Exhibit 99.1

SOUTH RAILROAD PROJECT

FORM 43-101F1 TECHNICAL REPORT


	22.8	ROYALTIES	412
	22.9	GOVERNMENT FEES	413
	22.10	INCOME TAX	413
	22.11	NET INCOME AFTER TAX	413
	22.12	PROJECT FINANCING	413
	22.13	ECONOMIC INDICATORS	413
	22.14	SENSITIVITY ANALYSIS	413
	22.15	DETAILED FINANCIAL MODEL	414
23	ADJACENT PROPERTIES		418
	23.1	RAIN	418
	23.2	EMIGRANT	419
	23.3	PONY CREEK PROPERTY	419
24	OTHER RELEVANT DATA AND INFORMATION		420
25	INTERPRETATION AND CONCLUSIONS		421
	25.1	PROJECT RISKS	421
	25.2	PROJECT OPPORTUNITIES	422
	25.3	EXPLORATION AND MINERAL RESOURCE EXPANSION	422
26	RECOMMENDATIONS		423
	26.1	EXPLORATION	423
	26.2	INFILL DRILLING	423
	26.3	EXPLORATION AND EXPANSION DRILLING	424
	26.4	CONDEMNATION DRILLING	424
	26.5	METALLURGICAL TEST WORK	424
	26.6	PERMITTING AND BASELINE STUDIES	424
	26.7	ENGINEERING STUDIES AT FEASIBILITY LEVEL	424
	26.8	HYDROLOGY	425
	26.9	GEOTECHNICAL SURVEY	425
	26.10	TOTAL COST OF RECOMMENDED STUDY PROGRAM	425
27	REFERENCES		426


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FORM 43-101F1 TECHNICAL REPORT

LIST OF FIGURES AND ILLUSTRATIONS


FIGURE	DESCRIPTION	PAGE
Figure 1-1: 2016-2017 Gold Extraction vs. Days Under Leach for Column-Leach Tests		6
Figure 1-2: 2019 Pinion Gold Extraction vs. Days under Leach for Column-Leach Tests		7
Figure 1-3: Conventional Crush vs. HPGR Gold Extraction Comparison		7
Figure 1-4: 2017 Dark Star Column-Leach Gold Extraction vs. Days under Leach		9
Figure 1-5: 2019 Dark Star Column-Leach Gold Extraction vs. Days under Leach		10
Figure 1-6: Conventional Crush vs. HPGR Gold Extraction Comparison, Dark Star Main		11
Figure 4-1: Location Map for the Railroad-Pinion Property		28
Figure 4-2: Railroad-Pinion Property with Ownership Percentages, Elko County, Nevada		28
Figure 4-3: Railroad-Pinion Property Map with Royalty Encumbrances		30
Figure 4-4: Property Map with Railroad- Pinion Permit Boundaries		32
Figure 7-1: Regional Geology of the Railroad-Pinion Property		48
Figure 7-2: Long Section through the Carlin Trend		49
Figure 7-3: Gold Standard Property Geologic Map		51
Figure 7-4 North Bullion Stratigraphic Column		53
Figure 7-5: Stratigraphic Column for the Pinion, Dark Star, and Jasperoid Wash Deposit Areas		55
Figure 7-6: North Bullion Cross Section N4488800		58
Figure 7-7: Pinion Deposit Geology, Section 4479380N		60
Figure 7-8: Dark Star Geologic Cross Section N4479600		61
Figure 7-9: Jasperoid Wash Geologic Cross Section 4473200N		62
Figure 8-1: Regional-Scale Carlin-Type Deposit Model		65
Figure 9-1: Ground-based Geophysical Surveys by Gold Standard 2009 to 2015		67
Figure 9-2: Rock and Soil Sample Locations 2010 - 2018		70
Figure 10-1: Railroad-Pinion Drill Hole Map (1969 – 2018)		73
Figure 10-2: Map of North Railroad Property Drill Collar Locations		81
Figure 12-1: Dark Star Assay Comparison - AAL vs. MBA - 1991 CDS Holes		102
Figure 12-2: Dark Star Assay Comparison - AAL vs Actlabs - 1991 CDS Holes		102
Figure 12-3: Dark Star Check Assays – ALS Assay vs. Bureau Veritas (Inspectorate), 2015		107
Figure 12-4: Dark Star Check Assays - ALS Assay vs. Bureau Veritas (Inspectorate) 2016		109
Figure 12-5: Scatter Plot of Twin-Hole Analysis – DC18-09 (core) vs DR18-44 (RC)		113
Figure 12-6: Control Chart for MEG-Au.11.34		116
Figure 12-7: Control Chart for MEG-S107007X		117


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FORM 43-101F1 TECHNICAL REPORT


Figure 12-8: Control Chart for MEG-Au.11.19 - 2018	120
Figure 12-9: Grade Ranges and Dates of 2018 Pinion CRMs	120
Figure 12-10: Gold Relative Percent Difference – Pinion Duplicate vs. Original	123
Figure 12-11: Gold Relative Percent Difference – ALS vs. Bureau Veritas, 2017 Pulps	124
Figure 12-12: Gold in Blanks and Preceding Samples 2014	126
Figure 12-13: Gold in Blanks and in Preceding Samples 2015	126
Figure 12-14: Gold in Pulp Blanks and in Preceding Samples 2017 – 2018	127
Figure 12-15: Gold in Coarse Blanks and in Preceding Samples 2017 – 2018	128
Figure 12-16: Histogram of 2018 Twin Drill-Hole Samples	129
Figure 12-17: Analyses of CRMs in North Railroad Drilling 2010 to 2015	132
Figure 12-18: Analytical Results from Standard MEG-S107007X	133
Figure 12-19: Analytical Results from Standard MEG-Au.11	133
Figure 12-20: 2010 to 2015 North Railroad Portion Field Duplicate Assays	134
Figure 13-1: Plot of Column P80 (microns) vs. Gold Extraction (%)	136
Figure 13-2: Pinion Zone Location Map for 2015 – 2016 Metallurgical Composites	139
Figure 13-3: 2016 – 2017 Pinion Metallurgical Core Hole Locations	141
Figure 13-4: 2016 – 2017 Gold Extraction vs. Days Under Leach for Column-Leach Tests	144
Figure 13-5: Conventional Crush vs. HPGR Gold Extraction Comparison	148
Figure 13-6: Conventional Crush vs. HPGR Silver Extraction Comparison	148
Figure 13-7: Pinion Deposit Metallurgical Core Location Map	150
Figure 13-8: 2019 Pinion Gold Extraction vs. Days under Leach for Column-Leach Tests	153
Figure 13-9: RC Drill Hole Locations for the 1991 Dark Star Bottle-Roll Tests	155
Figure 13-10 Location Map for 2017 Dark Star Metallurgical Composites	157
Figure 13-11: 2017 Dark Star Column-Leach Gold Extraction vs. Days under Leach	159
Figure 13-12: 2018 Dark Star Main - Conventional Crush vs. HPGR Gold Extraction	162
Figure 13-13: Dark Star North - Conventional Crush vs. HPGR Gold Extraction	163
Figure 13-14: Location Map for 2017-8 Dark Star Metallurgical Composites	165
Figure 13-15: 2019 Dark Star Column-Leach Gold Extraction vs. Days under Leach	168
Figure 13-16: Dark Star Composite #9C:P80 vs. Au Extraction (%)	171
Figure 13-17: Example Pinion Feed P80 vs. S/O Ratio Plot, 70% Recovery of Total Extractable Gold	174
Figure 13-18: Example Pinion Mineral Resource S/O Ratio vs. % Recovery of Total Extractable Gold	174
Figure 13-19: Pinion Head/Tails Grade	176
Figure 13-20: P80 vs. S/O Ratio for 70% Recovery of Total Extractable Gold	177
Figure 13-21: P80 vs. S/O Ratio for 90% Recovery of Total Extractable Gold	177


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Figure 13-22: Percent Recovery of Pinion Total Extractable Gold vs. Heap-Leach Feed Size	178
Figure 13-23: Pinion mlbx Pinion West Zone - Gold Recovery Model	180
Figure 13-24: Dark Star Head Grade vs. Tails Grade Plot	182
Figure 13-25: P80 vs. S/O Ratio for 90% Recovery of Total Extractable Gold	183
Figure 13-26: Percent Recovery of Pinion Total Extractable Gold vs. Heap-Leach Feed Size	183
Figure 13-27 Dark Star North (SI<2.0) - Gold Recovery Model	186
Figure 14-1: Dark Star Deposit Drill-Hole Map and Mineral Resource Outline	196
Figure 14-2: Cumulative Probability Plot of Dark Star Gold Assays	198
Figure 14-3: Dark Star Main Zone Gold Domains and Geology – Section N4479600	201
Figure 14-4: Dark Star North Zone Gold Domains and Geology – Section N4480080	202
Figure 14-5: Dark Star Spatial Relationship Between Estimation Areas and Drill Holes	206
Figure 14-6: Dark Star Main Zone Gold Domains and Block Model – Section N4479600	212
Figure 14-7: Dark Star North Zone Gold Domains and Block Model – Section N4480080	213
Figure 14-8: Cumulative Probability Plot of Dark Star AuCN/Au Ratios	214
Figure 14-9: Dark Star Optimized Pit and Additional Mineralization	221
Figure 14-10: Pinion Deposit Drill-Hole Map and Mineral Resource Outline	224
Figure 14-11: Cumulative Probability Plot of Pinion Deposit Gold Assays	226
Figure 14-12: Pinion Gold Domains and Geology – Section N4479230	229
Figure 14-13: Pinion Estimation Areas	232
Figure 14-14: Cumulative Probability Plot of Pinion Deposit Silver Assays	234
Figure 14-15: Pinion Silver Domains and Geology – Section N4479230	236
Figure 14-16: Pinion Gold Domains and Block Model– Section N4479230	242
Figure 14-17: Pinion Silver Domains and Block Model– Section N4479230	243
Figure 14-18: Cumulative Probability plot of Barium (NITON XRF) Sample Grades at Pinion	245
Figure 14-19: Pinion Barium Domains and Geology – Section N4479230	247
Figure 14-20: Cumulative Probability Plot of Pinion AuCN/AuFA Ratios	249
Figure 14-21: Pinion Optimized Pit and Additional Mineralization	256
Figure 14-22: Jasperoid Wash Deposit Drill-hole Map and Mineral Resource Outline	260
Figure 14-23: Cumulative Probability Plot of Jasperoid Wash Gold Assays	262
Figure 14-24: Jasperoid Wash Zone Gold Domains and Geology – Section N4473200	265
Figure 14-25: Jasperoid Wash Estimation Areas and Gold Domains in Cross Section	267
Figure 14-26 Jasperoid Wash Gold Domains and Block Model – Section N4473200	270
Figure 14-27: Cumulative Probability Plot of Jasperoid Wash AuCN/AuFA Ratios	271
Figure 14-28: Jasperoid Wash Deposit Rock Type and Metallurgical Models	272


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Figure 14-29: Jasperoid Wash Optimized Pits and Additional Mineralization	274
Figure 14-30: POD Deposit Geologic Cross Section	275
Figure 14-31: Plan View of North Bullion Deposits Mineralization and Structure Model	278
Figure 14-32: POD Zone Cross Section of Wire-Frame Interpretations	279
Figure 14-33: Sweet Hollow Zone Cross Section of Wire-Frame Interpretations	279
Figure 14-34: North Bullion Zone Cross Section of Wire-Frame Interpretations	280
Figure 14-35: Probability Plot of Sample Gold Grades in the Mineralized Zones	281
Figure 14-36: North Bullion Area Constrained Mineral Resource Blocks	287
Figure 15-1: Dark Star Slope Sectors	295
Figure 15-2: Pinion Slope Sectors	296
Figure 15-3: Dark Star Pit by Pit Graph	302
Figure 15-4: Pinion Pit by Pit Graph	305
Figure 15-5: Dark Star Ultimate Pit Design	308
Figure 15-6: Pinion Ultimate Pit Design	310
Figure 15-7: Pinion Phase 1 Pit Design	311
Figure 15-8: Pinion Phase 2 Pit Design	312
Figure 16-1: Dark Star Pit Design, Year -1	320
Figure 16-2: Dark Star Pit Design, Year 1	321
Figure 16-3: Dark Star Pit Design, Year 2	322
Figure 16-4: Dark Star Pit Design, Year 3	323
Figure 16-5: Dark Star Pit Design, Year 4	324
Figure 16-6: Dark Star Pit Design, Year 5	325
Figure 16-7: Pinion Pit Design, Year 4	326
Figure 16-8: Pinion Pit Design, Year 5	327
Figure 16-9: Pinion Pit Design, Year 6	328
Figure 16-10: Pinion Pit Design, Year 7	329
Figure 16-11: Pinion Pit Design, Year 8	330
Figure 17-1: Process Flowsheet for the Pinion-Dark Star Project	341
Figure 17-2: Crushing General Arrangement	343
Figure 17-3: ADR Recovery Plant General Arrangement	348
Figure 18-1 Site Plan Drawing	355
Figure 18-2: Water Management Process Flow Diagram	359
Figure 18-3: Pipeline Plan General Arrangement	360
Figure 18-4: Stormwater Controls General Arrangement	362


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Figure 18-5: Map Showing Locations of Dewatering Wells	366
Figure 18-6: Typical Dewatering Well Head Plan General Arrangement	367
Figure 18-7: Typical Dewatering Well Head Section	368
Figure 18-8: Plot of Historic Earthquake Events and Selected Seismic Source Zones within a 500 km Radius	377
Figure 18-9: Plot of PSHA Results and Comparison with DSHA Results	379
Figure 18-10: PSHA Results and Design Response Spectra	381


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FORM 43-101F1 TECHNICAL REPORT

LIST OF TABLES


TABLE	DESCRIPTION	PAGE
Table 1-1: Key Project Data		2
Table 1-2: Head Assays and Geo-metallurgical Characterization of Pinion Samples		5
Table 1-3: Summary of Column Leach Test Results for Pinion.		6
Table 1-4: Head Assays and Geo-metallurgical Characterization of Dark Star Samples		8
Table 1-5: Summary of Column Leach Test Results for Dark Star		8
Table 1-6: Dark Star, Pinion and Jasperoid Wash Estimated Mineral Resources		14
Table 1-7: North Bullion Mineral Resources		15
Table 1-8: Total Railroad-Pinion Mineral Resources		15
Table 1-9 Proven and Probable Mineral Reserves		16
Table 1-10: Capital Expenditure Schedule		19
Table 1-11. LOM Operating Costs		19
Table 1-12: Economic Analysis Summary		20
Table 1-13: Pre-Tax Cash Flow		21
Table 2-1: List of Qualified Persons		23
Table 2-2: Acronyms and Abbreviations		24
Table 6-1 Summary of Historical Exploration, Pinion Area		39
Table 6-2 Summary of Historical Exploration in the Dark Star Area		40
Table 6-3 Historical Pinion Deposit Estimated Mineral Resources		42
Table 6-4 1994 Dark Star Historical Crown Mineral Resource Estimate		44
Table 6-5 Dark Star Deposit 1995-1996 Cyprus Mineral Resource Estimate		45
Table 6-6: POD Deposit Historical Mineral Resource Estimates 1985 - 2003		46
Table 10-1: All Railroad-Pinion Drilling 1969 – 2019		72
Table 10-2: Historical Drilling Summary		74
Table 10-3: Summary of Gold Standard Drilling 2010 – 2018		79
Table 10-4: Bald Mountain Drilling Contractors and Methods		82
Table 10-5: Gold Standard’s Dark Star Drilling Contractors and Methods		83
Table 10-6: Gold Standard Pinion Area Drilling Contractors and Methods		84
Table 12-1 MDA Verification GPS Checks of Dark Star Drill Collars		97
Table 12-2 Dark Star Carbon and Sulfur Records Checked and Analytical Procedures		97
Table 12-3 Pinion Carbon and Sulfur Records Checked and Analytical Procedures		98
Table 12-4: Summary Counts of Dark Star QA/QC Analyses		101


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Table 12-5 Summary of Dark Star Results Obtained for Certified Reference Materials, 1997	104
Table 12-6: List of Dark Star Failed Certified Reference Materials, 1997	105
Table 12-7: Summary of Dark Star Results Obtained for Certified Reference Materials, 2015	106
Table 12-8: List of Dark Star Failed Certified Reference Materials, 2015	106
Table 12-9: Summary of Dark Star Results Obtained for Certified Reference Materials, 2016	108
Table 12-10: List of Dark Star Failed Certified Reference Materials, 2016	108
Table 12-11: Summary of Dark Star Results Obtained for Certified Reference Materials, 2017	110
Table 12-12: List of Dark Star Failed Certified Reference Materials, 2017	110
Table 12-13: Summary of Dark Star Results Obtained for Certified Reference Materials, 2018	111
Table 12-14: List of Dark Star Failed Certified Reference Materials, 2018	112
Table 12-15: Summary of Dark Star Results Obtained for Standards, 2019	114
Table 12-16: Summary of Results Obtained for CRMs, 2014 – 2015	115
Table 12-17: List of Failed CRM Analyses, 2014 – 2015	115
Table 12-18: Summary of Results Pinion for CRMs, 2016	118
Table 12-19: Summary of Results for Pinion CRMs, 2017	118
Table 12-20: Summary of Results for Pinion CRMs, 2018	119
Table 12-21: List of Failed Pinion CRMs, 2018	119
Table 12-22: Summary of 2019 Analyses of Silver CRMs	121
Table 12-23 List of Failed Silver CRMs	121
Table 12-24 Summary of Results for Pinion Field Duplicates	122
Table 12-25: Summary of Results for Duplicates in Silver Re-Assays	123
Table 12-26: Summary of Results for 2018 Re-Assays of 2017 Pinion Samples	125
Table 12-27: Comparison of Original Assays and Re-Runs in Part of PIN15-14	127
Table 12-28: Results of Silver Analyses of Pulp Blanks	128
Table 12-29: Summary of Counts of Jasperoid Wash QA/QC Analyses	129
Table 12-30: Summary Counts of Jasperoid Wash QA/QC Analyses	130
Table 13-1: Summary of Metallurgical Tests Prior to Gold Standard Ventures Tests.	137
Table 13-2: Summary of Nominal Feed P80 for Column and Bottle-Roll Leach Tests	142
Table 13-3: Pinion Main Zone HPGR Composite Head Assays	146
Table 13-4: 2018 Pinion Main Zone HPGR-Crushed Bottle-Roll Results	146
Table 13-5 2018 Pinion Main Zone HPGR-Crushed Column Leach Test Results	147
Table 13-6: Example: Dark Star Composite MDS #9C Gold Extraction	172
Table 13-7: Example: Dark Star Gold Extraction Model Results	172
Table 13-8: Modeled Pinion S/O Ratios at Various P80 Particle Size	178


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FORM 43-101F1 TECHNICAL REPORT


Table 13-9: mlbx Pinion West Zone - Modeled Head Grade vs. Gold Recovery	179
Table 13-10: ROM Pinion Gold and Silver Recovery Equations (Oxide)	180
Table 13-11: HPGR Pinion Gold and Silver Recovery Equations (Oxide)	181
Table 13-12: Modeled Pinion S/O Ratios at Various P80 Particle Size	184
Table 13-13: Dark Star North (SI<2.0) - Modeled Head Grade vs. Gold Recovery	185
Table 13-14: ROM – Dark Star Gold Recovery Equations (Oxide)	186
Table 13-15: HPGR - Dark Star Gold Recovery Equations (Oxide)	187
Table 13-16: ROM - Dark Star Gold Recovery Equations (Transition)	187
Table 13-17: HPGR - Dark Star Gold Recovery Equations (Transition)	187
Table 13-18: ROM - Pinion Gold Recovery Equations (Transition)	188
Table 13-19: HPGR - Pinion Gold Recovery Equations (Transition)	188
Table 13-20: Pinon Compacted Permeability Test Results	190
Table 13-21: Dark Star Compacted Permeability Test Results	191
Table 14-1: Summary of Drilling at Dark Star	194
Table 14-2: Descriptive Statistics of Sample Assays in Dark Star Drill-Hole Database	195
Table 14-3: Dark Star Descriptive Statistics by Domain	198
Table 14-4: Dark Star Capping Levels for Gold by Domain	203
Table 14-5: Dark Star Descriptive Composite Statistics by Domain	204
Table 14-6: Dark Star Estimation Areas, Search-Ellipse Orientations and Maximum Search Distances by Domain	205
Table 14-7: Dark Star Estimation Parameters	206
Table 14-8: Dark Star Classification Parameters	207
Table 14-9: Dark Star Total In-Pit Gold Mineral Mineral resources – Measured*	209
Table 14-10: Dark Star Total In-Pit Gold Mineral Resources – Indicated*	209
Table 14-11: Dark Star Total In-Pit Gold Mineral Resources - Measured and Indicated*	210
Table 14-12: Dark Star Total In-Pit Gold Mineral Resources – Inferred*	210
Table 14-13: Number of Samples and Mean Inorganic Carbon Values for Dark Star Estimation Categories	216
Table 14-14: Number of Samples and Mean Sulfide Sulfur Values for Dark Star Estimation Categories	216
Table 14-15: PAG/NAG Designation Criteria	217
Table 14-16: Density Values Applied to the Dark Star Block Model	218
Table 14-17: Drill Holes at Pinion	223
Table 14-18: Descriptive Statistics - Exploration and Mineral Resource Drill-Hole Database	225
Table 14-19: Pinion Deposit Descriptive Gold Statistics by Domain	227
Table 14-20: Pinion Gold Capping Levels for Gold by Domain	230
Table 14-21: Pinion Deposit Descriptive Gold Assay Composite Statistics by Domain	230


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Table 14-22: Pinion Estimation Areas	231
Table 14-23: Pinion Gold Estimation Parameters	233
Table 14-24: Pinion Deposit Descriptive Silver Statistics by Domain	235
Table 14-25: Pinion Capping Levels for Silver by Domain	237
Table 14-26: Pinion Deposit Descriptive Silver Assay Composite Statistics by Domain	237
Table 14-27: Pinion Silver Estimation Parameters	238
Table 14-28: Pinion Classification Parameters	239
Table 14-29: Pinion Measured Gold and Silver Resources*	240
Table 14-30: Pinion Indicated Gold and Silver Resources*	240
Table 14-31 Pinion Measured and Indicated Gold and Silver Resources*	241
Table 14-32 Pinion Inferred Gold and Silver Resources	241
Table 14-33: Pinion Samples Barium Statistics by Domain	245
Table 14-34: Pinion Composites Barium Statistics by Domain	246
Table 14-35: Pinion Barium Estimation Parameters	248
Table 14-36: Number of Samples and Mean Inorganic Carbon Values for Pinion Estimation Categories	250
Table 14-37: Number of Samples and Mean Sulfide Sulfur Values for Pinion Estimation Categories	251
Table 14-38: Assigned Inorganic Carbon and Sulfide Sulfur Values for Pinion Estimation Categories	252
Table 14-39: Sulfide Sulfur Capping Values for Pinion Estimation Categories	253
Table 14-40: Density Values Applied to the Pinion Block Models	254
Table 14-41: Comparison of 2019 Drilling Gold Assays to 2018 Coincident Model Block Grades	257
Table 14-42: Summary of Drilling at Jasperoid Wash	258
Table 14-43: Descriptive Statistics of Sample Assays in Jasperoid Wash Mineral Resource Database	258
Table 14-44: Jasperoid Wash Descriptive Statistics by Gold Domain	263
Table 14-45: Descriptive Composite Statistics by Domain for Jasperoid Wash	266
Table 14-46: Jasperoid Wash Search Ellipse Orientations and Maximum Search Distances by Estimation Area	267
Table 14-47: Jasperoid Wash Estimation Parameters	268
Table 14-48: Jasperoid Wash Inferred Gold Mineral Resources	269
Table 14-49: Density Values Applied to the Jasperoid Wash Block Model	273
Table 14-50: Summary Statistics Sample Gold Grades within the Mineralized Zones	281
Table 14-51: North Bullion Area Semi-Variogram Parameters of Composited Gold Grades	283
Table 14-52: North Bullion Area Density Measurements by Zone	283
Table 14-53: North Bullion Area Estimation and Search Ellipsoid Criteria (from Dufresne and Nicholls, 2018)	284
Table 14-54: North Bullion Area Estimated Gold Grades vs. Average Composite Grades	285
Table 14-55: North Bullion Deposits Classification Criteria	286


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Table 14-56: Parameters Used in Whittle Pit Optimization Studies	287
Table 14-57: North Bullion Mineral Resources with Cutoff Grades*	288
Table 14-58: Cutoff Sensitivity Analysis of the Sweet Hollow and POD Oxide Mineral Resources*	288
Table 14-59: Cutoff Sensitivity Analysis of the North Bullion, Sweet Hollow and POD Sulfide Mineral Resources*	289
Table 15-1: South Railroad Economic Parameters	292
Table 15-2: Dark Star ROM Recovery Equations for Gold	293
Table 15-3: Dark Star HPGR Recovery Equations for Gold	293
Table 15-4: Pinion ROM Recovery Equations	294
Table 15-5: Pinion HPGR Recovery Equations	294
Table 15-6: Dark Star Slope Recommendations by Sector	295
Table 15-7: Pinion Slope Recommendations by Sector	297
Table 15-8: Dark Star Cutoff Grades	298
Table 15-9: Pinon Breakeven Cutoff Grades	298
Table 15-10: Dark Star Pit Optimization Results	300
Table 15-11: Dark Star Pit by Pit Results	301
Table 15-12: Pinion Pit Optimization Results	303
Table 15-13: Pinion Pit by Pit Results	304
Table 15-14: Road and Ramp Design Parameters	307
Table 15-15: Dark Star In-Pit Proven and Probable Mineral Reserves	313
Table 15-16: Pinion In-Pit Mineral Resources	313
Table 15-17: Total Dark Star and Pinion Proven and Probable Mineral Reserves	314
Table 16-1: Waste Containment Requirements (Thousands, Cubic Meters)	315
Table 16-2: Dark Star Mine Production Schedule	317
Table 16-3: Pinion Mine Production Schedule	318
Table 16-4: Total Project Mine Production Schedule	319
Table 16-5: Recovery of Recoverable Gold by Month	331
Table 16-6: Railroad-Pinion Process Production Schedule	332
Table 16-7: ROM Stockpile Balance	333
Table 16-8: HPGR Stockpile Balance	333
Table 16-9: Dark Star Toll Roasting Stockpile Balance	333
Table 16-10: Schedule Efficiency	334
Table 16-11: Mine Equipment Purchases	335
Table 16-12: Personnel Requirements	337
Table 18-1: Current Modeled Pumping Rates for Dark Star Pit Dewatering System	358


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Table 18-2: Expected Pumping Rates for Contact Water Ponds	370
Table 18-3: Summary of Phased Liner Deployment	374
Table 18-4: Results Summary from the Deterministic Model – Typical/Average Range Cycle	375
Table 18-5: Summary of Total Required Emergency Storage Volume for Pond Sizing by Phase	376
Table 18-6: Mean Deterministic Pseudo-Acceleration Response Spectrum by Seismic Source Zone	380
Table 18-7: 84^thPercentile Deterministic Pseudo-Acceleration Response Spectrum by Seismic Source Zone	380
Table 19-1: Bloomberg Consensus Pricing for Gold	382
Table 19-2 Bloomberg Consensus Pricing for Silver	383
Table 20-1: Ministerial Permits, Plans, and Notifications	390
Table 21-1: Capital Cost Summary	392
Table 21-2: Operating Cost Summary	392
Table 21-3: Mining Capital Cost by Year	393
Table 21-4: Initial Capital Process Plant Cost Summary	395
Table 21-5: Expansion Capital Process Plant Cost Summary	396
Table 21-6: Yearly Mine Operating Cost Estimate	398
Table 21-7: Mine General Services Costs	399
Table 21-8: Yearly Mine Maintenance Costs	399
Table 21-9: Yearly Drilling Costs	400
Table 21-10: Yearly Blasting Costs	401
Table 21-11: Yearly Loading Costs	402
Table 21-12: Yearly Haulage Costs	403
Table 21-13: Yearly Mine Support Costs	403
Table 21-14: LOM Operating Costs by Pit and Process Type, US$/tonne ore	404
Table 21-15: Life of Mine Average Process Operating Cost by Year	405
Table 21-16: Power Requirements Summary	406
Table 21-17: Process Consumables Average Annual Consumptions	407
Table 21-18: Yearly G&A Costs	409
Table 22-1: Yearly Mine & Process Physicals	410
Table 22-2: Life of Mine Process Statistics	411
Table 22-3: Capital Expenditure Schedule	412
Table 22-4: LOM Operating Costs	412
Table 22-5: Key Economic Results	413
Table 22-6: Sensitivity Analysis	413
Table 22-7: Detailed Financial Model	415


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FORM 43-101F1 TECHNICAL REPORT


Table 26-1: Cost Estimate for the Recommended Study Program	425


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LIST OF APPENDICES


APPENDIX	DESCRIPTION
A	Preliminary Feasibility Study Contributors and Professional Qualifications – Certificates of Qualified Persons


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FORM 43-101F1 TECHNICAL REPORT

1 SUMMARY

This Technical Report (“Technical Report”) has been prepared by M3 Engineering and Technology Corporation (“M3”) with Gold Standard Ventures Corp. (“Gold Standard” or “GSV”) in accordance with the National Instrument 43-101 Standards of Disclosures for Mineral Projects (“NI 43-101”). The Technical Report presents the results of the South Railroad preliminary feasibility study (“PFS”) in support of updated mineral resource and mineral reserve estimates in the Dark Star and Pinion gold deposits.

Gold Standard’s Railroad Pinion property is located in the Bullion mining district of the southern Carlin trend in Nevada. The property has two adjacent parts, the North Railroad portion (“North Railroad”), which includes POD, Sweet Hollow, and North Bullion (collectively called the North Bullion deposits, or the North Bullion area), and the South Railroad portion (“South Railroad”), which includes Dark Star, Pinion, and Jasperoid Wash.

Extensive metallurgical testing has been completed for the Dark Star and Pinion deposits and is ongoing for the Jasperoid Wash deposit. On the other hand, the North Railroad portion of the property has not been tested comprehensively for metallurgical response.

Gold Standard reports mineral reserves for Dark Star and Pinion deposits for the first time in this Technical Report. The PFS, which includes the mine schedule, process-plant design, and financial analysis, covers only these two deposits.

The proposed project is an open-pit gold mine operation that will deliver ore to a 47.3-tonne heap leach facility over 8 years of mine life. The heap leach facility will treat the ore with three stages of crushing, agglomeration, and leaching on a dedicated leach pad. Tertiary crushing duty is planned to be accomplished with high-pressure grinding rolls (“HPGR”). About 40% of the ore will be stacked and leached as run-of-mine (“ROM”) ore after blasting optimized to produce a fragmentation that has a P80 of about 6 inches (150 mm).

Gold Standard selected M3 and other third-party consultants to prepare mineral resource/reserve estimates, mine plans, process plant design, and to complete environmental studies and cost estimates used for this Technical Report. All consultants have the capability to support the project, as required and within the confines of their expertise, from preliminary feasibility study to full operation.

1.1 PRINCIPAL FINDINGS

The key project parameters and findings are presented in Table 1-1, including a summary of the project size, productions, capital and operating costs, metal prices, and financial indicators.


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Table 1-1: Key Project Data


Mine Life	8 Years + pre-strip (8 months)
Mine Type	Open Pit
Process Description	24.3 M tonnes ROM heap leach; 28.5 M tonnes 3-stage crush with HPGR and agglomeration to heap leach, 0.4 M tonnes sulfide ores to toll milling. Gold recovery by carbon columns, ADR plant
Total Mineral Reserve Estimate	47.3 M Tonnes
Average Grade	0.82 g Au/t Au; 4.70 g Ag/t Ag (Pinion)
Contained Gold / Silver Ounces	1.248 M oz Au; 2.705 M oz Ag
Average Recovery	ROM: 69% Au, 22% Ag; HPGR: 77% Au, 43% Ag
Annual Tonnes Moved	24.3 Million Tonnes
Annual Mineral Reserve Estimate	5.9 Million Tonnes
Sulfide Ore to Toll Milling	0.4 Million Tonnes
Strip Ratio	3.1:1
Design Process Throughput (tonnes/day)	10,000 crushing/agglomeration; 22,500 stacking
Initial Capital Expenditures	$194.0 M
Expansion Capital Expenditures	$88.304 M
Sustaining Capital Expenditures	$20.380 M

Payable Metals
Gold	931,000
Silver	1,040,000

Unit Operating Costs
Average Life of Mine Mining Costs	$1.93/tonne
Average Life of Mine Processing Costs	$1.83/tonne ROM, $4.87/tonne HPGR
G & A	$0.71 /ore tonne
Cash Costs	$601/oz
Cash Costs After By-Product Credit	$582/oz
All in Sustaining Costs (AISC)	$657/oz


Financial Indicators	Base +150	Base +$50	Base Case	Base -50	Base -150
Gold Price (per troy oz)	$1,550	$1,450	$1,400	$1,350	$1,250
Silver Price (per troy oz)	$18.94	$17.72	$17.11	$16.50	$15.28
Pre-tax Cash Flow, $M	$549.50	$456.28	$409.66	$363.05	$269.83
Pre-tax Net Present Value (5%) in $M	$417.64	$340.60	$302.08	$263.56	$186.52
Pre-tax Internal Rate of Return (IRR)	40.49%	35.2%	32.4%	29.51%	23.4%
Pre-tax Payback (Years)	2.4	2.5	2.6	2.6	2.8
After-tax Cash Flow, $M	$448.12	$374.18	$337.11	$299.76	$222.85
After-tax Net Present Value (5%) in $M	$333.23	$272.11	$241.47	$210.61	$147.05
After-tax Internal Rate of Return (IRR)	34.70%	30.14%	27.77%	25.30%	19.95%
After-tax Payback (Years)	2.5	2.6	2.7	2.7	2.9

The effective dates of the Pinion and Dark Star databases on which the mineral resources described in this Technical Report are estimated are May 31, 2019 and April 26, 2019, respectively. The effective date of both the Pinion and Dark Star mineral resource estimates is August 7, 2019. The effective dates of the Jasperoid Wash database and resource estimate are October 6, 2018 and November 15, 2018, respectively. The effective date of the North Bullion mineral database and mineral resource estimates are August 18, 2017 and September 15, 2017, respectively.


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1.2 PROPERTY DESCRIPTION AND OWNERSHIP

The Railroad Pinion property is in the Piñon Range in Nevada, accessed primarily from U.S. Interstate 80 (“I-80”), approximately 442 km west of Salt Lake City, Utah, and 467 km east of Reno, Nevada. The property is located between 13 and 29 km south of I-80 and can be reached by a series of paved and gravel roads from Elko, Nevada (population 18,300). The property is centered approximately at UTM NAD27 Zone 11 coordinates of 585,000E and 4,480,000N.

Gold Standard’s contiguous North and South Railroad portions of the Railroad-Pinion property constitute a combined land position totaling 21,679 hectares in Elko County, Nevada, centered approximately at UTM NAD27 Zone 11 with coordinates of 585,000E and 4,480,000N. This includes 1,454 claims owned by Gold Standard and 207 claims held under lease, a total of 30 claims are patented. There is also a total of 9,562 gross hectares of private lands of which Gold Standard’s ownership of the subsurface mineral rights varies from 49.2% to 100%.

1.3 EXPLORATION AND MINING HISTORY

The Railroad–Pinion property is being explored on an ongoing basis by Gold Standard using geological mapping, geochemical and geophysical surveying, and drilling. Exploration work by Gold Standard commenced in 2010 and has resulted in the identification of 17 prospect areas or zones of mineralization within the property.

Twenty-one different historical operators are known to have drilled 1,084 holes, for a total of 152,566.1 m, from 1969 through 2008. As of the database effective dates, Gold Standard has drilled 848 holes for a total of 248,227.1 m. At least 74% of all drilling used RC methods. However, the amount of RC drilling may be understated because the hole-types are not known for a substantial number of holes drilled in the late 1980s and 1990s, when RC drilling was common.

1.4 GEOLOGY AND MINERALIZATION

The Railroad-Pinion property is located in the southern portion of the Carlin trend, centered on the Railroad dome in the Piñon Range, which is comprised of Ordovician through Permian marine sedimentary rocks. Eastern assemblage formations throughout the property include the Pogonip, Hanson Creek, Eureka Quartzite, Lone Mountain Dolomite, Oxyoke, Beacon Peak, Sentinel Mountain Dolomite, and Devils Gate Limestone and Tripon Pass formations. Siliceous clastic units include those of the Webb, Chainman, and Tonka formations. The north-south-striking Bullion fault corridor separates Tertiary volcanic rocks to the east from the Paleozoic sedimentary units in the range, which have been intruded by a complex of Eocene igneous rocks centered south of Bald Mountain, in the core and east flank of the range.

The gold-silver deposits within the Railroad-Pinion property that are the focus of this Technical Report are considered to be Carlin-type, sedimentary-rock-hosted deposits. Precious metal mineralization is generally submicroscopic, disseminated, and hosted principally in sedimentary rocks, with some mineralization in felsic dikes and sills as well.

In the South Railroad portion of the property, the Dark Star Main (“Dark Star Main”) and Dark Star North (“Dark Star North”) zones, which comprise the Dark Star deposit are hosted primarily within Pennsylvanian-Permian rocks, with minor amounts of gold mineralization found in the Chainman Formation and Tertiary conglomerates. The deposits are centered along the roughly north-south Dark Star fault corridor, within which is a horst block and associated silicified zone bounded by the West fault and Dark Star fault. Gold mineralization in the horst block is hosted in the middle, coarse-grained conglomeratic and bioclastic limestone-bearing unit of a Pennsylvanian-Permian undifferentiated sequence interpreted to be equivalent to the Tomera Formation. Mineralization dips steeply to the west near the surface at Dark Star Main and Dark Star North, but dips less steeply at depth at Dark Star Main.

Also, in the South Railroad portion of the property, the Pinion deposit is situated in a sequence of Paleozoic sedimentary rocks exposed within large horst blocks in which the sedimentary rocks have been broadly folded into a south- to


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southeastward-plunging, asymmetric anticline. The axis of this Pinion anticline trends approximately N50ºW to N60ºW and can be traced for approximately 3.2 km. The limbs of the anticline dip shallowly at 10° to 25° to the west, and more steeply at 35° to 50° to the east. Disseminated gold and silver mineralization at the Pinion deposit is strongly controlled by a 3 m to 120 m-thick dissolution-collapse breccia at the contact between calcarenite of the Devils Gate Limestone and the overlying silty micrite of the Tripon Pass Formation. Gold deposition was contemporaneous with breccia development, quartz veins formation, silica ± barite replacement and infill of open spaces.

The Jasperoid Wash disseminated gold deposit, also located in the South Railroad portion of the property, is hosted by altered Tertiary feldspar porphyry dikes and their host Pennsylvanian-Permian conglomeratic rocks of a Tomera Formation equivalent. The deposit has approximate extents of 1,400 m to the north and a width of about 1,100 m, and is partially contained within an elongate, north to south, steeply dipping structural corridor. Drilling shows the deposit dips steeply to the west nearby and within Tertiary dikes; east of the dikes, the deposit dips gently to the west. The gold is inferred to be submicroscopic in grain size, however, petrographic studies have yet to be performed.

In the North Railroad portion of the property, disseminated gold mineralization has been defined by drilling in the North Bullion, POD, and Sweet Hollow zones. The mineralization is focused in the footwall of the Bullion fault zone. Faults appear to be important controls on mineralization. In general, gold-silver mineralization is localized in gently to moderately dipping, strongly sheared rocks of the Webb and Tripon Pass formations, in dissolution-collapse breccia developed above and within silty micrite of the Tripon Pass Formation, and calcarenite of the Devils Gate Limestone. The top of gold mineralization varies from 105 m to 400 m below the surface and varies in dip from 10° to 45° to the east. Gold is associated with “sooty” sulfide minerals, silica, carbon, clay, barite, realgar, and orpiment.

1.5 DATA VERIFICATION

Mr. Ristorcelli is satisfied that the Pinion, Dark Star, and Jasperoid Wash drilling databases are in good condition. Various audits and checks were performed by MDA to verify collar coordinates, down-hole deviation surveys, geology and assay data in the drill-hole database. All Gold Standard gold assay data was verified using digital laboratory certificates. However, about one third of the Pinion assays and one quarter of the Dark Star assays from historical drill campaigns were unsupported with original assay certificates. Drill-hole data lacking adequate supporting documentation, as well as data from holes observed during sectional modeling to be inconsistent with surrounding holes, were treated as lower confidence, or excluded from use in modeling and estimation. Mr. Dufresne states the North Bullion area database has undergone extensive verification.

In 2019, Gold Standard supplemented their Pinion silver database with re-assayed individual samples for which composites of multiple intervals had previously been analyzed. Over 50% of the original certificates were available for all silver data and were used for verification. Quality assurance/quality control (“QA/QC”) data was also evaluated, and the silver data was deemed acceptable for use in estimation of classified mineral resources.

Cyanide-soluble gold assays at Dark Star and Pinion were not verified, and no QA/QC data was available for evaluation. Inorganic carbon and sulfide sulfur data were audited and determined to be adequate for use in their respective estimates done for metallurgical characterization. Carbonate carbon was also verified as adequate. No QA/QC data was associated with the carbon and sulfur analyses.

There is no evidence of significant historical QA/QC programs for drilling in the South Railroad portion of the property prior to 2014. For Gold Standard programs, the QA/QC program was minimal in 2014 through 2016 but was more comprehensive in 2017 and 2018. The results and amount of QA/QC data, as well as non-remedied QA/QC “failures,” were considered in mineral resource classification for the Dark Star, Pinion, and Jasperoid Wash deposits. Mr. Ristorcelli concludes that the Dark Star, Pinion, and Jasperoid Wash analytical data are adequate for the purposes used in this Technical Report, subject to issues described in Section 12.


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Mr. Dufresne concludes that the QA/QC protocols employed by Gold Standard throughout the 2010 to 2017 drilling programs at the North Bullion deposits were adequate and appropriate for ensuring high accuracy and precision in the sample assays. However, no QA/QC evaluations were done on historical drill campaigns. Mr. Dufresne considers the North Bullion area drill database sufficiently verified for use in the mineral resource estimation discussed in Section 14.5.

Barium was estimated in the Pinion deposit block model for metallurgical characterization. Barium analyses were done using pressed-powder energy-dispersive x-ray fluorescence (“XRF-ED”) and loose-powder NITON XRF analytical methods. These methods were evaluated by running additional analyses on duplicate pulp samples by various methods. Mr. Ristorcelli believes that the data can be used to estimated NITON XRF-derived barium grades into the model.

1.6 MINERAL PROCESSING AND METALLURGICAL TESTING

The current study of the South Railroad portion of the Railroad-Pinion project focuses on two main sources of ore, for which mineral reserves are declared in this study: The Pinion and Dark Star deposits. These deposits have different geo-metallurgical characteristics, which are briefly summarized as follows:

The Pinion deposit can be characterized as hard and abrasive material, with a steep feed P80 vs. gold recovery response. Much of the gold is contained in the rock ground mass and requires fine crushing (-1/4” inch) to liberate gold for the most efficient cyanide-leach extraction. Gold recovery has proven to be sensitive to high barite/silica content. Materials with higher barite content demonstrate significantly lower gold recovery than low-barite materials. Gold recovery from the high-barite materials benefits the most from fine crushing. This deposit can be heap leached without crushing, at low gold recovery, conventionally crushed and leached at modestly higher gold recovery, or HPGR-crushed at significantly higher gold recovery. Additional metallurgical testing and trade-off economic analysis of the three heap-leach flowsheet options will be required to make the final decision on how to proceed at Pinion.

The Dark Star deposit can be characterized as hard and moderately abrasive material, with a flat feed P80 vs. gold recovery response. Most of the gold is contained in fractures that have been oxidized and accessible to cyanide solutions that easily pass through the rock matrix. Consequently, high gold extractions are achieved at coarse particle size, requiring no crushing prior to heap leaching.

Due to the multiple material types, and the dependence of gold recoveries on head grades and crush size, 66 gold and silver recovery v head grade equations were developed, along with recovery v solution-to-ore ratio equations. These equations were used to develop ore routing, optimize the mine schedule, and estimate gold and silver production over the life of mine.

1.6.1 Pinion Deposit Tests 2016 – 2019

Gold Standard commissioned bottle-roll test and column leach test programs in 2015-2016, 2016-2017, 2018, and 2019 on the Pinion deposit. Table 1-2 shows the head assays of the samples used.

Table 1-2: Head Assays and Geo-metallurgical Characterization of Pinion Samples


Component	2015-2016		2016-2017		2018	2019
Component	Range	Average	Range	Average	Average	Range	Average
Au grade, ppm	0.19 - 4.4	0.81	0.23 - 1.82	0.76	0.736	0.25 - 2.87	0.85
Ag grade, ppm	0.62 - 72.3	6.9	3.3 - 38.7	10.4	4.53	0.5 - 29.1	7.7


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1.6.1.1 Column Leach Testing

Column leach tests were performed on 12.5-mm and 25-mm materials in 2016-2017, and in 2019. A summary of the tests results is shown in Table 1-3. Gold extraction plotted by days under leach for the column-leach tests are shown graphically in Figure 1-1 and Figure 1-2.

Table 1-3: Summary of Column Leach Test Results for Pinion.


	P80, mm	2016-2017		2019
	P80, mm	Range	Average	Range	Average
Au Recovery, %	12.5	55.8 – 90.4	70	29.8 - 80	63
Au Recovery, %	25	51.5 - 69.5	56.4	30.5 – 81.2	59.9
Ag Recovery, %	12.5	5.4 – 47.3	22.7	9.5 – 76.4	30.6
Ag Recovery, %	25	9.7 – 44.8	22.6	9.5 – 76.4	30.6
CN-, kg/t			1.0		0.76
Lime, kg/t			0.56		0.93

The plots show kinetics of heap leaching that could be expected, as well as the variability of the samples within the Pinion orebody.

1.6.1.2 Pinion Deposit High Pressure Grinding Roll (HPGR) Testing

Gold Standard commissioned KCA to perform HPGR-crush column-leach testing on a drill core composite sample from the Pinion Main zone. The results of the tests are included in Figure 1-2. HPGR crushing showed improvement in gold and silver extraction compares to conventionally crushed ore, as shown in Figure 1-3 for gold.

Cyanide consumption was low to moderate. Lime consumption ranged from 0.5 to 1.2 kg/tonne.

Figure 1-1: 2016-2017 Gold Extraction vs. Days Under Leach for Column-Leach Tests


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Figure 1-2: 2019 Pinion Gold Extraction vs. Days under Leach for Column-Leach Tests

Figure 1-3: Conventional Crush vs. HPGR Gold Extraction Comparison


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Gold Standard Dark Star Deposit Metallurgical Testing

Head assays and geo-metallurgical characterization analyses were obtained for 68 composites in 2017 and 50 composites in 2019, using a combination of four separate laboratories: KCA, ALS, UBC, and FLS. Table 1-4 shows the results of the head assays and geo-metallurgical characterization.

Table 1-4: Head Assays and Geo-metallurgical Characterization of Dark Star Samples


Component	2017		2019
	Range	Average	Range	Average
Au grade, ppm	0.177 - 7.35	1.59	0.182 - 5.62	1.23
Ag grade, ppm	0.27 - 5.07	0.71	0.50 - 3.50	1.01

Drill core composites from the Dark Deposit were subjected to bottle-roll leach testing at target P80 sizes of 75 µm and 1,700 µm in 2017 and 2019. Percentage of gold recoveries where in the low to mid 80s. Cyanide consumption ranged from 0.4 to 1.8 kg/tonne, while lime consumption ranged from 0.8 to 1.3 kg/tonne.

1.6.2.1 Dark Star Column-Leach Tests

Forty-one (41) 2017 composites and fifty 2019 composites were column leached utilizing material crushed to 100% passing 19 mm (target P80 = 12.5 mm), and 100% passing 37.5 mm (target P80 = 25 mm). The results of the tests, are summarized in Table 1-5 below, and plotted in Figure 1-4 and Figure 1-5, show good recoveries for transition and oxide ores.

Table 1-5: Summary of Column Leach Test Results for Dark Star


	P80, mm or type	2017		2019
		Range	Average	Range	Average
Au Recovery, %	12.5	15.0 - 94.8	78.9
Au Recovery, %	25
	Sulfide	15 – 25.5	20.3	28.4 – 39.9	35.1
	Transition	57.8 – 85.8	69.7	47.5 – 84.7	67.2
	Oxide	56.3 – 94.9	84.1	63.3 – 94.7	84.4
CN-, kg/t			1.07		0.95
Lime, kg/t			1.15		0.92


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Figure 1-4: 2017 Dark Star Column-Leach Gold Extraction vs. Days under Leach


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Figure 1-5: 2019 Dark Star Column-Leach Gold Extraction vs. Days under Leach

1.6.2.2 Dark Star HPGR Metallurgical Test Work

Column-leach tests were performed on eight HPGR composite-sample charges, four from Dark Star Main and four from Dark Star North. The Dark Star Main HPGR column-leach gold extractions are significantly higher than the conventional-crushed column charge as shown in Figure 1-6. The Dark Star North HPGR gold extractions are only marginally higher than the conventional-crushed composite at similar crush size (see Section 13).


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Figure 1-6: Conventional Crush vs. HPGR Gold Extraction Comparison, Dark Star Main

Comminution Characterization and Load Permeability Tests

The testing programs included comminution characterization using JK SMC testing and Bond abrasion measurements, as well as load permeability testing on column leach tailing. Details are included in Section 13.

Geo-metallurgy Characterization and Recovery Models

Large geo-metallurgy databases have been developed for the Pinion and Dark Star deposits to assist in evaluating material type selections, representing different Au and Ag recovery response. The corresponding geo-metallurgical analysis has identified key variables, within both deposits, that were used to select the different metallurgical recovery zones requiring separate gold recovery modeling.

1.6.4.1 Pinion Deposit Geo-Metallurgy

The following is a summary of the four gold and silver recovery zones in the Pinion Deposit:

Mtp (Tripon Pass) – Tripon Pass mineralization is a formation unit that sits on top of the multi-lithic breccia (mlbx) which hosts the majority of the Au mineralization at Pinion.
Mlbx Pinion East (Ba > 4.0%, Hi SiO2) – The Pinion East Zone is carved out of a larger mlbx zone that is characterized by high barium (Ba) > 4.0% and high quartz (SiO2) > 65%.
Mlbx Pinion West – The Pinion West Zone captures all the remaining Pinion mlbx zone of mineralization that is not contained within the Pinion East (Ba > 4.0%, Hi SiO2) zone.
Ddg (Devils Gate) – Devils Gate mineralization is stratigraphically positioned underneath the Pinion mlbx.


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1.6.4.2 Dark Star Deposit Geo-Metallurgy

The Dark Star mineralization is hosted in two connected zones: Dark Star North and Dark Star Main. Dark Star North can be characterized as a relatively high-grade heap leachable zone, whereas Dark Star Main is lower grade and contains more transitional mineralization. Within both zones, gold mineralization is mainly contained within three formation units: ST-U (upper siltstone), CGL (middle conglomerate), and ST-L (lower siltstone). Geo-metallurgical evaluations did not detect significant variation in gold recovery based upon the host formation but did identify a significant difference is gold recovery response in local regions of low and high silica intensity (SI), as logged by the geologists. SI is characterized by the geologists using a scale of 0 to 3, with 0 indicating no (or low) silica and 3 being the highest silica.

Recovery models for silver were not developed for Dark Star because of its low silver contents.

The following is a summary of the four gold recovery zones, in the Dark Star deposit:

Dark Star Main (SI<2.0)
Dark Star Main (SI>2.0)
Dark Star North (SI<2.0)
Dark Star North (SI>2.0)

1.6.4.3 Au and Ag Recovery Methodology

Four steps were used in developing final Au and Ag recovery models for Pinion and Dark Star:


Step 1:	Determining the gold extraction for each variability composite using a combination of fine grind/crush bottle rolls and medium/coarse crush column tests.

Step 2:	Develop head grade vs. tails grade models to use in final development of the gold recovery equations.

Step 3:	Build a database of the laboratory solution: ore (S/O) ratio data at various percentages of total extractable gold. A correction factor is applied to each laboratory S/O ratio data point to scale up the laboratory data to commercial scale. Typical laboratory S/O ratio data is tabulated for the following percentages of total extractable gold: 60%, 70%, 80%, 90%, 95%, and 99%.

Step 4:	Incorporating steps 1-3 into final recovery models that reflect commercial scale inefficiencies and deductions for solution losses, plus application of cumulative S/O ratios over the life of the project to predict timing of gold recovery.

Details of this methodology are included in Section 13 and will not be repeated here to save space. Sixty-four (64) recovery v grade equations were developed of the form

Recovery % = k1 ln(head grade) + k2,

which were used with the recovery v solution:ore ratio relationships to predict metal production over the life of mine. These equations are listed in Section 13 and in the Metallurgy Report (Simmons, 2019).

1.6.5 Reagent Consumption

Reagent consumptions and requirements, including cyanide, lime and cement were estimated by KCA based on metallurgical test work completed to date for the Pinion and Dark Star material. Reagent consumptions are summarized below.


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In KCA’s experience, field cyanide consumptions are typically 25% to 33% of observed laboratory consumptions. Cyanide consumptions for the ROM and HPGR-crushed Pinion and Dark Star material have been estimated at 33% of the laboratory consumptions for this study.

As there were no column leach tests performed on ROM material, cyanide consumptions have been estimated based on column leach tests on 37.5-mm crushed materials from Pinion and Dark Star. ROM cyanide consumption in the field are typically be 80% less than crushed ore consumption. Laboratory cyanide consumptions for Pinion material at 37.5 mm crush ranged from 0.66 kg/t to 1.19 kg/t, with an average consumption of 0.85 kg/t. Dark Star laboratory cyanide consumptions at 37.5 mm crush ranged from 0.46 kg/t to 1.31 kg/t, with an average consumption of 0.87 kg/t. Based on this data, field cyanide consumptions are estimated at 0.22 kg/t and 0.23 kg/t for ROM Pinion and Dark Star material, respectively.

Laboratory cyanide consumption for Pinion material ranged from 0.48 kg/t to 0.89 kg/t with an average consumption of 0.67 kg/t. Dark Star laboratory cyanide consumption ranged from 0.63 kg/t to 0.89 kg/t with an average consumption of 0.76 kg/t. Based on this data, field cyanide consumptions are estimated at 0.22 kg/t and 0.25 kg/t for Pinion and Dark Star HPGR crushed material, respectively.

Lime is required for pH control for the ROM ore during leaching. Because hydrated lime was utilized in the lab leach tests, the laboratory lime consumptions are adjusted to predict consumptions of quicklime (pebble lime, CaO) in the field. Estimated quicklime consumption for Pinion and Dark Star ROM ores is 1.0 kg/t of ore.

Cement is required for heap permeability and pH control during leaching for the Pinion and Dark Star HPGR crushed material. Based on compacted permeability tests, cement requirements are estimated at 2.0 kg/t ore and 7.0 kg/t ore for Pinion and Dark Star HPGR crushed material, respectively for a maximum heap height of 60 m.

1.7 RECOVERY METHODS

The process selected for recovery of gold and silver from the Pinion and Dark Star ore is a conventional heap-leach recovery circuit. Ore will be mined by standard open pit mining methods from two separate pits. Lower-grade Pinion and Dark Star ore will be truck-stacked on the heap as ROM ore directly, without crushing, at an average rate of 12,500 tonnes of ore per day in 9 m lifts; lime will be added directly to the haul trucks for pH control. Higher grade Pinion and Dark Star (low clay) ore will be processed in a three-stage crushing circuit with a high-pressure grinding roll (HPGR) at an average rate of 10,000 tonnes of ore per day. Ore will be crushed to 100% passing 14 mm, treated with cement and agglomerated, then conveyor-stacked onto heap leach pad in 7 m lifts. Lime will be added to crushed Pinion ore for additional pH control.

Stacked ore will be leached with a dilute cyanide solution using a drip irrigation system. After percolating through the ore, the pregnant gold and silver bearing solution will flow by gravity to a pregnant solution tank where it is pumped to a carbon adsorption circuit to recover the precious metal from solution. The gold and silver will be stripped from the loaded carbon using a desorption process, followed by electrowinning to produce a precipitate sludge. The precipitate sludge will be processed using a retort oven for drying and mercury recovery, and then refined in a melting furnace to produce gold and silver doré bars.

1.8 MINERAL RESOURCE ESTIMATE AND MINERAL RESERVE ESTIMATE

1.8.1 Mineral Resource Estimate

The estimated mineral resources presented in this Technical Report were classified in order of increasing geological and quantitative confidence into inferred, indicated, and measured categories to be in accordance with the “CIM Definition Standards - For Mineral Resources and Mineral Reserves” (2014) and therefore Canadian National


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Instrument 43-101. Mineral resources are reported at cutoffs that are reasonable for deposits of this nature given anticipated mining methods and plant processing costs, while also considering economic conditions, because of the regulatory requirements that a mineral resource exists “in such form and quantity and of such a grade or quality that it has reasonable prospects for eventual economic extraction.”

MDA modeled geology and metal domains for the Dark Star, Pinion, and Jasperoid Wash deposits, then estimated and classified gold mineral resources. A silver estimate was also produced for the Pinion deposit. Gold Standard provided the geologic modeling for the various deposits and were intimately involved with metal domain modeling. Block sizes were 9 m x 9 m x 9 m for Dark Star and Pinion, and 6 m x 6 m x 6 m for Jasperoid Wash. Estimation was done using inverse-distance methods with powers ranging from two to four. Multiple models were estimated in order to optimize the estimation parameters.

The estimate of mineral resources for the Railroad-Pinion property, excluding North Bullion, is the block-diluted inverse-distance estimate and is reported at variable cutoffs for open-pit mining. The cutoff for oxidized and transitional redox material is 0.14 g Au/t, whereas the cutoff for sulfide material is 1.0 g Au/t. Mineral resources were classified as Measured, Indicated or Inferred for each deposit separately. Factors considered for classification include results of data verification and QA/QC results, the level of geologic understanding of each deposit, and performance of past mineral resource block models with new drilling. Sulfide material at Dark Star was reported at a higher cutoff grade of 1.0 g Au/t. Table 1-6 presents the pit-constrained estimated mineral resources for the Dark Star, Pinion, and Jasperoid Wash deposits based on a $1,500/oz gold price. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 1-6: Dark Star, Pinion and Jasperoid Wash Estimated Mineral Resources


Dark Star Measured and Indicated*
Cutoff, g Au/t	Tonnes	g Au/t	oz Au
variable**	32,717,000	0.88	921,000
Dark Star Inferred
Cutoff, g Au/t	Tonnes	g Au/t	oz Au
variable**	2,479,000	0.70	56,000
*mineral resources are inclusive of mineral reserves
South Railroad - Cutoff for oxide and transitional mineral resources is 0.14 g Au/t, and for sulfide mineral resources at 1.0 g Au/t;*


Pinion Measured and Indicated*
Cut off, g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.140	28,925,000	0.58	544,000	4.22	3,929,000
Pinion Inferred
Cut off, g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.140	10,810,000	0.64	224,000	3.80	1,322,000
* mineral resources are inclusive of mineral reserves


Jasperoid Wash Inferred
Cut off, g Au/t	Tonnes	g Au/t	oz Au
0.140	10,569,000	0.33	111,000

Barium was estimated into the Pinion deposit block model for use in metallurgical characterization of the Pinion mineralized material. The average barium grade is ~1.7% for the gold mineralization grading at least 0.14 g Au/t. Factoring between barium analytical results were required, which added some uncertainty to the model.


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Cyanide-soluble gold block models were produced for the Pinion and Dark Star deposits. These estimates appear reasonable in areas with Gold Standard drilling, however, there is less confidence in some areas where cyanide-soluble gold data is lacking, such as where historical drilling is predominant.

An acid-base accounting (“ABA”) model was generated for Pinion and Dark Star to characterize waste material for mine planning and handling. Because of limited data, these estimates can only be considered as guides for environmental planning.

The North Bullion area gold mineral resources were estimated within modeled mineralized lode solids using variable search ellipsoids by Mr. Dufresne of APEX Geoscience Ltd. (“APEX”), of Edmonton, Alberta. The reported mineral resource was estimated using inverse distance squared, and the parent block size for the sub-blocked model was 10 m x 10 m x 3 m. Table 1-7 presents the pit-constrained estimated mineral resources for North Bullion based on a $1,350/oz gold price, and Table 1-8 combines all mineral resources for the entire Railroad-Pinion project.

Table 1-7: North Bullion Mineral Resources


North Bullion Indicated
Cut off, g Au/t	Tonnes	g Au/t	oz Au
0.140	2,920,000	0.96	90,100
North Bullion Inferred
Cut off, g Au/t	Tonnes	g Au/t	oz Au
variable**	10,970,000	2.28	805,800
*Cutoff for potential underground sulfide mineral resources at North Bullion are reported at 2.25 g Au/t*

Table 1-8: Total Railroad-Pinion Mineral Resources


All Railroad-Pinion Measured and Indicated*
Cut off, g Au/t	Tonnes	g Au/t	oz Au
variable**	64,562,000	0.75	1,555,100
All Railroad-Pinion Inferred
Cut off, g Au/t	Tonnes	g Au/t	oz Au
variable***	34,828,000	1.07	1,196,800
* mineral resources are inclusive of mineral reserves *South Railroad - Cutoff for oxide and transitional mineral resources is 0.14 g Au/t, and for sulfide mineral resources at 1.0 g Au/t;*
**Cutoff for potential underground sulfide mineral resources at North Bullion are reported at 2.25 g Au/t*

Mineral Reserve Estimate

Measured and Indicated mineral resources were used as the basis to define mineral reserves for both the Dark Star and Pinion deposits. Mineral reserve definition was done by first identifying ultimate pit limits using economic parameters and applying pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Modifying factors including mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social, and governmental factors have been considered in the estimate of mineral reserves.

MDA provided the final production schedule to M3 Engineering who developed the final cash-flow model which demonstrates that the Pinion and Dark Star deposits make a positive cash flow and are reasonable with respect to statement of mineral reserves for these deposits.


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The total Proven and Probable mineral reserves reported for the PFS are shown in Table 1-9. Within the designed pits there are a total of 147.3 million tonnes of waste associated with the in-pit mineral resources. This results in an overall project strip ratio of 3.11 tonnes of waste for each tonne of material processed.

Table 1-9 Proven and Probable Mineral Reserves


*Dark Star*	K Tonnes	g Au/t	K Ozs Au
Proven	5,434	1.39	243
Probable	24,023	0.83	641
P&P	29,456	0.93	884


*Pinion*	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag
Proven	1,081	0.66	23	5.49	191
Probable	16,806	0.63	341	4.65	2,514
P&P	17,887	0.63	364	4.70	2,705


*Consolidated Gold Reserves*
*Dark Star & Pinion*	K Tonnes	g Au/t	K Ozs Au
Proven	6,515	1.27	266
Probable	40,829	0.75	982
P&P	47,344	0.82	1,248

Note: cutoff grades are applied by material type as described in Section 15.2.3; and Pinion Proven and Probable mineral reserves for Pinion include silver as reported above.

1.9 MINING METHODS

The PFS includes mining at both the Dark Star and Pinion deposits; both are planned as open-pit, truck and shovel operations. The truck and shovel method provides reasonable costs and selectivity for these deposits.

The production schedule considers the processing of material by ROM, HPGR, and toll processing. HPGR stockpiles of lower-grade material will be utilized to ensure that higher-grade material is processed first. The stockpiles will be maintained near the crusher. All ROM material will be dumped in place directly on the ROM leach pad. Monthly periods were used to create the production schedule with pre-stripping starting in Dark Star at month -9. Start of ROM processing is assumed to be month 1. The maximum rate for ROM processing will be 12,500 tonnes per day or 4,562,500 tonnes per year on a 365-day basis. HPGR processing is started one year after the start of ROM processing. The maximum HPGR process rate is targeted at 10,000 tonnes per day or 3,650,000 tonnes per year on a 365 day per year basis.

Toll processing occurs from year 2 through year 5 and the rate is limited to 500 tonnes per day. Only Dark Star material with greater than 1.17 g Au/t would be toll processed and the material would be stockpiled near Dark Star prior to being loaded in a contractor’s over-the-road haul truck.

The total Dark Star mining rate would ramp up from 20,000 tonnes per day to about 90,000 tonnes per day over a period of 8 months. A maximum of 90,000 tonnes per day is used in the production schedule. Mining is to start in Dark Star North, then progress to Dark Star Main and then Pinion.


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the first principle hours required for primary mining equipment to achieve the production schedule. Primary mining equipment includes drills, loaders, hydraulic shovels, and CAT-785 style haul trucks.

Waste storage facility designs were created for the PFS to contain the material that is not processed. A 1.3 swell factor was assumed which provides for both swell when mined and compaction when placed into the facility.

1.10 INFRASTRUCTURE

Project infrastructure for South Railroad has been developed to support the mining and heap leaching operations. Access to the site will be provided by a new ten-mile access road to the facility following an existing dirt road east of highway 278. Electrical power will be supplied from upgraded power lines along the highway from Carlin, NV, and then to the site by Nevada Energy (“NVE”) adjacent to the new access road. Project buildings located at the site will include Security and Emergency services, Administration, Change House, Crushing, Truck Shop, ADR/Refinery Plant, and Laboratory buildings. These will mainly be located between Pinion and Dark Star pits for ease of access and be connected by local roads and haul routes.

1.11 ENVIRONMENT AND PERMITTING

Gold Standard has been conducting environmental baseline studies over the past several years as part of their ongoing permitting efforts and in preparation for the submittal of permit applications for conduct mining operations. The main portion for the project area has been surveyed for surface water resources, including Waters of the United States (WOTUS), biological resources, and cultural resources. The project access road, the powerline route, and the water management area remain to be surveyed. In 2018, Gold Standard commenced material characterization testing of the mineralized material and waste rock to determine the metal leaching and acid generation potential. Additionally, an evaluation of the groundwater resources was commenced to determine groundwater supply potential, as well as the potential impacts from groundwater pumping and pit lake development. Gold Standard had a meeting with the BLM in January 2019 to determine any additional baseline data collection needs for the permitting process.

Within and adjacent to the project area there are Greater Sage Grouse and Golden Eagles. These species will have an effect on how the project is permitted and what mitigation in required or proposed. Gold Standard is working with the BLM on the management of these species.

The review and approval process for the Plan Application by the BLM constitutes a federal action under the National Environmental Policy Act (“NEPA”) and BLM regulations. Thus, for the BLM to process the Plan Application the BLM is required to comply with the NEPA and prepare either an EA, or an Environmental Impact Statement (“EIS”). Gold Standard anticipates that the BLM will require an EIS, due to the mine dewatering and potential pit lake. Gold Standard will also need an Individual Section 404 Permit from the United States Army Corps of Engineers, and this agency will be a cooperating agency on the NEPA documents.

There are a number of environmental permits issued by the NDEP that are necessary to develop the project and which Gold Standard needs to permit the project. The NDEP issues permits that address water and air pollution, as well as land reclamation. The Nevada Division of Water Resources (“NDWR”) issues water rights for the use and management of water.

The SRMP (as defined below) is a previously explored minerals property with exploration related disturbance. However, there have been very long periods of non-operation. There are no known ongoing environmental issues with any of the regulatory agencies. Gold Standard has been conducting baseline data collection for a couple of years for environmental studies required to support the Plan Application and permitting process. The waste and mineralized material characterization and the hydrogeologic evaluation are currently in their latter stages of development. Material characterization indicates the need to manage a significant portion of the waste rock as potentially acid generating in engineered facilities. Additional results to date indicate limited cultural issues, air quality impacts appear to be within


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State of Nevada standards, traffic and noise issues are present but at low levels, and socioeconomic impacts are positive.

Social and community impacts have been and are being considered and evaluated for the Plan Amendment and Plan Application performed for the project in accordance with the NEPA and other federal laws. Potentially affected Native American tribes, tribal organizations and/or individuals are consulted during the preparation of all plan amendments to advise on the proposed projects that may have an effect on cultural sites, resources, and traditional activities.

Potential community impacts to existing population and demographics, income, employment, economy, public finance, housing, community facilities and community services are evaluated for potential impacts as part of the NEPA process. There are no known social or community issues that would have a material impact on the project’s ability to extract mineral resources. Identified socioeconomic issues (employment, payroll, services and supply purchases, and state and local tax payments) are anticipated to be positive.

A Tentative Plan for Permanent Closure (“TPPC”) for the project would be submitted to the NEDP with the WPCP application. In the TPPC, the proposed heap leach closure approach would consist of fluid management through evaporation, covering the heap leach growth media, and then revegetating. The design of the process components is not sufficiently advanced to determine the closure costs. Any residual heap leach drainage will be managed with evaporation cells.

1.12 WATER MANAGEMENT

Gold Standard developed a Water Management Plan for South Railroad in support of the PFS. The Water Management Plan formed the basis for evaluating the infrastructure and associated cost to manage water through the life cycle of the mine. The purpose of the Water Management Plan is to present the water management strategies that focus on water as an asset and allow Gold Standard to proactively plan and manage water from development to post-closure such that operational and stakeholder water needs are met, and that human health and the environment are protected.

To support the development of water management strategies for the project, the following pre-design studies/activities were completed:

Analytical and numerical groundwater model to estimate pit dewatering requirements and potential impacts for the Dark Star North pit;
Evaluation and modeling of long-term climate records and 24-hour design storms used as input for event- based stormwater modeling, continuous water balance modeling, and infiltration modeling;
Stormwater modeling and calculations for locating and sizing stormwater management infrastructure;
Infiltration modeling to predict the amount of seepage from the WRDFs that will require management during operation, closure, and post-closure periods;
Water balance modeling to evaluate the supplies of and demands of site water over the life of mine;
Evaluation of water disposal alternatives; and
Closure cover assessment to limit or eliminate post-closure water handling and treatment demands.

The water management strategy and technical investigations to support the Water Management Plan resulted in the following PFS level infrastructure:

Stormwater management and seepage collection facilities, such as channels, ponds, culverts, tanks, attenuation structures, down drains, and other related open-channel stormwater controls;
A groundwater dewatering system needed to mine ore below the groundwater table in the Dark Star pits; and
A Rapid Infiltration Basin (“RIB”) for water disposal to handle excess water generated by the Dark Star groundwater dewatering system.


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1.13 CAPITAL COST SUMMARY

The capital expenditure schedule for the life of mine is shown in Table 1-10 below.

Table 1-10: Capital Expenditure Schedule


Capital Expenditure, $000	Initial	Expansion	Sustaining
Capital Expenditure, $000	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7
Pre-stripping	$28,226
Mine	$69,143	$16,485	$3,506	$600	$583	$605	$642	$682
Process	$90,214	$67,171	$6,044	$5,621	$2,098
Owner's Cost	$6,420	$4,649
Total	$194,004	$88,304	$9,551	$6,220	$2,681	$605	$642	$682

1.14 OPERATING COST SUMMARY

The total production cost includes mine operations, process plant operations, general administration, reclamation and closure, and government fees. Table 1-11 below shows the operating costs over the LOM by area.

Table 1-11. LOM Operating Costs


LOM Operating Cost ($000)
Mining	$348,505
Process Plant	$156,936
G&A	$33,637
Refining	$4,679
Total Operating Cost	$543,757
Royalty	$16,282
Reclamation/Closure	$49,094
Total Production Cost	$609,133

1.15 CONCLUSIONS AND RECOMMENDATIONS

The results of this study indicate that South Railroad is both technically and economically feasible and demonstrate robust returns, even at the conservative metal prices. The authors recommend that the South Railroad be advanced to feasibility level, with a list of specific recommendations to achieve that goal (see Section 26).

Presently there are 1.25 million probable and proven ounces of gold in the Dark Star and Pinion deposits estimated mineral reserves combined, 1.55 million measured and indicated ounces of gold in the Dark Star, Pinion and North Bullion deposits estimated mineral resources combined, inclusive of mineral reserves in the Dark Star and Pinion deposits, and there are 1.20 million inferred ounces in the Dark Star, Pinion, Jasperoid Wash and North Bullion deposits estimated mineral resources combined. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

The PFS indicates an average gold production over the estimated 8-year life of mine of about 116,986 ounces per year, with peak production in Year 3 of 294,316 ounces of gold. Cash costs are estimated to be $582 per ounce of gold after by-product credit, and all-in sustaining-capital costs (AISC) are estimated to be $657 per ounce of gold. The resulting after-tax cash flow is $337.1 million, for an after-tax NPV (5%) of $241.5 million and an estimated payback period of 2.7 years. A summary of the pre-tax and after-tax PFS economic indicators is shown in Table 1-12. Pre-tax cash flow is summarized in Table 1-13.


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Table 1-12: Economic Analysis Summary


Indicators ($000)	Before Tax	After Tax
LOM Cash Flow	$409,665	$337,113
NPV @ 5%	$302,081	$241,474
NPV @ 10%	$217,392	$166,153
IRR	32.4%	27.8%
Payback (years)	2.6	2.7


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Table 1-13: Pre-Tax Cash Flow


	LOM	Year-1	Year1	Year2	Year3	Year4	Year5	Year6	Year7	Year8
Payable Metals
Gold(kozs)	931		31	165	293	135	120	79	64	38
Silver(kozs)	1,040		-	-	-	31	224	327	303	141
Revenues ($000)
Gold	$1,303,691		$43,880	$231,385	$409,983	$189,341	$168,111	$111,150	$89,830	$53,376
Silver	$17,796		$0	$0	$0	$539	$3,837	$5,595	$5,186	$2,409
Total Revenues	$1,321,487		$43,880	$231,385	$409,983	$189,880	$171,948	$116,745	$95,017	$55,784
Operating Cost ($000)
Mining	$348,505		$54,111	$54,301	$52,905	$53,790	$52,132	$46,731	$26,548	$7,987
Process Plant	$156,936		$7,896	$23,845	$26,652	$25,357	$22,395	$18,714	$19,270	$11,137
G&A	$33,637		$3,716	$3,704	$3,680	$3,680	$3,680	$3,680	$3,680	$3,651
Refining	$4,679		$158	$831	$1,472	$680	$603	$399	$322	$192
Total Operating Cost	$543,757		$65,880	$82,681	$84,709	$83,506	$78,811	$69,523	$49,821	$22,966
Royalty	$16,282		$541	$2,851	$5,051	$2,339	$2,119	$1,439	$1,171	$687
Salvage Value	$0		$0	$0	$0	$0	$0	$0	$0	$0
Reclamation/Closure	$49,094		$2,183	$2,183	$3,876	$3,876	$3,422	$3,422	$1,729	$1,729
Total Production Cost	$609,133		$68,604	$87,714	$93,636	$89,722	$84,351	$74,384	$52,721	$25,382
Operating Income	$712,353		-$24,725	$143,671	$316,347	$100,158	$87,597	$42,361	$42,296	$30,402
Depreciation ($000)
Initial Capital	$194,004		$27,723	$47,511	$33,931	$24,231	$17,325	$17,305	$17,325	$8,653
Sustaining Capital	$108,685		$12,619	$22,991	$18,672	$14,606	$10,909	$10,215	$9,988	$5,940
Total Depreciation	$302,688		$40,342	$70,502	$52,604	$38,837	$28,234	$27,520	$27,313	$14,593
Net Income after Depreciation	$409,665		-$65,066	$73,169	$263,743	$61,321	$59,364	$14,841	$14,983	$15,809
Cash Flow ($000)
Net Income before Taxes	$409,665	$0	-$65,066	$73,169	$263,743	$61,321	$59,364	$14,841	$14,983	$15,809
Add back Depreciation	$302,688	$0	$40,342	$70,502	$52,604	$38,837	$28,234	$27,520	$27,313	$14,593
Operating Cash Flow	$712,353	$0	-$24,725	$143,671	$316,347	$100,158	$87,597	$42,361	$42,296	$30,402
Initial Capital Expenditures ($000)
Pre-stripping	$28,226	$28,226	$0	$0	$0	$0	$0	$0	$0	$0
Mining	$69,143	$69,143	$0	$0	$0	$0	$0	$0	$0	$0
Process	$90,214	$90,214	$0	$0	$0	$0	$0	$0	$0	$0
Owner's Cost	$6,420	$6,420	$0	$0	$0	$0	$0	$0	$0	$0
Expansion Capital Expenditures ($000)
Mining	$16,485		$16,485	$0	$0	$0	$0	$0	$0	$0
Process	$67,171		$67,171	$0	$0	$0	$0	$0	$0	$0
Owner'sCost	$4,649		$4,649	$0	$0	$0	$0	$0	$0	$0
Sustaining Capital Expenditures ($000)
Mining	$6,617		$0	$3,506	$600	$583	$605	$642	$682	$0
Process	$13,763		$0	$6,044	$5,621	$2,098	$0	$0	$0	$0
Total Capital	$302,688	$194,004	$88,304	$9,551	$6,220	$2,681	$605	$642	$682	$0
Cash Flow before Taxes ($000)	$409,665	-$194,004	-$108,816	$130,364	$305,400	$103,409	$87,098	$42,468	$40,590	$29,269
Cumulative Cash Flow before Taxes ($000)		-$194,004	-$302,820	-$172,456	$132,944	$236,353	$323,451	$365,919	$406,509	$435,779
Financial Indicators before Taxes ($000)
NPV @0%	$409,665
NPV @5%	$302,081
NPV @10%	$217,392
IRR	32.4%
Payback (years)	2.6					-	-	-	-	-


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2 INTRODUCTION AND TERMS OF REFERENCE

2.1 PURPOSE OF REPORT

This NI 43-101 Technical Report was prepared by M3 for Gold Standard of Vancouver, British Columbia, a corporation that is listed in TSX Venture Exchange (TSX.V: GSV) and the New York Stock Exchange (NYSE: GSV).

Gold Standard owns the Railroad-Pinion project in the southern Carlin trend, in Elko County, Nevada, USA.

This Technical Report describes the preliminary feasibility of extracting and processing the oxide mineral reserve at the South Railroad property, which includes the Dark Star and Pinion gold deposits. This study includes the updated 2019 mineral resource and mineral reserve estimates for the Dark Star and Pinion gold deposits, and initial mineral resource estimate for the Jasperoid Wash deposit. The current estimate supersedes the 2018 mineral resource estimate, as well as previous mineral resource estimates reported by Dufresne et al. (2017) and Dufresne and Nicholls (2018), APEX of Edmonton, Alberta (Dufresne and Nicholls, 2016), Dufresne and Nicholls (2017a; 2018), and APEX (Dufresne et al., 2015).

The North Railroad portion of Gold Standard’s property includes the POD (formerly Railroad deposit), Sweet Hollow, and North Bullion cluster of gold deposits. Together these three deposits are referred to as the North Bullion deposits or North Bullion area. The first-time estimates of POD, Sweet Hollow, and North Bullion gold mineral resources were originally reported by Dufresne and Nicholls (2017b). The POD, Sweet Hollow, and North Bullion mineral resources, as amended and restated in Dufresne and Nicholls (2018) remain current and are presented herein.

Other targets mentioned in this Technical Report include Bald Mountain, in the North Railroad portion of the Railroad-Pinion property, and JR Buttes, Dixie, Irene, Sentinel, Ski Track, and East Jasperoid in the South Railroad portion of the Railroad-Pinion project.

References to Tomera Formation equivalent stratigraphy have been noted historically. However, recent work suggests these units in the Railroad-Pinion area may not be of equivalent age, so all usage of Tomera Formation equivalent in this Technical Report refer to units that are Pennsylvanian-Permian undifferentiated.

This Technical Report has been prepared in accordance with the disclosure and reporting requirements set forth in NI 43-101 Companion Policy 43-101CP, and Form 43-101F1, as well as with the Canadian Institute of Mining, Metallurgy and Petroleum’s “CIM Definition Standards - For Mineral Resources and Reserves, Definitions and Guidelines” (“CIM Standards”) adopted by the CIM Council on May 10, 2014.

2.2 SOURCES OF INFORMATION

In compiling the background information for this Technical Report, the authors fully relied on information provided by Gold Standard and on other references as cited in Section 3, including technical reports by APEX (Dufresne and Turner, 2014; Dufresne et al., 2014; Dufresne et al., 2015; Dufresne and Nicholls, 2016; Dufresne et al., 2017; and Dufresne and Nicholls, 2017a, 2017b, 2018).

The Pinion, Dark Star, and Jasperoid Wash mineral resource estimates presented in this Technical Report were estimated and classified under the supervision of Mr. Steven J. Ristorcelli, C.P.G. and Principal Geologist for MDA. Mr. Thomas L. Dyer, P.E., Senior Engineer for MDA, prepared the mining and economic studies for the PFS.

Table 2-1 is a list of qualified persons who contributed to this Technical Report.


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Table 2-1: List of Qualified Persons


QP Name	Company	Qualification	Site Visit Date	Area of Responsibility
Art S. Ibrado	M3 Engineering & Technology Corporation, Tucson AZ.	PE	September 25, 2019	Sections 1.1 to 1.3, 1.15, 21.5.3.3, 2 to 5, 21.5.4, 23 to 26
Mathew Sletten	M3 Engineering & Technology Corporation, Chandler AZ.	PE	No site visit	Sections 1.10, 1.12 to 1.14, 18.1 to 18.5, 19, 21.3, 21.5.4 & 22
Steven Ristorcelli	Mine Development Associates, Reno NV	CPG	November 18, 2016	Sections 1.4, 1.5, 1.8, 6, 7, 8, 9, 10, 11, 12, 14.1, 14.3, & 14.4
Michael Lindholm	Mine Development Associates, Reno NV	CPG	September 19, 2018	Sections 1.4, 1.5, 1.8, 6, 7, 8, 9, 10, 11, 12, 14.1 & 14.2
Michael B. Dufresne	APEX Geoscience Ltd., Edmonton, AB	P.Geol., P.Geo.	June 7-9, 2017	Sections 1.4, 1.5, 10.8, 11.5, 12.9, 10.2, 10.6.1, 10.6.3, 11.1, 11.2, 12.3, 12.8 & 14.5
Thomas Dyer	Mine Development Associates, Reno NV	PE	November 18, 2016	Sections 1.8, 1.9, 15, 16, 21.1, & 21.4
Gary L. Simmons	GL Simmons Consulting, LLC	QP-MMSA	June 22, 2017	Sections 1.6, & 13
Carl Defilippi	Kappes, Cassiday & Associates, Reno NV	RM-SME	August 28, 2019	Sections 1.6.5, 1.7, 13.10, 17.0 to 17.7, 17.9, 17.10, 21.2, & 21.5 except 21.5.3.3 & 21.5.4
Richard DeLong	EM Strategies, Inc., Reno NV	QP-MMSA, RG, PG	No site visit	Sections 1.11 & 20
Kenneth L. Myers	The MINES Group, Reno NV	PE	October 22, 2019	Sections 17.8, 18.6, 18.7, 18.8

2.3 PROJECT SCOPE AND TERMS OF REFERENCE

Gold Standard has been actively exploring the North Railroad portion of the property since 2010 and the South Railroad portion of the property since 2014 (Koehler et al., 2014; Turner et al., 2015).

The North Bullion deposits’ mineral resource estimates (POD, North Bullion, and Sweet Hollow deposits) were prepared and classified under the supervision of Mr. Michael B. Dufresne, M.Sc., P. Geol., of APEX.

The scope of this study includes a review of pertinent technical reports and data provided to MDA by Gold Standard relative to the general setting, geology, project history, exploration activities and results, methodology, quality assurance, interpretations, drilling programs, metallurgy, and estimated mineral resources.

The authors have relied almost entirely on data and information derived from work done by Gold Standard and its predecessor operators of the amalgamated South Railroad and North Railroad portions of the Railroad-Pinion property. The authors have reviewed much of the available data and made site visits and have made judgments about the general reliability of the underlying data. Where deemed either inadequate or unreliable, the data were either eliminated from use or procedures were modified to account for lack of confidence in that specific information. The authors have made such independent investigations as deemed necessary in their professional judgment to be able to reasonably present the conclusions discussed herein.

The effective date of this PFS is September 9, 2019, and the issue date of the Technical Report is October 24, 2019. The effective dates of the Pinion and Dark Star databases on which the mineral resources described in this Technical Report are estimated are May 31, 2019 and April 26, 2019, respectively. The effective dates of the Pinion and Dark Star mineral resource estimates are both August 7, 2019. The effective dates of the Jasperoid Wash database and


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mineral resource estimate are October 6, 2018 and November 15, 2018, respectively. The effective date of the North Bullion deposits database and mineral resource estimates are August 18, 2017 and September 15, 2017, respectively.

2.4 FREQUENTLY USED ACRONYMS, ABBREVIATIONS, DEFINITIONS, AND UNITS OF MEASURE

In this Technical Report, measurements are generally reported in metric units. Where information was originally reported in imperial units, MDA has made the conversions as shown below. In the case of metallurgical test data and historical mineral resource estimates the units are as originally reported in order to preserve historical accuracy and avoid errors that can result from rounding converted data.

Currency, units of measure, and conversion factors used in this Technical Report include:


Linear Measure
1 centimeter	= 0.3937 inch
1 meter	= 3.2808 feet	= 1.0936 yard
1 kilometer	= 0.6214 mile

Area Measure
1 hectare	= 2.471 acres	= 0.0039 square mile

Capacity Measure (liquid)
1 liter	= 0.2642 US gallons

Weight
1 ton	= 1 imperial short ton	=2,000 pounds
1 tonne	= 1.1023 short tons	= 2,205 pounds or
		= 1,000 kilograms
1 kilogram	= 2.205 pounds

Regarding currency, unless otherwise indicated, all references to dollars ($) in this Technical Report refer to currency of the United States.

Frequently used acronyms and abbreviations are as shown in Table 2-2.

Table 2-2: Acronyms and Abbreviations


Abbreviation	Description
2SD	two times the standard deviation
3SD	three times the standard deviation
AA	atomic absorption spectrometry
ABA	acid-base accounting
Ag	silver
AgCN	cyanide-soluble silver
AgFA	silver analysis by fire assay, total silver content
Au	gold
AuCN	cyanide-soluble gold
AuFA	gold analysis by fire assay, total gold content
Calc, calc	calculated
CINO	inorganic carbon
cm	centimeters
core	diamond core-drilling method
^°C	degrees Celsius


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Abbreviation	Description
Ext	extracted
°F	degrees Fahrenheit
FA	fire assay
ft	foot or feet
gal	gallon(s)
g	gram
gpl	grams per liter
g/t	grams per metric tonne
Ha	hectares
hd	head
Hr., hr., hrs	hour, hours
ICP	inductively-coupled plasma-emission spectrometric method
ICP-MS	inductively-coupled plasma-emission and mass spectrometry
in	inch or inches
kg	kilograms
km	kilometers
kWh/m³	kilowatt-hours per cubic meter
l	liter (L in metallurgical use)
lb or lbs.	Pounds
m	Meters
Ma	million years
mi	mile or miles
mm	millimeters
µm	micron or 10-6 meters
NAG	non-acid generating, (neutralizing potential)
NSR	net smelter return
opt	troy ounce per short ton
org	Organic
oz	troy ounce
P80	the theoretical square screen-opening, through which 80 weight percent of the particles will pass.
PAG	potential acid generating
ppm	parts per million
ppb	parts per billion
QA/QC	quality assurance and quality control
RC	reverse-circulation drilling method
RQD	rock-quality designation
SO4	Sulfate
st	Imperial short ton (2,000 pounds)
SSUL	sulfide sulfur
t	metric tonne or tonnes
tot	total


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3 RELIANCE ON OTHER EXPERTS

Mr. Ekins, who is an independent registered professional landman (RPL#32306) and president of GIS Land Services in Reno, Nevada, assisted with the preparation of the summary land description and property maps discussed below. Mr. Ekins and Gold Standard have relied upon title opinions prepared by Mr. Jeff N. Faillers of Erwin Thompson Faillers, of Reno, Nevada, Mr. Richard Thompson of Harris & Thompson, of Reno, Nevada, and Ms. Tracy Guinand, an independent registered professional landman of Tracy Guinand Land LLC, of Reno, Nevada. The most recent of these title opinions are dated September 5, 2018. The opinions provided on surface ownership and subsurface mineral ownership, along with royalty information, are current as of the effective date of this Technical Report. Additional details with respect to the surface and subsurface ownership are provided in Gold Standard’s most recent Annual Information Form (“AIF”), which can be found on the SEDAR website at www.sedar.com.

