage control will be accomplished by diverting stormwater, isolating facility runoff, and minimizing erosion;

 

   

Where suitable as a growth medium, surface soils and some alluvial material in the open pit will be managed as a growth media resource and removed, stockpiled, and used during reclamation; and

 

   

A Reclamation Plan will be implemented that addresses earthwork and recontouring, revegetation and stabilization, detoxification and disposal, and monitoring operations necessary to satisfactorily reclaim the proposed disturbance including roads, process ponds, heaps, waste rock facilities, buildings, and equipment.

 

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 i-80 Gold Corp  

Environmental Studies, Permitting and Plans,

Negotiations or Agreements with Local Individuals or

Groups

  Page 199 

 

Suitable growth media will be salvaged during development of site facilities as practical. Suitable alluvial material from the open pits also will be salvaged as growth media. Growth media will be placed in stockpiles within the proposed disturbance area (i.e., ancillary disturbance area or completed portions of the waste rock facilities) and will be located such that mining operations will not disturb the stockpiles. The stockpiles will be recontoured to slopes of 2.5H:1V and seeded with an interim seed mix to minimize wind and water erosion. BMPs (e.g., silt fences or staked weed-free straw bales) will also be used, as necessary, to control sediment transport. Alternatively, the growth media may be transported to, and redistributed on, mine-related surface disturbance areas undergoing concurrent reclamation (e.g., waste rock facilities).

A cost estimate for reclamation was developed using the Nevada Standardized Reclamation Cost Estimator. The estimated cost for reclamation including costs to reclaim existing facilities currently used or to be re-purposed to support the full-scale mining aspects will be determined once the Project permitting actions have been approved by the BLM and NDEP. Current bonding in place, excluding the full-scale mining aspects, for the entire McCoy Cove site is $13,841,592 (Table 17-2).

Table 17-2 McCoy Cove Reclamation Bonds

   
McCoy Cove Mine    $5,431,494  
   
McCoy Cove Exploration    $865,763  
   
Cove Helen UG Mine    $7,544,335  
   
Total    $13,841,592  

 

  17.4.

Closure and Reclamation

Reclamation bonding requirements during the pre-construction phase of the project are $166k per year and increase to $826k per year during construction and operation of the mine. Regulatory bonding requirements will be satisfied by the purchase of surety for an annual cost of 2% per year. Estimated reclamation costs net of salvage total $22.9M. Post closure monitoring is forecast to continue for 5 years following final reclamation at a cost of $250k per annum. Closure and reclamation costs on a per unit basis total $42.30 per gold ounce (Table 17-3).

Table 17-3 Annual Closure and Reclamation Costs ($M)

           
     2025 -

2027

   2128-

2035

   2036 –

2040

   2041-

2045

   Total  
   
Reclamation Bonding    0.2    0.8    -    -    7.1  
   
Reclamation    -    -    4.6    -    22.9  

 

 

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     2025 -

2027

   2128-

2035

   2036 –

2040

   2041-

2045

   Total  
   
Closure and Monitoring    -    -    -    0.3    1.3  
   
Total    0.4    1.9    4.6    0.3    31.3  

 

  17.5.

QP Statement

Practical Mining is of the opinion that the current plans for environmental compliance, permitting, closure are adequate to address any concerns of local individuals and groups.

 

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 i-80 Gold Corp   Capital and Operating Costs   Page 201 

 

  18.

Capital and Operating Costs

 

  18.1.

Capital Costs

Costs were generated from estimates provided by local suppliers and contractors and from similar work performed at other area mines. All cost estimates include Lander County and Nevada sales taxes of 7.1%, freight, contractor mobilization and demobilization, engineering procurement, and construction management. Capital cost estimates for the project are summarized in Table 18-1 and detailed in Table 18-2 through Table 18-5.

Contingencies for capital costs are 15% for development and drilling and 25% on the remaining capital expenditures. The estimated accuracy for capital costs are within an Initial Assessment level of +/-50%.

Table 18-1 Project Capital Costs ($M)

         
     Pre-Construction   Construction   Sustaining     
         
      2025     2026    2027   2028   2029   2030    2031     2032     2033    2034   Total
         
Mine Development   -   -   -   0.4   24.8   12.1   5.8   12.3   0.6   10.5   66.5
         
Dewatering   -   -   -   69.9   18.0   -   -   -   -   -   87.9
         
120kv Substation   -   -   -   3.1   -   -   -   -   -   -   3.1
         
Mine Facilities   -   -   -   1.3   0.9   1.0   0.1   0.2   -   -   3.5
         
Pre-Production   1.7   1.7   1.7   10.9   -   -   -   -   -   -   15.9
         
Env. Permitting & Feas.   4.0   2.0   1.0   -   -   -   -   -   -   -   7.0
         
Drilling   2.0   -   -   -   -   -   -   -   -   -   2.0
         
Contingency   1.7   0.9   0.7   19.5   8.5   2.1   0.9   1.9   0.1   1.6   37.8
         
Total   9.4   4.6   3.3   105.2   52.2   15.2   6.8   14.4   0.7   12   223.8
         
    17.3   157.4   49.1   223.8

Contingency - 15% development and drilling, 25% dewatering, substation, mine facilities and pre-production expenses.

The mine development unit costs shown in Table 18-2 are typical contractor costs in northern Nevada. These combined with the mine development schedule presented in Section 13 yield the development capital shown in Excludes contingency

Table 18-3.