The sample collection, security, transportation, preparation, and analytical procedures are judged by the authors to be acceptable and to have produced data suitable for use in the estimation of the mineral resources reported in Section 11, subject to those exclusions or modifications discussed in Section 14. The authors consider the procedures utilized by Gold Standard and the assay laboratories to be appropriate for use as described.


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4 PROPERTY DESCRIPTION AND LOCATION

4.1 LOCATION AND LAND AREA

The property that is the subject of this Technical Report comprises two contiguous areas of mineral tenure held by Gold Standard (Figure 4-1) that straddle the Piñon Range in the Railroad mining district at the southeast end of the Carlin trend, a northwest-southeast trending belt of prolific gold endowment in northern Nevada. In previous Technical Reports the northern portion of the land holdings, now referred to as the North Railroad portion of the property (Figure 4-1), has been referred to as the Railroad project and the Railroad property (Dufresne et al., 2017). The southern portion of the Railroad-Pinion property, now referred to as the South Railroad portion of the property (Figure 4-1), was referred to as the Pinion project and the Pinion property in previous technical reports (Dufresne et al., 2017). In November 2017, Gold Standard published a technical report on the Railroad-Pinion property, which included a mineral resource estimate for the North Bullion, POD, and Sweet Hollow gold deposits (Dufresne and Nicholls, 2017b), located in the North Railroad portion of the Railroad-Pinion property, approximately 10 km north of the Dark Star and Pinion deposits. Based on available information, North Bullion, POD, and Sweet Hollow would not likely share a common mining infrastructure with Dark Star and Pinion.

The North and South Railroad properties combined constitute a land position totaling 21,679 hectares, and with partial interests taken into consideration, 20,477 net hectares of land in Elko County, Nevada. The properties are located within Section 13 in Township (“T”) 28N, Range (“R”) 52E; Sections 10, 11, 14, 16, 17, 18, 23, and 24 in T28N, R53E; Sections 1 to 21, 23, 24, 25, 29, 30, 31, 33, 35, and 36 in T29N, R53E; Sections 7, 18, 19, and 30 in T29N, R54E; Section 12 in T30N, R52E; Sections 1 to 10, 13 to 33, and 36 in T30N, R53E; Sections 24 and 36 in T31N, R52E; and Sections 8, 10, 14 to 22 and 26 to 35 in T31N, R53E, as shown in Figure 4-2. Gold Standard owns, or otherwise controls 100% of the subsurface mineral rights on a total of 12,117 gross hectares of land held as patented and unpatented lode claims. This includes 1,455 unpatented claims owned by Gold Standard and 207 unpatented claims held under lease (Appendix B). Gold Standard also owns or leases 30 patented claims (Appendix B).

There is also a total of 9,562 gross hectares of private lands of which Gold Standard’s ownership of the subsurface mineral rights varies from 49.2% to 100% (Figure 4-2), for a net position of approximately 8,360 gross hectares.


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Figure 4-1: Location Map for the Railroad-Pinion Property	Figure 4-2: Railroad-Pinion Property with Ownership Percentages, Elko
	County, Nevada
(from Dufresne and Nicholls, 2017b)
	(from Dufresne and Nicholls, 2017b)


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Private surface and private mineral property are wholly owned and subject to lease agreement payments (see Section 4.2) and property taxes (paid on an annual basis) as determined by Elko County. The estimated holding cost for the patented claims, unpatented claims, and private lands held by Gold Standard is $1,103,402 per annum. Unpatented lode mining claims grant the holder 100% of the mineral rights and access to the surface for exploration activities which cause insignificant surface disturbance. Ownership of the unpatented mining claims is in the name of the holder (locator), subject to the paramount title of the United States of America, under the administration of the BLM. Under the Mining Law of 1872, which governs the location of unpatented mining claims on federal lands, the locator has the right to explore, develop, and mine minerals on unpatented mining claims without payments of production royalties to the U.S. government, subject to the surface management regulation of the BLM. Currently, annual claim-maintenance fees are the only federal payments related to unpatented mining claims. The mineral rights do not expire if the unpatented claims are maintained by paying an annual fee of $165 per claim to the U.S. Department of Interior, BLM prior to the end of the business day on August 31 every year. A notice of intent to hold must also be filed with the Elko County Recorder on or before November 1 annually, along with a filing fee of $12.00 per claim, plus a $4.00 document fee.

Gold Standard has completed its federal claim maintenance fee obligations for the owned and leased unpatented claims for 2019-2020 assessment year. The federal claim maintenance fees for the claims for the 2020-2021 assessment year are due on or before September 1, 2020. Gold Standard’s estimated claim maintenance fee cost for the owned and leased unpatented claims is $274,230 per annum, and the company’s total estimated annual cost to maintain its property package is $1,377,632.

AGREEMENTS AND ENCUMBRANCES

Portions of the unpatented and private lands are encumbered with royalties predominantly in the form of standard Net (or Gross) Smelter Return (“NSR” or “GSR”) and Mineral Production (“MP”) royalty agreements, or Net Profit Interest (“NPI”) agreements. The locations and aerial distribution of the currently relevant royalty encumbrances for the Railroad-Pinion property are shown in Figure 4-3. These are summarized as follows:

1.0% NSR royalty to Royal Standard Minerals, Inc. and Manhattan Mining Co. on the portion of the property acquired by statutory plan of arrangement;
1.5% MP royalty to Kennecott Holdings Corporation on claims noted as the Selco Group;
5.0% NSR royalty to the owners of the undivided private mineral interests;
Gold Standard owns an approximate 99.2% mineral interest in Sections 21 and 27 by way of several lease agreements. Pursuant to the terms of the relevant lease agreements, Sections 21 and 27 are subject to a 5.0% NSR royalty to the lessors of the leased property;
Section 22 is comprised of the TC 1 through 39, and TC 37R and 38R unpatented lode mining claims owned by Gold Standard. The TC claims are subject to the following royalties: (1) an unknown/unspecified NSR royalty to "GSI, Inc., of Virginia"; and (2) a 2.0%) NSR royalty to Waterton Global Value LP;
1.0% NSR royalty to Aladdin Sweepstake Consolidated Mining Company on the portion of the property acquired by statutory plan of arrangement, specifically the PIN#1 to PIN#12 lode mining claims;
4.0% NSR royalty to ANG Pony LLC for mining claims acquired by Gold Standard in Sections 34 and 36 in T30N, R53E, and Sections 2 and 4 in T29, R53E;
3.0% NSR royalty to Peter Maciulaitis for certain mining claims in Sections 24 and 26 in T30N, R53E;

A 3.0% NSR royalty (relative to mineral interest) to Linda Zunino and Tony Zunino, Trustees of the Delert J. Zunino and Linda Zunino Family Trusts dated October 11, 1994, and a 3.0% NSR royalty (relative to


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mineral interest) to John C. Carpenter and Roseann Carpenter, husband and wife, on Section 23 in T29N, R53E;

A 3.0% NSR royalty to Nevada Sunrise LLC on the 14 WMH claims situated in Sections 1, 2, 3, and 11 in T29N, R53E; and
A 3.5% NSR royalty (relative to mineral interest) to Dominek Pieretti and Tusca Sullivan on Sections 3, 5, 7, 8, 9, 10, 15, 17, 19, 21, 29, 31, and 33 in T29N, R53E, and Section 33 in T30N, R53E.

Thus, the estimated total annual cost for maintaining the Railroad-Pinion property is $1,377,632, including lease payments and claims maintenance.

(from Dufresne & Nicholls, 20147b)

Figure 4-3: Railroad-Pinion Property Map with Royalty Encumbrances


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4.3 ENVIRONMENTAL PERMITS

As of the effective date of this Technical Report, the authors are not aware of any significant factors or risks that may affect access, title, or the right or ability to perform work on the property. Gold Standard controls sufficient ground and has sufficient permitting in place to access the project and continue future exploration programs. Details on permitting are provided below. The following section discusses land use permitting and other regulatory information specific to the South Railroad portion of the property.

Gold Standard currently has three Plans of Operations (“PoO’s”) and two Notices of Intent (“NOI’s”) in place with the BLM for the Dark Star and Pinion areas of the South Railroad portion of the property (Figure 4-4). Gold Standard has drafted a Historic Properties Treatment Plan (“HPTP”) for archaeological sites within the Railroad Exploration Project Plan of Operations. The draft HPTP is currently being evaluated by the BLM.

Gold Standard represents that the PoO for the “South Railroad” portion of the Railroad-Pinion project was approved by the BLM on July 2016. The approved PoO covers a total of 3,422 ha with 2,119 ha of public land and 1,243 ha of private land located in Section 2 in T29N, R53E, and Sections 20, 21, 22, 23, 24, 25, 27, 28, 34, 35, and 36, and portions of Sections 14, 16, and 26 in T30N, R53E. Within the area of the PoO exploration-related disturbance and reclamation bonding can be conducted in two phases of up to 100 acres in Phase I and 150 acres in Phase II. A reclamation bond in the amount of $315,021 has been posted with the BLM.

The South Railroad PoO has encompassed two earlier Plans of Operation (Pinion and Greater Pinion), and two notices (Dark Star and Irene). These two plans and notices will be closed by the respective agency, either the BLM or the Nevada Department of Environmental Protection (“NDEP”), and the associated reclamation bonds will be returned to the Company. The bond amounts for these PoO’s and NOI’s are, respectively: $90,849, $40,241, $25,955, and $30,482.

Activities in Jasperoid Wash are permitted with an approved BLM NOI for an area situated in Section 16 in T29N, R53E. The bond for this NOI has been posted in the amount of $31,956.

4.3.1 Other Permits

Gold Standard has received a Water Pollution Control Permit (“WPCP”) issued by the NDEP for North Bullion that expires in November 2019. There is also a Surface Area Disturbance Permit (“SADP”) issued by NDEP for North Bullion.

4.3.1 Private Land Disturbance

As of the effective date of this Technical Report, Gold Standard has received a Reclamation Permit (“RP”) that includes the Pinion, Dark Star, and Irene reclamation plans. This RP covers both private land and public land disturbances. Previously approved reclamation plans associated with these areas will be closed by the respective permitting agency, either BLM or NDEP. These operated under an Interim Reclamation Permit (“IRP”) issued by the State of Nevada for disturbance greater than five acres on private land. The IRP allowed up to 11 acres of surface disturbance and covered portions of Sections 21 and 27 (not included in the PoO) in T30N, R53E. Gold Standard is in the process of applying for a permanent Reclamation Permit, and as of the effective date of this Technical Report, is permitted to operate in Section 25 in T30N, R53E at Dark Star as long as it keeps the total disturbance to less than five acres at any given time.

For the South Railroad portion of the Railroad-Pinion property, Gold Standard has received an SADP and Environmental Assessment (“EA”) (Figure 4-4) that covers approximately 19.3 km²and allows up to 101 ha of disturbance. The SADP EA expires on July 2022.


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(from Gold Standard, 2018)

Figure 4-4: Property Map with Railroad- Pinion Permit Boundaries


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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESS TO PROPERTY

The Railroad-Pinion property in the Piñon Range is accessed primarily from the four-lane transcontinental U.S. Interstate 80 (“I-80”), approximately 442 km west of Salt Lake City, Utah, and 467 km east of Reno, Nevada. The project is located between 13 and 29 km south of I-80 and can be reached by a series of paved and gravel roads from Elko, Nevada (population 18,300). From Elko, access is via I-80 west for 32 km to the town of Carlin (population 2,400), and then south on State Highway 278 for 24 km, turning east on a gravel road immediately north of Trout Creek and driving for another 13 km. Alternatively, the project area may also be reached during the summer and autumn months by traveling 48 km southwest from Elko on the Bullion Road, a county-maintained dirt and gravel road. Travel within the project area is via a network of historical and recently constructed dirt roads and four-wheel drive tracks.

5.2 CLIMATE

The project area has a relatively dry and cool, high-desert climate. Weather records from the Newmont Mining Corporation (“Newmont”) Carlin mine, 55 km to the north, indicate that from 1966 through 2002, the average January maximum and minimum temperatures were 1.3°C and -6.9°C, respectively. July average maximum and minimum temperatures were 28.4°C and 14.6°C, respectively.

Rainfall in the region is generally light and infrequent between May and October. At Emigrant Pass, 16 km west of the town of Carlin, Nevada and 20 km northwest of the property, average annual precipitation has been 32.8 cm with average precipitation on January and July of 3.7 cm and 0.6 cm, respectively (US Climate Data). Much of the annual precipitation occurs as snowfall during the winter months.

Precipitation can vary dramatically with changes in elevation and season. Moist airflow from the south brings summer thunderstorms from July through September. A small number of these storms may carry heavy rains that can cause localized flooding in creeks and drainages. Winter snow and spring runoff may temporarily limit local access with respect to drilling and other geological fieldwork activities between November and April each year but are not considered to be significant issues. Mining and exploration can be conducted year-round with adequate snow removal and maintenance of access roads.

5.3 PHYSIOGRAPHY

Northern Nevada is within the Basin and Range physiographic province, an area characterized by gently sloping valleys bounded by generally north-south-trending mountain ranges. The project area is located within and adjacent to the Piñon Range at elevations ranging from 1,770 m to nearly 2,650 m above sea level. Lower elevations consist of gentle, rolling hills with little to no bedrock exposure. Higher elevations are characterized by steeper slopes, deeply incised drainages, and an increase in bedrock exposure.

Vegetation largely consists of sagebrush, rabbit brush, small cacti, and bunch grass communities, consistent with a high-desert climate. Cottonwood trees are present in canyon and creek bottoms, and near springs. Pinyon pine, juniper, mountain mahogany, and aspen trees are present in some areas at higher elevations.

5.4 LOCAL RESOURCES AND INFRASTRUCTURE

Elko, Nevada is a small, full-service city based on mining, ranching, and transportation that has served as the center for northern Nevada mining and exploration for more than half a century. Housing, hotels, groceries, restaurants, clinics, and a hospital, industrial supplies, a regional airport with daily flights to and from Salt Lake City, Utah, interstate highway and railway, local, state and federal government offices, fuel, telecommunications, engineering services, light and heavy equipment sales and services, and a community college are all present.


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In this part of Nevada, there are local, regional, and international exploration and mining service companies, including assay laboratories, surveyors, suppliers, drilling contractors, and heavy equipment vendors supporting the exploration and mining industry. These companies are served by a skilled and experienced local labor force accustomed to the mining and exploration industries.

The North Railroad and South Railroad portions of the property are within 65 km of several large, active open-pit and underground mines operated by Newmont and Barrick Gold Corp. (“Barrick”) along the Carlin trend. These mine sites also include fully operational mill complexes designed to treat sulfide and/or carbon-sulfide refractory gold ores.

Water for drilling at Pinion, Dark Star, and Jasperoid Wash is available at the project site. For communications, a commercial cellular telephone and data network is available in select locations. High voltage electrical transmission lines are located 9.7 km from the property. There are sufficient and appropriate sites within the property to accommodate exploration and potential mining facilities, including waste rock disposal, heap-leach pads, and processing infrastructure. Surface rights controlled by Gold Standard are sufficient for potential mining operations.


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6 HISTORY

Historical exploration conducted at the North Railroad and South Railroad portions of the Railroad-Pinion property is summarized below and is largely derived from Dufresne and Nicholls (2016), Dufresne et al. (2017), Dufresne and Nicholls (2017b), Dufresne and Nicholls (2018), and other sources as cited. The authors have reviewed this information and believe it accurately represents the history of the property as presently understood. MDA has added details of drilling types, meterage and number of holes based on Gold Standard’s recently compiled project-wide database.

6.1 NORTH RAILROAD PORTION OF THE PROPERTY

This portion of the report is extracted and modified from Dufresne and Nicholls (2018) who provided the most recent summary of historical exploration in the North Railroad portion of the property using information taken largely from Hunsaker (2010, 2012a, 2012b), Shaddrick (2012), Koehler et al. (2014), and sources as cited. Details of types and amounts of drilling were derived from Gold Standard’s project-wide drill database.

The earliest prospecting and mineral exploration in the North Railroad portion of the property likely dates to the mid-1860s. In 1869, the Railroad mining district was established in the area of Bunker Hill and the district was also known as the Bullion or Empire City district (LaPointe et al., 1991). Initially silver, lead, and copper ore was shipped to Chicago and San Francisco. A smelter was built in 1872 at the nearby town of Bullion. Beginning in 1905, shipments from operating mines, old dumps, and slag were sent to Salt Lake City (Ketner and Smith, 1963).

Early production in the district was mainly silver, lead, and copper extracted from numerous underground mines on the northern flank of Bunker Hill. Emmons (1910) reported that the mines were reopened in 1906 and at the time of his review the Standing Elk, Tripoli, Red Bird, Copper Belle, and Delmas mines were accessible. The most important mines exploited replacement and skarn-type deposits in marbleized and dolomitized rocks in the vicinity of the Bullion stock (see Section 7.3). There were also minor, undeveloped gold veins in intrusive rocks.

Beginning in 1910, and until the mines quit production in the 1960s, zinc became the prominent metal mined (LaPointe et al., 1991). In 1905, the Davis tunnel was started from a location approximately 1,340 m northeast and 305 m below the 600 level of the Standing Elk mine. Many lessors worked at extending the tunnel, which was driven southwest to reach a zone beneath the Standing Elk. In 1959, the zone was reached but no ore was found. Numerous oxidized faults and oxidized zones of base-metal mineralization were crossed.

Modern-era exploration began in 1967 when American Selco optioned claims from Aladdin Sweepstake Consolidated Mining, launching a period of surface sampling, geophysics, geological mapping, and surface drilling in the Railroad district and the North Railroad portion of the property that has continued to the date of this Technical Report. Records are incomplete but historical exploration was likely conducted in various areas at various times by 15 companies. These companies collected 6,260 soil samples, 3,508 rock samples, and drilled 382 holes, according to Dufresne and Nicholls (2018). Drilling in the North Railroad portion of the property by these operators is discussed in Section 10.2.

American Selco, Placer Amex, and El Paso Natural Gas Company with Louisiana Land and Exploration Company explored for porphyry copper and molybdenum in the Bullion stock and Grey Eagle intrusive rocks. They also looked for replacement sulfide “lenses” in limestone and “unknown replacement or disseminated” mineralization west of the Bullion stock (American Selco, 1970). American Selco completed an induced potential (“IP”) and magnetic geophysical surveys and drilled seven core holes and seven holes of unknown type. Subsequent core holes, and several of the rotary holes completed to the desired depths, intersected up to 50% sulfides as well as molybdenum, copper, silver, and gold.

Placer Amex drilled a single 365.8 m hole in 1972. In 1974, El Paso Natural Gas Company drilled 671.5 m in five holes of unknown type with the Louisiana Land and Exploration Company. In 1975, AMAX Inc. (“AMAX”) optioned the claims and explored for tungsten, molybdenum, and base metals until 1980. Detailed mapping was completed in Sections 27,


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28, 29, 32, 33, and 34 in T31N, R53E, and in Sections 3, 4, 5, 6, 8, and 9 in T30N, R53E (Dufresne and Nicholls, 2017b). AMAX also conducted surface dump and rock chip sampling, soil sampling, a vegetation geochemical survey, a ground magnetometer survey and drilled two core holes and 13 holes of unknown type in 1977 through 1980.

In 1980, Homestake Mining Company (“Homestake”) entered into a joint venture arrangement with AMAX and exploration was focused on gold in the North Railroad area, particularly after Newmont discovered the Rain gold deposit about 10 km north of Bunker Hill. Homestake drilled 22 holes (Galey, 1983) and collected rock and soil samples. The Homestake work identified what later became known as the POD deposit.

NICOR Mineral Ventures, Inc. (“NICOR”) became AMAX’s joint venture partner in 1983. As operator, NICOR continued the geologic mapping, soil geochemistry, and drilling initiated by Homestake. NICOR drilled 102 rotary and reverse-circulation holes (“RC”) in 1983 through 1986, expanding the drill coverage at the POD deposit and estimating a mineral resource (see Section 6.3.3).

In 1986, Westmont Mining Inc. (“Westmont”) took over NICOR’s interests and operated until 1993. Some of NICOR’s rock and soil sample data are recorded as Westmont data. Westmont drilled 42 RC holes and 31 holes of unknown type in the POD, North Bullion, Bald Mountain, and north of North Bullion areas during 1987-1992 and collected rock and soil samples. They developed a detailed stratigraphic interpretation for the late Paleozoic sedimentary units and also began to recognize low-angle reverse and low-angle normal faults, as well as prominent north-south-trending and northwest-trending high-angle normal and reverse faults. The interplay of the Webb-Devils Gate contact and complex faulting as controls to the mineralization were also identified late in the Westmont tenure.

At some time prior to 1993, Corona Gold (“Corona”) reported on land held jointly with Pezgold mineral resource Corporation (“Pezgold”) in Sections 16 and 20 in T31N, R53E, and which later became part of the Railroad-Pinion property. Available data indicate that six holes were drilled and geologic mapping, soil and rock sampling, and geophysics were conducted. Gold Standard’s drilling database does not contain drill holes attributed to Corona or Pezgold. Specific drill locations are not known, and drill data indicates all the holes remained in the Mississippian Webb Formation above the target horizons. A northeast-trending corridor of subtle Carlin or Rain-type alteration and weak geochemistry was noted.

The Corona Gold area was acquired by Newmont in 1993. According to Dufresne and Nicholls (2017b), two holes were drilled, and additional geophysical surveys were conducted. The drilling reached as deep as 425 m, but this was not deep enough to reach the target horizon in those locations. These holes are not in Gold Standard’s drilling database. Gravity data outlined the northeast-trending zone and indicated a significant fault in the northeast corridor.

Ramrod Gold (“Ramrod”) became operator in 1993 and drilled 10 RC holes in the POD-North Bullion area in 1994. Newmont drilled one hole north of the POD-North Bullion deposit in 1995.

Mirandor Exploration (U.S.A.) Inc. (“Mirandor”) operated the project in 1996-1997 and drilled 42 RC holes in the POD-North Bullion, Bald Mountain, and north of North Bullion areas. The exploration for Ramrod and Mirandor was carried out by geologists who were previously employed by Westmont.

The Ramrod and Mirandor drilling tested greater depths than their predecessors and showed encouraging results along the northwest-trending POD zone. Elsewhere, the EMRR series of drill holes returned favorable results adjacent to the historic Sylvania mine which had historic production from replacement/skarn mineralization. Ramrod and Mirandor’s deeper drilling and drill hole placement encountered higher gold values than earlier drilling.

Kinross Gold U.S.A, Inc. (“Kinross”) took over the project during 1998 and 1999 under the terms of an earn-in agreement with Mirandor. Kinross drilled 64 RC holes and one core hole in the POD-North Bullion and Bald Mountain areas, according to the Gold Standard database and collected 871 rock and 2,531 soil samples according to Dufresne


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and Nicholls (2017b). The soil samples were collected using a uniform collection process and the analysis of both soils and rocks was consistent in analytical laboratory and procedures.

The Kinross surface-sample results were consistent with known mineralization geology across most of the historical project area. Gold in soil anomalies from the Kinross samples generally coincides with the known historical drilling. Ag, As, Sb, and Hg also gave a similar pattern and highlight the known areas. Cu, Pb, Zn, and Mo highlight the historical skarn and replacement area (Dufresne and Nicholls, 2017b). The Kinross drilling tested within the areas of historical estimates, on the extensions of those zones, as well as newer target areas. The results in the areas of known gold returned similar results (Bartels, 1999). Step-out drilling appeared to be encouraging. Kinross drilled deeper holes, and in many cases, tested more of the stratigraphy than had been tested by previous operators.

The authors have no information on historical exploration, if any, carried out in the North Railroad portion of the property from 1999 until 2007. In 2007 and 2008, Royal Standard Minerals (“RSM”) drilled four core holes in the Bald Mountain area.

RSM, or its North Railroad property, was acquired by Scorpio Gold Corporation (“Scorpio”). In 2009, Gold Standard acquired the North Railroad property of Scorpio Gold and various private investors. MDA is not aware if this was the entire North Railroad portion of the current property, or if parts of the current North Railroad portion were acquired subsequent to 2009.

Gold Standard commenced exploration in the North Railroad portion of the property in 2009 (see Section 9.1 and Section 9.3) and began drilling in the North Bullion area in 2010 (see Section 10.4.1.1).

6.2 SOUTH RAILROAD PORTION OF THE PROPERTY

Various parts of the current South Railroad portion of the property have been held by at least 15 different successive operators at various times. The summaries in Table 6-1 and Table 6-2 provide a timeline of the historical operators that held ownership of various portions and conducted historical exploration work. In some cases, historical project and property names, and boundaries have been applied in different forms than have been in use by Gold Standard over the last several years. Drilling by historical operators is summarized in Section 10.

6.2.1 Pinion Area Exploration History

Exploration activity at the Pinion area dates back to the discovery of the Pinion prospect by Newmont in 1980. Newmont referred to the prospect as Trout Creek. The majority of the historical work was conducted in the late 1980s and early to mid-1990s and overlaps somewhat with that of the adjacent North Railroad portion of the property. Historical exploration work conducted in the Pinion area is summarized in Table 6-1. This work identified a Carlin-type gold deposit at the Pinion prospect in Sections 22 and 27 in T30N and R53E, which for a time was known as the South Bullion deposit. An additional zone of gold-silver mineralization was discovered and partly delineated at the Dark Star prospect in Section 25 in T30N, R53E.

Historical drilling began in 1980 with RC methods. Drilling by historical operators is summarized in Section 10. The mineral resource estimates mentioned in Table 6-1 and in Section 6.3 were estimated prior to the introduction of the standards set forth in NI 43-101 and are not in accordance with NI 43-101. The authors of this Technical Report have referred to these estimates as “historical resources” and are not treating them, or any part of them, as current mineral resources or mineral reserves. There is insufficient information available to properly assess data quality, estimation parameters, and standards by which the estimates were categorized, and the authors have not done sufficient work to classify these historical mineral resources as current mineral resources. The historical mineral resource estimates described above should not be relied upon and are relevant only for historical completeness.


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Historical exploration in the Pinion area identified two discrete zones of mineralization (Main and North) with the majority of the historical drilling having been completed at the Main zone, including the testing of the jasperoid breccia outcrops located near the southern boundary of Section 22 in T30N and R53E. Historical drilling extended the Main zone gold mineralization well into Section 27 to the southeast. The north zone is located approximately ~300 m northeast of the Jasperoid outcrops of the Main zone.

In 2014, Gold Standard acquired a large portion of the Pinion and surrounding area mineral rights from Scorpio.

Subsequently, Gold Standard expanded their land position to include all of South Railroad.


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Table 6-1 Summary of Historical Exploration, Pinion Area

(modified from DeMatties, 2003; Dufresne and Nicholls, 2017; and with data from Gold Standard, 2018 and 2019)


Year	Company	Exploration Work Performed
1980	Newmont	- Conducted a regional stream sediment survey within the Piñon Range, which revealed a geochemical anomaly along Trout Creek. - Further prospecting and discovery of the baritic jasperoid, Pinion Main zone.
1980-1981	Cyprus Exploration Co. (“Cyprus”)/ AMOCO	- 31 RC drill holes in Au-bearing jasperoid outcrops and soil geochemical anomalies.
1983	Freeport-McMoRan (“Freeport”)	- 8 RC holes; each intersected gold.
1985	Santa Fe Mining (“Santa Fe”)	- 14 RC holes were drilled
1987-1989	Newmont	- 61 RC holes, estimation of historical mineral resource known as South Bullion, 20 million tons grading 0.89 g Au/t * (see discussion below in Section 6.2.4).
1988	Battle Mountain Gold (“Battle Mountain”)	12 holes of unknown type were drilled.
1987-1989	Teck Resources (“Teck”)	- 39 RC drill holes, geological mapping, and performed a soil geochemical survey.
1989-1991	Westmont Resources (“Westmont”)	11 holes of unknown type were drilled.
1990-1994	Crown Resources (“Crown”)	- 130 RC holes, conducted metallurgical testing, detailed mapping, rock chip sampling, 800 soil samples, controlled-source audio-magnetotelluric (“CSAMT”), and an airborne magnetic-electromagnetic-radiometric survey. - Defined two small and shallow mineralized zones: Pinion Main and “Central" zone, also known as Pinion North zone; estimated historical mineral resource of 8.1 million tons @ 0.026 oz Au/ton* (see discussion below in Section 6.2.4).
1994-1995	Cyprus Mining (“Cyprus”)	- 914 rock samples, compiled geochemical results of previous exploration, identified Au anomalies defining the Ridge zone and Northern Extension. - 74 RC holes in the South Bullion mineral resource area, expanded the historical mineral resource to 31 million tons at 0.89 g Au/t * (see discussion below in Section 6.2.4).
1996	Crown/Royal Standard Minerals Inc. (“RSM”)	- 225 rock chip samples along 100 ft spaced lines; conducted geologic mapping, drilled 7 diamond-core holes at the Main zone and North (Pod) zone not in the Gold Standard database; produced a historical mineral resource and preliminary scoping study.
1997-1999	Crown/RSM/Cameco	- Conducted geologic mapping, CSAMT surveys, rock chip sampling. Cameco drilled 18 RC holes and 8 core holes were completed; some may have started with RC.
1998-1998	Kinross Gold	- 1 RC hole and 1 hole of unknown type were drilled.
2002-2011	RSM	- 2003 drilled 10 RC holes, conducted metallurgy work with samples from drilling and trenches; obtained density measurements indicating historical mineral resources could have been understated. In 2007, 5 core holes drilled to determine the water table and to characterize the neutralization and acid generating potential of the mineralized and waste rocks. - Proposed leach pad drill testing was completed in 2007, holes not in database.


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6.2.2 Dark Star Area Exploration History

The Dark Star deposit is located approximately 3 km east of the Pinion Main zone (Figure 4-1). Historical exploration work was conducted at the Dark Star area from 1990 through 1999 by Crown, Westmont, Cyprus, Cameco and RSM, Mirandor, and Kinross, as summarized in Table 6-2. In 1990, Crown identified a surface gold anomaly through rock and soil sampling in what became the Dark Star deposit.

Drilling in 1991 confirmed the presence of subsurface gold mineralization at Dark Star. Further historical drilling identified an approximately north-south-trending mineralized zone that became known as the Dark Star Corridor. Historical drilling is summarized in Section 10.

Table 6-2 Summary of Historical Exploration in the Dark Star Area


Year	Company	Exploration
1984	Cyprus-AMAX	9 rotary holes drilled
1990	Crown	- Discovery and definition of Dark Star surface mineralization with rock chip and soil samples.
1991	Crown, Westmount Resources Inc.	- 38 holes; detailed rock and soil sampling; geologic mapping; drilled 6 reconnaissance holes peripheral to the Dark Star deposit. - 3 holes north of the Dark Star mineralized zone.
1992	Crown	- 33 holes; detailed CSAMT survey; detailed palynology studies to better define Dark Star stratigraphy; - Estimated mineral resource of 7.0 million tons at 0.75 g Au/t or 154,00 oz of Au* in Section 25 (Calloway, 1992).
1994	Crown	-updated estimated mineral resource of Pan Antilles Resources of 7.5 million tons 0.69 g Au/t or 151,000 oz Au* (McCusker and Drobeck, 2012).
1994-1995	Cyprus	- 9 holes drilled to the north of the Dark Star mineralized zone (not in Gold Standard’s database); - Estimated a mineral resource of 7.7 million tons at 0.69 g Au/t or 170,000 oz Au*.
1997	Cameco, RSM	- Gradient IP/Resistivity survey completed between Dark Star and Dixie
1997	Mirandor	- 11 holes drilled north and west of Dark Star mineralized zone.
1998	Kinross, Mirandor	- 7 holes drilled just north of mineralized zone.
1999	Kinross, Mirandor	- 6 holes drilled northwest of mineralized zone.

* The mineral resource estimates summarized in Table 6-2 were calculated prior to the introduction of the standards set forth in NI 43-101 and are not in accordance with that Instrument. The authors of this Technical Report have referred to these estimates as “historical resources” and are not treating them, or any part of them, as current mineral resources. There is insufficient information available to properly assess data quality, estimation parameters and standards by which the estimates were categorized, and the authors have not done sufficient work to classify these historical mineral resources as current mineral resources. The historical mineralresource estimates described above should not be relied upon and are relevant only for historical completeness.

6.2.3 Jasperoid Wash

The Jasperoid Wash prospect is located 7.5 km southwest of the Dark Star deposit (Figure 4-1). In 1988, Westmont conducted geologic mapping, and rock and soil sampling over the Jasperoid Wash and Black Creek regional area. The geochemical sampling identified a large anomalous mineralized system and a 13-hole RC drilling program followed in 1989. Nine of the 13 holes drilled in 1989 intersected intervals of ≥0.01 to 0.03 oz Au/ton (0.34 to 1.03 g Au/t). Follow-up drilling programs were conducted in 1990, 1991, and 1992 by drilling 34 RC and three core holes. Low-grade gold mineralization was intersected in 22 of the holes (Jones and Postlethwaite, 1992). This historical drilling is summarized in Section 10.1.

In 1997, Cameco collected 35 rock-chip samples to test the anomaly within the hydrothermally altered Diamond Peak and Chainman-Dale Canyon formations of the Jasperoid Wash prospect. Four RC holes were drilled, totalling 556.3


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m, targeting structural intersections. Significant gold mineralization was not intersected in the 1997 drilling at Jasperoid Wash, although two of the holes intersected low-grade, anomalous mineralization (Parr, 1998).

In 1998, Cameco completed gradient IP/Resistivity geophysical surveys over the Jasperoid Wash area and identified a large zone of low chargeability and high resistivity in the western part of the survey area. This was reportedly tested in 1998 by four RC holes totalling 677 m. Significant gold mineralization was not intersected in the drilling, although two of the drill holes intersected low-grade anomalous gold (Parr, 1999).

6.2.4 Other Prospects in the South Railroad Portion of the Property

Historical exploration has taken place intermittently since 1980 at several locations approximately 3.5 to 7.5 km southwest and south of the Dark Star and Pinion deposits as summarized below, and at the Irene prospect west of the Pinion deposit.

6.2.4.1 Dixie

The Dixie or Dixie Creek area, which is located 3.6 km south of the Dark Star deposit (Figure 4-1), has been explored intermittently since 1980 by various operators. The majority of the historical exploration work has been regional to semi-detailed in nature. In 1997, Cameco conducted rock sampling at the Pinion, Dark Star, and Dixie areas. The 1997 rock sampling at the Dixie area was intended to examine the nature of surface mineralization, in greater detail, and to compare this data with results of the then recently completed drill holes at the prospect. At the main Dixie area, a group of 32 rock samples defined a distinct, >1,500 ppb Hg anomaly with elevated Au, As, Sb, and Ag (Parr, 1999). This anomaly was found to roughly correspond with gold mineralization in the subsurface. Immediately to the north, a North Dixie anomaly was identified that was characterized by similar chemistry (elevated Hg, Au, As, and Sb). Farther north, a group of 15 rock samples collected between the Pinion and Dark Star areas defined a similar zone at the “CISS” area where six samples contained 20-135 ppb Au, including As values up to 940 ppm, Sb up to 161 ppm, and Hg up to 15 ppm.

In addition to the rock sampling, Cameco completed limited induced potential and resistivity (“IP/Resistivity”) geophysical surveys at several prospects including the Dixie area in 1997 and 1998. The IP/Resistivity surveys at Dixie identified broad zones of contrasting high and low resistivity, and corresponding zones of high chargeability (Parr, 1999).

The first documented drill program at the Dixie prospect was conducted by Freeport in 1988 and 1989, during which 26 holes were drilled in a joint venture with Crown. In 1991, Crown completed seven RC holes and later Cameco drilled 11 RC holes at the Dixie prospect. This historical drilling is summarized in Section 10. The drilling identified a zone of low-grade gold mineralization within Pennsylvanian siliciclastic and carbonate rocks above the contact between the Webb Formation and the underlying Devils Gate Limestone. This important contact between the Webb Formation and the underlying Devils Gate Limestone was not intersected by any of the historical Dixie Creek drilling (Redfern, 2002). The mineralization intersected at Dixie Creek is hosted in rocks that are similar in nature to the host rocks for the Dark Star gold mineralization (see Section 7.2.2.2).

6.2.4.2 JR Buttes

The JR Buttes prospect is located 4.5 km southwest of the Dark Star deposit. Geological mapping was completed over the JR Buttes area by an unknown company in 1977. This work outlined a zone of intense silicification over an interpreted graben structure (Dufresne and Nicholls, 2017a). In 1992, Westmount conducted rock chip, reconnaissance soil sampling, and detailed mapping, followed by a 19-hole RC drill program of 2,549.6 m. The drilling was designed to test for mineralization adjacent to the graben structural zone. Mineralization defined by silicification, arsenic, and gold concentrations was intersected along the western boundary fault of the graben. No mineralization was intersected


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along the eastern side of the graben (Jones and Postlethwaite, 1992). In 1994, Cyprus drilled three RC holes, and Cameco drilled one RC hole in 1998. The JR Buttes drilling is summarized in Section 10.3.

6.2.4.3 Irene

Newmont carried out drilling in the Irene area during 1981-1982 and 1987-1989. Altogether, a total of 42 holes were drilled as summarized in Section 10.3.2 and Section 10.3.6.

HISTORICAL MINERAL RESOURCE ESTIMATES

Several historical mineral resource estimates have been estimated by a variety of companies for the Pinion and Dark Star deposits prior to the implementation of NI 43-101. The reader is advised that the historical mineral resource estimates are not in accordance with NI 43-101 and should therefore not be relied upon. A qualified person has not done sufficient work to classify the historical mineral resources as current mineral resources or mineral reserves. Historical mineral resources at Dark Star and Pinion are superseded by the current mineral resources estimated by MDA and presented in this Technical Report. At POD, North Bullion, and Sweet Hollow the mineral resources by APEX presented in Section 13 of this Technical Report are current mineral resources. The historical mineral resources described below are relevant only for historical completeness and are not being treated as current mineral resources or mineral reserves by Gold Standard.

Pinion Deposit Historical Estimates

The first documented historical mineral resource estimate for the Pinion deposit was completed by Crown in 1991 (Calloway, 1992). The 1991 estimate included information from 194 drill holes in the Main zone and North zone. The estimate used a cross-sectional polygonal method, a gold cutoff grade of 0.34 g Au/t, and a density of 2.464 g/cm³(tonnage factor of 13.0 ft³/ton). A “geologic” mineral resource of 7.36 million tonnes of material averaging 0.89 g Au/t was calculated, containing approximately 210,000 troy ounces of gold (Table 6-3). The authors have not done sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Gold Standard is not treating these historical estimates as current estimates. These historical mineral resource estimates are superseded by the current mineral resource estimate described in Section 14.3 and are relevant only for historical completeness.

Table 6-3 Historical Pinion Deposit Estimated Mineral Resources


Mineral Resource*	Year	Tons (x10⁶)	Tonnes (x10⁶)	Gold Grade		Silver Grade		Cut-off Grade	Contained Ounces
Mineral Resource*	Year	Tons (x10⁶)	Tonnes (x10⁶)	(opt)	(g/t)	(opt)	(g/t)	(opt Au)	Au	Ag
Crown (Calloway, 1992a)	1991	8.11	7.36	0.026	0.891	-	-	0.01	210,860	-
Polygonal (Wood,1995)	1995	30.64	27.8	0.026	0.89	-	-	0.01	796,640	-
MIK (Wells, 1995)	1995	18.26	16.56	0.0269	0.92	-	-	0.01	491,194	-
Bharti (Bharti Eng., 1996)	1996	10.8	9.8	0.025	0.857	0.157	5.383	-	270,000	1,695,600

*The mineral resource estimates summarized in Table 6-3 are not consistent with current CIM standards for mineral resource estimation and classification. The authors have not done sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Gold Standard is not treating these historical estimates as current estimates. These historical mineral resource estimates are superseded by the current mineral resource estimate described in Section 14.3 and are relevant only for historical completeness. Calloway (1992a) in table is Calloway (1992) of this Technical Report.

Historical mineral resource estimates were updated for the Pinion deposit in 1995 by Cyprus (Table 6-3). They comprise a polygonal estimate (Wood, 1995) and a Multiple Indicator Kriging (“MIK”) estimate that used Mintec’s MED System software (Wells, 1995). The polygonal estimate incorporated high-density and low-density drilling at, and surrounding,


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the two zones of mineralization and utilized a density of 2.563 g/cm³(tonnage factor of 13.50 ft³/ton). Polygons were constructed using cross-sectional drill-hole information and were classified as “proven” in areas where drill density was 30.48 m, and where polygons were projected 15.24 m on either side of a section. Polygons with drill-hole spacing between 30 m and 60 m were classified as “probable” and those with drill hole spacing over 61 m, were classified as “inferred.” The mineral resource was calculated by summing all polygons with an average grade above a cutoff of 0.34 g Au/t. The original classification of the 1995 polygonal Pinion mineral resource is not consistent with CIM standards. The summary provided in Table 6-3 is taken from the original report and represents a summation of all three of the historical mineral resource categories into a global historical mineral resource.

The 1995 Cyprus polygonal mineral resource (Table 6-3) was calculated using ~350 drill holes, but the estimate incorporated very few density measurements and a very limited amount of quality control/quality assurance (“QA/QC”) data were available. The Cyprus historical mineral resources included drill-hole data and estimates for mineral resources in Section 27. Cyprus also produced an MIK estimate for the Pinion deposit in 1995 utilizing a similar database to that of the 1995 Cyprus polygonal mineral resource described above. The same 2.563 g/cm³as the polygonal mineral resource was used and grade was applied to a 15.24 m x 15.24 m x 6.1 m block model using Mintec’s MED System software and an MIK grade-estimation algorithm. Following the estimation process, Lerchs-Grossman pit models were run for $400/oz and $700/oz gold price scenarios using various parameters including: a) 45^omaximum pit slopes; b) a $2.51/short ton crushed ore cost (crushing processing, pad construction, and G&A); c) 48% recovery for ROM material; and d) 62% recovery for crushed material. A lower cutoff grade of 0.274 g Au/t was employed for the ROM material and 0.49 g Au/t was utilized for the crushed material for the $700/oz scenario. A lower cutoff of 0.31 g Au/t was utilized for the combined ROM/crush material for the $400/oz scenario. In a mineral resource summary document by Wells (1995), it is clearly stated that the mineral resource work relied on estimations for key factors such as density, recovery, and optimal crush size due to limited test work.

In 1996, RSM contracted Bharti Engineering (“Bharti”) of Toronto, Canada, to conduct mineral resource estimation on the Pinion Main and North zones within Section 22 in T30N, R53E and excluded data within Section 27 (Table 6-3; Bharti Engineering, 1996). The mineral resource estimate utilized GEMCOM mining software and although not clearly stated, it is thought that the Inverse Distance Squared (ID²) grade-estimation algorithm was used to apply grade to a 15.24 m x 15.24 m x 6.1 m block model. Samples (~1.524 m average length) were uncapped and composited (to 6.1 m), with a minimum of two and a maximum of 12 data points required for modeling. The Bharti estimate (Table 6-3) comprised a “global resource,” without cutoff grade, of 9.8 million tonnes at 0.86 g Au/t, representing a total of 273,800 contained ounces of gold. A qualified person has not done sufficient work to classify the historical mineral resources as current mineral resources or mineral reserves and the historical mineral resources are superseded by the current mineral resource estimates presented in Section 14.3 of this Technical Report. The historical mineral resources described above are relevant only for historical completeness and are not being treated as current mineral resources or mineral reserves by Gold Standard. This estimate incorporated more data but is otherwise comparable to the 1991 Crown polygonal estimate discussed above. A Whittle pit was run for the 1996 mineral resource estimate using a gold price of $390/oz, a recovery of 67%, total operating costs of $5.75/ton, a 4% Royalty, 50^°maximum pit slope and dilution estimated at 10%. Using these values, two potential pits were generated for the Main zone and North zone totaling only 2.99 million tonnes and averaging 0.89 g Au/t, which represented approximately 85,750 ounces of gold. Of that, 57,400 ounces was considered recoverable. A qualified person has not done sufficient work to classify the historical mineral resources as current mineral resources ore mineral reserves and historical mineral resources are superseded by the current mineral resource estimates presented in Section 14.3 of this Technical Report. The historical mineral resources described above are relevant only for historical completeness and are not being treated as current mineral resources or mineral reserves by Gold Standard. As with the previously discussed historical mineral resource estimates, this 1996 estimate incorporated limited density, QA/QC and recovery information and its geographic limitation to Section 22 in turn limited the applicability of the mineral resource estimate as it excluded a significant amount of drill data in the northern part of Section 27.


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6.3.2 Dark Star Deposit Historical Estimates

Based upon the 1991 to 1993 drilling results, Crown and Cyprus estimated mineral resources in 1992 and 1994, prior to the 1997 to 1999 drill holes completed by Mirandor and Kinross. The historical mineral resource estimates discussed below should not be relied upon and they are superseded by the current mineral resources estimate presented in Section 14.2 of this Technical Report.

Calloway (1992) described the 1992 Crown estimate for the Dark Star deposit as follows:

“Crown Resources has delineated a geologic resource in the Dark Star discovery area of approximately 7.0 MT @ 0.022 opt Au, or 154,000 oz of contained gold. Mineralization remains open in three directions. Calculations of the Dark Star geological resource utilized nearest neighbor and ordinary kriging methods, with a 0.010 opt cutoff, minimal 15 ft benches, and a 13.5 ft³/st density factor.”

There are no other details provided for the 1992 Crown estimate by Calloway (1992). The estimate is considered historical and should not be relied upon. The authors have not done sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Gold Standard is not treating these historical estimates as current estimates. These historical mineral resource estimates are superseded by the current mineral resource estimate described in Section 14.2 and are relevant only for historical completeness.

In 1994, a consultant on behalf of Crown constructed a polygonal mineral resource estimate for the Dark Star deposit (Table 6-4) using GEO-MODEL and PC-XPLOR modules of GEMCOM (Peek, 1994; McCusker and Drobeck, 2012). The estimated mineral resource was based upon a polygonal methodology using composited drill-hole intervals and cross sections at 30.48 m intervals. Tonnages, grade, and total ounces were calculated using polygons of 15.24 m width on either side of each cross-section. The 1994 estimate did not include any geostatistics on variability or capping, no geologic constraints, no down-hole surveys, no QA/QC data evaluation, and no mention of density measurements. Peek (1994) utilized an assumed density of 2.375 g/cm³for the estimate. No economic constraints other than a lower-grade cutoff were applied to the mineral resource estimate.

Table 6-4 1994 Dark Star Historical Crown Mineral Resource Estimate


Mineral Resource (Reference)	Tons (x 10⁶)	Tonnes (x 10⁶)	Gold Grade (opt)	Gold Grade (g/t)	Cut-off Grade (opt Au)	Cut-off Grade (g/t Au)	Contained Au (oz)
Polygonal (Peek, 1994)	11.5	10.43	0.0168	0.576	0.010	0.343	193,709
Polygonal (Peek, 1994)	7.55	6.85	0.0201	0.689	0.013	0.446	151,481
Note: The authors have not done sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Gold Standard is not treating these historical estimates as current estimates. These historical mineral resource estimates are superseded by the current mineral resource estimate described in Section 14.2, are relevant only for historical completeness and should not be relied upon.

In December 1995 to January 1996, Cyprus personnel estimated a polygonal mineral resource estimate for the Dark Star deposit with data from ~81 drill holes, utilizing a lower-grade cutoff, a pit shell, internal dilution, and a stripping ratio of 1.5:1, in a manner that was consistent with industry standards at that time (DeMatties, 2003). Polygons were constructed on cross sections using drill-hole information and were classified as “proven” in areas where drill density was 30.48 m and polygons were projected 15.24 m on either side of a section. Polygons with drill-hole spacing between 30 m and 60 m were classified as “probable” and those with spacing >61 m were classified as “inferred.” The Dark Star mineral resource was estimated by summing all polygons with an average grade 0.34 g Au/t (cutoff grade). A density of 2.563 g/cm³was used. Very few density measurements and little or no QA/QC data were incorporated (DeMatties, 2003).


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The 1995-1996 Cyprus estimate for Dark Star is summarized in Table 6-5. It represents a global historical “geological” mineral resource as of January 1996 and does not include the drilling by Mirandor and Kinross in Section 24. Although the Dufresne and Nicholls (2016) review established a high quality for the data used in the 1995-1996 estimate, there is insufficient information available to properly assess all of the estimation parameters and the standards by which the estimate for Dark Star was categorized. The authors have not done sufficient work to classify these historical estimates as current mineral resources or mineral reserves, and Gold Standard is not treating these historical estimates as current estimates. The historical mineral resource estimate for Dark Star should not be relied upon, it is relevant only for historical completeness, and it is superseded by the current mineral resource estimate presented in Section 14.2 of this Technical Report.

Table 6-5 Dark Star Deposit 1995-1996 Cyprus Mineral Resource Estimate


Mineral Resource	Tons	Tonne	Gold Grade		Cut-off Grade		Contained
(Reference)	(x 10⁶)	(x 10⁶)	(opt)	(g/t)	(opt Au)	(g/t Au)	Au (oz)
Polygonal DeMatties 2003; Cyprus 1995-1996	7.72	7.00	0.020	0.690	0.01	0.34	151,365

6.3.3 POD (Railroad) Deposit Historical Mineral Resources 1985 - 2003

The first estimate of gold mineral resources at the POD deposit was made by Kuhl (1985) using the data from NICOR’s drilling (Table 6-6). A rectangular-block polygonal estimate was used with the following parameters:

Data projected half way to the adjoining drill hole or 30.5 m;
Inclusion of intercepts less than 1.03 g Au/t if the outlying intervals brought the overall average to equal 1.03 g Au/t;
Minimum 3.05 m intercept in the drill hole;
All calculations made using fire assay intervals;
No stripping ratio calculated; and
No metallurgical recovery information utilized.

Bartels (1999) re-estimated the gold mineral resource at the POD (Railroad) deposit (Table 6-6) with a cross-sectional method based on 58 holes on 27 cross sections spaced 30.5 m apart using the following assumptions:

Tonnages were calculated using a density of 2.563 g/cm³;
Assay values include silver credits at a 60:1 ratio;
Compositing of assay values was done according to the following conventions:
Intervals of low grade (<1.030 g Au/t) up to 4.6 m thick, bound on both sides by >1.030 g Au/t values were included within the “ore” envelope only if the average of the low grade and the upper- and lower-bounding values were >=1.030 g Au/t;
No capping of high-grade assay values was done;
Volumes were determined by projecting the contoured “ore” areas 15.24 m on either side of the section plane;
An average grade was assigned to each area by determining the weighted average grade of all drill intercepts within the “ore” envelope; and
Average grade was assigned to the respective volume and contained ounces were calculated.

Masters (2003a) re-estimated the gold contained within the POD (a.k.a. Railroad) zone (Table 6-6) utilizing a cross-sectional polygonal method with 71 holes on 15 cross sections approximately spaced 30.5 m apart using the following methodology and assumptions:

Tonnages were calculated using a density of 2.563 g/cm³;


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Mineralization was categorized as oxidized (cyanide soluble gold within the Webb siltstone) and refractory gold (primarily within carbonaceous, sulfidic, unoxidized Webb siltstone);
The oxidized and refractory gold categories were sub-divided into grade shells of 0.34 g Au/t to 0.69 g Au/t and >0.69 g Au/t, and an additional category of mineralization for refractory gold at depths above 91.4 m depth was also considered; and
Each mineralization category was estimated separately for tons, ounces and grade.

Masters (2003b) also presented the first estimation for gold contained within the East Jasperoid zone (Table 6-6) located immediately to the east of the POD zone, located immediately to the east of the POD zone. Estimation was completed utilizing a cross-sectional method on four sections spaced 30.5 m apart.

Table 6-6: POD Deposit Historical Mineral Resource Estimates 1985 - 2003


Resource Area	Tons	Tonnes	Contained Ounces Au	Average Au Grade		Cutoff Au Grade		Reference
Resource Area	Tons	Tonnes	Contained Ounces Au	(opt)	(g/t)	(opt)	(g/t)	Reference
POD	1,197,400	1,086,280	107,766	0.090	3.09	0.030	1.03	Kuhl, 1985
POD	1,400,000	1,270,080	112,000	0.080	2.74	0.020	0.69	Kuhl, 1985
POD	1,006,665	913,250	89,731	0.089	3.05	0.030	1.03	Bartels, 1999
POD	2,654,112	2,407,810	134,445	0.0506	1.73	0.010	0.34	Masters, 2003a
East Jasperoid	1,013,808	919,727	31,742	0.031	1.06	0.010	0.34	Masters, 2003b

Note: The historical mineral resource estimates summarized in Table 6-4 were performed prior to the implementation of the standards set forth in NI 43-101 and are relevant only for historical completeness. There is insufficient information available to properly assess the estimation parameters and the standards used. The authors have not done sufficient work to classify these as current mineral resources, Gold Standard is not treating them as current mineral resources and they have been superseded by the current resources presented in Section 13. These historical mineral resources should not be relied upon.

6.4 HISTORICAL MINE PRODUCTION

6.4.1 North Railroad

The North Railroad portion of the property covers the historic Railroad district. Ketner and Smith (1963) suggested that historic production records for the district are not very reliable for the period between 1869 and 1905. Only the total volumes of tons mined, and commodity produced were reported, if they were reported. They estimated the total value of production through 1956 to be worth $2 million using the value of the commodity produced for the year it was produced. Ketner and Smith (1963) reported 43,940 total tons of ore were mined with mineral production distributed as follows:

Gold - 6,918 ounces
Silver - 382,000 ounces
Copper - 2,850,000 pounds
Lead - 4,340,000 pounds
Zinc - 372,000 pounds

6.4.2 South Railroad

There has been no mineral production reported for the South Railroad portion of the property.


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7 GEOLOGICAL SETTING AND MINERALIZATION

This section summarizes the geologic setting and mineralization of the Pinion-Railroad property, which includes the Dark Star, Pinion, Jasperoid Wash, and North Bullion area deposits. This section is based on the descriptions and information provided by Dufresne and Nicholls (2016), Hunsaker (2010; 2012a; 2012b), Koehler et al. (2014), Shaddrick (2012), and sources cited therein. The authors have reviewed this information and believe it accurately represents the geology and mineralization as currently understood.

References to Tomera Formation equivalent stratigraphy have been noted historically. However, recent work suggests these units in the Railroad-Pinion property may not be of equivalent age, so all usage of Tomera Formation equivalent in this Technical Report refer to units that are Pennsylvanian-Permian undifferentiated.

7.1 REGIONAL GEOLOGIC SETTING

The Railroad-Pinion property is located in the southern portion of the Carlin trend, a northwest-southeast alignment of sedimentary-rock hosted gold deposits and mineralization, as shown in Figure 7-1. The property is centered on the Railroad dome, or “window” in the Piñon Range (Mathewson, 2002) as shown in Figure 7-2. Such domes or “windows” consist of upright folds in horsts of Paleozoic rocks of the Roberts Mountains autochthon, exposed by erosion, that were favorable for the formation of Carlin-style gold deposits (Jackson and Koehler, 2014). In the case of the Railroad and other “windows” within the Carlin trend, pulses of Mesozoic and Cenozoic magmatism intruded the folds and related faults (Figure 7-2).

The Carlin trend was within the passive, western continental margin of North America during the early and middle Paleozoic time, which is the time of deposition of the oldest rocks observed in the area (Stewart, 1980). A westward-thickening wedge of sediments was deposited at and west of the continental margin, in which the eastern depositional facies tend to be coarser and carbonate-rich (shelf and slope deposits, carbonate platform deposits) while the western facies are primarily fine-grained siliciclastic sediments (deeper basin deposits). The Carlin trend is proximal to the shelf-slope break, although this break was not static over time.

In the Late Devonian through Middle Mississippian, east-west compression during the Antler Orogeny produced folds and thrust faults, the most significant manifestation of which is the Roberts Mountain Thrust. This regional, low-angle fault placed western facies siliciclastic rocks over eastern facies carbonate rocks across the region. In this Technical Report the western facies are referred to as allochthonous whereas the eastern facies are autochthonous. As a result of this tectonism, the Mississippian and Pennsylvanian overlap assemblage of clastic rocks was deposited across the region (Smith and Ketner, 1975). Late Paleozoic sedimentary rocks in the Piñon Range are interpreted as structurally interleaved allocthonous and autochthonous sequences (Longo et al., 2002; Mathewson, 2001; Rayias, 1999; Smith and Ketner, 1975).

Multiple igneous intrusions occur along the Carlin trend. The oldest igneous rocks are reported to be Late Triassic in age (Teal and Jackson, 2002).


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Figure 7-1: Regional Geology of the Railroad -Pinion Property

(from Dufresne and Nicholls, 2017a; after Crafford, 2007)

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Figure 7-2: Long Section through the Carlin Trend

Other igneous rocks include: a Late Jurassic dioritic intrusion documented at the Goldstrike gold deposit (Bettles, 2002); intermediate to mafic dikes of Jurassic and Cretaceous age; the Cretaceous, quartz monzonite Richmond stock; the Eocene age Welches Canyon stock; and hydrothermally altered and locally gold-bearing felsic to mafic dikes of Eocene age (Ressel, 2000). The Eocene-age Bullion stock (Henry et al., 2015) is situated between the North Bullion and Pinion gold deposits within the Railroad-Pinion property (Figure 7-1 and Figure 7-2).

Late Eocene and Miocene volcanic rocks were erupted over large areas of the region. These predominantly consist of ash-flow tuffs and lava flows, mainly of rhyolitic compositions, as well as volumetrically smaller amounts of andesitic and basaltic lavas. Sequences of lacustrine sedimentary and volcanic-sedimentary rocks, as young as Pliocene in age, interfinger with and overlie the Cenozoic volcanic cover rocks.

Major regional extension commenced in mid-Miocene time. The extension was generally east-west directed, has continued to the present, and is manifested in the Basin and Range topography. The extensional faulting varies from normal-displacement, block faulting to listric-style faulting with progressively greater extension. The significant consequence of extensional faulting has been the dismemberment and tilting of pre-existing rocks, and development of range-scale horsts and grabens.


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7.2 LOCAL AND PROPERTY GEOLOGY

The property is within the central part of the Piñon Range, which is comprised of Ordovician through Permian marine sedimentary rocks (Smith and Ketner, 1975; Figure 7-3) that form a structural dome. At least one large-scale, asymmetric anticline is present, but younger horst and graben structure developed within a framework of overprinted high-angle faults is a prominent feature of the range. Tertiary sedimentary rocks deposited in shallow, freshwater lakes and overlying intermediate to felsic Tertiary volcanic rocks are present on the flanks of the range and within adjacent grabens (Figure 7-3).

Four prominent high-angle fault directions have been identified including west-northwest, north-south, northwest, and northeast-striking faults. The north-south-striking Bullion fault corridor separates the Tertiary volcanic rocks to the east from the Paleozoic sedimentary units in the range. Northwest- and west-northwest-striking faults occur across the project area and include the South and Main faults at Pinion and the Saddle fault at Dark Star. Some of the faults are low-angle. Drilling indicates that multiple episodes of low-angle fault displacements have juxtaposed Devonian carbonate rocks and Mississippian rocks, resulting in interleaved sections of the Devonian Devils Gate Limestone and Webb Formation, as well as Webb age-equivalent rocks of the Tripon Pass Formation (Hunsaker, 2012b).

A complex of Eocene igneous rocks, centered south of Bald Mountain, have intruded the Paleozoic sedimentary units in the core and east flank of the range (Figure 7-3). Twenty-four samples of intrusive and volcanic rocks from the project area have been studied by Dr. Christopher Henry of the Nevada Bureau of Mines and Geology. Petrography, chemical analyses, and ⁴⁰Ar/³⁹Ar and U-Pb zircon age dates have led to an interpretation that at least 10 distinct igneous rock types at the project were emplaced during at least four distinct episodes between 38.9 and 37.5 Ma, associated with the Indian Well volcanic field (Henry et al., 2015).

The Railroad-Pinion area geology is summarized in two parts that correspond to the North Railroad portion of the property, and the South Railroad portion of the property.


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Figure 7-3: Gold Standard Property Geologic Map

(from Dufresne et al., 2017; modified after Smith and Ketner, 1978)


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7.2.1 North Railroad Portion of the Property

7.2.1.1 North Bullion Area Geology

The North Bullion horst is bounded to the east and northwest by younger, generally flat lying, dacitic to rhyolitic tuffs of the Indian Well Formation (Figure 7-3; Henry et al., 2015). The Indian Well Formation contains phenocrysts of quartz, sanidine, hornblende, and biotite within a pink to grey groundmass, and rests on top of an angular unconformity above the underlying, Eocene-age Elko Formation in the eastern hanging wall of the North Bullion fault zone (“NBFZ”). The Elko Formation is exposed within the eastern hanging wall of the NBFZ in the northern part of the property, as shown in Figure 7-3, and consists of thick- to thinly bedded mudstone, sandstone, chert pebble conglomerate, freshwater limestone, and tuffaceous sediments (Stewart, 1980; Smith and Ketner, 1976).

The North Bullion horst consists of thick bedded, fining upward, conglomerate, and mudstone of the Mississippian Chainman Formation which contains 1 m to 7 m thick dacite sills from 100 m to 200 m below the surface. Dacite dikes occur along steeply dipping faults within the NBFZ (Jackson et al., 2015). In between the upper and lower Chainman Formation is a sequence of mixed carbonate and siliciclastic rocks, which are interpreted to belong to the Mississippian Tripon Pass Formation (Longo et al., 2002; Matthewson, 2001; Oversby, 1973). Two limestones within the Tripon Pass Formation act as informal marker units. Limestone 1 is a dark-grey, laminated to thinly bedded micrite located at the top of the Tripon Pass Formation, and limestone 2 is a grey, medium- to thick-bedded calcisiltite to calcarenite located approximately 55 m below limestone1 (Figure 7-4). The Tripon Pass Formation hosts the upper gold zone of the North Bullion deposit and locally contains >6 g Au/t. The Tripon Pass Formation is underlain by the variably bedded sandstone, conglomerate, and silty mudstones of the Mississippian Chainman Formation.

Underlying the Chainman Formation, in low-angle fault contact, is the Devonian Devils Gate Limestone (Devils Gate Limestone of Figure 7-3). It is composed of grey, thick-bedded calcarenite and minor micrite, between 60 to 150 m in thickness. Dissolution-collapse breccia developed at the top of the Devils Gate Limestone is host to high-grade gold within the lower zone at North Bullion (Jackson et al., 2015). In the northern portion of the deposit, silty mudstone of the Mississippian Webb Formation and silty micrite of the Mississippian Tripon Pass Formation (Figure 7-4), are important hosts to gold, and are preserved along the low-angle fault contact between the Chainman Formation and the Devils Gate Limestone. Beneath the Devils Gate Limestone there is a transitional contact into the Sentinel Mountain Dolomite, which has an average thickness of 150 m, and is in transitional contact with calcareous sandstone of the underlying Oxyoke Formation (Oxyoke Sandstone in Figure 7-4). The cross-bedded Oxyoke is approximately 120 m in thickness and consists of well-rounded quartz grains, which are either matrix- or grain-supported. There is tectonic and dissolution-collapse breccia that extends from the lower contact of Tripon Pass limestone to the top of the Devils Gate Limestone between the Massif and West Strand faults. The deepest drill holes at North Bullion bottomed in thin-to thick-bedded dolomite of the Devonian Beacon Peak Dolomite (Figure 7-4).


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Figure 7-4 North Bullion Stratigraphic Column

(from Jackson et al., 2015)


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Jackson et al. (2015) described the structural effect on geology at North Bullion as follows:

“North Bullion (gold deposit) occurs in a triangular shaped horst in the footwall of the major north-striking, steeply east-dipping, North Bullion Fault Zone (NBFZ). The western edge of the horst is bounded by a northeast-striking, northwest-dipping fault. The NBFZ is 300 m wide and apparent normal displacement across the NBFZ is greater than 600 m, as constrained by the deepest holes into the Indian Well Formation volcanic rocks that fill the Bullion graben to the east. Chainman sandstone occupies the center of the horst, and the variable strikes and dips at the surface indicate an open fold is centered on the horst. The western edge of the horst is bounded by a N50E striking, northwest-dipping fault. The triangular shape of the horst is well represented in structure contours on the top of the Devils Gate Limestone.”...

...“Intrusive relationships and tilting of units indicate the deposit formed during an Eocene event with synchronous intrusion, hydrothermal activity and extensional movement on graben-bounding faults. Dacite sills, dated at 38.8–38.2 Ma, intruded steeply dipping faults within the NBFZ and low angle, bedding parallel faults, capping the gold system. The margins of dacite dikes and sills are commonly sheared and some dacite occurs as clasts within mineralized dissolution-collapse breccia, indicating continued movement along faults and hydrothermal activity after emplacement of the dacite. In fault steps within the NBFZ, the Eocene Elko Formation has the same moderate eastward dip as the underlying Paleozoic rocks. The collapse breccia generally exhibits a flat-tabular textural fabric subparallel to today’s surface. All of this evidence supports the Formation of North Bullion during a very dynamic, focused Eocene event with synchronous extension, intrusion and Carlin-style mineralization.”...

7.2.2 South Railroad Portion of the Property

7.2.2.1 Pinion Deposit Area Geology

The geological setting, stratigraphic units and the overall tectonic history of the Pinion area is the same as that described for the adjacent North Railroad area by Hunsaker (2010, 2012a, 2012bb), Shaddrick (2012), Koehler et al. (2014), Turner et al. (2015), Dufresne and Koehler (2016), and Dufresne and Nicholls (2018). The geology is illustrated in Figure 7-3. A stratigraphic column for the project area is presented in Figure 7-5.

The Pinion deposit area encompasses a sequence of Paleozoic sedimentary rocks exposed within large horst blocks in which the sedimentary rocks have been broadly folded into a south- to southeastward-plunging, asymmetric anticline. The axis of the Pinion anticline can be traced for approximately 3.2 km, trends N20°W, and plunges approximately 25° to 30° to the south-southeast (DeMatties, 2003). The apparent dip of the western fold limb ranges from 10° to 35° and the steeper eastern limb dips 25º to 50º. Eastern assemblage formations including the Oxyoke, Beacon Peak, Sentinel Mountain, and Devils Gate formations form the core of the anticline. Siliceous clastic units of the Tripon Pass, Webb, Chainman, and Tonka formations form its limbs (Calloway, 1992a).

The contact between the Devils Gate and Tripon Pass (Figure 7-5) is characterized by a multi-lithic dissolution-collapse breccia (“mlbx”) that ranges from 3 m to 120 m in thickness. The mlbx is characterized by multi-lithic clasts, barite, clay matrix with a silica overprint, and infrequent banded quartz veins. The breccia is thickest on the east limb of the fold and thins along the crest and along the west limb.


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Figure 7-5: Stratigraphic Column for the Pinion, Dark Star, and Jasperoid Wash Deposit Areas

(from Gold Standard 2019; Undifferentiated Pennsylvanian-Permian units are those at Dark Star and Jasperoid Wash)

The Pinion deposit is contained within a northwest-trending horst. Faults on the northeast horst margin are linking structures to the more northerly striking, range-bounding Bullion fault corridor (Norby et al., 2015) and include the locally named Bullion, Linkage, N10E, and Tonka faults. Older N50°W- to N60°W-striking faults (South and Main faults) transect the Pinion deposit and offset the anticline.

At depth, the Devils Gate, Tripon Pass, and Webb formations overlie Mississippian-aged Chainman Formation. This contact was defined by Norby et al. (2015) as gently west-dipping Pinion thrust fault between the overlying Devils Gate to Chainman sequence and the underlying Chainman sequence. On the east limb of the fold, additional localized thrust faults occur above the Pinion thrust fault, resulting in locally repeated sections of Chainman, Webb, and Tripon Pass.

Alteration associated with gold-silver mineralization is primarily silicification of the breccia. There are also zones of abundant disseminated and vein barite, with up to 75% barium determined from x-ray fluorescence analysis. Decalcification of the Tripon Pass and Devils Gate formations along the margins of the breccia have also been observed. Minor clay alteration can be seen along the Main, South, and Bullion faults. Elements associated with gold are silver, antimony, arsenic, barium, and mercury. A type of mineralization with more typically epithermal-like textures is also present at Pinion. Banded fine-grained to fine-cockscomb silica occurs throughout the deposit, locally with stibnite (or oxidized to stibiconite) and elevated silver to 70 ppm.


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Gold and silver mineralization at the Pinion deposit is strongly controlled by the dissolution-collapse breccia at the contact between calcarenite of the Devils Gate Limestone and the overlying silty micrite of the Tripon Pass Formation (Norby et al., 2015). Approximately 90% of the mineralization is hosted within the breccia and is defined locally as the Main zone. The Pinion deposit extends northwards, along the Bullion fault corridor, and is referred to as the North zone. The North zone appears to be a fault offset of the east limb of the Pinion anticline. Low-grade mineralization extends into the footwall of the Bullion fault and is hosted in Sentinel Mountain Dolomite.

7.2.2.2 Dark Star Geology

The Dark Star deposit is located east of the Pinion deposit (Figure 7-3) and occurs in a 400 m- to 600 m-wide structural block of Pennsylvanian-Permian rocks (Harp et al., 2016). A generalized stratigraphic column for the Dark Star area is illustrated in Figure 7-5.

Dark Star lies along the north-south Dark Star fault corridor that has Mississippian Chainman Formation and unconformably overlying Tertiary Conglomerate to the west, and Eocene Indian Wells Formation to the east. These formations are fault bounded by the West fault and Dark Star fault, respectively. Pennsylvanian-Permian undifferentiated conglomerate and calcareous bioclastic units are interpreted to be a Tomera Formation equivalent, a localized unit that occurs at Dark Star and possibly Jasperoid Wash, comprise the horst between these faults. The Pennsylvanian-Permian undifferentiated section is informally broken down into the uppermost unit of siltstone (generally calcareous), a middle unit of calcareous conglomerate (with minor interbedded sandstone), and a lower unit of calcareous siltstone (Figure 7-5). These units are gently folded in a north-south-trending syncline-anticline pair between the West and Dark Star faults.

The Dark Star fault corridor is a prominent north-south-trending fault system consisting of the West, Ridgeline, IDK, East, and Dark Star faults. The corridor has a surface expression of greater than 12 km in length. All but the West fault are steeply east-dipping normal faults with 15 m to 200 m of offset. The West fault is a moderately west-dipping fault with displacement of the Chainman Formation over the Pennsylvanian-Permian rocks and may be a continuation or age equivalent to the Pinion thrust fault.

An older set of N°40W- to N60°W-striking faults, the Saddle and Outcrop faults, transect the Dark Star deposit, and appear to offset the mineralization. These appear to be contemporaneous with the N60°W-striking faults at Pinion. Surface mapping has indicated the presence of regional N55°E- to N60°E-striking faults north and south of the Dark Star deposit.

Alteration at Dark Star is dominated by decalcification and silicification of the Pennsylvanian-Permian rocks. Small areas of clay alteration associated with faults have been observed, along with localized barite veins and widespread disseminated barite (Harp et al., 2016). Quartz veinlets, drusy-quartz lining fractures, and banded-quartz occur in the silicified rocks. Several stages of tectonic, collapse, and hydrothermal breccia are recognized throughout the mineralized zone. Alteration of the upper and lower siltstone units is characterized by decalcification, overprinted by argillic and weak silicic alteration.

7.2.2.3 Jasperoid Wash Geology

The Jasperoid Wash deposit is located south of the Pinion deposit in a structural block of Pennsylvanian-Permian rocks (Figure 7-3). These rocks are similar to those at the Dark Star deposit as illustrated in Figure 7-5.

The Jasperoid Wash deposit occurs along a linear, north-south-striking structural corridor which is bounded on its east and west sides by major faults. The west-bounding fault juxtaposes Mississippian Tonka Formation against the Pennsylvanian-Permian rocks to the east. A horst block of Pennsylvanian-Permian conglomerate and clastic units is between the two main faults. The Pennsylvanian-Permian rocks are informally assigned to an upper unit of silty limestone, a middle unit of calcareous sandstone and conglomerate, a lower unit of calcareous siltstone, and an


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underlying conglomerate composed of chert pebbles and a sandstone matrix. These rocks are similar to and may correlate with Tomera Formation equivalent units at Dark Star, but the stratigraphic position relative to known formations is not known at Jasperoid Wash.

At Jasperoid Wash, the middle sandstone and conglomerate unit crops out at the surface in small crags that are resistant to weathering. This unit is mostly composed of thick beds of debris-flow conglomerate containing clasts of chert and cherty bio-micrite in a silicified, sandy calcarenite to silty-micrite matrix. The lower calcareous siltstone unit is composed of varying thicknesses of interbedded calcisiltite, calcarenite, bioclastic limestone, calcareous sandstone, and minor beds of conglomerate. Outcrops of this unit tend to be less resistant to weathering and are smooth and low-lying.

Dikes of “quartz-eye” rhyolite and feldspar porphyry with a composition close to dacite are present within the Jasperoid Wash deposit and are inferred to be of Tertiary age. These intrusions occur within the structural corridor and at fault intersections. A third type of dike, composed of intensely silicified quartz-feldspar porphyry, crops out north of the deposit. At a fault intersection within the deposit, some outcrops consist of a multi-phased, hydrothermally altered breccia consisting of younger quartz-feldspar porphyry matrix and clasts of dacite and rhyolite. Also, strongly clay-altered and mineralized dacite porphyries with very fine-grained pyrite has been encountered in drilling.

Structurally, the Jasperoid Wash deposit is bounded to the west by the north-south-striking, 65°W-dipping Westport fault. This is interpreted to be a reactivated thrust fault and is similar to the West fault at Dark Star. The Eastport fault of the Jasperoid Wash structural corridor also strikes north-south, and dips 78°W. The Eastport fault truncates a syncline-anticline pair that also trends north within the structural corridor. There are also east-west-trending faults within the north-south fault corridor that bound a horst block and define the southern extent of the deposit.

Alteration of the middle conglomerate and lower siltstone units includes moderate to strong silicification, decalcification, and argillization. Quartz veinlets and drusy quartz on fractures occur with silicification. Small pods of unoxidized sulfide minerals are preserved within the sedimentary rocks where oxidizing fluids did not permeate the rock. Vugs formed by decalcification of limestone and dolostone are present. Hydrothermal alteration is mostly seen in the feldspar porphyry, calcisiltite, calcarenite, calcareous sandstone, and bioclastic limestone units, and is marked by strong clay development and/or disseminated sulfide grains with a sooty appearance that are mostly oxidized to limonite and hematite. Hydrothermal alteration of the feldspar porphyry dike is distinct and defined by disseminated sulfide grains with a sooty appearance. The lower siltstone unit is commonly decalcified and becomes more calcareous with depth.

7.3 MINERALIZATION

The Railroad-Pinion property includes demonstrated Carlin-type gold mineralization in at least four deposit areas: North Bullion, Pinion, Dark Star, and Jasperoid Wash. These deposits are similar in setting and style to that of other deposits in the region, including Rain and Emigrant (Koehler et al., 2014; Norby et al., 2015; Turner et al., 2015; Dufresne and Koehler, 2016). Mineralization occurs mainly as finely disseminated, submicroscopic gold in largely stratiform bodies in Devonian, Mississippian, and Pennsylvanian-Permian rocks. The following subsections describe the mineralization in the North Bullion, Pinion, Dark Star, and Jasperoid Wash deposits and are modified from Dufresne and Nicholls (2016; 2017a; 2017b; and 2018).

7.3.1 North Bullion Deposits

The North Bullion deposits, which includes North Bullion, POD, and Sweet Hollow zones, contains Carlin-type disseminated-gold mineralization that is largely not exposed at the surface. The bulk of the geological understanding and interpretation of the North Bullion deposits has come from core drilling that was guided by interpretations of gravity and CSAMT data. Gold mineralization is focused in the footwall of the NBFZ, a north-south-striking zone of normal faults with an overall down-to-the-east displacement. North-south-, northwest-, west-northwest-, and northeast-striking faults appear to be important controls on mineralization. In general, gold-silver mineralization is localized in gently to


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moderately dipping, strongly sheared Webb and Tripon Pass formation rocks, and dissolution-collapse breccia developed above and within silty micrite of the Tripon Pass Formation and calcarenite of the Devils Gate Limestone (Figure 7-6) (Jackson and Koehler, 2014; Jackson et al., 2015).

The upper limit of gold mineralization at the North Bullion deposit varies from 105 m to 400 m in depth. The dip of the mineralized material steepens from 10° to 45° to the east as the eastern strand of the NBFZ is approached. Gold is associated with sooty-looking, very fine-grained sulfide minerals, silica, carbon, clay, barite, realgar, and orpiment in addition to elevated arsenic, mercury, antimony, and thallium. Gold grades >6 g Au/t have been intercepted.

The North Bullion deposit, as currently defined, is approximately 750 m in length, 600 m in width and as much as 500 m in vertical extent.

Figure 7-6: North Bullion Cross Section N4488800

(from Jackson et al., 2015)

Mineralization at the nearby POD zone is restricted to a steeply dipping shear zone which trends west-northwest and is situated in rocks stratigraphically higher than the lower mineralization at North Bullion (Hunsaker, 2012b; Masters, 2003). Mineralization at POD is hosted by the upper siltstones of the Webb Formation. The core of the mineralized body contains carbon and fine-grained, disseminated pyrite, and accounts for approximately 15% of the mineralization.


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This is surrounded by strongly oxidized mineralization. Gold grains are in the range of 5 to 20 microns, and are associated with oxidized pyrite, stibnite, and arsenopyrite (Masters, 2003). Additionally, gold mineralization at POD is associated with silicified rock, including jasperoid, argillized rock, pyrite, barite, and some minor dolomite replacement of calcite (Hunsaker, 2012b).

As currently defined, the POD zone is approximately 650 m in length, 150 m in width, and as much as 200 m in vertical extent.

The Sweet Hollow zone is situated about 200 m southeast of the POD zone and about 600 m south of the North Bullion deposit. As currently defined, the Sweet Hollow zone is approximately 1,050 m in length, 250 m in width, and as much as 100 m in vertical extent.

7.3.2 Pinion Deposit

The Pinion gold deposit is located along the west-northwest-trending Pinion anticline and proximal to the Bullion fault. The Main zone trends approximately N50°W to N60°W, is approximately 1,000 m long by 1,000 m wide, and varies in thickness between ~15 to 150 m vertically. Mineralization at the Main zone has been intersected to a depth of 200 m below surface. Mineralization is hosted primarily along the crest of the Pinion anticline, but also along the east and west limbs. The multi-lithic dissolution-collapse breccia at the Devils Gate-Tripon Pass contact hosts the majority of mineralization, with minor amounts associated with decalcified limestone and dolostone above and below the breccia.

The North zone is approximately 1,100 m long, along a roughly north-northwest trend, varies from 45 m to 100 m in width, and ranges from 35 m to 135 m in vertical thickness. Lateral continuity of mineralization is shown in a representative Gold Standard cross section (Figure 7-7). Mineralization at the North zone is hosted primarily in multi-lithic breccia and appears to be an offset of the east limb of the anticline. Low-grade mineralization has also been noted in the Sentinel Mountain Dolomite.


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Figure 7-7: Pinion Deposit Geology, Section 4479380N

(from Gold Standard, 2018)

Mineralization at Pinion occurs mainly as submicroscopic disseminated gold in the largely stratiform, multi-lithic, dissolution-collapse breccia developed along the contact between silty micrite of the Tripon Pass Formation and calcarenite of the underlying Devils Gate Limestone (Figure 7-5). Important structural controls are west-northwest and north- to northeast-striking folds and faults. Gold deposition is thought to have been contemporaneous with breccia development and with quartz vein formation and silica ± barite replacement and infill of open spaces. Some free gold in 2 to 20 micron-size grains has been noted in 2018 mineral liberation studies (AMTEL, 2018). Barite was deposited as both massive and disseminated forms and is found most often in the multi-lithic, dissolution-collapse breccia. Barite appears to be paragenetically late, overprinting both the breccia and silica events.

Dark Star Deposit

The Dark Star deposit is hosted primarily within Pennsylvanian-Permian undifferentiated units possibly equivalent to the Tomera Formation, with minor amounts of gold mineralization found in the Chainman Formation. The deposit is centered along the north-south-striking Dark Star fault corridor and is elongate in the N5°E direction. As presently defined by drilling, the deposit consists of the Dark Star Main and Dark Star North zones, and has dimensions of approximately 1,400 m in length, up to 700 m in width, and to a depth of 450 m below surface. A representative geologic cross section is shown in Figure 7-8.


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Figure 7-8: Dark Star Geologic Cross Section N4479600

Gold mineralization at Dark Star is submicroscopic and disseminated within a north- to north-northeast-striking zone of silicification within the middle coarse conglomeratic and bioclastic limestone-bearing unit. This unit is between the upper and lower silty limestone and calcisiltite units (see stratigraphic column in Figure 7-5, Section 7.2.2.2, and geology cross sections in Figure 7-8). At Dark Star Main the mineralization dips steeply to the west near the surface to sub-horizontal at depth; at Dark Star North the mineralization dips steeply to the west.

Oxidation is pervasive at Dark Star Main to a depth of 450 m in the middle conglomeratic unit. At Dark Star North, oxidation is pervasive to a depth of 330 m in the middle conglomeratic and lower silty limestone and calcisiltite units. Oxidation products are primarily limonite with lesser hematite. However, thin zones of unoxidized sulfide minerals are present; pyrite is the principal sulfide mineral.

7.3.4 Jasperoid Wash Deposit

The Jasperoid Wash deposit has approximate extents of 1,400 m in a north direction and a width of about 1,100 m. Drilling shows the deposit dips west gently to steeply at least 400 m. Gold is disseminated within altered feldspar porphyry dikes and adjacent conglomeratic rocks, possibly the same units that host mineralization at Dark Star. The gold is inferred to be submicroscopic, though no petrographic studies have been done. Higher-gold grades are associated with drusy quartz in fractures, which have a varnish of limonite and/or hematite and with zones of very fine-grained disseminated sulfide minerals that have a sooty appearance in the argillized feldspar porphyry. A representative Gold Standard cross section is shown in Figure 7-9.


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Figure 7-9: Jasperoid Wash Geologic Cross Section 4473200N


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8 DEPOSIT TYPES

Gold deposits known and being explored for in the Railroad-Pinion property area are sedimentary-rock hosted, disseminated, Carlin-type gold deposits. These types of gold deposit were first recognized in the 1960s in northern Nevada, and were named for the town of Carlin, Nevada. Since then, over 100 geologically similar deposits, containing approximately 200 million ounces of gold, have been discovered in northern Nevada (Hofstra and Cline, 2000), making it one of the most significant gold regions in the world.

Carlin-type deposits are epithermal deposits with characteristics sufficiently different from typical epithermal deposits that they are considered a distinct deposit type. When first discovered, these deposits were often informally referred to as “no-see-um” gold deposits or “micron” gold deposits because the gold is rarely visible to the naked eye and cannot be recovered by panning.

These deposits are distinctive from typical epithermal deposits because they form replacement bodies with structural and stratigraphic controls, contain primary gold that is restricted to ionic substitution and sub-micron-sized grains in arsenian pyrite, and have alteration that is subtle but dominated by carbonate dissolution of calcareous host rocks (Cline, 2004). Gold did not precipitate in response to boiling or fluid cooling, but instead precipitated in response to sulfidation of iron in the host rock or in a second iron-bearing fluid (Muntean et al., 2011). Host rocks for Carlin-type deposits in Nevada are primarily Paleozoic carbonate rocks. Other host rocks include calcsilicate hornfels, chert, argillite, and igneous dikes.

Most systems exhibit a main stage of alteration and mineralization characterized by acid dissolution and replacement of the calcareous host rock. If the host rock is composed of relatively pure carbonate without quartz silt or sand-grain support, dissolution of the carbonate can result in the formation of open space, leading to collapse and breccia formation. Main-stage decarbonatization of carbonate host rocks is typically accompanied by clay alteration (argillization) of silicate minerals, sulfidation of available reactive iron, and silicification of limestone. Alteration is characterized by an assemblage of quartz, illite, and dolomite with the edges of the system marked by an increase in calcite (Kuehn and Rose, 1992). In gold-enriched zones, dissolution of carbonates, and argillization of silicate minerals is accompanied by sulfidation of iron released by mineral alteration, resulting in precipitation of disseminated auriferous-and arsenian-pyrite, marcasite, or arsenopyrite. These iron sulfide minerals commonly occur as rims on preexisting pyrite. The most important consequence of the pyrite-forming sulfidation reaction is the coupled precipitation of gold with this pyrite (Hofstra and Cline, 2000). It is well-documented that most of the gold in Carlin-type deposits initially resides in arsenian pyrite, arsenian marcasite, and arsenopyrite (Hofstra and Cline, 2000), occurring as sub-micron inclusions of native gold or as structurally bound gold. Pervasive silica replacement (silicification) of the various host rocks is also common.

A distinctive suite of late-stage minerals is commonly present in open cavities and fractures. Textural relationships demonstrate that these minerals precipitated after the main-stage alteration and mineralization. In proximal zones, open cavities and fractures may be filled with orpiment and/or realgar, in places accompanied by quartz, barite, fluorite, pyrite, marcasite, cinnabar, barite, or thallium and antimony sulfides. More distal veins are dominantly calcite ± orpiment and realgar. The geochemistry of Carlin-type deposits is characterized by a distinctive suite of gold, arsenic, antimony, thallium, and mercury ± tungsten (Hofstra and Cline, 2000). These elements are frequently used as pathfinder elements for surface geochemical surveys and as vectors toward mineralization in drill-hole geochemical studies.

Carlin-type deposits vary greatly in size and contained gold. Areal footprints of district deposit clusters range from about 20 to 120 km². Mineralization within a deposit can extend laterally more than 1,500 m and over vertical intervals greater than 1000 m. The larger deposits in Nevada occur within linear districts, or “trends” extending up to more than 20 km and are often controlled by regional structures. Some of these structures probably resulted from reactivation of much older basement normal faults that originated during Proterozoic rifting of western North America (Lund, 2008). These


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old faults are inferred to have served as conduits for deep-crustal hydrothermal fluids responsible for formation of Carlin deposits.

The varied forms of individual deposits reflect local zones of high porosity and permeability that result from favorable lithologic and structural features. Permeable features frequently associated with orebodies include high-angle faults, thrust faults, low-angle normal faults, hinge zones of anticlines, lithologic contacts, reactive carbonate units, debris-flow facies carbonate rocks, lithologic facies changes, breccia zones of all types, and contacts of sedimentary rock with metamorphic aureoles (Cline et al., 2005).

Carlin-type deposits share many features in common, that include (Muntean et al., 2011):

Middle to late Eocene ages (42 and 36 Ma.) (Cline, 2004), a time that corresponds to a change from tectonic compression to extension and renewed felsic to intermediate magmatism;
Deposits occur in linear clusters along old reactivated structures that are probably linked at depth to crustal- scale Proterozoic basement rift structures;
Deposits are preferentially hosted in carbonate rocks within or adjacent to structures in the lower plate of a regional thrust fault;
Deposits exhibit very similar paragenesis, characterized by decarbonatization, argillization, silicification, and sulfidation that results in the formation of gold-bearing arsenian pyrite, which initially hosts the vast majority of the gold in the deposits. This replacement mineralization was followed by open-space deposition of minor amounts of drusy quartz with pyrite, followed by orpiment, realgar, stibnite, and other sulfides. Oxidation often removes the initial sulfide formed in the deposit;
Deposits have low concentrations of silver and base metals, and have an elemental signature of predominantly Au-Tl-As-Hg-Sb;
Deposits were formed by non-boiling ore-forming fluids that ranged from 180°C to 240°C during mineralization, were of low to moderate salinity (mostly ≤6 wt% NaCl equivalent), and CO2-bearing (<4 mol%); kaolinite and illite indicate that fluids were acidic;
There is a lack of mineral or elemental zoning at the district scale that suggest minor temperature gradients. There are no coeval associated porphyry copper, skarn, or distal Au-Pb-Zn-Mn zones; and
Evidence suggests deposit formation by largely fracture-controlled fluid flow from multiple upwelling zones with little evidence for significant lateral fluid flow or spaced convection cells.

A schematic regional deposit model cross-section is shown in Figure 8-1 from Muntean (2018).