Table 18-2 Mine Development Unit Costs

   
Description    $/ft    
   
Primary Drifting (15 ft x 17 ft)      $2,000    
   
Secondary Horizontal Access (15 ft x 15 ft)      $2,000    
   
Raise Bore (10 ft dia.)      $4,000    

Excludes contingency

 

 

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Table 18-3 Mine Development Capital ($M)

         
     Pre-Construction   Construction   Sustaining     
         
      2025     2026     2027     2028     2029      2030    2031     2032     2033     2034     Total 
         
Helen   -   -   -   0.4   24.8   12.1   5.8   3.0         45.8
         
Gap   -   -   -   -   -   -   -   9.3   0.6   10.5   20.7

Total

  -   -   -   0.4   24.8   12.1   5.8   12.3   0.6   10.5   66.5

Excludes contingency

Dewatering capital includes 15 pumping wells and a pit lake barge. Well drilling and completion costs are approximately $3.4M per well. Costs include drilling, completion, and pumping equipment. Dewatering capital costs are listed in Table 18-4.

Table 18-4 Dewatering Capital ($M)

     
      Construction      
     
     2028   2029    Total 
     

Wells

  41.1   10.3   51.3
     

RIBS & Pipeline

  21.3   7.1   28.4
     

Electrical

  1.9   0.6   2.5
     

Monitor Wells

  0.2   -   0.2
     

Pit Barge

  5.5   -   5.5
     
Dewatering Total   69.9   18.0   87.9
  1.

Excludes contingency

  2.

(Montgomery & Assoc.2024)

Table 18-5 Facilities and Site General ($M)

         
     Pre-Construction   Construction   Sustaining     
         
      2025     2026     2027     2028     2029     2030     2031     2032     2033     2034     Total 
         
Environmental and Permitting   2.0   2.0   1.0   -   -   -   -   -   -   -   5.0
         
Metallurgical Testing and Feasibility Study   2.0   -   -   -   -   -   -   -   -   -   2.0
         
Backfill Plant   -   -   -   1.3   -   -   -   -   -   -   1.0
         
Substation   -   -   -   3.1   -   -   -   -   -   -   3.1
         
Administration and Management   1.7   1.7   1.7   3.6   -   -   -   -   -   -   8.7
         
Electrical Power   -   -   -   7.3   -   -   -   -   -   -   7.3
         
Escape Hoists   -   -   -   -   0.5   0.5   -   -   -   -   1.0
         
Fans and Load Centers   -   -   -   0.3   0.4   0.5   0.1   0.2   -   -   1.5
         
Facilities Total   5.7   3.7   2.7   15.4   0.9   1.0   0.1   0.2   -   -   29.5

Excludes contingency

 

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 i-80 Gold Corp   Capital and Operating Costs   Page 203 

 

 

  18.2.

Closure and Reclamation

Reclamation bonding requirements during the pre-construction phase of the project are $166k per year and increase to $826k per year during construction and operation of the mine. Regulatory bonding requirements will be satisfied by the purchase of surety for an annual cost of 2% per year. Estimated reclamation costs net of salvage total $22.9M. Post closure monitoring is forecast to continue for 5 years following final reclamation at a cost of $250k per annum. Closure and reclamation costs on a per unit basis total $42.30 per gold ounce (Table 18-6).

Table 18-6 Closure and Reclamation Costs ($M)

           
      2025 - 

2027

    2128- 

2035

    2036 – 

2040

    2041- 

2045

   Total  
   
Reclamation Bonding    0.2    0.8    -    -    7.1  
   
Reclamation    -    -    4.6    -    22.9  
   
Closure and Monitoring    -    -    -    0.3    1.3  
   
Total    0.4    1.9    4.6    0.3    31.3  

 

  18.3.

Operating Costs

The unit mining costs presented in Table 18-7 are typical contractor costs for the anticipated conditions at Cove. Operating cost estimates are suitable for Initial Assessments and are accurate to within +/- 50%.

Table 18-7 Unit Operating Costs

     
Item    Unit Cost    Units
   
Stope Development    $110.59    $/ ton
   
Production    $80.00    $ /ton
   
Cemented Backfill    $37.93    $ /fill ton
   
Gob Fill    $13.00    $ /fill ton
   
Expensed Waste    $110.59         $ /waste ton     
   
Hauling to Roaster    $24.38    $ /ton
   
Trucking to Lone Tree    $14.68    $ /ton
   
Toll Roasting    $75.00    $ /ton
   
Alkaline Pressure Oxidation    $70.81    $ /ton

Table 18-8 One Way Trucking Distance to Nevada Metallurgical Plants

   
Name and Description    Distance (miles) 
   
NGM Goldstrike Roaster    107
   
NGM Goldstrike Autoclave    106
   
Jerritt Canyon Roaster (Idle)    150

 

 

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  i-80 Gold Corp 

 

   
Name and Description    Distance (miles) 
   
NGM Gold Quarry Roaster    87
   
NGM Twin Creeks Autoclave    101
   
i-80 Lone Tree Autoclave (Idle)    55

Table 18-9 Operating and Capital Costs

       
Category   

  Total Cost  

($M)

    $/ton Processed       $/Au oz  
       
Mining    408   139   552
       
Transportation and Processing    270   91   365
       
Electrical Power    71   24   96
       
G&A, Royalties, and NV Taxes    138   49   186
       
By Product Credits    (3)   (1)   (4)
       
Total Operating Costs    883   303   1,194
       
Closure and Reclamation    31   11   42
       
Sustaining Capital    49   17   66
       
All in Sustaining Costs    963   330   1,303
       
Federal Income Tax    83   28   112
       
Construction Capital    157   54   213
       
All in Cost Excluding Pre-Construction Capex    1209   706   1,628
       
Pre-Construction Capex    17   6   23
       
All in Costs    1226   401   1,651

 

  18.4.