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Figure 8-1: Regional-Scale Carlin-Type Deposit Model

(from Muntean and Cline, 2018)

These features strongly suggest Carlin-type deposits, which formed over a broad region of northern Nevada during a relatively narrow time interval, shared common underlying processes for the formation and transport of gold-bearing hydrothermal fluids and the deposition of gold.

Carlin systems can be large deposits with high concentrations of gold. Deposits frequently occur in clusters and can occur at depth with subtle or no surface evidence. It is notable that although the original Carlin deposit in Nevada was discovered in 1960, exploration continues, and discoveries continue to be made.

The Dark Star, Pinion, Jasperoid Wash, and North Bullion gold deposits present characteristics similar to other Carlin-type gold deposits of the Carlin trend. Specific geologic features in these deposits include:

Deposits occur in relatively close proximity to a multi-phase Eocene igneous center with associated igneous stocks, dikes and sills; gold mineralization is of Eocene age;
Deposits occur in a linear zone;
Deposits are hosted in or adjacent to carbonate rock types;
Deposits exhibit strong structural control, localized in areas with greater fault density and occur in either hanging wall or footwall settings of high-angle faults;
Alteration is characterized by decarbonatization, dolomitization, argillization, silicification, barite, and sulfidation;
Gold generally occurs initially as a chemical impurity or as micron-scale particles of arsenian pyrite. Later oxidation has generally removed most sulfides at Dark Star, Pinion, and Jasperiod Wash.


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9 EXPLORATION

The Railroad–Pinion property is being explored on an ongoing basis by Gold Standard using geological mapping, geochemical and geophysical surveying, and drilling. This section of the report is largely drawn from Dufresne and Nicholls (2016), Dufresne et al. (2017), Dufresne and Nicholls (2017a), and Dufresne and Nicholls (2018). The authors have reviewed this information and believe it accurately represents the exploration work done by Gold Standard.

Prior to 2015, exploration activities by Gold Standard were focused in the North Railroad portion of the property. Work completed in 2015 was largely focused on the Pinion area in the South Railroad portion of the property, after its acquisition in 2014. A thorough discussion of these work programs and their results and interpretations is available in previous Technical Reports by Hunsaker (2010, 2012a, 2012); Shaddrick (2012); Koehler et al. (2014); Turner et al. (2015); Dufresne and Koehler (2016); and Dufresne et al. (2017).

Exploration work by Gold Standard since 2010 has resulted in the identification of 17 prospect areas or zones of mineralization within the overall property position, including the Bald Mountain area and North Bullion deposits in the North Railroad-Pinion portion of the property, the Pinion, Dark Star, and Jasperoid Wash deposits, and other areas of the South Railroad portion of the property. Drilling conducted by Gold Standard is summarized in Section 10.

9.1 2009 – 2019 GEOPHYSICS

There is a significant and growing body of geophysical information for the Railroad-Pinion property that includes gravity, controlled-source audio magneto-telluric (“CSAMT”), and ground magnetic surveys. These surveys have been employed to aid in identifying geological structures, key lithologies, and zones of hydrothermal alteration related to mineralization. Additionally, the geophysical surveys have aided in drill-hole targeting and have assisted in the definition of multiple exploration targets.

A ground magnetic survey was completed over the Bullion stock area in 2014 (Figure 9-1). A total of 197 line-km was surveyed with total magnetic intensity recorded in continuous mode at 2-second intervals on lines 100 m apart. The lines were oriented east-west.

Gold Standard completed six gravity surveys from 2009 to 2015, collecting measurements from 3,991 stations covering a large portion of the property as shown in Figure 9-1. The gravity surveys were designed to delineate structures, particularly those in areas lacking bedrock exposures, and/or those areas under cover, and to identify rock types and alteration related to sedimentary-rock hosted and skarn-type mineralization (Wright, 2013). During 2017, gravity measurements at an additional 1,027 stations were taken, covering 23 km²in the South Railroad portion of the property. The 2017 gravity survey was conducted by Magee Geophysical Services LLC and was interpreted by Wright Geophysics.

Seven CSAMT surveys were completed by Gold Standard from 2012 to 2016, covering the Bullion fault corridor, the North Bullion, Pinion, and Dark Star deposits, and the Dark Star fault corridor (Figure 9-1). A total of 85 line-km of CSAMT data were collected during the seven CSAMT surveys. The 2016 CSAMT survey involved 21.2 line-km focused on the Dark Star fault corridor, with nine east-west lines at variable spacing from 200 m to 500 m, that were oriented perpendicular to the main fault trend in the area.


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Figure 9-1: Ground-based Geophysical Surveys by Gold Standard 2009 to 2015

(from Dufresne et al. 2017 with 2016 property outline)


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During 2017, another 68 line-km of CSAMT were surveyed with 21 lines across the Dark Star fault corridor, Ski Track and Bullion to East Pine Mountain areas. The data were acquired by Zonge International Inc. and interpreted by Wright Geophysics.

James Wright of Wright Geophysics designed, supervised, and interpreted the 2016 CSAMT survey. An interpretation of the results by Wright (2016a) is summarized as follows:

A major north-south-oriented structural zone—the Dark Star fault corridor—exists along the east side of all 2016 sections, juxtaposing Tertiary rocks against older sedimentary rocks. The zone has two major normal faults bounding a predominantly Pennsylvania–Permian horst block. Both bounding faults have multiple parallel faults and lesser splays;
A north-south-oriented horst of Pennsylvania–Permian clastic rocks beneath approximately 80 m of Tertiary and Quaternary cover is bounded by two major faults and runs parallel to and 450 m west of the Dark Star fault corridor;
The above horst is terminated to the north by a north-northeast-trending fault and is divided to the south by a major cross-cutting west-northwest-trending fault. South of that the two horsts appear to merge to the south of this cross-cutting structure; and
The Dark Star Main and Dark Star North deposits correlate with high resistivity from a depth of 0 to 25 m to a depth of 200 to 400 m, respectively. The near-surface high resistivity features may be related to alteration.

In 2016, Gold Standard purchased a portion of an airborne magnetic survey from EDCON-PRJ that covered the entire Piñon Range including the North Railroad and South Railroad portions of the property and their surroundings. The Bullion stock forms a strong and large magnetic high, and several of the major structures were extended by the airborne interpretation of Wright (2016b).

Seismic surveys were performed in 2017 and 2018 at Pinion, Dark Star, and North Bullion. In total, three east-west-oriented lines for 37.2 line-km were surveyed. In 2019 three additional seismic lines, totaling 21 line-km, were surveyed directly over and to the north of the North Bullion deposit. The seismic data were acquired by Bird Seismic Services and processed and interpreted by Columbia Geophysical, Sterling Seismic Services Ltd., and Wright Geophysics.

9.2 2010 – 2018 GEOCHEMISTRY

Historical data and subsequent work by Gold Standard has shown there is a positive correlation between anomalous gold and arsenic concentrations in soil samples, and near-surface gold mineralization confirmed with drilling. Gold Standard collected approximately 7,450 soil samples from 2010 to 2015. These were collected over grids in six areas (Figure 9-2) with lines 50 m to 100 m apart and samples taken at spacings of 50 m. During 2017 and 2018, a total of 7,823 soil samples were collected from the South Railroad portion of the property in the Ski Track, Dixie, and Jasperoid Wash areas, and near the southern limit of the property. Samples were taken at intervals of 50 m along lines spaced 100 m apart.

To expand the rock geochemistry database in areas that lacked historical sampling, Gold Standard collected approximately 3,500 rock samples throughout the Dark Star, Pinion, and North Bullion deposit areas from 2010 to 2015 (Figure 9-2). Samples were collected from outcrops, road cuts, and field traverses parallel with topography. The majority of these rock samples comprise simple “grab” samples, but chip, channel and scoop sampling techniques were employed to a lesser degree.

Gold Standard did not collect any rock, soil, or scoop samples in 2016. During 2017 and 2018, a total of 1,550 rock samples were collected from the Ski Track, Dixie, and Jasperoid Wash areas of the property. The geochemical


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exploration work described above identified eight drill targets, some of which have returned significant intercepts of gold, silver, copper, lead, and zinc.

The authors have not analyzed the sampling methods, quality, and representativity of surface sampling at the Railroad-Pinion property because drilling results form the basis for the mineral resource estimates described in Section 14. Drilling is described in Section 10.

9.3 2009 – 2019 GEOLOGIC MAPPING

During 2009 through 2016, Gold Standard geologists carried out Anaconda-style, layer-based geological mapping that covers a total of 150 km²within and near the Railroad-Pinion property. The mapping was done at scales of 1:6,000 to 1:2,000. The cumulative results of that mapping, combined with published mapping by the U.S.G.S. and the Nevada Bureau of Mines and Geology, as well as certain mapping by historical operators, are shown in Figure 7-3. During 2016-2018, approximately 53.5 km²were mapped in the Dark Star, Dixie, Jasperoid Wash, Ski Track, Elliot Dome, and east Pine Mountain areas. Additional mapping was conducted at a scale of 1:2,000 in the Ski Track and LT areas during 2018.


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Figure 9-2: Rock and Soil Sample Locations 2010 - 2018


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9.4 2014 – 2016 DARK STAR AND PINION PETROGRAPHY

Petrographic analysis systematically describes mineralogical and textural details of rock samples, commonly using thin-section optical microscopy. Consultant Mark McComb of McComb Petrographics performed a petrographic analysis on one sample of Pinion area drill core in 2014, and 14 samples of Dark Star area drill core in 2016. The 2016 samples were from drill hole DS15-13 (Dufresne et al., 2017). McComb (2016) summarized his findings as follows:

“Rock types found in this suite of samples generally include silicified biomicrite, silicified silty to sandy biomicrite, silicified siltstone and sandstone, and decalcified siltstone and sandstone. Gold grades are the highest in samples that contain the most decalcified siltstone and sandstone and were logged as debris flow. Debris flow samples often contain clasts of silicified silty to sandy biomicrite in a decalcified siltstone/sandstone matrix. Decalcified siltstone/sandstone usually has wispy stylolaminated texture attesting to the removal of carbonate and generally comprises detrital quartz in a matrix of low birefringent clay that is often iron stained and contains extremely fine-grained iron oxides. Low birefringent clay appears to be kaolinite, where it is not highly iron stained. Gold mineralization is interpreted to occur in iron oxides, which are interpreted to be oxidized arsenian pyrite. Silica locked extremely fine-grained pyrite can still be observed locally. Mineralized debris flow samples are similar to what is described in the Roberts Mountain DSr3 unit in the northern Carlin Trend.” (pp.1)


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10 DRILLING

The information presented in Section 10 is derived from multiple sources, as cited. The authors have reviewed this information and believe this summary accurately represents the drilling conducted at the Railroad-Pinion property.

10.1 SUMMARY

MDA received from Gold Standard three separate mineral resource-area databases for the Pinion, Dark Star, and Jasperoid Wash deposits, as well as data from other drilling outside the mineral resource areas but within the South Railroad portion of the property. The effective dates of the Pinion and Dark Star databases on which the mineral resources described in this Technical Report are estimated are May 31, 2019 and April 26, 2019, respectively. The effective dates of the Pinion and Dark Star mineral resource estimates are both August 7, 2019. The effective dates of the Jasperoid Wash database and mineral resource estimate are October 6, 2018 and November 15, 2018, respectively. The effective dates of the North Bullion database and mineral resource estimates are August 18, 2017 and September 15, 2017, respectively.

On August 12, 2019 Gold Standard sent a Railroad-Pinion property-wide database containing records from a total of 400,793 m drilled in 1,932 holes since drilling commenced in 1969 (Table 10-1). These totals exclude two holes for which MDA has collar locations, but no depths drilled, hole type, company or assays. Twenty-one different historical operators are known to have drilled 1,084 holes, for a total of 152,566.1 m, from 1969 through 2008. As of the effective dates of the databases, Gold Standard has drilled 848 holes for a total of 248,227 m (Table 10-1). The drilling was done using Imperial units of measure. These were converted to metric measurements. Figure 10-1 shows the distribution of all known drill collar locations in the property.

Approximately 80% of the holes have records to indicate they were drilled with RC methods. There is a total of 10,167 m drilled in 88 historical holes for which MDA has no reliable information on the type of hole or drilling methods used. The authors believe the amount of RC drilling may be understated because the historical holes with no hole-type attribute were drilled in the late 1980s and 1990s when RC drilling was common.

Table 10-1: All Railroad-Pinion Drilling 1969 – 2019


Period	Rotary & RC Holes	Rotary & RC Meters	Core Holes	Core Meters	RC+Core Tail Holes	RC+Core Tail Meters	Unknown	Unknown	Total Holes	Total Meters
Historical Drilling 1969-2008	938	131,854.1	58	10,544.7			88	10,167.4	1,084	152,566.1
Gold Standard 2010-2019	636	164,627.5	171	61,085.0	41	22,514.6			848	248,227.1
Totals	1,574	296,481.6	229	71,629.7	41	22,514.6	88	10,167	1,932	400,793.2

A summary of historical drilling by operator, area and year is presented in Table 10-2. Unless given in the report, the authors are not aware of information on the drilling contractors, rig makes, bit diameters, or specific drilling, logging, and sampling methods and procedures used during any of the historical drilling from 1969 through 2008.


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Figure 10-1: Railroad-Pinion Drill Hole Map (1969 – 2018)

Note: For more detailed depictions of drill holes and mineral resource outlines, see Figure 14-1, Figure 14-10, and Figure 14-22 in Mineral Resource Estimates, Section 14.


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Table 10-2: Historical Drilling Summary


Year	Company	Area Drilled	Rotary Holes	Rotary Meters	RC Holes	RC Meters	Core Holes	Core Meters	? Type Holes	? Type Meters	Total Holes	Total Meters
1969-1970	American Selco	Bald Mountain					7	2,619.0	7	1,205.5	14	3,824.5
1972	Placer Amex	Bald Mountain			1	365.8					1	365.8
1974	El Paso-LLE	Bald Mountain, Pinion			1	254.5	4	618.6			5	873.1
1977-1980	AMAX	Bald Mountain					15	1,893.4			15	1,893.4
1980-1981	AMOCO	Pinion			31	2,897.1					31	2,897.1
1980-1981	Homestake	POD-N.Bullion, Bald Mountain			22	1,764.2					22	1,764.2
1981-1982	Newmont	Irene			6	381.0			23	2,016.9	29	2,397.9
1983	Freeport	Pinion			8	821.4					8	821.4
1983	NICOR	POD-N.Bullion, Bald Mountain			98	11,766.8					98	11,766.8
1984	Cyprus-AMAX	Dark Star	9	1,127.8							9	1,127.8
1985	Santa Fe Mining	Pinion			14	1,543.8					14	1,543.8
1985-1986	NICOR	POD-N.Bullion, Bald Mountain			12	1,880.6					12	1,880.6
1987-1989	Newmont	Irene, Pinion			65	11,314.8			11	559.3	76	11,874.1
1987-1989	Teck	Pinion			39	3,807.0					39	3,807.0
1987-1992	Westmont	POD-N.Bullion, Bald Mountain, Jasperoid Wash, Pinion, Dark Star, JR Buttes			144	18,348.4	3	294.6	9	1,150.6	156	19,793.6
1988	Battle Mountain	Pinion							12	1,159.8	12	1,159.8
1988-1989	Freeport	Dixie			26	3,730.7					26	3,730.7
1990-1993	Crown Resources	Pinion, Dark Star, Dixie			205	25,007.6					205	25,007.6
1993	Unknown	Pinion							2	378.0	2	378.0
1994	Ramrod	POD-N.Bullion, Pinion			13	2,831.6					13	2,831.6
1994-1995	Cyprus	JR Buttes, Pinion			77	13,102.4					77	13,102.4
1995	Newmont	N of N.Bullion							1	425.2	1	425.2
1996	Royal Standard	Pinion					6	358.1			6	358.1
1996-1997	Mirador	Bald Mountain, Pinion, Dark Star, POD-N.Bullion			53	7,743.3			4	283.5	57	8,017.8
1997-1999	Cameco	Dixie, Jasperoid Wash, Pinion, JR Buttes			36	8,533.2	8	3,006.2			44	11,539.3
1998-1999	Kinross	Dark Star, Pinion, POD-N.Bullion, Bald Mountain			68	13,842.5	2	329.2	12	2,639.6	82	16,811.3
2003	Royal Standard	Pinion			10	798.6	4	323.1	3	213.4	17	1,335.1
2005	Unknown	Pinion, POD-N.Bullion							4	135.6	4	135.6
2007-2008	Royal Standard	Pinion, Bald Mountain					9	1,102.5			9	1,102.5
	Grand Total		9	1,127.8	929	130,726.3	58	10,544.7	88	10,167.4	1,084	152,566.1


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10.2 HISTORICAL NORTH RAILROAD DRILLING

10.2.1 1969-1974 American Selco, Placer Amex and El Paso Gas Company

American Selco drilled 7 core holes and 7 holes of unknown type, for a total of 3,824.5 m, exploring for porphyry copper and molybdenum in the general Bald Mountain area in 1969-1970.

In 1972, Placer Amex drilled a single RC hole to a down-hole depth of 365.8 m in the Bald Mountain area exploring for porphyry-type mineralization.

The El Paso Natural Gas Company and Louisiana Land and Cattle Company drilled one RC hole and four core holes for 873.1 m in the Bald Mountain and Pinion areas in 1974.

10.2.2 1977-1980 AMAX

AMAX drilled 15 core holes in the Bald Mountain area in 1977-1980 for a total of 1,893.4 m (Table 10-2). Drill hole AR-7 intersected 29.9 m that averaged 3.77 g Au/t from 11.3 m to 41.1 m near the historic replacement and skarn mines.

10.2.3 1980-1981 Homestake

Homestake drilled 1,764.2 m in 22 RC holes in 1980 and 1981 (Table 10-2). Four of these were drilled in the Bald Mountain area and 18 holes were drilled in the POD-North Bullion area. Homestake’s drilling produced the first significant results in the North Bullion area when hole BDH05 returned 13.1 m with an average of 1.58 g Au/t starting at a down-hole depth of 2.1 m.

10.2.4 1983 and 1985-1986 NICOR

From 1983 through 1986, NICOR drilled a total of 110 RC holes for 13,647.4 m. This included 21 RC holes in the Bald Mountain area for 2,028.4 m. During this period NICOR also drilled 99 RC holes for 11,619 m in the North Bullion area and north of North Bullion. This drilling expanded the drill coverage at North Bullion and resulted in the first historical mineral resource estimate for the POD portion of the North Bullion deposits.

10.2.5 1987-1992 Westmont

Westmont drilled 58 RC holes for 6,616.6 m in the POD-North Bullion area from 1987 through 1992. Three RC holes for 330.7 m were drilled north of the North Bullion deposit area in 1987 and 1990. A total of 1,594.1 m were drilled in 12 RC holes in the Bald Mountain area in 1987-1992.

10.2.6 1994 Ramrod

Ramrod Gold drilled 13 RC holes in the POD-North Bullion area in 1994 for a total of 2,831.6 m.

10.2.7 1995 Newmont

One hole of unknown type was drilled by Newmont north of the deposits in 1995 for 425.2 m.

10.2.8 1996-1997 Mirandor

During 1996 and 1997, Mirandor drilled 28 RC holes in the POD-North Bullion and north of North Bullion areas for a total of 4,157.5 m. Fourteen RC holes were drilled in 1997 in the Bald Mountain area. Hole EMRR-9722 penetrated 21.3 m that averaged 3.81 g Au/t from 4.6 m to 25.9 m, including 13.7 m at a grade of 5.62 g Au/t from 10.7 m to 21.3


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m, and 6.1 m at 8.09 g Au/t from 16.8 m to 22.9 m. This hole was drilled near AMAX hole AR-7, adjacent to the historic Sylvania mine, which had historic production from replacement and/or skarn mineralization.


10.2.9	1998-1999 Kinross

Kinross drilled 37 RC holes and one core hole for 6,652.3 m in the POD-North Bullion deposit area in 1998 and 1999. During this period, 27 RC holes were drilled in the Bald Mountain area for 6,324.6 m. Hole K98-49 intersected 21.3 m with a grade of 3.70 g Au/t at 260.6 m to 281.9 m, including 1.52 m at 13.27 g Au/t from 268.2 m. Hole K99-19 returned a significant interval well away from any previously targeted areas with 3.05 m at 0.89 g Au/t from 185.6 m and 3.05 m at a grade of 0.62 g Au/t from 367.3 m.


10.2.10	2005-2008 Royal Standard Minerals

In 2005, RSM drilled a total of 536.5 m in four core holes and three holes of unknown type in the POD-North Bullion area. At the Bald Mountain area, RSM drilled three core holes in 2007 and one core in 2008 for 692.5 m.


10.3	HISTORICAL SOUTH RAILROAD DRILLING

10.3.1	1980-1981 AMOCO Minerals

AMOCO drilled 31 RC holes for 2,897.1 m in the Pinion area in 1980 and 1981.


10.3.2	1981-1982 Newmont

The Irene prospect was tested by Newmont in 1981 and 1982 when six RC holes and 21 holes of unknown type were drilled for 2,397.2 m.


10.3.3	1983 Freeport

In 1983, Freeport drilled eight RC holes for 821.4 m in the Pinion deposit area.


10.3.4	1984 Cyprus-AMAX

The Dark Star area was first tested by Cyprus-AMAX with nine rotary holes for 1,127.8 m in 1984.


10.3.5	1985 Santa Fe Mining

Santa Fe Mining drilled 14 RC holes for 1,543.8 m in the Pinion deposit in 1985.


10.3.6	1987-1989 Newmont

Newmont drilled four RC holes and 11 holes of unknown type for 1,371.6 m in the Irene prospect during 1987 through 1989. During this same time period, Newmont drilled 61 RC holes in the Pinion deposit and vicinity.


10.3.7	1987-1989 Teck Resources

Teck drilled 39 RC holes for 3,807 m in the Pinion deposit.


10.3.8	1988 Battle Mountain

A total of 12 holes of unknown type and 1,159.8 m were drilled at the Pinion area by Battle Mountain Gold Corp. (“BMGC”) or Battle Mountain Exploration Co. (“BMEC”) in 1988.


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10.3.9	1989-1992 Westmont

Westmont first drilled in the Jasperoid Wash area with 48 RC holes and two core holes for 6,800.5 m in 1989 through 1992. The Pinion area was drilled by Westmont in 1989 with nine holes of unknown type for 1,150.6 m. In 1991, Westmont drilled two RC holes at Pinion for 207.3 m. Three RC holes for 544.1 m were drilled at Dark Star by Westmont in 1991. Westmont tested the JR Buttes prospect in 1992 with 19 RC holes for 2,549.6 m.


10.3.10	1988-1989 Freeport

The Dixie prospect was tested by Freeport with 26 RC holes for 3,730.7 m drilled.


10.3.11	1990-1993 Crown Resources

In 1990, Crown began drilling in the Pinion deposit and by 1993 had drilled 12,297.1 m in 130 RC holes. Crown also drilled 11,235.2 m in 69 RC holes at the Dark Star deposit in 1991 through 1993. A total of 1,554.5 m in seven RC holes were also drilled by Crown at the Dixie prospect in 1991, following up on the drilling done there by Freeport.


10.3.12	1994-1995 Cyprus Mining

During 1994 and 1995, Cyprus drilled at total of 12,441 m in 73 RC holes in the Pinion deposit area. Cyprus also drilled three RC holes for a total of 464.8 m at the JR Buttes prospect.


10.3.13	1997 Mirandor

Mirandor drilled a total of 2,203.7 m in 11 RC holes at the Dark Star deposit in 1997. A total of 283.5 m in four holes of unknown type were also drilled in the Pinion deposit area.


10.3.14	1997-1999 Cameco

Cameco’s drilling during this period was focused at the Pinion deposit area with a total of 8,76.8 m drilled in 20 RC holes and eight core holes. A total of 2,685.3 m in 11 RC holes were drilled by Cameco in the Dixie prospect in 1997 and 1998, and one RC hole for 221.0 m was drilled in 1998 at JR Buttes. In 1997, Cameco also drilled 556.3 m in four RC holes at the Jasperoid Wash area.


10.3.15	1998-1999 Kinross

Kinross focused their 1998 and 1999 drilling in the South Railroad portion of the property at Dark Star with one core hole, three RC holes and 11 holes of unknown type for a total of 3,378.7 m. A total of 455.7 m were also drilled in two RC holes in the Pinion deposit area.


10.3.16	2003 and 2007 Royal Standard Minerals

In 2003, RSM drilled a total of 798.6 m in 10 RC holes in the Pinion deposit area. RSM subsequently drilled five core holes at the Pinion deposit area in 2007, for a total of 410.0 m.


10.4	GOLD STANDARD DRILLING, NORTH RAILROAD AREA 2010 - 2017

Gold Standard’s drilling in the North Railroad portion of the property commenced in 2010. As summarized in Table 10-3, a total of 79,550.4 m have been drilled in 154 holes as of the effective date of the database of this Technical Report. Approximately 31% of the meters and 34% of the holes were drilled with RC methods. Diamond-core drilling accounts for 43% of the meters and 42% of the holes; the balance of the drilling was done using RC followed by core tails.


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Gold Standard’s RC holes were drilled wet; water was always injected. Face-return bits were only used when interchanges were flooded out. Tri-cone bits were only used when the hammer bits were ineffective due to too much water.

For core drilling, Gold Standard geologists completed paper or digital logs on the whole core. The logs captured and illustrated core recovery, sample intervals, lithologic data, hydrothermal alteration, mineralogy, and structural features. Structural features were measured with respect to the core axis. When available, structural features were measured on core oriented using a Reflex Act 2 orienting device. Photographs were taken of all drill core, labeled with drill hole footages and sample intervals. RC drill chips were also logged on paper or digital logs by Gold Standard geologists. The data from the paper drill logs were later captured in electronic spreadsheets for both core and RC drill holes.


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Table 10-3: Summary of Gold Standard Drilling 2010 – 2018


Year	Area	RC Holes	RC Meters	Core* Holes	Core* Meters	RC + Core Holes	RC + Core Meters	Total Holes	Total Meters
*North Railroad*
2010	POD-N. Bullion	6	2,843.8	5	2,237.7	4	1,857.8	15	6,939.2
	N of N. Bullion	1	609.6					1	609.6
2011	POD-N. Bullion	5	1,693.6	5	2,897.1	7	4,063.9	17	8,654.6
	Bald Mountain			4	1,483.8			4	1,483.8
	N of N. Bullion	2	1,549.9	1	1,105.7			3	2,655.6
2012	POD-N. Bullion	4	1,824.2	25	13,267.5	2	1,396.9	31	16,488.6
	Bald Mountain			3	1,770.9			3	1,770.9
2013	POD-N. Bullion	5	2,308.9	15	8,202.5			20	10,511.3
2013	Bald Mountain	4	2,436.9	3	1,582.5			7	4,019.4
2014	Bald Mountain	5	1,895.9					5	1,895.9
2015	POD-N. Bullion			2	958.0	2	708.4	4	1,666.4
2016	Bald Mountain	9	5,010.9					9	5,010.9
2016	POD-N. Bullion	1	666.0			9	5,255.4	10	5,921.4
2017	Bald Mountain	4	1,623.1					4	1,623.1
2017	POD-N. Bullion	6	2,384.0	1	440.4	14	7,475.4	21	10,299.8
2010- 2017	*N. Railroad* *Totals*	52	*24,846.7*	64	*33,946.0*	38	*20,757.7*	*154*	*79,550.4*
*South Railroad*
2012	Pinion & Vicinity	6	3,026.7					6	3,026.7
2014	Pinion & Vicinity	53	12,608.1	4	482.8			57	13,090.9
	Pinion & Vicinity	23	8,804.2					23	8,804.2
2015	Dark Star	13	5,048.1					13	5,048.1
	Irene	1	605.0					1	605.0
	Pinion & Vicinity	17	5,952.7	6	752.2			23	6,704.9
2016	Dark Star	17	7,726.7	19	7,987.4	2	947.6	38	16,661.7
2016	Dixie	2	1,190.2					2	1,190.2
	Irene	2	1,356.4					2	1,356.4
	Pinion & Vicinity	16	1,917.2	3	420.6			19	2,337.8
2017	Dark Star	36	13,059.9	11	2,378.4			47	15,438.3
2017	Jasperoid Wash	10	3,555.5	2	790.0			12	4,345.5
	Dixie	16	7,287.8	2	850.1			18	8,137.9
	Pinion & Vicinity	109	13,214.6	35	4,128.5			144	17,343.1
	Dark Star	117	21,755.1	19	3,933.6			136	25,688.7
2018	Jasperoid Wash	44	8,945.2	4	4,706.7			48	13,652.0
	Dixie	25	10,962.1	1	475.8	1	809.2	27	12,247.2
	Ski Track	6	2,036.1					6	2,036.1
2019	Dark Star	71	10,729.2	1	232.9			72	10,962.0
2012- 2018	*S. Railroad* *Totals*	*584*	*139,780.8*	*107*	*27,139.1*	3	*1,756.9*	*694*	*168,676.7*
	*Grand Totals*	*636*	*164.627.5*	*171*	*61,085.0*	41	*22,514.6*	*848*	*548,227.1*


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10.4	North Bullion Deposits Drilling by Gold Standard

Drilling by Gold Standard in the North Bullion area commenced in 2010 and a total of 60,481.4 m had been drilled in 119 holes through the end of 2016. An additional 25 holes were completed in 2017 for a total of 10,299.8 m. Drill collar locations in the North Bullion area are shown in Figure 10-2.


10.4.1.1	2010-2013 North Bullion Deposits Drilling

From 2010 through 2013, Gold Standard drilled 83 holes totalling 42,593.8 m in the North Bullion area (Table 10-3; Figure 10-2; Hunsaker, 2012a, b; Shaddrick, 2012; Koehler et al., 2014). In 2010, Gold Standard utilized gravity data and geological models to identify an untested target that lead to intercepts of 32 m of 1.39 g Au/t and 43.6 m of 1.21 g/t Au in hole RR10-8 at the North Bullion deposit (Jackson et al., 2015). This discovery of blind, sedimentary-rock hosted, Carlin-style gold mineralization lead to additional drilling conducted from 2010 to 2013 within the North Bullion deposit area and eventually to the estimated gold mineral resources presented in Section 13. The true thickness of mineralization in the POD deposit and North Bullion deposit, and its relationship to drill interval lengths, is discussed in Section 13 of this Technical Report.

Gold Standard’s 2010 and 2013 RC drilling was conducted by Hard Rock Exploration Inc. (“Hardrock”) and National Exploration Wells and Pumps (“National”), using a TH75 and 685 Schramm, respectively. Bit sizes were 5 ¼ in. to 6 ½ in. diameter bits. The rig was operated on one or two 12 hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.

Core drilling in 2010 to 2013 was done by Redcor Drilling Inc. with an LF-230 rig. Core sizes were PQ3, HQ3, and NQ3.

No drilling was done in 2014.


10.4.1.2	2015 North Bullion Deposits Drilling

In 2015, Gold Standard drilled two core holes and two RC holes with core tail holes totalling 1,666.4 m (Table 10-3; Figure 10-2; Turner et al., 2015; Dufresne and Koehler, 2016). The RC drilling was conducted by National using a 685 Schramm. Bit sizes were 5 ¼ in. to 6 ½ in. diameter bits. The rig was operated on one or two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.


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Figure 10-2: Map of North Railroad Property Drill Collar Locations


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The 2015 core drilling was performed by Timberline Drilling (“Timberline”) of Elko Nevada using an LF90 drill rig Core sizes were PQ3, HQ3, and NQ3. Core was also drilled by TonaTec Exploration LLC (“TonaTec”) of Utah. The rig may have been a CS2000. Core sizes were PQ3, HQ3, and NQ3.


10.4.1.3	2016-2017 North Bullion Deposits Drilling

A total of 16,221.1 m were drilled in 31 holes in 2016 and 2017 (Table 10-3; Figure 10-2). Most of the RC drilling was conducted by National using a 685 Schramm. Bit sizes were 5 ¼ in. to 6 ½ in. diameter bits. The rig was operated on one or two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.

Boart Longyear of Elko, Nevada was the contractor for four RC holes drilled in 2017. A track-mounted drill of unknown type was used; specific methods and procedures are not reported.

The 2015 core drilling was performed by Timberline of Elko Nevada using an LF90 drill rig. Core sizes were PQ3, HQ3, and NQ3. Core was also drilled by First Drilling (“First Drilling”) of Elko Nevada. The rig was an LF90. Core sizes were PQ3, HQ3, and NQ3.

The results from drilling completed prior to August 18, 2017 were used to estimate the current gold mineral resources presented in Section 14.5 of this Technical Report.


10.4.2	Bald Mountain Drilling by Gold Standard

A total of 15,803.9 m were drilled by Gold Standard in 22 RC and 10 core holes in the Bald Mountain area from 2011 through 2017 (Table 10-3; Figure 10-2). Drilling contractors, rig types and diameters for the Bald Mountain area drilling are summarized in Table 10-4.

All 2011-2017 core drilling was done with two 12-hr shifts per day. The RC drills operated for one or two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone.

Table 10-4: Bald Mountain Drilling Contractors and Methods


Year	2011 to 2013	2014	2016	2017
RC Contractor	NA	Hardrock	National	Boart Longyear
RC Drill Rig	NA	TH75	685 Schramm	MPD 1500
RC Diameter	NA	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.
Core Contractor	Redcor	NA	NA	NA
Core Drill Rig	LF-230	NA	NA	NA
Core Diameter	PQ3, HQ3, and NQ3	NA	NA	NA


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10.5	GOLD STANDARD DRILLING, SOUTH RAILROAD AREA 2012-2019

Drilling in the South Railroad portion of the property by Gold Standard commenced in 2012. As summarized in Table 10-3, a total of 168,676.7 m were drilled in 694 holes (Figure 10-1). Approximately 83% of the meters and 84% of the holes were drilled with RC methods. Diamond-core drilling accounts for about 16% of the meters and 15% of the holes; the balance of the drilling was done using RC followed by core tails. Both angle and vertical drilling was done.

A Gold Standard representative checked each drill rig at least once per day during drilling to monitor sample collection. For core drilling, Gold Standard geologists completed paper or digital logs on the whole core. The logs captured and illustrated core recovery, sample intervals, lithologic data, hydrothermal alteration, mineralogy, and structural features. Structural features were measured with respect to the core axis. When available, structural features were measured on core oriented using a Reflex Act 2 orienting device. Photographs were taken of all drill core, labeled with drill hole footages and sample intervals. RC drill chips were also logged on paper or digital logs by Gold Standard geologists. The data from the paper drill logs were later captured in electronic spreadsheets for both core and RC drill holes.

Gold Standard’s RC holes were drilled with water injection. Face-return bits were utilized when not impeded by excess water. Tri-cone bits were only used when the hammer bits were unable to function due to excessive water pressure.


10.5.1	Dark Star Area Drilling by Gold Standard

In 2015, Gold Standard began drilling in the Dark Star deposit area to extend historically known shallow oxidized gold mineralization and to test other exploration targets. In 2015 through 2019, Gold Standard drilled a total of 73,798.9 m in 306 holes (Table 10-3). RC drilling accounts for about 83% of the holes and 79% of the meters drilled by Gold Standard. Collar locations for the Gold Standard drilling at Dark Star are shown in Figure 10-2 and in greater detail in Figure 14-1.

Drilling contractors, rig types and diameters for the Dark Star area drilling are summarized in Table 10-5. All 2015-2019 core drilling was done with two 12-hr shifts per day. The RC drills operated for one or two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone.

Table 10-5: Gold Standard’s Dark Star Drilling Contractors and Methods


Year	2015	2016	2017	2018
RC Contractor	National	National	National; Boart Longyear	National; Boart Longyear
RC Drill Rig	T450GT, 685 Schramm	685 Schramm	685 Schramm, T450GT; 685 Schramm, MPD1500	685 Schramm, T450GT, EDM95; 685 Schramm, MPD1500
RC Diameter	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.
Core Contractor	National	National; Timberline	First Drilling; National	First Drilling; National; Boart Longyear
Core Drill Rig	CT14	CT14; LF90	LF90; CT14	LF90; CT14; LF90
Core Diameter	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3
Core Diameter	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3


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Highlights from the 2016 drill program at Dark Star and an updated mineral resource estimate were presented by Dufresne and Nicholls (2017a). The true thickness of mineralization in the Dark Star deposit, and its relationship to drill interval lengths, is discussed in Section 14.5 of this Technical Report.


10.5.2	Pinion Area Drilling by Gold Standard

Gold Standard’s drilling in the Pinion deposit area (Figure 10-1) has totalled 51,307.5 m in 272 holes drilled from 2012 through the effective date of the database (Table 10-3). The great majority of the drilling, approximately 90% of the meters drilled, was done with RC methods. Contractors, rig types, and hole diameters for the Pinion area drilling by Gold Standard are summarized in Table 10-6.

Following acquisition of the Pinion deposit area in 2014, in the South Railroad part of the property, Gold Standard focused their drilling on the expansion and infill drilling of various zones of what is now the Pinion gold deposit. The 2014 drilling (Table 10-3) produced significant gold intervals at the Pinion deposit indicating that gold mineralization associated with multi-lithic breccia and certain structures remained open along and across strike. Further drilling of 23 holes in 2015 also provided significant gold intercepts indicating the mineralized system was still open in a number of directions.

Table 10-6: Gold Standard Pinion Area Drilling Contractors and Methods


Year	2014	2015	2016	2017	2018
RC Contractor	Hard Rock; Major	Hard Rock; National	National	Boart Longyear	National; Boart Longyear
RC Drill Rig	TH75; T450GT	TH75; T450GT, 685 Schramm	685 Schramm	685 Schramm, MPD1500	450 Schramm; 685 Schramm
RC Diameter	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.	5¼ in. to 6½ in.
Core Contractor	Major	NA	National; Timberline	National	First Drilling; Boart Longyear
Core Drill Rig	LF230	NA	CT14; LF90	CT14	LF90; LF90
Core Diameter	PQ3, HQ3, and NQ3	NA	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3	PQ3, HQ3, and NQ3

In 2016, Gold Standard drilled a total of 23 holes in the Pinion deposit area for a total of 8,804.2 m. This drilling was designed to extend known zones of mineralization, provide infill data for specific zones, and provide material for metallurgical testing. Several holes were drilled to test the Irene geological and geochemical target 2.0 km west of the Pinion deposit (Figure 10-1) and at the Sentinel target to the north of the Pinion deposit.

The 2016 Pinion drilling resulted in several significant gold intersections, defined as averaging greater than the 0.14 g Au/t cut-off grade that was used previously for the 2016 estimate of Pinion gold mineral resources (Dufresne and Nicholls, 2016). Most significantly, the 2016 drilling identified a new stratigraphic target called the Sentinel zone, which is located at the north end of the Pinion deposit area and comprises gold hosted within the Sentinel Mountain dolomite and the top of the underlying Oxyoke sandstone, below the Devils Gate Limestone. The Sentinel gold mineralization is shallow, oxidized, and open to the north and west.

Gold Standard’s 2014 through 2018 RC drilling was conducted on one or two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. The splitter reduced the samples to approximately 2.3 to 9.1 kg, which were collected in pre-numbered sample bags. A few grams of each 1.524 m interval were placed in chip trays for logging.


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Results from the 2014 through 2018 Gold Standard drilling were used with data from historical drilling to estimate the current gold mineral resources presented in Section 14.3 of this Technical Report. The true thickness of mineralization in the Pinion deposit, and its relationship to drill interval lengths, is discussed in Section 14 of this Technical Report.


10.5.3	Jasperoid Wash Area Drilling by Gold Standard

Gold Standard’s drilling at the Jasperoid Wash deposit area commenced in 2017. Since then a total of 17,997.5 m have been drilled in 60 holes (Table 10-3). RC drilling accounts for about 90% of the holes and 70% of the meters drilled by Gold Standard. Collar locations for the Gold Standard drilling at Jasperoid Wash are shown in Figure 10-1 (see Section 14.4 and Figure 14-22 for a detailed map).

The 2017 and 2018 RC drilling was conducted by National using a 450 Schramm, 685 Schramm, and an EDM 95. Major also drilled at Jasperoid Wash and used a 455 Schramm. Bit sizes were 5¼ in. to 6½ in. in diameter. The rig was operated on two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.

Core drilling in 2017 and 2018 was carried out by National and First Drilling using a CT14 and an LF90, respectively. Core sizes drilled were PQ3, HQ3, and NQ3.

The results of the Gold Standard drilling, together with historical drill data from Jasperoid Wash, have been used to estimate the current gold mineral resources presented in Section 14.4 of this Technical Report. The true thickness of mineralization in the Jasperoid Wash deposit, and its relationship to drill interval lengths, is shown in Section 14.4 of this Technical Report.

After the 2018 estimate was completed (reported in this Technical Report in 2019), 21 additional drill holes (5,210.3 m) were completed at Jasperoid Wash, two were core holes (610.8 m) and the remainder were RC (4,599.5 m). These holes’ data were received on April 24, 2019 and evaluated on April 29, 2019. No auditing or QA/QC evaluations were done on this post-model drill-hole data set.


10.5.4	Irene Area Drilling by Gold Standard

Three RC holes for a total of 1,961.4 m were drilled at the Irene prospect about 2 km west of the Pinion deposit in 2015 and 2016 (Table 10-3). Drilling done at Irene used drill rigs similar to those used for the Pinion drilling.


10.5.5	Dixie Area Drilling by Gold Standard

The Dixie prospect, including Arturus and Elliot Dome targets, located about 3 km south of Dark Star, was drilled by Gold Standard in 2016, 2017, and 2018. A total of 19,440.1 m were drilled in 43 RC holes, three core holes, and one RC pre-collar holes with a core tail (Table 10-3). This drilling was conducted by National using a 685 Schramm, 450 Schramm, and EDM 95, and Boart Longyear using a 685 Schramm or MPD1500. Major also drilled at Dixie in 2018 using a 455 and a 685 Schramm. Bit sizes were 5¼ in. to 6½ in. diameter. The rigs operated on two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.

Core drilling was conducted by National and First Drilling using a CT14 and LF90, respectively. Core sizes drilled were PQ3, HQ3, and NQ3.


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10.5.6	Ski Track Drilling by Gold Standard

Six RC holes were drilled in 2018 at the Ski Track prospect for a total of 2,036.1 m. This drilling was done by Major using a 685 Schramm. Bit sizes were 5¼ in. to 6½ in. in diameter. The rigs operated on two 12-hr shifts per day. RC samples were collected continuously over 1.524 m (5 ft) intervals and split with a rotating wet splitter located beneath the cyclone. A drilling technician placed a few grams of each 1.524 m interval in plastic chip trays for logging.


10.6	DRILL-HOLE COLLAR SURVEYS

10.6.1	Historical Collar Surveys, North Railroad Portion of the Property

APEX stated that collar locations were rectified to a satellite orthophoto with one-meter contours (Dufresne and Nicholls, 2017b). Elevations for all the remaining holes were adjusted to a topographic surface created from the orthophoto.


10.6.2	Historical Collar Surveys, South Railroad Portion of the Property

Mr. Ristorcelli has no information on the methods used to survey the locations of the historical drill collar locations in the South Railroad portion of the property. Coordinates for historical drill holes at the Pinion, Dark Star, and Jasperoid Wash deposits were obtained from old records, resurveying in the field, and taken from historical maps. Much work was done by Gold Standard and APEX resolving collar location issues. However, those few that did contradict surrounding holes, or whose geology and grades were improbable, were eliminated from use in modeling and estimation.


10.6.3	Gold Standard Collar Surveys, North Railroad Portion of the Property

Gold Standard has performed differential Global Positioning System (“GPS”) surveys of all collar locations for holes drilled from 2010 through 2017. The surveys were carried out by Apex Surveying LLC out of Spring Creek Nevada using a Trimble differential GPS. Where possible, the locations of historical drill collars were also surveyed. During the site visits, Mr. Dufresne located some historical and Gold Standard drill collars using a hand-held GPS, along with tracks representing drill roads and trails. Although unmarked in the field, several drill collars were ascertained due to their unique location, which were found to be consistent with historically recorded location information. Further work on refining the collar positions has been performed by Gold Standard personnel and reviewed by the author of this section of the report.

The most significant problem with the historical drill locations are collar elevations which initially had obvious errors. With near flat-lying mineralized zones it was imperative to obtain a reliable dataset of collar elevations that were internally consistent from one hole to the next. Once accurate real-world coordinates were obtained for the historical collars, elevations were obtained by projecting the collars to a digital elevation model that was generated by Pacific Geomatics from ortho-rectified satellite imagery with ~1 m elevation and horizontal resolution.


10.6.4	Gold Standard Collar Surveys, South Railroad Portion of the Property

As stated in Section 10.6.3, the collar locations for all Gold Standard holes drilled through 2018 were surveyed by differential GPS. After the holes were abandoned, the collars were marked by wooden lath with the hole name on a wire and aluminum tag placed in the cement collar plug. Apex Surveying, LLC, of Spring Creek, Nevada professionally surveyed the Gold Standard drill collars at the Pinion and Dark Star deposits using a “differential GPS” according to APEX.


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10.7	DOWN-HOLE SURVEYS

10.7.1	Historical Down-Hole Surveys, North and South Railroad Portions of the Property

APEX reported that most of the deeper historical drill holes in the Railroad-Pinion property were downhole surveyed (Dufresne and Nicholls, 2017b). Survey equipment used is unknown. During 1999, at least a portion of the Kinross drill holes in various areas of the property were surveyed down-hole by Silver State Surveys of Elko, Nevada (Jones et al., 1999), but the type of instrument and methods and procedures are not known.


10.7.2	Gold Standard Down-Hole Surveys, North and South Railroad Portions of the Property

Gold Standard contracted International Directional Services (“IDS”), who used Stockholm Precision Tools with a continuous-read, north-seeking gyro down-hole surveying tool named Memory North Seeking Gyroscopic Inclinometer. IDS has also used an Axis Champ Navigator, supplied by Axis Mine Tech. In 2017, Gold Standard contracted Minex, using a MEMS continuous-read, north-seeking gyro down-hole surveying tool. All holes longer than ~100 m were down-hole surveyed for azimuth and dip.


10.8	SUMMARY STATEMENT

The authors believe that the drilling, sampling, and logging methods and procedures provided samples that are representative and of sufficient quality for use in the mineral resource estimations subject to the elimination of some drill holes and some samples, and to the downgrading of mineral resource classification when blocks were dominantly estimated by historical drilling (discussed in Section 14). The authors are aware of sampling or recovery factors that impact the reliability of the samples for use in a mineral resource estimate. Those samples were removed from use in estimation (discussed in Section 14).


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11	SAMPLE PREPARATION, ANALYSES AND SECURITY

The information presented in Section 11 is derived by MDA from Dufresne et al., 2017, Dufresne and Nicholls (2017b), data received directly from Gold Standard and other sources, as cited. The authors have reviewed this information and believe this summary accurately represents the methods, procedures and analyses used for the drilling samples on which the estimated mineral resources presented in Section 14 of this Technical Report are based.

Documentation of the methods and procedures used for historical surface and drilling sample collection, preparation, analyses, and sample security at the Railroad-Pinion property is incomplete and in many cases is not available. MDA recommends that Gold Standard compile and evaluate the information contained in records that are available.

Methods and procedures used for the security, preparation, and analysis of surface samples collected by historical operators and Gold Standard have not been evaluated for this Technical Report because the results have not been used in the estimation of the mineral resources presented in Section 14. While useful for identifying drilling targets and planning exploration drilling, the results and representativity of the Gold Standard surface sampling are not of material importance to the interpretations and conclusions of this Technical Report. The reader is referred to Koehler et al. (2014), Dufresne et al. (2014; 2015; 2017) and references cited in those reports for information on Gold Standard’s soil- and rock-sample collection, security, preparation, and analyses.


11.1	HISTORICAL OPERATORS’ DRILLING SAMPLES - NORTH RAILROAD PORTION OF THE PROPERTY

Historical drill logs and reports in the possession of Gold Standard have not been evaluated. MDA recommends that Gold Standard extract and compile information from available documents regarding logging methods, and where available information on core diameters, RC-bit diameters, and sample splitting prior to shipment to the analytical laboratories.

The authors and Gold Standard are not aware of the methods and procedures used by American Selco, Placer Amex, El Paso, AMAX, Homestake, and NICOR for historical drill-sample collection, splitting, preparation, analyses, and sample security during drilling at Bald Mountain and North Bullion from 1969 through 1986.

Samples from the Westmont drilling in the North Bullion area in 1987 were analyzed for gold and silver by fire assay methods at Universal Laboratory, Inc. (“Universal”), in Elko, Nevada. It is not known if this laboratory was independent of Westmont, or if any certifications were held. Samples from Westmont’s drilling at North Bullion in 1990 and 1992 were analyzed at Cone Geochemical Inc. (“Cone”), in Lakewood, Colorado. Gold was determined by fire-assay fusion of 30 g aliquots. Cone was independent of Westmont, but MDA is not aware if any certifications were held by Cone at that time. MDA is not aware of sample security measures taken or the details of transport from the drill sites to the laboratories.

Samples from Ramrod’s drilling in the North Bullion area in 1994 were assayed at Cone and at Monitor Geochemical Laboratory Inc. (“Monitor), in Elko, Nevada. At Cone, gold was determined by fire-assay fusion of 20 g and 30 g aliquots with an atomic adsorption (“AA”) finish. At Monitor, Ramrod’s samples were analyzed for gold and silver by 30 g fire-assay fusion and some were analyzed by cyanide-leach with an AA finish. Some composited pulps representing 8.7 m (25 ft) lengths were analyzed for arsenic, antimony and mercury by unspecified method(s). Monitor and Cone were independent of Ramrod. It is not known if any certifications were held by these laboratories at the time. MDA is not aware of sample-security measures taken or the details of transport from the drill sites to the laboratories.

In 1997, Mirandor’s drill samples from north of North Bullion and the Bald Mountain areas were analyzed by Interteck Testing Services, a division of Bondar-Clegg & Company Ltd. (“Bondar-Clegg”), in North Vancouver, British Columbia. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some samples were re-analyzed for gold by 30 g fire assay with a gravimetric finish. Silver was determined by AA and inductively-coupled plasma-emission spectrometry (“ICP). Some samples were analyzed for copper, lead, zinc, molybdenum, arsenic, and antimony by AA,


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and for mercury by cold-vapor AA (“CVAA”). Bondar-Clegg was independent of Mirandor. MDA is not aware if any certifications were held by Bondar-Clegg at that time. MDA is not aware of sample-security measures taken or the details of transport from the drill sites to the laboratory.

Samples from Kinross’ drilling in 1998 and 1999 at North Bullion and Bald Mountain were analyzed at Chemex Labs, Inc. (“Chemex”), in Sparks, Nevada. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some samples were re-analyzed for gold by 30 g fire assay with a gravimetric finish. Composited pulps representing 8.7 m (25 ft) sample lengths were analyzed ICP for 35 minor, major, and trace elements, including silver. Chemex was independent of Kinross. MDA is not aware if any certifications were held by Chemex at that time. MDA is not aware of sample-security measures taken and the details of transport from the drill sites to the laboratory.


11.2	GOLD STANDARD’S DRILLING SAMPLES - NORTH RAILROAD PORTION OF THE PROPERTY

Commencing in 2010, drilling company employees collected Gold Standard’s RC samples at the rig. Those samples were then picked up at the drill sites by representatives of ALS Minerals (“ALS”) or Inspectorate America Corporation (“Inspectorate”), a division of Bureau Veritas Mineral Laboratories USA (“Bureau Veritas”) and transported by truck to their respective laboratories in either Elko or Reno, Nevada (for ALS), or Elko (for Bureau Veritas). Excessively wet samples were kept at the drill sites for a few days to drain and dry prior to collection by the laboratory staff.

ALS and Bureau Veritas were, and continue to be, commercial laboratories independent of Gold Standard. ALS is accredited to the standard ISO/IEC 17025:2005 for specific analytical procedures, while most of their laboratories have attained ISO 9001:2008 certification. Bureau Veritas’ laboratories in Sparks, Nevada is accredited to the standard ISO/IEC 17025:2017, RG- MINERAL:2017. The Bureau Veritas laboratory in Vancouver, British Columbia is accredited to the standard ISO/IEC 17025:2005 and ISO 9001:2008.

Core samples were transported daily from the drill sites to Gold Standard’s logging and core-cutting facility in Elko by Gold Standard personnel. After logging and marking core-sample intervals by Gold Standard geologists, the core was photographed prior to being sawed lengthwise by contractor technicians. Whole HQ-size core was sawed in half. Whole PQ-size core was sawed in quarters. One half of the HQ core, and three quarters of the PQ core, were returned to the core boxes and the remainder was placed in pre-numbered sample bags that were closed with ties. Following insertion of quality assurance/quality control (“QA/QC”) blanks and certified reference materials (“CRMs”), the core samples were transported by representatives of ALS or Bureau Veritas to their respective laboratories for preparation and analysis.

Samples from Gold Standard’s RC and core drilling at North Bullion in 2010 through 2014, and at Bald Mountain in 2014, were prepared at the ALS laboratories in Elko and Reno, Nevada. The samples were dried and crushed in their entirety to 70% at less than 2 mm. The crushed samples were riffle-split to obtain 250 g subsamples that were pulverized to 85% less than 75 microns. The pulps were shipped by air freight by ALS to the ALS laboratory in North Vancouver, British Columbia, for analysis. Gold was determined by 30 g fire-assay fusion with an AA finish (method code Au-AA23). Samples assayed at ≥10 g Au/t were re-analyzed with a second 30 g aliquot by fire-assay fusion and gravimetric finish (method code Au-GRA21). Separate aliquots of 0.5 g were analyzed for silver and 34 major, minor and trace elements by ICP following an aqua regia digestion. In some cases, the ICP analyses were conducted on pulps from 1.524 m drill samples. In other cases, ICP analyses were conducted on composited pulps representing 6.008 m (20 ft) drill intervals. Samples that assayed >10,000 ppm for silver or zinc by ICP were re-analyzed using AA following aqua regia digestion of 0.1 g aliquots.

A minority of the 2010 through 2012 drill samples were analyzed by SGS Canada Inc. (“SGS”) of Vancouver, British Columbia. The assay certificates do not indicate how or where the samples were prepared for analysis. At the SGS laboratory in Burnaby, British Columbia, gold was determined by 30 g fire-assay fusion with an AA finish and separate


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aliquots were analyzed by ICP for 35 major, minor and trace elements. SGS was a commercial laboratory independent of Gold Standard. MDA is not aware of certifications held by SGS at that time.

In 2013, pulps from previously prepared samples from North Bullion were analyzed by Bureau Veritas in Sparks, Nevada. Gold was determined by 30 g fire-assay fusion with an AA finish. Some of the samples were analyzed using a 30 g aliquot by fire-assay fusion and gravimetric finish. In 2014, some of the Bald Mountain drill sample pulps were re-analyzed at Bureau Veritas’ laboratory in Vancouver, British Columbia for copper by cyanide-H2SO4 leach. Other pulps were analyzed for 45 major, minor and trace elements by a combination of ICP and mass spectrometry (“ICP-MS”) after 4-acid digestion.

Samples from the 2015, 2016. and 2017 drilling at North Bullion and Bald Mountain were analyzed at ALS and Bureau Veritas. At ALS the methods and procedures of preparation were the same as those used in 2010 through 2014. Gold was determined using ALS method code Au-AA23 and Au-GRA21 principally in the ALS laboratory in North Vancouver. Most gold assays on 2017 North Bullion samples were performed in the ALS laboratory in Reno with the same methods (Au-AA23; Au-GRA21). Separate aliquots of 0.5 g were analyzed for silver and 34 major, minor and trace elements by ICP following an aqua regia digestion in the North Vancouver laboratory. In some cases, these were composited pulps representing 6.008 m (20 ft) drill intervals.

A significant portion of the samples from the 2016 North Bullion drilling, and the majority of the 2017 North Bullion samples, were prepared and analyzed by Bureau Veritas. These samples were prepared in the Bureau Veritas laboratory in Elko. Following crushing, a 250 g riffle-split subsample was obtained from each drill sample. These subsamples were pulverized to 200-mesh size and the pulps were shipped to the Bureau Veritas laboratory in Sparks, Nevada. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. The pulps were shipped via air freight by Bureau Veritas to their analytical laboratory in Vancouver where they were analyzed for 45 major, minor and trace elements by ICP-MS after four-acid digestion.


11.3	HISTORICAL OPERATORS - SOUTH RAILROAD PORTION OF THE PROPERTY

AMOCO and Cyprus’ drilling samples from the Pinion area in 1980 and 1981 were mainly analyzed at Barringer Resources, Inc. (“Barringer”) in Sparks, Nevada. Gold and silver were determined by fire-assay fusion of 30 g aliquots. Some samples were also analyzed for arsenic and mercury, but no other information is available. In 1980, some of AMOCO’s samples were analyzed for silver and gold at Monitor, but the methods of analysis are not available. Barringer and Monitor were independent of AMOCO and Cyprus. MDA is not aware of any certifications that may have been held by these laboratories at that time.

In 1981, Newmont’s drilling samples from the Irene area were analyzed at Monitor in Elko. Gold and silver were determined by fire-assay fusion, but MDA has no other information on the methods and procedures used. Newmont’s 1982 drilling samples from the Pinion area were analyzed at Skyline Labs Inc. (“Skyline”), in Tucson, Arizona. Gold was determined by fire-assay fusion, but no other information is available. Skyline and Monitor were independent of Newmont, but MDA is not aware of any certifications that may have been held by these laboratories at that time.

Santa Fe’s samples from their 1985 drilling in the Pinion area were analyzed by Monitor in Elko. Gold was determined by fire-assay fusion of 30 g aliquots, but no other information is available. Monitor was independent of Santa Fe, but MDA is not aware of any certifications that may have been held by Monitor at that time.

Samples from Teck Resource’s drilling in the Pinion area in 1987 and 1989 were analyzed by Chemex in Sparks, Nevada. Gold was determined by fire-assay fusion with an AA finish. Some samples were analyzed for silver using AA after an aqua regia digestion. In 1988, Teck’s samples from Pinion were analyzed at American Assay Laboratories (“AAL”) in Sparks. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Silver was determined by AA following aqua regia digestion. Some samples were analyzed for gold by fire-assay fusion of 60 g aliquots.


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Chemex and AAL were independent of Teck, but MDA is not aware of certifications held by these laboratories at that time.

Newmont’s 1987 and 1988 drilling samples from the Pinion area, and some of their 1989 Pinion samples, were analyzed at Geochemical Services, Inc. (“GSI”). MDA is not aware of the location(s) of the GSI laboratory. Gold was determined by fire-assay fusion of 30 g aliquots with both gravimetric and AA finish. Samples were also analyzed for silver, arsenic and antimony by ICP. In 1989, Newmont also sent drilling samples from the Pinion area to be analyzed at Bondar-Clegg in Sparks. Following crushing, a subsample was pulverized to -150 mesh. Gold was determined by fire-assay fusion of 30 g aliquots with and AA finish. Silver, arsenic, antimony, molybdenum, and thallium were analyzed by direct-current plasma emission (“DCP”) and mercury was determined by CVAA. Bondar-Clegg and GSI were independent of Newmont, but MDA is not aware of certifications held by these laboratories at that time.

In 1989, Westmont’s drilling samples from the Pinion area were analyzed at Universal in Elko, Nevada. Gold and silver were analyzed by fire-assay fusion, but MDA has no further information on the methods and procedures used. Westmont’s 1991 and 1992 drill samples from the JR Buttes, Jasperoid Wash, and Black Rock areas were analyzed by Cone in Lakewood, Colorado. Gold was determined by fire-assay fusion of 30 g aliquots with a gravimetric finish. Silver, arsenic, antimony, and mercury were determined by AA. Universal and Cone were independent of Westmont, but MDA is not aware of certifications held by these laboratories at that time.

Crown Resources’ samples from their 1991 drilling at Pinion, Dixie, and Dark Star were in part analyzed for gold at AAL in Sparks using fire-assay fusion of 30 g aliquots. Arsenic and antimony were also analyzed, but MDA has no information on the methods and procedures used. Some of the samples from Crown’s drilling at Dark Star in 1991 were analyzed at Activation Laboratories Ltd (“ActLabs”). Composited pulps from prior assays were analyzed for gold, silver and 34 other elements. MDA is not aware of the location of the ActLabs laboratory or the methods and procedures used for the analyses. Samples from Crown’s drilling at the Dark Star and Pinion areas in 1993 were analyzed for gold at AAL in Sparks using fire-assay fusion of 30 g aliquots. AAL and ActLabs were independent of Crown, but MDA is not aware of certifications held by these laboratories at that time.

In 1995, samples from the Cyprus drilling in the Pinion area were analyzed at Chemex in Sparks. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some 1.524 m samples and composited pulps of up to 15.24 m (50 ft) lengths were analyzed for silver, arsenic, antimony, mercury, and barium by AA following digestion in aqua regia. Chemex was independent of Cyprus, but MDA is not aware of certifications held by Chemex at that time.

RSM’s 1996 drill samples from the Pinion area were analyzed at Chemex in Sparks. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Silver was determined by AA following digestion in aqua regia. In 2014, pulps from some of these 1996 RSM Pinion area samples were re-analyzed by ALS in North Vancouver, British Columbia. At ALS, gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Separate aliquots of 0.5 g were analyzed for silver and 34 major, minor and trace elements by ICP following an aqua regia digestion. Portions of remaining drill core from RSM’s 1996 drilling at Pinion were also analyzed at ALS in 2014. These samples were crushed in their entirety to 70% at less than 2 mm. The crushed samples were riffle-split to obtain 250 g subsamples that were pulverized to 85% at less than 75 microns. Gold was determined by 30 g fire-assay fusion with an AA finish. Separate aliquots of 0.5 g were analyzed for silver and 34 major, minor and trace elements by ICP following an aqua regia digestion. Chemex was independent of RSM, but MDA is not aware of certifications held by Chemex at that time.

In 1997, Mirandor’s drilling samples from the Pinion and Dark Star areas were analyzed at Intertek Testing Services (“ITS”) in North Vancouver, British Columbia. At that time, ITS was a division of Bondar-Clegg. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some samples were analyzed for gold by fire-assay fusion of 30 g aliquots with a gravimetric finish. Arsenic, antimony, and barium were determined in some of the samples by AA. Mercury was determined by CVAA. ITS and Bondar-Clegg were independent of Mirandor, but MDA is not aware of certifications held by ITS or Bondar-Clegg at that time.


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Cameco’s 1997 drill samples from the Pinion and Dixie areas were analyzed at Chemex and AAL, both in Sparks. At both laboratories, gold was determined by fire-assay fusion of 30 g aliquots. At Chemex these fire assays were finished with AA. Copies of the AAL assay records do not indicate the type of finish. The samples assayed at AAL were also analyzed for silver and 29 major, minor and trace elements by ICP following aqua regia digestion of 0.5 g aliquots. In 1999, Cameco’s drill samples from the Pinion area were analyzed for gold at AAL by fire-assay fusion of 30 g aliquots. Chemex and AAL were independent of Cameco, but MDA is not aware of certifications held by Chemex or AAL at that time.

In 1998 and 1999, the Kinross drill samples from Dark Star and Pinion were analyzed at Chemex in Sparks. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Composited pulps representing 7.62 m (25 ft) drill intervals were analyzed for 34 major, minor and trace elements by ICP. Chemex was independent of Kinross, but MDA is not aware of certifications held by Chemex at that time.

RSM’s 2003 drill samples from the Pinion area were analyzed by ALS Chemex in North Vancouver, British Columbia. The samples were prepared in the ALS Chemex laboratory in Elko, Nevada, where they were crushed in their entirety to 70% at less than 2 mm. The crushed samples were riffle-split to obtain 250 g subsamples that were pulverized to 85% at less than 75 microns. Gold was determined by 30 g fire-assay fusion with an AA finish. In 2007, RSM’s drill samples from the Pinion area were also analyzed by ALS Chemex. MDA is not aware of how or where these samples were prepared, but silver plus 34 major, minor and trace elements were assayed by ICP following aqua regia digestion of 0.5 g aliquots. Pulps from the 2007 RSM drilling at Pinion were re-analyzed in 2014 at ALS in North Vancouver for gold by 30 g fire-assay fusion with an AA finish.


11.4	GOLD STANDARD - SOUTH RAILROAD PORTION OF THE PROPERTY

MDA has not reviewed and evaluated the methods and procedures used for the collection and analysis of surface samples by Gold Standard as these samples were not used to prepare the mineral resource estimates and mineral reserve estimates presented in later sections of this Technical Report. While useful for purposes of exploration, the surface soil and rock samples of Gold Standard are not material to the interpretations and conclusions of this Technical Report.

Commencing in 2012, Gold Standard’s RC samples stored by the drill rig were collected at the drill sites by representatives of ALS or Bureau Veritas and transported via truck to their respective laboratories in Elko, Nevada. Excessively wet samples were kept at the drill sites for a few days to drain and dry prior to collection by the laboratory staff.

Core samples were transported daily from the drill sites to Gold Standard’s logging and core cutting facility in Elko by Gold Standard personnel. After logging and marking core-sample intervals by Gold Standard geologists, the core was photographed prior to being sawed lengthwise by contractor technicians. Whole HQ-size core was sawed in half. Whole PQ-size core was sawed in quarters. One half of the HQ core, and three quarters of the PQ core, were returned to the core boxes and the remainder was placed in pre-numbered sample bags that were closed with ties. Following insertion of QA/QC blanks and CRM, the core samples were transported by representatives of ALS or Bureau Veritas to their respective laboratories for preparation and analysis.


11.4.1	Pinion Deposit Area Drill Samples

Samples from Gold Standard’s drilling in 2012, 2014, 2015, 2016, and 2017 were analyzed by ALS. The samples were prepared at the ALS laboratory in Elko, Nevada. The samples were dried and crushed in their entirety to 70% at less than 2 mm. The crushed samples were riffle-split to obtain 250 g subsamples that were pulverized to 85% at less than 75 microns. The pulps were shipped via air freight by ALS to the ALS laboratory in North Vancouver, British Columbia, for analysis. Gold was determined by 30 g fire-assay fusion with an AA finish (method code Au-AA23). Samples assayed at ≥10 g Au/t were re-analyzed with a second 30 g aliquot by fire-assay fusion and gravimetric finish (method


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code Au-GRA21). Separate aliquots of 0.5 g were analyzed for silver and 34 major, minor and trace elements by ICP following an aqua regia digestion. In some cases, the ICP analyses were conducted on pulps from 1.524 m drill samples. In other cases, ICP analyses were conducted on composited pulps representing 6.096 m (20 ft) drill intervals. Some samples in 2014 were analyzed for silver by fire-assay fusion of 30 g aliquots with a gravimetric finish. In 2014, some samples were also assayed for 48 major, minor and trace elements by ICP-MS after four-acid digestions. During 2017, samples were analyzed for gold by cyanide leach with an AA finish.

In 2018, Pinion area drill samples were analyzed at Bureau Veritas and AAL. At the Bureau Veritas laboratory in Sparks, Nevada, samples were crushed in their entirety and riffle-split to obtain 250 g subsamples. These subsamples were pulverized to 200-mesh size. Gold was determined by 30 g fire-assay fusion with an AA finish. Some samples were analyzed for gold by cyanide leach with an AA finish. The pulps were shipped to the Bureau Veritas laboratory in Vancouver, British Columbia. Carbon, CO2 and sulfur were determined by induction-furnace infrared absorption and thermal conductivity (“LECO”) analyses of 0.1 g aliquots. Gold, silver and 35 major, minor and trace elements were assayed by ICP following aqua regia digestion of 0.5 g aliquots. Additional silver assays were completed in 2019 at Bureau Veritas using drill-sample pulps from previous analyses. Silver was determined by AA following four-acid digestion of 1.0 g aliquots.

At AAL in Sparks, Nevada, composited pulps of 2018 Pinion area drill samples were analyzed for gold by 30 g fire-assay fusion with an AA finish, and in some cases, with a gravimetric finish. Some of the samples were analyzed for gold by cyanide leach and an AA finish. Gold, silver and 49 major, minor and trace elements were determined in some samples by ICP-MS following digestion in aqua regia.

AAL also analyzed selected, previously assayed drill-sample pulps for elemental barium using an energy-dispersive, x-ray fluorescence (“XRF-ED”) procedure. Pressed-powder pellets made from 2 g aliquots of sample pulps were used for the XRF-ED analyses, which were performed in 2018 and 2019. Other selected sample pulps were analyzed for barium using XRF-ED with 2 g pressed-powder pellets. Some of these were also analyzed for barite using wave-length dispersive x-ray fluorescence (“XRF-WD”) following lithium metaborate fusion of 0.5 g aliquots. Other sample pulps were analyzed for elemental barium by NITON hand-held XRF on both loose-powder aliquots. These were also analyzed by x-ray diffraction (“XRD”) for barite, witherite and calcite, as well as sulfur and carbon by induction-furnace infrared (LECO).

Gold Standard also performed assays of elemental barium together with 39 major, minor and trace elements using hand-held NITON XRF analyzers. These assays were done in 2018 in Elko, Nevada by independent contractor Rangefront Geological using selected drill-sample pulps in loose powder form.


11.4.2	Dark Star Deposit Area Drill Samples

Gold Standard’s 2015 drilling samples from the Dark Star area were mostly analyzed by Bureau Veritas after preparation in the Bureau Veritas laboratory in Elko, Nevada. The samples were crushed in their entirety and riffle-split to obtain 250 g subsample. These subsamples were pulverized to 200-mesh size. Gold was determined by 30 g fire-assay fusion with an AA finish in Bureau Veritas’ laboratory in Sparks, Nevada. Composited pulps were analyzed in Bureau Veritas’ laboratory in Vancouver, British Columbia, for gold, silver and 35 major, minor and trace elements by ICP-MS following aqua regia digestion of 0.5 g aliquots. Some of the 2015 pulps were re-analyzed by ALS in in North Vancouver, British Columbia, for gold by 30 g fire-assay fusion with an AA finish.

The 2016 and 2017 drilling samples from the Dark Star area were analyzed in part by Bureau Veritas and in part by ALS, with sample preparation in their respective laboratories in Elko, Nevada, using the same procedures that were used for the Pinion area samples as summarized in Section 11.4.1. The ALS assays were carried out in their Reno and North Vancouver laboratories where gold was determined by 30 g fire-assay fusion with an AA finish. Samples


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with ≥10 g Au/t were re-analyzed with a second 30 g aliquot by fire-assay fusion and gravimetric finish. Silver and 34 major, minor, and trace elements were assayed by ICP following aqua regia digestion of 0.5 g aliquots.

The Bureau Veritas assays of the 2016 and 2017 Dark Star drilling samples were performed in Bureau Veritas’ laboratories in Sparks, Nevada, and Vancouver, British Columbia. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish and in some cases with a gravimetric finish. Some samples were analyzed for gold by cyanide leach and an AA finish, and some samples were analyzed for gold with a screen-fire assay procedure. Gold, silver, and 35 major, minor, and trace elements were assayed in the Vancouver laboratory by ICP-MS following aqua regia digestion of 0.5 g aliquots.

The 2018 and 2019 drilling samples from the Dark Star area were prepared in either Bureau Veritas’ Elko or Sparks, Nevada, laboratories and analyzed in their Sparks and Vancouver laboratories. Gold and multi-element assays were carried out with the same methods and procedures used for the 2016-2017 samples. In addition, some samples were analyzed for carbon, sulfur and CO2 by LECO methods.


11.4.3	Jasperoid Wash Area Drill Samples

The 2017 drilling samples from the Jasperoid Wash area were analyzed in part by Bureau Veritas and in part by ALS following preparation at their respective laboratories in Elko, Nevada. Gold and multi-element analyses were performed at their respective laboratories in Sparks, Nevada, Vancouver and North Vancouver, British Columbia, using the same methods and procedures used for the 2016-2018 Dark Star samples as summarized in Section 11.4.2.

All of the 2018 drill samples from Jasperoid Wash were prepared and analyzed by Bureau Veritas in Sparks, Nevada and Vancouver, British Columbia, using the same methods and procedures used for the 2016-2019 Dark Star samples as summarized in Section 11.4.2.


11.4.4	Dixie Area Drill Samples

Gold Standard’s 2017 and 2018 drilling samples from the Dixie area were prepared by Bureau Veritas in Sparks, Nevada and Elko, Nevada. Analyses were conducted in the Bureau Veritas Sparks and Vancouver laboratories. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some samples were analyzed for gold by cyanide leach and an AA finish. Gold, silver and 35 major, minor and trace elements were assayed in the Vancouver laboratory by ICP-MS following aqua regia digestion of 0.5 g aliquots. Composited pulps from the 2018 drilling were analyzed for carbon, sulfur and CO2 by LECO methods in the Vancouver laboratory.


11.4.5	Ski Track Area Drill Samples

Most RC samples from Gold Standard’s 2018 drilling at the Ski Track area were prepared by Bureau Veritas in Sparks, Nevada and Elko, Nevada. Analyses were conducted in the Bureau Veritas Sparks and Vancouver laboratories. Gold was determined by fire-assay fusion of 30 g aliquots with an AA finish. Some samples were analyzed for gold by cyanide leach and an AA finish. Gold, silver, and 35 major, minor and trace elements were assayed in the Vancouver laboratory by ICP-MS following aqua regia digestion of 0.5 g aliquots. Composited pulps from the 2018 drilling were analyzed for carbon, sulfur, and CO2 by LECO methods in the Vancouver laboratory.


11.4.6	AUTHOR’S OPINION

The sample collection, security, transportation, preparation, and analytical procedures are judged by the authors to be acceptable and to have produced data suitable for use in the estimation of the mineral resources reported in Section 14, subject to those exclusions or modifications discussed in Section 14. The authors consider the procedures utilized by Gold Standard and the assay laboratories to be appropriate for use as described.


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12	DATA VERIFICATION

Data verification, as defined in NI 43-101, is the process of confirming that data have been generated with proper procedures, have been accurately transcribed from the original sources and are suitable to be used. Additional confirmation of the drill data’s reliability is based on the authors’ evaluations of the Dark Star, Pinion, Jasperoid Wash, and North Bullion area QA/QC procedures and results, as described below, and in general working with the data. No separate evaluations of QA/QC procedures and results were done on data from drilling outside the mineral resource areas.

Prior to MDA’s involvement, as part of the data verification process, APEX visited the Railroad-Pinion property in May 2013, April 2014, and October 2014. Mr. Dufresne conducted several additional site visits from May 31 to June 4, 2015, August 30 to September 2, 2015, and most recently June 7 to 9, 2017. During all site visits, the project geology was reviewed, which included: a) a field tour of the deposit area; b) visual inspection of core holes; and c) discussion with Gold Standard personnel of the current geologic interpretations. Drill site and mineralization verification procedures were conducted, and core drilling and sampling procedures were appraised.

Mr. Dyer and Mr. Ristorcelli visited the Pinion and Dark Star deposit sites on November 18, 2016. This site visit included reviews of core, examination of drill-hole cross sections with the geologic model, and investigations of representative exposures in road cuts and outcrops. Mr. Ristorcelli also visited the Gold Standard office in Elko, Nevada on June 21, 2018. Mr. Lindholm and Mr. Mijal, Senior Geologists with MDA, visited the Dark Star and Jasperoid Wash sites, respectively, on September 18 and 19, 2018. Their work included review of core, checking collar locations, and visiting the site to see the geology.


12.1	DARK STAR AND PINION DATABASE AUDITS

Beginning in March 2018, MDA conducted verification of Gold Standard’s Pinion and Dark Star drilling databases. The databases consisted of Excel spreadsheets, exported from Micromine’s GeoBank secure database software, with collar, survey, assay, and geologic information. MDA decided the best approach to validation was to import the data into a SQL database (GeoSequel) and use the built-in data validation routines. Collar, survey, assay interval, geologic interval, and geologic measurements were imported into GeoSequel directly from the spreadsheet data for both Dark Star and Pinion (see Section 12.2 for Jasperoid Wash and Section 12.3 for North Bullion). The following validation tests were conducted to identify:

Collars: collars with missing depths, collars with missing coordinates, coordinates that might be swapped, drill holes without assay intervals, drill holes without collar survey information, drill holes with nearly duplicate coordinates, drill holes without assays, drill holes without geology, and drill holes with geotechnical information (core holes only);
Surveys: survey depths greater than total depth, survey points missing azimuth or dip values, surveys where azimuth readings were not between 0° and 360°, surveys with flat dip angles (< ~ -45°); and
Assay interval: “excessively” large or small sample intervals, “excessively” large or small geologic intervals, assay intervals that are greater than collar total depth, geologic intervals that are greater than total depth, geotechnical intervals that are greater than total depth, gaps and overlaps in sample intervals, gaps and overlaps in geologic intervals, gaps and overlaps in geotechnical intervals.

Errors found during these tests were iteratively corrected in the database by Gold Standard staff or by MDA with input from Gold Standard.

The next step was to verify the assay data by comparison to the original assay certificates. Because Gold Standard provided electronic copies of certificates for their own drilling, and electronic copies of historical certificates for the pre-Gold Standard holes were incomplete, MDA split the assay validation into two parts. About 58.9% of Pinion drilling


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assays were backed by certificates, and 73.5% of the Dark Star assay data could be tracked back to scanned copies of physical certificates.

The digital data were verified against any physical data that Gold Standard had obtained. Collar information and collar coordinate data were largely validated in their entirety, while the assay data was validated using a representative subset of the data. Collar data were found to be reasonably accurate.

Down-hole survey data from original sources were available for the Gold Standard core holes and some of the historical drill holes, and these too were loaded into GeoSequel for comparison. Eight core holes were evaluated for improbable rates of change of azimuth and dip in down-hole surveys, however, none could be shown to be incorrect and were left in the database.

The first portion of the assay verification was comparing the databases to a random sampling of 10% of the certificate-backed assays. For the Pinion database, thirty certificates with 3,120 sample intervals were randomly selected and checked against the database. The database entries largely compared very well, with only three significant errors, all of which were in the silver values. For the Dark Star database, MDA randomly selected fourteen certificates with 2,391 sample intervals and compared these to the database. These database entries also held up very well, with no significant errors. Insignificant errors for both deposits were found, including rounding the detection limit values to half the detection limit, rounding errors due to conversion of units, and not storing the values in the original reported units.

The second portion of the assay verification involved a random selection of 10% of the drill holes for the two deposits and checking the database entries against all available information in the Gold Standard files. For the Pinion property, this involved 35 drill holes with 2,756 sample intervals. For the Dark Star property, 11 holes with 887 sample intervals were compared. No significant assay errors were found.

For both deposits, MDA found omitted assay values for Ag, As, and other geochemical analyses, and numerous issues with rounding errors. These were not corrected but other, albeit minor, errors with the gold data were corrected with Gold Standard.

In May of 2019, MDA was given a database containing 47,550 silver values for the Pinion project. Using digital certificates supplied by GSV, 24,523 rows of silver data (51.6%) had an error rate of less than 0.01%. Of these errors, all were rounding errors and very minor. These errors were all corrected in the database being used by MDA.

MDA randomly selected a group of certificates to manually spot-audit the remaining rows. Most of these certificates were supplied by Gold Standard as pdf files. Over five percent of these remaining rows were audited with an error rate of 1.1%, of which only a few were important. Most of the errors were in rounding or dropping the detection limit negative sign.

Additional checking was accomplished by working with the data, particularly while working on cross sectional modeling. Suspect data included samples with no gold detected in the middle of intervals of mineralization, assay values where NS (“no sample”) was recorded, and potential down-hole contamination. Issues were resolved most often by considering the samples not reliable and therefore not usable in estimation. While working with the cross sections, MDA also found significant discrepancies between some TCX holes (drilled by Amoco) and surrounding drill holes. The TCX holes were determined to be unreliable and were not used in either domain modeling or estimation.

All the above issues were presented to Gold Standard for corrections in their databases. MDA found that the updated databases contained the corrections and made a few more minor changes.

For non-analytical field data, Gold Standard has instituted protocols to ensure data integrity. For example, during surface geochemical sampling (rock grab and soil sampling), samplers are required to enter sample locations and descriptive information into computers daily and locations are checked to eliminate data input errors. For non-analytical


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drill hole information, Gold Standard employs a similar protocol of continuous data checking to ensure the accurate recording within the project drilling database, which includes all geological and geotechnical information from both core and RC chip logging. The procedures employed are considered reasonable and adequate with respect to insuring data integrity.

During the Dark Star site visit in September 2018, Gold Standard, with MDA present, took GPS measurements on six drill pads and seven drill-hole collars in the field to spot-check coordinates in Gold Standard’s collar tables (see Table 12-1). A Garmin - Rino 530 non-differential GPS was used to measure coordinates at the drill collars. The Garmin website indicates it is accurate to within 3 m to 5 m. Only one easting exceeded the maximum range of accuracy of the GPS by less than a meter; all other readings were within acceptable limits.

Table 12-1 MDA Verification GPS Checks of Dark Star Drill Collars


	MDA GPS Location			Surveyed Location			Difference (GPS - Survey)
Drill Hole	East	North	Elev.	East	North	Elev.	East	North	Elev.
DR18-71	588,001	4,480,319	2,067	588,000.3	4,480,316.1	2,067.7	0.70	2.90	-0.70
DS17-37	588,063	4,480,104	2,049	588,064.9	4,480,106.4	2,048.4	-1.90	-2.40	0.60
DR18-68	588,015	4,479,846	2,015	588,014.9	4,479,844.0	2,012.4	0.10	2.00	2.60
DC18-15	587,920	4,479,606	2,072	587,921.4	4,479,605.3	2,073.2	-1.40	0.70	-1.20
DR18-58	587,783	4,479,476	2,078	587,788.8	4,479,472.6	2,077.3	-5.80	3.40	0.70
DR18-95	587,865	4,479,361	2,108	587,865.8	4,479,360.4	2,110.8	-0.80	0.60	-2.80
DR18-96	587,861	4,479,329	2,110	587,861.4	4,479,328.0	2,111.6	-0.40	1.00	-1.60


12.1.1	2019 Audit of Dark Star Carbon, CO2 and Sulfur Data

Gold Standard provided MDA with assay tables containing 7,081 records of analyses and calculated values for carbon and sulfur in the chemical forms listed in Table 12-2. Most of the analyses were done by Bureau Veritas in Vancouver, British Columbia. A smaller number were done by AAL in Sparks, Nevada. The records verified are summarized in Table 12-2.

Table 12-2 Dark Star Carbon and Sulfur Records Checked and Analytical Procedures


Laboratory	No. of Records	C Total % method	CO2 % method	C InOrganic % method	C Organic % method	S Total % method	S Sulfide % method
Bureau Veritas	7,062	TC003	TC006	calculated	calculated	TC003	TC009
AAL	19	ELTRA C	n/a	calculated	ELTRA C	ELTRA C	n/a


Note:	n/a indicates “not applicable” as in not analyzed and not calculated; TC003 and TC006 are Bureau Veritas method codes for LECO analyses. ELTRA C is AAL method code for LECO-type analyses. On the Bureau Veritas certificates, the TC00x codes on the data listings and cover pages are not the same. The codes listed in the table above are from the cover pages.

MDA compared the measured values in the assay tables from Gold Standard to copies of the laboratory certificates.

Gold Standard’s calculated values were checked using the equations of Gold Standard as follows:

When C Inorganic was not directly assayed

C Inorganic = CO2 Percent / 3.666

Or
C_Inorganic = C Total – C Organic


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C Organic = C Total – C Inorganic

MDA found that all the measured values in the assay tables were the same as those in the laboratory certificates, and all the calculations were done correctly.


12.1.2	2019 Audit of Pinion Carbon, CO2 and Sulfur Data

Gold Standard provided MDA with assay tables containing 4,050 records of analyses and calculated values for carbon and sulfur in the chemical forms listed in Table 12-3. Most of the analyses were done by Bureau Veritas in Vancouver, British Columbia. A smaller number were done by AAL if Sparks, Nevada. The records verified are summarized in Table 12-3.

Table 12-3 Pinion Carbon and Sulfur Records Checked and Analytical Procedures


Laboratory	No. of	C Total %	CO2 %	C InOrganic	C Organic	S Total %	S Sulfide
	Records	method	method	% method	% method	method	% method
Bureau Veritas	3,941	TC003	TC006	calculated	calculated	TC003	TC009
AAL	93	ELTRA C	n/a	calculated	ELTRA C	ELTRA C	ELTRA C
AAL	16	ELTRA C	n/a	calculated	ELTRA C	ELTRA C	ELTRA C

Note: n/a indicates “not applicable” as in not analyzed and not calculated; TC003, TC006 and TC009 are Bureau Veritas method codes for LECO analyses. ELTRA C is AAL method code for LECO-type analyses. On the Bureau Veritas certificates, the TC00x codes on the data listings and cover pages are not the same. The codes listed in the table above are from the cover pages.

MDA compared the measured values in the assay tables from Gold Standard to copies of the laboratory certificates.

Gold Standard’s calculated values were checked using Gold Standard’s equations as follows:

When C Inorganic was not directly assayed
- C Inorganic = CO2 Percent / 3.666
- C Inorganic = C Total – C Organic
C Organic = C Total – C Inorganic

MDA found that all the measured values in the assay tables were the same as those in the laboratory certificates, and all the calculations were done correctly. The only errors that MDA found were 36 assay intervals from hole DS18-07 for which the starting and/or ending depths had been entered incorrectly. MDA corrected these in consultation with Gold Standard.


12..2	JASPEROID WASH DATABASE AUDIT

The drilling database for Jasperoid Wash contained 10,147 assay intervals in 97 drill holes. Gold Standard had good documentation for the 40 of those holes drilled by Gold Standard. Fourteen of the 40 Gold Standard holes did not have assay data associated with them. Historical data for 43 holes came from several sources, mostly written reports, database printouts and previous compilations. Because of the differences in the data, the audit of the Jasperoid Wash assay data involved two parts. In the first part, MDA compared digital certificates supplied by Gold Standard to the existing data compilation using database queries written to compare the two sets. No significant issues were found in the Jasperoid Wash drilling database.

In the second part of the test, MDA compared some of the older drill samples with two annual reports from Westmont, an assay compilation from Cameco, and an assay compilation called PHOLASAY.txt (origin unknown), which contains


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the JW-8910 and JW-8911, and JW-9001 to JW-9014 holes. Only one sample was identified to be a minor typographical error and corrected, and other than minor rounding issues, the verification demonstrated the database properly reflects the underlying data sources.

The drill-hole survey data for the 2017 and 2018 Gold Standard holes were verified against down-hole survey instrument files obtained from Gold Standard. No issues were found between the compiled data set and the source survey data.


12.3	NORTH BULLION DEPOSITS DATABASE AUDIT

This subsection is modified from Dufresne and Nicholls (2017b).


12.3.1	Drill-Hole Collar Locations and Down-Hole Survey Data

Gold Standard’s North Bullion-Bald Mountain database has undergone extensive verification prior to the 2017 mineral resource estimate (Dufresne, 2017b). The confirmation and verification of historical drill hole collar locations was addressed during site visits conducted by Mr. Dufresne. During the site visits, Mr. Dufresne located a number of historical and Gold Standard drill collars using a hand-held GPS, along with tracks representing drill roads and trails. Although unmarked in the field, the identity of several drill collars was ascertained due to their unique location, which were found to be consistent with historically recorded location information. Further work on refining the collar positions and validation of the drill hole database has been performed by Gold Standard personnel and reviewed by Mr. Dufresne.

A total of 170 of the 503 drill holes in the North Bullion-Bald Mountain database have down-hole survey data, including Gold Standard’s 128 holes completed between 2010 and 2017. The database contains 42 holes drilled by Kinross during 1998 and 1999. All available historical documentation was searched with respect to identifying down-hole orientation survey data. No additional down-hole orientation survey data were obtained. Of the 503 drill holes in the database, 297 drill holes were drilled vertically (-85° to -90° dip), with the remaining inclined holes (206 holes) drilled along easterly or westerly azimuths or 040°/ 220° in the case for the POD zone with dips (inclinations) ranging from -30° to -85° from horizontal. A total of 33 of the 333 historical holes without down-hole survey information were drilled to depths greater than 200 m and only 13 of the 33 holes were drilled to greater than 300 m. Based upon the abundance of shallow vertical holes in the historical drill hole database, the amount of down-hole deviation from collar setups for the historical holes is uncertain but is considered likely to be minimal. As such, this has been taken into consideration with the classification of the North Bullion area mineral resources.