Cutoff Grade

Cut off grades vary for each mineralized lens depending on process and lens specific recovery. Cutoff grade calculations for both processes are shown in. (Table 18-10)

Table 18-10 Cutoff Grades for Alkaline POX and Roaster

     
      Alkaline POX    Roaster

Gold Price ($/oz)

   $2,175

Nevada Commerce and Excise Tax

   1.151%

Refining and Sales ($/oz)

   $1.85

Royalty

   2%

Recovery1

   78.5 – 95.1%    73.2 – 93.3%

 

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 i-80 Gold Corp   Capital and Operating Costs   Page 205 

 

 

     
      Alkaline POX    Roaster

Process Capacity (tpd)

   2,500    750

Mine Capacity (tpd)

   1,300

Mining Costs ($/ton)

   $139.11

Haulage Cost

   $14.71    $23.94

Process Cost

   $70.81    $75

Incremental Cutoff Grade (opt)

   0.043 - 0.052    0.050 – 0.064

Mine Limited Cutoff Grade (opt)

   0.112 - 0.136    0.121 – 0.155

Fixed Costs ($ /ton)

   $18.36    $61.19

Process Limited Cutoff Grade (opt)

   0.131 – 0.159    0.152 - 0.194

 

 

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  19.

Economic Analysis

The analysis of the Cove Underground project presented herein is based solely on mineral resources and not mineral reserves. Mineral resources which are not mineral reserves do not have a demonstrated economic viability. This financial analysis includes inferred mineral resources. Inferred mineral resources are too speculative to be classified as mineral reserves and the estimated quantity or grade may not be realized.

The project timeline is shown in Figure 19-1. The pre-construction work is necessary to reach a go-ahead decision. All costs during this period are being treated as sunk costs and they have been excluded from the financial analysis.

Figure 19-1 Project Timeline

LOGO

Constant dollar cash flow analysis is presented in Table 19-1 through Table 19-6 and graphically in Figure 19-3. Royalties include both the 11⁄2 % Newmont NSR and the 2% Summa Corporation NSR. The Summa royalty applies only to a portion of the mine production.

Federal income taxes of 21% apply to taxable income after appropriate deductions for depreciation and depletion. The gold percentage depletion rate is 15%. Nevada’s commerce tax is 0.051% on all revenue above 4M per annum. The excise tax is 0.75% on all revenue above $20M and less than $150M and 1.1% on revenues over $150M.

 

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Table 19-1 Income Statement Includes Inferred) (Millions $US except Unit Cost per Ounce)

         
     

Pre-

Const.

   Construction   Production     
  

 2025- 

 2027 

    2028      2029     2030     2031     2032     2033     2034     2035   

 2036- 

 2037 

   Total 
         
Gold Sales    0    0    87   99   151   180   392   325   222   154   1,609
         
Silver Sales    0    0    0   0   0   0   1   1   1   0   3
         
Total Revenue    0    0    87   99   151   180   393   325   222   155   1,612
         
Mining Cost    0    0    (26)   (24)   (67)   (67)   (67)   (58)   (61)   (38)   (816)
         
Haulage and Processing    0    0    (15)   (16)   (27)   (28)   (60)   (59)   (38)   (25)   (536)
         
Electrical Power    0    0    (8)   (9)   (9)   (8)   (9)   (9)   (9)   (8)   (71)
         
Site Administration    0    0    (7)   (7)   (7)   (7)   (11)   (11)   (7)   (7)   (65)
         
Refining and Sales    0    0    (0)   (0)   (0)   (0)   (0)   (0)   (0)   (0)   (1)
         
Royalties    0    0    (2)   (2)   (1)   (3)   (6)   (5)   (5)   (4)   (27)
         
Nevada Net Proceeds    0    0    (1)   (2)   (1)   (3)   (11)   (8)   (4)   (3)   (34)
         
Total Cash Cost    0    0    (59)   (60)   (115)   (119)   (170)   (154)   (127)   (86)   (891)
         
Cash Cost per Ounce1 ($/oz)    0    0    1,482   1,327   1,658   1,435   936   1,027   1,239   1,212   1,201
         
EBITDA    0    0    28   38   36   61   223   171   95   68   720
         
Reclamation Accrual    0    0    (2)   (2)   (3)   (3)   (8)   (6)   (4)   (3)   (31)
         
Depreciation    0    0    (13)   (16)   (25)   (32)   (70)   (64)   (44)   (30)   (295)
         
Total Cost    0    0    (74)   (78)   (143)   (154)   (247)   (224)   (175)   (120)   (1,217)
         
Income Tax    0    0    (3)   (5)   (2)   (6)   (32)   (22)   (11)   (7)   (83)
         
Net Income    0    0    10   16   6   20   113   79   37   28   312

Net of Byproduct Sales

Table 19-2 Cash Flow Statement (Includes Inferred)

         
     

Pre-

Const.