12.3.2	Drill-Hole Assay Audit

The assay portion of the North Bullion-Bald Mountain drilling database has also undergone rigorous verification. To date APEX personnel have validated the drill-hole assay database data with all available original assay certificates that could be located. Assay certificates were available for all 135 Gold Standard holes completed between 2010 and 2017. Gold Standard and APEX personnel located the original laboratory assay certificates for 140 of the 368 historical drill holes. The original assay data for all historical drill holes were compared to the Gold Standard database and other historical databases in Gold Standard’s possession that had been compiled by companies such as Newmont, Kinross, and Royal Standard. No assay issues were identified with Gold Standard’s historical assay database. This provided APEX confidence in the validity of the Gold Standard assay database used for the mineral resource estimation presented in Section 14.5.

In order to understand the geological model and place the mineralization in context with the geology, Gold Standard personnel initiated a re-logging program using a) historical drill logs and b) archived RC drilling chips from the historical drill holes. This has resulted in an increased level of confidence in the geological controls of mineralization. Out of the


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503 drill holes in the database, 101 historical holes have been “re-interpreted” by Gold Standard personnel in 2016 and 2017 as of the date of this Technical Report, seven historical holes have been re-interpreted from cross sections by Gold Standard personnel, and 135 holes have been drilled and logged by Gold Standard from 2010 through 2017. Based upon the core drilling and the improved understanding of the geology, the 2010 to 2017 and historical drill holes have been re-logged and re-interpreted.


12.4	GOLD STANDARD QA/QC PROCEDURES

No QA/QC data was available or evaluated for historical drilling programs in the South Railroad portion of the property. The analytical portion of the QA/QC program employed by Gold Standard aimed to provide a means by which the accuracy and precision of the assaying that was performed on the drilling samples (core and RC chip) can be assessed to ensure the highest possible data quality. In order to achieve this goal, Gold Standard personnel inserted samples of CRMs (standards), which are commercially available pulverized materials certified to contain a known concentration of an element (or elements) - in this case gold. The Gold Standard protocol was to use several CRMs of varying gold concentration during a drilling campaign and randomly insert one CRM sample pulp into the stream of actual drill samples at a rate of approximately one in 10. These were alternately inserted with a blank material with gold below detectable limits. The analytical QA/QC measures employed by Gold Standard are sufficient to properly monitor analytical accuracy and precision, and possible in-lab contamination.

CRMs refers to “certified reference materials.” Those used in mineral exploration are usually powders comprised of rock-forming minerals, including the metal of interest in known concentrations. They are analyzed along with batches of samples, and the resulting analyses are evaluated using criteria for passing or failing. CRMs are usually obtained from commercial suppliers. The suppliers provide specifications including the average of many analyses by multiple labs, and the standard deviation of the analyses. In the years 2014 through 2018 Gold Standard has used CRMs obtained from MEG Inc., of Reno, Nevada.

A typical criterion for accepting the analyses of CRMs in the mineral industry is that they should fall within a range determined by the average or expected value ± three standard deviations. Gold Standard uses a stricter criterion, the expected value ± two standard deviations. In the evaluation described here, MDA has used the expected value ± three standard deviations.

Blanks are samples known or thought to contain little or no gold. They are inserted into the sample stream and the results are monitored to be sure that the lab does not report significant gold values when little or no gold should be present.


12.5	DARK STAR DRILL PROGRAM QA/QC

MDA has QA/QC data for the years 1997 and 2015 through 2019, and a very small amount of data for 1991. The types of QA/QC data vary from year to year, but in general there is a substantial suite of QA/QC data available to support the assays used in the Dark Star mineral resource estimate. Table 12-4 summarizes the quantities of each type of data by year.


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Table 12-4: Summary Counts of Dark Star QA/QC Analyses


QA/QC Type	1991	1997	2015	2016	2017	2018	2019
Standard
Number in Use		14	6	5*	5	5	3
Number of Analyses		285	150	708*	310	594	201
Number of Failures		2	1	2	3	3	0

Field Duplicate		56**			322	714	301
Coarse (Preparation) Duplicate		105	58	185
Pulp Duplicate or Replicate		248	59	198
External Check	133		443	1,376	175

Pulp Blank		300	148	1107	170	364	153
Coarse Blank				205	111	158	10


Notes:	* A single analysis of a sixth standard is not included in the counts for 2016.
	** A description of the 1997 duplicates is not available to MDA, so it is only an assumption that they are field duplicates.

The QA/QC data summarized in Table 12-4 are comprised of some QA/QC samples that were part of the project operators’ QA/QC programs, and some that were part of the internal QA/QC protocols of the laboratories that were used.

The QA/QC data available to MDA, including “historical” data inherited from the former project operators of 1997, are adequate to support the use of the Dark Star assay database in a mineral resource estimate. Current QA/QC protocols are adequate to support on-going exploration. MDA has not seen any coarse duplicate data for 2017, 2018, and 2019. Data for coarse duplicates may be available from the laboratories for those years, and if so, MDA suggests that they be acquired and compiled by Gold Standard. Coarse duplicates would be a useful addition to future QA/QC protocols.

During the 2018 and 2019 drilling, Gold Standard has submitted only pulp blanks with samples from RC drilling and only coarse marble blanks with samples of drill core. It would be ideal to submit both types of blanks with both types of samples. If only one type of blank is used, coarse blanks are the best choice.

The following sections contain brief summaries of the QA/QC results by year.


12.5.1	Dark Star Drill Program QA/QC 1991

Very little QA/QC data are available for any holes drilled prior to 1997. However, for 1991 there is a comparison between assays of composited intervals by AAL, which was apparently the original laboratory used, MBA Lab and Actlabs, each using a different analytical method. The composites were made from material drawn from 10 of the 63 holes known to have been drilled that year. AAL and MBA used variations of the atomic absorption analytical method, and their results compare well. Actlabs used the instrumental neutron activation method, a very different analytical method, and obtained results biased significantly high relative to the other two labs (Figure 12-1 and Figure 12-2). Thus, for a small subset of the 1991 drill holes there is some validation of the assay results, based on the AAL vs. MBA comparison. The analyses in the assay table used for estimation are those of AAL.


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Figure 12-1: Dark Star Assay Comparison - AAL vs. MBA - 1991 CDS Holes

Figure 12-2: Dark Star Assay Comparison - AAL vs Actlabs - 1991 CDS Holes


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12.5.2	Dark Star Drill Program QA/QC 1997


12.5.2.1	Standards

The quantity of data available for CRMs in 1997 is ample, although much of the data came from the laboratory’s internal QA/QC. Only two failures were noted, and they are not from holes that are included in the Dark Star mineral resource estimate. Results for CRM analyses are summarized in Table 12-5, and the two failed analyses are detailed in Table 12-6.


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Table 12-5 Summary of Dark Star Results Obtained for Certified Reference Materials, 1997


Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct	Comment
Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct	Comment
C1	n/a	1.931	2.095	1.668	18	29-Aug-97	13-Nov-97	0	0	n/a	target value not known
C2	n/a	1.541	1.830	1.181	29	29-Aug-97	13-Nov-97	0	0	n/a	target value not known
C3	n/a	0.765	0.910	0.663	13	29-Aug-97	07-Nov-97	0	0	n/a	target value not known; too few samples to chart
C4	n/a	0.718	0.767	0.663	2	16-Sep-97	14-Nov-97	0	0	n/a	target value not known
Gannet_192	0.192	0.188	0.262	0.166	49	29-Aug-97	13-Nov-97	1	0	-2.08	target value known, spec. limits not known
Gannet_394	0.394	0.391	0.436	0.358	31	29-Aug-97	07-Nov-97	0	0	-1.27	target value known, spec. limits not known
Gannet_415	0.415	0.456	0.473	0.435	3	14-Nov-97	13-Nov-97	0	0	9.88	target value known, spec. limits not known; too few samples to chart
Gannet_1585	1.585	1.557	1.709	1.365	48	29-Aug-97	13-Nov-97	0	0	-1.70	target value known, spec. limits not known
Gannet_1050	1.05	1.038	1.185	0.944	48	29-Aug-97	13-Nov-97	1	0	-1.14	target value known, spec. limits not known
Gannet_2450	2.45	2.405	2.593	2.263	45	29-Aug-97	13-Nov-97	0	0	-1.24	target value known, spec. limits not known
Gannet_9900	9.9	9.641	10.29	9.06	8	3-Oct-97	14-Nov-97	0	0	-1.84	target value known, spec. limits not known
Gannet_13800	13.8	14.055	14.39	13.72	2	26-Oct-97	9-Nov-97	0	0	1.85	target value known, spec. limits not known; too few samples to chart
BCC_Gold_STD_90-1	6.32	6.693	7.32	5.67	3	25-Sep-97	21-Oct-97	0	0	5.9	target value known, spec. limits not known; too few samples to chart
FA_Synthetic	n/a	1.47	1.47	1.47	1	10-Oct-97	10-Oct-97	0	0	n/a	target value not known; too few samples to chart

Count or Sum	14				285			2	0
Percent					100			0.7	0


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Table 12-6: List of Dark Star Failed Certified Reference Materials, 1997


Standard ID	Drill Hole ID	Target for Std	Fail Type High/Low	Fail Limit	Failed Value	Comment
Gannet_192	EMRR-9714	0.192	High	0.26	0.262	This drill hole is not in MDA’s data set.
Gannet_1050	EMRR-9713	1.05	High	1.158	1.185	This drill hole is not in MDA’s data set.

The main issue with the CRMs is that the available records do not include specifications for the CRMs used by Mirandor the operator. MDA had to interpret how many different CRMs had been used. The ten used by the laboratory Intertek have known expected values, but the expected standard deviations are not known. MDA used statistics derived from the gold assay data set to characterize the expected standard deviations.


12.5.2.2	Field Duplicates

The 1997 assay certificates available to MDA include results for 56 samples with a suffix “D.” It is assumed that these are field duplicates, but specific information is lacking.

Based on relative differences, at grades below about 0.036 g Au/t, the “D” duplicates are on average biased 26% high relative to the original samples. At higher grades, the high bias of the duplicates averages only about 3.8%, within the range of biases that MDA typically finds in such data sets.


12.5.2.3	Preparation Duplicates

The 1997 assay certificates contain results for 105 samples described as “Prep Duplicate”, which MDA takes to mean preparation or coarse crush duplicates. MDA’s evaluation of these revealed no cause for concern.


12.5.2.4	Pulp Duplicates

The 1997 assay certificates contain results for 58 pulp duplicates analyzed using a gravimetric finish and 190 pulp duplicates analyzed using an atomic absorption finish. MDA’s evaluation of these showed the results to be acceptable.


12.5.2.5	Comment on Grade Ranges

Two subsets of gold grade ranges were recognized and evaluated for each of the three duplicate types that were analyzed using an AA finish in 1997, namely the “D” duplicates, preparation duplicates and most of the pulp duplicates. The subsets were selected based on visual inspection of relative difference graphs. Notably, the divisions between lower- and higher-grade subsets chosen in this way all fall in the range 0.036 to 0.040 g Au/t. It appears that the relative precision of the analytical method was substantially better at grades higher than approximately 0.040 g Au/t than it was at lower grades. This is expected, and any likely mineral resource cutoff grades would fall in the higher-grade range in which the analyses are more precise.


12.5.2.6	Mirandor “B” Blanks

In 1997, Mirandor inserted blanks into the sample stream at intervals of approximately 76.2 m (250 ft), for 62 insertions.

MDA does not know the nature of this blank material. The results showed no issues of note or concern.


12.5.2.7	Intertek Analytical Blanks

The Intertek assay certificates from 1997 contain results for 238 analyses of material that Intertek labelled “Analytical Blank.” MDA reviewed these and found no issues of note or concern.


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12.5.3	Dark Star Drill Program QA/QC 2015


12.5.3.1	Dark Star CRMs

Gold Standard used Bureau Veritas as its primary laboratory in 2015, and for the Bureau Veritas assays the quantity of analyses of CRMs is ample, with only one failure recorded. The single failure is not material to the mineral resource estimate. Results for CRM analyses are summarized in Table 12-7, and the failed analysis is detailed in Table 12-8.

Table 12-7: Summary of Dark Star Results Obtained for Certified Reference Materials, 2015


Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct	Comment
Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct	Comment
MEG-Au.10.02	0.035	0.034	0.054	0.024	31	Jun-15	Nov-15	1	0	-2.86
MEG-Au.10.04	0.078	0.078	0.095	0.063	24	Jun-15	Nov-15	0	0	0
MEG-Au.11.29	3.689	3.702	4.166	3.478	16	Jun-15	Nov-15	0	0	0.35
MEG-Au.13.02	0.746	0.753	0.811	0.697	31	Jun-15	Nov-15	0	0	0.94
MEG-S107007X	1.526	1.560	1.665	1.475	27	Jun-15	Nov-15	0	0	2.23
MEG-Au.11.17	2.693	2.723	3.033	2.510	21	Jun-15	Oct-15	0	0	1.11

Count or Sum	6				150			1	0
Percent					100			0.67	0

Table 12-8: List of Dark Star Failed Certified Reference Materials, 2015


Standard ID	Drill Hole ID	Target for Std	Fail Type High/Low	Fail Limit	Failed Value	Comment
MEG-Au.10.02	EMRR-9714	0.035	High	0.047	0.054

In 2017, “A comprehensive assay check (umpire) program” was completed by ALS on original sample pulps from the Gold Standard’s 2015 and 2016 drilling at the Dark Star deposit which had reported values at or above the 0.14 g Au/t cut-off grade…” On reviewing the results, Gold Standard elected to use the assays from ALS for the 2015 and 2016 samples.

Consequently, while most of the assays in MDA’s database for the 2015 drill holes are the original Bureau Veritas assays, the majority of the assays at or above 0.14 g Au/t are those from ALS. There are 376 such assays, out of a total of 3,426 assays from the 2015 drill holes. MDA’s review of standards for 2015 applies only to the Bureau Veritas assays. MDA has no QA/QC data for the assays of 2015 samples from ALS, which is to say, no QA/QC data applying to most of the mineral resource-grade samples from 2015.


12.5.3.2	Bureau Veritas Duplicates

One of two Excel files provided to MDA, containing QA/QC data for 2015, contains a compilation of analytical results for Bureau Veritas’ internal-preparation duplicates and Bureau Veritas’ pulp duplicates, relating to holes DS15-06 through DS15-12. MDA does not have any duplicate data for other holes drilled in 2015.

MDA evaluated the results for these two types of duplicates and found them to be generally as expected. One somewhat surprising finding is that the within-laboratory pulp duplicates or replicates have, on average, a stronger than expected negative bias with respect to the original analyses, based on the relative differences. Pulp duplicates at grades exceeding 0.04 g Au/t have an average difference, relative to the originals, of -5.5%.


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12.5.3.3	ALS vs. Bureau Veritas Checks

In 2017, Gold Standard obtained re-analyses of pulps from the 2015 samples at ALS, for comparison with the original Bureau Veritas assays. MDA evaluated these as check assays, as shown in Figure 12-3. They amount to 443 pairs. ALS’ analyses are biased on average about 4.7% high relative to Bureau Veritas’.

Figure 12-3: Dark Star Check Assays – ALS Assay vs. Bureau Veritas (Inspectorate), 2015


12.5.3.4	Blanks

In 2015, Gold Standard used pulp blanks obtained from a vendor of standard reference materials. No issues of note or concern were revealed by the 148 analyses of this blank.


12.5.3.5	Assay Substitution

The QA/QC data available for 2015 support the original assays for that year by Bureau Veritas. Subsequent to 2015, later check assays by ALS were substituted for some of the original Bureau Veritas assays. MDA has little or no QA/QC data that directly relates to the ALS assays, but they compare well enough to the Bureau Veritas assays, albeit with a high bias in the order of 4%, to be accepted.


12.5.4	Dark Star Drill Program QA/QC 2016

12.5.4.1	CRMs

The number of CRMs in use and the number of analyses from 2016 are ample. Two failures occurred but are not of material concern with respect to the mineral resource estimate. Results for CRM analyses are summarized in Table 12-9, and the two failed analyses are detailed in Table 12-10.


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Table 12-9: Summary of Dark Star Results Obtained for Certified Reference Materials, 2016


Laboratory	Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct
Laboratory	Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
Inspectorate	MEG-Au.10.02	0.035	0.035	0.041	0.028	156			0	0	0
ALS	MEG-Au.10.02	0.035	0.036	0.041	0.032	11			0	0	2.86
Inspectorate	MEG-Au.10.04	0.078	0.078	0.092	0.057	142			0	2	0
ALS	MEG-Au.10.04	0.078	0.082	0.084	0.079	8			0	0	5.13
Inspectorate	MEG-Au.13.02	0.746	0.749	0.804	0.7	141			0	0	0.40
ALS	MEG-Au.13.02	0.746	0.759	0.768	0.74	6			0	0	1.74
Inspectorate	MEG-S107007X	1.526	1.557	1.691	1.378	114			0	0	2.03
ALS	MEG-S107007X	1.526	1.503	1.55	1.475	7			0	0	-1.51
Inspectorate	MEG-Au.11.17	2.693	2.773	3.017	2.543	110			0	0	2.97
ALS	MEG-Au.11.17	2.693	2.824	2.93	2.71	13			0	0	4.86
Inspectorate	MEG-Au.11.29	3.689	4.135	4.135	4.135	1			0	0	12.09
ALS	MEG-Au.11.29	3.689	n/a	n/a	n/a	0			0	0	n/a

Inspectorate						664
ALS						45

Count or Sum	6					263			0	2
Percent						100			0	0.28

Table 12-10: List of Dark Star Failed Certified Reference Materials, 2016


Standard ID	Laboratory	Drill Hole ID	Target for Std	Fail Type High/Low	Fail Limit	Failed Value	Comment
MEG-Au.10.04	Inspectorate	DS16-08 651A	0.078	low	0.06	0.058
MEG-Au.10.04	Inspectorate	DS16-38 1650A	0.078	low	0.06	0.057

In addition to the analyses of CRMs by Bureau Veritas, some analyses of the CRMs by ALS are available in the 2016 data set. While the ALS analyses of CRMs are few relative the number of Bureau Veritas analyses, and thus statistics for the ALS assays are less robust than those for the Bureau Veritas assays, it is noteworthy that ALS’ biases relative to the expected values for four of five CRMs were higher than Bureau Veritas’. Bureau Veritas’ assays were, on average, closer to the expected values for the CRMs. Note that in 2016, some ALS check assays have been substituted for the original Bureau Veritas assays, and for these MDA has no data for CRMs.


12.5.4.2	Bureau Veritas Duplicates

MDA was provided with a compilation of Bureau Veritas’ internal preparation duplicate and replicate data, comprised of 185 preparation duplicate pairs and 198 pulp duplicate or replicate pairs. MDA’s evaluation of these revealed no issues of concern.


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12.5.4.3	ALS vs. Bureau Veritas Checks on Pulps

In 2017, Gold Standard obtained re-analyses of the 2016 samples at ALS, for comparison with the original Bureau Veritas assays. MDA evaluated these as check assays, as shown in Figure 12-4. They amount to 1,376 pairs. ALS’ analyses are biased on average about 3.8% high relative to Bureau Veritas.

Figure 12-4: Dark Star Check Assays - ALS Assay vs. Bureau Veritas (Inspectorate) 2016


12.5.4.4	Gold Standard Pulp Blanks

In 2016, Gold Standard used a commercial pulp blank obtained from a vendor of CRMs, obtaining 572 analyses. No issues were revealed by these analyses.


12.5.4.5	Bureau Veritas Pulp Blanks

The data package for 2016 contains 535 analyses of a pulp blank used by Bureau Veritas as part of their internal QA/QC program. MDA evaluated these and found no issues of note or concern.


12.5.4.6	Bureau Veritas Coarse Blanks

The data available to MDA include a compilation of results from 205 coarse blanks analyzed by Bureau Veritas as part of their internal QA/QC protocol. MDA has no information as to where in the sequence of analyses these blanks were placed, so the data are not useful for checking on the possibility of contamination in the sample preparation process. Viewed without sequential context, the analyses revealed no issues of concern.


12.5.4.7	Assay Substitution

The QA/QC data available for 2016 support the original assays for that year by Bureau Veritas. Subsequent to 2015, later check assays by ALS were substituted for some of the original Bureau Veritas assays. MDA has only a small


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amount of QA/QC data that directly relates to the ALS assays, but they compare well enough to the Bureau Veritas assays, albeit with a high bias in the order of 4%, to be accepted.


12.5.5	Dark Star Drill Program QA/QC 2017

12.5.5.1	CRMs

The number of CRMs in use and the number of analyses from 2017 are sufficient. Of 310 analyses of CRMs, 180 were done at ALS and 130 were done at Bureau Veritas. Both ALS and Bureau Veritas analyses of the lowest-grade CRM, which has an expected value of 0.078 g Au/t, were biased high by more than 6%, and three of ALS’s analyses were high-side failures. Because the expected value and even the highest-grade of the failures are below likely mineral resource cutoff grades, the failures and the high biases have little if any impact on confidence in the mineral resource estimate. Results for CRM analyses are summarized in Table 12-11, and the three failed analyses are detailed in Table 12-12.

Table 12-11: Summary of Dark Star Results Obtained for Certified Reference Materials, 2017


Laboratory	Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct
Laboratory	Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
ALS	MEG-Au.10.04	0.078	0.083	0.099	0.073	135	16-Jul-17	27-Oct-17	3	0	6.41
Inspectorate	MEG-Au.10.04	0.078	0.084	0.095	0.071	64	8-Aug-17	2-Nov-17	0	0	7.69
ALS	MEG-Au.13.02	0.746	0.751	0.763	0.742	8	2-Aug-17	10-Jan-18	0	0	0.67
Inspectorate	MEG-Au.13.02	0.746	0.749	0.768	0.723	6	12-Jan-18	27-Feb-18	0	0	0.4
ALS	MEG-Au.12.11	1.465	1.518	1.53	1.495	3	30-Jul-17	30-Jul-17	0	0	3.62
Inspectorate	MEG-Au.12.11	1.465	1.469	1.544	1.349	11	19-Jan-18	27-Feb-18	0	0	0.27
ALS	MEG-Au.12.21	0.143	0.138	0.147	0.127	34	30-Dec-17	15-Jan-18	0	0	-3.5
Inspectorate	MEG-Au.12.21	0.143	0.139	0.155	0.121	44	27-Nov-17	27-Feb-18	0	0	-2.8
Inspectorate	MEG-Au.11.19	0.12	0.120	0.123	0.112	5	27-Feb-18	27-Feb-18	0	0	0
Totals or Averages
ALS	4					180			3	0	1.80
Inspectorate	5					130			0	0	1.11

All	9					310			3	0
Percent						100			0.97	0

Table 12-12: List of Dark Star Failed Certified Reference Materials, 2017


Standard ID	Laboratory	Sample ID	Target for Std	Fail Type High/Low	Fail Limit	Failed Value	Comment
MEG-Au.10.04	ALS	DS17-15 2045-2050-A2	0.078	high	0.096	0.099	no follow-up
MEG-Au.10.04	ALS	DS17-15 1045-1050-A2	0.078	high	0.096	0.096	no follow-up
MEG-Au.10.04	ALS	DS17-15 1245-1250-A2	0.078	high	0.096	0.098	no follow-up

For three of four CRMs that have data allowing comparisons between results from Bureau Veritas and results from ALS, Bureau Veritas’ assays are closer, on average, to the expected values than ALS’. This differs from the comparison for the 2016 analyses of CRMs when ALS was biased low compared to Bureau Veritas.


12.5.5.2	Gold Standard Duplicates

The 322 field duplicates available in the 2017 data set reveal no issues of note or concern.


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12.5.5.3	External Checks of 2017 Assays in 2018

In April and August 2018, Gold Standard selected pulps from three holes drilled in 2017 and sent them to a lab other than the original for check assays. One hole was originally assayed by ALS and was sent to Bureau Veritas (a.k.a. Inspectorate) for check assays. The other two holes were originally assayed by Bureau Veritas and were sent to ALS for checks. In total, 175 usable original/check assay pairs were produced. MDA evaluated the results as a comparison between ALS and Bureau Veritas.

At grades up to about 3 g Au/t, the two labs’ results compare well. Between about 3 g Au/t and 10 g Au/t, which is the upper limit for ALS’ Au-AA23 analytical method, ALS is biased about 6.8% low relative to Bureau Veritas, based on relative differences of 14 pairs of analyses. Conversely, at grades above 10 g Au/t, for which both labs used a gravimetric finish, ALS is biased on average about 4.6% high relative to Bureau Veritas, based on relative differences of only five pairs of analyses. With such small sample populations, no important conclusions can be drawn based on these results.


12.5.5.4	Gold Standard Pulp Blanks

In 2017, Gold Standard inserted 170 pulp blanks, obtained from a supplier of CRMs, into the sample stream. The analyses of these revealed no important issues.


12.5.5.5	Gold Standard Coarse Blanks

Gold Standard inserted 111 samples of a coarse marble blank into the sample stream in 2017. In three holes analyzed on October and November 2017, there is a significant correlation between analyses of blanks that reported detectable gold, and high gold values in adjacent real samples. The highest gold assay in a marble blanks is 0.021 g Au/t; not high enough to cause concern about using related assays of real samples in a mineral resource estimate. Such correlations between low levels of detectable gold in coarse blanks and adjacent samples containing high gold grades are common and are not necessarily a problem but must be monitored.


12.5.6	Dark Star Drill Program QA/QC 2018

12.5.6.1	CRMs

The number of CRMs used, and the number of analyses is ample. Three total failures occurred, two on the high-side, and one on the low-side. The latter may be due to a sample mix-up rather than an analytical issue. Results for CRM analyses are summarized in Table 12-13, and the three failed analyses are detailed in Table 12-14.

Table 12-13: Summary of Dark Star Results Obtained for Certified Reference Materials, 2018


Laboratory	Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct
Laboratory	Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
Inspectorate	MEG-Au.11.19	0.12	0.115	0.139	0.091	76	14-Mar-18	16-Apr-18	0	0	-4.17
AAL	MEG-Au.11.19	0.12	0.099	0.105	0.003	16	24-Apr-18	30-Apr-18	0	1	-17.5
Inspectorate	MEG-Au.17.06	0.098	0.102	0.123	0.082	376	26-Jun-18	14-Feb-19	1	0	4.08
Inspectorate	MEG-Au.13.02	0.746	0.750	0.788	0.727	10	17-Aug-18	06-Sep-18	0	0	0.54
Inspectorate	MEG-Au.12.11	1.465	1.486	1.595	1.358	66	17-Aug-18	14-Feb-19	0	0	1.43


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Inspectorate	MEG-Au.17.07	0.188	0.201	0.225	0.184	50	06-Sep-18	14-Feb-19	1	0	6.91
Totals or Averages
Inspectorate	5					578			2	0	0.47
AAL	1					16			0	1	-17.5

All	6					594			2	1
Percent						100			0.34	0.17

Table 12-14: List of Dark Star Failed Certified Reference Materials, 2018


Standard ID	Lab	Sample ID	Target for Std	Fail Type	Fail Limit	Failed Value	Comment
MEG-Au.11.19	AAL	DR18-25 545-550 A9	0.120	low	0.081	<0.005	blank?
MEG-Au.17.06	Insp.	DS18-02 1845-1850-L1	0.098	high	0.119	0.123	insufficient sample
MEG-Au.17.07	Insp.	DC18-04 490-495-A12	0.188	high	0.221	0.225	deemed OK by GSV*

Note: * This failure was accepted because it is in an unmineralized geotechnical drill hole, not material to the block model.

Most of the analyses in 2018 were done by Bureau Veritas, but there are sixteen analyses by AAL of a CRM having an expected value of 0.12 g Au/t. The average of AAL’s analyses of this CRM are biased 17.5% low. This magnitude of bias is unusual. Only one of the AAL analyses is a failure, and as mentioned in the preceding paragraph, that may not be an analytical failure, but such a low bias merits investigation.

Two high-side failures occurred in Inspectorate’s analyses of CRMs in 2018. One of these involved a case of insufficient sample material. Both failed analyses were 0.004 g Au/t above the upper failure limit. Gold Standard did not take any corrective action in either case.


12.5.6.2	Gold Standard Duplicates

The 714 field duplicates available in the 2018 data set reveal no issues of note or concern.


12.5.6.3	Gold Standard Pulp Blanks

During the 2018 drill program, Gold Standard inserted pulp blanks into the stream of samples from RC drilling. Pulp blanks have not been used with the samples from core drilling.

Most of the 364 analyses of pulp blanks have reported acceptable results. Gold Standard flagged one pulp blank, with a gold analysis of 0.028 g Au/t, for re-analysis of part of the sample batch. Successful re-analyses were received in September 2018.


12.5.6.4	Gold Standard Coarse Blanks

The 158 analyses of coarse marble blanks in 2018 show that during the mid-July to mid-October period, there was a correlation between higher-grade samples and instances of detectable gold in the marble blank, similar to that which was noted in the 2017 samples. The same conclusion applied in 2018 as in 2017; if the detectable gold in the marble blank was due to contamination from adjacent gold-bearing samples, the degree of contamination was not so high as to cause concern about using the assays in a mineral resource estimate, but this issue should be closely monitored.


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12.5.6.5	Twin-Hole Analysis

Gold Standard drilled one core hole twin (DC18-09) of a RC drill hole (DR18-44). The holes were collared 5.7 m apart and intersected a significant amount of low- and high-grade mineralization. Core intervals were composited to 3.05 m to match the RC intervals to facilitate a more direct comparison of the data. Viewed on-screen, the higher-grade intercepts are wider in the core hole, although the relative positions of mineralization are similar in the two holes. Average gold grade is higher in the core hole at 1.95 g Au/t, compared to 1.50 g Au/t in the RC hole (Figure 12-5). The core hole roughly confirms the data in the RC hole, but general conclusions about the rest of the drilling from this one pair of holes cannot be made. The single twin-hole comparison also suggests that grades of mineralization can vary rapidly over short distances within the Dark Star deposit.

Figure 12-5: Scatter Plot of Twin-Hole Analysis – DC18-09 (core) vs DR18-44 (RC)


12.5.7	Dark Star Drill Program QA/QC 2019

12.5.7.1	CRMs

The number of CRMs used and the number of analyses during the first quarter of 2019 were ample. No failures occurred. Results for one CRM were on average biased close to 6% above the expected value. The same standard had a similar bias in 2018. While the results for the CRM are within acceptable limits, MDA suggests doing some test analyses of the CRM at a different laboratory, to see if a similar bias is obtained. It may be that the characterization of the standard should be modified. Results for CRM analyses are summarized in Table 12-15.


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Table 12-15: Summary of Dark Star Results Obtained for Standards, 2019


Laboratory	Standard ID	Grades in ppm Au				Count	Dates Used		Failure Counts		Bias pct
Laboratory	Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
Inspectorate	MEG-Au.17.06	0.098	0.100	0.113	0.090	92	08-Feb-19	24-Apr-19	0	0	2.04
Inspectorate	MEG-Au.12.11	1.465	1.496	1.578	1.385	66	08-Feb-19	24-Apr-19	0	0	2.12
Inspectorate	MEG-Au.17.07	0.188	0.199	0.215	0.181	43	08-Feb-19	24-Apr-19	0	0	5.85
Totals or Averages
All	3					201			0	0
Percent						100			0	0


12.5.7.2	Duplicates

The field duplicates collected in 2019 returned results within the normal range that MDA expects for such field duplicates. The results were acceptable.


12.5.7.3	2019 Pulp Blanks

The results for the pulp blanks analysed during the first quarter of 2019 were acceptable, revealing no causes for concern.


12.5.7.4	2019 Coarse Blanks

Results are available for 10 analyses of coarse blanks analysed in a shipment of samples from one core hole. The results were acceptable.


12.6	GOLD STANDARD’S PINION DRILL PROGRAM QA/QC

The lack of QA/QC before 2014 has impacted the mineral resource classification as described in Section 14.3.

During the period 2014 through 2016, Gold Standard’s QA/QC program involved the use of pulp blanks and CRMs. No coarse blanks or duplicates were used during those years. In 2017 and 2018, Gold Standard’s QA/QC program was the same as in 2014-2016, but with the addition of coarse blanks and RC rig (field) duplicates. MDA’s evaluation of Gold Standard’s QA/QC data revealed these issues:

In the latter half of 2014, six of 52 analyses of one CRM, having a target grade of close to 2 g Au/t, were high- side failures for a high-side failure rate of about 12%;
In April 2018, seven sequential analyses of a CRM, with a target grade of 0.12 g Au/t, were biased low by an average of about 18.5%;
Among samples from hole PIN15-14, ten of the blanks assayed gold in the range 0.032 to 0.083 g Au/t. Gold Standard obtained re-analyses of 14 mineralized samples analyzed in the same batch as the blanks in question. In a 16.76 m interval, the re-run assays averaged 0.35 g Au/t, whereas the original assays averaged 0.41 g Au/t. The rerun assays are used in the project database;
There are some indications in the data that occasional sample mix-ups of QA/QC samples have occurred. This is, however, unprovable; and
The data provided to MDA, particularly prior to 2016, do not contain consistent records of actions that may have been taken to investigate QA/QC failures. The example of PIN15-14 shows that Gold Standard does


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monitor and take action on some QA/QC failures, but records of these activities were not readily accessible in the data sets provided to MDA.

The issues that MDA has identified are not sufficient to preclude the use of Gold Standard’s gold assays in a mineral resource estimate. Their effect on the estimate would not be material. However, the lack of QA/QC from before 2014, which still is the majority of drilling, and the minimal QA/QC samples prior to 2016 are considered in mineral resource classification.

MDA strongly suggests that, in the future, Gold Standard’s QA/QC program should include the use of coarse (preparation) duplicates, to monitor the consistency of the laboratory’s sample preparation procedures and possible contamination during preparation.


12.6.1	Pinion Drill Program QA/QC CRMs

12.6.1.1	CRMs 2014 - 2015

For drilling during 2014 and 2015, Gold Standard supplied MDA with the analyses of CRMs but did not have evaluations available. MDA did the evaluations, putting the data for 2014 and 2015 together to make a single two-year time period, and preparing control charts for each of the eight CRMs used in those two years. The results are summarized in Table 12-16. Details of the 11 failures are listed in Table 12-17.

Table 12-16: Summary of Results Obtained for CRMs, 2014 – 2015


CRM ID		Grades in g Au/t			Count	Dates Used		Failure Counts		Bias pct
CRM ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
MEG-Au.10.02	0.035	0.034	0.079	0.022	180	Apr 2014	Dec 2015	1	1	-2.86
MEG-Au.10.04	0.078	0.079	0.089	0.064	123	Apr 2014	Dec 2015	0	0	1.28
MEG-Au.11.19	0.120	0.117	0.137	0.096	20	Apr 2014	July 2014	0	0	-2.50
MEG-Au.11.29	3.689	3.730	4.470	3.170	86	Apr 2014	Dec 2015	0	0	1.11
MEG-Au.13.02	0.746	0.759	0.825	0.679	164	July 2014	Dec 2015	0	0	1.74
MEG-S107007X	1.526	1.535	1.660	1.185	112	Apr 2014	Dec 2015	0	3	0.59
MEG-Au.11.34	2.114	2.300	5.410	1.840	52	July 2014	Dec 2014	6	0	-0.95
MEG-Au.11.17	2.693	2.798	2.990	2.630	36	June 2015	Dec 2015	0	0	3.90

Sum					773			7	4
Percent					100			0.91	0.52

Table 12-17: List of Failed CRM Analyses, 2014 – 2015


CRM ID	Sample ID	Cert. Grade g Au/t	Fail Type High/Low	Fail Limit g Au/t	Failed Value g Au/t	Comment
CRM ID	Sample ID	Cert. Grade g Au/t	Fail Type High/Low	Fail Limit g Au/t	Failed Value g Au/t	Comment
MEG-Au.10.02	PIN14-20 650A	0.035	Low	0.023	0.022
MEG-Au.10.02	PIN14-44 100B	0.035	High	0.047	0.079	mis-identification?
MEG-S107007X	PIN14-06 302A	1.526	Low	1.322	1.310
MEG-S107007X	PIN14-09 350A	1.526	Low	1.322	1.310
MEG-S107007X	PIN14-11 150A	1.526	Low	1.322	1.185
MEG-Au.11.34	PIN14-11 450A	2.114	High	2.630	2.850


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CRM ID	Sample ID	Cert. Grade g Au/t	Fail Type High/Low	Fail Limit g Au/t	Failed Value g Au/t	Comment
CRM ID	Sample ID	Cert. Grade g Au/t	Fail Type High/Low	Fail Limit g Au/t	Failed Value g Au/t	Comment
MEG-Au.11.34	PIN14-18 250A	2.114	High	2.630	3.310
MEG-Au.11.34	PIN14-34 250A	2.114	High	2.630	2.780
MEG-Au.11.34	PIN14-38 250A	2.114	High	2.630	5.410	mis-identification?
MEG-Au.11.34	PIN14-52 250A	2.114	High	2.630	5.340	mis-identification?
MEG-Au.11.34	PIN14-56 250A	2.114	High	2.630	3.590

Multiple high-side failures were observed with one CRM, MEG-Au.11.34, for which six analyses were failures, amounting to almost 12% of the analyses. This is a high failure rate. The control chart for this CRM is shown in Figure 12-6. MDA suspects, but cannot prove, that some of the very high failures listed in Table 12-17 and illustrated in Figure 12-6 reflect sample numbering mix-ups rather than analytical failures.

Another control chart, for MEG-S107007X, is shown in Figure 12-7. It is included to point out the first 10 analyses, highlighted in red on the chart. They are biased consistently low, and three are low enough that they qualify as analytical failures, as listed in Table 12-17. MDA suspects, but again cannot prove, that the CRM analyzed in those 10 cases has been mis-identified. An alternative interpretation is that for a period in 2014, the laboratory had a significantly low analytical bias. However, such a consistent low bias wasn’t identified in other CRMs analyzed during the same time period.

Regardless of the explanations for the failures described above, whether they are due to analytical issues, mis-identifications or other unidentified causes, they were failures.

Figure 12-6: Control Chart for MEG-Au.11.34

Note: data points shown as hollow squares were not used in calculating the average and bias listed in Table 12-16.


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Figure 12-7: Control Chart for MEG-S107007X


Explanations for Figure 12-6 and Figure 12-7
Items Obtained from Certificate for CRM
USL	Upper Specification Limit	Target + 3 Std Dev
Target	Expected Value
LSL	Lower Specification Limit	Target - 3 Std Dev
Items Calculated using Gold Standard Data
UCL	Upper Control Limit	Avg + 3 Std Dev
Avg	Mean Value
LCL	Lower Control Limit	Avg - 3 Std Dev


12.6.1.2	CRMs Used in 2016, 2017 and 2018

Gold Standard has provided MDA with evaluations containing sets of control charts for the CRMs that were used during the 2016, 2017, and 2018 drilling campaigns. The results obtained are summarized in Table 12-18, Table 12-19, and Table 12-20.

No failures were identified in the data for CRMs in 2016 and 2017. Three high-side failures occurred in 2018, as indicated in Table 12-20 and listed in Table 12-21. One of those is exactly at the failure limit and is accepted by Gold Standard and MDA. The other two high-side failures do not relate to any samples that influence the mineral resource, as indicated in the “Comment” column of Table 12-21.

The biases listed in Table 12-18 and Table 12-19 are within the range that MDA typically finds. There is, however, an issue with bias for CRM MEG-Au.11.19 in 2018, shown in Table 12-20. Two sets of data for this CRM are shown in the table, because the first 13 analyses of MEG-Au.11.19 are on average biased 2.3% low, within the range of biases that MDA expects. However, the seven latest analyses of MEG-Au.11.19 are biased 18.5% low, well outside the expected range. This is illustrated on Figure 12-8.

MDA cannot explain the very low bias in the seven most recent analyses of MEG-Au.11.19. Possible explanations include a change in the character of the CRM itself, or a change in some aspect of the laboratory’s analytical process. While none of the seven low-biased analyses is technically a failure using the usual criteria of expected value ± three


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standard deviations, a low bias of this magnitude could indicate a systematic low analytical bias that could potentially have had a negative effect on the mineral resource estimate.

The period during which the low bias in analyses of MEG-Au.11.19 occurred was April 10 through April 29, 2018. Another CRM having a gold grade with a similar order of magnitude, MEG-Au.17.06, was analyzed during the period April 11 through June 13, 2018. Analyses of MEG-Au.17.06 have a slight positive bias. This suggests that the laboratory was not producing results with a consistent strong low bias in the grade range near 0.1 g Au/t.


12.6.1.3	Grade Ranges of CRMs Used in 2018

Although the number of CRMs in use has varied over the years, generally with fewer CRMs in use as time progressed, Gold Standard has typically used one or more low-grade CRMs, one or more mid-grade CRMs, and one or more high-grade CRMs, relative to the grades of mineralized samples encountered at Pinion. In 2018, except for a period in mid-April, Gold Standard used only one of two low-grade CRMs, both having grades below the likely mineral resource cutoff. In mid-April of 2018, Gold Standard also used a high-grade CRM. Since April 20, 2018 the only CRM in use at Pinion has a certified grade below the likely mineral resource cutoff. This is illustrated in Figure 12-9.

The low-grade CRMs are useful in that they test the analytical method used for most of the mineral resource-grade samples. However, a single analytical method can yield different accuracies and precisions over its useful range. Using at least one each of mid-grade and high-grade CRMs, along with the low-grade CRM, would yield greater confidence in the related assays.

Table 12-18: Summary of Results Pinion for CRMs, 2016


CRM ID	Grades in g Au/t				Count	Dates Used		Failure Counts		Bias pct
CRM ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
MEG-Au.10.02	0.035	0.034	0.039	0.029	73	June 2016	Jan 2017	0	0	-2.86
MEG-Au.10.04	0.078	0.080	0.090	0.070	68	June 2016	Jan 2017	0	0	2.32
MEG-Au.11.29	3.689	3.708	4.120	3.520	15	June 2016	July 2016	0	0	0.52
MEG-Au.13.02	0.746	0.759	0.782	0.731	25	June 2016	Jan 2017	0	0	1.73
MEG-S107007X	1.526	1.526	1.630	1.420	39	June 2016	Oct 2016	0	0	0
MEG-Au.11.17	2.693	2.786	3.000	2.430	43	June 2016	Jan 2017	0	0	3.44

Sum					263			0	0
Percent					100			0	0

Table 12-19: Summary of Results for Pinion CRMs, 2017


CRM ID	Grades in g Au/t				Count	Dates Used		Failure Counts		Bias pct
CRM ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
MEG-Au.10.02	0.035	0.034	0.040	0.008	11	Nov 2017	Nov 2017	0	1	-2.86
MEG-Au.10.04	0.078	0.081	0.085	0.076	8	Nov 2017	Nov 2017	0	0	3.85
MEG-Au.12.21	0.143	0.136	0.151	0.122	31	Nov 2017	Dec 2017	0	0	-4.90

Sum					50			0	1
Percent					100			0	2


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Table 12-20: Summary of Results for Pinion CRMs, 2018


CRM ID	Grades in g Au/t				Count	Dates Used		Failure Counts		Bias pct
CRM ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
MEG-Au.17.06	0.098	0.100	0.126	0.084	258	11 Apr 2018	6 Jul 2018	3	0	2.32
MEG-Au.11.19	0.120	0.117	0.138	0.096	13	29 Mar 2018	9 Apr 2018	0	0	-2.31
MEG-Au.11.19	0.120	0.098	0.102	0.091	7	10 April 2018	29 Apr 2018	0	0	-18.45
MEG-Au.11.29	3.689	3.849	4.348	3.392	16	5 April 2018	20 Apr 2018	0	0	5.41

Sum					115			0	0
Percent					100			0	0

Table 12-21: List of Failed Pinion CRMs, 2018


CRM ID	Sample ID	Target for Std g Au/t	Fail Type High/Low	Fail Limit g Au/t	Failed Value	Comment
MEG-Au.17.06	PR18-78 245-250-L1	0.098	High	0.119	0.119	accepted by Gold Standard
MEG-Au.17.06	PR18-89 45-50-L1	0.098	High	0.119	0.126	accepted by Gold Standard; does not affect any mineral resource blocks
MEG-Au.17.06	PC18-03 32-34.5-L1	0.098	High	0.119	0.122	samples re-run by Gold Standard; related assays not used in mineral resource estimate


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Figure 12-8: Control Chart for MEG-Au.11.19 - 2018

Note: Limiting lines such as “USL” etc. are explained immediately following Figure 12-7.

Figure 12-9: Grade Ranges and Dates of 2018 Pinion CRMs


12.6.1.4	CRMs for Drill Sample Silver Analyses in 2019

In 2019, Gold Standard sent drill-sample pulps for silver assays. The samples consisted of pulps from 1.524 m (5.0 ft) intervals of holes drilled by Gold Standard. These intervals were previously analyzed for gold at either ALS or Bureau Veritas. In conjunction with the silver analyses done in March and April of 2019, Gold Standard inserted 765 silver CRMs. Three different CRMs certified for silver were utilized. They are listed, with their target values, in the first two


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columns of Table 12-22. MDA evaluated the results for the CRMs using a variation of Shewhart-type charts, prepared by Gold Standard. The results of the 765 analyses of standards for silver are summarized in Table 12-22 and Table 12-23.

Table 12-22: Summary of 2019 Analyses of Silver CRMs


Standard ID	Grades in Ag g/t				Count	Dates Used		Failure Counts		Bias pct
Standard ID	Target	Average	Maximum	Minimum	Count	First	Last	High	Low	Bias pct
MEG-LWA-34	1.9	1.3	3	0.5	331	20-Mar-19	10-Apr-19	0	33*	-31.6
MEG-Au.11.29	13.4	13.8	20	2	336	20-Mar-19	15-Apr-19	7	1	3.0
MEG-Au.13.03	4.5	4.8	6	2	98	18-Apr-19	27-Apr-19	0	1	6.7
Totals
counts					765			7	2*
percentages					100			0.9	0.3


Note:	* the 33 “failures” of MEG-LWA-34 are not included in the calculations of failure percentages. See the discussion following Table 12-23.
	MEG-LWA-34 and MEG-Au.11.29 were used with samples from holes drilled in 2014 through 2018.
	MEG-Au.13.03 was used with samples from holes drilled in 2014 through 2016.

Table 12-23 List of Failed Silver CRMs


Standard ID	Sample ID	Target for Std (g/t)	Fail Type	Fail Limit (g/t)	Failed Value (g/t)	Comment
Standard ID	Sample ID	Target for Std (g/t)	Fail Type	Fail Limit (g/t)	Failed Value (g/t)	Comment
MEG-LWA-34	33 samples	1.9	low	0.7	<1	below detection limit; not material
MEG-Au.11.29	PR18-52 245-250-S2	13.4	high	16.1	20
MEG-Au.11.29	PR18-01 245-250-S2	13.4	high	16.1	17
MEG-Au.11.29	PR18-80 245-250-S2	13.4	high	16.1	20
MEG-Au.11.29	PIN16-19 245-250-S2	13.4	high	16.1	20
MEG-Au.11.29	PIN15-09 1845-1850-S2	13.4	high	16.1	18
MEG-Au.11.29	PIN15-19 245-250-S2	13.4	high	16.1	20
MEG-Au.11.29	PIN14-07 645-650-S2	13.4	high	16.1	19
MEG-Au.11.29	PR18-74 645-650-S2	13.4	low	10.7	2	sample mix-up?
MEG-Au.13.03	PIN14-27 45-50-S3	4.5	low	2.7	2

MEG-LWA-34 is a low-grade standard whose expected value is close to the 1.0 g Ag/t lower detection limit of the analytical method. The 33 low-side “failures” are not material in light of the precision of the analytical method, and they are not included in the calculations of low-side failure percentages in Table 12-22.

MDA evaluated the seven high-side failures of MEG-Au.11.29 in the context of adjacent silver assays in the same drill holes and the locations of mineral domains. MDA concluded that the failures are not material with respect to the silver estimate.


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The QA/QC data includes another 17 silver analyses of a CRM certified for gold but not for silver. The information characterizing the CRM notes an expected silver value of 0.07 g Ag/t. The 17 analyses, reported in ppb silver and converted to g Ag/t, fell in the range 0.015 to 0.030 g Ag/t. Given the low silver grades involved and the fact that the standard is not certified for silver, MDA does not believe that significant conclusions should be drawn from these results.

Gold Standard did not re-analyze any of the analytical batches containing failed silver standards.


12.6.2	Pinion Drill Program QA/QC Field Duplicates

In 2017 and 2018, Gold Standard collected field duplicates or rig duplicates, at what appear to be 30.48 m (100-foot) intervals. In most holes these amount to two or occasionally three duplicates per hole. Duplicates were obtained by collecting two samples simultaneously from a rotating wet splitter. Gold Standard did not collect duplicate samples in prior years.

MDA prepared three types of charts for the duplicates:

A scatterplot, showing an RMA regression;
A quantile/quantile plot; and
Several relative difference plots (see explanation, below).

MDA used a relative difference expressed as a percentage for each duplicate pair calculated as follows:

Table 12-24 summarizes the results for the field duplicates. The average of the relative difference listed in Table 12-24 is based on Equation 1 above and is an indication of the bias between the duplicates and the originals. The “Abs Rel Pct Diff” is the average of the absolute relative differences and gives an indication of the degree of variability between the duplicates and originals.

Table 12-24 Summary of Results for Pinion Field Duplicates


Type	Period	Corr. Coeff.*	Counts			RMA Regression	Averages as Percent
Type	Period	Corr. Coeff.*	All	Used	Outliers	(y = dup, x = orig)	Rel Pct Diff	Abs Rel Pct Dif
Field Dup	2017 - 2018	0.95	331	277	2	y = 1.059x - 0.010	-4.8	27.6

MDA also performed an alternative calculation as part of the evaluation of duplicates, but whose results are not listed in Table 12-24, using the following:

The disparity in Table 12-24 between the total number of pairs (“All”) and the number of pairs used (“Used”) exists because MDA did not include in calculations those pairs in which one or both analyses fell below the analytical detection limit. Two “outlier” pairs were also excluded because their differences were so great as to skew the statistics of the data set. Note that the average reported in the table is for all grades above the detection limit, excluding the outliers. Reporting single averages for the entire set of duplicates masks different responses in different grade ranges. See the chart in Figure 12-10 for a more complete view of the relative difference data.

As indicated by the relative difference shown in Table 12-24, and shown in more detail by the red moving average line in Figure 12-10, there is a tendency for the field duplicate samples to have slightly lower grades than the originals.


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Figure 12-10 shows the low bias of the duplicates to be most pronounced at mean grades below about 0.2 g Au/t, and almost absent at higher grades.

MDA does not have information on which to base any opinion as to the cause of the low bias in the duplicates at lower grades. MDA suggests that Gold Standard review procedures used for sampling, sample preparation, and analysis to determine if a cause can be identified and take corrective action if necessary.

Figure 12-10: Gold Relative Percent Difference – Pinion Duplicate vs. Original

The 2019 silver assays of pulps from earlier Gold Standard drill-hole samples included 309 samples having the suffix “dup.” MDA matched these to the original sample assays and evaluated the matched duplicate pairs. A summary of the evaluation appears in Table 12-25.

The set of silver duplicate assays includes 202 pairs for which both analyses were below the lower detection limit of the analytical method, and 12 pairs for which one of the analyses is below the lower detection limit. These 214 pairs were not used in the calculations of the statistics that appear in Table 12-25, leaving 95 pairs containing detectable silver in the evaluation. The results indicated in Table 12-25 are acceptable.

Table 12-25: Summary of Results for Duplicates in Silver Re-Assays


Type	Comment	Corr. Coeff.	Ag Grade Averages (g/t)		Counts			RMA Regression (y = dup, x = orig)	Averages as Percent
Type	Comment	Corr. Coeff.	Mean of Pair	Dup – Original	All	Used	Outliers	RMA Regression (y = dup, x = orig)	Rel Pct Diff	Abs Rel Pct Dif
Field Dup	all available excluding outliers	0.93	6.1	-0.2	30 9	95	0	y = 0.987x – 0.219	-3.2	30.3


Notes:	The differences between the numbers of duplicate pairs available (“All”) and those “Used” occurs because pairs in which one or both analyses fell below the method detection limit were excluded.
	Mop indicates mean of pair
	Relative differences shown in the last two columns of Table 12-25 are averages of those calculated using Equation 1. A negative relative difference indicates that, on average, the duplicate analyses were lower than the originals.


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12.6.3	External Check Assays for Pinion Drilling

In April and August 2018, Gold Standard selected pulps from two holes drilled in 2017 that were originally assayed by ALS and sent them to Bureau Veritas for check assays. In total, 95 usable original and check assay pairs were produced. MDA evaluated the results as a comparison between ALS and Bureau Veritas. MDA evaluated these using the same suite of charts and statistics used to evaluate the other types of duplicates described herein. Differences were calculated as the ALS value less the value reported by Bureau Veritas; in other words, where an ALS assay was higher than a Bureau Veritas assay, the difference is positive, and vice-versa.

The results of MDA’s evaluation are summarized in Table 12-26. The relative differences are illustrated in Figure 12-11. Thirteen pairs having a mean grade less than 0.033 g Au/t were excluded from the comparison because at such low grades small differences are large in relative terms and skew the statistics of the data set.

As with the other types of duplicates, MDA looked at grade subranges, in this case two, each having its own distinct characteristic statistics. Pairs having mean grades less than 0.033 g Au/t were ignored.

For sample pairs having mean grades between 0.033 g Au/t and 0.2 g Au/t, ALS was, on average, biased higher than Bureau Veritas, with an average relative difference of +4.7%. For pairs having mean grades greater than 0.2 g Au/t, a range comprising about 60% of the pairs, there was effectively no overall bias, although smaller subsets of the data can be seen on Figure 12-11 with biases of a few percentage points high or low.

In summary, the check assays done in 2018 on 2017 assay pulps revealed no issues of concern.

Figure 12-11: Gold Relative Percent Difference – ALS vs. Bureau Veritas, 2017 Pulps


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Table 12-26: Summary of Results for 2018 Re-Assays of 2017 Pinion Samples


Type	Comment			Corr. Coeff.	Grade Averages as g Au/t
Type	Comment			Corr. Coeff.	Mean of Pair	Dup – Original
Check	all available excluding Au < 0.033			0.999	1.072	-0.004
Check	subrange 0.040 ≤ mop < 0.2			0.986	0.101	0.006
Check	subrange mop > 0.20			0.999	1.473	-0.009
Type	Counts			RMA Regression (y = ALS, x = bv)	Averages as Percent
Type	All	Used	Outliers	RMA Regression (y = ALS, x = bv)	Rel Pct Diff	Abs Rel Pct Dif
Check	95	82	-	y = 0.976x + 0.021	1.4	4.4
Check	24	24	-	y = 1.136x – 0.007	4.7	7.2
Check	58	58	-	y = 0.971x + 0.034	0.0	3.3

Notes:

The differences between the number of duplicate pairs available (“All”) and those “Used” occurs because very low-grade pairs were excluded from statistical calculations, as were outliers. “mop” indicates mean of pair.
Pairs in which one or both assays are below detection limit are not used in statistical calculations.
< means “less than”; > means “greater than”; ≤ means “less than or equal to”; ≥ means “greater than or equal to”.
Relative differences in these tables are those calculated using Equation 1.


12.6.4	Pinion Drill Program QA/QC Blanks

In the period 2014–2018, Gold Standard used pulp blanks, obtained from a supplier of CRMs, and certified as containing gold below the usual laboratory detection limits. This type of blank tests the analytical process in the laboratory, but not the sample preparation process.

In 2017 and 2018, Gold Standard also used a blank consisting of coarse marble. This type of blank undergoes the full sample preparation and analytical process and can show if any sample-to-sample contamination takes place in the sample preparation process.


12.6.4.1	Blanks Used in 2014

In 2014 drilling at Pinion, Gold Standard used pulp blanks. In the data available to MDA there are 422 analyses of these blanks. The blanks were inserted into the sample stream every 30.48 m (100 feet). Five of the blanks, all in drill hole PIN14-44, are marked in the database as “labelled wrong,” so MDA did not use them in this evaluation. Among the 417 blank analyses that MDA did evaluate, six were reported to have detectable gold, with the maximum gold analysis reported being 0.017 g Au/t.

Figure 12-12 shows the gold analyses of the blanks, and in the adjacent drill samples in the sample stream. There is no meaningful evidence that the grades reported for the blanks are related to the grades in the adjacent samples, so there is no evidence of between-sample contamination in the analytical process. Pulp blanks are not useful for checking for contamination in the sample preparation process.


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Figure 12-12: Gold in Blanks and Preceding Samples 2014


12.6.4.2	Blanks Used in 2015

In 2015, Gold Standard used a pulp blank for drilling at Pinion. There are 296 analyses of the blank available, at downhole intervals of 30.48 m (100 feet). A chart of the blank analyses plotted along with the immediately preceding samples (pulps are assumed to be analyzed in sequence, but this was not directly assessed), is shown in Figure 12-13.

Figure 12-13: Gold in Blanks and in Preceding Samples 2015

Nineteen of the 296 analyses of pulp blanks in 2015 reported some detectable gold. In nine cases this was in the range 0.005 to 0.007 g Au/t. However, ten of the analyses of blanks, among samples from drill hole PIN15-14, reported gold in the range 0.032 to 0.083 g Au/t. These are noted on Figure 12-13, but do not seem to correlate with gold-rich samples. Gold Standard obtained re-analyses of fourteen mineralized samples analyzed in the same batch as the blanks in question. Table 12-27 summarizes a comparison between the original gold analyses and the re-run assays.


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Table 12-27: Comparison of Original Assays and Re-Runs in Part of PIN15-14


From-To in feet	Length in feet	Average grade original (Au-AA23?)	Average grade re-runs (Au-AA23)
710 – 765	16.68 m (55 ft)	0.415 g Au/t	0.35 g Au/t
800 – 815	4.57 m (15 ft)	0.098 g Au/t	0.10 g Au/t

Table 12-27 shows that the re-run assays for the interval from 216.41 to 233.17 m (710 to 765 ft) are on average lower than the original assays. In the Pinion database used for the estimation of mineral resources, the rerun assays were used.

12.6.4.3 Blanks Used in 2016

During the 2016 drilling campaign Gold Standard used a pulp blank having a certified value of “<0.003 ppm Au.” This pulp was inserted into the analytical stream every 30.48 m (100 feet), for 255 analyses of the pulp. Only once was detectable gold reported, at 0.006 g Au/t. The analyses of blanks in the 2016 campaign revealed no causes for concern.

12.6.4.4 Blanks Used in 2017 and 2018

During the 2017–2018 period two blanks were used. One was a pulp blank obtained from a commercial supplier, labelled as MEG-Blank.14.03. The other is described as “Gold Standard marble”, which MDA understands to be a coarse blank. Both have an expected value less than the lower detection limit for the analytical method in use, which is 0.005 g Au/t.

MDA prepared charts for both of these blanks. The charts are shown in Figure 12-14 and Figure 12-15. In the case of the pulp blanks shown in Figure 12-14, only four of 159 analyses reported gold exceeding the detection limit, and the highest grade reported was 0.009 g Au/t. There is no evidence that the analyses of the pulp blanks are affected by the grades of the samples that numerically precede them.

Eleven of the 58 analyses of coarse blanks were reported to contain detectable gold, with the highest reported grade being 0.011 g Au/t. Figure 12-15 gives a visual impression that blanks that follow higher-grade samples in numerical sequence are more likely to have reported gold analyses exceeding the lower detection limit. The correlation coefficient between the blanks and the preceding samples is a statistically significant 0.5. This suggests low-level contamination of blanks that follow higher-grade samples through the crushing and grinding process in the laboratory. The degree of contamination is not sufficient to be of concern in terms of using the gold analyses for a mineral resource estimate.

Figure 12-14: Gold in Pulp Blanks and in Preceding Samples 2017 – 2018


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Figure 12-15: Gold in Coarse Blanks and in Preceding Samples 2017 – 2018

12.6.4.5 Blanks Used in Silver Analyses

Pulp Blanks

The QA/QC data for silver include 646 analyses of pulp blanks, analysed with batches of samples from the 2014 through 2018 drill campaigns. The results are summarized in Table 12-28. They are acceptable.

Table 12-28: Results of Silver Analyses of Pulp Blanks


Analytical Result	Count of Analyses
below detection limit of 1 g Ag/t	632
at detection limit of 1 g Ag/t	13
2 g Ag/t	1

Coarse Blanks

Data for 15 silver analyses of coarse blanks are included in the QA/QC package that MDA obtained from Gold Standard. One of the 15 was from a series of samples in a 2014 drill hole. It was analysed using Bureau Veritas’ method MA401, with a lower detection limit of 1 g Ag/t and returned a result below the detection limit.

The other 14 analyses of coarse blanks were from a series of samples from 2018 core holes. The analytical method was Bureau Veritas’ AQ250, having a lower detection limit of 0.002 g Ag/t. All the 14 samples returned results above the lower detection limit, in the range 0.005 to 0.026 g Ag/t (0.005 g Au/t to 0.026 g Ag/t). MDA compared the results for the analyses of coarse blanks to the results for the immediately preceding real samples and found no statistically meaningful correlation.

12.6.5 Twin Holes

In 2018, four core holes were drilled into the Pinion deposit to obtain material for metallurgical testing. These four holes were twins of previously drilled holes. A comparison of length and grade of the intersected mineralization was made between these four pairs of twin holes. For the 172 m of drilling in the mineralized zones, the grade was 21% higher in the newer drill holes (Table 12-29). A histogram of the two sets of data show more low-grade samples in the older four holes (Figure 12-16).


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Table 12-29: Summary of Counts of Jasperoid Wash QA/QC Analyses


	Older Holes	Diff.	2018 holes	Units
Count	185		120
Length	172.0	0%	172.5	g Au/t
Grade	0.5	21%	0.6	g Au/t
Metal	84.4	21%	102.3	(Model/Actual)

Figure 12-16: Histogram of 2018 Twin Drill-Hole Samples

12.6.6 Pinion Drill Program QA/QC on Barite

Prompted by metallurgical data showing that the presence of barite affects gold recovery, Gold Standard analyzed samples for additional and different barium analyses. Specifically, barium grades of over around four percent were found to negatively affect metallurgical recovery (see Section 13). Consequently, there is a need to assess the quality of the barium analyses which are used to define the locations and quantity of barite in the Pinion deposit.

Initial barium analyses by ICP methods with two-acid digestion have been shown to be incorrect in grades above ~0.1% to ~0.2%. Later barium analyses were done at AAL on existing pulps using a pressed-powder XRF-ED analysis (method Ba ED-XRF E5 with a lower detection limit of 0.003% Ba). There were 938 barium assays done at AAL with this method. In addition, 21,747 loose-powder NITON XRF measurements of barium were made on drill-sample pulps by independent contractor Rangefront Geological.


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A total of 4,235 duplicate readings of barium content by the NITON XRF instrument were also taken by independent contractor Rangefront Geological. MDA compared 4,091 of these duplicate loose-powder NITON XRF readings to determine variability of results. Seventeen pairs were deemed to be extreme outliers and removed from the calculations. No significant biases were noted, and reproducibility was shown to be just over 10%.

For comparison to the Gold Standard loose-powder NITON XRF data, only 32 sample pulps were analyzed at AAL by a) ICP following a two-acid digestion, by b) ICP following a five-acid digestion, c) loose-powder NITON-XRF, d) pressed-powder XRF-ED, and e) XRF-WD (lithium metaborate fusion). The two-acid ICP analyses were 95% lower than the loose-powder NITON-XRF measurements, and the five-acid ICP analyses were 91% lower. The pressed-powder XRF-ED and XRF-WD analyses were 86% and 87% higher than the corresponding loose-powder NITON-XRF measurements. While 32 samples is not a statistically significant data set, the results do indicate good correlation with the XRF-ED analyses but with a slope to the regression line of 0.55. The XRF-ED analyses are being applied in the metallurgical test work (see Section 13.1).

12.7 GOLD STANDARD’S JASPEROID WASH DRILL PROGRAM QA/QC

Table 12-30 summarizes the types and quantities of QA/QC data that are available to MDA for the Jasperoid Wash drill samples.

Table 12-30: Summary Counts of Jasperoid Wash QA/QC Analyses


QA/QC Type	2017	2018
CRM
Number in Use	2	1
Number of Analyses	93	93
Number of Failures	1	1
Field Duplicate	113	153
Pulp Blank	66	75
Coarse Blank	-	10

The QA/QC data from Jasperoid Wash for 2017 and 2018 do not reveal any issues that would preclude use of any assays from the same time period in a mineral resource estimate, or that would reduce the confidence in those assays. The main issues noted by MDA are as follows:

In 2017 and 2018, although three separate CRMs were used, only one CRM was in use at Jasperoid Wash at any given time. The expected values for all three CRMs were either below or very close to any likely cutoff grade for a mineral resource. It would be prudent to use more than one CRM at a time, and to use CRMs with expected values that span the range of grades of likely economic importance;
In addition to field duplicates, the following additional types of duplicates, replicates or check assays are useful:
- Preparation duplicates, also called coarse crush duplicates, that are useful for monitoring the laboratory’s sample-preparation circuit;
- Analytical duplicates, sometimes called replicates, which are second splits from the original pulp; and
- Check assays done at a different lab than the original assays.


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Only pulp blanks were used at Jasperoid Wash in 2017. In 2018, pulp blanks and coarse blanks were both used, but the latter were only used with samples from one drill hole. Both types of blanks are useful, but if only one type is to be used, coarse blanks are the most useful, in that they provide information about potential contamination in the sample preparation process, which cannot be obtained from pulp blanks.

12.8 NORTH BULLION DEPOSITS DRILL PROGRAM QA/QC

All North Railroad drilling, including exploration and within the three deposit areas collectively called the North Bullion deposits, were evaluated together. Approximately 43% of the historical drill holes at North Railroad have paper lab reports/certificates that have been utilized in part to validate the assay database. A number of these lab reports/certificates contain obvious QA/QC data. However, none were in digital format and in many cases are of unknown origin and quality, and therefore were not evaluated for the historical drilling programs. These data should be compiled and evaluated where possible.

Drilling between 2010 and 2017 at North Railroad by Gold Standard, including exploration and deposit drilling, included a substantial number of blanks, CRMs, and duplicates (including samples inserted by Gold Standard and the laboratory). There has been no umpire assaying/sampling conducted with the 2010 to 2017 North Bullion, Sweet Hollow, and POD mineral resource drilling. It is not known if any failures occurred, or if failed batches were re-assayed and corrected in the database used for mineral resource estimation.

12.8.1 Blanks

A total of 1,244 blank pulp samples were inserted in the drilling sample stream from the North Railroad portion of the property between 2010 and 2017 by Gold Standard. In total, 114 of the 1,244 blank pulp samples returned a gold assay result above detectable limits (<5 ppb Au), or 9.2 % of the blank samples. Blanks were purchased from MEG Inc.

The 2010 to 2015 North Railroad drilling included 976 blank pulp samples inserted by Gold Standard personnel. There were 114 “failures” (11.7 %), which is much higher than one would normally expect in such a dataset. The maximum value detected was only 0.037 g Au/t, well below the mineral resource lower cut-off grade, therefore the results were considered immaterial. There were no issues with the blank pulps during the 2016 to 2017 North Railroad drilling.

12.8.2 Certified Reference Materials

A total of 1,477 CRMs was inserted in the sample stream for the North Railroad 2010 to 2017 drilling.

12.8.2.1 2010 to 2015 North Railroad CRMs

Between 2010 and 2015 a total of 1,073 CRM samples were inserted in the sample stream. Results from these CRMs are presented in Figure 12-17, which illustrates that there were no significant issues in fire assaying of the ten CRMs inserted by Gold Standard during the 2010 to 2015 North Railroad drilling programs. It is likely that at least some of the failures shown in Figure 12-17 were sample handling, labelling, and control problems since their assay values correspond with values of other CRMs. The total number of “failures” among CRMs was 35 out of the total of 1,073 samples, or 3.3%. The fail rate drops to 2.7% if one considers the data for the five CRMs with a “statistically significant” number of analyses (i.e., >50 analyses).


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Figure 12-17: Analyses of CRMs in North Railroad Drilling 2010 to 2015

12.8.2.2 2016 to 2017 North Railroad Standards

During the 2016 and 2017 drilling at the North Railroad portion of the property, a total of 404 CRMs were inserted in the drill-sample stream by Gold Standard. The analytical data for each of the six CRMs have been graphed separately but one graph is presented here for an example (Figure 12-18). The total number of outside-two standard deviations “failures” was 13 of 404 assays (3.2%). A slight positive bias was observed for this CRM; however, the results overall are considered acceptable.

The most significant issue recognized in the 2016 to 2017 North Railroad CRM-sample dataset was a slight, but consistent, positive bias with a 15% “failure” rate in CRM MEG-Au.11.17 (Figure 12-19). The 60 analyses were completed at both ALS and Bureau Veritas and both laboratories showed a similar positive bias thereby making the CRM sample suspect and minimizing the significance of the “failures” and “bias.” The results overall are considered acceptable.


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Figure 12-18: Analytical Results from Standard MEG-S107007X

Figure 12-19: Analytical Results from Standard MEG-Au.11

12.8.2.3 Duplicate Core-Sample Assays

There were 352 field duplicate samples analyzed during the 2010 to 2015 North Railroad drilling programs. The one-half cut core was cut in half and both quarter-core samples were sent for analysis. The duplicate assay data are illustrated in Figure 12-20, which shows good correlation (0.9777) between the duplicate and original samples.


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Figure 12-20: 2010 to 2015 North Railroad Portion Field Duplicate Assays

12.8.2.4 Lab-Inserted Standard Reference Materials

Bureau Veritas in Sparks, NV inserted internal laboratory QA/QC samples into the sample stream. Bureau Veritas reported the results for 45 CRMs, 35 blank pulps, and 17 replicate pulps during the 2010 to 2015 North Bullion-Bald Mountain drilling programs. During the 2016 drilling Bureau Veritas reported results for 144 blank pulps, 57 coarse blanks, 293 CRMs, 100 replicates, and 99 duplicates. While these lab-controlled QA/QC programs are not independent, they do add information. After detailed statistical evaluations and utilizing graphical displays, APEX concluded that the results yield acceptable failure rates, reasonable precision, and excellent correlation.

12.8.2.5 Summary of North Railroad Portion QA/QC

The QA/QC protocols employed by Gold Standard throughout the 2010 to 2017 drilling programs at the North Railroad portion of the property were adequate and appropriate for ensuring high accuracy and precision in the sample assays. Although there appear to have been sporadic and generally minor issues with QA/QC results, the 2010 to 2017 QA/QC sampling did not identify any significant issues with respect to the overall accuracy and precision of analytical results in the North Railroad drill assay database. APEX considers the North Railroad drill database sufficiently validated for use in the mineral resource estimation discussed Section 14.

12.9 SUMMARY STATEMENT ON DATA VERIFICATION

It is Mr. Ristorcelli’s opinion that the Dark Star, Pinion, and Jasperoid Wash analytical data are adequate for the purposes used in this Technical Report, subject to those samples removed and issues described above. The issues


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described above have been considered in assigning levels of confidence and the classification of the mineral resources estimated in Section 14.

There is no evidence of significant QA/QC programs in the North Railroad portion of the property prior to 2014 when Gold Standard began drilling. The QA/QC program was minimal in 2014 through 2016 having only had CRMs and pulp blanks. In 2017 and 2018, field duplicates and coarse blanks were inserted into the sample stream. In 2016 and 2017, some Dark Star pulps were sent out to secondary laboratories. The level of, and amount of QA/QC data, as well as non-remedied QA/QC “failures” was considered in mineral resource classification.

It is Mr. Dufresne’s opinion that the QA/QC protocols employed by Gold Standard throughout the 2010 to 2017 drilling programs at North Railroad are adequate and appropriate for ensuring high accuracy and precision in the sample assays. However, no QA/QC evaluations were done on historical drill campaigns. Mr. Dufresne considers the North Railroad drill database sufficiently validated for use in the mineral resource estimation discussed in Section 14.


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13 MINERAL PROCESSING AND METALLURGICAL TESTING

The current study of the South Railroad portion of the Railroad-Pinion property focuses on two main sources of ore: The Pinion and Dark Star deposits. Prior to acquisition of the property by GSV, numerous bottle roll and column leach tests were performed on these deposits using RC cuttings, diamond drill hole samples, and trench samples. A summary of these early tests is presented in Table 13-1.

Column leach tests on Pinion samples attained gold recoveries as high as 69 % (trench samples, -¼” crush). In general, bottle roll tests achieved higher maximum gold recoveries: 80.6% for Pinion and 82.2% for Dark Star.

Bottle roll and column leach recoveries for Pinion trench samples were inversely proportional to logarithm of particle size, as shown in Figure 13-1.

Figure 13-1: Plot of Column P80 (microns) vs. Gold Extraction (%)


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Table 13-1: Summary of Metallurgical Tests Prior to Gold Standard Ventures Tests.


Company	Lab and Test		Sample	Time h	P80 (P100) mm	Au Rec %	Ag Rec %	Calc Heads		NaCN lb/t	Lime lb/t	Comments
Company	Lab and Test		Sample	Time h	P80 (P100) mm	Au Rec %	Ag Rec %	oz/t Au	oz/t Ag	NaCN lb/t	Lime lb/t	Comments
PINION 1994-1995
Cyprus	McClelland	BR	35 Composite of RC cuttings		1.68	66.10%						Rapid reaction, low CN, moderate lime
					0.21	60.6 - 68
					0.074	3.8% Incr						Fine grind had little effect
		COL	880 kg bulk surface samples		(-2" & -3/4")	52.8 - 61.5
		BR	880 kg bulk surface samples		(-1/2"-100 mesh)	55.9- 80.6
PINION 2004
RSM	KCA	BR	5 trench samples	72	0.075	78%	54%	0.048	0.670	0.63	4
		COL	5 trench samples		0.53 " (-1.5")	57%	31%	0.046	0.29	1.35	2
					0.35" (-0.5")	59%	33%	0.049	0.42	1.08	2
					0.04 (-0.25)	69%	62%	0.048	0.37	1.88	2
DARK STAR 1991
Crown	McClelland	BR	158 RC cuttings (1.52 m drill intervals), 8 comps	96	59.8% -10 mesh	82.2		0.011- 0.043		0.27	10.5	most of Au leached after 24 h


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13.1 2015 – 2016 GOLD STANDARD PINION DEPOSIT CYANIDE BOTTLE-ROLL LEACH

Gold Standard commissioned three related bottle-roll test programs at KCA on a large number of samples extracted from composites made from Pinion drill intervals. The results were documented in three separate reports as follows: KCA (2016a), KCA (2016b), and KCA (2016c).

KCA (2016c) documented test results for 90 RC variability composites and KCA (2016b) reported results for 10 of the original 90 composites that were selected for re-run cyanide bottle-roll leach testing due to insufficient leach time. KCA (2016a) reported results on an additional 12 RC variability composites. Composites consisted of mostly oxide materials with some transition and sulfide samples.

13.1.1 2015 – 2016 Pinion Head Assays

Head assays and geo-metallurgical characterization were obtained for all 90 composites using a combination of three separate laboratories: KCA, ALS, and FL Schmidt (Simmons, 2019, Appendices 1,2, and 3), with the following results:

Gold grade ranged from 0.19 to 4.41 ppm and averaged 0.81 ppm;
Silver grade ranged from 0.62 to 72.3 ppm and averaged 6.9 ppm;
Organic carbon (not preg-robbing) ranged from 0.02 to 3.68% and averaged 0.18%;
Sulfide sulfur ranged from <0.01 to 4.18% (in the sulfide sample) and averaged 0.19%;
Preg-robbing analysis ranged from -1.70 to 35.2% and averaged 2.2%, which is considered non-preg robbing;
Copper values by ICP were very low, ranging from 5 to 39 ppm;
Cyanide solubility of gold ranged from 7.4 to 100% and averaged 78.3%;
Concentrations of the deleterious elements by ICP were: <5 ppm selenium, mercury ranged from 0.02 to 7.7 ppm, and arsenic was low at 47 to 1,360 ppm and averaged 280 ppm;
Concentrations of the primary cyanide consumers were low and suggest minimum potential for affecting cyanide-consumption rates. Copper averaged 17 ppm, nickel averaged 22 ppm, and zinc averaged 67 ppm; and
Silica content ranged from 28.1 to 96.7% by whole-rock analysis and averaged 81.4%.

13.1.2 2015 – 2016 Pinion Bottle-Roll Test Results

Bottle-roll leach cyanidation testing was conducted on 102 drill-core composites to evaluate the general leachability character of the Pinion geologic mineral resource. By design, these composites are not constrained by any pit shapes and therefore many of the composites may be located outside of any future economic pit limit. Bottle-roll testing was conducted at two targeted particle sizes: 80% passing 1,700 µm (10 mesh) and 80% passing 75 µm (200 mesh). Initially, retention times were 48-hrs for the 75 µm samples and 96-hrs for the 1,700 µm samples. Gold extraction results revealed that a significant number of samples were not completely leached in the allotted time frames.

Obvious under-leached samples were selected for re-leaching. The 75 µm samples were re-leached for 96 hours and the 1,700 µm samples were re-leached for 144 hours. All subsequent bottle-roll testing in a later program, KCA (2016a), were conducted at the longer retention times.

The 1,700 µm bottle-roll testing followed a standard procedure that is described in detail by the final KCA reports (KCA 2016a). The 75 µm bottle-roll procedure was the same as for the 1,700 µm bottle rolls, except the retention time was reduced to 96 hours. Results for the 1,700 µm bottle-roll test and 75 µm bottle-roll procedure are shown in Appendix 4 and 5 from the Metallurgy Report (Simmons, 2019).

For metallurgical testing, the Pinion mineral resource was divided into 12 zones. These are the Far North Zone (“FNZ”), North Zone North (“NZ-N”), North Zone Central (“NZ-C”), North Zone South (“NZ-S”), Main Zone (“MZ”), South East


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Central Zone (“SE-CZ”), East Pinion North Zone (“EP-NZ”), SE South Zone (“SE-SZ”), Central South Zone (“C-SZ”), NW South Zone (“NW-SZ”), NW Pinion West Zone (“NWP-WZ”), and the NW Pinion and East Zone (“NWP-EZ”). The zones from which each of the 102 composited sample material originated are shown in Figure 13-2 and listed in Metallurgical Report (Simmons, 2019, Appendices 4 and 5).

(bottle-roll cyanide-leach average gold recoveries, 200 and 10 mesh tests; composites from 2014 and 2015 Gold Standard drill holes at the Pinion deposit)

Figure 13-2: Pinion Zone Location Map for 2015 – 2016 Metallurgical Composites

Direct agitated cyanidation (bottle roll) tests were conducted on each of the 102 drill-core composites at particle size 80% passing 1.7 mm (10 mesh) and 75 µm (200 mesh), to determine gold extraction, extraction rate, reagent consumption, and sensitivity to feed size. The following is a summary of the findings from the bottle-roll test results:

13.1.2.1 10-Mesh Bottle-Roll Results 2015 - 2016

Gold head grades for the composites ranged from 0.15 to 4.65 ppm Au (average = 0.74 ppm Au). Gold extraction results ranged between 0.0 and 86.2 % (average = 65.0%). Three of the composites were sulfide (74852L, 74852M, and 74863I), and after removing them from the data set, the remaining transition and oxide composites ranged from 40.7 to 86.2% gold extraction (average = 66.7%).

Silver head grades for the composites ranged from 0.53 to 67.97 ppm Ag (average = 6.70 ppm Ag). Silver extraction results ranged from 3.1 to 69.4% (average = 24.3%). Three of the composites were sulfide (74852L = 10.5%, 74852M = 8.9%, and 74853I = 12.3%), and after removing them from the data set, the remaining transition and oxide composites averaged 24.7% silver extraction.

Cyanide consumption averaged 0.48 kg/t and lime consumption averaged 1.66 kg/t, with the three sulfide composites excluded from the averages.


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13.1.2.2 200-Mesh Bottle-Roll Results

Gold head grades for the composites ranged from 0.16 to 4.19 ppm Au (average = 0.75 ppm Au). Gold extraction results ranged from 0.0 to 94.0% (average = 76.1%). Three of the composites were sulfide (74852L, 74852M, and 74863I), and after removing them from the data set, the remaining transition and oxide composites had gold extractions from 44.3 to 94.0% (average = 77.9%).

Silver head grades for the composites ranged from 0.55 to 53.3 ppm Ag (average = 6.37 ppm Ag). Silver extraction results ranged between 13.0 and 83.0% (average = 46.8%). Three of the composites were sulfide (74852L = 23.5%, 74852M = 20.0%, and 74853I = 24.5%), and after removing them from the data set, the remaining transition and oxide composites averaged 47.5% silver extraction.

Cyanide consumption averaged 3.15 kg/t and lime consumption averaged 1.18 kg/t, with the three-sulfide composited excluded from the averages.

13.2 2016 - 2017 GOLD STANDARD PINION DEPOSIT METALLURGICAL TESTING

In 2016 - 2017, a total of 33 composites were made from intervals selected from 10 core holes, on two cross-sections, located in the Pinion North and NW Pinion Main zones. The drill hole locations for the 2016 – 2017 composites are shown in Figure 13-3. These composites were used for column-leach, bottle-roll, and load permeability testing at KCA in Reno, Nevada, and results are documented in a final report by KCA (2017a).

Fourteen of the 33 composites were selected and shipped to Hazen Research, Inc. (“HRI”) in Golden, Colorado, for SMC testing (SMC Test®) and Ai testing. Comminution and abrasion final test results were reported in KCA (2017a) and in a separate letter report from HRI (Stepperud, 2017a).


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Figure 13-3: 2016 – 2017 Pinion Metallurgical Core Hole Locations

(from Gold Standard 2017)

13.2.1 2017 Pinion Head Assays

Head assays and geo-metallurgical characterization were conducted on all composites using a combination of three separate laboratories: KCA, ALS, and UBC.

Head assays are tabulated for gold, silver, copper, cyanide gold solubility, carbon and sulfur species, and preg-robb analysis (Simmons, 2019, Appendix 6). ICP multi-element analyses and whole-rock analyses are shown in Appendix 7 and 8, respectively, in Metallurgical Report (Simmons, 2019). Gold cyanide-solubility (“AuCN”) assays presented are the average of two ALS assay procedures: AuAA13 and AuAA13s. The results for the 2016 – 2017 drill core composites are summarized below:

Gold grades ranged from 0.23 to 1.82 ppm and averaged 0.76 ppm;
Silver grades ranged from 3.3 to 38.7 ppm and averaged 10.4 ppm;
Organic carbon ranged from 0.04 to 0.218% and averaged 0.10%;
Sulfide sulfur ranged from <0.01 to 0.11% and averaged 0.03%;
Preg-robb analyses ranged from -6.20 to 18.2% and averaged 2.8% (considered non-preg robbing);
Copper values were very low, ranging from 1.5 to 74.8 ppm and averaged 6.1 ppm;
Gold cyanide solubility ranged from 70.2 to 94.4% and averaged 84.2%;
Concentrations of the deleterious elements were: selenium averaged 7 ppm, mercury ranged from 0.3 to 10.1 ppm with an average of 3.6 ppm, and arsenic levels were low ranging from 63 to 815 ppm with an average of 277 ppm;
Concentrations of the primary cyanide consumers were low and suggest minimum potential for affecting cyanide consumption rates. Copper averaged 22 ppm, nickel averaged 46 ppm, and zinc averaged 139 ppm;
Whole-rock silica content ranged from 25.7 to 89.1% and averaged 66.6 %.


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13.2.2 2016 – 2017 Bottle Roll and Column Leach Testing (KCA)

Twenty-four of the 33 drill core composites were subjected to bottle-roll leach testing at target P80 sizes of 75 µm and 1,700 µm, and to column-leach testing at either 12.5 mm or 25.0 mm crush sizes. The remaining nine composites were only bottle-roll leached at target P80 sizes of 75 µm and 1,700 µm. The testing program is summarized in Table 13-2. The main objective of these tests was to evaluate the laboratory-scale leachability character of the Pinion mineral resource in terms of gold extraction, extraction rate, reagent consumption, and sensitivity to feed size.

Table 13-2: Summary of Nominal Feed P80 for Column and Bottle-Roll Leach Tests


Pinion North Zone				Pinion NW Main Zone
Columns		Bottle Rolls		Columns		Bottle Rolls
12.5 mm	25 mm	75 µm	1,700 µm	12.5 mm	25 mm	75 µm	1,700 µm
13	1	20	20	9	3	13	13

The bottle-roll testing used a standard procedure that is described in the final laboratory report (KCA 2017), using 144 hours of retention time for 1,700 µm tests, and 96 hours for 75 µm tests.

Column-leach tests were conducted utilizing material crushed to target P80’s and placed in columns of 10 and 15 cm diameters. During testing the material was leached for 60, 90 or 121 days with a dilute NaCN solution. After leaching, each column was washed for four days with water. A portion of the leached and washed material (“tailings”) from each column was assayed for “tail screen” analyses by size fraction.

Tailings material from 12 columns was utilized for compacted permeability test work. Additionally, tailings material from seven columns was submitted to Western Environmental Testing Laboratory (“WETLAB”) in Sparks, Nevada, for acid-base accounting (“ABA”) and meteoric-water mobility tests (“MWMT”).

Geologic information for selected metallurgical composites, together with feed sizes, retention times, reagent consumptions, and gold and silver extraction balances can be found in the Metallurgical Report (Simmons, 2019, Appendix 9). The geologic information provided is part of the geo-metallurgical characterization of the Pinion mineral resource.

The following is offered as a summary of the findings from the 2016 – 2017 column and bottle-roll test results:

13.2.2.1 2017 Bottle-Roll Tests on 1,700 µm Composite Samples

Gold head grades for the composites ranged from 0.064 to 1.78 ppm Au, with an average of 0.74 ppm Au. From this material the gold extraction ranged from 49.0 to 86.0%, with an average extraction rate of 68.4%.

Silver head grades for the composites ranged from 3.4 to 40.4 ppm Ag, with an average of 10.4 ppm Ag. Silver extraction from this material ranged from 5.0 to 85.0%, with an average extraction rate of 26.9%.

Cyanide consumption averaged 0.18 kg/t and lime consumption averaged 0.78 kg/t.

13.2.2.2 2017 Bottle-Roll Tests on 75 µm Composite Samples

Gold head grades for the composites ranged from 0.13 to 1.85 ppm Au, with an average of 0.78 ppm Au. Gold extraction from this material ranged from 66.0 to 90.0%, with an average of 81.3%.


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Silver head grades for the composites ranged from 3.49 to 103.1 ppm Ag, with an average of 13.7 ppm Ag. Silver extraction from this material ranged from 16.0 to 95.0%, with an average of 49.0%.

Cyanide consumption averaged 0.88 kg/t and lime consumption averaged 0.60 kg/t.

13.2.2.3 2017 Column-Leach Tests on Composite Samples

Column-leach test extraction results were calculated based upon loaded carbon assays and tails assays. Gold head grades for the twenty-two 12.5 mm column composites ranged from 0.26 to 1.88 ppm Au (Average = 0.76 ppm Au). Gold extraction results ranged between 55.8 and 90.4%, with an average of 70.0%.

Silver head grades for the twenty-two 12.5 mm column composites ranged from 1.44 to 41.6 ppm Ag, with an average of 9.54 ppm Ag. Silver extraction results ranged between 5.4 and 47.3%, with an average of 22.7%

Cyanide consumption averaged 0.96 kg/t and lime consumption averaged 0.59 kg/t.

Gold head grades for the four 25.0 mm columns ranged from 0.44 to 0.90 ppm Au, with an average of 0.67 ppm Au. Gold extraction results ranged from 51.5 to 69.5%, with an average of 56.4%.

Silver head grades for the four 25.0 mm column composites ranged from 6.0 to 11.9 ppm Ag, with an average of 8.3 ppm Ag. Silver extraction results ranged between 9.7 and 44.8%, with an average of 22.6%

Cyanide consumption averaged 1.0 kg/t and lime consumption averaged 0.56 kg/t.

KCA advises that commercial-scale, operational cyanide consumption typically runs in the range of 25 to 33% of laboratory consumption.

Gold extraction plotted by days under leach for the column-leach tests are shown graphically in Figure 13-4.