  Construction   Production     
         
     

 2025- 

 2027 

   2028     2029     2030     2031     2032     2033     2034     2035      2036     2037   

 2038- 

 2045 

   Total 
         
Net Income    0   0   10   16   6   20   113   79   37    28   0   3   312
         
Depreciation    0   0   13   16   25   32   70   64   44    30   0   0   295
         
Reclamation    (0)   (1)   1   1   2   3   7   5   3    (2)   (5)   (15)   0
         
Working Capital    0   0   (7)   (0)   (6)   (0)   (6)   2   3    5   10   0   0
         
Operating Cash Flow    (0)   (1)   17   33   27   54   185   150   87    61   6   (12)   606
         
Capital Costs    (17)   (105)   (52)   (15)   (7)   (14)   (1)   (12)   0    0   0   0   (224)
         
Net Cash Flow    (18)   (106)   (35)   18   20   40   184   138   87    61   6   (12)   382
         
All in Cost 1,2 ($/oz)        0   2,893   1,804   1,828   1,723   1,159   1,300   1,387    1,353   0   0   1,658
  1.

Net of Byproduct Sales

  2.

Note: All in Cost Exclusive of Corporate Taxes, Exploration, Corporate G&A and Interest on Debt

 

 

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Table 19-3 Income Statement (Excludes Inferred) (Millions $US except Unit Cost per Ounce)

 

         
     

Pre-

Const.

   Construction   Production     
  

 2025- 

 2027 

    2028      2029     2030     2031     2032     2033     2034     2035   

 2036- 

 2037 

   Total 
         
Gold Sales    0    0    26   29   45   41   72   60   31   17   322
         
Silver Sales    0    0    0   0   0   0   0   0   0   0   0
         
Total Revenue    0    0    26   29   45   41   72   60   31   17   323
         
Mining Cost    0    0    (7)   (6)   (18)   (12)   (5)   (2)   (3)   (0)   (106)
         
Haulage and Processing    0    0    (5)   (5)   (8)   (7)   (11)   (11)   (5)   (3)   (109)
         
Electrical Power    0    0    (8)   (9)   (9)   (8)   (9)   (9)   (9)   (8)   (71)
         
Site Administration    0    0    (7)   (7)   (7)   (7)   (11)   (11)   (7)   (7)   (65)
         
Refining and Sales    0    0    (0)   (0)   (0)   (0)   (0)   (0)   (0)   (0)   (0)
         
Royalties    0    0    (0)   (1)   (0)   (0)   (1)   (1)   (1)   (0)   (4)
         
Nevada Net Proceeds    0    0    0   0   0   0   (1)   (1)   0   0   (2)
         
Total Cash Cost    0    0    (27)   (28)   (43)   (35)   (39)   (35)   (25)   (19)   (252)
         
Cash Cost per Ounce1 ($/oz)    0    0    2,267   2,064   2,059   1,854   1,178   1,266   1,714   2,404   1,697
         
EBITDA    0    0    (1)   2   2   6   33   25   7   (2)   71
         
Reclamation Accrual    0    0    (3)   (3)   (4)   (4)   (7)   (6)   (3)   (2)   (31)
         
Depreciation    0    0    (20)   (24)   (38)   (37)   (65)   (61)   (32)   (18)   (295)
         
Total Cost    0    0    (50)   (55)   (85)   (76)   (111)   (103)   (59)   (38)   (578)
         
Income Tax    0    0    0   0   0   0   0   0   0   0   4
           
Net Income    0    0    (23)   (25)   (40)   (35)   (39)   (42)   (28)   (21)   (251)

Net of Byproduct Sales

Table 19-4 Table 19-5 Cash Flow Statement (Excludes Inferred)

         
     

Pre-

Const.

  Construction   Production     
         
     

 2025- 

 2027 

   2028     2029     2030     2031     2032     2033     2034     2035     2036     2037   

 2038- 

 2045 

   Total 
         
Net Income    0   0   (23)   (25)   (40)   (35)   (39)   (42)   (28)   (21)   0   3   (251)
         
Depreciation    0   0   20   24   38   37   65   61   32   18   0   0   295
         
Reclamation    (0)   (1)   2   2   4   3   6   5   2   (3)   (5)   (15)   0
         
Working Capital    0   0   (3)   (0)   (2)   1   (0)   0   1   1   2   0   (0)
         
Operating Cash Flow    (0)   (1)   (5)   1   (0)   6   32   25   7   (6)   (2)   (12)   44
         
Capital Costs    (17)   (105)   (52)   (15)   (7)   (14)   (1)   (12)   0   0   0   0   (224)
         
Net Cash Flow    (18)   (106)   (57)   (15)   (7)   (8)   31   13   7   (6)   (2)   (12)   (180)
         
All in Cost 1,2 ($/oz)        0   6,842   3,394   2,600   2,828   1,410   1,910   1,925   2,615   0   0   3,389
  1.

Net of Byproduct Sales

  2.

Note: All in Cost Exclusive of Corporate Taxes, Exploration, Corporate G&A and Interest on Debt

 

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The Helen and Gap zones contain 70% and 89% inferred mineral resources respectively. The results without inferred are the result of a gross factorization of the production stream. There has been no adjustment to capital development, dewatering capital or mine facilities capital. Furthermore, there has not been any recalculation of productivities or operating costs due to the lower production rates.