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Figure 13-4: 2016 – 2017 Gold Extraction vs. Days Under Leach for Column-Leach Tests

13.2.3 2017 Pinion Comminution Characterization at HRI

Fourteen drill core samples were selected for comminution test work. These samples were limited to where sufficient material was available from the 2016 – 2017 metallurgical composites and represented major material types. They were subjected to the modified SMC Test at HRI to generate data for SAG mill comminution parameters, crushing index (“Mic”) by JKTech, and Ai testing. A final letter report was issued by HRI (Stepperud, 2017a).

13.2.3.1 2017 SMC Test Results

The 2017 HRI SMC Test® results for the 14 samples are given in the Metallurgical Report (Simmons, 2019, Appendix 10). This table includes the average rock density, A x b (a measure of resistance to impact breakage) and drop-weight index (“Dwi”) values that are the direct result of the SMC Test® procedure. The values determined for the Mia, Mih, and Mic parameters and the definitions of these abbreviations developed by SMCT are also presented in this table.

The DWi ranged from 2.13 to 8.02 kWh/m³, indicating soft to medium-hard material, and is tabulated along with other parameters of the SMC evaluation in the Metallurgical Report (Simmons, 2019, Appendix 10). In summary:

The Pinion samples A x b and Dwi values can be categorized as soft to moderate in comparison to the SMC worldwide database values. Although the Pinion oxide mineral resource material is not envisioned to require a milling circuit, the SAG comminution parameters are a primary component (output) of the SMC test, which also provides crushing parameters that can be used to design conventional crushing circuits.


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13.2.3.2 2017 Pinion Bond Abrasion Index (Ai) Tests

Bond Abrasion index tests were performed at HRI on 14 composite samples. The Metallurgical Report (Simmons, 2019, Appendix 11) lists the Ai values for the 14 composites that were tested. Ai values ranged from a low of 0.4591 g to a high of 1.5548 g, indicating moderate to very high abrasiveness of the materials tested. The silica content of the Pinion mineral resource is the inferred rock component that contributes to the corresponding high Ai test results.

13.2.3.3 2017 Pinion Comminution Test Summary

The Pinion comminution samples tested can be considered amenable to conventional, multi-stage crushing and screening circuit design. Mic, the SMC crusher component value, with an average of 5.9 kWh/t, would be ranked in the lower mid-range of the SMC worldwide database.

The Ai values (average = 0.9725 g) are modest to very high (see Simmons, 2019, Appendix 11) and represent the potential for high rates of wear on crusher liners, screen panels and conveyor drop boxes. The high Ai values of this material will likely translate into high wear rates on all ground-engaging equipment used for mining, including dozer tracks and blades, blast-hole drills, shovel and loader buckets, bucket teeth, and haul truck tires and bed liners.

13.2.4 2017 Pinion Load Permeability Test Work on Column Tailings

A portion of tailings material from each column-leach test was utilized for load permeability test work. The purpose of the load permeability test work was to examine the permeability of the crushed material under compaction loading equivalent to heap heights of 25 m, 50 m, 75 m, and 100 m.

The test cell utilized for modeling the permeability of stacked material at various heap heights, was a steel column or cell. Staged axial (vertical) loading of the test material was utilized to simulate the incrementally increased pressure obtained when loading the heap. Drainage layers were installed at the top and at the base of the column. External load was applied to the charge of material in the column utilizing a perforated steel plate that moved freely within the walls of the column.

A brief version of the guidelines that KCA utilizes when reviewing the results from this type of test are as follows:

1.	A slump of over 10% is generally an indication of failure.
2.	A measured flow of 10 times the heap design flow (10 to 12 li/h/m²) is considered a pass for a bed of agglomerate material. However, lower flows are not necessarily a failure if there are enough consistently passing tests.
3.	“Pellet breakdown” within the column of about 15% is marginally acceptable and anything higher is a failure. However, in general, a higher range may be allowable due to the subjective nature of the test, being based on visual observation. The tests only apply to materials agglomerated with cement.
4.	Solution color and clarity is typically an indicator of agglomerate failure and fines migration. This information is utilized in coordination with both slump as well as pellet breakdown to determine if the test column passes.

All twelve column residues that were tested passed using KCA’s criteria. The results of the load permeability test work are summarized in the Metallurgical Report (Simmons, 2019, Appendix 12).

13.3 GOLD STANDARD 2018 PINION DEPOSIT HIGH PRESSURE GRINDING ROLL (HPGR) TESTING

Gold Standard commissioned KCA to perform bottle roll, conventional-crush column-leach and HPGR-crush column-leach testing on a drill core composite sample from the Pinion Main zone, here termed the “HPGR composite.” Test results were documented in KCA (2018a).


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13.3.1 2018 Head Assays Pinion Main Zone HPGR Composite

The HPGR composite sample was comprised of intervals from two PQ-diameter core holes: PIN17-12, 42.7 m to 53.8 m and PIN17-13, 114.3 m to 159.1 m. Head assays are presented in Table 13-3. The sulfide sulfur (S⁼) head assay of 0.02% demonstrates the oxide character of this sample. The presence of C(org) (0.11%) and the preg-robb assay of 9.5% indicate that this composite may be mildly preg-robbing.

Table 13-3: Pinion Main Zone HPGR Composite Head Assays


KCA Sample No.	Description	Au & Ag Assays				Sulfur and Carbon Species					Preg-robb, %
		Au	Ag	AuCN	AgCN	C(tot)	C(org)	S (tot)	S=	SO4
		ppm	ppm	%	%	%	%	%	%	%
Pinion HPGR Composite Sample
78508C	Pin 17-12 140’ to 176.5’ and Pin 17-13 375’ to 522’	0.736	4.53	80.2	73.5	0.19	0.11	3.07	0.02	3.05	9.5%

13.3.2 2018 Pinion Main Zone HPGR Bottle-Roll and Column-Leach Testing

The Pinion Main zone HPGR composite was also subjected to bottle-roll leach testing at target P80 sizes of 38 µm, 75 µm and 1,700 µm. Conventional column-leach testing was conducted at target P80 of 12.5 mm and HPGR column-leach testing was done on sub-samples subjected to low, medium, and high HPGR press forces. The main objective of these bottle-roll and column-leach tests was to evaluate the differences in gold extraction, comparing conventional-crush laboratory column-leach results to those from material crushed using HPGR.

13.3.2.1 2018 Bottle-Roll Tests, Pinion Main Zone HPGR Composite Sample

Bottle-roll leach testing was performed on 500 g or 1,000 g portions of head material comminuted to a P80 target size of 1,700 microns (1.70 mm), 75 microns (0.075 mm), and 38 microns (0.038 mm). Bottle-roll testing, wet screening and assay methods were performed utilizing the same procedures as outlined in Section 13.2.2. Bottle-roll cyanide-leach test results are shown in Table 13-4.

Table 13-4: 2018 Pinion Main Zone HPGR-Crushed Bottle-Roll Results

KCA
Sample
No.

Test No

Comp ID

GSV Geology

Feed Size

Leach
Time
(hrs)

Au Balance

Ag Balance

Reagents

Zone

Subunit

Rock
Type 1

Vein ¹

Target
P (µm)
80

Screen
P (µm)
80

Au Ext
%

Calc Hd
Au
(ppm)

Ag Ext
%

Calc Hd
Ag
(ppm)

Na CN
kg/t

Lime
kg/t

78508C

78525 A

HPGR Comp

Pinion Main

CGL

car

qzv

1,700

1,860

144

53.8

0.630

32.5

4.640

0.11

0.50

78508C

78526 A

HPGR Comp

Pinion Main

CGL

car

qzv

72.5

0.803

58.3

4.940

0.50

78508C

78526 B

HPGR Comp

Pinion Main

CGL

car

qzv

69.8

0.758

59.9

4.810

0.27

0.50

The reported bottle-roll cyanide-leach gold extractions are low for an oxide sample. This is an indication of refractoriness due to factors other than sulfide sulfur or C(org) contents.

13.3.2.2 Column-Leach Tests on Pinion Main Zone HPGR Composite

Column-leach tests were performed on four samples of the HPGR composite that were prepared in the following manner:


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1 – Conventional crush to target P80 = 12.5 mm;

2 – HPGR crushed at low press force (2.20 N/mm²) setting, P80 = 7,000 µm;

3 – HPGR crushed at medium press force (3.35 N/mm²) setting, P80 = 6,500 µm;

4 – HPGR crushed at high press force (4.30 N/mm²) setting, P80 = 5,000 µm

The column-leach tests were conducted for 65 days with a dilute sodium cyanide solution utilizing the same procedures as outlined in Section 13.2.2. The results are summarized in Table 13-5.

Table 13-5 2018 Pinion Main Zone HPGR-Crushed Column Leach Test Results

KCA
Sample
No.

Test No

Comp ID

GSV Geology

Feed Size

Leach
Time
(days)

Au Balance

Ag Balance

Reagents

Zone

Subunit

Rock
Type 1

Vein 1

Target
P ₈₀(µm)

Screen
P ₈₀(µm)

Au Ext,
%

Calc Hd
Au
(ppm)

Ag Ext
%

Calc Hd
Ag
(ppm)

NaCN
kg/t

Lime
kg/t

78509B

78516

HPGR - Low

Pinion
Main

CGL

car

gzv

N/A

7,000

53.3

0.846

37.4

4.600

0.54

1.01

78510B

78519

HPGR - Med

Pinion
Main

CGL

car

gzv

N/A

6,500

65.5

0.722

39.9

4.560

0.57

1.01

78511B

78522

HPGR - High

Pinion
Main

CGL

car

gzv

N/A

5,000

64.0

0.708

42.8

4.070

0.61

1.01

78508C

78513

Conventional
Crush

Pinion
Main

CGL

car

gzv

12,500

12,200

43.8

0.864

21.6

3.990

0.54

1.02

The column-test extractions in Table 13-5 are based upon pregnant solution carbon assays using the calculated head (carbon assays + tails assays), which ranged from 0.71 g Au/t to 0.85 g Au/t. Gold extractions ranged from 44% (conventional crush) to 66% (HPGR medium pressure). Sodium cyanide consumption ranged from 0.54 to 0.61 kg/t and hydrated lime consumption ranged from 1.01 to 1.02 kg/t.

Graphical comparisons for gold and silver extraction between conventionally crushed and HPGR-crushed sample charges are shown in Figure 13-5 and Figure 13-6.


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Figure 13-5: Conventional Crush vs. HPGR Gold Extraction Comparison

Figure 13-6: Conventional Crush vs. HPGR Silver Extraction Comparison

The data demonstrate that the HPGR-crushed column charges, at medium and high press force, provide a significant gold extraction advantage over the conventionally crushed sample. While it is relatively simple to design a flowsheet


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to produce any specific P80 particle size from conventional crushing, it is not for HPGR comminution. The P80’s shown in Figure 13-5 and Figure 13-6 represent a close approximation to the product size that would be produced in a commercial HPGR comminution circuit.

13.3.3 2018 Pinion Main Zone HPGR Agglomeration and Load Permeability Testing

Preliminary agglomeration testing was performed on the low, medium, and high press-force HPGR-comminuted samples before being loaded into columns. All charges passed the KCA agglomeration criteria except the medium press-force sample at “0” kg/t cement addition. It was decided to column leach the HPGR samples without any cement addition or agglomeration for this phase of test work.

All column-leach residue charges were subjected to evaluation of percent slump, maximum percolation rate and load permeability tests. The results are shown respectively in Appendix 13, 14, and 15 from the Metallurgical Report (Simmons, 2019).

The medium and high press-force HPGR column-leach residues failed load permeability testing at 50 m height. Based upon these results it is recommended that future testing continue to evaluate cement agglomeration on HPGR-comminuted samples to support heap heights of at least 50 m and possibly 75 m.

13.4 2019 GOLD STANDARD PINION DEPOSIT METALLURGICAL TEST WORK

Gold Standard drilled additional met core holes in the Pinion North and Main zones in 2017-2018, that were tested in 2019. A total of 26 composites were made from intervals selected from 22 core holes. Metallurgical core drill hole locations, for all phases of work, is shown in Figure 13-7, and the 2017- 2018 composites are shown in green and blue. These composites were used for geo-metallurgical characterization, comminution testing, column-leach, bottle-roll, load permeability testing, and environmental characterization, at KCA in Reno, Nevada, and results are documented in a final report by KCA (2019a).

Nine of the 26 composites were selected and shipped to HRI”) in Golden, Colorado, for SAG mill comminution (“SMC”) testing (SMC Test®) and Bond Abrasion index (“Ai”) testing. Comminution and abrasion final test results were reported in KCA (2019a) and in a separate letter report from HRI (Stepperud, 2019a).


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Figure 13-7: Pinion Deposit Metallurgical Core Location Map 2019 Pinion Head Assays

Head assays and geo-metallurgical characterization were conducted on all composites using a combination of three separate laboratories: KCA, ALS, and University of British Columbia (“UBC”).

Head assays are tabulated for gold, silver, copper, cyanide gold solubility, carbon and sulfur species, and preg-robb analysis (Simmons, 2019, Appendix 16). In the Metallurgical Report, ICP multi-element analyses are shown in Appendix 17, whole-rock analyses are shown in Appendix 18 and QXRD analysis in Appendix 19 (Simmons, 2019).


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Gold cyanide-solubility (“AuCN”) assays presented are the average of two ALS assay procedures: AuAA13 and AuAA13s. The results for the 2017 – 2018 drill core composites are summarized below:

Gold grades ranged from 0.25 to 2.87 ppm and averaged 0.85 ppm;
Silver grades ranged from 0.5 to 29.1 ppm and averaged 7.7 ppm;
Organic carbon ranged from 0.08 to 0.45% and averaged 0.25%;
Sulfide sulfur ranged from 0.005 to 0.67% and averaged 0.078%;
Preg-robb analyses ranged from -4.0 to 16.7% and averaged 4.5% (considered non-preg robbing);
Copper values were very low, ranging from 1.2 to 38.9 ppm and averaged 6.0 ppm;
Gold cyanide solubility ranged from 43.6 to 87.3% and averaged 74.8%;
Concentrations of the deleterious elements were: selenium averaged 8.4 ppm, mercury ranged from 0.2 to 4.3 ppm with an average of 1.4 ppm, and arsenic levels were low ranging from 1 to 607 ppm with an average of 273 ppm;
Concentrations of the primary cyanide consumers were low and suggest minimum potential for affecting cyanide consumption rates. Copper averaged 119 ppm (with 1 composite containing 2260 ppm), nickel averaged 29 ppm and zinc averaged 134 ppm;
Whole-rock SiO2 content ranged from 11.0 to 94.8% and averaged 73.3%.

13.4.2 2019 Pinion Bottle Roll and Column Leach Testing at KCA

Twenty-six drill core composites were subjected to bottle-roll leach testing at target P80 sizes of 75 µm and 1,700 µm, and to column-leach testing at 12.5 mm or 25.0 mm crush sizes. The main objective of these tests was to evaluate the laboratory-scale leachability character of the Pinion mineral resource in terms of gold extraction, extraction rate, reagent consumption, and sensitivity to feed size.

Geologic information for selected metallurgical composites, together with feed sizes, retention times, reagent consumptions, and gold and silver extraction balances are shown in the Metallurgical Report (Simmons, 2019, Appendix 20).

The bottle-roll testing used a standard procedure that is described in the final laboratory report (KCA 2019a), using 144 hours of retention time for 1,700 µm tests, and 96 hours for 75 µm tests.

Column-leach tests were conducted utilizing material crushed to target P80’s and placed in columns of 10 and 15 cm diameters. During testing the material was leached for 59, 70, 94 or 130 days with a dilute NaCN solution. After leaching, each column was washed for four days with water. A portion of the leached and washed material (“tailings”) from each column was assayed for “tail screen” analyses by size fraction.

Tailings material from 19 columns was utilized for compacted permeability test work. Additionally, tailings material from the same 19 columns was submitted to Western Environmental Testing Laboratory (“WETLAB”) in Sparks, Nevada for environmental characterization.

The following is offered as a summary of the findings from the 2019 column and bottle-roll test results:

13.4.2.1 2019 Pinion Bottle Roll Tests on 75-µm Composite Samples

Gold head grades for the composites ranged from 0.138 to 2.63 ppm Au, with an average of 0.81 ppm Au. From this material the gold extraction ranged from 30.1 to 87.4%, with an average extraction rate of 68.4%.

Silver head grades for the composites ranged from 0.60 to 32.0 ppm Ag, with an average of 7.6 ppm Ag. Silver extraction from this material ranged from 31.7 to 84.3%, with an average extraction rate of 56.2%.


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Cyanide consumption averaged 0.30 kg/t and lime consumption averaged 0.64 kg/t.

13.4.2.2 2019 Pinion Bottle-Roll Tests on 1,700 µm Composite Samples

Gold head grades for the composites ranged from 0.156 to 2.59 ppm Au, with an average of 0.78 ppm Au. Gold extraction from this material ranged from 37.5 to 79.0%, with an average of 61.5%.

Silver head grades for the composites ranged from 0.55 to 31.8 ppm Ag, with an average of 7.63 ppm Ag. Silver extraction from this material ranged from 12.0 to 82.2%, with an average of 30.7%.

Cyanide consumption averaged 0.22 kg/t and lime consumption averaged 1.07 kg/t.

13.4.2.3 2019 Pinion Column-Leach Tests on Conventional Crushed Composite Samples

Column-leach test extraction results were calculated based upon loaded carbon assays and tails assays. Gold head grades for the sixteen 12.5 mm column composites ranged from 0.198 to 3.19 ppm Au (average = 0.95 ppm Au). Gold extraction results ranged between 29.8 and 80.0%, with an average of 63.0%.

Silver head grades for the sixteen 12.5 mm column-leach composites ranged from 0.65 to 27.9 ppm Au, with an average of 7.84 ppm Ag. Silver extraction results ranged between 9.5 and 76.4%, with an average of 30.6%.

Cyanide consumption averaged 0.92 kg/t and lime consumption averaged 1.06 kg/t.

Gold head grades for the ten 25.0 mm column-leach composites ranged from 0.313 to 1.65 ppm Au, with an average of 0.72 ppm Au. Gold extraction results ranged between 30.5 and 81.2%, with an average of 59.9%.

Silver head grades for the ten 25.0 mm column-leach composites ranged from 1.94 to 22.2 ppm Ag, with an average of 7.0 ppm Ag. Silver extraction results ranged between 9.3 and 25.0%, with an average of 18.2%.

Cyanide consumption averaged 0.76 kg/t and lime consumption averaged 0.93 kg/t.

KCA advises that commercial-scale, operational cyanide consumption typically runs in the range of 25 to 33% of laboratory consumption.

Gold extraction plotted by days under leach for the column-leach tests are shown graphically in Figure 13-8.

13.4.2.4 2019 Pinon Column Leach Tests on HPGR Crushed Composite Samples

Column-leach test extraction results were calculated based upon loaded carbon assays and tails assays. Gold head grades for the five HPGR crush column-leach composites ranged from 0.467 to 0.891 ppm Au (average = 0.65 ppm Au). Gold extraction results ranged between 56.6 and 78.3%, with an average of 70.3%.

Silver head grades for the five HPGR crush column-leach composites ranged from 1.72 to 28.0 ppm Au, with an average of 7.27 ppm Ag. Silver extraction results ranged between 27.9 and 58.3%, with an average of 40.4%.

Cyanide consumption averaged 0.73 kg/t and lime consumption averaged 0.31 kg/t.

Gold extraction plotted by days under leach for the column-leach tests are shown graphically in Figure 13-8.


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Figure 13-8: 2019 Pinion Gold Extraction vs. Days under Leach for Column-Leach Tests

13.4.3 2019 Pinion Comminution Characterization at HRI

Nine drill core samples were selected for comminution test work and were subjected to the modified SMC Test at HRI to generate data for SAG mill comminution parameters, crushing index (“Mic”) by JKTech, and Ai testing. A final letter report number 12635 was issued by HRI, March 6, 2019 (Stepperud, 2019a).

13.4.3.1 2019 Pinion SMC Test Results

The 2019 HRI SMC Test® results for the 9 samples are given in the Metallurgical Report (Simmons, 2019, Appendix 21). This table includes the average rock density, A x b (a measure of resistance to impact breakage) and drop-weight index (“Dwi”) values that are the direct result of the SMC Test® procedure. The values determined for the Mia, Mih and Mic parameters, and the definitions of these abbreviations developed by SMCT, are also presented in this table.

The DWi ranged from 5.34 to 7.30 kWh/m³, indicating soft to medium-hard material, and is tabulated along with other parameters of the SMC evaluation.


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The Pinion samples A x b and DWi values can be categorized as moderate in comparison to the SMC worldwide database values. Although the Pinion oxide mineral resource material is not envisioned to require a milling circuit, the SAG comminution parameters are a primary component (output) of the SMC test, which also provides conventional crushing parameters that can be used to design conventional crushing circuits.

The 2019 Pinion comminution samples can be considered in line with previous testing and is amenable to conventional, multi-stage crushing and screening circuit design. Mic, the SMC crusher component value, with an average of 6.8 kWh/t, would be ranked in the mid-range of the SMC worldwide database.

13.4.3.2 2019 Pinion Bond Abrasion Index (Ai) Tests

Bond Abrasion index tests were performed at HRI on 9 composite samples. The Metallurgical Report (Simmons, 2019, Appendix 22) lists the Ai values for the 9 composites that were tested. Ai values ranged from a low of 0.4005 g to a high of 0.8481 g, indicating moderate to above average abrasiveness of the materials tested. The silica content of the Pinion mineral resource is the inferred rock component that contributes to the corresponding high Ai test results.

The 2019 Pinion Ai values (average = 0.6948 g) are lower than the Phase one samples (0.9725 g) and can be considered as moderate to above average (see Appendix 22, Simmons, 2019) and represent the potential for slightly elevated rates of wear on crusher liners, screen panels, and conveyor drop boxes.

13.4.4 2019 Pinion Load Permeability Test Work on Column Tailings

A portion of tailings material from ten 12.5 mm, five 25 mm, and five 4 HPGR column-leach test residues was utilized for load permeability test work. The purpose of the load permeability test work was to examine the permeability of the crushed material under compaction loading equivalent to heap heights of 25 m, 50 m, 75 m, and 100 m.

Load Permeability Test procedures and guidelines have been described earlier in this Technical Report. Refer to Section 13.2.4 for details.

All ten 12.5 mm and five 25 mm conventional crush columns passed load permeability test criteria up to 100-meter heap height, except for PM #59, which failed at 100 meters. The conventional crushed column load permeability test results are summarized in the Metallurgical Report, Appendix 23 (Simmons, 2019).

All five of the HPGR crushed columns passed load permeability test criteria at 75 meters and two of the five (PM #37 and PM #56) failed at 100 meters. 2.0 kg/t of cement was added to PM #51 and 6.0 kg/t to PM #59. The HPGR crushed column load permeability test results are summarized in the Metallurgical Report, Appendix 24 (Simmons, 2019).

13.5 1991 DARK STAR DEPOSIT METALLURGICAL TESTING

Figure 13-9 shows drill hole locations for samples used for Dark Star bottle roll tests conducted by McClelland Laboratories for Crown Resources in 1991. Bottle roll test results are summarized in Table 13-1.


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Figure 13-9: RC Drill Hole Locations for the 1991 Dark Star Bottle-Roll Tests

13.6 2017 GOLD STANDARD DARK STAR DEPOSIT METALLURGICAL TESTING

In 2017 Gold Standard commissioned KCA to complete a metallurgical testing program on drill core composite samples from the Dark Star Main and North mineral resources. Test results were documented in KCA (2017b).

13.6.1 2017 Dark Star Head Assays for Bottle-Roll and Column-Leach Tests

Head assays and geo-metallurgical characterization analyses were obtained for 68 composites using a combination of four separate laboratories: KCA, ALS, UBC, and FLS. The head assays are tabulated in Appendix 25, 26, and 27 (Simmons, 2019) and show:

Gold grade ranged from 0.177 to 7.35 ppm and averaged 1.59 ppm;
Silver grade ranged from 0.27 to 5.07 ppm and averaged 0.71 ppm;
Organic carbon ranged from <0.10 to 2.14% (sulfide sample) and averaged 0.24%;
Sulfide sulfur ranged from <0.01 to 2.14% (sulfide sample) and averaged 0.21%;
Preg-robb analysis ranged from 0.0 to 19.2% and averaged 1.6%;
Copper values were very low, ranging from 5 to 42 ppm;
Gold cyanide solubility ranged from 25.7% (sulfide sample) to 100% and averaged 88.3%;


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Concentrations of deleterious elements by ICP were low: <5 ppm selenium on average, mercury ranged from 0.99 to 127.4 ppm (sulfide sample) and averaged 9.54 ppm, and arsenic ranged from 61 to 605 ppm with an average of 196 ppm;
Concentrations of the primary cyanide consumers were low and suggest minimum potential for effecting cyanide consumption rates. Copper averaged 18 ppm, nickel averaged 32 ppm and zinc averaged 126 ppm.
Whole-rock quartz (SiO2) analyses were high ranging from 42.4 to 95.8% and averaged 85.0%.

13.6.2 2017 Dark Star Bottle-Roll and Column-Leach Tests at KCA

Sixty-eight drill core composites were subjected to bottle-roll leach testing at target P80 sizes of 75 µm and 1,700 µm. A subset of 41 of the 68 composites were subjected to column-leach testing at crush sizes of 12.5 mm (on all 41 composites) and 25.0 mm (six of the 41 composites), depending upon available mass. The main objective of the bottle-roll and column-leach testing was to evaluate laboratory-scale leachability of the Dark Star mineral resource in terms of gold extraction, extraction rate, reagent consumption, and sensitivity to feed size.

13.6.2.1 2017 Dark Star Bottle Roll Tests

Bottle-roll leach testing was conducted on portions of material from each of the 68 composites. A 500 or 1,000 g portion of head material was crushed to a nominal size of 1,700 µm (1.70 mm) and utilized for leach testing. A second portion of material was milled in a laboratory rod mill to a target size of 80% passing 75 µm (0.075 mm). The milled slurry was then utilized for leach testing. The tests, which are described in detail by the laboratory report (KCA 2017), employed retention times of 144 hours for the 1,700 µm material and 72 hours for the 75 µm material.

The tailing material from the 1,700 µm tests was wet screened at 0.075 mm. The undersized material was dried and set aside. The oversized material was dried and dry screened at 4.75, 3.35, 2.36, 1.70, 1.18, 0.850, 0.600, 0.425, 0.300, 0.212, 0.150, 0.106, and 0.075 mm. The dry-screened -0.075 mm material was then combined with the wet screened material. Each separate size fraction was then weighed and reported. The material was then recombined. From the recombined material, three portions were split out and individually ring and puck pulverized to 80% passing 0.075 mm. The pulverized portions were then assayed for residual gold and silver content. The reject material was stored.

The tailing material from the 75 µm tests was wet screened at 0.038 mm. The undersized material was dried and set aside. The oversized material was dried and dry screened at 0.212, 0.150, 0.106, 0.075, 0.053, and 0.038 mm. The dry-screened, -0.038 mm material was then combined with the wet-screened material. Each separate size fraction was then weighed and reported. The material was then recombined. From the recombined material, three portions were split out and individually ring and puck pulverized to 80% passing 0.075 mm. The pulverized portions were then assayed for residual gold and silver content. The reject material was stored.

Gold Standard has divided the Dark Star deposit into two zones for metallurgical testing: Dark Star Main and Dark Star North. The drill holes, shown with numbers in Figure 13-10, are core holes from 2015-2016 drilling, from which metallurgical composites were compiled. Dark Star bottle-roll gold and silver extraction results are summarized in Appendix 28 and 29 from the Metallurgical Report (Simmons, 2019). The zones from which the 2015-2016 composite sample material originated are listed in Appendix 30 and 31 (Simmons, 2019).


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Figure 13-10 Location Map for 2017 Dark Star Metallurgical Composites

Note: BR = bottle roll test; COL = column-leach test.

The following is a summary of the findings from the Dark Star bottle roll test results.

13.6.2.2 2017 Dark Star 1.70 mm (10 Mesh) Bottle-Roll Results

Dark Star 10-mesh bottle-roll gold and silver extraction results are shown in the Metallurgical Report (Simmons, 2019, Appendix 28). Gold head grades for the 10-mesh composite samples ranged from 0.18 to 6.22 ppm Au, with an average of 1.56 ppm Au. Gold extraction ranged between 26.1 and 97.7% and averaged 81.8%. Five of the composites were sulfide/carbon refractory, with gold cyanide solubility <60%, nine of the composites were transitional with gold cyanide solubility >60% and <85%, and 54 of the composites were oxide with AuCN solubility >85%.

Silver grades are very low at Dark Star. Silver head grades for the 10-mesh composites ranged from 0.31 to 5.01 ppm Ag with an average of 0.71 ppm Ag. Silver extraction ranged from 0.0 to 83.6% and averaged 20.2%. Cyanide consumption averaged 0.42 kg/t and lime consumption averaged 1.11 kg/t.

13.6.2.3 2017 Dark Star 0.74 mm (200 Mesh) Bottle-Roll Results

Dark Star 200-mesh bottle roll gold and silver extraction results are shown in the Metallurgical Report (Simmons, 2019, Appendix 29). Gold head grades for the 200-mesh composite samples ranged from 0.22 to 6.48 ppm Au with an average of 1.55 ppm Au. Gold extraction ranged between 30.9 and 97.9% and averaged 85.6%. Five of the composites were sulfide/carbon refractory with gold cyanide solubility <60%, nine of the composites were transitional with gold cyanide solubility >60% and <85%, and 54 of the composites were oxide with gold cyanide solubility >85%.

Silver grades are very low at Dark Star. Silver head grades for the 200-mesh composites ranged from 0.24 to 5.06 ppm Ag with an average of 0.69 ppm Ag. Silver extraction ranged from 5.6 to 85.8% and averaged 31.5%.


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Cyanide consumption averaged 1.79 kg/t and lime consumption averaged 0.77 kg/t.

13.6.2.4 2017 Dark Star 12.5 mm and 25.0 mm Column Leach Results

Forty-one of the 2017 composites were column leached utilizing material crushed to 100% passing 19 mm (target P80 = 12.5 mm), and six of the 41 composites were crushed to 100% passing 37.5 mm (target P80 = 25 mm). During testing the material was leached for 60, 90 or 121 days with dilute NaCN solution and placed, respectively, in columns of 100 mm and 150 mm diameters. After leaching, each test was washed for four days with water. A portion of the tailings material from each column-leach test was utilized for tail screen analyses with assays by size fraction. Column-leach gold and silver extraction results are summarized in the Metallurgical Report (Simmons, 2019, Appendix 32).

None of the columns required agglomeration in the laboratory column set-up. Column-leach extraction results were calculated based upon loaded carbon assays and tails assays. Calculated gold head grades for the 41 columns ranged from 0.18 to 6.39 ppm Au with an average of 1.58 ppm Au. Gold extraction ranged between 15.0 and 94.8% with an average of 78.9%.

Two of the column-leach composites were sulfide/carbon refractory with gold cyanide solubility <60% and gold extraction ranged from 15.0 to 25.5% and averaged 20.3%.
Eight of the column-leach composites were transitional with gold cyanide solubility >60% and <85%. Gold extraction for these columns ranged from 57.8 to 85.8% and averaged 69.7%.
Thirty-seven of the column-leach composites were oxide with gold cyanide solubility >85%. Gold extraction for these columns ranged from 56.3 to 94.9% and averaged 84.1%.

Calculated silver head grades for the 47 columns ranged from 0.30 to 2.54 ppm Ag with an average of 0.58 ppm Ag. Silver extraction ranged between 14.3 and 68.0% with an average of 31.1%. Silver head grades for Dark Star are very low and of minimal economic significance.

Cyanide consumption averaged 1.07 kg/t and lime consumption averaged 1.15 kg/t. Commercial scale ROM cyanide consumptions are expected to be in the range of 25 to 33% of laboratory-scale test results. Laboratory lime consumptions are assumed to be similar to commercial-scale consumptions.

Gold extraction versus days under leach for the 47 column-leach tests are shown graphically in Figure 13-11. The two low gold extraction plots show in Figure 13-11, are for the sulfide composites discussed in the first bullet above.


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Figure 13-11: 2017 Dark Star Column-Leach Gold Extraction vs. Days under Leach

13.6.3 2017 Dark Star Comminution Characterization at HRI

Twelve Dark Star drill core samples were selected for comminution test work. These samples were splits from metallurgical composites and represent major material types. They were subjected to the modified SMC Test at HRI to generate data for SMC parameters; Mic by JKTech and Ai testing was also completed. A final letter report was issued: Comminution Testing, Hazen Project 12391 Report and Appendices A and B – July 5, 2017 (Stepperud, 2017b).

13.6.3.1 2017 Dark Star SMC Test Results

The 2017 Hazen SMC Test® results for the twelve samples are given in the Metallurgical Report (Simmons, 2019, Appendix 33). The table includes the average rock density, A x b and drop-weight index values that are the direct result of the SMC Test® procedure. The values determined for the Mia, Mih, and Mic parameters and the definitions of these abbreviations developed by SMCT are also presented in the table.

13.6.3.2 2017 Dark Star SAG Mill Comminution Test

The drop weight index ranged from 2.57 to 8.53 kWh/m³, indicating soft to medium-hard material, and is tabulated along with other parameters of the SMC evaluation in Appendix 33 (Simmons, 2019). The range of A x b for the 12 composites spanned a low of 30.7 (moderately hard) to a high of 99.4 (soft) and averaged 49.6.


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13.6.3.3 2017 Dark Star Bond Abrasion Index (Ai) Tests

Bond Abrasion index testing was performed at Hazen on 12 Dark Star composite samples. The Metallurgical Report (Simmons, 2019, Appendix 34) lists the Ai values for the 12 composites that were tested. Ai values ranged from a low of 0.2432 g to a high of 1.2381 g, indicating moderate to high abrasiveness of the materials tested. The silica content of the Dark Star mineralized material is the inferred rock component that contributes to the corresponding high Ai test results.

13.6.3.4 2017 Dark Star Comminution Test Summary

The Dark Star comminution samples tested can be considered amenable to conventional, multi-stage crushing and screening circuit design. Mic, the SMC crusher-component value (average = 6.8 kWh/t), would be ranked in the midrange of the SMC worldwide database.

The Ai values are modest to high (average = 0.7864 g) and represent the potential for above average rates of wear on crusher liners, screen panels, and conveyor drop boxes.

13.6.4 2017 Dark Star Load Permeability Testing

A portion of tailings material from twenty-four (24) column-leach test was utilized for load permeability test work. The purpose of the load permeability test work was to examine the permeability of the crushed material under compaction loading equivalent to heap heights of 25, 50, 75, and 100 m.

The test cell utilized for modeling the permeability of stacked material at various heap heights was a steel column or cell. Staged axial (vertical) loading of the test material was utilized to simulate the incrementally increased pressure obtained when loading the heap.

Drainage layers were installed at the top and at the base of the column. External load was applied to the charge of material in the column utilizing a perforated steel plate that moved freely within the walls of the column.

Guidelines that KCA utilizes when reviewing the results from this type of test were listed in Section 13.2.4. The results of the Dark Star load permeability test work are summarized in the Metallurgical Report (Simmons, 2019, Appendix 35).

Twenty of the 24 column residues that were tested passed using KCA’s criteria at all simulated heap heights. One sample failed at the 100 m simulated height and three samples failed at the 25 m simulated height.

The Metallurgical Report (Simmons, 2019, Appendix 36) summarizes geologic information and column-residue screen analysis data for the three column residue samples that failed load permeability testing at the 25 m height. Of specific note, these three column-residue samples had the highest percentage of -200-mesh (75 µm) fines reported in the column residue screen analysis, of all 24 residue samples that were tested, and geologic logging of two of the samples identified appreciable amounts of fault and clay material. It is unknown at this time how much of the total mineral resource tonnage may be represented by these three samples, but it is believed to be minor and it is assumed that this material can be blended during mining and processing.

13.7 2018 GOLD STANDARD DARK STAR HPGR METALLURGICAL TEST WORK

Two Dark Star HPGR composite samples were comprised of selected core samples (Simmons, 2019, Appendix 37) remaining from the 2017 Dark Star bottle-roll and column-leach test program.


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13.7.1 2018 Dark Star HPGR Head Assays

Geological information and head assays for the two 2018 Dark Star HPGR Master Composites are shown in Appendix 37 and 38 (Simmons, 2019).

13.7.2 2018 Dark Star HPGR Composite Bottle-Roll and Column-Leach Tests

The Dark Star HPGR composite samples were subjected to bottle-roll leach testing at target P80 sizes of 38 µm, 75 µm, and 1,700 µm. Column-leach testing was conducted on conventional-crushed material at a P80 of 12.5 mm and on HPGR-crushed samples subjected to low, medium, and high HPGR press forces. The objective of this bottle-roll and column-leach testing was to evaluate the differences in gold extraction, comparing conventional-crush results to HPGR-crush results.

13.7.2.1 2018 Dark Star Bottle-Roll Tests on HPGR Composite Samples

Bottle-roll leach tests were performed on 500 g or 1,000 g portions of head material comminuted to a P80 target size of 1,700 microns (1.70 mm), 75 microns (0.075 mm), and 38 microns (0.038 mm). Bottle-roll tests, wet screening and assay methods were performed with the same procedures outlined in Section 13.6.2. The 2018 bottle-roll results are shown in the Metallurgical Report (Simmons, 2019, Appendix 39).

Bottle-roll cyanide-leach gold extractions are lower for the Dark Star Main composite but appear to be in line with the lower gold head grade. Silver extractions are low for both Dark Star Main and North composites, this is expected for low silver head grades which are of minimal economic significance.

13.7.2.2 2018 Dark Star Column-Leach Tests on HPGR Composite Samples

Column-leach tests were performed on eight HPGR composite-sample charges, four from Dark Star Main and four from Dark Star North, prepared in the following manner sample:

Conventional crush to target P80 = 12.5 mm
HPGR crush at low press force (2.20 N/mm²) setting, P80 = 7,000 µm
HPGR crush at medium press force (3.35 N/mm²) setting, P80 = 6,500 µm
HPGR crush at high press force (4.30 N/mm²) setting, P80 = 5,000 µm

The column tests were leached for 80 days with a dilute sodium cyanide solution, utilizing the same procedures as outlined in Section 13.2.2. Column-leach test results are summarized in the Metallurgical Report (Simmons, 2019, Appendix 40); extractions results are based upon the calculated head derived from the loaded carbon assays + tails assays.

For the Dark Star “Main Master Composite #1” gold extractions ranged from 81% (conventional crush) to 86% (HPGR average, all pressure settings) based upon calculated heads ranging from 0.709 g Au/t to 0.736 g Au/t. Sodium cyanide consumption ranged from 0.76 kg/t to 0.84 kg/t and hydrated-lime consumption ranged from 1.01 kg/t to 1.04 kg/t.

For the Dark Star “North Master Composite #2” gold extractions ranged from 86% (conventional crush) to 91% (HPGR high pressure) based upon calculated heads ranging from 1.20 g Au/t to 1.70 g Au/t. Sodium cyanide consumption ranged from 0.59 kg/t to 0.89kg/t and hydrated lime consumption ranged from 1.00 kg/t to 1.03 kg/t.

A graphical comparison of gold extraction from conventionally-crushed versus HPGR-crushed sample charges, from the Dark Star Main Master Composite #1, is shown in Figure 13-12.


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Figure 13-12: 2018 Dark Star Main - Conventional Crush vs. HPGR Gold Extraction

The green line in Figure 13-12 is provided for benchmarking purposes and represents the original 2017 Phase 1 conventional-crush gold extraction results by weight-averaging the variability composite samples that were included in the Dark Star Main Master Composite #1. The blue line represents conventional-crush gold extraction results on HPGR Master Composite #1 from the 2018 HPGR tests. The magenta triangles represent gold-extraction results for the three HPGR Master Composite #1 column-leach tests at low, medium, and high HPGR press forces.

A graphical comparison of gold extraction from conventionally-crushed versus HPGR-crushed sample charges from the Dark Star North Master Composite #2 is shown in Figure 13-13. The green and blue lines and the magenta triangles represent, respectively: the original 2017, Phase 1, conventional-crush gold-extraction results by weight averaging the variability composite samples that were included in the Dark Star North Master Composite #2, the conventional-crush gold extraction results from HPGR Master Composite #2 in the 2018 HPGR tests, and gold extraction results from the three HPGR Master Composite #2 column-leach tests at low, medium, and high HPGR press forces.


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Figure 13-13: Dark Star North - Conventional Crush vs. HPGR Gold Extraction

The Dark Star Main HPGR column-leach gold extractions are significantly higher than the conventional-crushed column charge. The Dark Star North HPGR gold extractions are only marginally higher than the conventional-crushed composite at similar P80’s. While it is relatively simple to design a flowsheet to produce any specific P80 particle size from conventional crushing, it is not for HPGR comminution. The P80’s shown in Figure 13-12 and Figure 13-13 represent a close approximation to the product size that would be produced in a commercial HPGR comminution circuit.

13.7.3 2018 Dark Star Main & North HPGR-Crushed Load Permeability Testing

All column-leach charges were leached without cement addition or agglomeration for this phase of testing. Column-leach residues were subjected to evaluation of percent slump maximum percolation rate and load permeability tests. Results are shown respectively in Appendix 41, 42, and 43 (Simmons, 2019).

The medium press-force column-leach residue from Dark Star Main HPGR Master Composite #1 failed load permeability testing at all heights. The medium and high press-force column-leach residues from two of the Dark Star North HPGR Master Composite #2 charges failed at all heights. All other column-leach residues passed at all heights tested. It is recommended that future testing continue to evaluate cement agglomeration on HPGR-comminuted samples to support heap heights of at least 50 m, and possibly 75 m.

13.8 2019 GOLD STANDARD DARK STAR DEPOSIT METALLURGICAL TEST WORK

In 2018 Gold Standard commissioned KCA to complete a bottle roll, conventional crush and HPGR crush column leach metallurgical test program on 2017-2018 drill core composite samples from the Dark Star Main and North deposits. Test results are documented in KCA (2019b).


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13.8.1 2019 Dark Star Head Assays for Bottle-Roll and Column-Leach Tests

Head assays and geo-metallurgical characterization analyses were obtained for 50 composites using a combination of four separate laboratories: KCA, ALS, UBC, and FLS. The head assays are tabulated in Appendix 44 through 47 (Simmons, 2019) showing that:

Gold grade ranged from 0.182 to 5.62 ppm and averaged 1.23 ppm.
Silver grade ranged from 0.50 to 3.50 ppm and averaged 1.01 ppm.
Organic carbon ranged from 0.01 to 1.13% (sulfide sample) and averaged 0.16%.
Sulfide sulfur ranged from <0.01 to 0.83% (sulfide sample) and averaged 0.18%.
Preg-robbing analysis ranged from 0.0 to 5.3% and averaged 0.7% (non-preg-robbing).
Copper values were very low, ranging from 9 to 43 ppm and averaged 18 ppm.
Gold cyanide solubility ranged from 33.4% (sulfide sample) to 100% and averaged 83.2%.
Concentrations of deleterious elements by ICP were low: <5 ppm selenium on average, mercury ranged from 0.1.3 to 63.0 ppm (sulfide sample) and averaged 7.8 ppm, and arsenic ranged from 1.1 to 562 ppm with an average of 198 ppm.
Concentrations of the primary cyanide consumers were low and suggest minimum potential for effecting cyanide consumption rates. Copper averaged 18 ppm, nickel averaged 36 ppm, and zinc averaged 131 ppm.
Whole-rock quartz (SiO2) analyses were high, ranging from 51.4 to 93.7% and averaged 88.3%.

13.8.2 2019 Dark Star Bottle-Roll and Column-Leach Tests at KCA

Fifty drill core composites were subjected to bottle-roll leach testing at target P80 sizes of 75 µm and 1,700 µm, conventional crush column-leach testing at crush sizes of 12.5 mm and 25.0 mm and six of the fifty composites were HPGR crushed (at medium press) force and column leached. The main objective of the bottle-roll and column-leach testing was to evaluate laboratory-scale leachability of the Dark Star mineral resource in terms of gold extraction, extraction rate, reagent consumption, sensitivity to feed size, and to evaluate comparative differences between conventional crush and HPGR crush Au recovery.

13.8.2.1 2019 Dark Star Bottle Roll Tests

Bottle-roll leach testing was conducted on portions of material from each of the 50 composites. A 500 or 1,000 g portion of head material was crushed to a nominal size of 1,700 µm (1.70 mm) and utilized for leach testing. A second portion of material was milled in a laboratory rod mill to a target size of 80% passing 75 µm (0.075 mm). The milled slurry was then utilized for leach testing. The tests which are described in detail by the laboratory report (KCA 2019b), employed retention times of 144 hours for the 1,700 µm material and 72 hours for the 75 µm material.

Gold Standard has divided the Dark Star deposit into two zones for metallurgical testing: Dark Star Main and Dark Star North. Dark Star metallurgical core holes are color coded by year in Figure 13-14 (below). The 2017 (green) and 2018 (blue) core holes were used in the 2019 bottle-roll and column leach test work.

Gold and silver extraction results are summarized in Appendix 48(75 µm Bottle Rolls), Appendix 49(1,700 µm Bottle Rolls), Appendix 50 (Conventional Crush Columns), and Appendix 51(HPGR Crush Columns) from the Metallurgical Report (Simmons, 2019). Dark Star zones from which composite sample material originated is shown in Figure 13-10.


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Figure 13-14: Location Map for 2017-8 Dark Star Metallurgical Composites


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The following is a summary of the findings from the 2019 Dark Star bottle roll test results:

13.8.2.2 2019 Dark Star 75 µm (200 Mesh) Bottle-Roll Results

Dark Star 200-mesh bottle-roll gold and silver extraction results are shown in the Metallurgical Report (Simmons, 2019, Appendix 48). Gold head grades for the 200-mesh composite samples ranged from 0.18 to 5.85 ppm Au with an average of 1.19 ppm Au. Gold extraction ranged between 24.7 and 96.1% and averaged 80.8%. Three of the composites were sulfide/carbon refractory with gold cyanide solubility <60%, 20 of the composites were transitional with gold cyanide solubility >60% and <85%, and 27 of the composites were oxide with AuCN solubility >85%.

Silver grades are very low at Dark Star. Silver head grades for the 75 µm composites ranged from 0.31 to 2.81 ppm Ag with an average of 0.87 ppm Ag. Silver extraction ranged from 7.3 to 85.4% and averaged 42.7%. Cyanide consumption averaged 0.90 kg/t and lime consumption averaged 1.08 kg/t.

13.8.2.3 2019 Dark Star 1,700 µm (10 Mesh) Bottle-Roll Results

Dark Star 10-mesh bottle roll gold and silver extraction results are shown in the Metallurgical Report (Simmons, Appendix 49). Gold head grades for the 1,700 µm composite samples ranged from 0.15 to 5.89 ppm Au with an average of 1.15 ppm Au. Gold extraction ranged between 26.0 and 94.5% and averaged 75.0%. Three of the composites were sulfide/carbon refractory with gold cyanide solubility <60%, 20 of the composites were transitional with gold cyanide solubility >60% and <85%, and 27 of the composites were oxide with AuCN solubility >85%.

Silver grades are very low at Dark Star. Silver head grades for the 1,700 µm composites ranged from 0.38 to 2.77 ppm Ag with an average of 1.25 ppm Ag. Silver extraction ranged from 4.3 to 67.1% and averaged 24.9%.

Cyanide consumption averaged 0.59 kg/t and lime consumption averaged 1.27 kg/t.

13.8.2.4 2019 Dark Star 12.5 mm and 25.0 mm Conventional Crush Column Leach Results

Eleven of the 2019 composites were column leached utilizing material crushed to 100% passing 19 mm (target P80 = 12.5 mm), and thirty-nine composites were crushed to 100% passing 37.5 mm (target P80 = 25 mm) for column leach testing. During testing, the material was leached for 66, 95, 98 or 99 days with NaCN solution and placed, respectively, in columns of 100 mm and 150 mm diameters. After leaching, each test was washed for four days with water. A portion of the tailings material from each column-leach test was utilized for tail screen analyses with assays by size fraction. Column-leach gold and silver extraction results are based upon pregnant solution carbon assays and tails screen assays and are summarized in the Metallurgical Report (Simmons, 2019, Appendix 50).

Seven of the columns were agglomerated with 2 kg/t of cement in the laboratory column set-up. Column-leach extraction results were calculated based upon loaded carbon assays and tails screen assays. Calculated gold head grades for the 50 columns ranged from 0.19 to 5.39 ppm Au with an average of 1.32 ppm Au. Gold extraction ranged between 28.4 and 94.7% with an average of 74.5%.

Three of the column-leach composites were sulfide/carbon refractory with gold cyanide solubility <60%. Gold extraction ranged from 28.4 to 39.9% and averaged 35.1%.
Twenty of the column-leach composites were transitional with gold cyanide solubility >60% and <85%. Gold extraction for these columns ranged from 47.5 to 84.7% and averaged 67.2%.
Twenty-seven of the column-leach composites were oxide with gold cyanide solubility >85%. Gold extraction for these columns ranged from 63.3 to 94.7% and averaged 84.4%.


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Calculated silver head grades for the 50 columns ranged from 0.41 to 2.90 ppm Ag with an average of 1.17 ppm Ag. Silver extraction results ranged between 12.0 and 76.1% with an average of 35.5%. Silver head grades for Dark Star are very low and of minimal economic significance.

Cyanide consumption averaged 0.95 kg/t and lime consumption averaged 0.92 kg/t. Commercial scale ROM cyanide consumptions are expected to be in the range of 25 to 33% of laboratory-scale test results. Laboratory lime consumptions are assumed to be similar to commercial-scale consumptions.

Gold extraction versus days under leach for the 50 column-leach tests are shown graphically in Figure 13-15.

13.8.2.5 2019 Dark Star HPGR Crush (Medium Press Force) Column Leach Results

Seven duplicate splits from the fifty 2019 composites were HPGR Crushed using medium press force conditions and column leached under the same conditions as their conventional crush column pairs. After leaching, each test was washed for four days with water. A portion of the tailing material from each column-leach test was utilized for tail screen analyses with assays by size fraction.

Column-leach gold and silver extraction results are based upon pregnant solution carbon assays and tails screen assays and are summarized in the Metallurgical Report (Simmons, 2019, Appendix 51).

Plots of the laboratory column leach gold extractions, for the HPGR and conventionally crushed composites are shown in Figure 13-15.


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Figure 13-15: 2019 Dark Star Column-Leach Gold Extraction vs. Days under Leach

13.8.3 2019 Dark Star Comminution Characterization at HRI

Thirteen of the 2019 Dark Star drill core samples were selected for comminution test work. These samples were splits from metallurgical composites and represent major material types. They were subjected to the modified SMC Test at HRI to generate data for SMC parameters; Mic by JKTech and Ai testing was also completed. A final letter report was issued: Comminution Testing, Hazen Project 12620 Report and Appendices A and B – February 11, 2019 (Stepperud, 2019b).

13.8.3.1 2019 Dark Star SMC Test Results

The 2019 Hazen SMC Test® results for the thirteen samples are given in the Metallurgical Report (Simmons, 2019, Appendix 52). This table includes the average rock density, A x b and drop-weight index values that are the direct result


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of the SMC Test® procedure. The values determined for the Mia, Mih, and Mic parameters, and the definitions of these abbreviations developed by SMCT are also presented in this table.

13.8.3.2 2019 Dark Star SAG Mill Comminution Test

The drop weight index ranged from 2.05 to 9.62 kWh/m³, indicating soft to medium-hard material, and is tabulated along with other parameters of the SMC evaluation in Appendix 52 (Simmons, 2019). The range of A x b for the 12 composites spanned a low of 34.6 (moderately hard) to a high of 123.3 (soft) and averaged 49.3.

13.8.3.3 2019 Dark Star Bond Abrasion Index (Ai) Tests

Bond Abrasion index testing was performed at Hazen on thirteen Dark Star composite samples. Appendix 53 from the Metallurgical Report (Simmons, 2019) lists the Ai values for the 13 composites that were tested. Ai values ranged from a low of 0.0306 g to a high of 1.1656 g, indicating very soft to high abrasiveness of the materials tested. The silica content of the Dark Star mineralized material is the inferred rock component that contributes to the corresponding high Ai test results.

13.8.3.4 2019 Dark Star Comminution Test Summary

The Ai values range from low to high (average = 0.6895 g) and represent the potential for average to above average rates of wear on crusher liners, screen panels and conveyor drop boxes.

13.8.4 2019 Dark Star Load Permeability Testing

A portion of material from fifteen (15) conventionally crushed column-leach residues and four (4) HPGR crushed column residues were utilized for load permeability testing. The purpose of the load permeability test work was to examine the permeability of the crushed material under compaction loading equivalent to heap heights of 25, 50, 75, and 100 m.

Test cell set up and guidelines for interpreting load permeability results have been described earlier in this Technical Report. Refer to Section 13.2.4 for details.

All fifteen conventional crush column residues passed at simulated heap heights up to 100 meters except for the column residue from composite DS17-07 #94, which failed at 75 and 100-meter simulated heap height, using the KCA criteria. See Appendix 54 in the Metallurgical Report (Simmons, 2019) for a summary of the conventional crush load permeability test results.

All four HPGR crush column residues passed at simulated heap heights of 75 meters. Three of the four column residues were agglomerated with 6 kg/t of cement and failed at 100 meters, using the KCA criteria. See Appendix 55 in the Metallurgical Report (Simmons, 2019) for a summary of the HPGR load permeability test results.

13.9 EO-METALLURGY CHARACTERIZATION

13.9.1.1 Pinion Deposit Geo-Metallurgy


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analysis has identified key variables, within both deposits, that were used to select the different metallurgical recovery zones requiring separate gold recovery modeling.

The following is a summary of the four gold and silver recovery zones in the Pinion Deposit:

1.	Mtp (Tripon Pass) – Tripon Pass mineralization is a formation unit that sits on top of the multi-lithic breccia (mlbx) which hosts the majority of the Au mineralization at Pinion.
2.	Mlbx Pinion East (Ba > 4.0%, Hi SiO2) – The Pinion East Zone is carved out of a larger mlbx zone that is characterized by high barium (Ba) > 4.0% and high quartz (SiO2) > 65%.
3.	Mlbx Pinion West – The Pinion West Zone captures all the remaining Pinion mlbx zone of mineralization that is not contained within the Pinion East (Ba > 4.0%, Hi SiO2) zone.
4.	Ddg (Devils Gate) – Devils Gate mineralization is stratigraphically positioned underneath the Pinion mlbx.

13.9.1.2 Dark Star Deposit Geo-Metallurgy

The Dark Star mineralization is hosted in two connected deposits: Dark Star North and Dark Star Main. Dark Star North can be characterized as a relatively high-grade heap leachable deposit, whereas Dark Star Main is lower grade and contains more transitional mineralization. Within both deposits, gold mineralization is mainly contained within three formation units: ST-U (upper siltstone), CGL (middle conglomerate), and ST-L (lower siltstone). Geo-metallurgical evaluations did not detect significant variation in gold recovery based upon the host formation but did identify a significant difference is gold recovery response in local regions of low and high silica Intensity (SI), as logged by the geologists. Silica Intensity (SI) is characterized by the geologists using a scale of 0 to 3, with 0 indicating no (or low) silica and 3 being the highest silica.

Recovery models for silver were not developed for Dark Star because of its low silver contents.

The following is a summary of the four gold recovery zones, in the Dark Star deposit:

1.	Dark Star Main (SI<2.0)
2.	Dark Star Main (SI>2.0)
3.	Dark Star North (SI<2.0)
4.	Nrth Dark Star North (SI>2.0)

13.9.2 Au and Ag Recovery Methodology

The following is a brief description of the methodology used to derive the Pinion and Dark Star gold-recovery models. Four steps were used in developing final Au recovery models for Pinion and Dark Star:

Step 1: Determining the gold extraction for each variability composite using a combination of fine grind/crush bottle rolls and medium/coarse crush column tests.

Step 2: Develop head grade vs. tails grade models to use in final development of the gold recovery equations.

Step 3: Build a database of the laboratory solution: ore (S/O) ratio data at various percentages of total extractable gold. A correction factor is applied to each laboratory S/O ratio data point to scale up the laboratory data to commercial scale. Typical laboratory S/O ratio data is tabulated for the following percentages of total extractable gold: 60%, 70%, 80%, 90%, 95%, and 99%.

Step 4: Incorporating steps 1-3 into final recovery models that reflect commercial scale inefficiencies and deductions for solution losses, plus application of cumulative S/O ratios over the life of the project to predict timing of gold recovery.


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13.9.2.1 Step 1: Determining the Gold Extraction for Each Composite

For each composite sample perform a series of bottle-roll and column-leach tests at a range of reasonable P80 particle sizes, typically 75 um, 1,700 m, 12.5 mm (minimum) or larger column feed sizes;
Determine gold extraction as a percent (%) of the fire assay and cyanide-solubility value at each feed size and plot gold extraction versus particle size on a log-normal graph;
Fit an equation through the data; and
Use the equation to calculate and extrapolate gold extraction to the desired ore-particle size, e.g. ROM size at 150 mm (150,000 microns).

An example log-normal plot for the Dark Star Main composite (MDS) #9C is provided in Figure 13-16.

Figure 13-16: Dark Star Composite #9C:P80 vs. Au Extraction (%)

Gold extractions at various P80 particle sizes are calculated (projected) using the gold fire assay (black) and gold cyanide solubility (green) equations in Figure 13-16. The results for composite MDS #9C are presented in Table 13-6. Only the fire assay gold extraction model equation is used for projection of heap-leach gold recovery for the Pinion and Dark Star mineral resources.


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Table 13-6: Example: Dark Star Composite MDS #9C Gold Extraction


P80		Au Rec	Au Rec
µm	Inches	% of FA	% of AuCN
75		86.2	93.9
1700		85.3	92.9
12,500	0.50	84.7	92.3
25,000	1.0	84.5	92.0
50,000	2.0	84.3	91.8
150,000	6.0	84.0	91.5

Using the example Figure 13-16 and Table 13-6, projected gold extraction will be 84.0% at a ROM size P80 of 150,000 microns (150 mm). Corresponding gold extraction at a P80 of 12,500 microns (12.5 mm) will be 84.7%. Note that the gold extractions shown here have not been corrected for commercial-scale operation inefficiencies, which typically fall in the range of negative 1 to 3% depending upon head grade, solution to ore (“S/O”) application ratios, and process soluble losses.

As an example, the summary of the gold extraction model results derived from the 2017-8 Dark Star Main bottle roll and column-leach testing is presented in Table 13-7.

Table 13-7: Example: Dark Star Gold Extraction Model Results


Comp ID	Zone	Mat'l Type	Modeled Au Extraction %
Comp ID	Zone	Mat'l Type	Calc Hd Au(ppm)	P = 80 75 µ	P = 80 1700µ	P80= 12.5 mm	P = 80 25 mm	P = 50 80 mm	P 80 150 mm
DS16-17 #1(C)	DS Main	Oxide	0.330	82.5	81.1	80.3	80.0	79.7	79.2
DS16-17 #2(C)	DS Main	Oxide	0.911	93.5	91.8	90.7	90.3	89.9	89.3
DS16-17 #3(C)	DS Main	Oxide	0.915	84.4	73.0	65.7	63.2	60.6	56.6
DS16-17 #5(C)	DS Main	Oxide	0.899	92.5	92.5	92.5	92.5	92.5	92.5
DS16-18 #9(C)	DS Main	Oxide	0.877	86.2	85.3	84.7	84.5	84.3	84.0
DS16-18 #11(C)	DS Main	Oxide	1.173	89.8	89.2	88.9	88.7	88.6	88.4

13.9.2.2 Step 2: Head and Tail Grades and Recovery Models

Head grade vs. Tails grade plots and the resulting gold recovery equations are first developed in the absence of consideration for commercial scale inefficiencies and system losses.

Final gold recovery models are derived by applying commercial scale heap leach inefficiencies and system losses to the gold extraction models discussed in the Methodology Section of this Technical Report, which explains the development of the head and tails relationship, S/O Ratio models, commercial scale heap leach inefficiencies, and solution losses.

Using this approach, a head grade and recovery relationship can be developed by incorporating the following components:


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Equation 1:


*Au Rec =*	*HG – TG*	=	*HG – (a x HG + b) + system losses)*	(1)
	HG		HG

Where HG = head grade in ppm (g/t);
TG = tail grade in ppm (g/t);
Where a and b are constants from the tails grade fitted equation. This example assumes a linear equation, but it can also be a natural log, power function, or polynomial equation whichever provides the best fit of the data; and
System losses are the losses due to solution hold up in the heap at the end of economic gold recovery from the heap.

Equation 1 calculates the expected gold recovery that can be achieved from a reasonably well operated heap leach. Under commercial heap-leach conditions, recovery takes place over an extended period of time and must be known as a function of time for planning purposes. To achieve this, the solution:ore ratio (S/O) concept is used.

13.9.2.3 Step 3: Solution to Ore Ratio Models

Determine the % recovery of extractable gold, as a function of time, as represented by the S/O ratio.

Since commercial heap-leach, coarse-particle material sizes, are not normally tested under laboratory conditions, S/O ratio requirements vs. P80 particle size, need to be established to derive a value at the operational coarse-particle or ROM size;
Select percentages of the total extraction that will be used, typically 60%, 70%, 80%, 90%, 95%, and 99% of total gold extraction;
Construct a graph of particle size vs. S/O (normal/normal) and obtain a graphical relationship.
From this, the S/O ratio required at a specific particle size to achieve a target % of total gold extraction can be calculated;
Data points are collected for all percentages of total extraction for all ore types;

These values are graphed on a P80 (particle size in microns or mm) versus S/O ratio plots for the various percentages of total extractable gold content, i.e., 60%, 70%, 80%, 90%, 95%, and 99% (ultimate recovery) of total extractable gold content. An example S/O Ratio plot for achieving 70% recovery of total extractable gold is shown in Figure 13-17.


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Figure 13-17: Example Pinion Feed P80 vs. S/O Ratio Plot, 70% Recovery of Total Extractable Gold

By using the equations from all the plots at 60%, 70%, 80%, 90%, 95%, and 99% (ultimate recovery) of total extractable-gold content, a relationship can be developed to predict the percent recovery of total extractable gold as a function of S/O ratio, for all potential heap-leach feed P80 particle sizes. An example plot is provided in Figure 13-18.

Figure 13-18: Example Pinion Mineral Resource S/O Ratio vs. % Recovery of Total Extractable Gold


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13.9.2.4 Step 4: Bringing the Two Together

To relate recovery as a function of time, the above two parameters (recovery as a function of head grade and recovery as a function of time, S/O Ratio) need to be incorporated together; and
From the % recovery of total extractable gold vs. S/O ratio graphs, a linear or natural log equation can be developed of the format used in Equation 2 to represent the ROM heap-leach option.

Equation 2:


*% of maximum recovery achieved at a given time = d x (S/O) + f (linear shown)*	(2)

The variables d and f are constants from the fitted equation, in Figure 13-18, with d = 0.1385 and f = 0.4514 in the example for a ROM heap leach with P80 = 150,000 microns (150 mm).

To determine the S/O ratio at any given time, the loading schedule, application rates, and application schedules need to be combined and interpreted.

The final recovery equation is then of the format:

Equation 3:


*Recovery % =*	*HG – ((a x HG +b) + system losses) x (d x (S/O) + f)*	(3)
	HG

13.9.3 Pinion Deposit Recovery Models

Metal recovery, head grade vs. tail grade, and S/O ratio models were developed from the data derived from reconstructed historical column-leach testing and from the 2016-2019 metallurgical test programs commissioned by Gold Standard and carried out by KCA.

13.9.3.1 Head Grade vs. Tail Grade Models

Metallurgical data used for developing the overall Pinion mineral resource head-tails gold-recovery relationship and S/O ratio models have been reported in the Pinion deposit metallurgical section in this Technical Report (above) and details can be found in the Appendix tables.

Head grade versus tails grade models were developed for four geo-metallurgy recovery zones, based upon differences in lithology/formation, barium content, silica intensity (SI), and gold recovery response:

Mtp – Tripon Pass Formation;
Ddg – Devils Gate Formation;
Mlbx Pinion West Zone, and
Mlbx Pinion East Zone (Ba >4.0%, Hi SiO2)

The mlbx Pinion West models (and graphs) were developed for multiple crush and ROM P80 feed sizes. The model for ROM P80 = 150,000 microns (150 mm) is shown in Figure 13-19. Similar models were developed for the mlbx Pinion East (Ba>4.0%, Hi SiO2), Mtp and Ddg geo-metallurgy material types.


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Figure 13-19: Pinion Head/Tails Grade

13.9.3.2 Pinion Solution:Ore Ratio Model

The laboratory S/O data reported by KCA was adjusted to 9.0 m to represent commercial-scale ROM heap height. The column-leach data from two bulk-sampling programs designated as composites SBBO and PMZ, tested in 1995 by Cyprus and in 2004 by RSM, were reconstructed and used to assist in projecting S/O ratio requirements at coarser particle size.

Figure 13-20 and Figure 13-21: are plots of heap-leach feed P80 versus S/O ratio required to achieve 70% and 90% recovery of total extractable gold. These plots show a typical progression of the increased S/O ratio required to achieve higher recovery of total extractable gold content at increasing feed size. The fitted straight-line equations obtained from the graphs in Figure 13-20 and Figure 13-21: , with similar graphs plotted for 60%, 80%, 95%, and 99% recovery of total extractable gold, are all used to model S/O ratios for the deposit at any heap-leach feed particle size. The results of this model are shown graphically in Figure 13-22.


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Figure 13-20: P80 vs. S/O Ratio for 70% Recovery of Total Extractable Gold

Figure 13-21: P80 vs. S/O Ratio for 90% Recovery of Total Extractable Gold


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Figure 13-22: Percent Recovery of Pinion Total Extractable Gold vs. Heap-Leach Feed Size

Note: feed sizes shown by symbols are in micrometers.

For a ROM heap leach, 99% recovery of total extractable gold is achieved at S/O ratio = 3.56, as shown in Table 13-8. If crushing were to be used at Pinion, S/O ratios required to recover 99% of total extractable gold would be significantly lower depending upon the degree of particle size reduction.

Table 13-8: Modeled Pinion S/O Ratios at Various P80 Particle Size


% of Extractable	S/O Ratios: P80 Particle Size (microns)
Au	12,500	50,000	100,000	150,000
60%	0.04	0.23	0.48	0.73
70%	0.06	0.35	0.73	1.11
80%	0.12	0.54	1.09	1.65
90%	0.29	0.81	1.50	2.19
95%	0.44	1.14	2.08	3.03
99%	0.85	1.59	2.57	3.56

The ROM (P80 = 150 mm) trend line equation fitted to the data in Table 13-8, and shown graphically in Figure 13-22 above, was used to obtain the linear equation constants d and f to be used in Equation 2: and Equation 3:. In the Pinion case d = 0.250 and f = 0.6798.

Since the commercial scale design criteria for the Pinion ROM heap leach uses a S/O ratio of 3.6 - by default Equation 2 = 0.98 (by assuming a 2.0% loss of solubilized gold in the heap leach that is not recovered). This simplifies Equation 3 as shown in Equation 4 below:

Equation 4:


*Recovery % =*	**HG – ((aHG***^b) + system losses) x 0.98	(4)
	HG


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13.9.3.3 Summary of Pinion Gold and Silver Recovery Models

The heap leach values for the constants a, b, d, and f, which are needed for Equation 4, are summarized in Appendix 56 in the Metallurgical Report (Simmons, 2019) for the Mtp, Ddg, mlbx Pinion West, and mlbx Pinion East (Ba>4.0%, Hi SiO2) material types. The values associated with solution losses in the heap have been estimated to be 0.0008 g/t Au.

Constants in (Simmons, 2019, Appendix 56) were used to calculate commercial-scale gold recovery for a heap leach using ROM P80 = 150,000 microns for all material types. This takes into account the models derived for tails grade, S/O ratio losses, commercial scale inefficiencies, and system solution losses. The gold recovery model predicted results for mlbx Pinion West is shown in Table 13-9. Similar tables were developed for the Mtp, Ddg, and mlbx Pinion East (Ba>4.0%, Hi SiO2) recovery zones.

Table 13-9: mlbx Pinion West Zone - Modeled Head Grade vs. Gold Recovery


Au (g/t)	ROM HL (150 mm)
Au (g/t)	Tail Grade Au(g/t)	Au Rec %
0.100	0.0539	44.3
0.200	0.0930	52.0
0.400	0.1605	58.5
0.600	0.2207	61.8
0.800	0.2767	64.0
1.000	0.3298	65.6
1.500	0.4537	68.3
1.750	0.5121	69.3
2.000	0.5688	70.1
3.000	0.7824	72.4

13.9.3.4 ROM Pinion Gold Recovery Equations (Oxide)

ROM – mlbx Pinion West zone oxide gold recoveries shown in Table 13-9 are presented graphically in Figure 13-23 Final gold model recovery equations are shown for two gold grade ranges: <0.40 g Au/t (natural log function) and >0.40 g Au/t (natural log function). See Table 13-10 for all of the Pinion (Oxide) gold recovery zone equations.

ROM Pinion Silver Recovery Equations (Oxide)

Pinion silver recovery models were developed using the same techniques as for gold recovery and are not shown here for brevity. Refer to Table 13-10 for all of the Pinion (Oxide) silver recovery zone equations.

HPGR Pinion Gold and Silver Recovery Equations (Oxide)

The Pinion HPGR gold and silver recovery equations were developed by testing column pairs, one column feed was prepared using laboratory jaw and/or cone crushing equipment and a second column feed prepared by using laboratory scale HPGR equipment. The differences in gold extraction between the conventionally crushed and HPGR crushed composite samples, after commercial scale modeling, was used to factor the HPGR Au recovery model/equations. See Table 13-11 for the Pinion (Oxide) gold and silver recovery equations using HPGR crushing.


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Figure 13-23: Pinion mlbx Pinion West Zone - Gold Recovery Model

The Pinion ROM and HPGR gold and silver recovery equations tabulated below should be used for mine and process modeling and for economic evaluation.

Table 13-10: ROM Pinion Gold and Silver Recovery Equations (Oxide)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
ROM - mlbx Pinion West	5	=10.203*ln(HG) + 68.038	Au HG < 0.40 g/t
ROM - mlbx Pinion West	6	=6.9059*ln(HG) + 65.295	Au HG ‡ 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	7	=8.3765*ln(HG) + 52.119	Au HG < 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	8	=1.612*ln(HG) + 45.803	Au HG ‡ 0.40 g/t
ROM - Mtp (Tripon Pass)	9	=6.7859*ln(HG) + 66.608	Au HG < 0.40 g/t
ROM - Mtp (Tripon Pass)	10	=1.3059*ln(HG) + 61.492	Au HG ‡ 0.40 g/t
ROM - Ddg (Devils Gate)	11	=11.823*ln(HG) + 81.691	Au HG < 0.40 g/t
ROM - Ddg (Devils Gate)	12	=2.257*ln(HG) + 72.763	Au HG ‡ 0.40 g/t

Geomet Rec Zone	Equation	Silver Recovery, %	Range
ROM-mlbx Pinion West	13	=6.1173*ln(HG) + 0.1598	Ag HG < 6.0 g/t
ROM-mlbx Pinion West	14	=1.5839*ln(HG + 8.264)	Ag HG ‡ 6.0 g/t
ROM-mlbx Pinion East (Ba>4.0%, Hi SiO2)	15	=3.7645*ln(HG) - 1.686	Ag HG < 6.0 g/t
ROM-mlbx Pinion East (Ba>4.0%, Hi SiO2)	16	=0.9747*ln(HG) + 4.8186	Ag HG ‡ 6.0 g/t
ROM - Mtp (Tripon Pass)	17	=0.2926*ln(HG) + 4.2102	All Ag grades
ROM - Ddg (Devils Gate)	18	=8.0095*ln(HG) + 27.973	All Ag grades


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Table 13-11: HPGR Pinion Gold and Silver Recovery Equations (Oxide)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
HPGR - mlbx Pinion West	19	=10.203*ln(HG) + 75.119	Au HG < 0.40 g/t
HPGR - mlbx Pinion West	20	=6.9059*ln(HG) + 2.292	Au HG ‡ 0.40 g/t
HPGR - mlbx Pinion East (Ba>4.0%, Hi SiO2)	21	=8.3765*ln(HG) + 75.119	Au HG < 0.40 g/t
HPGR - mlbx Pinion East (Ba>4.0%, Hi SiO2)	22	=1.612*ln(HG) + 69.023	Au HG ‡ 0.40 g/t
HPGR - Mtp (Tripon Pass)	23	=6.7859*ln(HG) + 79.446	Au HG < 0.40 g/t
HPGR - Mtp (Tripon Pass)	24	=1.3059*ln(HG) + 74.330	Au HG ‡ 0.40 g/t
HPGR - Ddg (Devils Gate)		(High clay, not suitable for HPGR)

Geomet Rec Zone	Equation	Silver Recovery, %	Range
HPGR - mlbx Pinion West	25	=3.043*ln(HG) + 37.246	All Ag grades
HPGR - mlbx Pinion East (Ba>4.0%, Hi SiO2)	26	=3.7645*ln(HG) + 38.289	Ag HG ‡ 6.0 g/t
HPGR - mlbx Pinion East (Ba>4.0%, Hi SiO2)	27	=0.8298*ln(HG) + 43.623	Ag HG < 6.0 g/t
HPGR - Mtp (Tripon Pass)	28	=0.2926*ln(HG) + 19.546	All Ag grades
HPGR - Ddg (Devils Gate)		(High clay, not suitable for HPGR)

The Ddg formation is high clay and is not suitable for HPGR crushing.

13.9.4 Dark Star Main and Dark Star North Deposit Recovery Models

Dark Star head grade vs. tails grade, S/O ratio and metal recovery models were developed using the same methodology applied to the Pinion deposit. Actual data extracted for developing the models were taken from the 2017-2019 metallurgical test programs commissioned by Gold Standard and carried out by KCA.

13.9.4.1 Head Grade vs. Tail Grade Models

Metallurgical data used for developing the overall Dark Star mineral resource head-tails gold-recovery relationship and S/O ratio models have been reported in the Dark Star metallurgical sections of this Technical Report (above) and details can be found in the corresponding Appendices to this Technical Report.

Dark Star head grade vs. tails grade models were developed for four geo-metallurgy recovery zones based upon deposit location and silica Intensity (SI):

Dark Star Main (SI<2.0)
Dark Star Main (SI>2.0)
Dark Star North (SI<2.0)
Dark Star North (SI>2.0)

The Dark Star North (SI<2.0) zone Head Grade vs. Tails Grade models (graphs) was developed for multiple crush size, up to ROM feed sizes but only the ROM P80 = 150,000 microns (150 mm) is shown below in Figure 13-24. Similar models were developed for the Dark Star Main (SI<2.0), Dark Star Main (SI>2.0), and Dark Star North (SI>2.0) gold recovery zones and are not shown for brevity.


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Figure 13-24: Dark Star Head Grade vs. Tails Grade Plot

13.9.4.2 Dark Star Solution: Ore Ratio Model

The laboratory S/O data, reported by KCA, was adjusted to 9.0 m to represent commercial-scale ROM heap height.

Figure 13-25 is a plot of heap-leach feed P80 versus S/O ratio required to achieve 90% recovery of total extractable gold. The largest columns tested were crushed to a target P80 = 25 mm. It is recommended that future test work include columns with larger feed P80 size to develop a more accurate S/O ratio model. The fitted straight-line equations obtained from the graphs in Figure 13-25, along with accompanying graphs plotted for 60%, 70%, 80%, 95%, and 99% recovery of total extractable gold, are all used to model S/O ratio for the Dark Star deposit at any feed particle size. The results of this model are shown graphically in Figure 13-26.


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Figure 13-25: P80 vs. S/O Ratio for 90% Recovery of Total Extractable Gold

Figure 13-26: Percent Recovery of Pinion Total Extractable Gold vs. Heap-Leach Feed Size

For a ROM heap leach, 99% recovery of total extractable gold is achieved at S/O ratio = 3.49, as shown in Table 13-12. If crushing were to be used at Dark Star, S/O ratios required to recover 99% of total extractable gold would be significantly lower, depending upon the degree of particle size reduction.


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Table 13-12: Modeled Pinion S/O Ratios at Various P80 Particle Size


% Extractable Au	S/O Ratios: P80 Particle Size (microns)
% Extractable Au	12,500	25,000	50,000	100,000	150,000
60%	0.03	0.05	0.10	0.20	0.29
70%	0.05	0.09	0.18	0.36	0.54
80%	0.07	0.18	0.39	0.81	1.23
90%	0.18	0.37	0.75	1.50	2.25
95%	0.44	0.68	1.15	2.10	3.05
99%	0.89	1.13	1.60	2.54	3.49
P80 (Inches)	0.5	1.0	2.0	4.0	6.0

The ROM (P80 = 150,000 mm) trend line equation fitted to the data in Table 13-12, and shown graphically in Figure 13-22, was used to obtain the power function equation constants d and f to be used in Equation 2: and Equation 3:. In the Dark Star North (SI<2.0) case d = 0.7717 and f = 0.1935.

Since the design criteria for the Pinion ROM heap leach uses a S/O ratio of 3.5 - by default Equation 2: = 0.98 (assuming a 2.0% commercial scale heap leach inefficiency vs. laboratory testing results). Assuming that the LOM S/O ratio function (d x (S/O) + f ) is never less than 3.5. This simplifies Equation 3: as shown in Equation 4 below:

Equation 4:


*Recovery % =*	**HG – ((aHG***^b) + system losses) x 0.98
	HG

13.9.4.3 Summary of Dark Star Gold Recovery Models

The heap leach values for the constants a, b, d, and f, which are needed for Equation 4, are summarized in the Metallurgical Report (Simmons, 2019, Appendix 57) for Dark Star North (SI<2.0) and Dark Star North (SI>2.0) gold recovery zones. Appendix 58 in the Metallurgical Report (Simmons, 2019) is a summary for the Dark Star Main (SI<2.0) and Dark Star Main (SI>2.0) gold recovery zones. The values associated with solution losses in the heap have been estimated to be 0.0008 ppm (g/t) Au.

Constants (Simmons, 2019, Appendix 57 and 58) were used to calculate commercial-scale gold recovery for a heap leach using ROM P80 = 150,000 microns for all gold recovery zones. This takes into account the models derived for head grade vs. tails grade, S/O ratio, commercial heap inefficiencies, and system solution losses. The gold recovery model predicted results for Dark Star North (SI<2.0) is shown in Table 13-13. Similar tables were developed for Dark Star North (SI>2.0), Dark Star Main (SI<2.0), and Dark Star Main (SI>2.0).


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Table 13-13: Dark Star North (SI<2.0) - Modeled Head Grade vs. Gold Recovery


Au (g/t)	ROM HL (150mm)
Au (g/t)	Tail Grade Au(g/t)	Au Rec %
0.100	0.0172	80.4%
0.200	0.0310	82.4%
0.400	0.0560	84.1%
0.600	0.0791	84.9%
0.800	0.1011	85.5%
1.000	0.1223	85.9%
1.500	0.1728	86.7%
2.000	0.2208	87.1%
3.000	0.3119	87.8%
4.000	0.3986	88.2%
5.000	0.4820	88.5%

ROM Dark Star Silver Recovery Equations (Oxide)

Silver grades are very low in the Dark Star deposit and are considered to have no economic value. Therefore, silver recovery models were not developed for Dark Star.

HPGR Dark Star Gold and Silver Recovery Equations (Oxide)

The Dark Star HPGR gold recovery equations were developed by testing column pairs; one column feed was prepared using laboratory jaw and/or cone crushing equipment and a second column feed prepared by crushed using laboratory scale HPGR equipment. The differences in gold extraction between the conventionally crushed and HPGR crushed composite samples, after commercial scale modeling, was used to develop the HPGR Au recovery model/equations.

13.9.4.4 ROM Dark Star Gold Recovery Equations (Oxide)

ROM – Dark Star North (SI<2.0) oxide gold recovery is shown in Table 13-13 and presented graphically in Figure 13-27. Final gold recovery equations are shown for two gold grade ranges: <0.40 g/t (natural log function) and >0.40 g/t (natural log function).


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Figure 13-27 Dark Star North (SI<2.0) - Gold Recovery Model

ROM and HPGR equations for all Dark Star oxide gold recovery zones are provided in the tables below and should be used for mine and process modeling and for economic evaluation.

Table 13-14: ROM – Dark Star Gold Recovery Equations (Oxide)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
ROM – Dark Star North (SI<2.0)	29	=2.675*ln(HG) + 86.591	Au HG < 0.40 g/t
ROM – Dark Star North (SI<2.0)	30	=1.746*ln(HG) + 85.848	Au HG ‡ 0.40 g/t
ROM – Dark Star North (SI>2.0)	31	=13.047*ln(HG) + 79.527	Au HG < 0.40 g/t
ROM – Dark Star North (SI>2.0)	31	=6.736*ln(HG) + 74.5923	Au HG ‡ 0.40 g/t
ROM – Dark Star Main (SI<2.0)	33	=3.656*ln(HG) + 90.398	Au HG < 0.40 g/t
ROM – Dark Star Main (SI<2.0)	34	=0.559*ln(HG) + 87.606	Au HG ‡ 0.40 g/t
ROM - Dark Star Main (SI>2.0)	35	=2.544*ln(HG) + 78.065	Au HG < 0.40 g/t
ROM - Dark Star Main (SI>2.0)	36	=0.389*ln(HG) + 76.104	Au HG ‡ 0.40 g/t

Silver grade at Dark Star is very low and of no economic importance. Silver recovery was not modelled.


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Table 13-15: HPGR - Dark Star Gold Recovery Equations (Oxide)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
HPGR – Dark Star North (SI<2.0)	37	=2.6752*ln(HG) + 88.622	Au HG < 0.40 g/t
HPGR – Dark Star North (SI<2.0)	38	=1.7497*ln(HG) + 87.881	Au HG ‡ 0.40 g/t
HPGR – Dark Star North (SI>2.0)	39	=13.047*ln(HG) + 83.459	Au HG < 0.40 g/t
HPGR – Dark Star North (SI>2.0)	40	=6.7285*ln(HG) + 78.602	Au HG ‡ 0.40 g/t
HPGR – Dark Star Main (SI<2.0)	41	=3.6583*ln(HG) + 91.795	Au HG < 0.40 g/t
HPGR – Dark Star Main (SI<2.0)	42	=0.5637*ln(HG) + 89.007	Au HG ‡ 0.40 g/t
HPGR - Dark Star Main (SI>2.0)	43	=2.5449*ln(HG) + 83.174	Au HG < 0.40 g/t
HPGR - Dark Star Main (SI>2.0)	44	=0.3879*ln(HG) + 81.244	Au HG ‡ 0.40 g/t

Silver grade at Dark Star is very low and of no economic value. HPGR silver recovery was not modelled.

13.9.5 ROM and HPGR Dark Star Gold Recovery Equations (Transition)

ROM and HPGR gold recovery equations were developed for Dark Star North and Dark Star Main transition zones (mixed oxide and sulfide). These are presented in Table 13-16 and Table 13-17 below. Mine planning used the oxide equations to calculate Ag recovery from transition ores. The approach is conservative because transition silver typically leaches better than oxide silver.

Table 13-16: ROM - Dark Star Gold Recovery Equations (Transition)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
ROM – Dark Star North (SI<2.0)	45	=5.7255*ln(HG) + 69.711	Au HG < 0.40 g/t
ROM – Dark Star North (SI<2.0)	46	=0.8755*ln(HG) + 65.341	Au HG ‡ 0.40 g/t
ROM – Dark Star North (SI>2.0)	47	=3.1157*ln(HG) + 59.947	Au HG < 0.40 g/t
ROM – Dark Star North (SI>2.0)	48	=2.4395*ln(HG) + 59.396	Au HG ‡ 0.40 g/t
ROM – Dark Star Main (SI<2.0)	49	=4.665*ln(HG) + 70.373	Au HG < 0.40 g/t
ROM – Dark Star Main (SI<2.0)	50	=0.7134*ln(HG) + 66.812	Au HG ‡ 0.40 g/t
ROM - Dark Star Main (SI>2.0)	51	=8.8543*ln(HG) + 66.87	Au HG < 0.40 g/t
ROM - Dark Star Main (SI>2.0)	52	=5.8832*ln(HG) + 64.6	Au HG ‡ 0.40 g/t

Table 13-17: HPGR - Dark Star Gold Recovery Equations (Transition)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
HPGR – Dark Star North (SI<2.0)	53	No Data Available	Au HG < 0.40 g/t
HPGR – Dark Star North (SI<2.0)	54	No Data Available	Au HG ‡ 0.40 g/t
HPGR – Dark Star North (SI>2.0)	55	=3.1163*ln(HG) + 69.073	Au HG < 0.40 g/t
	56	=2.43575*ln(HG) + 68.524	Au HG ‡ 0.40 g/t
HPGR – Dark Star Main (SI<2.0)	57	No Data Available	Au HG < 0.40 g/t
HPGR – Dark Star Main (SI<2.0)	58	No Data Available	Au HG ‡ 0.40 g/t
HPGR - Dark Star Main (SI>2.0)	59	=8.8549*ln(HG) + 71.84	Au HG < 0.40 g/t
HPGR - Dark Star Main (SI>2.0)	60	=5.8732*ln(HG) + 69.683	Au HG ‡ 0.40 g/t


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13.9.6 ROM and HPGR Pinion Gold Recovery Equations (Transition)

ROM and HPGR gold recovery equations were developed for Pinion transition zones (mixed oxide and sulfide). These are presented in Table 13-18 and Table 13-19 below. As in the Dark Star transition ores, mine planning used the oxide equations to calculate Ag recovery from transition ores – a conservative approach because transition silver typically leaches better than oxide silver.

Table 13-18: ROM - Pinion Gold Recovery Equations (Transition)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
ROM - mlbx Pinion West	61	=10.203*ln(HG) + 57.371	Au HG ‡ 0.40 g/t
ROM - mlbx Pinion West	62	=6.9059*ln(HG) + 54.627	Au HG < 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	63	=8.3765*ln(HG) + 42.53	Au HG ‡ 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	64	=1.612*ln(HG) + 36.214	Au HG < 0.40 g/t
ROM - Mtp (Tripon Pass)	65	=6.7859*ln(HG) + 58.057	Au HG ‡ 0.40 g/t
ROM - Mtp (Tripon Pass)	66	=1.3059*ln(HG) + 52.941	Au HG < 0.40 g/t
ROM - Ddg (Devils Gate)	67	=11.823*ln(HG) + 53.897	Au HG ‡ 0.40 g/t
ROM - Ddg (Devils Gate)	68	=2.257*ln(HG) + 44.969	Au HG < 0.40 g/t

Table 13-19: HPGR - Pinion Gold Recovery Equations (Transition)


Geomet Rec Zone	Equation	Gold Recovery, %	Range
ROM - mlbx Pinion West	69	=10.203*ln(HG) + 64.368	Au HG ‡ 0.40 g/t
ROM - mlbx Pinion West	70	=6.9059*ln(HG) + 61.624	Au HG < 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	71	=8.3765*ln(HG) + 65.75	Au HG ‡ 0.40 g/t
ROM - mlbx Pinion East (Ba>4.0%, Hi SiO2)	72	=1.612*ln(HG) + 59.434	Au HG < 0.40 g/t
ROM - Mtp (Tripon Pass)	73	=6.7859*ln(HG) + 70.896	Au HG ‡ 0.40 g/t
ROM - Mtp (Tripon Pass)	74	=1.3059*ln(HG) + 65.779	Au HG < 0.40 g/t
ROM - Ddg (Devils Gate)	75	No Data Available	Au HG ‡ 0.40 g/t
ROM - Ddg (Devils Gate)	76	No Data Available	Au HG < 0.40 g/t

13.10 REAGENT CONSUMPTIONS SOUTH RAILROAD PROPERTY

Reagent consumptions and requirements, including cyanide, lime, and cement were estimated by KCA based on metallurgical test work completed to date for the Pinion and Dark Star material. Reagent consumptions are summarized below.

13.10.1 Cyanide

The column leach test cyanide consumptions were studied for the ROM and HPGR crushed Pinion and Dark material and adjusted to provide a basis for the expected field cyanide consumptions. In KCA’s experience, field cyanide consumptions are typically 25% to 50% of observed lab consumptions and have been estimated at 33% of the lab consumptions for this study.