Table 19-6 Financial Statistics1

     
      With Inferred    Without Inferred
   

Gold price (US$/oz)

   $2,175
   

Silver price (US$/oz)

   $27.25
   

Mine life (years)

   8
     

Average mineralized mining rate (tons/day)

   1,010    200
     

Average grade (oz/t Au)

   0.305    0.313
     

Average gold recovery (roaster %)

   79%    79%
     

Average gold recovery (autoclave %)

   86%    86%
     

Average annual gold production (koz)

   92    19
     

Total recovered gold (koz)

   740    148
     

Pre-development capital ($M)

   $17.3    $17.2
     

Mine construction capital ($M)

   $157.4    $157.4
     

Sustaining capital (M$)

   $49.1    $49.1
   

Construction Start Date

   1/1/2028
 
Economic Indicators Post Construction Decision 3
     

Cash cost (US$/oz) 1

   $1,194    $1,697
     

All-in sustaining cost (US$/oz)2

   $1,303    $2,240
     

All in cost (US$/oz) 5

   $1,635    $3,302
     

Project after-tax NPV5% (M$)

   $271    ($160)
     

Project after-tax NPV8% (M$)

   $216    ($159)
     

Project after-tax IRR

   30%    NA
     

Payback Period

   5.5 Years    NA
     

Profitability Index 5%4

   2.4    0.2
 
Financial Statistics Including Pre-Construction Period (1/1/2025 – 12/31/2027) 6
     

All in cost (US$/oz) 5

   $1,658    $3,419

 

 

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      With Inferred    Without Inferred
     

Project after-tax NPV5% (M$)

   $256    ($178)
     

Project after-tax NPV8% (M$)

   $198    ($177)
     

Project after-tax IRR

   25%    NA
     
Payback Period    6.8 years    NA
     

Profitability Index 5%4

   1.9    0.2

Notes:

  1.

Net of byproduct sales;

  2.

Excluding income taxes, construction capital, corporate G&A, corporate taxes and interest on debt;

  3.

Discounted to 2028, Construction Start;

  4.

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates breakeven;

  5.

Excluding corporate G&A, corporate taxes and interest on debt;

 

  6.

Discounted to 2025;

  7.

This IA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the IA will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability;

  8.

Inferred mineral resources constitute 70 % of the Helen zone and 89% of the Gap zone. The “Without Inferred” statistics presented are a gross factorization of the mine plan without any redesign of mine excavations or recalculation of productivities and costs. Capital costs are the same for the “With Inferred” and “Without Inferred” scenarios. The “Without Inferred” scenario is presented solely to illustrate the projects dependence on inferred mineral resources.

  9.

The financial analysis contains certain information that may constitute “forward-looking information” under applicable Canadian and United States securities regulations. Forward-looking information includes, but is not limited to, statements regarding the Company’s achievement of the full-year projections for ounce production, production costs, AISC costs per ounce, cash cost per ounce and realized gold/silver price per ounce, the Company’s ability to meet annual operations estimates, and statements about strategic plans, including future operations, future work programs, capital expenditures, discovery and production of minerals, price of gold and currency exchange rates, timing of geological reports and corporate and technical objectives. Forward-looking information is necessarily based upon a number of assumptions that, while considered reasonable, are subject to known and unknown risks, uncertainties, and other factors which may cause the actual results and future events to differ materially from those expressed or implied by such forward looking information, including the risks inherent to the mining industry, adverse economic and market developments and the risks identified in Premier’s annual information form under the heading “Risk Factors”. There can be no assurance that such information will prove to be accurate, as actual results and future events could differ materially from those anticipated in such information. Accordingly, readers should not place undue reliance on forward-looking information. All forward-looking information contained in this Presentation is given as of the date hereof and is based upon the opinions and estimates of management and information available to management as at the date hereof. Premier disclaims any intention or obligation to update or revise any forward-looking information, whether as a result of new information, future events or otherwise, except as required by law;

 

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 i-80 Gold Corp   Economic Analysis   Page 211 

 

Figure 19-2 Gold Production and Unit Costs (With Inferred)

 

LOGO

Figure 19-3 Cash Flow Waterfall Chart Including Pre-Construction Costs (With Inferred)

 

LOGO

 

 

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Figure 19-4 Gold Production and Unit Costs (Without Inferred)

 

LOGO

Figure 19-5 Cash Flow Waterfall Chart Including Pre-Construction Costs (Without Inferred)

 

LOGO

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Economic Analysis   Page 213 

 

Figure 19-6 NPV 5% Sensitivity

 

LOGO

Figure 19-7 Profitability Index 5% Sensitivity

 

LOGO

 

 

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Figure 19-8 IRR Sensitivity

 

LOGO

 

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 i-80 Gold Corp   Adjacent Properties   Page 215 

 

20. Adjacent Properties

There are no adjacent properties with a similar geologic setting to the McCoy/Cove project.

 

 

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21. Other Relevant Data and Information

The authors are not aware of any other relevant technical data or information pertaining to the Cove Project necessary to make this Technical Report Summary understandable and not misleading.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Interpretation and Conclusions   Page 217 

 

22. Interpretation and Conclusions

The Cove Project is in a politically stable mining friendly jurisdiction with a long history of Mineral Resource extraction. The Project is potentially economic. Results from this IA indicate a post construction decision NPV 5% of $274M (excluding $17.3M pre-development capital) and an IRR of 30%. The project should proceed with the completion of environmental baseline studies, permitting activities, underground delineation diamond drilling, and feasibility study in support of a construction decision.

Metallurgical Testing

 

  1.

Head assays for the both the Helen Zone and Gap indicated that the gold in the two resources will likely be finely disseminated and not amenable to gravity gold recovery;

 

  2.

The mineralogy of the Helen and Gap resources differ in two significant areas, the first being that the Helen appears to be lower in arsenic content than the Gap resource and that the Gap resource appears to be lower on average in TCM and TOC than the Helen resource;

 

  3.

The Helen composite arsenic assays indicate the resource is lower in arsenic content that the Gap resource;

 

  4.

The Helen and Gap resources based on the composites tested appear to be doubly refractory to conventional cyanidation and require both sulfide oxidation and passivation of active carbonaceous mineralization to significantly increase gold extractions;

 

  5.

Based on the composites tested the Helen Zone appears to generally be more amenable to roasting and CIL processing;

 

  6.

Based on the composites tested, the Gap resource appears to generally be more amenable to pressure oxidation and CIL processing;

 

  7.

The data set was too small to establish any clear relations between mineralogy and metal head grade and extractions for either resource although it is clear that mineralogy factors such as arsenic content and TCM or TOC are influencing extractions using either roasting and calcine cyanidation or pressure oxidation and residue cyanidation.

Toll Processing

 

  1.

The feed specifications appear to be somewhat rigid and could preclude some material being sent to the toll processor. Blending may allow shipment of some off-specification material provided appropriate material is available for onsite blending prior to shipping to the toll processor;

 

  2.

The terms appear to be consistent and typical with those encountered in the industry; and

 

 

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  3.

The recovery terms appear to be the result of analyzing the metallurgical data provided by i-80 Gold.

Mining and Infrastructure

 

  1.

Mining conditions typical for sedimentary deposits in the north-east Nevada extensional tectonic environments are anticipated;

  2.

Helen Zone dewatering will require five wells and reach pumping rates of 10,500 gpm; and

  3.

Gap Zone dewatering will require ten wells and reach pumping rates of 26,000 gpm for a total projected pumping rate of 36,500 gpm.

Financials

The financial analysis presented in this IA is an evaluation of the Cove Mineral Resource. Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. This financial analysis includes inferred mineral resources. Inferred mineral resources are too speculative to be classified as mineral reserves and the quantity or grade estimated may not be realized.

 

  1.

Capital requirements total $206.5M excluding $17.3M in pre-construction capital;

  2.

The project achieves NPV 5% of $273.7M and NPV 8% of $215.6M;

  3.

When including the pre-construction capital, the NPV 5% reduces to $255.9M and the NPV 8% is $197.8M; and,

  4.

The estimated payback period is 6.8 years with an IRR of 30%.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Interpretation and Conclusions   Page 219 

 

 

  22.1.

Risks and Opportunities

The authors have identified the following risks and opportunities to the project.

Table 22-1 Project Risks

 

Risks

 

Impact

 

Mitigation Measure

   
Agencies may identify deficiencies in the baseline environmental data   Project delays   Proceed with baseline data collection and engineering to support both possibilities
   
Dewatering rates may increase   Additional facilities required   Additional hydrology testing and modeling
   
Water quality levels above Tier I standards for infiltration   Water treatment required, increased capital costs   Geochemical study of RIBs to ascertain the possibility of attenuation
   
Water rights Availability   Project delays and increased costs   Continue water rights acquisition and seek agreements with local ranches

Table 22-2 Opportunities

 

Opportunities

 

Impact

   
Senior level government initiative to streamline the permitting process   Earlier production and increased NPV
   
Resource additions in the Gap Extension area   Increased ounce production and improved project economics
   
2201 Zone could add higher grade mineralization to the mine plan utilizing common infrastructure   Increased ounce production and improved project economics
   
Develop grade-thickness mineralization model   Optimize mine design

 

  22.2.

Work Program

Activities at Cove are structured to complete resource definition drilling to a level that will support the feasibility study. Secondly the program will advance the project to a record of decision on the environmental impact statement. Lastly the program will complete the feasibility study and make a recommendation on the construciton decision. Program costs are listed in Table 22-3.

 

 

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Table 22-3 Work Program Estimated Costs (US$M)

 

 

Description

  2025     2026     2027     Total  
       

Environmental, Permitting and Feasibility

  4.0     2.0     1.0     7.0  
       

Preproduction Expense

  1.7     1.7     1.7     5.1  
       

Resource Conversion Drilling

  2.0     -          2.0  
       

Contingency

  1.7     0.9     0.7     3.3  
                 
       

Program Total

  9.4     4.6     3.4     17.3  

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   Recommendations   Page 221 

 

23. Recommendations

The project pre-feasibility or feasibility study should address the following components. The work should be planned to minimize the permitting time required to achieve positive cash flow.

Resource Delineation and Exploration

 

  1.

Completion of the underground drilling program;

  2.

Update the resource model;

  3.

Expansion of the Gap and 2201 Zones could add high grade mineralization to the project which would be accessed through the planned Helen and Gap infrastructure.

Dewatering

 

  1.

Detailed review and selection of well locations, and;

  2.

Fine tune hydrogeologic model, pumping rates and drawdown rates.

Mining

 

  1.

A geotechnical characterization program should be implemented along with resource delineation:

  a.

The objectives of the program are to characterize the mining horizons using the Rock Mass Rating (RMR) system;

  b.

Collect downhole Acoustic Tele Viewer (ATV) drill logs to collect joint orientation data for mine designs and accurately estimate ground support requirements; and,

  c.

Collect full core samples for physical rock property testing.

  2.

Complete additional testing of potential back fill sources to optimize the Cemented Rock Fill (CRF) mix design;

  3.

Run trade off studies between aquifer pumping rates and alternative mining scenarios; and,

  4.