ROM cyanide consumptions have been estimated based on column leach tests at 37.5 mm crush size for the Pinion and Dark Star materials. Because there are no ROM column leach test data available and ROM cyanide consumptions in the field are typically less than crushed ore consumptions, the estimated field cyanide consumptions for the ROM material is considered to be 80% of the crushed material cyanide consumptions. Lab cyanide consumptions for Pinion


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material at 37.5 mm crush ranged from 0.66 kg/t to 1.19 kg/t with an average consumption of 0.85 kg/t. Dark Star lab cyanide consumptions at 37.5 mm crush ranged from 0.46 kg/t to 1.31 kg/t with an average consumption of 0.87 kg/t. Based on this data, field cyanide consumptions are estimated at 0.22 kg/t and 0.23 kg/t for ROM Pinion and Dark Star material, respectively.

Cyanide consumptions for the Pinion and Dark Star HPGR crushed material has been estimated based on the average cyanide consumptions from the HPGR column leach tests; only the “Master Composite” HPGR column leach tests were considered for the Dark Star material as these tests are considered to be most representative. Lab cyanide consumptions for Pinion material ranged from 0.48 kg/t to 0.89 kg/t with an average consumption of 0.67 kg/t. Dark Star lab cyanide consumptions ranged from 0.63 kg/t to 0.89 kg/t with an average consumption of 0.76 kg/t. Based on this data, field cyanide consumptions are estimated at 0.22 kg/t and 0.25 kg/t for Pinion and Dark Star HPGR crushed material, respectively.

13.10.2 Lime

Lime is required for pH control for the ROM and Pinion crushed ore during leaching. Because hydrated lime was utilized in the lab leach tests, the laboratory lime consumptions are adjusted to accurately predict consumptions of quicklime (pebble lime, CaO) in the field. Estimated quicklime consumptions for the Pinion and Dark Star ROM ores are 1.0 kg/t of ore and 0.5 kg/t of ore for Pinion crushed ore.

13.10.3 Cement

Cement is required for heap permeability and pH control during leaching for the Pinion and Dark Star HPGR crushed material. Compacted permeability test work results for HPGR crushed Pinion and Dark Star material is presented in Table 13-20 and Table 13-21. Cement consumption in the field for Pinion and Dark Star HPGR-crushed ores are expected to 2.0 kg/t and 7.0 kg/t, respectively.


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Table 13-20: Pinon Compacted Permeability Test Results


KCA Sample No.	Sample Description	Crush Type	Test Phase	Cement Added, kg/MT	Effective Height, meter	Flow Rate, LpHr/m²	Flow Result Pass/Fail	Saturated Permeability, cm/sec	Incremental Slump, %	Cum. Slump, % Slump	Slump Result Pass/Fail	Overall Pass/Fail
78516	Pin 17-12 140' to 176.5' and Pin 17-13 375' to 522'	HPGR - Low	Primary	0	25	364	Pass	1.0E-02	3%	3%	Pass	Pass
			Stage Load		50	336	Pass	9.3E-03	2%	5%	Pass	Pass
			Stage Load		75	329	Pass	9.1E-03	1%	6%	Pass	Pass
			Stage Load		100	260	Pass	7.2E-03	2%	8%	Pass	Pass
78519	Pin 17-12 140' to 176.5' and Pin 17-13 375' to 522'	HPGR - Med	Primary	0	25	134	Pass	3.7E-03	3%	3%	Pass	Pass
			Stage Load		50	93	Fail	2.6E-03	3%	6%	Pass	Fail
			Stage Load		75	46	Fail	1.3E-03	1%	7%	Pass	Fail
			Stage Load		100	33	Fail	9.2E-04	2%	9%	Pass	Fail
78522	Pin 17-12 140' to 176.5' and Pin 17-13 375' to 522'	HPGR - High	Primary	0	25	38	Fail	1.1E-03	5%	5%	Pass	Fail
			Stage Load		50	3	Fail	8.3E-05	3%	8%	Pass	Fail
			Stage Load		75	0	Fail	0.0E+00	N/A	N/A	N/A	Fail
			Stage Load		100	0	Fail	0.0E+00	N/A	N/A	N/A	Fail
84134 A	PM#51-mlbx-ls/HiBa	HPGR	Primary	2	25	3,976	Pass	0.110	2%	2%	Pass	Pass
			Stage Load		50	984	Pass	0.027	2%	4%	Pass	Pass
			Stage Load		75	461	Pass	0.013	2%	6%	Pass	Pass
			Stage Load		100	293	Pass	0.008	2%	8%	Pass	Pass
84137 A	PM#59-MTP=stmic, stls/LowBa	HPGR	Primary	4	25	1,344	Pass	0.037	3%	3%	Pass	Pass
			Stage Load		50	261	Pass	0.007	4%	7%	Pass	Pass
			Stage Load		75	112	Pass	0.003	2%	9%	Pass	Pass
			Stage Load		100	73	Fail	0.002	1%	10%	Pass	Fail


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Table 13-21: Dark Star Compacted Permeability Test Results


KCA Sample No.	Sample Description	Crush Type	Test Phase	Cement Added, kg/MT	Effective Height, meter	Flow Rate, LpHr/m²	Flow Result Pass/Fail	Saturated Permeability, cm/sec	Incremental Slump, %	Cum. Slump, % Slump	Slump Result Pass/Fail	Overall Pass/Fail
79268	DSMain-HPGR Master Comp #1	HPGR - Low	Primary	0	25	403	Pass	1.1E-02	5%	5%	Pass	Pass
			Stage Load		50	472	Pass	1.3E-02	2%	7%	Pass	Pass
			Stage Load		75	355	Pass	9.9E-03	2%	9%	Pass	Pass
			Stage Load		100	327	Pass	9.1E-03	1%	10%	Pass	Pass
79271	DSMain-HPGR Master Comp #1	HPGR - Med	Primary	0	25	93	Fail	2.6E-03	5%	5%	Pass	Fail
			Stage Load		50	64	Fail	1.8E-03	2%	7%	Pass	Fail
			Stage Load		75	44	Fail	1.2E-03	1%	8%	Pass	Fail
			Stage Load		100	39	Fail	1.1E-03	1%	9%	Pass	Fail
79274	DSMain-HPGR Master Comp #1	HPGR - High	Primary	0	25	314	Pass	8.7E-03	5%	5%	Pass	Pass
			Stage Load		50	221	Pass	6.1E-03	2%	7%	Pass	Pass
			Stage Load		75	181	Pass	5.0E-03	2%	9%	Pass	Pass
			Stage Load		100	151	Pass	4.2E-03	1%	10%	Pass	Pass
79277	DSNorth-HPGR Master Comp #2	HPGR - Low	Primary	0	25	338	Pass	9.4E-03	2%	2%	Pass	Pass
			Stage Load		50	358	Pass	9.9E-03	2%	4%	Pass	Pass
			Stage Load		75	288	Pass	8.0E-03	1%	5%	Pass	Pass
			Stage Load		100	347	Pass	9.6E-03	1%	6%	Pass	Pass
79280	DSNorth-HPGR Master Comp #2	HPGR - Med	Primary	0	25	81	Fail	2.3E-03	6%	6%	Pass	Fail
			Stage Load		50	43	Fail	1.2E-03	1%	7%	Pass	Fail
			Stage Load		75	44	Fail	1.2E-03	1%	8%	Pass	Fail
			Stage Load		100	43	Fail	1.2E-03	1%	9%	Pass	Fail
79283	DSNorth-HPGR Master Comp #2	HPGR - High	Primary	0	25	3	Fail	8.3E-05	8%	8%	Pass	Fail
			Stage Load		50	0	N/A	N/A	N/A	N/A	N/A	N/A
			Stage Load		75	0	N/A	N/A	N/A	N/A	N/A	N/A
			Stage Load		100	0	N/A	N/A	N/A	N/A	N/A	N/A
84446 A	DS17-34 #102	HPGR	Primary	4	25	912	Pass	0.025	1%	1%	Pass	Pass
			Stage Load		50	148	Pass	0.004	2%	3%	Pass	Pass
			Stage Load		75	0	Fail	0.000	2%	5%	Pass	Fail
			Stage Load		100	0	Fail	0.000	N/A	N/A	N/A	Fail
84446 A	DS17-34 #102	HPGR	Primary	6	25	2,187	Pass	0.061	1%	1%	Pass	Pass


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KCA Sample No.	Sample Description	Crush Type	Test Phase	Cement Added, kg/MT	Effective Height, meter	Flow Rate, LpHr/m²	Flow Result Pass/Fail	Saturated Permeability, cm/sec	Incremental Slump, %	Cum. Slump, % Slump	Slump Result Pass/Fail	Overall Pass/Fail
			Stage Load		50	165	Pass	0.005	2%	3%	Pass	Pass
			Stage Load		75	0	Fail	0.000	2%	5%	Pass	Fail
			Stage Load		100	0	N/A	N/A	N/A	N/A	N/A	N/A
84447 A	DS17-22 #104	HPGR	Primary	4	25	8,341	Pass	0.232	0%	0%	Pass	Pass
			Stage Load		50	1,014	Pass	0.028	2%	2%	Pass	Pass
			Stage Load		75	515	Pass	0.014	1%	3%	Pass	Pass
			Stage Load		100	298	Pass	0.008	1%	4%	Pass	Pass
84447 A	DS17-22 #104	HPGR	Primary	2	25	8,429	Pass	0.234	1%	1%	Pass	Pass
			Stage Load		50	3,196	Pass	0.089	1%	2%	Pass	Pass
			Stage Load		75	2,225	Pass	0.062	1%	3%	Pass	Pass
			Stage Load		100	695	Pass	0.019	2%	5%	Pass	Pass
84448 A	DS18-03 #109	HPGR	Primary	4	25	4,951	Pass	0.138	0%	0%	Pass	Pass
			Stage Load		50	422	Pass	0.012	3%	3%	Pass	Pass
			Stage Load		75	199	Pass	0.006	2%	5%	Pass	Pass
			Stage Load		100	105	Fail	0.003	1%	6%	Pass	Fail
84448 A	DS18-03 #109	HPGR	Primary	6	25	1,947	Pass	0.054	1%	1%	Pass	Pass
			Stage Load		50	210	Pass	0.006	3%	4%	Pass	Pass
			Stage Load		75	111	Fail	0.003	1%	5%	Pass	Fail
			Stage Load		100	0	Fail	0.000	3%	8%	Pass	Fail
84449 A	DS18-07 #116	HPGR	Primary	4	25	7,208	Pass	0.200	0%	0%	Pass	Pass
			Stage Load		50	1,934	Pass	0.054	3%	3%	Pass	Pass
			Stage Load		75	999	Pass	0.028	2%	5%	Pass	Pass
			Stage Load		100	589	Pass	0.016	1%	6%	Pass	Pass


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13.11 METALLURGICAL TESTING ON JASPEROID WASH AND NORTH BULLION SAMPLES

13.11.1 Jasperoid Wash Deposit Metallurgical Testing

In 2017, Gold Standard commissioned KCA to complete metallurgical testing of composited core samples from the Jasperoid Wash deposit (KCA 2018c). Drill-core composites were subjected to bottle-roll cyanide-leach testing at target P80 sizes of 75 µm (200 mesh) and 1,700 µm (10 mesh), column-leach testing at eighty percent passing (P80) 12.5 mm, and one column leach testing at P80 = 25 mm. Additionally, three (3) metallurgical core holes were drilled in 2018, from which composites will be tested at a later date. Jasperoid Wash was not included in the current financial model. Accordingly, only a brief summary of the 2017 test results is presented below.

Gold extraction in the 200-mesh bottle rolls ranged from 67.7 and 96.6%, while Silver extraction ranged from 15.6 to 43.1%. Cyanide consumption averaged 0.46 kg/t for the eight oxide composites.

Gold extraction from the 10-mesh bottle rolls ranged between 52.6 and 93.7% (average = 76.1%). Silver extraction ranged from 12.8 to 83.6%. Silver grades are considered low at Jasperoid Wash. Cyanide consumption averaged 0.24 kg/t for the eight oxide samples.

A composite that was classified as sulfide carbon refractory had one of the lowest recoveries and the highest cyanide consumption.

Column-leach gold extractions ranged between 65.3 and 95.3% and averaged 82.9%. Silver head grades for Jasperoid Wash are low and of minimal economic significance. Cyanide consumption averaged 1.01 kg/t and lime consumption averaged 1.23 kg/t for the five oxide composites.

One of the composites, despite being a sulfide/carbon refractory material (AuCN = 38.2%), achieved a high gold extraction of 90.1%. NaCN and lime consumptions were high, 3.12 kg/t and 10.95 kg/t respectively, which is expected due to its high sulfide sulfur content (1.6%).

Other tests on the Jasperoid Wash samples were performed to characterize the comminution, abrasion, and load permeability properties of the materials. Details on these may be found in the metallurgical report (KCA 2018c).

13.11.2 North Railroad Deposits Metallurgical Testing

Two separate preliminary metallurgical tests were performed on the North Bullion (POD deposit) and Bald Mountain areas, which are part of the North Railroad portion of the property (Dufresne et al. 2017b).

In 2006, a total of 63 bottle-roll tests and three column-leach tests were completed by KCA on core material from the POD prospect located in the North Railroad portion of the property (KCA 2006). The results of the 63 individual bottle roll tests were highly variable, yielding gold extractions from 0% to 83%. The high variability of the extraction results was attributed to carbonaceous materials in the samples. The column-leach tests at 1.5, 0.5, and 0.25-inch crusher resulted in an average gold recovery of 85%.

Bench-top roasting tests were conducted by Newmont on North Bullion drill-core samples. Gold recoveries from three calcined samples were 83%, 90%, and 79%, with high lime demands of 15 to 22 lb/ton (Arthur, 2013).

Fourteen agitated cyanide leach tests were performed on samples from one drill hole in the Bald Mountain target.

The average gold recovery attained was 82.2%, with better recoveries resulting from higher-grade samples.


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14 MINERAL RESOURCE ESTIMATES

14.1 INTRODUCTION

The statistical analysis, geological modeling, and mineral resource estimation were performed under the supervision of Mr. Ristorcelli (Pinion and Jasperoid Wash deposits), Mr. Lindholm (Dark Star deposit), and Mr. Dufresne (North Bullion deposits). These estimated mineral resources were classified in order of increasing geological and quantitative confidence into inferred, indicated, and measured mineral resource categories to be in accordance with the “CIM Definition Standards - For Mineral Resources and Mineral Reserves” (2014) and therefore Canadian National Instrument 43-101. CIM mineral resource definitions are given below, with CIM’s explanatory material shown in italics:

The authors of this section report mineral resources at cutoffs that are reasonable for deposits of this nature given anticipated mining methods and plant processing costs, while also considering economic conditions, because of the regulatory requirements that a mineral resource exists “in such form and quantity and of such a grade or quality that it has reasonable prospects for eventual economic extraction.”

14.2 DARK STAR MINERAL RESOURCES

The Dark Star gold mineral resource estimate was completed on May 15, 2019 based on data derived from drilling completed in 2019 up through drill holes DR19-74, DC19-01, DS19-01, and MW19-01. The drill-hole database has an effective date of April 26, 2019. The Dark Star mineral resource estimate has an effective date of August 7, 2019.

References to Tomera Formation equivalent stratigraphy have been noted historically. However, recent work suggests these units in the Railroad-Pinion deposit may not be of equivalent age, so all usage of Tomera Formation equivalent in this Technical Report refer to units that are Pennsylvanian-Permian undifferentiated.

14.2.1 Dark Star Database

Six companies have conducted exploration drilling programs in the Dark Star deposit area since 1984, including Gold Standard, which began drilling in 2015. In all, 425 holes totaling 95,748.9 m have been drilled (see Table 14-1). These drill holes, as well as Gold Standard’s property limits and the Dark Star mineral resource outline, are shown in Figure 14-1. RC drill holes account for 82% of the meters drilled and core holes account for the balance.

Table 14-1: Summary of Drilling at Dark Star


Type of hole	Count	Drilled meters
Core	60	16,851.0
RC	365	78,897.9
Grand Total	425	95,748.9

Table 14-2 presents descriptive statistics of all Dark Star drill-hole analytical sample data audited and imported into MineSight by MDA. Measured density and core geotechnical data are also summarized. Rejected sample assay data have been excluded from the table. Trace element and whole-rock geochemical data have also been provided by Gold Standard but are not shown in Table 14-2.


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Table 14-2: Descriptive Statistics of Sample Assays in Dark Star Drill-Hole Database
(accepted sample data only)


	Valid	Median	Mean	Std. Dev.	C. of Var.	Minimum	Maximum	Units
From	61,049					0.0	944.9	m
To	61,049					0.3	946.4	m
Length	61,049	1.5	1.6			0.0	45.7	m
Au	58,969	0.020	0.236	0.860	3.7	0.003	22.3	g/t
Ag	25,425	0.171	0.307	2.158	7.0	0.000	136.4	g/t
AuCN	12,681	0.250	0.709	1.394	2.0	0.015	22.4	g/t
AuCN/AuFA ratio	12,681	86.0	76.8	24.9	0.3	0.000	100.0	g/t
Specific Gravity	1,023	2.520	2.456	0.247	0.1	1.540	4.5	g/cm³
Core Recovery	6,238	100.000	90.680	20.160	0.2	0.000	409.3	%
RQD	6,238	40.000	50.860	55.150	1.1	0.000	409.3	%


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Figure 14-1: Dark Star Deposit Drill-Hole Map and Mineral Resource Outline


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The Dark Star database contains 58,969 accepted gold assay records (Table 14-2). The total number of rejected gold assays is 425. These records from five Dark Star North RC drill holes were rejected due to suspected down-hole contamination as demonstrated by cyclicity of assay grades relative to depths of drill-rod changes.

Only 25,425 (43%) of the accepted gold assay samples were analyzed for silver, and 12,681 samples (22%) were analyzed for AuCN. Of the silver assays, 21,403 (84%) are repeated values. A few of these could be individual assays with coincidentally the same assay value, but nearly all represent assays of composited samples for which the silver assay was assigned to multiple individual sample intervals. The composites with a single silver value are generally about 6 m long and composed of four samples.

Collar locations, downhole survey data, and gold analyses were audited for verification purposes. Logged core recovery and RQD were loaded into the database but were not verified. A few RQD values greater than 100% were noted, but not investigated. The database also contains logged geologic features, including rock types, formations, faults, vein type, silicification, clay, dolomite, barite, limonite, hematite, carbonate, sulfide percent, and percent reduced (unoxidized), all of which were imported. The logged geology was reviewed and used in modeling the gold domains.

Analyses of various carbon and sulfur species were also provided by Gold Standard, verified, and loaded into the mineral resource database. Metallurgical bottle-roll, column-leach, comminution, specific gravity, and flotation test results were compiled and loaded, but not verified.

14.2.2 Dark Star Geologic Model

Gold Standard provided geologic interpretations as surfaces and solids for faults, formation contacts, silicification, and metallurgically refractive material. MDA combined the formation and fault surfaces into solids that represent each formation, which includes the Chainman Formation (Mississippian), undifferentiated section of Pennsylvanian-Permian units, and Tertiary conglomerates and Indian Well Formation tuffs and sediments. The Pennsylvanian-Permian undifferentiated is further divided into lower siltstone, middle conglomerate (which is the primary host for Dark Star mineralization), and upper siltstone units. All formational units and faults are summarized in Section 7 of this Technical Report. MDA determined that Quaternary colluvium is present in sufficient quantities to be distinguished from heavier bedrock, so it was modeled on section and as solids to potentially improve stripping costs.

MDA reviewed silicification solids provided by Gold Standard. The solids compare well with logged silicification values of ‘2’ and ‘3’ (‘3’ representing the strongest silicification). Continuity in the modeled solids was broadly established by default as a function of the logged data, although continuity was lacking somewhat between sections where silicification was more localized.

All geologic interpretations, in combination with assays and logged data, were used to guide metal domain modeling and to define metallurgical domains.

14.2.3 Dark Star Gold Modeling and Estimation

14.2.3.1 Gold Domain Model

Gold domains defined from sample assay ranges were explicitly modeled on sections spaced 30 m apart, oriented east-west and looking north. Domains were defined based on population breaks on the cumulative probability plot (“CPP”) for all gold data (Figure 14-2). The following grade ranges were identified and used to model gold:

Outer shell domain: ~0.04 g Au/t to ~0.3 g Au/t;
Low-grade domain: ~0.3 g Au/t to ~2.5 g Au/t; and
High-grade domain: >~2.5 g Au/t.


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A higher-grade domain >~10.5 g Au/t was considered, but there was insufficient continuity for modeling, and it would contain less than 0.1% of the assays. Descriptive statistics of assays by the modeled domains are presented in Table 14-3.

Figure 14-2: Cumulative Probability Plot of Dark Star Gold Assays

Table 14-3: Dark Star Descriptive Statistics by Domain
(accepted sample data only)


Outer Shell Gold Domain
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	11,939	1.5	1.5			0.0	45.7	m
Au	11,848	0.088	0.106	0.074	0.69	0.003	2.3	g/t
Au capped	11,848	0.088	0.106	0.074	0.69	0.003	2.3	g/t
AuCN	4,250	0.120	0.125	0.076	0.61	0.015	1.6	g/t
AuCN/AuFA ratio	4,250	80	74	24	0.30	4	100	%
Specific Gravity	194	2.54	2.46	0.23	0.10	1.78	4.00	g/cm3
Core Recovery	1,278	100	89	21	0.23	0	120	%
RQD	1,278	40	49	51	1.03	0	344	%


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Low-Grade Gold Domain
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	8,643	1.5	1.6			0.0	8.6	m
Au	8,613	0.494	0.675	0.538	0.80	0.016	8.2	g/t
Au capped	8,613	0.494	0.675	0.538	0.80	0.016	8.2	g/t
AuCN	6,686	0.400	0.565	0.528	0.93	0.015	7.6	g/t
AuCN/AuFA ratio	6,686	87	78	25	0.30	1	100	%
Specific Gravity	203	2.56	2.54	0.26	0.10	1.72	4.36	g/cm3
Core recovery	1,178	100	92	18	0.19	0	313	%
RQD	1,178	38	55	61	1.13	0	409	%
High-Grade Gold Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	1,469	1.5	1.5			0.3	3.4	m
Au	1,461	3.352	4.266	3.042	0.71	0.155	22.3	g/t
Au capped	1,461	3.352	4.266	3.042	0.71	0.155	22.3	g/t
AuCN	1,339	2.750	3.463	2.848	0.82	0.015	22.4	g/t
AuCN/AuFA ratio	1,339	92	82	26	0.30	3	100	%
Specific Gravity	92	2.56	2.54	0.19	0.07	1.93	3.43	g/cm3
Core Recovery	466	100	93	16	0.18	0	100	%
RQD	466	54	69	69	1.01	0	409	%
Outside Modeled Gold Domains
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	38,998	1.5	1.6			0.0	39.6	m
Au	37,047	0.008	0.016	0.059	3.73	0.003	3.8	g/t
Au capped	37,047	0.008	0.015	0.028	1.92	0.003	0.6	g/t
AuCN	406	0.040	0.113	0.244	2.15	0.015	2.6	g/t
AuCN/AuFA ratio	406	85	71	33	0.50	0	100	%
Specific Gravity	534	2.48	2.41	0.24	0.10	1.54	4.47	g/cm3
Core Recovery	3,316	100	91	21	0.23	0	409	%
RQD	3,316	38	48	52	1.08	0	409	%

MDA reviewed core from DC18-05, DC18-07, and DC18-09 during a site visit on September 18 and 19, 2018 in an effort to determine the geologic characteristics of each domain. Gold Standard staff geologists provided guidance and expertise with respect to the geology of the deposits and the nature of gold mineralization. The following characteristics were observed with respect to gold domains, and mineralization in general:

The middle conglomerate of the Pennsylvanian-Permian undifferentiated (or Tomera Formation age equivalent) is the primary host for mineralization. The upper and lower siltstone units are mineralized as well, but to a lesser degree.
One of the primary characteristics associated with gold grade is the presence and quantity of limonite on fractures.
Gold grade increases with increased fracture permeability (structural preparation).
More porous, coarser-grained sedimentary lithologies tend to be better hosts. Some porous zones were created by decalcification of calcareous sedimentary rocks.


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Gold mineralization is commonly confined between less permeable lithologies, such as argillized fault gouges or stratigraphic horizons.
Grade decreases from relatively coarse-grained rocks in the low-grade domain, to more fine-grained micritic lithologies in the outer-shell domains.
Barite, scorodite, and jarosite were observed at moderate to higher grades, above ~1.0 g Au/t.
Degree of silicification does not seem to be associated with strong gold mineralization. Where rocks are silicified, grades of ~1.0 to 6.0 g Au/t were found in zones of increased limonite on fractures; and
Some pervasive, very fine-grained pyrite was observed with moderate gold grades, particularly in gouge zones.

To summarize, gold mineralization increases with increasing limonite on fractures, and increasing porosity. More favorable porosity is inherent in coarser-grained sedimentary lithologies or developed by structural preparation and/or decalcification. Structural preparation ranges from localized fractures to wider gouge zones, and to broad zones of fractures and stockwork breccias. Silicification and argillic alteration may be indirectly associated with gold grade, i.e. clay can be abundant in structurally deformed zones, but may or may not be related to gold deposition.

As noted in the previous section, geologic logging and interpretations, along with observations of core directly or in photos, were used to guide mineral-domain modeling. Mineral domains were generally drawn parallel to stratigraphic contacts, per guidance from Gold Standard. Gold domains were offset across faults according to sense-of-movement indicated by Gold Standard interpretations. Schematic cross sections in the Dark Star Main zone and Dark Star North zone are given in Figure 14-3 and Figure 14-4, respectively.


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Figure 14-3: Dark Star Main Zone Gold Domains and Geology – Section N4479600

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Figure 14-4: Dark Star North Zone Gold Domains and Geology – Section N4480080


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The relationship between gold mineralization and major faults mapped on the surface or interpreted on section is not well understood. The primary bounding structures of the major horst block are the West and Dark Star faults, although some mineralization does cross the West Fault into the Chainman Formation and appears to terminate against an unrecognized barrier somewhere to the west of the East fault. The Ridgeline and IDK faults are located within the deposits, and gold grades appear to be strongest and more widespread between them.

Some significant gold grades have been intercepted in multiple drill holes extending downward between the Ridgeline and IDK faults in the Dark Star North zone (see Figure 14-4). Gold Standard describes and interprets the mineralization in this area as follows:

“The zone between the Ridgeline and IDK faults appear to be [a] highly brecciated structural corridor. The gold zone follows down between these two faults, but generally has a floor at/near the Conglomerate and underlying ST-L [lower Tomera Formation equivalent siltstone] contact. The contact is likely a chemistry change from high to low carbonate, causing mineralization above the contact, and much weaker below. We suspect both faults are feeders and long term might see a small breccia pipe or feeder along one or both faults to some depth”

The unusual occurrence and precise geometry of mineralization in this deeper area is still not fully understood, however, new drilling between the faults continues to confirm the existence of relatively high gold grades in the zone.

Gold grade decreases in intensity and thickness down-dip and up-dip along stratigraphy from the Ridgeline and IDK faults. The relationship between mineralization in the footwall and hanging wall of the Ridgeline fault in particular is not well understood. In the current model, domains were drawn as if the fault is a hard boundary to mineralization, with no continuity across the fault, although as noted above, Gold Standard suspects the faults may ultimately prove to be feeders. No domains were drawn along the fault, because it is unclear at this time whether gold was deposited along the structure. The IDK fault does not appear to be a barrier to mineralization as significantly in Dark Star North, so domains were drawn more continuously across it.

After gold domain interpretations were completed on 30 m spaced cross sections oriented east-west, the domain interpretations were snapped to drill holes in three dimensions and sliced for modeling on mid-bench level plans. The modeled level plans are spaced at 9 m and are located at the midpoint of each bench. Silver was not modeled or estimated.

14.2.3.2 Gold Sample and Composite Statistics

The modeled gold mineral domains were used to assign codes to drill-hole samples. Quantile plots were made of the coded assays. Potential capping levels for each domain were assessed by identifying the grade above which outlier values occur. Applied capping grades (Table 14-4) were then determined after reviewing the outlier samples on screen with respect to grade and proximity of surrounding samples, geology, general location, and materiality. Descriptive statistics of sample assays by domain were also considered to evaluate the necessity for capping of assays (Table 14-3).

Table 14-4: Dark Star Capping Levels for Gold by Domain


Domain	Capping Grade (g Au/t)
Outer Shell	0.85
High-Grade	None
Low-Grade	None
Outside Domains	0.55


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Once the capping was completed, the drill holes were down-hole composited to 3.05 m intervals honoring domain boundaries. The 3.05 m length was chosen to avoid de-compositing small fractions of the original 5 ft (1.524 m) drilled sample intervals, which represent the vast majority of the sample lengths. Descriptive statistics by domain of the composited database are given in Table 14-5.

Table 14-5: Dark Star Descriptive Composite Statistics by Domain


Outer Shell Gold Domains
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	6,353	3.05	2.83			0.00	3.05	m
Au	6,281	0.093	0.107	0.064	0.60	0.003	1.89	g/t
Au capped	6,281	0.09	0.11	0.06	0.57	0.00	0.85	g/t
AuCN	3,043	0.120	0.121	0.064	0.53	0.015	1.28	g/t
AuCN/AuFA ratio	3,043	80	74	24	0.30	4	110	%
Low-grade Gold Domains
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	4,720	3.05	2.85			0.00	3.05	m
Au	4,715	0.510	0.668	0.468	0.70	0.050	5.66	g/t
Au capped	4,715	0.51	0.67	0.47	0.70	0.05	5.66	g/t
AuCN	3,521	0.410	0.560	0.465	0.83	0.015	5.34	g/t
AuCN/AuFA ratio	3,521	87	78	24	0.30	1	110	%
High-grade Gold Domains
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	787	3.05	2.70			0.00	3.05	m
Au	786	3.476	4.205	2.507	0.60	0.629	20.30	g/t
Au capped	786	3.476	4.205	2.507	0.60	0.629	20.30	g/t
AuCN	715	2.850	3.427	2.396	0.70	0.115	14.83	g/t
AuCN/AuFA ratio	715	92	82	25	0.30	4	110	%
Outside Modeled Gold Domains
	Valid	Median	Mean	Std Dev	C. of V.	Minimum	Maximum	Units
Length	20,119	3.05	2.80			0.00	3.05	m
Au	19,063	0.008	0.016	0.042	2.70	0.003	3.03	g/t
Au capped	19,063	0.008	0.015	0.024	1.59	0.003	0.52	g/t
AuCN	287	0.060	0.108	0.214	1.99	0.015	2.08	g/t
AuCN/AuFA ratio	287	83	72	35	0.50	3	110	%

Correlograms were generated from the composited gold grades to evaluate grade continuity. Correlogram parameters were determined and applied to the kriged estimate, against which the reported inverse distance estimate was compared. The evaluated continuity of grade also contributed to classification of mineral resources. The correlogram results by domain are summarized as follows:


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Outer shell gold domain – The nugget is 45% of the total sill. The first sill is 30% of the total sill with a range of 10 m to 40 m depending on direction. The remaining 25% of the total sill has a range of 110 m to 230 m depending on direction.

Low-grade gold domain – The nugget is 50% of the total sill. The first sill is 35% of the total sill with a range of 20 m to 30 m depending on direction. The remaining 15% of the total sill has a range of 160 m to 280 m depending on direction.

High-grade gold domain – The nugget is 40% of the total sill. The first sill is 40% of the total sill with a range of 10 m to 45 m depending on direction. The remaining 20% of the total sill has a range of 30 m to 60 m depending on direction

14.2.3.3 Gold Estimation

The mineral resource block model is not rotated, and the blocks are 9 m north-south by 9 m vertical by 9 m east-west. Four gold estimates were completed: a polygonal, nearest neighbor, inverse distance, and kriged, with the inverse-distance estimate being reported. All the estimates, excluding the polygonal, were run several times in order to determine sensitivity to estimation parameters, and to evaluate and optimize results. The inverse distance power was three (“ID³”). The model was divided into nine estimation areas (“ESTAR”) to control search anisotropy, orientation, and distances according to the differing geometries of mineralization in each area during estimation. Table 14-6 summarizes the estimation areas and associated search orientations and maximum search distances by domain. Figure 14-5 depicts the spatial relationship of the estimation areas to the drilling and the gold domains.

Table 14-6: Dark Star Estimation Areas, Search-Ellipse Orientations and Maximum Search Distances by Domain


Estimation Area	Search Ellipse Orientation			Maximum Search Distance
Estimation Area	Azimuth	Dip	Rotation	Outer Shell	Low-grade	High-grade	Outside Domains
1	12.5	0	0	250	200	150	50
2	12.5	0	27.5	250	270	150	50
3	12.5	0	52.5	250	220	150	50
4	12.5	0	77.5	200	150	150	50
5	0	0	50	200	150	150	50
6	0	0	0	200	150	150	50
7	0	0	27.5	200	150	150	50
8	0	0	52.5	200	150	150	50
9	0	0	77.5	200	150	150	50
Note: Semi-major search distance = major search distance ÷ 1, 1.5 or 2, and the vertical search distance = major search distance ÷ 4


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Figure 14-5: Dark Star Spatial Relationship Between Estimation Areas and Drill Holes

One estimation pass was run for each domain, up to a maximum anisotropic search distance of 270 m along the major axis. Search ellipse anisotropy varies from 1:1:4 to 1:2:4 (major versus semi-major versus minor axes). Composite-length weighting was applied to all estimation runs. Estimation parameters for each domain are given in Table 14-7.

Table 14-7: Dark Star Estimation Parameters

(for search orientations and maximum distances, see Table 14-6)


Description	Parameter
Outer Shell Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search anisotropies: major/semimajor/minor (vertical)	1 / varies 0.5 to 1 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g/t, distance in m)	None
Low-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search anisotropies: major/semimajor/minor (vertical)	1 / varies 0.5 to 1 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g/t, distance in m)	2.7 / half max search
High-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 4
Search (m): major/semimajor/minor (vertical)	1 / 0.5* / 0.25
Inverse distance power	3
High-grade restrictions (grade in g/t, distance in m)	10.0 / 75


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Outside Modeled Gold Domains
Samples: minimum/maximum/maximum per hole	2 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	2
High-grade restrictions (grade in g/t, distance in m)	0.1 / 9
* - Exception, ESTAR 5 major to semi-major axis search anisotropy is 1

14.2.4 Dark Star Gold Mineral Resources

Mr. Lindholm classified the Dark Star mineral resources giving consideration to confidence in the underlying database, sample integrity, analytical precision/reliability, QA/QC results, and confidence in geologic interpretations. The classification parameters are given in Table 14-8.

Table 14-8: Dark Star Classification Parameters

Measured

In modeled domain, and

*Drill-hole confidence code ≥ 0.9, and

Number of holes ≥ 3, and average distance ≤ 35 m; Or

Number of samples ≥ 3, and closest distance ≤ 15 m

Indicated

In modeled domain and Main area, and

Isotropic distance ≤ 20 m; Or

Number of Samples ≥ 7 and isotropic distance ≤ 60 m; Or

Number of Samples ≥ 4 and average distance ≤ 70 m; Or

Number of Samples ≥ 4 and closest distance ≤ 25 m; Or

Number of Samples ≥ 2 and closest distance ≤ 15 m

In modeled domain and North area, and

Isotropic distance ≤ 20 m; Or

Number of Samples ≥ 7 and isotropic distance ≤ 50 m; Or

Number of Samples ≥ 4 and average distance ≤ 60 m; Or

Number of Samples ≥ 4 and closest distance ≤ 20 m; Or

Number of Samples ≥ 2 and closest distance ≤ 10 m

Measured Reduced to Indicated if:

Metallurgy code indicates refractory or uncategorized material (METC = 100-199)

Measured and Indicated Reduced to Inferred if:

Reduced-class area code 3 and closest distance ≥ 15 m; Or

In modeled domain and closest distance ≥ 30 m and drill-hole confidence code ≤ 0.5*; Or

In Tertiary conglomerates and modeled domain, and closest distance ≥ 30 m

Inferred

In modeled domain that is not Measured or Indicated; Or

All estimated blocks outside modeled domains, and isotropic distance ≤ 20 m**

*Confidence code of '1' assigned to holes drilled by Gold Standard with collar surveys, '0.5' to Gold Standard holes with no collar surveys, and '0' to non-Gold Standard drill holes

**A strong search restriction on composites ≥0.1 g Au/t within this distance (at 9 m) was applied

Holes drilled at the end of 2018 and early 2019 were compared to an unpublished mineral resource block model produced by MDA in September of 2018. The majority of the drill holes that were considered as infill confirmed the September 2018 model, showing at most small increases or decreases in the volume of domains. As a result of the favorable comparisons of more recent drilling to the 2018 block model, confidence was increased, and therefore the amount of Measured and Indicated mineral resource material in the model was increased.


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As described in Table 14-8, the amount of influence that historical data has on a given block decreases confidence in the estimated grade and consequently the classification. For a block to be classified as Measured mineral resources, 90% or more of the estimating composite grades must be derived from Gold Standard data. Similarly, block grades estimated with all composites beyond 30 m based on 50% or more historical data are classified as Inferred mineral resources.

The results of the QA/QC evaluation warrant additional comment. The major project risk shown by these evaluations is that there is no historical QA/QC except for 11 Mirandor drill holes. Consequently, the reliability of pre-Gold Standard data, and therefore model block grades derived predominantly from historical data, is diminished and supports the reduction in classification. Gold Standard did infill drill areas where historical drilling dominated, so the risk is mitigated in these areas.

Due to excessive snow conditions following the 2019 drilling program, none of the newer 2018-2019 drill collars were surveyed. In all, 81 drill holes in the Dark Star database do not have surveyed collars. The assays associated with these holes were assigned confidence codes of 0.5. The net effect for classification is that Measured and Indicated mineral resources beyond 30 m from a composite is reduced to Inferred mineral resource if the block is estimated using a combination of unsurveyed Gold Standard and historical drill holes.

Gold Standard surveyed 69 of the unsurveyed holes after the current mineral resource estimate was completed. Although the model was not updated with the new information due to time constraints, MDA did compare the new hole locations to modeled domains. Only seven holes would have caused more than very small adjustments to domains boundaries, essentially all within the 9 m³model block dimension. Of the few cases where domain changes would extend more than a few blocks, none were significant, and would probably cause only localized changes to pit shapes, if any. In some cases, a volume of mineralized material would be added, in others material would be lost, but the net impact on the mineral resources would likely be negligible.

The exact nature of deep high-grade mineralization protruding down between the Ridgeline and IDK faults in the Dark Star North zone is not completely understood. Gold Standard interprets the zone as a possible breccia pipe or feeder for gold mineralization, although drilling does not yet confirm the hypothesis. However, despite this uncertainty, drilling consistently intersected mineralization in deep Dark Star North, and continues to confirm the presence of relatively high-grade mineralization in the zone.

Greater restrictions were applied to Measured and Indicated mineral resource material in specific areas of the model due to locally limited understanding of geology and/or gold mineralization (excluding Dark Star North area discussed in the previous paragraph), or suspected (but not proven) down-hole contamination. For example, classification was restricted for mineralization associated with deep, isolated intercepts on the West fault.

A small amount of mineralization has been intercepted in drilling near the surface in Tertiary conglomerates at the southwest end of Dark Star Main. Although the mineralization is present in rocks younger than the bulk of the Dark Star deposit, Gold Standard has observed similar occurrences in Tertiary rocks in other areas of the district. No metallurgical test work has been performed on this material, although there are cyanide-soluble assays that provide a measure of gold recovery. The existence and shape of this mineralization has been confirmed in numerous drill holes, but because the exact nature of gold mineralization in Tertiary conglomerates is not understood, Indicated material was limited to within 30 m of a composite.

MDA reported the Dark Star mineral resources at cutoffs that are reasonable for Carlin-type deposits of comparable size and grade. Technical and economic factors likely to influence the requirement “in such form and quantity and of such a grade or quality that it has reasonable prospects for eventual economic extraction” were evaluated using the best judgement of the author responsible for this section of the report. For evaluating the open-pit potential, MDA modeled a series of optimized pits using variable gold prices, mining costs, processing costs, and anticipated


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metallurgical recoveries. MDA used costs appropriate for open-pit mining in Nevada, estimated processing costs and metallurgical recoveries related to heap leaching, and G&A costs. The factors used in defining cutoff grades are based on a gold price of $1,500/oz.

The Dark Star mineral resource estimate is the fully block diluted ID³estimate and is reported at variable cutoffs for open-pit mining. The cutoff for oxidized and transitional material is 0.14 g Au/t, whereas the cutoff for sulfide material is 1.0 g Au/t. No reported sulfide material is classified as Measured mineral resources. Table 14-9 through Table 14-12 present the estimates of the Measured, Indicated, combined Measured, and Indicated and Inferred gold mineral resources within the $1,500/oz Au pits. The breakdown of mineral resources by area and oxidation state is given in Appendix C. Representative cross sections of the gold block model in the Dark Star Main and North zones are given in Figure 14-6 and Figure 14-7, respectively. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 14-9: Dark Star Total In-Pit Gold Mineral Mineral resources – Measured*


Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.100	6,732,000	1.15	249,000
0.120	6,227,000	1.23	247,000
0.140	5,857,000	1.31	246,000
0.160	5,585,000	1.36	245,000
0.180	5,373,000	1.41	243,000
0.200	5,189,000	1.46	243,000
0.220	5,051,000	1.48	241,000
0.240	4,887,000	1.53	240,000
0.260	4,755,000	1.56	239,000
0.280	4,638,000	1.60	238,000
0.300	4,530,000	1.63	237,000
0.400	3,939,000	1.82	230,000
0.500	3,510,000	1.98	224,000
0.600	3,110,000	2.17	217,000
0.700	2,793,000	2.35	211,000
0.800	2,538,000	2.50	204,000
0.900	2,355,000	2.63	199,000
1.000	2,197,000	2.76	195,000

*mineral resources are inclusive of mineral reserves.

Table 14-10: Dark Star Total In-Pit Gold Mineral Resources – Indicated*


Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.100	30,176,000	0.73	706,000
0.120	28,039,000	0.77	698,000
0.140	26,320,000	0.82	690,000
variable	26,860,000	0.78	675,000
0.160	24,929,000	0.85	684,000
0.180	23,735,000	0.89	676,000
0.200	22,668,000	0.92	669,000
0.220	21,728,000	0.95	663,000
0.240	20,803,000	0.98	655,000
0.260	20,083,000	1.01	649,000
0.280	19,354,000	1.03	642,000
0.300	18,644,000	1.06	635,000


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Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.400	15,017,000	1.23	592,000
0.500	11,999,000	1.42	546,000
0.600	9,552,000	1.63	501,000
0.700	7,766,000	1.85	463,000
0.800	6,458,000	2.07	429,000
0.900	5,531,000	2.27	403,000
1.000	4,832,000	2.45	381,000

*mineral resources are inclusive of mineral reserves.

Table 14-11: Dark Star Total In-Pit Gold Mineral Resources - Measured and Indicated*


Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.100	36,908,000	0.80	955,000
0.120	34,266,000	0.86	945,000
0.140	32,177,000	0.90	936,000
variable	32,717,000	0.88	921,000
0.160	30,514,000	0.95	928,000
0.180	29,108,000	0.98	919,000
0.200	27,857,000	1.02	911,000
0.220	26,779,000	1.05	904,000
0.240	25,690,000	1.08	895,000
0.260	24,838,000	1.11	887,000
0.280	23,992,000	1.14	880,000
0.300	23,174,000	1.17	869,000
0.400	18,956,000	1.35	820,000
0.500	15,509,000	1.54	768,000
0.600	12,662,000	1.76	717,000
0.700	10,559,000	1.98	672,000
0.800	8,996,000	2.19	632,000
0.900	7,886,000	2.37	602,000
1.000	7,029,000	2.55	576,000

*mineral resources are inclusive of mineral reserves.

Table 14-12: Dark Star Total In-Pit Gold Mineral Resources – Inferred*


Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.100	2,901,000	0.61	57,000
0.120	2,719,000	0.65	57,000
0.140	2,579,000	0.69	57,000
0.160	2,471,000	0.69	55,000
0.180	2,370,000	0.72	55,000
0.200	2,249,000	0.75	54,000
0.220	2,154,000	0.78	54,000
0.240	2,044,000	0.82	54,000
variable	2,479,000	0.70	56,000
0.260	1,952,000	0.83	52,000
0.280	1,883,000	0.86	52,000
0.300	1,795,000	0.88	51,000
0.400	1,446,000	1.01	47,000


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Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.500	1,148,000	1.19	44,000
0.600	853,000	1.35	37,000
0.700	690,000	1.53	34,000
0.800	571,000	1.69	31,000
0.900	499,000	1.81	29,000
1.000	430,000	2.03	28,000

*mineral resources are inclusive of mineral reserves.


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Figure 14-6: Dark Star Main Zone Gold Domains and Block Model – Section N4479600


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Figure 14-7: Dark Star North Zone Gold Domains and Block Model – Section N4480080


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Although the authors are not experts with respect to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political matters, the authors are not aware of any unusual factors relating to these matters that may materially affect the Dark Star mineral resources as of the effective date of this Technical Report.

14.2.5 Dark Star Cyanide-Soluble Gold and Geo-Metallurgical Models

A cyanide-soluble gold block model was produced to characterize the spatial variability of cyanide solubility of gold at Dark Star. The model was estimated using the ratio of cyanide-soluble gold assays to fire-assay gold contents (“AuCN/AuFA”). These ratios are graphically depicted in the cumulative probability plot in Figure 14-8, were capped at 110% in samples because using data capped at 100% would introduce a low bias in the estimated ratio values. Composites were also not modified, but all estimated values in the block model were capped at 100%. Two distinct AuCN/AuFA ratio populations, separated by a broad gradational zone from 65% to 90% cyanide-solubility, are apparent in the plot.

Figure 14-8: Cumulative Probability Plot of Dark Star AuCN/Au Ratios

AuCN/AuFA ratios were estimated by rock units separately within the Chainman Formation, within each of the lower siltstone, middle conglomerate, and upper siltstone units of the Pennsylvanian-Permian undifferentiated, and within the Tertiary conglomerates. Ratios were not estimated in the post-mineralization Tertiary Indian Well Formation and Quaternary rocks, which contain no gold. ID³methodology was used, and only AuCN/AuFA ratios with fire-assay gold grades >0.05 g Au/t were included in the estimate. Maximum major and semi-major search distances applied were 350 m, with strong anisotropy of 4:1 relative to the minor search axis. Estimated block AuCN/AuFA ratios were capped at 100%.

Refractory solids were modeled by Gold Standard to segregate zones in the deposit for which gold will not likely be extractable by cyanide heap-leach methods. MDA evaluated the solids and determined that they appear reasonable


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compared to AuCN/AuFA ratios, assayed sulfide-sulfur percent, and logged redox and sulfide percentages. Assayed total-sulfur percent correlates moderately well, but there is relatively high total sulfur with correspondingly low sulfide sulfur percent (presumably representing sulfate minerals) outside the refractory solid. The correlation between refractory solids and logged oxide minerals in drill holes is not as good, because there are zones of mixed iron oxide and sulfide material outside the solids that do not represent completely non-refractory material. In summary, the refractory solids represent material that contains little or no oxidation, whereas the areas outside the solids are mixed oxide and sulfide, or predominantly oxidized rock.

As per metallurgical guidance provided in Section 13, unique metallurgical codes were assigned to the block model based on estimated AuCN/AuFA ratios, refractory zones, rock units, and silicification solids (discussed in Section 14.2.2). Cyanide solubilities and refractory zones were used to define the base metallurgical code group, whereas rock units and silicification were used to further sub-divide those groups of codes. Metallurgical codes were assigned as follows:

Sulfide, low gold recovery: AuCN/AuFA ratios less than 60% or greater than 50% of block is in refractory solid; gold recovery is low;
Transitional, moderate gold recovery: AuCN/AuFA ratios between 60% and 85%, moderate gold recovery; and
Oxide, high gold recovery: AuCN/AuFA ratios greater than 85%.

14.2.6 Dark Star Acid-Base Accounting Model and Estimation

An acid-base accounting (“ABA”) block model was produced to characterize the spatial variability of potential acid-generating (“PAG”) or neutralizing potential (“NAG”) for mine planning and handling of mined material. MDA estimated inorganic carbon (“CINO”) and sulfide sulfur (“SSUL”) into this block model, and designated model blocks as either PAG or NAG. All calculations and PAG/NAG designation criteria were provided by Stantec.

Gold Standard provided LECO analyses of carbon and sulfur species for samples that varied between those on original core intervals (1 ft to 6 ft) to RC sample composites (10 ft to 35 ft). Assayed CINO values were used, or the values were converted from assayed CO2%. The relationship between total organic and inorganic carbon was applied as well where necessary. In the data received from Gold Standard, below-detection limit values were substituted for assays below detection. MDA modified the below-detection assays per Stantec guidance, so that carbon species assays were equal to one-half the below-detection value, and sulfur species assays below detection were set to ‘0’.

MDA evaluated CINO and SSUL statistics by rock unit, refractory zone and silicified zone (Table 14-13 and Table 14-14). The statistics in the tables are summarized according to categories chosen for estimation into the block model.


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Table 14-13: Number of Samples and Mean Inorganic Carbon Values for Dark Star Estimation Categories

(by rock unit, zones inside [refractory] or outside [oxide and transitional] refractory solids, and in/out of silicified zones)


Estimation Category	Chainman Formation		Lower Siltstone
Estimation Category	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)
Oxide and Transitional, not silicified	73	0.167	417	0.946
Oxide and Transitional, Silicified	73	0.167	101	0.269
Refractory, not silicified	285	0.796	247	2.351
Refractory, Silicified	285	0.796	22	0.060


Estimation Category	Middle Conglomerate		Upper Siltstone		Tertiary Conglomerates
	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)

Not silicified	961	0.776	490	0.189	118	0.176
Silicified	3,449	0.102	520	0.035	110	0.008


Estimation Category	Indian Wells Formation		Quaternary Alluvium
Estimation Category	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)
All Data	110	0.051	27	0.148

Table 14-14: Number of Samples and Mean Sulfide Sulfur Values for Dark Star Estimation Categories

(by rock unit, zones inside [refractory] or outside [oxide and transitional] refractory solids)


Estimation Category	Chainman Formation		All Tomera Formation eq.		Tertiary Conglomerates
	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)

Oxide and Transitional	73	0.539	5,562	0.093	215	0.194
Refractory	285	2.034	645	0.877	13	0.946


Estimation Category	Indian Wells Formation		Quaternary Alluvium
Estimation Category	# of Samples	Mean Value (%)	# of Samples	Mean Value (%)
All Data	110	0.037	27	0.051

CINO statistics varied systematically by rock unit in combination with silicification for the middle conglomerate, upper siltstone, and Tertiary conglomerate. This correlation is indicative of the inverse relationship between silica and carbonate contents in increasingly altered and mineralized rocks due to silicification and decarbonization. CINO in the lower siltstone showed similar trends, but statistics also indicated differences inside and outside the modeled refractory solids. In the Chainman Formation, which is only locally mineralized, the variability observed was by refractory zone only. SSUL statistics indicated strong relationships by refractory zone within each of the Chainman Shale, all units of the Tomera Formation equivalent together, and the Tertiary conglomerate. No systematic differences were observed in CINO or SSUL for the Indian Well Formation or the Quaternary colluvium, so each was estimated using all respective contained data.

CINO and SSUL were estimated independently into the block model, according to the categories described above.

CPPs for each species estimated were evaluated by category for potential capping of assays, but none was warranted.


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Nearly half the sample composites are 9.144 m (30 ft) in length. Given the model block dimension of 9 m³, and the adverse effect of de-compositing to shorter interval lengths, assay data were composited to 9.14 m.

All estimates were done using the same search orientations and associated estimation areas as applied to the gold estimate (Table 14-6). The maximum search distance applied for both CINO and SSUL estimates was 300 m. Search ellipses were moderately anisotropic, with major, semi-major and minor search distances at 300 m, 240 m, and 120 m, respectively, and inverse distance squared methodology was used. Due to the relatively long composite length, the maximum number of composites, and maximum composites per hole allowed to estimate a block were limited to five and two, respectively. Review of CPP’s did justify search restrictions for a limited number of the estimated CINO categories, which were applied; however, none were necessary for SSUL estimates.

Correlograms were generated to evaluate continuities in the data with respect to distance. These demonstrated reasonable continuity at ranges up to 400 m, depending on rock unit, refractory type, and/or silicification zones. However, the LECO data is not evenly distributed within the deposits. The data at Dark Star Main is relatively well-distributed, but at Dark Star North, data is concentrated in the central portion of the gold mineralization. As a result, there are significant volumes of rock within potentially mined areas, particularly to the east and west of Dark Star North, where data is sparse or absent. Estimated grades of CINO and SSUL in these areas are relatively far from data. To flag model blocks that are at relatively greater distances from data, MDA assigned confidence codes (value of ‘0’) to all estimated blocks with closest composite greater than 180 m away. Because CINO and SSUL were estimated according to different criteria, these codes were assigned separately for each, and a combined code was assigned if either CINO or SSUL confidence codes was ‘0’.

Model blocks were designated as PAG (code of ‘1’) or NAG (code of ‘2’) according to criteria as defined by Stantec. First, acid-neutralizing potential (“ANP”), acid-generating potential (“AGP”), and net neutralizing potential (“NNP”) values were calculated from estimated CINO and SSUL values. Next, PAG/NAG designation was assigned according to criteria for three potential waste-characterization scenarios in Table 14-15.

Table 14-15: PAG/NAG Designation Criteria

PAG/NAG Designation - Scenario 1

Designate as NAG if

NNP ≥ 20 and ANP/AGP ≥ 3

Designate as PAG if

NNP < 20 or ANP/AGP < 3

PAG/NAG Designation - Scenario 2

Designate as NAG if

SSUL ≥ 0.1% and NNP ≥ 20 and ANP/AGP ≥ 3; Or

SSUL < 0.1% and ANP/AGP ≥ 3

Designate as PAG if

SSUL ≥ 0.1%, and NNP < 20 or ANP/AGP < 3; Or

SSUL < 0.1% and ANP/AGP < 3

PAG/NAG Designation - Scenario 3

Designate as NAG if

NNP ≥ 0.92 and ANP/AGP ≥ 0.77

Designate as PAG if:

NNP < 0.92 or ANP/AGP < 0.77

In Dark Star North there are areas in the upper reaches of potentially mineable pits, along the east and west sides, where no CINO or SSUL composite data was within 300 m, and either or both of these species remained un-estimated. As a result, designation as PAG or NAG was not possible using the above criteria. In agreement with Stantec, MDA assigned PAG or NAG designations for each of the three options described by rock unit, based on the PAG/NAG


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designation of adjacent blocks. The assignments were only necessary for blocks in Upper Siltstone, Tertiary conglomerate, and Quaternary colluvium. These assigned designations represent about one percent of the model tonnage within potential pits, nearly all of which is in Dark Star North.

14.2.7 Dark Star Density

Application of density values to the block model is dependent on numerous modeled criteria that have been discussed in various prior sections. There are 1,023 density measurements in the Dark Star database. All samples were measured using the immersion method by an independent laboratory. The values assigned to the model, by rock unit (Section 14.2.2), gold domains (Section 14.2.3), and refractory zone (Section 14.2.5), are summarized in Table 14-16. Spatially, the Dark Star North zone is well represented; however, there is no density data in the northern 200 m of the deposit. The Dark Star Main zone is moderately well-represented, although core holes are somewhat clustered locally so that there are areas with no density data.

Table 14-16: Density Values Applied to the Dark Star Block Model


Formation	Gold Domains	Refractory Zone	Number of Samples	Density (g/cm³)
Chainman Fm	All	All	27	2.45
Tomera Fm eq. - STL	OS and Outside Domains	Out	72	2.28
Tomera Fm eq. - STL	LG and HG	Out	4	2.40
Tomera Fm eq. - STL	OS and Outside Domains	In	167	2.48
Tomera Fm eq. - STL	LG and HG	In	0	2.60
Tomera Fm eq. - CGL	OS and Outside Domains	Out	290	2.40
Tomera Fm eq. - CGL	LG and HG	Out	238	2.50
Tomera Fm eq. - CGL	OS and Outside Domains	In	30	2.42
Tomera Fm eq. - CGL	LG and HG	In	17	2.56
Tomera Fm eq. - STU	OS and Outside Domains	Out	93	2.40
Tomera Fm eq. - STU	LG and HG	Out	20	2.54
Tomera Fm eq. - STU	OS and Outside Domains	In	0	2.47
Tomera Fm eq. - STU	LG and HG	In	14	2.59
Tertiary conglomerates	All	All	29	2.45
Tertiary Indian Well Formation	All	All	20	2.27
Quaternary colluvium	All	All	2	1.90
Formation acronyms: STL - lower siltstone, CGL - middle conglomerate, STU - upper siltstone
Gold Domain acronyms: OS - outer shell, LG - low-grade, HG - high-grade

The middle conglomerate unit of the Pennsylvanian-Permian undifferentiated (possibly Tomera Formation equivalent), the primary host of gold at Dark Star, is well-represented with nearly 600 density samples. There are at least 70 density samples within the outer shell/outside domains in the lower siltstone unit, a secondary host. However, there are only four samples in the low- or high-grade domains of the lower siltstone. Where a low number of density samples (<~20) were measured for a given category, the density values were evaluated and modified using data from units with similar geological characteristics that are based on more density measurements. A density value of 2.45 g/cm³was assigned to the Chainman Formation based on 27 measurements; a similar value was assigned to the same unit for the Pinion deposit, which was based on more measurements.


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Lower densities are associated with clay alteration. However, Gold Standard has indicated that clay zones are not common or pervasive in the Dark Star deposit. Therefore, although there are some density measurements of clay material that have been included in the statistical groupings in Table 14-16; density values that represent clay zones were not assigned locally in the block model. As a result, there are likely some inaccuracies with respect to tonnages in parts of the block model. Significant clay alteration or weathering is considered to be responsible for the variable density values observed for the Tertiary Indian Well Formation. However, drilling is limited in the unit, and although Gold Standard believes the unit consists primarily of unwelded tuffs that are weathered to clays, the clay zones and associated densities cannot be properly represented. Of 20 samples measured, seven density values ranged from 1.73 to 2.07 (presumably clay) and seven between 2.34 and 2.58 (presumably unaltered). A value of 2.27 was assigned to the unit as a whole, based on data localized in one area over the deposit. However, given the actual variability in densities in the formation, and since the rock unit is entirely waste material, local tonnages are probably not well-defined, and total waste tonnes in the model may be overstated.

14.2.8 Discussion of Dark Star Estimated Gold Mineral Resource and Supporting Models

Holes drilled at the end of 2018 and early 2019 were compared to an earlier, unpublished gold block model produced by MDA in September of 2018. Of the drill holes that are considered only as infill, the block model was confirmed in most holes in the Main area. The gold block model was confirmed in the majority of cases at Dark Star North, but to a lesser extent. As a result of the comparisons of drilling done since September 2018 to the older gold block model, confidence in the current model was increased, and the distances for assigning Indicated classification were increased by 5 m to 10 m in the Dark Star Main zone. Distances for Measured mineral resource classification were extended by 5 m in all areas as well.

Dark Star has a long history of exploration drilling dating back to 1984, and consequently there are many drill holes of varying quality and reliability, and with varying amounts of supporting documentation. In all, six companies, including Gold Standard, have performed exploration drilling on the property. Nearly 70% of the holes were drilled by Gold Standard, for which QA/QC procedures were consistently performed. About 75% of the assay certificates exist for all data, and MDA had access to essentially 100% of Gold Standard certificates. There is a lack of documentation for historical drilling, and QA/QC exists for only 11 holes drilled by Mirandor. As a result, classification of the mineral resource was reduced in areas relying predominantly on historical data. Overall, this reduction did not significantly affect the mineral resource because Gold Standard compensated for the lack of confidence by infill drilling in areas that are predominantly defined by historical drilling. However, there are still a few areas, e.g. the southeast part of Dark Star Main, where little or no Gold Standard drilling exists, and classification is consequently lower.

In general, the geology of gold mineralization is well understood. The geometry of mineralized zones is generally well defined, particularly in shallow areas between the Ridgeline and IDK faults in the Dark Star Main and North zones, as well as in the footwall of the Ridgeline fault in the Main zone, where drilling is relatively dense. However, the relationship between mineralization and the Ridgeline fault is not well understood.

Some significant gold grades have been intercepted in multiple drill holes extending downward between the Ridgeline and IDK faults in Dark Star North. Although the geometry and occurrence of this mineralization is not fully understood, drilling has continued to intersect relatively high-grade mineralization in the area. Measured and Indicated mineral resource classification consistent with the bulk of the Dark Star deposit has been applied to most of this deep Dark Star North mineralization.

Classification as Indicated mineral resource was made more restrictive in the deepest zones, where the general depth below the water table and the presence of anomalous cyanide-soluble gold ratios support the possibility of down-hole contamination. Because potential contamination is suspected by some geologists, it remains a risk, which is represented by the slightly stricter classification criteria.


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One obvious association between faults and mineralization is the consistent occurrence of gold along the West fault. Mineralization has been intercepted in drill holes down-dip along this fault and represents potential for additional mineralization at depth.

The cyanide-soluble gold block model appears reasonable in areas with Gold Standard drilling. In some areas, such as where historical drilling is predominant, AuCN assays are lacking and there is less confidence in the block model. Also, the AuCN data has not been verified and lacks QA/QC support. The refractory solids as modeled are sufficient for use in the block model to define refractory material. It is believed that there is enough data to further refine the refractory model by delineating transitional oxide/sulfide from generally completely oxidized material.

The ABA block model estimate is reasonable within data limits, although the estimate may be too smooth because of long 9.14 m composites. This was somewhat offset by limiting the maximum number of composites to estimate a block. Distribution of LECO data in Dark Star Main is reasonable. However, there are significant areas in Dark Star North that are at extreme distances from assayed samples. To help qualify risks relative to distance from data, estimated model blocks >180 m from the nearest LECO composite are flagged with a confidence code. Also, CINO and/or SSUL are not estimated at all in some of areas of the model, and therefore, blocks cannot be designated as PAG or NAG using the criteria applied to the rest of the model. PAG or NAG was assigned to these unclassified blocks according to the designation of the nearest groups of blocks with similar geologic characteristics where CINO and SSUL were both estimated.

For all classified material, MDA’s mineral resource tonnes at 0.14 g Au/t were larger by ~8.5%, gold grade was lower by ~10.5%, and total gold ounces were lower by ~3% compared to the last Dark Star mineral resource estimate reported by Dufresne and Nicholls (2018). Approximately 265 holes, a significant number of which are considered infill and delineation holes, have been drilled since the 2018 mineral resource estimate was published. However, that estimate (in $1,250/oz gold pit, reported at 0.2 g Au/t) and MDA’s reported mineral resources (in $1,500 gold pit, reported at 0.14 g Au/t) are not directly comparable. Still, it has been demonstrated at Dark Star that optimized pits increase in size only incrementally with changes in gold price, generally less than 1% for each $25 increase in the price of gold, so the difference between the $1,250 and $1,500 pits is relatively small. More significantly, the block dimensions were changed from 6 m x 6 m x 6 m in the APEX model, to 9 m x 9 m x 9 m in MDA’s update of this Technical Report. Additional dilution was incorporated in the current mineral resource estimate as a result of using the larger block sizes. MDA performed a bench-height study on composite data to evaluate the potential changes to the mineral resource attributed to the additional dilution with the changed bench height, and showed that, at the cutoff of 0.2 g Au/t, the gold grade would decrease by about 2% and tonnes would increase by about 6%.

In addition to the mineral resources reported herein, there is mineralization that continues beyond and contiguous with the reported mineral resources. The reported mineral resources are pit-constrained and therefore most of the estimated contiguous mineralization outside the pits (tonnes, grade, and ounces) is unreported. That additional mineralization is shown graphically in Figure 14-9.


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Figure 14-9: Dark Star Optimized Pit and Additional Mineralization

Note: dark lines are drill holes; blue solid is the 0.14 g Au/t grade shell; red is the mineral resource pit shell.

Although most drilling performed during the 2018-2019 campaign was infill of the known deposit area, some step-out drilling was conducted. Only small areas of low-grade mineralization were defined by the step-out holes. This was primarily along strike to the north and south, and on the east, up-dip side of Dark Star Main and along the conglomerate/lower siltstone contact within a Tomera Formation equivalent. A small amount of low-grade mineralization was also delineated in the Tertiary conglomerates at the southwest end of Dark Star Main.

The Dark Star deposit has clustered drill data, which lies primarily within the optimized-pit limits where mining would likely take place. This area also contains a large proportion of the highest-grade material, particularly in the Dark Star North zone. Gold grades from clustered data will tend to project into areas with sparse, non-clustered data during estimation, and a large number of block grades are attributed to only a small number of samples. This effect, which was noted to some extent during gold domain model checking, is mitigated somewhat by estimating with ID³rather


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than ID². De-clustering of composite data was not necessary because the majority of the adverse effect in the estimate occurs outside potential open pits and is not part of the reported mineral resource. Also, new drilling in 2018 and 2019 has mitigated the effects of clustered data, although it is still evident.

Significant clay alteration or weathering is likely responsible for the variable density values observed for the Tertiary Indian Well Formation. Of 20 samples measured, seven values ranged from 1.73 to 2.07 (presumably clay) and seven between 2.34 and 2.58 (presumably unaltered). A value of 2.27 was assigned to the unit as a whole, based on data localized in one area over the deposit. Given the variability in densities in the Indian Well Formation, local tonnages of the unit are probably not well-defined. However, Gold Standard believes the unit consists primarily of unwelded tuffs that are weathered to clays, so total waste tonnes in the model may be overstated.

There are likely some inaccuracies with respect to tonnages in parts of the block model associated with clay material. Drilling is limited in the Tertiary Indian Well Formation, and although Gold Standard believes the unit consists primarily of unwelded tuffs that are weathered to clays, the clay zones and associated densities cannot be properly represented. A value representative of relatively unaltered tuffs was assigned to the unit as a whole, however, given the actual variability and potential abundance of clay alteration in the formation, local tonnages are probably not well-defined, and total tonnes in the current gold block model may be overstated.

A minor reduction in classification was applied due to the lack of surveys for 81 drill collars. However, a comparison of newly received surveys of 69 of these collars to the gold domains revealed only a few cases where the location of domain boundaries would change more than the dimensions of one model block. None of these cases were significant, and the net impact on pit shapes and the total mineral resources would likely be negligible. Still, it is recommended that the gold block model be updated when the remaining new drill collars have been surveyed.

14.3 PINION DEPOSIT MINERAL RESOURCES

This Pinion mineral resource estimate is based on data derived from drilling initially completed in 2018, up through drill hole PR18-100. The 2018 drilling database was received on July 5, 2018, and the initial Pinion mineral resource estimate was completed on July 24, 2018. On May 9, 2019, new silver assays were received from Gold Standard. Also, on May 31, 2019, the Pinion database was updated with gold and silver assays from 12 additional drill holes totaling 2,716 m. Gold and silver mineral resources are herein reported.

This Pinion estimate is based on data derived from drilling initially completed in 2018, up through drill hole PR18-100. The 2018 drilling database was received on July 5, 2018, and the initial Pinion mineral resource estimate was completed on July 24, 2018. On May 9, 2019, new silver assays were received from Gold Standard. Also, on May 31, 2019, which is the effective date of the database, the Pinion database was updated with gold and silver assays from 12 additional drill holes totaling 2,716 m. Gold and silver resources are herein reported, and have an effective date of August 7, 2019.

14.3.1 Pinion Database

The Pinion drilling mineral resource database received from Gold Standard and then audited by MDA contains 689 drill holes with 136,430 m of drilling (Table 14-17). That drilling was done by twelve companies since 1981, including Gold Standard, which began drilling in 2014. Of those holes, 87% are RC and 12% are core; the remainder is of unknown type. A drill-hole map is given in Figure 14-10.


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Table 14-17: Drill Holes at Pinion


Hole Type	Count	Drilled meters
Core	66	16,596
RC	611	118,674
Other	12	1,160
Grand Total	689	136,430


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Figure 14-10: Pinion Deposit Drill-Hole Map and Mineral Resource Outline


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Table 14-18 presents descriptive statistics of all usable analytical or measured Pinion drill-hole sample data audited and imported into MineSight by MDA, except for the geochemical trace elements.

The Pinion drill database contains 60,273 assay records, of which 59,635 were accepted and are summarized in Table 14-18. There were 638 records rejected due to suspected down-hole contamination, core recovery of less than 50% or intervals having geology and mineralization that conflicted with surrounding holes. There are fewer silver assays than gold because many previous operators, and in some cases Gold Standard, did not analyze for silver. Besides gold and silver, other elements were analyzed. Logged core recovery and RQD were loaded into the database but were not audited. A few recoveries and RQD values >100% exist. The database also contains logged geologic features, of which rock types, formation, faults, vein type and intensity, silicification, clay, dolomite, barite, limonite, hematite, carbon, sulfide percent, and percent reduced were imported and at least reviewed, if not used in modeling. Only the collar location and downhole survey data, and the gold analyses, were verified.

Table 14-18: Descriptive Statistics - Exploration and Mineral Resource Drill-Hole Database

(accepted sample data only)


	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
FROM	61,087					0.0	777.2	m
To	61,087					0.9	778.8	m
Length	61,087	1.5	1.7			0.0	57.0	m
Au	59,635	0.012	0.144	0.439	3.1	0.000	12.3	g/t
Ag	47,101	0.253	1.242	8.614	6.9	0.000	1531.0	g/t
AuCN	6,132	0.210	0.388	0.595	1.5	0.015	10.8	g/t
AgCN	3,265	0.510	1.593	4.129	2.6	0.001	144.8	g/t
Density	443	2.600	2.578	0.233	0.1	1.750	4.0	g/cm3
Core Rec.*	3,235	98.300	91.260	15.600	0.2	0.000	166.6	%
RQD	3,235	24.600	34.080	35.030	1.0	0.000	204.5	%
*Core recovery and RQD data have not been audited and contain values exceeding 100%.

14.3.2 Pinion Geologic Model

Gold Standard built interpretations for faults, formations, rock units, occurrence of logged barite, silicification, and metallurgically refractive material. MDA made solids of the formation and rock unit surfaces and rebuilt new barite solids. Silicification solids provided by Gold Standard were used to define a silicified zone inside the solids, and zones where silicification is weak or absent outside the solids. These geologic interpretations were used to guide the metal domain modeling and define metallurgical domains.

Gold Standard’s geologic model of the formations defines contacts of the multi-lithic breccia, Sentinel Mountain, Devils Gate, Webb, Chainman, and Tripon Pass formations. Alluvial cover at Pinion is minimal and was not modeled. Fault surfaces were also defined by Gold Standard. Those fault surfaces were used to explain offsets in the formations, guide geometries of the gold domains, and in some cases explain deeper mineralization that may be structurally controlled. These units and faults are summarized in Section 7 of this Technical Report.


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14.3.3 Pinion Gold Modeling and Estimation

14.3.3.1 Gold Domain Model

Gold domains based on sample assays were modeled on cross sections spaced 30 m apart, oriented east-west and looking north. The geologic model guided interpretation and explicit modeling of the gold domains. These domains were defined based on population breaks on cumulative probability plots of the gold assays prior to compositing (Figure 14-11). The following grade ranges were identified and used for the gold domains:

Low-grade gold domain: ~0.04 g Au/t to ~0.3 g Au/t, and
High-grade gold domain >~0.3 g Au/t.

Descriptive statistics are presented in Table 14-19. Core photos, where available, were reviewed, and proved to be helpful in interpretations.

Figure 14-11: Cumulative Probability Plot of Pinion Deposit Gold Assays


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Table 14-19: Pinion Deposit Descriptive Gold Statistics by Domain

(accepted sample data only)


Low-grade Gold Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	6,191	1.5	1.5			0.3	6.1	m
TYPE	6,191					1	9
Au	6,084	0.086	0.114	0.189	1.7	0.000	6.7	g/t
Au capped	6,084	0.086	0.110	0.119	1.1	0.000	1.3	g/t
AuCN	1,403	0.110	0.152	0.300	2.0	0.015	6.4	g/t
AuCN/AuFA ratio	1,403	80	77	30	0.4	2	253	%
Density	57	2.58	2.54	0.19	0.1	1.88	2.79	g/cm3
Core Rec.	382	96	88	20	0.2	0	125	%
RQD	382	40	39	33	0.9	0	125	%
High-grade Gold Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	10,817	1.5	1.5			0.0	4.6	m
TYPE	10,816					1	9
Au	10,597	0.497	0.742	0.849	1.1	0.000	12.3	g/t
Au capped	10,597	0.497	0.742	0.849	1.1	0.000	12.3	g/t
AuCN	3,713	0.340	0.538	0.676	1.3	0.015	10.8	g/t
AuCN/AuFA ratio	3,713	82	77	23	0.3	1	253	%
Density	136	2.62	2.69	0.30	0.1	2.00	4.00	g/cm3
Core Rec.	883	96	88	18	0.2	0	127	%
RQD	883	25	34	34	1.0	0	100	%
Outside Gold Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	44,079	1.5	1.8			0.0	57.0	m
TYPE	44,073					1	9
Au	42,954	0.008	0.020	0.092	4.6	0.000	9.4	g/t
Au capped	42,954	0.008	0.019	0.055	2.9	0.000	0.9	g/t
AuCN	1,016	0.110	0.187	0.406	2.2	0.015	6.6	g/t
AuCN/AuFA ratio	1,016	80	86	55	0.6	1	253	%
Density	250	2.55	2.53	0.18	0.1	1.75	2.88	g/cm3
Core Rec.	1,970	100	93	13	0.1	0	167	%
RQD	1,970	19	33	36	1.1	0	205	%
*Core recovery and RQD data have not been audited and contain values exceeding 100%.

In spite of the CPP plot showing such a prominent domain beginning around 0.02 g Au/t, the low-grade domain was modeled excluding many 0.02 to 0.05 g Au/t samples, particularly beneath the deposit where the boundary of the mineralization is not defined by abrupt grade changes. It is difficult to determine whether or not this deep halo of low-


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grade mineralization is real or due to drilling conditions or both, because the grades are so low. This material, deliberately left outside the modeled domains, is classified as Inferred mineral resource and is estimated with strong restrictions placed on the rare high-grade sample assays found within it. The gold grades are mostly low and sub-economic under current economic conditions.

The high-grade domain greater than ~0.3 g Au/t lies almost exclusively within the multi-lithic breccia. It shows excellent continuity between drill holes, although the continuity of the actual grades within this domain is more variable, and, based on variography studies (Section 14.2.3.2), that continuity ranges from 45 m to 60 m. The highest grades within the high-grade domain are not sufficiently continuous to be explicitly modeled, so such grades are estimated with the rest of this domain. There is little risk in not explicitly modeling these high-grade values because they are not extreme in grade and this domain has a low coefficient of variation (Table 14-3), even without excluding these higher-grade assays. There is high confidence in this zone based on its geologic support and on analytical distributions lying within it. A typical cross section is given in Figure 14-12.

There are some zones of mineralization that seem to follow high-angle structures. The modeled fault surfaces were used to guide definition of high-angle mineralized domains. Because these are poorly defined and poorly understood, these high-angle volumes are classified as Inferred.

A number of holes have significant, often isolated intersections below the multi-lithic breccia contact and within the Devils Gate Formation. The lack of continuity of this mineralization, coupled with the lack of drill density in the Devils Gate requires that this mineralization in almost all cases be projected short distances and be classified as Inferred.

After sectional interpretations were completed, the gold domains were snapped to drill holes and sliced for modeling on north-south-oriented long sections. The long sections are spaced at 9 m, are located at each midblock in the block model, and are perpendicular to the 30 m spaced sections. In June of 2019, the 30 m cross sections and 9 m long sections were updated with new assay data from 12 drill holes. The changes made were generally minor, and a few low- and high-grade domains were extended downdip to the northeast.


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Figure 14-12: Pinion Gold Domains and Geology – Section N4479230


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14.3.3.2	Gold Sample and Composite Statistics

Once the mineral domains were defined and modeled on 30 m spaced cross sections, the domains were used to assign gold-domain codes to drill-hole samples. Quantile plots were made of the coded assays. Capping for each domain was determined by first assessing the grade above which the outliers occur. Then the outlier grades were reviewed on screen to determine materiality, grade, and proximity of the closest samples and general location. Descriptive statistics were generated and considered with respect to capping levels. Capping values were determined for each of the gold domains separately. Capping levels and number of samples capped are presented in Table 14-20.

Table 14-20: Pinion Gold Capping Levels for Gold by Domain


Domain	Number*	g Au/t
Low grade	18	1.3
High grade	none	na
Outside	53	0.9
* Excludes No Use samples (USEG = 1)

Once the capping was completed, the assays were down-hole composited to 3.05 m intervals honoring domain boundaries. The composite length of 3.05 m was selected because the majority of samples are 1.524 m in length, and the slight deviation from an even three meters was done to honor more precisely the conversion from the original interval unit in feet. Descriptive statistics of the composite database are given in Table 14-21.

Table 14-21: Pinion Deposit Descriptive Gold Assay Composite Statistics by Domain


Low-grade Gold Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	3,442	3.05	2.70			0.00	3.05	m
Au	3,407	0.091	0.114	0.169	1.5	0.000	5.8	g/t
Au capped	3,407	0.09	0.11	0.10	0.9	0.00	1.3	g/t
AuCN	953	0.120	0.150	0.246	1.6	0.015	5.5	g/t
AuCN/AuFA ratio	953	80	77	25	0.3	7	253	%
High-grade Gold Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	5,505	3.05	2.86			0.00	3.05	m
Au	5,467	0.530	0.742	0.752	1.0	0.003	11.0	g/t
Au capped	5,467	0.53	0.74	0.75	1.0	0.00	11.0	g/t
AuCN	1,883	0.363	0.532	0.579	1.1	0.015	6.7	g/t
AuCN/AuFA ratio	1,883	82	77	21	0.3	2	199	%
Outside Gold Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	25,698	3.05	2.85			0.00	3.05	m
Au	24,571	0.008	0.020	0.068	3.5	0.000	4.035	g/t
Au capped	24,571	0.008	0.019	0.049	2.6	0.000	0.900	g/t
AuCN	740	0.115	0.167	0.255	1.5	0.015	3.4	g/t
AuCN/AuFA ratio	740	80	82	47	0.6	1	253	%


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Low-grade gold domain – The nugget is 40% of the total sill. The first sill is 85% of the total sill with a range of 7 to 15 m depending on direction. The remaining sill (15%) has a range of around 25 m to 40 m depending on direction.

High-grade gold domain – The nugget is 55% of the total sill. The first sill is 90% of the total sill with a range of 16 to 20 m depending on direction. The remaining sill (10%) has a range of around 45 m to 60 m depending on direction.


14.3.3.3	Gold Estimation

The block model is not rotated, and the blocks are 9 m north-south by 9 m vertical by 9 m east-west.

Four estimates were completed: a polygonal, nearest neighbor, inverse distance, and kriged, with the inverse-distance estimate being reported. The nearest neighbor, inverse distance and kriged estimates were run several times in order to determine sensitivity to estimation parameters, and to evaluate and optimize results. The inverse distance power was three (“ID³”) and four (“ID⁴”) for the low- and high-grade domain estimates, respectively. The model was divided into 11 estimation areas (“ESTAR”) to control search anisotropy, orientation and distances according to the differing geometries of mineralization in each area during estimation. Table 14-22 lists these areas along with the search orientations and the maximum search per area by low-grade and high-grade domains. Figure 14-13 presents the spatial relationship of those estimation areas to the drilling and the gold domains.

Table 14-22: Pinion Estimation Areas


Area	Azimuth	Dip	Rotation	LG-Max Search	HG-Max Search
1	320	0	35	350	350
2	310	0	35	300	300
3	0	0	0	200	200
4	0	0	-20	300	300
5	30	0	-35	250	250
6	320	8	0	150	100
7	340	0	-20	200	150
8	295	0	-40	100	100
9	0	0	10	200	150
10	340	0	-30	200	100
11	15	0	-60	200	150
Note: maximum distance allowable for Indicated is 60 m


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Figure 14-13: Pinion Estimation Areas

One estimation pass was run for each domain ranging up to 350 m along the primary axis with a 4:1 anisotropy (major axis versus minor axis). All estimates and estimation runs weighted the samples by the sample lengths. Estimation parameters are given in Table 14-23.


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Table 14-23: Pinion Gold Estimation Parameters
(for all rotations/dip/tilt values, see Table 14-22)


Domain	Parameter
Low-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search anisotropies: major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g Au/t)	0.3 / 0.5 x max search
High-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	4
High-grade restrictions (grade in g Au/t)	6.0 / 0.66 x max search
Outside Modeled Gold Domains
Samples: minimum/maximum/maximum per hole	2 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	2
High-grade restrictions (grade in g Au/t and distance in m)	0.1 / 9


14.3.4	Pinion Silver Modeling and Estimation

14.3.4.1	Silver Domain Model

Silver domains based on sample assays were modeled on cross sections spaced 30 m apart, oriented east-west and looking north. The geologic model and gold domains guided the explicit modeling of the silver domains. Domains were defined based on population breaks on cumulative probability plots (Figure 14-11). The following grade ranges were identified and used for silver domains:

Low-grade silver domain: ~0.04 g Ag/t to ~2.0 g Ag/t, and
High-grade silver domain >~2.0 g Ag/t.

Descriptive statistics are presented in Table 14-24. Core photos, where available, were reviewed, and proved to be helpful in interpretations.


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Figure 14-14: Cumulative Probability Plot of Pinion Deposit Silver Assays


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Table 14-24: Pinion Deposit Descriptive Silver Statistics by Domain
(accepted sample data only)


Low-grade Silver Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	6,532	1.5	1.5			0.0	6.1	m
TYPE	6,531					1	9
Ag	4,497	0.896	1.116	1.345	1.2	0.000	22.0	g/t
Ag capped	4,497	0.900	1.093	1.106	1.0	0.000	10.0	g/t
AgCN	1,173	0.361	0.487	0.475	1.0	0.007	7.1	g/t
AgCN/AgFA ratio	1,173	40	42	25	0.6	1	253	%
Density	65	2.59	2.61	0.26	0.1	1.88	3.53	g/cm3
Core Rec.	429	94	86	19	0.2	0	127	%
RQD	429	26	34	32	1.0	0	125	%
High-grade Silver Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	7,935	1.5	1.5			0.0	4.6	m
TYPE	7,934					1	9
Ag	5,562	4.203	7.668	23.826	3.1	0.000	1531.0	g/t
Ag capped	5,562	4.200	7.085	8.582	1.2	0.000	60.0	g/t
AgCN	1,126	2.270	3.824	6.316	1.7	0.047	144.8	g/t
AgCN/AgFA ratio	1,126	50	48	15	0.3	1	116	%
Density	101	2.63	2.69	0.29	0.1	2.06	4.00	g/cm3
Core Rec.	622	95	86	20	0.2	0	117	%
RQD	622	3	26	33	1.3	0	100	%
Outside Silver Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	46,620	1.5	1.8			0.0	57.0	m
TYPE	46,615					1	9
Ag	37,042	0.253	0.401	3.196	8.0	0.000	213.4	g/t
Ag capped	37,042	0.250	0.312	0.513	1.6	0.000	4.0	g/t
AgCN	966	0.149	0.438	1.871	4.3	0.001	45.8	g/t
AgCN/AgFA ratio	966	32	36	27	0.8	0	253	%
Density	277	2.56	2.53	0.18	0.1	1.75	2.88	g/cm3
Core Rec.	2,184	100	93	13	0.1	0	167	%
RQD	2,184	32	36	36	1.0	0	205	%
*Core recovery and RQD data have not been audited and contain values exceeding 100%.

Gold Standard re-assayed pulps from the original, un-composited intervals for all samples within the modeled deposit area. The horizontal shift in Figure 14-14 at 0.25 g Ag/t represents an abundance (~19,000) of values at one-quarter of the 1 g Ag/t detection limit of the re-assayed samples. Original silver assays were performed using different analytical procedures at various detection limits of 0.5 g Ag/t or less.

For the most part, silver grades are generally similar in morphology and location to the gold and therefore also the multi-lithic breccia. However, the silver domains swell and pinch or diminish in some areas where gold is continuous. Some low-grade to anomalous silver mineralization exists below the multi-lithic breccia but is modeled as distinct zones in only one area. Elsewhere the drill-hole spacing is too wide to define continuity.


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Figure 14-15: Pinion Silver Domains and Geology – Section N4479230


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14.3.4.2	Silver Sample and Composite Statistics

Once the silver mineral domains were defined and modeled on 30 m spaced cross sections, the domains were used to assign silver-domain codes to drill-hole samples. Quantile plots were made of the coded assays. Capping for each domain was determined by first assessing the grade above which the outliers occur. Then the outlier grades were reviewed on screen to determine materiality, grade and proximity of the closest samples, and general location. Descriptive statistics were generated and considered with respect to capping levels. Capping values were determined for each of the silver domains separately. Capping levels and number of samples capped are presented in Table 14-25.

Table 14-25: Pinion Capping Levels for Silver by Domain


Domain	Number Capped*	g Ag/t
Low grade	4	10
High grade	43	60
Outside	91	4
Excludes No Use samples (USES = 1)

Once the capping was completed, the silver assays were down-hole composited to 3.05 m intervals honoring domain boundaries. The composite length of 3.05 m was selected because the majority of samples are ~1.524 m in length. The slight deviation from an even three meters was done to honor more precisely the conversion from the original interval unit in feet. Descriptive statistics of the composite database are given in Table 14-26.

Table 14-26: Pinion Deposit Descriptive Silver Assay Composite Statistics by Domain


Low-grade Silver Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	26,284	3.05	2.26			0.00	3.05	m
Ag	19,811	0.250	0.334	3.114	9.3	0.000	213.4	g/t
Ag capped	19,811	0.25	0.28	0.33	1.2	0.00	4.0	g/t
AgCN	639	0.185	0.458	1.340	2.9	0.001	26.3	g/t
AgCN/AgFA ratio	639	33	37	25	0.7	0	140	%
High-grade Silver Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	3,506	3.04	1.87			0.00	3.05	m
Ag	2,415	0.951	1.040	0.781	0.8	0.000	12.7	g/t
Ag capped	2,415	0.95	1.03	0.73	0.7	0.00	8.6	g/t
AgCN	674	0.390	0.500	0.433	0.9	0.008	6.4	g/t
AgCN/AgFA ratio	674	40	43	23	0.5	2	253	%
Outside Silver Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	4,014	3.05	2.01			0.00	3.05	m
Ag	2,829	4.654	7.711	17.639	2.3	0.025	770.0	g/t
Ag capped	2,829	4.65	7.11	7.24	1.0	0.03	60.0	g/t
AgCN	1,567	82.000	79.700	15.300	0.2	2.000	199.0	%
AgCN/AgFA ratio	586	49	48	14	0.3	1	93	%
*Core recovery and RQD data have not been audited and contain values exceeding 100%.

Correlograms were built from the composited silver grades in order to evaluate grade continuity. For the combined low-grade and high-grade domains, the nugget is 40% of the total sill. The first sill is 85% of the total sill, with a range of 5 m to 55 m depending on direction. The remaining sill (15%) has a range of around 100 m to 250 m depending on direction.


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14.3.4.3	Silver Estimation

Four estimates were completed for silver as was done for gold: a polygonal, nearest neighbor, inverse distance, and kriged, with the inverse-distance estimate being reported. The nearest neighbor, inverse distance and kriged estimates were run several times in order to determine sensitivity to estimation parameters, and to evaluate and optimize results. ID³and ID⁴for the low and high-grade domain estimates, respectively. The same 11 estimation areas used for gold to control search anisotropy, orientation and distances during estimation were used for silver (Table 14-22). One estimation pass was run for each domain ranging up to 300 m along the primary axis with a 4:1 anisotropy (major axis versus minor axis). Composite assay values were weighted by interval lengths for all silver estimation runs. Estimation parameters are given in Table 14-27.

Table 14-27: Pinion Silver Estimation Parameters
(for all rotations/dip/tilt values, see Table 14-22)


Domain	Parameter
Low-grade Silver Domain
Samples: minimum/maximum/maximum per hole	1 / 9 / 3
Search anisotropies: major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g Ag/t)	3.0 / 0.33 x max search
High-grade Silver Domain
Samples: minimum/maximum/maximum per hole	1 / 9 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	4
High-grade restrictions (grade in g Ag/t)	None
Outside Modeled Silver Domains
Samples: minimum/maximum/maximum per hole	2 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	2
High-grade restrictions (grade in g Ag/t and distance in m)	0.8 / 9


14.3.5	Pinion Density and Silver Resources

Mr. Ristorcelli classified the Pinion mineral resources considering the confidence in the underlying database, sample integrity, analytical precision/reliability, QA/QC results, and confidence in geologic interpretations. The gold classification was applied to the reported mineral resources, including for silver. The classification parameters for gold are given in Table 14-28. Although the author of this section is not an expert with respect to environmental, permitting, legal, title, taxation, socio-economic, marketing or political matters, the author is not aware of any unusual factors relating to these matters that may materially affect the Pinion mineral resources as of the effective date of this Technical Report.

As described in the table, the amount of influence that historical data has on a block affects the classification. For a block to be classified as Measured mineral resources, more than 90% of the sample influence must be derived from Gold Standard data. On the other hand, no block with the closest sample beyond 30 m and entirely based on historical data may be classified as Measured or Indicated mineral resources. Under most circumstances the confidence of a block would be lower if it were based entirely on historical data. However, the drilling is very dense in areas dominated by historical drill holes, the suspect holes and samples have been culled, and the multiple drill campaigns support each


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other. There are also areas where the geology and domains are more speculative, e.g., the northern area and root zones. The classification in these areas, which are defined in the block model using solids, is reduced.

The results of the QA/QC evaluation warrant additional comment. The major project risk shown by these evaluations, and taken into account in mineral resource classification, is that there is no QA/QC for the historical drilling, and some of those data do not have supporting documentation. Consequently, the reliability of those data is based on their corroboration from nearby Gold Standard data, support from 10 other exploration company drilling campaigns, and the rather loose statement that most of the previous exploration companies are well known.

Table 14-28: Pinion Classification Parameters


Measured
Inside a domain	Yes	Yes
Minimum holes	3	N/A
Minimum composites	3	3
Average distance	≤30	N/A
Closest distance	N/A	<10
Gold Standard drill hole influence	≥90%	≥90%
Indicated		or	or	or
Inside a domain	Yes	Yes	Yes	Yes
Minimum holes	3	2	2	1
Minimum composites	7	4	4	2
Closest distance	≤50	NA	≤20	≤10
Average distance	N/A	≤60	N/A	N/A
Inferred		or
Inside a domain	Yes	No*
Minimum composites	N/A	1
Closest distance	N/A	≤20
Measured and Indicated Reduced to Inferred if:		or
Closest distance	≥30	north area; high-angle areas
Gold Standard drill hole influence	≤1%	N/A
*extreme pullbacks are applied on higher grades outside domains

For reporting, the technical and economic factors likely to influence the requirement “reasonable prospects for eventual economic extraction” were evaluated using the best judgement of the author responsible for this section of the report. For evaluating the open-pit potential, MDA modeled a series of optimized pits using variable gold prices. MDA used costs appropriate for open-pit mining in Nevada, estimated processing costs and metallurgical recoveries related to heap leaching, and G&A costs. The cutoff grades are based on a gold price of $1,500/oz.

The reported Pinion mineral resource estimate is the fully block diluted ID³and ID⁴estimate. The blocks are 9 m cubed. The mineral resources are reported at a cutoff of 0.14 g Au/t for open-pit mining. No sulfide mineralization is reported at Pinion. Table 14-29 to Table 14-32 present the estimated Measured, Indicated, and Inferred gold and silver mineral


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resources at Pinion within the $1,500/oz Au pits. The breakdown of mineral resources by area and oxidation state is given in Appendix C. Representative cross sections of the gold and silver block models are shown in Figure 14-16 and Figure 14-17, respectively. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Table 14-29: Pinion Measured Gold and Silver Resources*


Cutoff
g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.100	1,368,000	0.61	27,000	4.98	219,000
0.120	1,341,000	0.63	27,000	5.06	218,000
0.140	1,304,000	0.64	27,000	5.15	216,000
0.160	1,274,000	0.66	27,000	5.25	215,000
0.180	1,239,000	0.65	26,000	5.32	212,000
0.200	1,197,000	0.65	25,000	5.43	209,000
0.220	1,139,000	0.68	25,000	5.60	205,000
0.240	1,089,000	0.71	25,000	5.77	202,000
0.260	1,044,000	0.72	24,000	5.81	195,000
0.280	996,000	0.75	24,000	5.90	189,000
0.300	959,000	0.78	24,000	6.00	185,000
0.400	765,000	0.89	22,000	6.30	155,000
0.500	585,000	1.01	19,000	6.75	127,000
0.600	460,000	1.15	17,000	7.30	108,000
0.700	363,000	1.20	14,000	7.71	90,000
0.800	303,000	1.33	13,000	7.70	75,000
0.900	247,000	1.51	12,000	7.93	63,000
1.000	205,000	1.67	11,000	8.19	54,000

*mineral resources are inclusive of mineral reserves.

Table 14-30: Pinion Indicated Gold and Silver Resources*


Cutoff
g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.100	29,496,000	0.55	524,000	4.00	3,797,000
0.120	28,518,000	0.57	521,000	4.09	3,754,000
0.140	27,621,000	0.58	517,000	4.18	3,713,000
0.160	26,751,000	0.60	512,000	4.27	3,673,000
0.180	25,959,000	0.61	508,000	4.36	3,637,000
0.200	25,112,000	0.62	503,000	4.45	3,593,000
0.220	24,295,000	0.64	497,000	4.54	3,545,000
0.240	23,502,000	0.65	492,000	4.62	3,488,000
0.260	22,774,000	0.66	486,000	4.69	3,434,000
0.280	21,838,000	0.68	478,000	4.77	3,350,000
0.300	21,015,000	0.70	470,000	4.85	3,274,000
0.400	16,897,000	0.78	424,000	5.23	2,841,000
0.500	13,176,000	0.87	370,000	5.54	2,348,000
0.600	9,859,000	0.98	311,000	5.80	1,840,000
0.700	7,318,000	1.10	259,000	5.96	1,402,000
0.800	5,421,000	1.22	213,000	5.94	1,035,000
0.900	4,131,000	1.33	177,000	6.04	802,000
1.000	3,103,000	1.46	146,000	6.11	610,000

*mineral resources are inclusive of mineral reserves.


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Table 14-31 Pinion Measured and Indicated Gold and Silver Resources*


Cutoff
g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.100	30,864,000	0.56	551,000	4.05	4,016,000
0.120	29,859,000	0.57	548,000	4.14	3,972,000
0.140	28,925,000	0.58	544,000	4.22	3,929,000
0.160	28,025,000	0.60	539,000	4.32	3,888,000
0.180	27,198,000	0.61	534,000	4.40	3,849,000
0.200	26,309,000	0.62	528,000	4.49	3,802,000
0.220	25,434,000	0.64	522,000	4.59	3,750,000
0.240	24,591,000	0.65	517,000	4.67	3,690,000
0.260	23,818,000	0.67	510,000	4.74	3,629,000
0.280	22,834,000	0.68	502,000	4.82	3,539,000
0.300	21,974,000	0.70	494,000	4.90	3,459,000
0.400	17,662,000	0.79	446,000	5.28	2,996,000
0.500	13,761,000	0.88	389,000	5.59	2,475,000
0.600	10,319,000	0.99	328,000	5.87	1,948,000
0.700	7,681,000	1.11	273,000	6.04	1,492,000
0.800	5,724,000	1.23	226,000	6.03	1,110,000
0.900	4,378,000	1.34	189,000	6.15	865,000
1.000	3,308,000	1.48	157,000	6.24	664,000
*mineral resources are inclusive of mineral reserves.

Table 14-32 Pinion Inferred Gold and Silver Resources


Cutoff
g Au/t	Tonnes	g Au/t	oz Au	g Ag/t	oz Ag
0.100	11,607,000	0.61	228,000	3.64	1,358,000
0.120	11,254,000	0.62	226,000	3.72	1,345,000
0.140	10,810,000	0.64	224,000	3.80	1,322,000
0.160	10,452,000	0.66	222,000	3.89	1,306,000
0.180	10,091,000	0.68	221,000	3.97	1,289,000
0.200	9,769,000	0.70	219,000	4.04	1,270,000
0.220	9,454,000	0.71	216,000	4.12	1,252,000
0.240	9,134,000	0.73	214,000	4.20	1,234,000
0.260	8,816,000	0.75	212,000	4.27	1,211,000
0.280	8,512,000	0.76	209,000	4.33	1,186,000
0.300	8,203,000	0.78	206,000	4.38	1,156,000
0.400	6,830,000	0.87	191,000	4.64	1,019,000
0.500	5,624,000	0.96	173,000	4.81	870,000
0.600	4,525,000	1.05	153,000	4.89	711,000
0.700	3,609,000	1.16	135,000	4.88	566,000
0.800	2,871,000	1.27	117,000	4.84	447,000
0.900	2,187,000	1.39	98,000	4.69	330,000
1.000	1,697,000	1.52	83,000	4.60	251,000


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Figure 14-16: Pinion Gold Domains and Block Model– Section N4479230


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Figure 14-17: Pinion Silver Domains and Block Model– Section N4479230


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14.3.6	Pinion Geo-Metallurgical Model

Three additional models, collectively called the Pinion geo-metallurgy model, were produced based on guidelines given from metallurgical test work and interpretations presented in Section 13: barium concentration (estimated within modeled domains), cyanide-soluble gold grade (estimated by rock units), and refractory material (modeled solids).


14.3.6.1	Pinion Barium Modeling and Estimation

The occurrence of barite and silicification seems to have significant impacts on gold recoveries. Consequently, a barium concentration (in lieu of barite) model was necessary for assigning gold recoveries to the deposit. The estimation of a silicification block model was also considered, but the available qualitative and logged geologic data was determined to be insufficient. There was no correlation demonstrated in a comparison of SiO2 assays from metallurgical composites and the relatively larger XRF data set.

Metallurgical testing of drill samples included the ED-XRF-E5 method of analysis for barium; there are 938 analyses of this type that were performed at AAL on press powder pulp material. In addition, 21,747 NITON XRF analyses of barium were taken by independent contractor Rangefront Geological on loose powder pulp material. A significant low bias was noted in the NITON XRF compared to the ED-XRF-E5 analyses (Section 12.6.6) but since there are substantially more NITON XRF values, the larger data set was chosen for modeling and estimation. MDA developed a regression equation to factor the 938 ED-XRF-E5 measurements to NITON XRF equivalents and merge them with the 21,747 NITON XRF barium analyses as follows:

NITON XRFeq = 0.5682 x ED-XRF-E5

The R² for this equation is 0.96 but there are only 32 samples from which the relationship was built. After estimation into the geo-metallurgical block model, the estimated NITON XRF barium grades were refactored to ED-XRF-E5 equivalents to be comparable to the metallurgical data, using the following equation:

ED-XRF-E5eq = 1.760 x NITON XRF

There is a total of 22,653 samples analyzed for barium by either NITON XRF, ED-XRF-E5, or by both methods, which compares to 57,738 gold samples. All NITON XRF barium analyses were plotted in a cumulative probability plot (Figure 14-18) and was used to define domains. No values factored from ED-XRF-E5 analyses are included on the plot. The resulting high-grade (>~6% Ba) and low-grade (~0.3 to ~6% Ba) domains were then modeled on 30 m spaced east-west sections, as was done for gold and silver. The geologic model was the primary guide for barium domain modeling. Barium is spatially related to the multi-lithic breccia, which is generally tabular and folded into the Pinion anticline. Gold domains correlate reasonably well with the barium domains and were used as guides as well. The high-grade domain is spatially restricted to the with a north- to northwest-trending axis of the Pinion anticline, and within the mineral resource pit. Logged barite intensity data was used to augment the analytical barium data in supporting the domain interpretations. Sectional interpretations were then snapped to drill holes and sliced to north-south sections on every 9 m mid-block in the block model.


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Figure 14-18: Cumulative Probability plot of Barium (NITON XRF) Sample Grades at Pinion

Descriptive statistics of the sample barium grades by domain are given in Table 14-33. These samples were composited to 3.05 m lengths. Descriptive statistics of the composited barium grades by domain are given in Table 14-34. A representative cross section showing geology and barium domains is given in Figure 14-19. Estimation parameters are presented in Table 14-35.

Table 14-33: Pinion Samples Barium Statistics by Domain


Low-grade Barium Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	4,294		1.49			0.34	15.1	%
Ba	4,294	0.36	0.83	1.19	1.4	0.00	14.0	%
Ba Capped	4,294	0.36	0.76	0.90	1.2	0.00	3.5	%
High-grade Barium Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	259		1.35			2.45	52.2	%
Ba	259	8.32	9.41	4.55	0.5	1.19	29.3	%
Ba Capped	259	8.32	8.98	3.49	0.4	1.19	15.0	%
Outside Barium Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	17,325		1.52			0.16	2.6	%
Ba	17,325	0.07	0.16	0.52	3.3	0.00	18.1	%
Ba Capped	17,325	0.07	0.15	0.35	2.3	0.00	4.0	%


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Table 14-34: Pinion Composites Barium Statistics by Domain


Low-grade Barium Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	2,273		2.80			0.0	3.1	m
Ba	2,273	0.42	0.84	1.08	1.3	0.00	11.4	%
Ba capped	2,273	0.42	0.77	0.83	1.1	0.00	3.5	%
High-grade Barium Domain
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	143		2.44			0.43	3.1	m
Ba	143	8.40	9.45	3.90	0.4	2.87	24.8	%
Ba capped	143	8.40	9.03	3.10	0.3	2.87	15.0	%
Outside Barium Domains
	Valid	Median	Mean	Std Dev	CV	Minimum	Maximum	Units
Length	8,805		2.97			0.02	3.1	m
Ba	8,805	0.07	0.16	0.45	2.8	0.00	11.2	%
Ba capped	8,805	0.07	0.15	0.32	2.1	0.00	4.0	%


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Figure 14-19: Pinion Barium Domains and Geology – Section N4479230


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Table 14-35: Pinion Barium Estimation Parameters
(for all rotations/dip/tilt values, see Table 14-22)


Domain	Parameter
Low-grade Barium Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search anisotropies: major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	3
High-grade restrictions (grade in %Ba)	N/A
High-grade Barium Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	4
High-grade restrictions (grade in %Ba)	N/A
Outside Modeled Barium Domains
Samples: minimum/maximum/maximum per hole	2 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 1 / 0.5
Inverse distance power	3
High-grade restrictions (grade in %Ba and distance in m)	0.15 / 9

The average barium grade for the gold mineralization grading at least 0.14 g Au/t in potentially mineable material is ~1.7%. There are substantially fewer barium analyses than gold analyses, so the barium estimate has lower confidence than the gold estimate. If precision of barium grades is critical to the economics of the deposit, then additional samples with barium grades should be obtained.


14.3.6.2	Pinion Cyanide-Soluble Gold Model

A cyanide-soluble gold block model was produced using cyanide-recoverable gold shaker test results and fire assays of sample pulps (Figure 14-20). AuCN/AuFA ratios were calculated from these two types of assays. ID³was used to estimate the AuCN/AuFA ratio grades. Only AuCN/AuFA ratios in which the fire-assay gold grades were greater than or equal to 0.05 g Au/t were used in the estimation. Due to the relatively few cyanide-shaker tests and the fact that no quality control or database auditing was done on these analyses, the AuCN/AuFA ratio block model is lower in confidence than the gold and silver block models. Otherwise, the estimation procedures, block dimensions, and methodology were generally the same as those used for the gold and silver models, with the exception of the items noted below in this section.


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Figure 14-20: Cumulative Probability Plot of Pinion AuCN/AuFA Ratios

The AuCN/AuFA ratio block model augments the barite block model to further define metallurgical domains applicable to estimating gold recovery. Referencing Section 13 again, blocks having estimated AuCN/AuFA ratios of less than 50% are deemed to have gold that is not sufficiently recoverable with cyanide processing; this material is not reported in the estimated mineral resources. Blocks with estimated AuCN/AuFA ratios of between 50 and 70% are categorized as transitional, are considered be to have moderate recovery, and are included in the reported mineral resources. Blocks with estimated AuCN/AuFA ratios of 70% or greater are classified as having relatively good recovery and are categorized as oxidized. These three categories are further subdivided using estimated barium values above or below 4%.


14.3.6.3	Refractory Solids Model

The term “refractory” technically refers to material that contains sulfur and carbon species that render gold extraction difficult with cyanide processing. Refractory solids were therefore modeled in order to provide input into metallurgical characterization and potential acid-generating properties. The refractory solids modeled at Pinion delineate unoxidized, sulfide-bearing material with carbon. Refractory zones within the solids generally correlate with material from which gold recovery is difficult as defined above using estimated barite grades and AuCN/AuFA ratios. Zones outside the solids are generally consistent with the oxide and transitional metallurgical domains defined above.

Gold Standard initially modeled solids using a combination of logged data, which represents the most abundant data set, augmented by AuCN/AuFA ratios. MDA modified these solids to include LECO sulfide-sulfur analyses. The contact of the resulting refractory solids is commonly abrupt and readily defined by the multiple data sets, which are rarely contradictory. Within the refractory solids, logged data indicates material is 30% or more refractory, AuCN/AuFA ratios are generally much lower than 50%, and sulfide-sulfur grades are mostly in the tenths of a percent or higher. By far the largest volume of refractory material is deep, below the multi-lithic breccia and outside the pit defining potentially minable mineral resources. Within the volume of the potential open-pit, refractory material is mostly coincident with the Chainman and to a lesser extent the Tripon Pass Formation, but a small amount is also in the Webb Formation in the southwest part of the pit. All refractory material within a potential pit lies above the multi-lithic breccia that hosts the


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gold mineralization. There is some known refractory material which is not modeled within the solids below mineralization in the Devils Gate Limestone, because of the limited drilling, but that material lies below and is immaterial to the estimated mineral resources.

The refractory solids model and the data on which it is based support the inference that potentially lower-recovery material, or material with the potential for having acid producing qualities, are properly represented.


14.3.7	Pinion Acid-Base Accounting Model and Estimation

An acid-base accounting block model was produced to characterize the acid-generating or neutralizing potential of mined waste material. MDA estimated CINO and SSUL into the ABA block model, and designated model blocks as either PAG or NAG. All ABA calculations and PAG/NAG designation criteria were provided by Stantec.

Gold Standard provided LECO analyses of carbon and sulfur species. The analyses were done on samples that varied at 1 ft (0.3048 m) to 6 ft (1.829 m) between those on original core intervals, and between RC sample composites at 10 ft (3.048 m) to 35 ft (10.668 m).

MDA evaluated the CINO and SSUL statistics by rock unit, barium domain and in/out of the refractory solids (Table 14-36 and Table 14-37). The statistics in the tables are summarized according to categories chosen for estimation. Because relationships between silica and barium contents relative to CINO and SSUL are similar, subsequent discussions regarding statistics and estimates in terms of barium domain also apply to the silica zones.

Table 14-36: Number of Samples and Mean Inorganic Carbon Values for Pinion Estimation Categories
(by rock unit, barium domain, and zones inside [refractory] or outside [oxide and transitional] refractory solids)


Estimation Category	Multi-lithic Breccia
Estimation Category	# of Samples	Mean Value (%)
Low-Grade Barium, and Outside Barium Domains, Oxide and Transitional	914	1.958
High-Grade Barium, Oxide and Transitional	113	0.495

Estimation Category	Sentinel Mountain Dolomite and Devil's Gate
Estimation Category	# of Samples	Limestone Mean Value (%)
Low-Grade Barium, and Outside Barium Domains	762	9.084

Estimation Category	Chainman and Webb Formations
Estimation Category	# of Samples	Mean Value (%)
Low-Grade Barium, and Outside Barium Domains, Oxide and Transitional	1,250	0.553
Low-Grade Barium, and Outside Barium Domains, Refractory	259	2.653

Estimation Category	Tripon Pass Formation
Estimation Category	# of Samples	Mean Value (%)
Low-Grade Barium, and Outside Barium Domains	740	4.299


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Table 14-37: Number of Samples and Mean Sulfide Sulfur Values for Pinion Estimation Categories
(by rock unit, barium domain, and zones inside [refractory] or outside [oxide and transitional] refractory solids)


Estimation Category	Multi-lithic Breccia
Estimation Category	# of Samples	Mean Value (%)
Oxide and Transitional	1,020	0.026

Estimation Category	Sentinel Mountain Dolomite and Devil's Gate Limestone
Estimation Category	# of Samples	Mean Value (%)
All Data	770	0.006

Estimation Category	Webb Formation
Estimation Category	# of Samples	Mean Value (%)
Oxide and Transitional, Outside Barium Domains	313	0.022
Oxide and Transitional, Low- and High-Grade Barium Domains	8	0.000
Refractory, Outside Barium Domains	37	0.101
Refractory, Low- and High-Grade Barium Domains	3	0.457

Estimation Category	Chainman Formation
Estimation Category	# of Samples	Mean Value (%)
Oxide and Transitional, Outside Barium Domains	917	0.069
Oxide and Transitional, Low- and High-Grade Barium Domains	10	0.000
Refractory, Outside Barium Domains	215	0.403
Refractory, Low- and High-Grade Barium Domains	4	1.445

Estimation Category	Tripon Pass Formation
Estimation Category	# of Samples	Mean Value (%)
Oxide and Transitional, Outside Barium Domains	432	0.024
Oxide and Transitional, Low- and High-Grade Barium Domains	122	0.038
Refractory, Outside Barium Domains	160	0.145
Refractory, Low- and High-Grade Barium Domains	25	0.158

Categories represented by only a small number of samples were evaluated on-screen with respect to location and volume of material to be estimated. If the volume in the block model to be estimated could be reasonably estimated without projecting CINO and SSUL grades over extreme distances, they were estimated and are included in Table 14-36 and Table 14-37. However, if unreasonable distances were required to estimate grades into model blocks, then values were assigned to those blocks rather than estimated. The assigned values (Table 14-38) were determined based on relationships between mean CINO and SSUL values for categories that are well represented by data.


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Table 14-38: Assigned Inorganic Carbon and Sulfide Sulfur Values for Pinion Estimation Categories
(by formation, barium domain and zones inside [refractory] or outside [oxide and transitional] refractory solids)


Assigned CINO
Formation	Barium Domain	Refractory Zone	Assigned Value
Multi-lithic Breccia	Low-Grade	Refractory	2.67
Multi-lithic Breccia	High-Grade	Oxide and Transitional	0.49
Multi-lithic Breccia	High-Grade	Refractory	0.74
Multi-lithic Breccia	Outside Domains	Refractory	3.86
Sentinel Mountain Dolomite and Devil's Gate Limestone	High-Grade	All	3.07
Chainman and Webb Formations	High-Grade	All	0.46
Chainman and Webb Formations	Low-Grade and Outside Domains	All	1.28
Tripon Pass Formation	High-Grade	All	1.39

Assigned SSUL
Formation	Barium Domain	Refractory Zone	Assigned Value
Multi-lithic Breccia	Low-Grade	Refractory	0.07
Multi-lithic Breccia	High-Grade and Outside Domains	Refractory	0.16

CINO statistics varied inversely and systematically by rock unit in combination with barium domain and silica zone (in/out of modeled solids) for the all units. The inverse correlation is indicative of increasingly altered and mineralized rocks due to baritization, silicification, and decarbonization. CINO values in the multi-lithic breccia and Webb and Chainman Formations differ in each barium domain. In low-grade barium and outside the barium domains, CINO also varies by refractory zone (in/out of modeled solids). CINO contents in low-grade barium domains and outside modeled barium domains behave similarly compared to high-grade barium domains in the Sentinel Mountain Dolomite, Devil’s Gate Limestone, and Tripon Pass Formation, and there is no distinction by refractory zone. SSUL statistics show strong relationships by refractory zone within the multi-lithic breccia. In the Webb, Chainman, and Tripon Pass Formations, SSUL varies by both refractory zone and barium domain. Statistics for SSUL in low- and high-grade barium domains are similar compared to outside barium domains in these units. No systematic differences were observed in SSUL values for the Sentinel Mountain Dolomite or Devil’s Gate Limestone, so both were estimated together using all respective contained data.

CINO and SSUL contents were estimated independently into the ABA block model, according to the categories described above. CPPs for each species estimated were evaluated by category for potential capping of assays. None was warranted for CINO, but several caps were applied to the SSUL data (Table 14-39). Half the sample composites are ~1 m in length. However, about one-quarter of the lengths are 9.144 m (30 ft). Given the model block dimension of 9 m³, assay sample data were composited to 9.14 m.


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Table 14-39: Sulfide Sulfur Capping Values for Pinion Estimation Categories
(by formation, barium domain, and zones inside [refractory] or outside [oxide and transitional] refractory solids)


Capped SSUL
Formation	Barium Domain	Refractory Zone	Capping Value (%)
Multi-lithic Breccia	All	Oxide and Transitional	0.50
Sentinel Mountain Dolomite and Devil's Gate Limestone	All	All	0.40
Webb Formation	Outside Domains	Oxide and Transitional	0.60
Webb Formation	Outside Domains	Refractory	0.30
Chainman Formation	Outside Domains	Oxide and Transitional	0.70
Tripon Pass Formation	Low- and High-Grade	Oxide and Transitional	0.30
Tripon Pass Formation	Outside Domains	Oxide and Transitional	0.40

All estimates were done using the same search orientations and associated estimation areas as were applied to the gold and silver estimates (Table 14-6). The maximum search distance applied to most estimates for both CINO and SSUL was 300 m. The maximum distance for a few runs were extended to 350 m on a limited basis to fill in a small number of unestimated blocks. Search ellipses were strongly anisotropic, with major, minor, and vertical search distances at 300 m, 300 m, and 75 m, respectively, and ID²methodology was used. Due to the relatively long composite length, the maximum number of composites, and maximum composites per hole allowed to estimate a block were limited to five and two, respectively. No search restrictions were applied to either CINO or SSUL estimates.

Correlograms were generated to evaluate continuities in the data with respect to distance. These demonstrated reasonable continuity for CINO at ranges up to 370 m data in low-grade and outside the barium domains. There was not enough data to build meaningful correlograms in the high-grade barium domain.

Correlograms of SSUL data indicate continuity to a maximum of 100 m, depending on refractory zone. As noted above, the maximum search distance applied to most estimates for CINO and SSUL was 300 m. The maximum distance for estimation applied to SSUL was the same as applied to CINO. The relatively short continuity indicated by correlograms might preclude the application of longer search distances, but PAG/NAG designation is dependent on the estimated grades of both CINO and SSUL, and a significant portion of blocks would not be characterized as PAG or NAG. So, although there is lower confidence in the SSUL estimated values beyond distances of 100 m, most blocks within potential open pits are able to be designated as PAG or NAG. This added risk was recorded as a block attribute.

The LECO data is relatively well-distributed within the deposit in potentially mineable pits at lower gold prices, but there are localized areas that lack data. Also, significant areas of pits at higher gold prices contain no LECO data at all. Estimated grades of CINO and SSUL in these areas can be relatively far from data. To flag model blocks that are at relatively greater distances from data, MDA assigned a confidence code of ‘0’ to all estimated blocks with closest composite more distant than 130 m. This confidence code compensates for the shorter continuities demonstrated in correlograms for SSUL. Because CINO and SSUL were estimated according to different criteria, these codes were assigned separately for each, and a combined code was assigned if either CINO or SSUL confidence codes was ‘0’.

Like Dark Star, model blocks were designated as PAG (code of ‘1’) or NAG (code of ‘2’) according to criteria as defined by Stantec. First, ANP, AGP, and NNP values were calculated from estimated CINO and SSUL values. Next, a PAG/NAG designation was assigned according to criteria for three potential waste characterization scenarios, as shown in Table 14-15 located in Dark Star Section 14.2.6.


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14.3.8	Pinion Density

All densities were measured using the immersion method by an independent laboratory. There are 440 density-sample measurements in the Pinion database within assayed intervals. However, once these samples are parsed out by formation/rock unit and barite domains, the geologic features that most affect density, there are only a small number (in the tens) of samples representing each category. The mean density values, and the values assigned to the units in the model, are summarized in Table 14-40. Application of density values to the Pinion gold block model was dependent on numerous supporting modeled (formation/rock unit and silicification zone in/out of solids) and estimated (barium grade) criteria that have been discussed in various prior sections.

Table 14-40: Density Values Applied to the Pinion Block Models


Formation/Rock Unit	Barite Domain	Silicification Zone	Density applied to model
Multi-lithic breccia	Outside	Outside	2.52
Multi-lithic breccia	Low-barite	Any	2.60
Multi-lithic breccia	Low- and High-barite*	Any	2.83
Multi-lithic breccia	High-barite	Any	3.05
Multi-lithic breccia	Outside	High-silica	2.54
Sentinel Mountain	Any	Any	2.56
Devils Gate	Outside	Outside	2.62
Devils Gate	Low-barite	Any	2.62
Devils Gate	Low- and High-barite*	Any	2.68
Devils Gate	High-barite	Any	2.74
Devils Gate	Outside	High-silica	2.56
Webb	Outside	Outside	2.42
Webb	Low-barite	Any	2.70
Webb	Low- and High-barite*	Any	2.71
Webb	High-barite	Any	2.72
Webb	Outside	High-silica	2.53
Chainman	Outside	Outside	2.45
Chainman	Low-barite	Any	2.40
Chainman	Low- and High-barite*	Any	2.55
Chainman	High-barite	Any	2.70
Chainman	Outside	High-silica	2.46
Tripon	Outside	Outside	2.49
Tripon	Low-barite	Any	2.57
Tripon	Low- and High-barite*	Any	2.64
Tripon	High-barite	Any	2.70
Tripon	Outside	High-silica	2.61
(* in the same block)


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14.3.8	Discussion of Pinion Estimated Mineral Resources and Supporting Models

Pinion has a long history of exploration drilling dating back to 1981 and consequently the project has many drill holes of varying quality and reliability. On the one hand, having twelve different companies drilling with twelve different procedures, documentation, and general quality makes the database difficult to verify and reduces the confidence. On the other hand, results from so many companies drilling over such a long period of time that corroborate each other add significant confidence. Consequently, the estimators spent much time auditing, evaluating QA/QC, sample integrity, and comparing drill campaigns through explicit modeling of domains. Because of these tasks, many of the TCX holes drilled by Amoco in 1981 were eliminated from use in estimation, for example.

None of the historical drilling has supporting QA/QC and not all have supporting assay certificates. This lower-confidence data set was taken into account during mineral resource classification by downgrading some of the blocks that were entirely dependent on the historical data. However, the downgrade was not severe because all historical data was mutually supportive, except for the 1981 Amoco drilling.

There is some contamination noted and those samples deemed likely contaminated by both MDA and Gold Standard were eliminated from estimation. However, below the main mineralized multi-lithic breccia body, there are many samples with low grades of gold. It is presently impossible to determine if these represent contaminated samples, so many of these blocks were classified as Inferred. Little of this is economic under today’s economic conditions.

The AuCN/AuFA ratios are based on relatively fewer cyanide-shaker test assays relative to fire assays that are unaudited and lack QA/QC. The cyanide-soluble gold block model, while likely reasonable, is a global estimate and the ability to produce reliable estimates on a block by block basis is improbable. The AuCN/AuFA ratio block model is lower in confidence than the gold and silver block models.

In addition to the mineral resources reported herein, there is mineralization that continues beyond, and is contiguous with the reported mineral resources. The reported mineral resource estimate is pit-constrained and therefore some of the estimated mineralization (tonnes, grade and ounces) is unreported. That additional mineralization is shown graphically in Figure 14-21.


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Figure 14-21: Pinion Optimized Pit and Additional Mineralization
(gray lines are drill holes; blue solid is the 0.14 g Au/t grade shell; red is the mineral resource pit shell)

Where silver was modeled, the ratio of silver grade to gold grade is around 7:1.

The Pinion deposit has clustered drill data, which if not taken into consideration during estimation, imparts risk to the estimate. Fortunately, the clustering of the data lies within the open-pit limits and where mining will potentially take place. This area is also, for the most part, the highest-grade area. For that reason, any estimation, except for nearest-neighbor type, will have a tendency to “push” those clustered-sample grades out beyond where they should be. Because that effect was noted during checking, the inverse-distance power for the low-grade gold domain estimate was changed to three, and to four for the high-grade gold domain. De-clustered composite data is better reflected in a nearest-neighbor grade than the inverse-distance estimated grade. There is a chance that these outer-area Inferred mineral resource grades may end up being slightly lower than what is presented herein.


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Previously reported mineral resources for Pinion were smaller than what is reported herein. Globally estimated mineralization, but not reported, is similar. However, the constraints placed on the estimated tonnes and grade in order to determine “…reasonable prospects for eventual economic extraction” are substantially more rigorous in this Technical Report than were used previously. It should also be noted that the mineral resource contained within any given optimized pit can increase significantly with small increases in gold price. For example, with a price increase of $25/ounce, the mineral resources increase by 50%, 81%, and 18% in the $1,150, $1,400, and $1,700 optimized pits, respectively. The step-like pit expansions are due primarily to the flat-lying, anticlinal nature of the deposit, and occur in a roughly southward direction along the axis of the anticline.

The block dimensions were changed from 6 m x 6 m x 6 m in the APEX model, to 9 m x 9 m x 9 m in MDA’s preparation of this Technical Report. Additional dilution was incorporated in the current mineral resource estimate as a result of using the larger block sizes. MDA performed a bench-height study on composite data to evaluate the potential changes to the mineral resource attributed to the additional dilution with the changed bench height, and showed that, at the cutoff of 0.2 g Au/t, the gold grade would decrease by about 3% and tonnes would increase by about 2%.

Since Gold Standard became the operator, a QA/QC program was initiated, but there is no evidence of a QA/QC program prior to that. The QA/QC program was minimal in 2014 – 2016, with only CRMs and pulp blanks. In 2017 and 2018, field duplicates and coarse blanks were inserted into the sample stream. In 2018, pulps were sent out to secondary laboratories for check assays. As of May 2018, the percent of drill holes, including historical drilling, with QA/QC was 23% and the percent of total meters drilled in those campaigns with QA/QC was 37%. The lack of QA/QC data were accounted for during classification of the current mineral resources.

Gold grades from four metallurgical holes drilled in 2019 and totaling 552.3 m were compared to estimated gold grades in coincident model blocks from a previous, unpublished block model produced by MDA in 2018. The mean and median gold grades of the 2018 model blocks intersected by the 2019 drill holes was 6% higher and 11% lower, respectively, compared to newly drilled grades. A few blocks with relatively extreme gold grades are skewing the mean higher. For blocks and drill-hole intervals with grades over 0.14 g Au/t, the mean and median gold grades of the drill holes were 7% higher and 10% lower, respectively, than estimated block model grades.

Table 14-41: Comparison of 2019 Drilling Gold Assays to 2018 Coincident Model Block Grades


All Coincident Blocks
	Valid	Median	Mean	Std. Dev.	CV	Min.	Max.	Units
Au	58	0.011	0.187	0.278	1.484	0.004	1.119	g Au/t
Au-ID	58	0.010	0.199	0.334	1.677	0.002	1.655	g Au/t
Diff		-11%	6%	20%	13%	-63%	48%	(Model/Actual)
All Coincident Blocks >=0.14 g Au/t
	Valid	Median	Mean	Std. Dev.	CV	Min.	Max.	Units
Au	21	0.504	0.499	0.245	0.491	0.008	1.119	g Au/t
Au-ID	21	0.454	0.533	0.366	0.686	0.058	1.655	g Au/t
		-10%	7%	49%	40%	586%	48%	(Model/Actual)


14.4	JASPEROID WASH MINERAL RESOURCES

The Jasperoid Wash mineral resource estimate was completed on November 15, 2018, which is the effective date of the estimate. The Jasperoid Wash mineral resource estimate is based on drilling through September 6, 2018; however, a minor number of drill holes were updated with new collar surveys and geology as late as October 6, 2018, which makes it effective date for the Jasperoid Wash database.

A total of 21 additional drill holes (5,210.3 m) were drilled at Jasperoid Wash after the effective date of the mineral resource estimate. Two were core holes for 610.8 m, and the remainder were RC for 4,599.5 m. Data for these holes


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were received on April 24, 2019 and evaluated on April 29, 2019. No auditing or QA/QC evaluations were done on this data set. The impacts of these holes on the reported mineral resource estimate are described in 14.4.3.


14.4.1	Jasperoid Wash Database

Since 1989, three companies have conducted exploration drilling at Jasperoid Wash. Gold Standard began drilling in 2017. In all, 91 RC holes (92% of meterage) and 6 core holes (8% of meterage) totaling 17,407 m have been drilled (see Table 14-42 and Figure 14-22). There are no historical QA/QC data for the historical holes, which currently represent 46% (8,031 m) of the holes in the mineral resource database.

Descriptive statistics of all Jasperoid Wash drill-hole analytical data audited and imported into MineSight by MDA are summarized in Table 14-43. There are no density measurements at Jasperoid Wash. Because there are so few core holes, core recovery and RQD data were not imported.

Table 14-42: Summary of Drilling at Jasperoid Wash


Company	Type	Number	Total meters
Cameco	RC	7	1,230
	Total	7	1,230
Westmont	Core	3	295
	RC	47	6,506
	Total	50	6,801
Gold Standard	Core	3	1,070
	RC	37	8,306
	Total	40	9,376
Total	Core	6	1,365
	RC	91	16,042
Grand Total		97	17,407

Table 14-43: Descriptive Statistics of Sample Assays in Jasperoid Wash Mineral Resource Database


	Valid	Median	Mean	Std Dev	CV	Min	Max	Units
From	10,387					0	589.8	m
To	10,387					1.52	589.8	m
Length	10,387	1.52	1.68			0.01	46.5	m
Type	10,341					1	2.0
Au	10,147	0.043	0.101	0.178	1.77	0.003	2.88	g/t
CN-sol	1,498	0.130	0.194	0.241	1.24	0.015	2.80	g/t
CN:FA Au	1,498	74	63	30	0.50	2	100	%

The Jasperoid Wash database contains 10,147 gold assay records (Table 14-43). No explicit determination of sample reliability was made because the intent here is to only report Inferred mineral resources, and there was nothing


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indicating serious problems with sample reliability. However, three holes have long intercepts of mineralization that are somewhat anomalous to adjacent holes. These were used in the mineral resource estimate, but additional work or drilling should be done to ensure that these results are reliable. If deemed not reliable, the impact on the mineral resource estimate would be small.

Gold Standard’s drill-hole collar locations, downhole survey data, and gold analyses were audited for verification. There are few supporting certificates for any historical drilling. The database also contains logged geologic features, including rock types, formations, faults, vein type, silicification, clay, dolomite, barite, limonite, hematite, carbonate, sulfide percent, and percent reduced, all of which were imported. The logged geology was reviewed and used in modeling but was not audited.


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Figure 14-22: Jasperoid Wash Deposit Drill-hole Map and Mineral Resource Outline
Note: hachured area shows third-party inlier claims not controlled by Gold Standard.


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14.4.2	Jasperoid Wash Geologic Model

Gold Standard provided geologic interpretations as surfaces and solids for faults, formation contacts, alteration and shapes defining areas of high AuCN/AuFA ratios. All geologic surfaces were interpreted on east-west cross-sections by use of surface maps and downhole drill data. MDA reviewed all sections and models provided by Gold Standard, and when problematic areas were encountered, MDA worked with Gold Standard geologists to produce a coherent, agreed upon cross-section or model.

MDA combined appropriate upper and lower geologic rock unit surfaces, fault surfaces, and intrusive cross-sectional interpretations to produce geologic solids for coding the block model. Coded rock units include: the Mississippian Tonka Formation (a conglomerate), the Pennsylvanian-Permian undifferentiated units (from oldest to youngest - lower conglomerate, lower siltstone, middle conglomerate, and upper siltstone), and Tertiary intrusive bodies. The middle conglomerate of the undifferentiated Pennsylvanian-Permian units, which may correlate with units at Dark Star that are possibly Tomera Formation age equivalent rocks, is the primary host for mineralization. The Tertiary conglomerates and Elko Formation is recognized as a secondary host, and the lower siltstone contains some less extensive gold mineralization. Limited mineralization is found in the lower conglomerate, and additional mineralization can be encountered within the intrusive bodies. MDA determined that Quaternary colluvium exists in insufficient quantities to impact mining and mineral resources, so it was not modeled. All geologic interpretations, in combination with assays and logged data, were used to guide metal domain modeling and to define metallurgical domains of cyanide solubility and clay zones.


14.4.3	Jasperoid Wash Gold Modeling and Estimation

14.4.3.1	Gold Domain Model

Gold domains based on sample assay ranges were interpreted on sections spaced 100 m apart, oriented east-west and looking north. Domains were defined based on population breaks on the CPP for all gold data (Figure 14-23). Two domains, even if indistinct on the CPP plots, were needed to control what was clearly a higher-grade portion of the deposit and maintain its likely geometry. The lowest-grade domain limit is at about 0.05 g Au/t, but its definition is unclear because of the high and variable gold-assay detection limits. The higher-grade domain boundary is ~0.15 g Au/t to 0.2 g Au/t, where a very subtle break occurs in the line on the CPP plot in Figure 14-23. While there is a higher-grade domain greater than ~1.5 g Au/t, there are so few samples that there is no evidence of continuity. There are no outlier grades in either domain at Jasperoid Wash. Descriptive statistics of assays by the modeled domains are presented in Table 14-44.


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Figure 14-23: Cumulative Probability Plot of Jasperoid Wash Gold Assays

During a site visit in September 2018, Mr. Lindholm reviewed core from JW17-01 and JW18-01. Gold Standard staff geologists provided guidance and expertise with respect to the geology of the deposit and the nature of gold mineralization. As is common with Carlin-type, sedimentary-rock hosted epithermal gold deposits, the relationships between gold mineralization and rock, alteration and/or mineral assemblages can be subtle and inconsistent. However, the following characteristics were commonly observed with respect to gold mineralization:

In Carlin-type systems, higher porosity can be attributed to decalcification of calcareous sedimentary rocks and coarser-grained sedimentary units. At Jasperoid Wash, the middle conglomerate is typically more decalcified than other units above and below;
Gold mineralization is commonly confined between less permeable units, such as argillized fault gouges or stratigraphic horizons;
Argillized areas often occur adjacent to felsic intrusive bodies and related zones of structural movement or weakness. No visible sulfides were observed in argillized areas; however, a distinct and strong sulfur smell was noted in argillized zones of JW17-01; and
Mineralized areas outside of argillic zones are dominated by limonite in fractures and moderate hematization of host rock.


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Table 14-44: Jasperoid Wash Descriptive Statistics by Gold Domain


Low-Grade Gold Domain
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
From	2,062						313.64	m
To	2,062					1.52	315.16	m
Length	2,062	1.52	1.54			0.01	22.65	m
Type	2,050					1	2
Au	2,037	0.101	0.111	0.070	0.63	0.003	0.720	g/t
Au Capped	2,037	0.101	0.111	0.070	0.63	0.003	0.720	g/t
Au_CN	633	0.100	0.101	0.064	0.63	0.015	0.630	g/t
Au_CN/Au_FA ratio	633	71	64	28	0.40	6	100	%
High-Grade Gold Domain
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
From	1,345						242.32	m
To	1,345					1.52	243.84	m
Length	1,345	1.52	1.50			0.02	2.59	m
Type	1,343					1	2	0
Au	1,338	0.290	0.402	0.327	0.81	0.010	2.884	g/t
Au Capped	1,338	0.290	0.402	0.327	0.81	0.010	2.884	g/t
Au_CN	774	0.200	0.282	0.302	1.07	0.015	2.800	g/t
Au_CN/Au_FA ratio	774	78	65	32	0.50	2	100	%
Outside Gold Domains
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
From	6,980					0	589.79	m
To	6,980					1.52	589.8	m
Length	6,980	1.52	1.75			0.01	46.46	m
Type	6,948					1	2	0
Au	6,772	0.026	0.038	0.057	1.49	0.003	1.543	g/t
Au Capped	6,772	0.026	0.038	0.048	1.27	0.003	0.600	g/t
Au_CN	91	0.070	0.090	0.122	1.35	0.015	1.090	g/t
Au_CN/Au_FA ratio	91	45	45	26	0.60	2	100	%

Geologic interpretations provided guidance into the definition of the domains. The mineralization is very much stratiform on the east, whereas it dips gently to the west. The mineralization becomes more steeply dipping to the west where faults cause the stratigraphy to drop down. This steep-to-the-west-dipping structural corridor, defined by surface mapping, seems to have controlled the Tertiary intrusions. Often, mineralization is found within the intrusive bodies and along their contacts. Gold mineral domains were generally drawn parallel to stratigraphic contacts in the east and parallel to the intrusions to the west. Silver was not modeled.

The MT thrust fault is located west of the mineralization and is considered a hard boundary, with no mineralization extending beyond it to the west. The MT thrust fault is interpreted to dip about 60^°to the west, parallel to the interpreted dips of the intrusive rocks and other faults.


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After sectional interpretations were completed, gold domains were snapped to drill holes in three dimensions and sliced to mid-bench level plans for modeling. The modeled level plans are spaced at 6 m and are located the middle of each bench. A cross section showing the interpreted gold domains is given in Figure 14-24.


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Figure 14-24: Jasperoid Wash Zone Gold Domains and Geology – Section N4473200


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14.4.3.2	Gold Composites Statistics and Capping

Jasperoid Wash gold domains were defined and modeled on 100 m spaced cross sections and each domain was used to code the drill-hole samples. Cumulative probability plots were made of the coded assays, which were reviewed to determine appropriate capping limits. Capping for each domain was determined by first assessing the grade above which outliers occur. Capping values were determined for each of the gold domains separately. Assays in the Jasperoid Wash gold domains required no capping but samples outside of the domains were capped to 0.6 g Au/t.

Once capping was completed, drill-hole samples were down-hole composited to 3.05 m to respect the original 5 ft drilled intervals, which honors domain boundaries. The composite length was chosen because the majority of samples are 1.52 m in length. Descriptive statistics were generated for all composites and were considered with respect to capping levels (Table 14-45).

Table 14-45: Descriptive Composite Statistics by Domain for Jasperoid Wash


Low-Grade Domain
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
To	1,091						315.16	m
Length	1,091					0.01	3.05	m
Au	1,091	0.103	0.111	0.058	0.52	0.0025	0.583	g/t
Au Capped	1,091	0.103	0.111	0.058	0.52	0.0025	0.583	g/t
Au_CN/Au_FAratio	481	71	64	28	0.4	7	100	%
High-Grade Domain
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
To	702						243.84	m
Length	702					0.57	3.05	m
Au	702	0.291	0.396	0.296	0.75	0.02	2.51	g/t
Au Capped	702	0.291	0.396	0.296	0.75	0.02	2.51	g/t
Au_CN/Au_FA ratio	416	78	65	31	0.5	2	100	%
Outside Gold Domains
	Valid	Median	Mean	Std. Dev.	CV	Min	Max	Units
To	3,398					1.52	589.8	m
Length	3,398					0.01	3.05	m
Au	3,398	0.028	0.038	0.050	1.30	0.00	1.14	g/t
Au Capped	3,398	0.028	0.038	0.043	1.14	0.00	0.60	g/t
Au_CN/Au_FA ratio	87	44	43	26	0.6	2	100	%

Correlograms were built from the composited gold grades in order to evaluate grade continuity. Correlogram parameters were used in the kriged estimate, which was used as a check on the reported inverse distance estimate, and also to give guidance to the classification of mineral resources. The correlograms for the mineralized domains have a nugget at 30% of the total sill. The first sill is 30% of the total sill with a range of 25 m to 30 m depending directions. The second sill is 40% of the total sill with a range of 35 m to 55 m depending directions.


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14.4.3.3	Gold Estimation

The block model is not rotated, and the blocks are 6 m north-south by 6 m vertical by 6 m east-west. The block dimensions are smaller than those for Pinion and Dark Star because the deposit is both smaller and more lenticular. Only gold was estimated and is being reported.

Multiple iterations of four types of estimates were completed: polygonal, nearest neighbor, inverse distance, and kriged with the inverse-distance estimate being reported. The nearest neighbor, inverse distance and kriged estimates were run several times in order to determine the optimum estimation parameters. ID³was used for the outside and low-grade domains. ID²was used for high-grade domains.

The model was divided into three estimation areas (Figure 14-25) to control search anisotropy, orientation, and distances according to the differing geometries of mineralization in each area. Table 14-46 lists these areas along with the search orientations and the maximum search per area by low-grade and high-grade gold domains. Figure 14-22 presents the spatial relationship of those estimation areas to the drilling and the gold domains.

Figure 14-25: Jasperoid Wash Estimation Areas and Gold Domains in Cross Section
(looking north)

Table 14-46: Jasperoid Wash Search Ellipse Orientations and Maximum Search Distances by Estimation Area


Estimation Area	Search Ellipse Orientation			Maximum Search Distance
Estimation Area	Azimuth	Dip	Rotation	Low- Grade	Mid- Grade	Outside Domains
1	90^o	30^o	0^o	300	250	50
2	90^o	75^o	0^o	300	250	50
3	90^o	15^o	0^o	300	250	50
Note: Semi-major search distance = major search distance; vertical (or minor) search distance = major search distance ÷ 2 (area 1) and ÷ 4 (areas 2 and 3)


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One estimation pass of up to 300 m was run for each domain. All estimation runs weighted the samples by the sample lengths. Estimation parameters are given in Table 14-47.

Table 14-47: Jasperoid Wash Estimation Parameters
(for search orientations and maximum distances, see Table 14-6)


Description	Parameter
Low-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 /2
Search anisotropies: major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g/t, distance in m)	NA
High-grade Gold Domain
Samples: minimum/maximum/maximum per hole	1 / 12 / 2
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / varies 0.5 to 0.25
Inverse distance power	2
High-grade restrictions (grade in g/t, distance in m)	NA
Outside Modeled Gold Domains
Samples: minimum/maximum/maximum per hole	1 / 12 / 3
Search (m): major/semimajor/minor (vertical)	1 / 0.5 / 0.25
Inverse distance power	3
High-grade restrictions (grade in g/t, distance in m)	0.1 / 6


14.4.4	Jasperoid Wash Gold Mineral Resources

Mr. Ristorcelli reports mineral resources at cutoffs that are reasonable for deposits of this nature, given anticipated mining and processing methods and approximate though current operating costs, while also considering economic conditions, because of the regulatory requirements that a mineral resource exists “in such form and quantity and of such a grade or quality that it has reasonable prospects for eventual economic extraction.” Although the author of this section is not an expert with respect to environmental, permitting, legal, title, taxation, socio-economic, marketing, or political matters, the author is not aware of any unusual factors relating to these matters that may materially affect the Jasperoid Wash mineral mineral resources as of the date of this Technical Report.

Mr. Ristorcelli classified the Jasperoid Wash mineral resources giving consideration to the confidence in the underlying database, sample integrity, analytical precision/reliability, QA/QC results, and confidence in geologic interpretations. The classification parameters are given in Table 14-8. Since there is a large amount of historical data, and because the geologic model is still evolving, all mineral resources at Jasperoid Wash are classified as Inferred.

For reporting, technical, and economic factors likely to influence the “reasonable prospects for eventual economic extraction” were evaluated using the best judgement of the author responsible for this section of the report. For evaluating the open-pit potential, MDA modeled a series of optimized pits using variable gold prices, mining costs, processing costs, and anticipated metallurgical recoveries. MDA used costs appropriate for open-pit mining in Nevada, estimated processing costs and metallurgical recoveries related to heap leaching, and G&A costs. The cutoff grades are based on $1,500/oz Au.


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The Jasperoid Wash reported mineral resource estimate is the block diluted ID estimate, comprised of ID³estimates for outside and low-grade domains, and by ID²for high-grade domains. The mineral resources are reported at a cutoff of 0.14 g Au/t for open-pit mining. Table 14-48 presents the estimate of the Inferred gold mineral resources at Jasperoid Wash. A representative cross section of the gold block model is shown in Figure 14-26.

Table 14-48: Jasperoid Wash Inferred Gold Mineral Resources


Cutoff g Au/t	Tonnes	g Au/t	oz Au
0.100	12,955,000	0.29	120,000
0.120	11,637,000	0.31	115,000
0.140	10,569,000	0.33	111,000
0.160	9,681,000	0.34	107,000
0.180	9,237,000	0.35	105,000
0.200	8,825,000	0.36	102,000
0.220	8,139,000	0.37sf	97,000
0.240	6,627,000	0.40	86,000
0.260	5,330,000	0.44	75,000
0.280	4,580,000	0.47	69,000
0.300	4,024,000	0.49	64,000
0.400	2,219,000	0.63	45,000
0.500	1,529,000	0.69	34,000
0.600	1,010,000	0.77	25,000
0.800	322,000	0.97	10,000
1.000	80,000	1.17	3,000


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Figure 14-26 Jasperoid Wash Gold Domains and Block Model – Section N4473200


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14.4.5	Jasperoid Wash Geo-Metallurgical Model

A cyanide-soluble gold block model was estimated using AuCN/AuFA ratios calculated from AuCN shaker test and total fire-assay gold assays. The AuCN/AuFA ratios are plotted in the CCP shown in Figure 14-27. First, a cyanide-soluble gold domain was interpreted on east-west cross sections spaced at 100 m intervals. The logged percent reduced (not oxidized) attribute was used when no AuCN values were available. A cross section showing the cyanide-soluble gold domains is given in Figure 14-28.

Figure 14-27: Cumulative Probability Plot of Jasperoid Wash AuCN/AuFA Ratios

Only about 15% of all fire-assay gold values in the database have corresponding AuCN analyses. Of the samples with AuFA assays inside modeled AuCN/AuFA ratio domains, approximately 23% have AuCN analyses. Within the high-grade gold domain, which is a proxy for economic mineralization, 58% of the gold assays have corresponding AuCN analyses.

ID³was used to estimate the ratios. Only AuCN/AuFA ratios were used in the estimate for samples whose fire-assay gold grades were >=0.05 g Au/t. Due to the relatively few cyanide-shaker assays, and the fact that no quality control or database auditing was done on these analyses, the AuCN/AuFA ratio block model is lower in confidence than the gold and silver block models.


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Figure 14-28: Jasperoid Wash Deposit Rock Type and Metallurgical Models
(Note: clay zones tend to follow faults and intrusives)


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As per metallurgical guidance given in Section 13, blocks:

lying outside the cyanide-soluble solid were coded refractory with no gold recovery;
with estimated Au_CN/Au_FA ratios of less than 50%, were also coded to no gold recovery and assigned to “sulfide”;
blocks with estimated Au_CN/Au_FA ratios of >=50% and <70%, were coded moderate recovery, and assigned to “transitional”; and
blocks with estimated Au_CN/Au_FA ratios of >=70%, were coded good recovery, and assigned to “oxide”.

Search ellipses, orientations, and distances similar to those used for the Jasperoid Wash gold block model were used to estimate the cyanide-soluble ratios. The geo-metallurgical model can only be considered preliminary and is not sufficiently reliable to be used for mineral reserves.


14.4.6	Jasperoid Wash Clay Model

Clay contents were logged in intensities of 1 through 3, with 3 being the highest. Metallurgical work indicates that logged clay of 2 and 3 may require agglomeration. Consequently, Gold Standard constructed a 3D solids model that delineates the majority of samples with logged clay intensities of 2 and 3. These solids project the high-clay zones in a manner consistent with the geology at Jasperoid Wash. Typically, these clay zones parallel the steeply dipping modeled dikes and faults. Overall, this model is considered adequate for a geologically Inferred mineral resource, but confirmation of the clay geometries is needed.


14.4.7	Jasperoid Wash Density

There were no density measurements at Jasperoid Wash as of receipt of the database. Consequently, MDA assigned density values to the gold block model based on similar rock units with measurements at Dark Star. The values assigned to the units in the block model are presented in Table 14-49.

Table 14-49: Density Values Applied to the Jasperoid Wash Block Model


Formation	Au_CN/Au_FA Domain	Density g/cm³
Tomera Fm eq. - Siltstone	In	2.45
Tomera Fm eq. - Siltstone	Out	2.55
Tomera Fm eq. – Conglomerate	In	2.5
Tomera Fm eq. – Conglomerate	Out	2.55
Intrusive Rocks	In	2.4
Intrusive Rocks	Out	2.5
Tonka Fm – Conglomerate	Out	2.5

Following the application of the density values given in Table 14-49, MDA reduced those values to account for the clay alteration, which in some cases is both strong and pervasive. The density value of those blocks lying 50% within the clay solid (Section 14.4.4) were assumed to have 50% clay at a density of 2.2 g/cm³.


14.4.8	Discussion of Jasperoid Wash Estimated Mineral Resources

The Inferred mineral resource classification reflects the current level of geologic understanding and support for Jasperoid Wash. It is likely, however, that the estimated mineral resources are fairly estimated in the area of drilling. However, the mineral resources are likely understated because the deposit is open to the south, north and east. All


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checks – volumes, cumulative probability plots, quantile plots, polygonal check, kriged, and nearest neighbor checks – indicated that the mineral resource is reliable. The biggest risk in this estimate is the geologic interpretation and in particular the orientations and continuity of the dikes. While the total metal may not change with further updates, the location due to geologic interpretations may change. The same risk exists with the clay zones. There are no density measurements and relatively few cyanide-soluble gold assays.

Optimized pits increase in size incrementally with gold price, generally 1% to 7% for each $25 increase in price per ounce. A significant increase in contained ounces of gold occurs in the pit using a $1,725/oz Au price.

Figure 14-29 shows the pit surface above which mineral resources are reported and the grade shell of the reported Inferred mineral resources. The figure also shows the extent of mineralization that exists below the reporting surface.

Figure 14-29: Jasperoid Wash Optimized Pits and Additional Mineralization
(blue lines are drill holes; blue solid is the 0.14 g Au/t grade shell; orange is the mineral resource pit shell; green is the optimized pit shell)

After the 2018 estimate was completed, 21 additional drill holes (5,210.3 m) were completed at Jasperoid Wash; two were core holes (610.8 m) and the remainder were RC (4,599.5 m). These holes were received on April 24, 2019 and evaluated on April 29, 2019. MDA did not audit or review the QA/QC evaluations done on these data because these data were not used to update the mineral resource estimate.

Those post-2018 model drill holes located north and south of the current mineral resource extended mineralization by 100 m to the south, 200 m southeast and 250 m north. New drilling internal to the current block model substantially confirmed the estimation, although with some local changes in location. It is interesting to note that two new core holes essentially twinning two 2018-model RC holes did return higher grades than their corresponding RC holes. The defined area of mineralization would likely be slightly larger if the new drilling was incorporated, but not materially different within the mineral resource pit from the current estimate in this Technical Report.


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14.5	NORTH BULLION DEPOSITS MINERAL RESOURCES

This sub-section is modified from Dufresne and Nicholls (2017b), where additional details were reported. The North Bullion area mineral resource estimate was completed under the direct supervision of Mr. Michael B. Dufresne, M.Sc., P. Geol., with APEX. The North Bullion mineral resource estimate was completed on September 15, 2017, which is the effective date of the estimate, and is based on drill data as of August 18, 2017.

A parent block size of 10 m (X) by 10 m (Y) by 3 m (Z), with sub-blocking down to 5 m (X) by 5 m (Y) by 1.5 m (Z), was applied in order to best honor interpreted wireframe solids that were constructed for the gold mineralization. Only gold was estimated and is being reported.


14.5.1	North Bullion Deposits Data

The North Bullion area database consists of 503 drill holes, many of which were used to guide the geological and mineralization interpretation, of which, 232 drill holes completed from 1980 to 2017 were used in the mineral resource estimation. The extensive historical data compilation and data validation process described in Section 12.3 resulted in a compiled drill database for the North Bullion area that is considered by Mr. Dufresne to be sufficiently reliable for use in the mineral resource estimation described below. The database incorporates all available RC drilling and core drilling completed. Those drill data were imported into Micromine where modeling was initially conducted in cross-sectional view (Figure 14-30).

Figure 14-30: POD Deposit Geologic Cross Sectio
(looking northwest, from Dufresne and Nicholls, 2018)


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A total of 503 drill holes guided the mineralization interpretation and estimation of the North Bullion area estimate. This total comprises 95 core holes completed by Gold Standard from 2010 to 2017, 40 RC drill holes also completed by Gold Standard from 2010 to 2014, along with 40 historical core holes and 328 historical RC holes that were completed from 1969 to 1999. Spacing between drill holes over the main mineral resource area varies from 2 m to 120 m. Drilling has been completed on roughly east-west sections that range in spacing from 25 to 120 m, with an average spacing of around 25 m through Sweet Hollow. The exception to this is the POD zone of mineralization, for which drilling was completed on 040° oriented drill sections that were roughly spaced 25 m apart. All drill holes were used to guide the mineralization model that was ultimately used in the mineral resource estimation.

The North Bullion area assay file comprises 75,890 analyses of variable lengths, of which 75,724 samples have been assayed for gold. Of the 75,890 samples in the database, roughly nine percent (6,941 assays) are in the gold mineralized zones. Within the original assay database there were assay intervals with codes describing the interval as a “missing sample” (e.g., “NS”, “MS” or “#value!”). There were eight NS samples situated with mineralized wireframes and were assigned a nominal gold value of 0.0025 g Au/t, which corresponded to half the lower detection limit.

The drill hole database was validated using the validation functions within the Micromine modeling software. No significant errors or issues were noted.


14.5.2	North Bullion Deposits Lithological Setting of Mineralized Zones

14.5.2.1	North Bullion Zone

Carlin-style disseminated gold mineralization in the North Bullion zone is not exposed at surface. The bulk of the geological understanding and interpretation of the North Bullion zone has come from core drilling of gravity and CSAMT features. Gold mineralization is focused in the footwall of the NBFZ, a north-south-striking zone of normal faults with an overall, down-to-the-east sense of motion. The footwall is a horst of Paleozoic siliciclastic and carbonate rocks, whereas the hanging wall is a deep graben filled with Tertiary volcanic rocks. In the footwall north-south-, northwest-, west-northwest-, and northeast-striking faults appear to be important controls on mineralization.

The North Bullion mineralization occurs in a triangular shaped horst in the footwall of the major north-south-striking, steeply east-dipping, NBFZ. The western edge of the horst is bounded by a northeast-striking, northwest-dipping fault. The deposit is capped by gently east-dipping dacite sills. In general, gold is hosted in two principal subzones, a gently-to moderately-dipping upper subzone of strongly sheared siliciclastic and carbonate rocks (a mixed composite of Mississippian Webb and Tripon Pass formations), and a flat-lying, lower subzone of dissolution-collapse breccia developed above and within silty micrite of the Mississippian Tripon Pass Formation and calcarenite of the Devonian Devils Gate Limestone (Jackson and Koehler, 2014; Jackson et al., 2015). Another, smaller volume of mineralization along the Massif fault is known as the Massif subzone. Between strands of the NBFZ, breccia with both collapse and tectonic features propagated upwards through the Mississippian section incorporating Webb Formation silty mudstone, Tripon Pass Formation silty micrite and Chainman Formation sandstone.

Gold mineralization ranges from 105 to 400 m in depth and steepens from nearly flat (10°) to moderate dip of 45° to the east as the subzones approach the eastern strand of the NBFZ. Gold is associated with sulfide minerals with a sooty appearance, silica, carbon, clay, barite, realgar and orpiment, in addition to elevated arsenic, mercury, antimony, and thallium. High-grade gold (> 6 g Au/t) has been intercepted in both the upper and lower subzones.


14.5.2.2	POD Zone

Gold mineralization at the POD zone is restricted to a steeply-dipping shear zone which trends west-northwest and is situated in rocks stratigraphically higher than the lower zone at North Bullion. The POD gold zone is hosted by silty mudstone of the Mississippian Webb Formation (Masters, 2003; Hunsaker, 2012b). Additionally, gold mineralization at


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POD is associated with silicification including jasperoid, argillic alteration, pyrite, barite, and some minor dolomitic alteration (Hunsaker, 2012b).


14.5.2.3	Sweet Hollow Zone

The Sweet Hollow gold zone is the up-plunge, shallow portion of the upper gold zone of the North Bullion deposit and is located in the footwall of the NBFZ. Sweet Hollow mineralization is hosted in strongly sheared siliciclastic and carbonate rocks of the Mississippian Webb and Tripon Pass formations. Alteration consists of decalcification, silicification, and minor argillization. Dissolution-collapse breccia is developed in carbonate rocks.


14.5.3	North Bullion Deposits Geological Models

Using Micromine 3-D software, a sectional approach was utilized for the initial examination of the drill hole database. A series of east-west and transform geology sections looking north and northwest were compiled by Gold Standard at 20 to 60 m intervals through the POD zone, at 100 m intervals through the Sweet Hollow zone and at 20 to 90 m intervals through the North Bullion zone. These were registered into Micromine to guide the interpretation of geology and mineralization. In addition, a series of surface and subsurface string files of the interpreted faults and top of formation boundaries were also compiled by Gold Standard personnel. From these sections and strings a geological model was built for the North Bullion, Sweet Hollow, and POD zones of mineralization.

In order to construct mineralized envelopes for mineral resource estimation, the gold assay data were examined relative to the geological model. A sectional approach was utilized for the initial examination of the data with a 30 m transform spacing, looking northwest, through the more densely drilled POD zone, a 20 m to 60 m spacing for the Sweet Hollow zone, and 40 to 60 m spacings at the northern end of the deposit (North Bullion zone). These sectional spacings correlate roughly to the 30 m to 60 m drill spacings. The interpretation of the mineralized horizons was guided by the revised geological model and restricted, where necessary, by late high-angle, cross-cutting faults as illustrated in Figure 14-31.


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Figure 14-31: Plan View of North Bullion Deposits Mineralization and Structure Model
(from Dufresne and Nicholls, 2018)

Note: red shows X.Xg Au/t grade shell

Upon closer examination of the sample population and anticipated economic parameters, it was decided to model the mineralization into two domains. These include a greater than 0.5 g Au/t higher-grade envelope and a lower-grade 0.1 to 0.5 g Au/t envelope. The higher-grade of 0.5 g Au/t is thought to better reflect the anticipated mining constraints for the sulfide-rich mineralization. This was completed by interpreting the entire deposit at a 0.1 g Au/t lower cutoff grade. Following this, a higher-grade >0.5 g Au/t envelope was created within this >0.1 g Au/t interpretation. Wireframe solids of both the >0.1 g Au/t low-grade and >0.5 g Au/t interpretations were then created. This produced a set of >0.5 g Au/t wireframes and an outer skin of 0.1 to 0.5 g Au/t wireframes (Figure 14-32, Figure 14-33, and Figure 14-34). For the North Bullion mineralized zone, which only contains sulfide material, a 0.3 to 0.5 g/t Au lower grade envelope was constructed around the >0.5 g/t Au mineralization.

The lower- and higher-grade interpreted wireframes were used to constrain the mineralization for the open-pit-constrained and underground-constrained potential mineral resources. The wireframes were interpreted using a minimum down-hole width of 3 m. Due to the grade, depth and the sulfidic nature of the North Bullion deposits mineralization, it was decided to create another set of wire frames at a 1 g Au/t lower cutoff that could be used for the underground-constrained potential mineral resource.


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Figure 14-32: POD Zone Cross Section of Wire-Frame Interpretations
(looking northwest, showing the interpreted 0.1 to 0.5 g Au/t and >0.5 g Au/t Au shells, from Dufresne and Nicholls, 2018)

Figure 14-33: Sweet Hollow Zone Cross Section of Wire-Frame Interpretations
(looking north, cross 4488015N showing the interpreted 0.1 to 0.5 g Au/t and >0.5 g Au/t gold shells, from Dufresne and Nicholls, 2018)


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Figure 14-34: North Bullion Zone Cross Section of Wire-Frame Interpretations

(looking north, cross section 4488700N showing the interpreted 0.3 to 0.5 g Au/t and >0.5 g Au/t gold shells, from Dufresne and Nicholls, 2018)

14.5.4 North Bullion Cyanide-Soluble Model

Little to no cyanide-soluble gold assay data exists in the North Bullion area database. Re-logging and re-interpretation of the historical and more recent Gold Standard drill-hole logs was completed with a focus on the oxidation state of the material logged. The degree of oxidation within mineralized zones was a key aspect of the re-logging and reinterpretation. Due to the block size of 10 m x 10 m x 3 m, a generalized approach of mapping zones of oxide, mixed or sulfide material was completed with minimum thicknesses of at least 3 m. Where there was no cyanide-soluble gold-recovery information available, the re-logged sulfide mineralization in conjunction with the intensity of hematization, and/or iron limonite alteration were utilized to determine the oxidation state of the material.

The entire North Bullion zone is classified as sulfide (reduced). Drill intersections at the Sweet Hollow and POD deposits comprise both oxide, mixed, and sulfide (reduced) mineralization. For Sweet Hollow and POD, historical logs along with RC chips and core, where available, were reviewed and re-interpreted to determine the percent down-hole sulfide, and intensity of hematite and iron limonite being the primary focus. Based on the current information, the mineralized wire-frame volumes at Sweet Hollow and POD comprise 64% oxide, 1% mixed, and 35% sulfide mineralization.

The areas of sulfide mineralization at Sweet Hollow and POD were modeled with the aim of identifying the proportion that is sulfide and would likely require higher-cost processing in the future. These areas of sulfide mineralization are mainly situated in the deeper portion of the POD zone, and in isolated areas of the Sweet Hollow zone.


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14.5.5 North Bullion Deposits Sample and Composite Statistics and Grade Capping

A probability plot and summary statistics for the North Bullion area un-composited samples lying within the interpreted mineralized zone are presented in Figure 14-35 and Table 14-50. The gold assays exhibit a single population of data. Due to the single population present, linear estimation techniques are deemed suitable for statistical estimation of the North Bullion, Sweet Hollow, and POD gold zones.

Figure 14-35: Probability Plot of Sample Gold Grades in the Mineralized Zones

(from Dufresne and Nicholls, 2018)

Table 14-50: Summary Statistics Sample Gold Grades within the Mineralized Zones

(from Dufresne and Nicholls, 2018)


Statistic	Global	Units
Mean	0.945	g Au/t
Median	0.377	g Au/t
Std Dev	1.900	g Au/t
Variance	3.609
Std Error	0.023	g Au/t
Coeff Var	2.01
Minimum	0.00	g Au/t
Maximum	30.199	g Au/t
Total data	6,941


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Drill hole samples situated within the mineralized wireframe solids were selected and coded with the wireframe name. The coded samples were checked visually next to the drill hole to check that the automatic flagging process worked correctly. All samples were correctly flagged and there was no need to manually flag or remove any samples.

A review of the sample lengths was conducted on the samples that were situated within the mineralized wireframes. The drill-hole sample-width analysis results showed a variable sample length from 0.15 m to 27.43 m. Looking at all sample widths, there is one dominant sample length population at 1.52 m. Of the 6,941 un-composited samples within the mineralization wireframes, 99.84 % of the samples were less than 3.05 m in length. In order to honor the wire-frame zones of mineralization, with some thinning to 3 m, composting was done to 3.05 m lengths. Length-weighted composites were calculated for all of the North Bullion area assay samples. The compositing process started from the first point of intersection between the drill hole and the mineralized wire frame and was stopped at the end of the mineralized wireframe.

The “orphan” composites greater than 1.5 m in length were retained and used in the estimate. This provided the best reflection of grade from the un-composited assays to the composite file. The composited samples were used for all sample statistics, capping, estimation input file, and validation comparisons.

Outlier grades in the gold composite file were used for the capping analysis. The composited gold grades were displayed using a log probability plot (very similar to the probability plot shown in Figure 14-35) and a log histogram plot. Both graphs show gold values belong to one single population. This is evident from the straight nature of the line in the log probability plot below. There is a suggestion of a smaller higher-grade population but due to limited data the composite dataset has been treated as one population. Inflection points along the log probability plot line are normally used to govern an appropriate capping level to apply. There is an inflection point located at the 99th percentile which is 9.0 g/t Au. After analysis of the composite file distribution, examination of the log histogram, log probability plot and the coefficients of variations of the composite statistics it was decided not to cap the composite file.

14.5.6 North Bullion Deposits Grade Continuity

Variography to examine grade continuity was conducted on the North Bullion area composite assays located within the mineralized wireframes and log-spherical semi-variogram’s were produced. The variography of the gold composites in the POD zone suggest a maximum continuity of grade along a 296° strike orientation, with a 0° plunge and a 70° tilt to the east. This is in line with the geological model. The variography of the gold composites in the Sweet Hollow zone suggest a maximum continuity of grade along a 002° strike orientation, with a 0° plunge and a 10° tilt to the west. This is in line with the geological model. The variography of the gold composites in the North Bullion zone suggest a maximum continuity of grade along a 006° strike orientation, with a 15° plunge to the north and a 11° tilt to the east. This is in line with the geological model. The variography of the gold composites in the Massif subzone suggest a maximum continuity of grade along a 172° strike orientation, with a 31° plunge to the north and a 50° tilt to the east. This is in line with the geological model.

A variable search orientation approach was adopted for the orientation of the search ellipsoids during grade estimation. This variable search was adopted to account for varying dips and strikes of the mineralization. This was accomplished by creating a three-dimensional surface of the general orientation of the folded/linear mineralization wire frames across the entire deposit area. This surface showing the trend of the mineralization was then used to orient the search ellipsoid for the nearest block. This was done by obtaining the dip and dip direction of the individual triangles for this trend surface, which were then used to code the blank block model using inverse distance to the power of one, which essentially coded the model with the dip and dip direction information from the trend surface. This dip and dip direction information in the block model was then used to orient the search ellipsoid based upon the orientation of the mineralized zone. That is, if a mineralized volume was folding over, then the search ellipsoid would cater to this change for each block in the block model. The ranges of the search ellipsoid were sourced from the suggested ranges from the variographic analysis and are summarized in Table 14-51.


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Table 14-51: North Bullion Area Semi-Variogram Parameters of Composited Gold Grades

(from Dufresne and Nicholls, 2018)


Zone	ID #	C0	Azimuth	Dip	Tilt	Structure 1					Structure 2
						Type	C1	Direction Ranges			Type	C2	Direction Ranges
						Type	C1	1	2	3	Type	C2	1	2	3
Pod	24	0	296	0	70	sph	0.43	22.5	10	25	sph	2.183	150	85	80
Sweet Hollow	25 & 26	0	2	0	10	sph	0.792	70	50	22
North Bullion	18, 19 & 21	0	6	15	11	sph	0.945	60	70	23
Massif	17	0	172	31	50	sph	0.539	69	69	33	sph	0.611	160	160	160

sph - spherical; C0 - nugget effect; C1 - covariance contribution of Structure 1; and C2 - covariance contribution of Structure 2.

14.5.7 North Bullion Deposits Density

A total of 988 density samples were collected from a variety of sources including bulk sampling, surface rock-chip samples, and drill core. A total of 272 density samples were situated within the mineralized wireframes. Those within the mineralized zones were examined on a zone by zone basis. The average density determined from the analytical work for each zone was assigned to all the blocks within that zone. The average density ranges from 2.33 g/cm³to 2.71 g/cm³with an average density of 2.68 g/cm³for the North Bullion, Sweet Hollow, and POD zones. The density values were determined using the Archimedes (weight in air/weight in water) method. Summary statistics of this analysis is shown in Table 14-52. Most density samples were collected from the North Bullion zone. Further measurements are recommended for collection in the other zones of mineralization. Assigned density values of the individual zones are presented in Table 14-52.

Table 14-52: North Bullion Area Density Measurements by Zone

(from Dufresne and Nicholls, 2018)


Mineralized Zone	NB Massif	NB Lower	NB Mid	NB North	NB Top	NB Upper	POD	Sweet Hollow
Lode Code	1, 17, & 42	2, 18 & 43	3, 19, 44 & 45	4, 20 & 47	5 & 22	6, 21 & 46	23, 24 & 41	25 to 30, 32, 33,35 to 38, 39 & 40
Normal Statistics
Mean	2.676	2.707	2.702	2.573	2.622	2.682	2.326	2.411
Median	2.665	2.695	2.67	1.27	2.58	2.65	2.33	2.42
Std Dev	0.14	0.188	0.156	0.095	0.035	0.152	0.064	0.144
Variance	0.02	0.035	0.024	0.009	0.001	0.023	0.004	0.021
Std Error	0.019	0.033	0.015	0.055	0.014	0.024	0.02	0.046
Coeff Var	0.052	0.069	0.058	0.037	0.014	0.057	0.028	0.06
Minimum	2.42	2.27	2.53	2.5	2.58	2.21	2.21	2.16
Maximum	3.15	3.31	3.79	2.68	2.67	3.18	2.43	2.57
Number of Points	55	33	115	3	6	40	10	10

14.5.8 North Bullion Deposits Grade Estimation

A parent block size of 10 m x 10 m x 3 m is deemed appropriate based on the drill-hole spacing, which ranges from 2 to 120 m in the mineral resource area. Sub-blocking was used to more effectively honor the volumes and shapes created during the geological interpretation of the mineralized zones. Grade was interpolated for the parent blocks and


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assigned to the sub-blocks. Each block was coded with a “lode” number, representing the different mineralized wire-frame volumes, so that grade could be estimated as hard boundaries.

The estimation of gold grades into the blocks was done using inverse distance squared (“ID²”) for each of the zones, and with ordinary kriging. Due to the resultant validation of the block models showing a better correlation of the input composites grades, it was decided to use the ID²estimation technique for reporting. Each wire-frame zone was estimated with ‘hard boundaries’, which means that only composite assays located within each zone were used to estimate the grade of the blocks within that zone.

There were four estimation passes run for each zone. The size of the anisotropic search ellipsoid was based on the ranges obtained from the variography. The orientation of the search ellipsoid was based on the orientation of the mineralized zones. Estimation runs 1 to 3 equated to ranges less than or equal to the maximum range observed in the variography. The final estimation run search range was expanded to ensure all blocks were estimated with grade. The criteria for the number of composites selected from the number of drill holes decreased with each run, as the search ellipsoid size increased. The estimation criteria for each pass are provided in Table 14-53. The number of composites used for the estimation of each block was capped to 20 composites and a maximum of three composites from any given hole was used.

Table 14-53: North Bullion Area Estimation and Search Ellipsoid Criteria

(from Dufresne and Nicholls, 2018)


Zone	Run No.	Minimum No. of Holes	Minimum No. of Samples	Maximum No. of Composites from any drill hole	Max No. of Composites	Search Ellipsoid Radius (m)
North Bullion	1	6	12	3	20	50 x 50 x 15
	2	6	12	3	20	50 x 50 x 15
	3	2	2	3	20	70 x 70 x 23
	4	1	1	3	20	280 x 280 x 100
Massif Subzone	1	6	12	3	20	120 x 1200x 25
	2	6	12	3	20	120 x 1200x 25
	3	2	2	3	20	160 x 160 x 33
	4	1	1	3	20	320 x 320 x 66
Sweet Hollow	1	6	12	3	20	50 x 40 x 15
	2	6	12	3	20	50 x 40 x 15
	3	2	2	3	20	70 x 50 x 22
	4	1	1	3	20	280 x 200 x 88
POD	1	6	12	3	20	110 x 65 x 60
	2	6	12	3	20	110 x 65 x 60
	3	2	2	3	20	150 x 85 x 80
	4	1	1	3	20	300 x 170 x 160

14.5.9 North Bullion Block Model Validation

The blocks were visually validated on cross sections comparing block grades to the sample grades for all sections and drill holes. In addition, the block and sample data were compared by zone, easting, northing, and elevation in swath plots. Table 14-54 shows the average gold grade of the composited sample data versus the estimated block-model grade by zone. It can be concluded that the mean grade of the ID²block model data is very close to, or generally slightly lower than, the composited sample data.


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Table 14-54: North Bullion Area Estimated Gold Grades vs. Average Composite Grades

(from Dufresne and Nicholls, 2018)


Zone	No. of Composites	Composite Mean Au g/t	ID2 Model Au g/t
1	45	0.32	0.34
2	67	0.33	0.34
3	269	0.33	0.34
4	3	0.37	0.36
5	15	0.36	0.38
6	43	0.36	0.36
17	183	2.02	1.31
18	53	0.84	0.81
19	405	1.52	1.41
20	15	3.15	2.66
21	132	2.51	1.98
22	2	0.90	0.90
23	2	0.79	0.79
24	410	2.23	1.97
25	69	0.64	0.68
26	213	1.06	1.00
27	9	1.25	1.02
28	7	1.10	1.08
29	2	0.59	0.59
30	11	0.78	0.73
32	51	0.81	0.79
33	12	0.95	0.80
34	286	0.25	0.25
35	860	0.20	0.20
36	107	0.17	0.17
37	26	0.21	0.21
38	21	0.19	0.18
40	86	0.22	0.23
41	6	0.29	0.32
42	139	2.44	1.65
43	12	1.45	1.46
44	221	2.29	2.24
45	7	1.44	1.41
46	88	3.39	2.71
47	12	3.73	3.32


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14.5.10 North Bullion Deposits Mineral Resources

The North Bullion area mineral resource estimate has been classified in accordance with guidelines established by the CIM as comprising both Indicated and Inferred mineral resources. The classification was based on geological confidence, data quality and grade continuity (Table 14-55). The most relevant factors used in the classification process were:

Drill hole spacing;
Level of confidence in the geological interpretation;
Estimation parameters (i.e., continuity of mineralization);
Proximity to the recently completed Gold Standard drill holes; and
Drill-hole database data density.

Table 14-55: North Bullion Deposits Classification Criteria

(from Dufresne and Nicholls, 2018)


Criteria	Pod		Sweet Hollow		North Bullion (open pit & UG)	Massif subzone (open pit & UG)
Criteria	Indicated	Inferred	Indicated	Inferred	Inferred	Inferred
Gold Search Distance Ranges (m)	110 x 65 x 60	150 x 85 x 80 to 300 x 170 x 160	50 x 40 x 15	70 x 50 x 22 to 280 x 200 x 88	50 x 50 x 15 to 280 x 280 x 100	120 x 120 x 25 to 320 x 320 x 66
Estimation Run Number	1 or 2	3 or 4	1 or 2	3 or 4	1 to 4	1 to 4
Minimum number of composites used to guide the estimation	12	1 or 2	12	1 or 2	12 to 1	12 to 1
Minimum number of holes used to guide the estimation	3 or 6	1 or 2	3 or 6	1 or 2	1 to 6	1 to 6

Based on the criteria noted above, a polygonal area was chosen for the area with Indicated mineral resources at Sweet Hollow and POD. The blocks in this area were mostly made up of blocks estimated in runs 1 and 2, which utilized a search ellipsoid with a range of three quarters of the maximum range of continuity based on the variography and the highest number of composites and drill holes (Table 14-55). This, in conjunction with the proximity of Gold Standard drilling and level of confidence in the geological interpretation, allowed for assigning a classification of Indicated to this area. The remaining mineral resource was classified as Inferred mineral resources. Although the author of this section is not an expert with respect to environmental, permitting, legal, title, taxation, socio-economic, marketing or political matters, the author is not aware of any unusual factors relating to these matters that may materially affect the North Bullion mineral resources as of the effective date of this Technical Report.

In order to demonstrate that the North Bullion deposits have potential for future economic extraction, the estimate was subjected to a series of Whittle pit optimizations and potential underground mining scenarios. The criteria used in the Whittle pit optimizer were considered reasonable for deposits in Nevada using open-pit mining and heap-leach processing as well as conventional sulfide processing. All reported mineral resources lie within an optimized pit shell and/or underground shell constrained using $1,350/ounce for gold. The criteria used for the $1,350/ounce pit shell and underground optimization are shown in Table 14-56. The open-pit and oxide-constrained mineral resource utilized a >0.1 g Au/t interpreted zone and a 0.14 g Au/t lower cutoff for the POD and Sweet Hollow deposits. The open-pit constrained sulfide (reduced) mineral resource utilized a >0.5 g Au/t interpreted wireframe cutoff and a 1.25 g Au/t lower cutoff. For the North Bullion deposit the open-pit, sulfide-constrained mineral resource utilized a >0.3 g Au/t


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interpreted wireframe/zone cutoff and a 1.25 g Au/t lower cutoff. The underground mineral resource was constrained using a 1 g Au/t interpreted lower cutoff zone. The blocks within this were then estimated and reported at a cutoff of 2.25 g Au/t, which represents expected underground costs associated with a potential block and cave mining scenario. The mineral resource block model and optimized pit shell is illustrated in Figure 14-36.

Table 14-56: Parameters Used in Whittle Pit Optimization Studies

(from Dufresne and Nicholls, 2018)


Parameter	Unit	Pit Oxide Cost	Pit Sulfide Cost	UG Sulfide Cost
Gold price	$/ounce	$1,350	$1,350	$1,350
Gold recover	%	80 %	88 %	88 %
Pit wall angles	50°	50°	50°	NA
Mining Cost	$/tonne	$1.75	$1.75	$50
Ore Density	kg/m³	Ranged from 2.33 to 2.71	2.33 to 2.71	2.33 to 2.71
Waste Density	kg/m³	2.60	2.60	2.60
Processing Rate	5 Mtpa	5 Mtpa	5 Mtpa	5 Mtpa
Processing Cost & Hauling	$/tonne	$3.25	$45	$45
G & A Cost	$/tonne	$0.50	$0.50	$0.50
Discount Rate	5 %	5 %	5 %	5 %

Figure 14-36: North Bullion Area Constrained Mineral Resource Blocks

(within the $1,350/oz Au pit and underground mineral resource, from Dufresne and Nicholls, 2018)

The Indicated and Inferred mineral resources are constrained within a drilled area that extends approximately 2.75 km along strike to the north, 0.95 km to the east and 600 m below surface. The North Bullion area mineral resource estimate is reported at a range of gold cutoff grades in Table 14-57.


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Table 14-57: North Bullion Mineral Resources with Cutoff Grades*

(from Dufresne and Nicholls, 2018)


Classification	Type	Au Cutoff Grade (g/t)	Tonnes (millions)	Au Grade (g/t)	Contained Au (troy oz)*
Pit Constrained
Indicated	Oxide (SH,POD)	0.14	2.92	0.96	90,100
Inferred	Oxide (SH,POD)	0.14	3.36	0.43	46,600
Inferred	Sulfide (SH,POD, NB)	1.25	2.05	2.60	171,400
Inferred Subtotal		0.14, 1.25	5.42	1.25	218,000
Underground
Inferred	Sulfide	2.25	5.55	3.29	587,700
Pit Constrained and Underground Inferred Total
Oxide/Sulfide		0.14, 1.25, 2.25	10.97	2.28	805,800

* Mineral resources are not mineral reserves. Mineral resources which are not mineral reserves do not have demonstrated economic viability. There has been insufficient exploration to define the Inferred mineral Resources tabulated above as an Indicated or Measured mineral resource, 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.

**The recommended reported mineral resources are highlighted in bold and have been constrained within a $US1,350/ounce of gold optimized pit shell and/or an underground mining scenario utilizing a 2.25 g/t Au lower cutoff.

***Contained troy ounces may not add due to rounding.

The Sweet Hollow and POD oxide Indicated and Inferred mineral resources use a cutoff grade of 0.14 g Au/t, which is constrained within an optimized pit shell. The base-case cutoff grades of 0.14 g Au/t for open-pit oxide and 1.25 g Au/t for open-pit sulfide material, and 2.25 g Au/t for underground sulfide material are highlighted in Table 14-57. Other cutoff grades are presented for review ranging from 0.14 g Au/t to 0.5 g Au/t for oxide are presented in Table 14-58. Other cutoff grades are presented in Table 14-59 for review ranging from 1.0 g Au/t to 2.0 g Au/t for sulfide open-pit-constrained material and 2.0 g Au/t to 3.0 g Au/t for sulfide underground-constrained material. The open pit and underground mineral resources are constrained within a $1,350/oz Au pit shell and a 2.25 g Au/t cutoff for the underground.