Complete a ventilation simulation to predict Diesel Particulate Matter (DPM), carbon monoxide, and other contaminate concentrations.

Metallurgical Testing

 

  1.

Additional metallurgical testing will be needed to thoroughly investigate the variability and viability of Helen and Gap resources to evaluate pressure oxidation with CIL cyanidation under Lone Tree conditions. Testing should also include baseline CIL tests and roasting testing as a comparison. Sampling objectives will include:

 

 

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Samples from GAP and Helen zones and their major lithological units; Favret, Panther Dolomite. The samples should also address spatial variability within each zone.

 

   

Sample intrusive formations in each zone.

 

   

Assess variability of the responses to roasting and calcine cyanidation across the resources;

 

   

Assess variability of the responses to pressure oxidation and residue cyanidation across the resources;

 

   

Testing should attempt to establish head grade and extraction relations for use in more detailed resource modelling;

 

   

Mineralogy impacts need to be established and geologic domains within each resource need to be determined, and;

 

   

Additional comminution data should be collected to assess hardness variability within the zones and any potential impacts on throughput in the Lone Tree process plant.

 

  2.

The resource model should be advanced to include arsenic, TCM, TOC, mercury, lead, zinc, total copper selenium, barium, cobalt, nickel, and cadmium as these will be important for predicting grades if toll process offsite is used and potentially for estimating extractions within the resources;

  3.

The estimated cost for the suggested next phase metallurgical program is to $850,000 based on current market pricing.

  5.

Development of a preliminary or conceptual onsite blending program is recommended to evaluate if on specification material can consistently be supplied to a toll processor; and,

Permitting and Development Decision

 

  1.

Baseline data collection in support of the Helen EA and GAP EIS should be done simultaneously to reduce the Project’s critical path and bring forward production; and.

  2.

The project should proceed directly with a feasibility study to support a construction decision.

 

 Practical Mining LLC   March 26, 2025 


 i-80 Gold Corp   References   Page 223 

 

24. References

Amendment to Minerals Lease and Agreement between Newmont McCoy Cove Limited and Victoria Resources (US) Inc, September 29, 2008.

Briggs, D. F., McCoy-Cove Complex: Mining operations report prepared by Geomineinfo, 2001.

Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014.

Emmons, D. L., and Eng, T. L., Geology and Mineral Resources of the McCoy Mining District, Lander County, Nevada: Text to accompany Nevada Bureau of Mines Map 103, 1995.

Exhibit A to McCoy Cove Earn in Agreement between Premier Gold Mines Ltd. And Barrick Gold Corp., January 10, 2018

Hydrologic Consultants, Inc. 1990. Conceptual hydrogeologic model of Cove pit area as of June, 1990. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1992. Updated conceptual hydrogeologic model and estimates of future dewatering costs as of September 1992. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1993. Status report Cove pit dewatering program. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1994. Updated conceptual hydrogeologic model and estimate of future dewatering costs for Cove pit. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1995. Updated conceptual hydrogeologic model and estimate of future dewatering costs for Cove pit as of April 1995. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1997a. Hydrogeologic framework and numerical ground-water flow modeling of McCoy/Cove Mine, Lander county, Nevada. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1997b. Updated conceptual hydrogeologic model and predicted dewatering requirements for remaining life of McCoy/Cove mine. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 1999. 1999 update of hydrogeologic framework and numerical ground-water flow modeling of McCoy/Cove Mine, Lander County, Nevada. Hydrologic Consultants, Inc, Denver.

Hydrologic Consultants, Inc. 2001. 2001 update of numerical ground-water flow modeling for McCoy/Cove mine, Lander county, Nevada. Hydrologic Consultants, Inc, Denver.

Itasca 2016. Numerical groundwater model and predictions of Cove pit-lake infilling, McCoy Cove Mine. Itasca, Denver.

Johnston, M. K., Geology of the Cove Mine, Lander County, Nevada, and a Genetic Model for the McCoy-Cove Magmatic-Hydrothermal System, University of Nevada, Reno, Ph.D. dissertation, May 2003.

 

 

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John, D.A., Henry, C.D., and Colgan, J.P., Magmatic and tectonic evolution of the Caetano caldera, north-central Nevada: A tilted, mid-Tertiary eruptive center and source of the Caetano Tuff, Geosphere, v.4, no. 1, 2008

Kappes Cassiday & Associates, Cove Project Report of Metallurgical Test Work, December 2008.

Kappes Cassiday & Associates, Cove Project Report of Metallurgical Test Work, August 2009.

Kappes Cassiday & Associates, Cove Project Report of Metallurgical Test Work, November 2009.

Kuyper, B. A., Mach, L. E., Streiff, R. E., and Brown, W. A., Geology of the Cove Gold-Silver Deposit: Society for Mining, Metallurgy, and Exploration, Inc., 1991.

Madrid, R. J., Anatomy of the Helen Gold System, a Carlin –Type Intersection Zone, McCoy-Cove Mining District, North Central Nevada. Victoria Resources, 2009.

McDonald, M.G, and A.W. Harbaugh 1988. Techniques of water-resources investigations of the United States Geological Survey, Chapter A1, A modular three-dimensional finite-difference ground-water flow model. USGS.

Memorandum of Agreement between Newmont Corporation and Victoria Resources (US) Inc., June 15, 2006.

Minerals Lease and Agreement between Newmont McCoy Cove Limited and Victoria Resources (US) Inc June 15, 2006.

Mining Deed from Echo Bay Exploration Inc. to Newmont Mining Corporation, February 7, 2003.

Mine Development Associates, Technical Report Cove Project Lander County, Nevada, U.S.A., October 24, 2008.

Montgomery and Associates, Report Summary of Field Program at Helen and Gap Deposits, Cove Helen Underground, April 10, 2020.

Montgomery and Associates 2023a. Baseline spring and seep survey. Montgomery and Associates, Reno. January 2023.

Montgomery and Associates 2023b. Flow modeling for the McCoy-Cove Project. Montgomery and Associates, Reno. December 2023.

Montgomery and Associates 2024a. 2023 Baseline seeps, springs, and streams survey report. Montgomery and Associates, April 2024.

Montgomery and Associates 2024b. Flow modeling for the Cove Gap/Helen Project. Montgomery and Associates Reno.

Montgomery and Associates 2024c. Analysis of 2023 long-term, multi-well aquifer test, October 2023 through November 2023. Montgomery and Associates, Reno. March 2024.

 

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 i-80 Gold Corp   References   Page 225 

 

Montgomery and Associates 2025. 2024 Baseline seeps, springs, and streams survey report. Montgomery and Associates, January 2025.

Piteau Associates USA, Cove Helen Hydrogeologic Characterization, Numerical Model Update and Preliminary Dewatering Plan, Luther, A., March 23, 2018.

Quantum Electric, Cove Helen – Electrical System Report, Elquist, J. A. August 8, 2017.

Purchase Agreement Between Newmont USA Ltd. And Premier Gold Mines Ltd., July 31, 2014

Roscoe Postle Associates Inc., Preliminary Assessment of the Cove Project, Nevada, Prepared by Valliant, W., Evans, L., and Bergen, R.D., for Premier Gold Mines Limited, October 3, 2012.

Roscoe Postle Associates Inc., Technical Report on the McCoy-Cove Gold Project, Lander County, State of Nevada, U.S.A., Evans, L., and Tudorel, C., April 15, 2017.

Royalty Deed from The Howard Hughes Corporation to Echo Bay Inc., December 17, 2007.

Second Amendment to Minerals Lease and Agreement between Newmont McCoy Cove Limited and Victoria Resources (US) Inc, March 29, 2009.

Silberling, N. and Roberts, R.J., Pre-Tertiary stratigraphy and structure of northwestern Nevada: Geol. Soc. America Special Paper 72, 58 p., 1962

Struhsacker, D. W., Overview of Permitting Requirements for Mineral Projects in Nevada, July 2009.

United States Securities and Exchange Commission, Rule 1300 of Regulation S-K promulgated under the U.S. Securities Act of 1933, as amended, Federal Register, Vol. 83, No. 246, December 26, 2018.

Wolverson, N. J., Review of Cove-McCoy Drilling QA/QC and Procedures, December 9, 2007.

 

 

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25. Reliance on Information Provided by the Registrant

The Consultant’s opinion contained herein is based on information provided to the Consultants by i-80 throughout the course of the investigations. Table 24-1 of this section of the Technical Report Summary will:

 

  1)

Identify the categories of information provided by the registrant;

  2)

Identify the particular portions of the Technical Report Summary that were prepared in reliance on information provided by the registrant pursuant to Subpart 1302 (f)(1), and the extent of that reliance; and

  3)

Disclose why the qualified person considers it reasonable to rely upon the registrant for any of the information specified in Subpart 1302 (f)(1).

Table 24-1 Reliance on Information Provided by the Registrant

 

Category    Section    Portion of Technical Report
Summary
   Disclose why the Qualified Person considers it
reasonable to rely upon the registrant
   
Claims List    3.2    Mineral Title    i-80 provided PM with a current listing of claims. The information was sourced from i-80’s Land Manager and is backed by the Purchase Agreement Between Newmont USA Ltd. And Premier Gold Mines Ltd., July 31, 2014.
   
Holding Costs    3.3    Property Holding Costs    Property holding costs are calculated with appropriate rates from the BLM and County. Property taxes are verified with online county records.
   
Environmental Liabilities    3.4    Environmental Liabilities    i-80 provided Practical with the status of current environmental liabilities and the extent of planned future liabilities. These are consistent with other mines in the area.
   
Permits    3.5    Permits and Licenses    i-80 provided the requirements for existing and future permits. These are similar to other mines in the area,
   
Hydrogeology Testing and Modelling    7.2    Hydrogeology    i-80 Provided Internal studies completed by a hydrogeology consulting firm.
   
Marketing Studies    16.1    Precious Metal Markets    I-80 provided Practical with precious metal market research and forecast from CIBC. This information was used to select the long term pricing for the estimation of mineral resources.

 

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 i-80 Gold Corp   Reliance on Information Provided by the Registrant   Page 227 

 

 

Category    Section    Portion of Technical Report
Summary
   Disclose why the Qualified Person considers it
reasonable to rely upon the registrant
   
Material Contracts    16.3    Previous Financing Agreements    Information on financing agreements is consistent w
   
Material Contracts    16.4    Roaster Toll Milling Agreement    i-80 provided Practical with copies of the existing agreement.
   
Current Environmental Programs    17    Environmental Studies, Permitting and Plans, Negotiations or Agreements with Local Individuals or Groups    i-80 Provided the status of current environmental programs, permitting and reclamation requirements. Practical believes these to be consistent with the requirements of other operations in the area.

 

 

 Practical Mining LLC   March 26, 2025