Table 14-58: Cutoff Sensitivity Analysis of the Sweet Hollow and POD Oxide Mineral Resources*

(from Dufresne and Nicholls, 2018)


Classification*	Au Cutoff Grade (g/t)	Tonnes (millions)	Average Au Grade (g/t)	Contained Au (troy oz)*
Indicated (Oxide)	0.14**	2.92	0.96	90,100
	0.2	2.51	1.09	87,800
	0.3	1.64	1.53	80,800
	0.4	1.41	1.73	78,400
	0.5	1.40	1.74	78,200
Inferred (Oxide)	0.14**	3.36	0.43	46,600
	0.2	2.71	0.49	43,000
	0.3	1.47	0.70	33,200
	0.4	1.14	0.81	29,700
	0.5	1.09	0.82	28,800

* Mineral resources are not mineral reserves. Mineral resources which are not mineral reserves do not have demonstrated economic viability. There has been insufficient exploration to define the Inferred mineral resources tabulated above as an Indicated or Measured mineral resource, 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.

**The recommended reported mineral resources are highlighted in bold and have been constrained within a $US1,350/ounce of gold optimized pit shell.

***Contained troy ounces may not add due to rounding.


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Table 14-59: Cutoff Sensitivity Analysis of the North Bullion, Sweet Hollow and POD Sulfide Mineral Resources*

(from Dufresne and Nicholls, 2018)


Classification	Au Cutoff Grade (g/t)	Tonnes (millions)	Average Au Grade (g/t)	Contained Au (troy ounces)****
Inferred (Near Surface Sulfide)	1.0	2.20	2.50	176,700
	1.25**	2.05	2.60	171,400
	1.5	1.85	2.73	162,200
	1.75	1.66	2.86	152,400
	2.0	1.35	3.09	133,800
Inferred (Underground Sulfide)	2.0	6.87	3.07	678,000
	2.25***	5.55	3.29	587,700
	2.5	4.31	3.55	492,800
	2.75	3.15	3.90	394,900
	3.0	2.52	4.16	336,300

* Mineral resources are not mineral reserves. Mineral resources which are not mineral reserves do not have demonstrated economic viability. There has been insufficient exploration to define the Inferred mineral resources tabulated above as an Indicated or Measured mineral resource, 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.

** The recommended reported mineral resources are highlighted in bold and have been constrained within a $US1,350/ounce of gold optimized pit shell.

*** The recommended reported mineral resources are highlighted in bold and have been constrained within a $US1,350/ounce of gold underground mining scenario.

**** Contained troy ounces may not add due to rounding.


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15 MINERAL RESERVE ESTIMATES

15.1 INTRODUCTION

Mr. Dyer classifies mineral reserves in order of increasing confidence into Probable and Proven categories to be in accordance with the “CIM Definition Standards - For Mineral Resources and Mineral Reserves” (2014), and therefore NI 43-101. Mineral reserves for the Pinion and Dark Star deposits were developed by applying relevant economic criteria in order to define the economically extractable portions of the current mineral resources. CIM standards require that modifying factors be used to convert mineral resources to mineral reserves. The standards define modifying factors and Proven and Probable mineral reserves with CIM’s explanatory material shown in italics as follows:

Mineral Reserve

Mineral reserves are sub-divided in order of increasing confidence into Probable mineral reserves and Proven mineral reserves. A Probable mineral reserve has a lower level of confidence than a Proven mineral reserve.

A mineral reserve is the economically mineable part of a Measured and/or Indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at preliminary feasibility or feasibility level as appropriate that include application of modifying factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which mineral reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

The public disclosure of a mineral reserve must be demonstrated by a preliminary feasibility study or feasibility study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of diluting material that will be mined in conjunction with the Mineral Reserves and delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the Qualified Person prepares a Mineral Reserve. For example, most metal deposits disclose mineral reserves with a “mill feed” reference point. In these cases, mineral reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. In contrast, coal reserves have traditionally been reported as tonnes of “clean coal”. In this coal example, mineral reserves are reported as a “saleable product” reference point and include reductions for plant yield (recovery). The Qualified Person must clearly state the ‘reference point’ used in the Mineral Reserve estimate.

Probable Mineral Reserve


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A Probable mineral reserve is the economically mineable part of an Indicated mineral resources, and in some circumstances, a Measured mineral resource. The confidence in the modifying factors applying to a Probable mineral reserve is lower than that applying to a Proven mineral reserve.

The Qualified Person(s) may elect, to convert Measured Mineral Resources to Probable Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a preliminary feasibility study.

Proven Mineral Reserve

A Proven mineral reserve is the economically mineable part of a Measured mineral resource. A Proven mineral reserve implies a high degree of confidence in the modifying factors.

Application of the Proven mineral reserve category implies that the Qualified Person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a preliminary feasibility study. Within the CIM Definition standards the term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.

Modifying Factors

Modifying Factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted to mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

The author of this section has used Measured and Indicated mineral resources as the basis to define mineral reserves for both the Dark Star and Pinion deposits. Mineral reserve definition was done by first identifying ultimate pit limits using economic parameters and pit optimization techniques. The resulting optimized pit shells were then used for guidance in pit design to allow access for equipment and personnel. Mr. Dyer then considered mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social, and governmental factors for defining the estimated mineral reserves.

Dark Star has been designed using two pit phases. Pinion has been designed using three pit phases. Mr. Dyer used the phased pit designs for both deposits to define the project production schedule, which was then used for cash-flow analysis for the preliminary feasibility study. The final cash-flow model was produced by M3 Engineering and demonstrates that the deposits make a positive cash flow and are reasonable with respect to statement of mineral reserves for those deposits.

15.2 PIT OPTIMIZATION

Pit optimizations were completed by first identifying economic and geometrical parameters. This was followed by defining cutoff grades to evaluate, and then running pit optimizations and economic analysis within various optimized pit shells.


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15.2.1 Economic Parameters

Economic parameters were used to generate optimized pits using a Lerchs-Grossman algorithm within Whittle™ software (Version 4.7). The economic parameters include mining costs, process cost, general and administrative costs (“G&A”), refining costs, royalties, and metal recoveries. Mine planning is an iterative process, and initial costs and recoveries were assumed to determine how large pits would be. The economic parameters were refined as concepts were developed on how material would be processed from the different deposits. The methods for processing that were determined include:

Use of ROM (no crushing) for lower grade oxide and transition material from Dark Star and Pinion;
Using HPGR, or primary crushing followed by HPGR, for higher grade oxide and transition material from Dark Star and Pinion; and
Toll processing of high-grade sulfide material from Dark Star that would be sent to a roaster-type process facility.

The economic parameters used are shown in Table 15-1. The overall process rate is assumed to be 20,000 tonnes per day or 7,200,000 tonnes per year. This is used for each process, though the final processing rates vary between the various process types and deposits on a monthly basis. The assumption here is only used to convert the fixed G&A component to a cost per tonne for the purpose of pit optimization. The G&A cost is later applied as a fixed cost in the cash-flow model.

Table 15-1: South Railroad Economic Parameters


	Dark Star			Pinion
	ROM	HPGR	Toll Roasting	ROM	HPGR	Units
Mining - Waste	$2.00	$2.00	$2.00	$2.00	$2.00	$/tonne Mined
Incremental Ore Mining Cost	$0.20	$0.20	NA	$0.20	$0.20	$/tonne Processed
Crushing & Stack	NA	$0.20	NA	NA	$0.20	$/tonne Processed
Leaching	$1.90	$4.28	NA	$1.90	$4.28	$/tonne Processed
Toll Processing	NA	NA	$38.00	NA	NA	$/tonne Processed
G&A Cost per Tonne	$0.56	$0.56	$0.56	$0.56	$0.56	$/tonne Processed
Refining - Au	$5.00	$5.00	$ -	$5.00	$5.00	$/oz Produced
Refining - Ag	NA	NA	NA	$0.50	$0.50	$/oz Produced
Royalty	By Area	By Area	By Area	By Area	By Area

Royalties were applied by royalty area or region as provided by Gold Standard. These are described in Section 4.2

Recoveries were applied in detail based on recommendations by Mr. Gary Simmons, the Qualified Person for Section 13 of this Technical Report. Most of the recoveries used are based on grade-dependent equations. In order to simplify the equations, they were separated into various ROM and HPGR equations for the different deposits and material types.

15.2.1.1 Dark Star Recoveries

Off-site roaster toll processing is assumed for the Dark Star sulfide material, which would be shipped to one of the Nevada mines that has capacity. This uses a straight 85% recovery for gold.

Dark Star equations were provided based on mineral resource model blocks classified as low- and high- silica in the deposit. Separate equations were provided for both Dark Star North and Dark Star Main and were also varied for oxide and transition material. Thus, there are eight separate gold recovery equations for Dark Star material referred to as ROM1, ROM2, etc.


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The same scheme was used for the recoveries for HPGR material, thus defining equations for HPGR1, HPGR2, etc. The definitions follow those of the ROM for the material shown above.

The resulting ROM equations are shown in Table 15-2 and the HPGR equations are provided in Table 15-3. “HG” in the equations to follow equals “head grade”.

Table 15-2: Dark Star ROM Recovery Equations for Gold


*North Dark Star*	Oxidation	Lith/Material	Equation
ROM1	Oxide	Low Silc	IF(HG<=0.4,0.0267LN(HG)+0.8659,0.01746LN(HG)+0.85848)
ROM2	Oxide	High Silc	IF(HG<=0.4,0.1305LN(HG)+0.7953,0.06736LN(HG)+0.74592)
ROM3	Transition	Low Silc	IF(HG<=0.4,0.0573LN(HG)+0.6971,0.0088LN(HG)+0.6534)
ROM4	Transition	High Silc	IF(HG<=0.4,0.0312LN(HG)+0.5995,0.0244LN(HG)+0.5940)
*Dark Star Main*
ROM5	Oxide	Low Silc	IF(HG<=0.4,0.0366LN(HG)+0.9040,0.00559LN(HG)+0.87606)
ROM6	Oxide	High Silc	IF(HG<=0.4,0.0254LN(HG)+0.7805,0.00389LN(HG)+0.76104)
ROM7	Transition	Low Silc	IF(HG<=0.4,0.0467LN(HG)+0.7037,0.0071LN(HG)+0.6681)
ROM8	Transition	High Silc	IF(HG<=0.4,0.0885LN(HG)+0.6687,0.0588LN(HG)+0.646)

Table 15-3: Dark Star HPGR Recovery Equations for Gold


*North Dark Star*	Oxidation	Lith/Material	Equation
HPGR1	Oxide	Low Silc	IF(HG<=0.4,0.0268LN(HG)+0.8862,0.0175LN(HG)+0.8788)
HPGR2	Oxide	High Silc	IF(HG<=0.4,0.1305LN(HG)+0.8346,0.0673LN(HG)+0.786)
HPGR3	Transition	Low Silc	IF(HG<=0.4,0.0573LN(HG)+0.6971,0.0088LN(HG)+0.6534)+0.091
HPGR4	Transition	High Silc	IF(HG<=0.4,0.0312LN(HG)+0.6907,0.0244LN(HG)+0.6852)
*Dark Star Main*
HPGR5	Oxide	Low Silc	IF(HG<=0.4,0.0366LN(HG)+0.918,0.0056LN(HG)+0.8901)
HPGR6	Oxide	High Silc	IF(HG<=0.4,0.0254LN(HG)+0.8317,0.0039LN(HG)+0.8124)
HPGR7	Transition	Low Silc	IF(HG<=0.4,0.0467LN(HG)+0.7037,0.0071LN(HG)+0.6681)+0.051
HPGR8	Transition	High Silc	IF(HG<=0.4,0.0885LN(HG)+0.7194,0.0587LN(HG)+0.6968)

15.2.1.2 Pinion Recoveries

Pinion recoveries are based on rock types, block model modeled barium content, modeled silica zones, and oxidation types. Zero recoveries are assumed for all Pinion sulfide materials. Block model rock type codes used to define the various recovery equations include MLBX, Devil’s Gate (“DgD”), MTP, and Other (not MLBX, DgD, or MTP). For Pinion ROM material, a total of four oxide equations and four transition equations were used. Table 15-4 shows the recovery equation names and a description of the material they are applied to along with the equations used.


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Table 15-4: Pinion ROM Recovery Equations


Equation	Oxidation	Lith/Material	Equation
ROM1	Oxide	DgD	IF(HG<0.4,11.823LN(HG)+81.691,2.257ln(HG)+72.763)/100
ROM2	Oxide	MlBx Lo Si	IF(HG<0.4,10.203LN(HG)+68.038,6.9059ln(HG)+65.295)/100
ROM3	Oxide	MlBx Hi Si	IF(HG<0.4,8.3765LN(HG)+52.119,1.612ln(HG)+45.803)/100
ROM4	Oxide	MTP	IF(HG<0.4,6.7859LN(HG)+66.608,1.3059ln(HG)+61.492)/100
ROM5	Transition	DgD	IF(HG<0.4,11.823LN(HG)+53.897,2.257ln(HG)+44.969)/100
ROM6	Transition	MlBx Lo Si	IF(HG<0.4,10.203LN(HG)+57.371,6.9059ln(HG)+54.627)/100
ROM7	Transition	MlBx Hi Si	IF(HG<0.4,8.3765LN(HG)+42.53,1.612ln(HG)+36.214)/100-0.10
ROM8	Transition	MTP	IF(HG<0.4,6.7859LN(HG)+58.057,1.3059ln(HG)+52.941)/100-0.10

HPGR equations were provided using the same material criteria as the ROM recoveries, with the exception of DgD material which is only considered for ROM processing. Table 15-5 shows the equations names, material criteria, and equations used for HPGR recoveries.

Table 15-5: Pinion HPGR Recovery Equations


Equation	Oxidation	Lith/Material	Equation
HPGR1	Oxide	DgD	None to use on DgD
HPGR2	Oxide	MlBx Lo Si	IF(HG<0.4,10.203LN(HG)+75.036,6.9059ln(HG)+72.292)/100
HPGR3	Oxide	MlBx Hi Si	IF(HG<0.4,8.3765LN(HG)+75.339,1.612ln(HG)+69.023)/100
HPGR4	Oxide	MTP	IF(HG<0.4,6.7859LN(HG)+79.446,1.3059ln(HG)+74.33)/100
HPGR5	Transition	DgD	None to use on DgD
HPGR6	Transition	MlBx Lo Si	IF(HG<0.4,10.203LN(HG)+64.368,6.9059ln(HG)+61.624)/100
HPGR7	Transition	MlBx Hi Si	IF(HG<0.4,8.3765LN(HG)+42.53,1.612ln(HG)+36.214)/100
HPGR8	Transition	MTP	IF(HG<0.4,6.7859LN(HG)+70.896,1.3059ln(HG)+65.779)/100

15.2.2 Geometric Parameters

Geometric parameters include land constraints and slope parameters. No land boundaries were used other than royalty areas as required to apply NSR royalties to the economics.

Slope recommendations were provided by Golder Associates (“Golder”) (Golder, 2019). These were given using different sectors for both the Dark Star and Pinion deposits. Golder provided two sets of recommendations for each deposit based on whether or not best-case blasting practices are used. MDA has applied the recommendations assuming that best blasting practices will be used to protect high walls from damage.

15.2.2.1 Dark Star Slope Recommendations

Dark Star slope sectors provided by Golder (2019) are shown in Figure 15-1 and the recommended bench heights, catch bench widths, bench face angles (“BFA”), and inner-ramp slope angles (“IRA”) are shown in Table 15-6.

The recommended bench heights and bench widths were provided in feet; however, the project and designs are in metric units. MDA used 9 m bench heights for design with 2 benches between catch benches making the geotechnical bench height 18 m. Catch benches 8.2 m wide were applied along with the BFA to achieve the corresponding IRA from the Golder recommendations.


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The slope sectors were flagged into the mineral resource model and exported to Whittle. For pit optimizations, the sectors were used with the recommended IRA, however the slopes on the west side of Dark Star North were flattened by 5° to represent the flattening due to inclusion of ramps on the west side of the pit.

Figure 15-1: Dark Star Slope Sectors

(from Golder, February 2019)

Table 15-6: Dark Star Slope Recommendations by Sector

(from Golder, February 2019)


Sector	Bench Height *	Bench Width *	BFA (°)	IRA (°)
NP1	60	27	63	46
NP2	60	27	72	52
NP3	60	27	69	52
NP4	60	27	72	52
NP5	60	27	69	50
SP6	60	27	69	50
SP7	60	27	72	52
SP8	60	27	67	49
SP9	60	27	63	46

*Bench height and withs were provided in feet


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15.2.2.2 Pinion Slope Recommendations

Pinion slope sectors provided by Golder are shown in Figure 15-2 and the recommended bench heights, catch bench widths, BFA, and IRA are shown in Table 15-7: Pinion Slope Recommendations by Sector.

As with Dark Star, the recommended bench heights and bench widths were provided in feet, however the project and designs are in metric units. MDA used 9 m bench heights for deign with 2 benches between catch benches making the geotechnical bench height 18 m. Catch benches 8.2 m wide were applied along with the BFA to achieve the corresponding IRA from Golder recommendations.

For Whittle pit optimizations, the IRA was applied. Unlike Dark Star North, the final designs in Pinion will be done without leaving a ramp in the high wall. Thus, no additional flattening of the IRA was required for the pit optimization runs.

Figure 15-2: Pinion Slope Sectors

(from Golder, February 2019)


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Table 15-7: Pinion Slope Recommendations by Sector

(from Golder, February 2019)


Sector	Bench Height	Bench Width	BFA	IRA
1	60	27	69	50
2	60	27	72	52
3	60	27	72	52
4	60	27	63	46
5	60	27	72	52
6	60	27	72	52
7	60	27	63	46
8	60	27	69	50
9	60	27	72	52

Sector 9 was not established as there was not significant highwall. Slopes are assumed based on the steeper sectors in the table above.

15.2.3 Cutoff Grades

Cutoff grades were calculated based on the economic parameters shown in Table 15-1. These were calculated for the different deposits and material types for the various potential processing methods. ROM, crush and agglomerate, and toll processing cutoff grades were each calculated as internal break-even cutoffs. The internal cutoff grade calculation eliminates the mining cost in the calculation. The pit designs are based on economical pits and the materials inside of the pits are assumed to be mined whether the material is waste or ore. Thus, the decision on whether to process the material is made at the point where the truck needs to turn either to the waste dump or the process facility. Thus, the mining cost is a sunk cost. The basic equation for the cutoff grade calculation is shown in Equation 5.

Equation 5 Breakeven Cutoff Grade Calculation (g Au/t)

Where costs are all processing costs plus G&A costs in $/t, RefCst is the refining cost in $/oz gold produced, Roy% is the NSR royalty, and Rec% is the calculated recovery at the cutoff grade.

HPGR break-even cutoff grades were also calculated. Most material fed through the HPGR process could also be fed through the ROM process. Therefore, the cutoff grade for HPGR material was estimated based on where the material would create the most value. As HPGR has higher recoveries, higher-grade material will likely benefit by being processed using HPGR. However, when the higher costs are considered, there will typically be some material that will benefit more by being processed by ROM.

For this reason, the “crossover” cutoff grade has been calculated. The crossover cutoff grade is the grade where the value of material sent to the HPGR process is equal to the value that is created if the same grade is sent to the ROM pad. This is based on the delta of the operating costs and the recoveries at a specified metal price.

Equation 6 Crossover Cutoff Grade Calculation


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Of note, when calculating the breakeven cutoff grades for ROM material, the cutoff grade can be very low and approach assay detection limits. Processing material with grades at the detection limits runs the risk that material may be sent to the leach pad that will incur more costs than the value it creates. Due to this lack of confidence in assays for the lower grades, the PFS uses a minimum grade of 0.17 g Au/t.

The calculated ROM breakeven cutoff grades are between 0.08 and 0.12 g Au/t, so the reporting cutoff grades used are 0.17 g Au/t for all Dark Star ROM material processed. While the HPGR breakeven cutoff grades are still fairly low, the crossover cutoff grades were used. Table 15-7 shows the crossover cutoff grades for Dark Star.

Table 15-8: Dark Star Cutoff Grades


*North Dark Star*	Oxidation	Lith/Material	COG g Au/t
HPGR1	Oxide	Low Silc	0.83
HPGR2	Oxide	High Silc	0.36
HPGR3	Transition	Low Silc	0.39
HPGR4	Transition	High Silc	0.48
*Dark Star Main*
HPGR5	Oxide	Low Silc	0.82
HPGR6	Oxide	High Silc	0.69
HPGR7	Transition	Low Silc	0.55
HPGR8	Transition	High Silc	0.41

The Pinion cutoff grades are shown in Table 15-8 by oxidation, rock type, barite content, and silica reference. ROM cutoff grades are shown as either the breakeven cutoff grades or the 0.17 g Au/t minimum cutoff, whichever is greater. The HPGR cutoff grades are calculated as crossover cutoff grades. As noted with recoveries, DgD material is not to be processed in the HPGR, thus no cutoff grade is given for this material.

Table 15-9: Pinon Breakeven Cutoff Grades


				COG (g Au/t)
ROM Eq	HPGR Eq	Oxidation	Lith/Material	ROM	HPGR COG
ROM1	HPGR1	Oxide	DgD	0.17	NA
ROM2	HPGR2	Oxide	MlBx Lo Si	0.17	0.37
ROM3	HPGR3	Oxide	MlBx Hi Si	0.17	0.24
ROM4	HPGR4	Oxide	MTP	0.17	0.33
ROM5	HPGR5	Transition	DgD	0.19	NA
ROM6	HPGR6	Transition	MlBx Lo Si	0.21	0.30
ROM7	HPGR7	Transition	MlBx Hi Si	0.29	0.49
ROM8	HPGR8	Transition	MTP	0.18	0.36

The ROM cutoff grades described above were used for minimum values in the Whittle optimizations. Both the ROM and HPGR cutoff grades above were used for final mineral reserve definition.


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15.2.4 Pit Optimization Methods and Results

Pit optimizations were run using Whittle™ software (version 4.7). Inputs into Whittle included the mineral resource block model along with the economic and geometric parameters previously discussed. Pit optimizations used for mineral reserve definition used only Measured and Indicated mineral resources for processing and all Inferred material is considered as waste. Each deposit was run separately, and ultimate pit shells were selected from the Whittle results for final design. For Dark Star and Pinion, additional pit shells were considered for guidance of interior pit phases.

The selections of ultimate pits and pit phases were done as a two-step process. The first step was to optimize a set of pit shells based on varying a revenue factor. This was done in Whittle using a Lerchs-Grossman algorithm. The revenue factor was multiplied by the recovered ounces and the metal prices, creating a nested set of pit shells based on different metal prices. Revenue factors for each of the deposits were varied from 0.30 to 2.0 in. increments of 0.025 with a base price of $1,000 per ounce of gold, so the resulting pit shells represent gold prices from $300 to $2,000 per ounce in increments of $25.00. This has the potential of generating up to 69 different pit shells that can be used for analysis.

The second step of the process was to use the Pit by Pit (“PbP”) analysis tool in Whittle to generate a discounted operating cash flow (note that capital is not included). This used a rough scheduling by pit phase for each pit shell to generate the discounted value for the pit. The program develops three different discounted values: best, worst, and specified. The best-case value uses each of the pit shells as pit phases or pushbacks. For example, when evaluating pit 20, there would be 19 pushbacks mined prior to pit 20, and the resulting schedule takes advantage of mining more valuable material up front to improve the discounted value. Evaluating pit 21 would have 20 pushbacks; pit 22 would have 21 pushbacks and so on. Note that this is not a realistic case as the incremental pushbacks would not have enough mining width between them to be able to mine appropriately, but this does help to define the maximum potential discounted operating cash flow.

The worst case does not use any pushbacks in determining the discounted value for each of the pit shells. Thus, each pit shell is evaluated as if mining a single pit from top to bottom. This does not provide the advantage of mining more valuable material first, so it generally provides a lower discounted value than that of the best case.

The specified case allows the user to specify pit shells to be used as pushbacks and then schedules the pushbacks and calculates the discounted cash flow. This is more realistic than the base case as it allows for more mining width, though the final pit design will have to ensure that appropriate mining width is available. The specified case has been used for each mine to determine the ultimate pit limits to design to, as well as to specify guidelines for designing pit phases.

15.2.4.1 Dark Star Pit Optimization

The previously discussed parameters were used along with gold prices varying from $300 to $2,000 per ounce to create the pit optimization results. These results are shown in Table 15-10 using $100 gold price increments with the addition of the $1,275 pit shell which is highlighted as the base price used for pit designs. The pit optimization used the IRA slopes provided by Golder, but the slopes on the west side of the pit were flattened by 5° to better represent the ramp design.

Table 15-11 lists the PbP results by process type and these are also shown graphically in Figure 15-3. Pit 39 is highlighted as having the best discounted operating cash flow for the specified case. This pit was used as a guide for the ultimate pit design. The final design was done using two pit phases, one for Dark Star North, which has the higher value, and another for Dark Star Main.


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Table 15-10: Dark Star Pit Optimization Results


	Price $/os Au	Material Processed			Waste K Tons	Total K Tons	Strip Ratio
Pit	Price $/os Au	K Tons	g Au/t	K Ozs Au	Waste K Tons	Total K Tons	Strip Ratio
1	$300	7,564	1.65	401	22,698	30,262	3.00
5	$400	11,532	1.54	570	35,702	47,234	3.10
9	$500	14,049	1.39	628	38,472	52,522	2.74
13	$600	17,085	1.24	682	41,226	58,312	2.41
17	$700	19,126	1.17	722	45,202	64,328	2.36
21	$800	22,149	1.08	769	49,760	71,909	2.25
25	$900	23,943	1.04	799	53,932	77,875	2.25
29	$1,000	26,980	0.97	845	62,163	89,142	2.30
33	$1,100	27,747	0.96	860	65,596	93,343	2.36
37	$1,200	28,656	0.95	876	69,362	98,018	2.42
40	$1,275	29,122	0.95	887	72,505	101,626	2.49
41	$1,300	29,276	0.95	892	73,809	103,086	2.52
45	$1,400	29,696	0.94	899	75,148	104,844	2.53
49	$1,500	30,105	0.93	905	76,674	106,779	2.55
53	$1,600	30,724	0.93	915	80,394	111,118	2.62
57	$1,700	30,971	0.92	919	81,435	112,406	2.63
61	$1,800	31,102	0.92	921	81,893	112,995	2.63
65	$1,900	31,260	0.92	924	82,670	113,930	2.64
69	$2,000	31,468	0.92	928	83,928	115,395	2.67


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Table 15-11: Dark Star Pit by Pit Results


	HPR Material Processed			ROM Material Processed			Toll Material Processed			Total Material Processed			Waste K Tons	Total K Tons	Strip Ratio	Disc. Op Cash Flow (M USD)
Pit	K Tons	g Au/t	K Ozs Au	K Tons	g Au/t	K Ozs Au	K Tons	g Au/t	K Ozs Au	K Tons	g Au/t	K Ozs Au	Waste K Tons	Total K Tons	Strip Ratio	Best	Specified	Worst
22	2,911	3.14	294	19,765	0.73	461	310	2.85	28	22,987	1.06	783	50,623	53,534	2.20	$ 510.02	$ 486.49	$ 476.48
23	2,943	3.12	295	19,981	0.72	464	314	2.85	29	23,238	1.05	788	51,349	54,292	2.21	$ 511.19	$ 487.65	$ 476.86
24	2,950	3.12	296	20,371	0.72	469	319	2.82	29	23,641	1.04	794	52,190	55,141	2.21	$ 512.57	$ 488.98	$ 476.69
25	2,996	3.10	298	20,725	0.71	475	330	2.80	30	24,051	1.04	803	53,824	56,820	2.24	$ 514.52	$ 490.87	$ 477.21
26	3,028	3.08	300	20,910	0.71	477	334	2.81	30	24,271	1.03	807	54,747	57,775	2.26	$ 515.44	$ 491.77	$ 477.41
27	3,030	3.08	300	22,138	0.69	493	338	2.79	30	25,505	1.00	824	57,522	60,552	2.26	$ 518.45	$ 494.46	$ 474.42
28	3,068	3.06	302	22,394	0.69	497	345	2.79	31	25,807	1.00	830	58,947	62,014	2.28	$ 519.55	$ 495.53	$ 474.37
29	3,084	3.05	303	23,593	0.68	512	358	2.81	32	27,035	0.97	847	62,108	65,192	2.30	$ 522.35	$ 498.13	$ 472.68
30	3,137	3.03	305	23,767	0.67	514	365	2.79	33	27,269	0.97	852	63,329	66,466	2.32	$ 523.11	$ 498.85	$ 472.89
31	3,137	3.03	305	23,838	0.67	515	365	2.79	33	27,340	0.97	853	63,422	66,559	2.32	$ 523.19	$ 498.93	$ 472.78
32	3,139	3.03	305	23,965	0.67	517	365	2.79	33	27,469	0.97	855	63,778	66,917	2.32	$ 523.39	$ 499.09	$ 472.52
33	3,198	3.00	308	24,221	0.67	520	376	2.76	33	27,796	0.96	862	65,546	68,745	2.36	$ 524.14	$ 499.79	$ 472.31
34	3,198	3.00	308	24,380	0.67	522	376	2.76	33	27,955	0.96	864	65,988	69,186	2.36	$ 524.32	$ 499.91	$ 471.78
35	3,200	2.99	308	24,486	0.67	524	380	2.75	34	28,066	0.96	865	66,350	69,550	2.36	$ 524.43	$ 499.98	$ 471.40
36	3,236	2.97	309	24,828	0.66	528	389	2.74	34	28,453	0.95	872	68,081	71,316	2.39	$ 524.84	$ 500.32	$ 470.46
37	3,270	2.96	311	25,015	0.66	531	400	2.72	35	28,685	0.95	877	69,333	72,603	2.42	$ 525.04	$ 500.47	$ 469.81
38	3,283	2.95	312	25,036	0.66	531	404	2.71	35	28,723	0.95	878	69,535	72,818	2.42	$ 525.07	$ 500.48	$ 469.72
39	3,365	2.92	316	25,285	0.66	535	426	2.68	37	29,076	0.95	887	72,399	75,763	2.49	$ 525.20	$ 500.53	$ 468.74
40	3,365	2.92	316	25,327	0.66	535	430	2.66	37	29,122	0.95	887	72,505	75,869	2.49	$ 525.21	$ 500.51	$ 468.54
41	3,399	2.90	317	25,418	0.66	536	444	2.64	38	29,262	0.95	892	73,824	77,223	2.52	$ 525.16	$ 500.45	$ 468.05
42	3,401	2.90	318	25,505	0.66	537	446	2.64	38	29,352	0.95	893	74,149	77,550	2.53	$ 525.13	$ 500.41	$ 467.57
43	3,404	2.90	318	25,603	0.65	539	446	2.64	38	29,454	0.94	894	74,564	77,968	2.53	$ 525.08	$ 500.34	$ 467.03
44	3,406	2.90	318	25,770	0.65	541	448	2.63	38	29,625	0.94	896	75,169	78,575	2.54	$ 524.97	$ 500.20	$ 466.08
45	3,406	2.90	318	25,783	0.65	541	452	2.62	38	29,641	0.94	897	75,202	78,608	2.54	$ 524.95	$ 500.18	$ 466.00
46	3,410	2.90	318	26,003	0.65	543	452	2.62	38	29,864	0.94	899	75,918	79,328	2.54	$ 524.74	$ 499.94	$ 464.77
47	3,411	2.90	318	26,079	0.65	544	452	2.62	38	29,943	0.94	900	76,372	79,783	2.55	$ 524.62	$ 499.81	$ 464.21
48	3,417	2.90	318	26,149	0.65	545	452	2.62	38	30,018	0.93	902	76,736	80,153	2.56	$ 524.51	$ 499.68	$ 463.83
49	3,417	2.90	318	26,153	0.65	545	452	2.62	38	30,022	0.93	902	76,758	80,175	2.56	$ 524.50	$ 499.68	$ 463.80


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Figure 15-3: Dark Star Pit by Pit Graph


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15.2.4.2 Pinion Pit Optimization

The Pinion optimization parameters were used along with variable gold prices to create the pit optimization results. These results are shown in Table 15-14 using $100 gold price increments with the addition of the $1,275 pit shell, which is highlighted as the base price used for the PFS study. Pit optimizations used the previously discussed Golder IRA slope criteria.

Table 15-15 shows the PbP results and these are also shown graphically in Figure 15-4. This shows the material processing type as selected by Whittle. Pit 40 is highlighted as having the best discounted operating cash flow for the specified case and pit 40 is highlighted as the $1,275 Au price pit shell which was chosen as the basis for pit designs.

It is worth noting the various steps in Figure 15-4 which illustrates the difficulty in overcoming stripping at certain metal prices. One of the larger jumps is between pit shells 35 and 36. The incremental contained ounces of gold between those pits are approximately 100,000 ounces.

Table 15-12: Pinion Pit Optimization Results


			Material Processed					Waste K Tons	Total K Tons	Strip Ratio
Pit	Au Price	Ag Price	K Tons	oz Au/ton	K Ozs Au	oz Ag/ton	K Ozs Ag	Waste K Tons	Total K Tons	Strip Ratio
1	$300	$3.88	709	1.341	31	6.29	143	466	1,175	0.66
5	$400	$5.18	1,635	1.072	56	6.38	335	1,099	2,734	0.67
9	$500	$6.47	2,474	0.947	75	6.27	499	1,691	4,165	0.68
13	$600	$7.76	3,321	0.874	93	6.28	671	2,692	6,013	0.81
17	$700	$9.06	4,235	0.812	111	6.03	821	3,865	8,100	0.91
21	$800	$10.35	4,919	0.780	123	5.74	908	5,233	10,152	1.06
25	$900	$11.65	5,732	0.748	138	5.38	991	7,129	12,860	1.24
29	$1,000	$12.94	8,027	0.712	184	5.01	1,292	14,864	22,892	1.85
33	$1,100	$14.24	9,459	0.706	215	4.93	1,499	21,666	31,125	2.29
37	$1,200	$15.53	17,303	0.635	353	4.65	2,584	50,828	68,131	2.94
40	$1,275	$16.50	18,444	0.623	370	4.60	2,729	54,416	72,861	2.95
41	$1,300	$16.82	18,659	0.621	373	4.61	2,763	55,066	73,725	2.95
45	$1,400	$18.12	19,652	0.611	386	4.53	2,863	58,127	77,779	2.96
49	$1,500	$19.41	20,389	0.601	394	4.47	2,931	59,965	80,354	2.94
53	$1,600	$20.71	21,296	0.593	406	4.40	3,015	63,186	84,483	2.97
57	$1,700	$22.00	22,600	0.581	423	4.31	3,129	68,300	90,900	3.02
61	$1,800	$23.29	23,856	0.575	441	4.23	3,245	75,311	99,167	3.16
65	$1,900	$24.59	26,621	0.566	485	4.15	3,556	94,217	120,839	3.54
69	$2,000	$25.88	27,683	0.558	497	4.10	3,652	99,127	126,810	3.58


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Table 15-13: Pinion Pit by Pit Results


	HPGR Processed					ROM Processed					Total Material Processed					Waste K Tonnes	Total K Tonnes	Strip Ratio	Disc. Op Cash Flow (M USD)
Pit	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag	Waste K Tonnes	Total K Tonnes	Strip Ratio	Best	Specified	Worst
22	2,841	0.97	89	7.91	723	2,335	0.49	36	2.64	198	5,176	0.75	125	5.53	921	5,042	10,218	0.97	$ 67.01	$ 67.00	$ 67.00
23	2,929	0.96	91	7.83	737	2,479	0.48	38	2.59	206	5,408	0.74	129	5.42	943	5,498	10,905	1.02	$ 68.11	$ 68.10	$ 68.10
24	3,034	0.96	94	7.75	756	2,633	0.48	40	2.55	216	5,667	0.74	134	5.34	972	6,269	11,936	1.11	$ 69.40	$ 69.37	$ 69.37
25	3,164	0.95	97	7.62	775	2,739	0.48	42	2.52	222	5,903	0.73	139	5.25	997	6,953	12,856	1.18	$ 70.54	$ 70.49	$ 70.49
26	3,240	0.95	99	7.56	788	2,796	0.47	43	2.51	226	6,036	0.73	141	5.22	1,013	7,306	13,342	1.21	$ 71.09	$ 71.03	$ 71.03
27	3,410	0.95	104	7.52	824	3,117	0.47	47	2.43	244	6,527	0.72	151	5.09	1,068	8,904	15,431	1.36	$ 73.04	$ 72.94	$ 72.94
28	3,475	0.94	105	7.51	839	3,192	0.46	48	2.42	248	6,667	0.71	153	5.07	1,087	9,183	15,850	1.38	$ 73.45	$ 73.33	$ 73.33
29	4,330	0.91	127	7.21	1,004	3,836	0.47	58	2.36	291	8,166	0.70	184	4.93	1,295	14,629	22,795	1.79	$ 78.72	$ 78.49	$ 78.25
30	4,370	0.91	128	7.19	1,011	3,878	0.47	58	2.36	294	8,248	0.70	186	4.92	1,305	14,966	23,214	1.81	$ 78.99	$ 78.75	$ 78.49
31	4,885	0.91	143	7.15	1,122	4,203	0.46	63	2.38	322	9,088	0.70	205	4.94	1,444	19,161	28,248	2.11	$ 81.76	$ 81.45	$ 80.88
32	5,015	0.91	147	7.11	1,147	4,358	0.46	65	2.41	338	9,373	0.70	211	4.93	1,484	20,665	30,038	2.20	$ 82.60	$ 82.27	$ 81.61
33	5,080	0.91	149	7.09	1,157	4,483	0.46	66	2.39	345	9,562	0.70	215	4.89	1,502	21,474	31,036	2.25	$ 82.98	$ 82.65	$ 81.94
34	5,148	0.91	151	7.04	1,165	4,622	0.46	68	2.36	351	9,769	0.70	219	4.83	1,516	22,246	32,016	2.28	$ 83.31	$ 82.96	$ 82.20
35	5,218	0.91	153	7.01	1,177	4,720	0.45	69	2.34	356	9,938	0.69	221	4.80	1,532	22,698	32,636	2.28	$ 83.50	$ 83.14	$ 82.33
36	8,481	0.82	223	6.58	1,795	7,126	0.43	99	2.58	591	15,607	0.64	322	4.75	2,386	43,678	59,284	2.80	$ 87.56	$ 87.17	$ 84.19
37	8,759	0.82	231	6.53	1,840	7,728	0.43	107	2.58	641	16,487	0.64	338	4.68	2,481	47,283	63,770	2.87	$ 88.22	$ 87.52	$ 84.27
38	9,103	0.82	240	6.52	1,907	8,402	0.43	115	2.56	692	17,505	0.63	355	4.62	2,599	51,246	68,751	2.93	$ 88.86	$ 87.98	$ 84.61
39	9,122	0.82	241	6.51	1,909	8,490	0.42	116	2.55	696	17,612	0.63	357	4.60	2,605	51,417	69,029	2.92	$ 88.88	$ 87.99	$ 84.61
40	9,394	0.82	247	6.57	1,984	8,840	0.42	120	2.57	730	18,234	0.63	367	4.63	2,714	54,113	72,347	2.97	$ 88.89	$ 87.69	$ 84.13
41	9,444	0.82	249	6.57	1,995	8,925	0.42	121	2.57	737	18,370	0.63	369	4.63	2,732	54,630	73,000	2.97	$ 88.86	$ 87.63	$ 84.05
42	9,515	0.82	250	6.58	2,012	9,022	0.42	122	2.57	744	18,538	0.62	372	4.62	2,756	55,345	73,883	2.99	$ 88.80	$ 87.47	$ 83.85
43	9,726	0.82	255	6.53	2,041	9,294	0.42	125	2.56	764	19,019	0.62	380	4.59	2,805	57,287	76,307	3.01	$ 88.51	$ 86.93	$ 83.17
44	9,732	0.82	255	6.53	2,042	9,350	0.42	125	2.56	770	19,081	0.62	381	4.58	2,812	57,407	76,489	3.01	$ 88.48	$ 86.90	$ 83.12
45	9,789	0.82	257	6.52	2,052	9,562	0.42	128	2.57	790	19,351	0.62	384	4.57	2,842	58,365	77,716	3.02	$ 88.23	$ 86.52	$ 82.68
46	9,808	0.82	257	6.51	2,054	9,622	0.41	128	2.56	793	19,430	0.62	386	4.56	2,847	58,691	78,121	3.02	$ 88.15	$ 86.41	$ 82.54
47	9,839	0.81	258	6.51	2,060	9,741	0.41	130	2.56	802	19,580	0.62	387	4.55	2,862	59,093	78,673	3.02	$ 87.99	$ 86.21	$ 82.32
48	9,875	0.81	258	6.51	2,067	9,886	0.41	132	2.56	813	19,761	0.61	390	4.53	2,881	59,819	79,580	3.03	$ 87.73	$ 85.85	$ 81.91
49	9,895	0.81	259	6.51	2,070	9,943	0.41	132	2.56	818	19,838	0.61	391	4.53	2,888	59,966	79,804	3.02	$ 87.64	$ 85.74	$ 81.79


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Figure 15-4: Pinion Pit by Pit Graph


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15.3 PIT DESIGNS

Detailed pit designs were completed for Dark Star and Pinion using Surpac™ software (version 6.7). Each of the designs utilize 9 m benches with a catch bench installed every other bench, or 18 m. All catch benches were designed with a width of 8.2 m and the BFA’s used are shown in Table 15-6 and Table 15-7.

15.3.1 Road and Ramp Design

Road designs have been completed for the PFS to allow primary access for people, equipment, and consumables to the site. This includes haul roads between the designed pits, dumps, and proposed leach facility. Within the pit designs, ramps have been established for haul truck and equipment access. The in-pit ramps will only require a single berm. Ramps outside of the pit will require two safety berms. The design parameters for ramps and roads are shown in Table 15-14. Note that these also show parameters for one-lane traffic. These would be used near the bottom of pits where the strip ratio is minimal, and the traffic requirements are low.

The ramps and haul roads assume the use of CAT-785 haul trucks with an operating width of 6.64 m. For two-way access the goal of the road design is to allow a running width of near 3.5 times the width of the trucks. Mine Safety and Health Administration (“MSHA”) regulations specify that safety berms be maintained with heights at least ½ of the diameter of the tires of the haul trucks that will travel on roads. The ½ height of the CAT-785 haul trucks tires is 1.52 m. An extra 10% was added to berm height design to ensure that all berms are a sufficient height.

Safety berms assume a slope of 1.5 horizontal to 1.0 vertical. Considering that ramps in the pit only need one berm, the road width of 28 m was determined for two-lane traffic, which allows for 3.42 times the operating width of the haul trucks. Single-lane traffic roads are estimated to require 18 m which allows 1.92 times the operating width of the CAT-785 haul trucks.

Roads outside of the pit will require two berms and widths are estimated to be 34 m allowing 3.53 times the width of the CAT-785 haul trucks.

Road designs are intended to have a maximum of 10% gradient, though some may exceed this for short distances around inside turns. Where switchbacks are utilized, the centerline gradient is reduced to about 8%. This keeps the inside gradient approximately 12%. Switchback designs have not added the detail for super elevation through the curves, but is it assumed that this will be done when they are constructed.


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Table 15-14: Road and Ramp Design Parameters


	Two-Lane	Two-Lane	One-Lane
	In-Pit	Ex-Pit	In-Pit
	Meters	Meters	Meters
Truck Width	6.64	6.64	6.64
Running / Truck Width Ratio	3.50	3.50	2.00
Road Running Width	23.24	23.24	13.28
Tire Size	33.00R51	33.00R51	33.00R51
Tire 1/2 Height	1.52	1.52	1.52
Berm Height	1.67	1.67	1.67
Berm Top Width	0.25	0.25	0.25
Berm Slope	1.50	1.50	1.50
Berm Bottom Width	5.27	5.27	5.27
# Berms	1.00	2.00	1.00
Total Berm Width	5.27	10.53	5.27
Overall Width	28.51	33.77	18.55
*Design Width*	*28.00*	*34.00*	*18.00*
Running Width After Berms	22.73	23.47	12.73
Running Width / Truck Width	3.42	3.53	1.92

15.3.2 Dark Star Pit Designs

Dark Star pit designs were completed using two pit phases. Phase 1 mines Dark Star North and phase 2 mines Dark Star Main. Due to the nature of these zones, they are each individual pit phases. Dark Star North has higher grades and better value, which is why it will be mined first. Figure 15-5 shows the ultimate Dark Star pit design.


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Figure 15-5: Dark Star Ultimate Pit Design


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15.3.3 Pinion Pit Designs

The Pinion ultimate pit design was achieved using three pit phases. The ultimate pit design is shown in Figure 15-6. Pinion phase 1 pit is located in the north part of the deposit and mines near surface oxide materials. Due to the lower strip ratio in this area, the phase 1 pit provides good initial value from the deposit. The Pinion phase 1 pit design is shown in Figure 15-7.

The Pinion phase 2 pit is located just south of phase 1 and mines into the major portion of the deposit from north to south. This pit phase was designed based on the optimized pit shell number 35 as described in 15.2.4.2. The Pinion phase 2 design is shown in Figure 15-8.


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Figure 15-6: Pinion Ultimate Pit Design


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Figure 15-7: Pinion Phase 1 Pit Design


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Figure 15-8: Pinion Phase 2 Pit Design


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15.4 DILUTION

The mineral resource block models were completed for both deposits using 9 m x 9 m x 9 m block sizes which is appropriate for use as a selective mining unit. The estimates for gold (and silver at Pinion) have been block diluted to the mineral resource block size. MDA believes that this dilution is appropriate to represent the dilution and ore loss that will be experienced when the blocks are mined.

15.5 PROVEN AND PROBABLE MINERAL RESERVES FOR DARK STAR AND PINION

In-pit Measured and Indicated mineral resources above the cutoff grades used were converted to Proven and Probable mineral reserves respectively. Dark Star Proven and Probable mineral reserves are shown in Table 15-15. The Dark Star pits have a total of 84.3 million tonnes of waste associated with the mineral reserves, and thus have an overall strip ratio of 2.86 tonnes of waste per tonne processed. The in-pit oxide and transition mineral resources are reported using the 0.17 g Au/t cutoff grade. The sulfide mineral resources are reported using a 1.17 g Au/t cutoff grade. Mineral reserve estimates are based on assumptions that include mining, metallurgical, infrastructure, permitting, taxation, and economic parameters. Increasing costs and taxation and lower metal prices will have a negative impact on the quantity of estimated mineral reserves. There are no other known factors that may have a material impact on the mineral reserve estimates at Dark Star and Pinion.

Table 15-15: Dark Star In-Pit Proven and Probable Mineral Reserves


	HPGR Material			ROM Material			Toll Material			Total Processed
Phase	K Tonnes	g Au/t	K Ozs Au	K Tonnes	g Au/t	K Ozs Au	K Tonnes	g Au/t	K Ozs Au	K Tonnes	g Au/t	K Ozs Au
North Dark Star	11,176	1.67	599	5,252	0.30	50	357	2.75	32	16,786	1.26	680
Dark Star Main	4,608	0.74	110	8,047	0.36	93	15	1.50	1	12,671	0.50	204
Total	15,785	1.40	709	13,300	0.33	143	372	2.70	32	29,456	0.93	884

Pinion Proven and Probable mineral reserves are shown in Table 15-14. The Pinion mineral reserves are associated with a total of 63.0 million tonnes of waste, resulting in a stripping ratio of 3.52 waste tonnes to processed tonnes. Cutoff grades used for reporting are variable based on the material type, oxidation, barite, and silica content.

The reference point for the Dark Star Proven and Probable mineral reserves is at the process facility. The Dark Star Proven and Probable mineral reserves are entirely within the current Measured and Indicated Dark Star mineral resources.

For the Pinion Proven and Probable mineral reserves the reference point is at the process facility, and the mineral reserves are entirely within the current Measured and Indicated Pinion mineral resources.

Table 15-16: Pinion In-Pit Mineral Resources


	HPGR Material					ROM Material					Total Processed
Phase	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag
Pin_Ph_1	1,734	0.81	45	4.60	256	910	0.36	11	2.09	61	2,644	0.66	56	3.73	317
Pin_Ph_2	5,334	0.80	136	6.12	1,049	1,988	0.41	26	2.70	172	7,322	0.69	162	5.19	1,222
Pin_Ph_3	5,683	0.67	123	5.26	961	2,238	0.31	23	2.85	205	7,922	0.57	146	4.58	1,167
Total	12,752	0.74	305	5.53	2,267	5,136	0.36	59	2.66	439	17,887	0.63	364	4.70	2,705

The total Proven and Probable mineral reserves reported for the PFS are shown in Table 15-17. Within the designed pits there are a total of 147.3 million tonnes of waste associated with the in-pit mineral resources. This results in an overall project strip ratio of 3.11 tonnes of waste for each tonne of material processed.


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Table 15-17: Total Dark Star and Pinion Proven and Probable Mineral Reserves


*Dark Star*	K Tonnes	g Au/t	K Ozs Au
Proven	5,434	1.39	243
Probable	24,023	0.83	641
P&P	29,456	0.93	884


*Pinion*	K Tonnes	g Au/t	K Ozs Au	g Ag/t	K Ozs Ag
Proven	1,081	0.66	23	5.49	191
Probable	16,806	0.63	341	4.65	2,514
P&P	17,887	0.63	364	4.70	2,705

Consolidated Gold Reserves


*Dark Star & Pinion*	K Tonnes	g Au/t	K Ozs Au
Proven	6,515	1.27	266
Probable	40,829	0.75	982
P&P	47,344	0.82	1,248

Note: cutoff grades are applied by material type as described in Section 15.2.3;
Pinion Proven and Probable mineral reserves for Pinion include silver as reported above; and
Due to lack of silver at Dark Star, consolidated gold mineral reserves are reported without silver to avoid reporting erroneous average silver grade.


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16 MINING METHODS

The PFS for the Railroad-Pinion project includes mining at both the Dark Star and Pinion deposits. These are planned as open-pit, truck and shovel operations. The truck and shovel method provides reasonable costs and selectivity for these deposits.

The methodology used for mine planning to define the economics for the PFS includes:

Define assumptions for the economic parameters;
Define geometric parameters and constraints;
Run pit optimizations;
Define road and ramp parameters;
Create pit designs;
Create dump designs;
Produce mine and process production schedules;
Define personnel and equipment requirements;
Estimate mining costs; and
Perform an economic analysis.

Parameters, pit optimizations, and pit designs is discussed in Section 15.

16.1 WASTE ROCK STORAGE AREAS AND LEACH PADS

Dump designs were created for the PFS to contain the material that is not processed. MDA has defined NAG and PAG waste, and coded it into the mineral resource block models, based on definitions provided by Stantec. This material has been handled separately to avoid storage issues with potential acid drainage. A 1.3 swell factor was assumed which provides for both swell when mined and compaction when placed into the facility. The total requirements for containment of waste and leach material are shown in Table 16-1.

Table 16-1: Waste Containment Requirements (Thousands, Cubic Meters)


*Dark Star*	PAG	NAG	UnKnown	Total
North Dark Star	27,078	7,604	0	34,682
Dark Star Main	9,796	198	6	10,000
Total Dark Star	36,874	7,802	6	44,682
*Pinion*
Phase 1	1,054	4,393	112	5,558
Phase 2	6,844	6,561	3	13,408
Phase 3	7,859	6,066	181	14,106
Total Pinion	15,757	17,020	296	33,073
*Total Project*
Dark Star	36,874	7,802	6	44,682
Pinion	15,757	17,020	296	33,073
Total	52,631	24,822	302	77,754

Waste Storage Facilities (“WSF”) designs were completed for both Dark Star and Pinion. There is sufficient room for waste storage.


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The PFS for Dark Star assumes that two waste WSF’s will be constructed, one on the east and one on the west side of the deposit. These are shown in Figure 15-5 along with the ultimate pit designs.

Pinion uses a single exterior WSF and will also use some minimal storage as backfill in phase 2 and phase 2. The WSF designs for Pinion are shown in Figure 15-6.

16.2 STOCKPILES

Stockpiles of lower-grade material will be utilized to ensure that higher-grade material is processed first. The stockpiles will be maintained near the crusher. It is expected that a smaller portion of the stockpile will be utilized as a location for trucks to dump material when the crusher becomes temporarily unavailable.

All ROM material will be dumped in place directly on the ROM leach pad.

16.3 MINE-PRODUCTION SCHEDULE

Production scheduling was completed using Geovia’s MineSched™ (version 9.1) software. Proven and Probable mineral reserves were scheduled to process facilities or stockpiles, while waste material was scheduled to WSF’s or backfill locations.

The production schedule considers the processing of material by ROM, HPGR, and toll processing. Monthly periods were used to create the production schedule with pre-stripping starting in Dark Star at month -9. Start of ROM processing is assumed to be month 1. The maximum rate for ROM processing will be 12,500 tonnes per day or 4,562,500 tonnes per year on a 365-day basis. This represents the rate that material can be sprayed and processed. Note that during the first year of processing the maximum rate is not met. However, in years 3 and 4 more material is mined than can be processed. This is reflected as a stockpile. However, the material is assumed to be placed on the leach pad until spraying can be done to liberate the gold. This is to be incorporated into the stacking plan done by process engineers and assumes that sufficient area is available to leave the material in a single lift until spraying capacity is available.

HPGR processing is started in month 13, or 1 year after the start of ROM processing. The delayed start of HPGR processing is due to the higher cutoff grades associated with Dark Star HPGR material. The cutoff grade is higher due to good recoveries from ROM material, especially in Dark Star North. Thus, the amount of HPGR material is minimal during initial mining. Some of the HPGR material is processed as ROM material prior to starting the HPGR process.

The maximum HPGR process rate is targeted at 10,000 tonnes per day or 3,650,000 tonnes per year on a 365-day per year basis. A material stockpile is used to store some initial HPGR material before the process is started, and then through the mine life. The maximum size of the stockpile is 100,000,000 tonnes on a yearly basis.

The toll processing rate is limited to 500 tonnes per day as the amount to be hauled on the highways to another facility with a roaster. Only Dark Star material above 1.17 g Au/t would be toll processed and the material would be stockpiled on site near Dark Star prior to being loaded in a contractor’s over-the-road haul truck. Toll processing occurs from year 2 through year 5.

The monthly mining production for Dark Star and Pinion is summarized yearly in Table 16-2 and Table 16-3 respectively. Table 16-4 summarizes the yearly total mine production schedule. Figures showing yearly pit and WSF position maps are presented in Figure 16-1 through Figure 16-11.


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Table 16-2: Dark Star Mine Production Schedule


		Units Pre-Prod		Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Total
Schedule Production Mine Star Dark	Rom Mined	K Tonnes	390	1,580	2,643	3,292	4,288	1,106	-	-	-	-	13,300
		g Au/t	0.27	0.27	0.31	0.35	0.36	0.36	-	-	-	-	0.33
		K Ozs Au	3	14	26	37	49	13	-	-	-	-	143
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Pit to HPGR StkPl	K Tonnes	400	1,854	2,244	2,385	550	151	-	-	-	-	7,583
		g Au/t	0.96	1.07	0.76	1.08	0.46	0.46	-	-	-	-	0.92
		K Ozs Au	12	64	55	83	8	2	-	-	-	-	224
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Pit to HPGR StkPl	K Tonnes	-	-	2,822	3,372	1,671	337	-	-	-	-	8,202
		g Au/t	-	-	2.18	2.18	0.79	0.79	-	-	-	-	1.84
		K Ozs Au	-	-	197	236	43	9	-	-	-	-	485
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Total HPGR	K Tonnes	400	1,854	5,066	5,757	2,220	489	-	-	-	-	15,785
		g Au/t	0.96	1.07	1.55	1.72	0.71	0.69	-	-	-	-	1.40
		K Ozs Au	12	64	252	319	51	11	-	-	-	-	709
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Pit to Su StkPl	K Tonnes	-	15	67	280	11	0	-	-	-	-	372
		g Au/t	-	1.36	1.95	3.00	1.45	1.20	-	-	-	-	2.70
		K Ozs Au	-	1	4	27	1	0	-	-	-	-	32
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Total Mined Above COG	K Tonnes	789	3,449	7,775	9,328	6,520	1,595	-	-	-	-	29,456
		g Au/t	0.62	0.71	1.13	1.28	0.48	0.46	-	-	-	-	0.93
		K Ozs Au	16	78	283	384	100	23	-	-	-	-	884
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	PAG to Dumps NAG to Dumps Un to Dumps	K Tonnes	12,185	22,837	11,374	13,016	7,378	730	-	-	-	-	67,520
		K Tonnes	3,011	6,408	5,921	1,227	147	-	-	-	-	-	16,714
		K Tonnes	63	1	-	2	8	-	-	-	-	-	74
	Total to Dumps	K Tonnes	15,259	29,246	17,295	14,246	7,533	730	-	-	-	-	84,308
	Total Mined	K Tonnes	16,048	32,695	25,070	23,574	14,053	2,325	-	-	-	-	113,764
	Strip Ratio	K Tonnes	19.34	8.48	2.22	1.53	1.16	0.46					2.86


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Table 16-3: Pinion Mine Production Schedule


		Units Pre-Prod		Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Total
Schedule Production Mine Pinion	Rom Mined	K Tonnes	-	-	-	-	524	1,850	841	1,435	486	-	5,136
		g Au/t	-	-	-	-	0.38	0.41	0.31	0.31	0.34	-	0.36
		K Ozs Au	-	-	-	-	6	24	9	14	5	-	59
		g Ag/t	-	-	-	-	2.32	2.43	3.22	2.94	2.08	-	2.66
		K Ozs Ag	-	-	-	-	39	144	87	136	32	-	439
	Pit to HPGR StkPl	K Tonnes	-	-	-	-	500	3,132	2,735	3,171	898	-	10,436
		g Au/t	-	-	-	-	0.62	0.63	0.60	0.54	0.54	-	0.58
		K Ozs Au	-	-	-	-	10	63	53	55	16	-	196
		g Ag/t	-	-	-	-	3.75	5.64	5.71	5.02	3.71	-	5.21
		K Ozs Ag	-	-	-	-	60	568	502	512	107	-	1,749
	Pit to HPGR StkPl	K Tonnes	-	-	-	-	127	1,037	485	479	188	-	2,315
		g Au/t	-	-	-	-	1.42	1.42	1.42	1.45	1.82	-	1.46
		K Ozs Au	-	-	-	-	6	47	22	22	11	-	109
		g Ag/t	-	-	-	-	8.88	6.61	8.13	7.09	4.10	-	6.95
		K Ozs Ag	-	-	-	-	36	220	127	109	25	-	517
	Total HPGR	K Tonnes	-	-	-	-	627	4,169	3,219	3,650	1,086	-	12,752
		g Au/t	-	-	-	-	0.78	0.82	0.72	0.66	0.76	-	0.74
		K Ozs Au	-	-	-	-	16	110	75	77	27	-	305
		g Ag/t	-	-	-	-	4.79	5.88	6.08	5.29	3.78	-	5.53
		K Ozs Ag	-	-	-	-	97	788	629	621	132	-	2,267
	Pit to Su StkPl	K Tonnes	-	-	-	-	-	-	-	-	-	-	-
		g Au/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Au	-	-	-	-	-	-	-	-	-	-	-
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Total Mined Above COG	K Tonnes	-	-	-	-	1,151	6,019	4,061	5,085	1,571	-	17,887
		g Au/t	-	-	-	-	0.60	0.70	0.64	0.56	0.63	-	0.63
		K Ozs Au	-	-	-	-	22	135	84	91	32	-	364
		g Ag/t	-	-	-	-	3.67	4.82	5.49	4.63	3.25	-	4.70
		K Ozs Ag	-	-	-	-	136	933	716	757	164	-	2,705
	PAG to Dumps NAG to Dumps Un to Dumps	K Tonnes	-	-	-	-	5,524	9,791	13,945	394	39	-	29,693
		K Tonnes	-	-	-	-	7,434	11,031	9,642	3,792	830	-	32,730
		K Tonnes	-	-	-	-	189	33	274	62	0	-	558
	Total to Dumps	K Tonnes	-	-	-	-	13,147	20,856	23,862	4,247	870	-	62,981
	Total Mined	K Tonnes	-	-	-	-	14,297	26,875	27,923	9,333	2,441	-	80,869
	Strip Ratio	K Tonnes					11.42	3.46	5.88	0.84	0.55		3.52


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Table 16-4: Total Project Mine Production Schedule


		Units Pre-Prod		Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Total
Schedule Production Mine Project Railroad Total	Rom Mined	K Tonnes	390	1,580	2,643	3,292	4,812	2,956	841	1,435	486	-	18,435
		g Au/t	0.27	0.27	0.31	0.35	0.36	0.39	0.31	0.31	0.34	-	0.34
		K Ozs Au	3	14	26	37	56	37	9	14	5	-	201
		g Ag/t	-	-	-	-	0.25	1.52	3.22	2.94	2.08	-	0.74
		K Ozs Ag	-	-	-	-	39	144	87	136	32	-	439
	Pit to HPGR StkPl	K Tonnes	400	1,854	2,244	2,385	1,050	3,283	2,735	3,171	898	-	18,019
		g Au/t	0.96	1.07	0.76	1.08	0.53	0.62	0.60	0.54	0.54	-	0.73
		K Ozs Au	12	64	55	83	18	65	53	55	16	-	420
		g Ag/t	-	-	-	-	1.79	5.38	5.71	5.02	3.71	-	3.02
		K Ozs Ag	-	-	-	-	60	568	502	512	107	-	1,749
	Pit to HPGR StkPl	K Tonnes	-	-	2,822	3,372	1,797	1,375	485	479	188	-	10,517
		g Au/t	-	-	2.18	2.18	0.84	1.27	1.42	1.45	1.82	-	1.76
		K Ozs Au	-	-	197	236	48	56	22	22	11	-	594
		g Ag/t	-	-	-	-	0.63	4.99	8.13	7.09	4.10	-	1.53
		K Ozs Ag	-	-	-	-	36	220	127	109	25	-	517
	Total HPGR	K Tonnes	400	1,854	5,066	5,757	2,848	4,658	3,219	3,650	1,086	-	28,536
		g Au/t	0.96	1.07	1.55	1.72	0.73	0.81	0.72	0.66	0.76	-	1.10
		K Ozs Au	12	64	252	319	67	121	75	77	27	-	1,014
		g Ag/t	-	-	-	-	1.06	5.26	6.08	5.29	3.78	-	2.47
		K Ozs Ag	-	-	-	-	97	788	629	621	132	-	2,267
	Pit to Su StkPl	K Tonnes	-	15	67	280	11	0	-	-	-	-	372
		g Au/t	-	1.36	1.95	3.00	1.45	1.20	-	-	-	-	2.70
		K Ozs Au	-	1	4	27	1	0	-	-	-	-	32
		g Ag/t	-	-	-	-	-	-	-	-	-	-	-
		K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-
	Total Mined Above COG	K Tonnes	789	3,449	7,775	9,328	7,671	7,614	4,061	5,085	1,571	-	47,344
		g Au/t	0.62	0.71	1.13	1.28	0.50	0.65	0.64	0.56	0.63	-	0.82
		K Ozs Au	16	78	283	384	123	158	84	91	32	-	1,248
		g Ag/t	-	-	-	-	0.55	3.81	5.49	4.63	3.25	-	1.78
		K Ozs Ag	-	-	-	-	136	933	716	757	164	-	2,705
	PAG to Dumps	K Tonnes	12,185	22,837	11,374	13,016	12,902	10,521	13,945	394	39	-	97,213
	NAG to Dumps	K Tonnes	3,011	6,408	5,921	1,227	7,581	11,031	9,642	3,792	830	-	49,444
	Un to Dumps	K Tonnes	63	1	-	2	197	33	274	62	0	-	632
	Total to Dumps	K Tonnes	15,259	29,246	17,295	14,246	20,679	21,586	23,862	4,247	870	-	147,289
	Total Mined	K Tonnes	16,048	32,695	25,070	23,574	28,350	29,200	27,923	9,333	2,441	-	194,633
	Strip Ratio	K Tonnes	19.34	8.48	2.22	1.53	2.70	2.83	5.88	0.84	0.55		3.11


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Figure 16-1: Dark Star Pit Design, Year -1


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Figure 16-2: Dark Star Pit Design, Year 1


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Figure 16-3: Dark Star Pit Design, Year 2


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Figure 16-4: Dark Star Pit Design, Year 3


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Figure 16-5: Dark Star Pit Design, Year 4


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Figure 16-6: Dark Star Pit Design, Year 5


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Figure 16-7: Pinion Pit Design, Year 4


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Figure 16-8: Pinion Pit Design, Year 5


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Figure 16-9: Pinion Pit Design, Year 6


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Figure 16-10: Pinion Pit Design, Year 7


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Figure 16-11: Pinion Pit Design, Year 8


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The process production schedule was created by MDA based on the mine production schedule and using recoveries and lag times estimated by Mr. Simmons. The recoveries are used to estimate recoverable gold. The lag time is generated by estimating the recovery of the recoverable ounces on a month by month basis after placement of material. Table 16-5 shows the assumed rate of recovery of the recoverable ounces by month for ROM, HPGR, crush and agglomeration, and toll roasting. In each case the recovery of gold is 0% for the month placed. This allows material to be placed, ripped as required, and then start leach spraying. The month after placement generally sees the most gold production with the exception of toll milling. ROM recovery is assumed to take place between the month after placement through the 31^stmonth after placement. HPGR and crush and agglomerate gold is anticipated to be recovered more quickly through a 13-month period. Toll processing is assumed to be paid to account during the third month after mining. This allows time for settlement of payments.

Table 16-5: Recovery of Recoverable Gold by Month


	ROM	HPGR	Crush & Agg	Toll Roasting
Mth Placed	0.0%	0.0%	0.0%	0%
Mth 1	46.9%	67.5%	67.5%	0%
Mth 2	2.5%	8.4%	8.4%	100%
Mth 3	1.9%	4.8%	4.8%	0%
Mth 4	1.9%	3.5%	3.5%	0%
Mth 5	1.9%	2.7%	2.7%	0%
Mth 6	1.9%	2.1%	2.1%	0%
Mth 7	1.9%	1.9%	1.9%	0%
Mth 8	1.9%	1.6%	1.6%	0%
Mth 9	1.9%	1.4%	1.4%	0%
Mth 10	1.9%	1.3%	1.3%	0%
Mth 11	1.9%	1.3%	1.3%	0%
Mth 12	1.9%	1.3%	1.3%	0%
Mth 13	1.7%	1.3%	1.3%	0%
Mth 14	1.7%	0.9%	0.9%	0%
Mth 15	1.7%	0.0%	0.0%	0%
Mth 16	1.7%	0.0%	0.0%	0%
Mth 17	1.7%	0.0%	0.0%	0%
Mth 18	1.7%	0.0%	0.0%	0%
Mth 19	1.7%	0.0%	0.0%	0%
Mth 20	1.7%	0.0%	0.0%	0%
Mth 21	1.7%	0.0%	0.0%	0%
Mth 22	1.7%	0.0%	0.0%	0%
Mth 23	1.7%	0.0%	0.0%	0%
Mth 24	1.7%	0.0%	0.0%	0%
Mth 25	1.7%	0.0%	0.0%	0%
Mth 26	1.7%	0.0%	0.0%	0%
Mth 27	1.7%	0.0%	0.0%	0%
Mth 28	1.7%	0.0%	0.0%	0%
Mth 29	1.7%	0.0%	0.0%	0%
Mth 30	1.4%	0.0%	0.0%	0%
Mth 31	1.3%	0.0%	0.0%	0%
Total	100.0%	100.0%	100.0%	100.0%

Table 16-6 shows the yearly process production summary by process type. The rows labeled “K Au Rec” shows the thousands of recoverable ounces of gold and the rows labeled “K Au Prod” are the thousands of ounces of gold


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produced. MDA has put together the gold production plan, but ultimately the metallurgical and processing consultants are responsible for the final production numbers, which may differ from what is in the cash-flow model.

The PFS total LOM gold production is estimated to be 936,000 ounces with a LOM average recovery of 75%. From Pinion a total of 1,061,000 ounces of silver is recovered with an overall LOM recovery of 39%.

Table 16-6: Railroad-Pinion Process Production Schedule


	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Yr 10	Total
Total ROM	K Tonnes	-	4,013	4,563	4,562	4,575	3,722	841	1,435	486	-	-	24,197
	g Au/t	-	0.67	0.51	0.57	0.36	0.39	0.31	0.31	0.34	-	-	0.48
	K Ozs Au	-	86	75	84	54	47	9	14	5	-	-	373
	K Au Rec	-	61	52	61	37	31	5	8	3	-	-	259
	K Au Prod	-	32	42	58	46	40	18	12	6	2	1	259
	g Ag/t	-	-	-	-	0.24	1.23	3.22	2.94	2.08	-	-	0.56
	K Ozs Ag	-	-	-	-	36	148	87	136	32	-	-	439
	K Ag Rec	-	-	-	-	10	43	16	20	7	-	-	96
	K Ag Prod	-	-	-	-	5	25	21	22	15	6	2	96
Total HPGR	K Tonnes	-	-	2,993	3,650	3,660	3,638	3,650	3,650	1,534	-	-	22,775
	g Au/t	-	-	2.10	2.22	0.70	0.91	0.70	0.66	0.66	-	-	1.15
	K Ozs Au	-	-	202	260	83	106	82	77	32	-	-	842
	K Au Rec	-	-	168	215	60	75	56	52	23	-	-	650
	K Au Prod	-	-	124	230	77	72	61	52	32	1	-	650
	g Ag/t	-	-	-	-	0.82	5.16	6.27	5.48	3.78	-	-	3.10
	K Ozs Ag	-	-	-	-	97	604	736	643	186	-	-	2,267
	K Ag Rec	-	-	-	-	37	257	318	274	78	-	-	965
	K Ag Prod	-	-	-	-	27	204	313	287	128	6	-	965
Dark Star Toll Process	K Tonnes	-	-	23	91	175	83	-	-	-	-	-	372
	g Au/t	-	-	1.72	2.70	2.80	2.76	-	-	-	-	-	2.70
	K Ozs Au	-	-	1	8	16	7	-	-	-	-	-	32
	K Au Rec	-	-	1	7	13	6	-	-	-	-	-	27
	K Au Prod	-	-	0	6	12	9	-	-	-	-	-	27
	g Ag/t	-	-	-	-	-	-	-	-	-	-	-	-
	K Ozs Ag	-	-	-	-	-	-	-	-	-	-	-	-
	K Ag Rec	-	-	-	-	-	-	-	-	-	-	-	-
	K Ag Prod	-	-	-	-	-	-	-	-	-	-	-	-
Total Processed	K Tonnes	-	4,013	7,579	8,304	8,410	7,442	4,491	5,085	2,019	-	-	47,344
Total Processed	g Au/t	-	0.67	1.14	1.32	0.56	0.67	0.63	0.56	0.58	-	-	0.82
	K Ozs Au	-	86	278	352	152	160	91	91	38	-	-	1,248
	K Au Rec	-	61	222	283	110	112	61	61	26	-	-	936
	K Au Prod	-	32	166	294	136	121	80	64	38	4	1	936
	g Ag/t	-	-	-	-	0.49	3.14	5.70	4.77	3.37	-	-	1.78
	K Ozs Ag	-	-	-	-	132	752	823	779	219	-	-	2,705
	K Ag Rec	-	-	-	-	48	300	334	294	85	-	-	1,061
	K Ag Prod	-	-	-	-	32	229	334	309	144	12	2	1,061

Table 16-7, Table 16-8, and Table 16-9 show the ROM, HPGR, and toll roasting stockpile balance sheets. As previously mentioned, the ROM stockpile is anticipated being placed on the pad and awaiting spray, so there would not be any physical stockpile or re-handle for the ROM material.


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Table 16-7: ROM Stockpile Balance


*Total Leach Stockpiled*	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9
Added to StkPl	K Tonnes	390	-	-	106	400	244	-	-	-	-
	g Au/t	0.27	-	-	0.32	0.36	0.35	-	-	-	-
	K Ozs Au	3	-	-	1	5	3	-	-	-	-
	g Ag/t	-	-	-	-	0.25	0.38	-	-	-	-
	K Ozs Ag	-	-	-	-	3	3	-	-	-	-
Removed from StkPl	K Tonnes	-	390	-	-	-	750	-	-	-	-
	g Au/t	-	0.27	-	-	-	0.35	-	-	-	-
	K Ozs Au	-	3	-	-	-	8	-	-	-	-
	g Ag/t	-	-	-	-	-	0.26	-	-	-	-
	K Ozs Ag	-	-	-	-	-	6	-	-	-	-
StkPl Balance	K Tonnes	390	-	-	106	505	-	-	-	-	-
	g Au/t	0.27	-	-	0.32	0.35	-	-	-	-	-
	K Ozs Au	3	-	-	1	6	-	-	-	-	-
	g Ag/t	-	-	-	-	0.20	-	-	-	-	-
	K Ozs Ag	-	-	-	-	3	-	-	-	-	-

Table 16-8: HPGR Stockpile Balance


*Total HPGR Stockpiled*	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9
Added to StkPl	K Tonnes	400	1,348	1,003	1,830	515	3,164	2,059	1,599	438	-
	g Au/t	0.96	1.27	1.03	1.24	0.62	0.62	0.67	0.67	0.70	-
	K Ozs Au	12	55	33	73	10	64	44	35	10	-
	g Ag/t	-	-	-	-	2.85	5.47	6.05	6.13	3.93	-
	K Ozs Ag	-	-	-	-	47	557	401	315	55	-
Removed from StkPl	K Tonnes	-	1,538	850	1,099	1,490	2,404	2,490	1,599	886	-
	g Au/t	-	1.22	1.15	1.74	0.61	0.68	0.64	0.67	0.55	-
	K Ozs Au	-	60	31	62	29	52	51	35	16	-
	g Ag/t	-	-	-	-	0.98	4.81	6.23	6.13	4.09	-
	K Ozs Ag	-	-	-	-	47	372	499	315	117	-
StkPl Balance	K Tonnes	400	210	364	1,094	119	879	448	448	-	-
	g Au/t	0.96	1.08	0.78	0.59	0.50	0.46	0.41	0.41	-	-
	K Ozs Au	12	7	9	21	2	13	6	6	-	-
	g Ag/t	-	-	-	-	-	6.53	5.98	5.98	-	-
	K Ozs Ag	-	-	-	-	-	184	86	86	-	-

Table 16-9: Dark Star Toll Roasting Stockpile Balance


*Total HPGR Stockpiled*	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9
Added to StkPl	K Tonnes	400	1,348	1,003	1,830	515	3,164	2,059	1,599	438	-
	g Au/t	0.96	1.27	1.03	1.24	0.62	0.62	0.67	0.67	0.70	-
	K Ozs Au	12	55	33	73	10	64	44	35	10	-
	g Ag/t	-	-	-	-	2.85	5.47	6.05	6.13	3.93	-
	K Ozs Ag	-	-	-	-	47	557	401	315	55	-
Removed from StkPl	K Tonnes	-	1,538	850	1,099	1,490	2,404	2,490	1,599	886	-
	g Au/t	-	1.22	1.15	1.74	0.61	0.68	0.64	0.67	0.55	-
	K Ozs Au	-	60	31	62	29	52	51	35	16	-
	g Ag/t	-	-	-	-	0.98	4.81	6.23	6.13	4.09	-
	K Ozs Ag	-	-	-	-	47	372	499	315	117	-
StkPl Balance	K Tonnes	400	210	364	1,094	119	879	448	448	-	-
	g Au/t	0.96	1.08	0.78	0.59	0.50	0.46	0.41	0.41	-	-
	K Ozs Au	12	7	9	21	2	13	6	6	-	-
	g Ag/t	-	-	-	-	-	6.53	5.98	5.98	-	-
	K Ozs Ag	-	-	-	-	-	184	86	86	-	-


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16.4 EQUIPMENT SELECTION AND PRODUCTIVITIES

The PFS has assumed owner mining in order to keep the cost lower than it would be with contract mining. The production schedule was used along with additional efficiency factors, cycle times, and productivity rates to develop the first principle hours required for primary mining equipment to achieve the production schedule. Primary mining equipment includes drills, loaders, hydraulic shovels, and CAT-785 style haul trucks.

The mine is anticipated to operate 24 hours per day utilizing four crews of workers, each working four days on and four days off. It is anticipated that these crews would rotate between day shift and night shift. The daily shift schedule would be 12 hours per day, reduced to account for standby time including startup/shutdown, lunch, breaks, and operational delays totaling 3.0 hours per day. This allows for 21 work hours in each day or 87.5% schedule efficiency. The estimated schedule efficiency is shown in Table 16-10.

Table 16-10: Schedule Efficiency


	Units	Value
Shifts per Day	shift/day	2
Hours per Shift	hr/shift	12
Theoretical Hours per Day	hrs/day	24
Shift Startup / Shutdown	hrs/shift	0.5
Lunch	hrs/shift	0.5
Breaks	hrs/shift	0.25
Operational Standby	hrs/shift	0.25
Total Standby / shift	hrs/shift	1.50
Total Standby / day	hrs/day	3.00
Available Work Hours	hrs/day	21.00
Schedule Efficiency	%	87.5%

16.5 EQUIPMENT REQUIREMENTS

Mine equipment is planned to be purchased over a period of 3 years (pre-production through year 2). This equipment is to be used through the LOM. Table 16-11 shows the yearly purchase schedule for mining equipment.


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Table 16-11: Mine Equipment Purchases


*Primary Mining Equipment*	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Total
Pioneer Drill	#	2	-	-	-	2
Production Drill	#	4	-	-	-	4
26-yrd Loader	#	-	1	-	-	1
23 cu m Hyd. Shovel	#	2	-	-	-	2
136-tonne Haul Trucks	#	9	5	1	-	15
*Support Equipment*
430 Kw Dozer (D10)	#	2	-	-	-	2
300 Kw Dozer (D9)	#	1	-	-	-	1
230 Kw Dozer (D8)	#	1	-	-	-	1
16' Motor Grader (14M)	#	2	-	-	-	2
Water Truck - 20,000 Gallon	#	2	-	-	-	2
Pit Pumps (5299 lpm)	#	2	-	-	-	2
50 ton Crane	#	1	-	-	-	1
Flatbed	#	2	-	-	-	2
*Blasting*
Sanding/Stemming Truck	#	-	-	-	-	-
Explosives Truck	#	1	-	-	-	1
Skid Loader	#	1	-	-	-	1
*Mine Maintenance*
Lube/Fuel Truck	#	1	-	-	-	1
Mechanics Truck	#	2	-	-	-	2
Tire Truck	#	2	-	-	-	2

16.5.1 Drilling Equipment

Pioneer drills would be smaller air-track drills with contained cabs and the production drills are anticipated to be 45,000lb-pulldown, track-mounted, rotary blast-hole drills. An 83% efficiency factor was used for pioneer drilling and 85% efficiency was used for production and controlled blast hole drilling. Penetration rates of 31.6, 31.6, 33.1, and 36.5 meters per hour were used along with 2.8, 2.8, 2.8, and 3.0 minutes per hole of non-drilling times for waste production, ore production, trim-rows, and pioneer drilling, respectively.

Based on the parameters used, two pioneer drills and four production drills are estimated to be needed. It is assumed that these drills will last through the LOM with an availability that is assumed to be 85% for the life of the drill.

Drilling patterns were adjusted by material. The adjustments were made based on studies by Barr Engineering (“Barr”) (2019) to create a nominal size distribution with a P80 of 152 mm (6”) minus. Based on that work, blast patterns where ore is anticipated is estimated to use 5.2 m spacing and 4.5 m burden with 1.52 m sub drill. With 196 mm diameter drill holes and stemming of 3 m, this results in a powder factor of 0.368 kg of explosive per tonne of material blasted. This was determined to be beneficial for gold recovery.

Waste patterns are assumed to have 6.1 m spacing and burden and 0.213 kg of explosive per tonne of material blasted.

The increase in areas of waste is not needed as this material would not be processed.


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During pioneering operations at the start of each deposit, smaller drills will be used due to uneven terrain. At the start of Dark Star mining it is assumed that 50% of blasting will be done as pioneering for the first two months. At the start of Pinion, 10% of the blasting for the first two months is assumed.

Trim row shot patterns are used with lower powder factors and tighter spacing of drill holes near pit high walls to minimize damage to the walls. The PFS assumes that 5% of the waste blasted will be in the form of trim row blasting.

16.5.2 Loading Equipment

Loading equipment is anticipated to include one 16.7 cubic yard type loader and two 23 cubic yard type hydraulic shovels. The theoretical productivity for the loader was estimated to be 2,345 tonnes per hour or 1,950 tonnes per hour after an operating efficiency of 83%. The loader is primarily used for back-up mining production and re-handle of material from stockpiles. The assumed availability starts at 90% and is reduced 1% per year until it reaches 85%, and then is held constant through the life of the shovels. No replacement loaders were assumed. The overall use of available hours is 31%.

Two hydraulic shovels are used as the primary loading tool. The initial shovel starts operating in month -9 and the second shovel starts working in month -5. The theoretical productivity was estimated to be 3,326 tonnes per hour or 2,760 tonnes per hour after applying 83% efficiency. As with the loader, the assumed availability starts at 90% and declines at 1% per year to a low of 85% and then remains the same through the life-of-mine. The overall use of operating hours is 66%.

16.5.3 Haulage Productivity

Haul trucks are assumed to be CAT-875 type, 136-tonne capacity, rigid frame trucks. Haulage profiles were used inside of MineSched with speeds attached to each segment. Speeds were based on percent gradients and haul truck performance curves. The speeds were estimated for up-hill loaded, down-hill loaded, up-hill empty, and down-hill empty. A rolling resistance of 3% was also used for the haulage speed calculations. In addition, bench haulage strings were created which depict the planned haulage routes on each bench where mining occurs.

Certain speed limits were set for the haulage profiles. They include:

Loaded speed limit = 48 kph
Empty speed limit = 48 kph
Loaded down-hill speed limit = 24 kph
Empty down-hill speed limit = 30 kph
Bench travel speed = 24 kph

Hydraulic shovel loading time of 2.67 minutes was used, plus 0.5 minutes and a spot and dump time of 1.5 minutes was added. Loading time was adjusted in spreadsheets to 3.73 minutes for trucks that would be loaded using a loader.

A capacity of 131 tonnes per load was used as dry tonnage to reflect the dry densities in the mineral resource block model. The number of trucks was calculated to increase over time due to farther haulage with some pit phases. A total of 14 haul trucks are purchased to maintain the production schedule. This assumes a 1% per year declining availability from 90% down to 85%.

Support and Maintenance Equipment

Support equipment is used to maintain the roads, pits, and dumps in order for the mining equipment to operate in an efficient manner. The maintenance equipment is used on site to maintain the mining equipment. The total number of equipment to be purchased for use on the site is shown in Table 16-11.


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16.6 MINING PERSONNEL AND STAFFING

Table 16-12 shows the estimated personnel requirements. This is based on the number of people that will be required to operate, supervise, maintain, and plan for operations to achieve the production schedule.

Table 16-12: Personnel Requirements


*Mining General Personnel*	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Max
Mine Superintendent	#	1	1	1	1	1	1	1	1	1	-	1
Mine General Foreman	#	1	1	1	1	1	1	1	1	1	-	1
Mine Foremen	#	4	4	4	4	4	4	4	4	4	-	4
Chief Mine Engineer	#	1	1	1	1	1	1	1	1	1	-	1
Mine Engineer	#	2	2	2	2	2	2	2	2	2	-	2
Chief Surveyor	#	1	1	1	1	1	1	1	1	1	-	1
Surveyor	#	2	2	2	2	2	2	2	2	2	-	2
Chief Geologist	#	1	1	1	1	1	1	1	1	1	-	1
Ore Control Geologist	#	1	1	1	1	1	1	1	1	1	-	1
Samplers	#	2	2	2	2	2	2	2	2	2	-	2
Total Mine General	#	16	16	16	16	16	16	16	16	16	-	16
*Mine Operations Hourly Personnel Operators*
Blasters	#	2	2	2	2	2	2	2	2	2	-	2
Blaster's Helpers	#	2	2	2	2	2	2	2	2	2	-	2
Drill Operators	#	16	16	16	16	20	16	16	12	8	-	20
Loader Operators	#	8	12	12	12	12	12	12	8	6	-	12
Haul Truck Operators	#	36	56	60	60	56	56	44	36	12	-	60
Support Equipment Operators	#	19	19	19	19	19	19	19	19	19	-	19
General Mine Labors	#	-	-	-	-	-	-	-	-	-	-	-
Total Operators	#	83	107	111	111	111	107	95	79	49	-	111
*Mechanics*
Mechanics - Drilling	#	8	8	8	8	10	8	8	6	4	-	10
Mechanics - Loading	#	8	8	8	8	10	8	8	6	4	-	10
Mechanics - Haulage	#	18	28	30	30	28	28	22	18	6	-	30
Mechanics - Support	#	10	10	10	10	10	10	10	10	10	-	10
Total Mechanics	#	44	54	56	56	58	54	48	40	24	-	58
*Maintenance*
Maintenance Superintendent	#	1	1	1	1	1	1	1	1	1	-	1
Maintenance Foreman	#	4	4	4	4	4	4	4	4	4	-	4
Maintenance Planners	#	2	2	2	2	2	2	2	2	2	-	2
Light Vehicle Mechanic	#	2	2	2	2	2	2	2	2	2	-	2
Welder	#	4	4	4	4	4	4	4	4	4	-	4
Servicemen	#	4	4	4	4	4	4	4	4	4	-	4
Tireman	#	2	2	2	2	2	2	2	2	2	-	2
Maintenance Labor	#	4	4	4	4	4	4	4	4	4	-	4
Total Maintenance	#	23	23	23	23	23	23	23	23	23	-	23
Total Personnel - Mining Personnel	#	166	200	206	206	208	200	182	158	112	-	208


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17 RECOVERY METHODS

The process selected for recovery of gold and silver from the Pinion and Dark Star ore is a conventional heap-leach recovery circuit. The ore will be mined by standard open pit mining methods from two separate pits. Lower-grade Pinion and Dark Star ore will be truck-stacked on the heap as ROM ore directly, without crushing. Higher grade Pinion and Dark Star (low clay) material will be processed in a three-stage crushing circuit with a high-pressure grinding roll (HPGR), treated with cement and agglomerated, then conveyor-stacked onto heap leach pads.

Oxide and transition material types will be leached with a dilute cyanide solution, and the leached gold and silver will be recovered from solution using a carbon adsorption circuit. The gold and silver will be stripped from carbon using a desorption process, followed by electrowinning to produce a precipitate sludge. The precipitate sludge will be processed using a retort oven for drying and mercury recovery, and then refined in a melting furnace to produce gold and silver doré bars.

The Pinion and Dark Star deposits have a total estimated mineral reserve of 47.3 million tonnes. Based on material grades within the Pinion and Dark Star mineral reserves, 24.3 million tonnes will be selected for ROM processing and 28.5 million tonnes will be selected for HPGR crushing. The total estimated mine life is 8 years. The nominal processing rate through the crushing circuit is 10,000 tonnes per day, and the design ROM processing rate (for the leaching and adsorption design basis) is 12,500 tonnes per day.

17.1 GOLD RECOVERY

The gold and silver recoveries for heap leaching of the Pinion and Dark Star ore have been taken from the recommendations detailed in Section 13 of this Technical Report.

For the Pinion and Dark Star mineral resources, the overall life-of-mine average gold recovery for the lower grade ROM ore is estimated at 69 percent. The overall life-of-mine average gold recovery for heap leaching of HPGR-crushed material is estimated at 77 percent.

17.2 REAGENTS AND CONSUMPTIONS

The major reagent consumptions for heap leaching of Pinion and Dark Star ore have been taken from available metallurgical test results from column leach tests on crushed material. No test data exists at the ROM particle size, so the selected reagent consumptions have been estimated based on test results on the coarsest samples tests, - 37 mm.

17.2.1 Sodium Cyanide

Sodium cyanide (NaCN) will be used in the leaching process and will be delivered in tanker trucks as a liquid at 30% concentration by weight (1.15 SG). Sodium cyanide will be stored in a 98 m³steel tank at the ADR area within concrete containment and will be distributed by metering pumps to points of use.

All cyanide distribution lines will be double-containment, either by “pipe-within-pipe” or “pipe-over-liner” containment systems. Cyanide consumptions have been estimated as follows:

Pinion ROM – 0.22 kg/tonne ore
Dark Star ROM – 0.23 kg/tonne ore
Pinion HPGR Crushed – 0.22 kg/tonne ore


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Dark Star HPGR Crushed – 0.25 kg/tonne ore

17.2.2 Lime

Pebble quicklime (CaO) will be used to treat the ROM and Pinion crushed ore prior to cyanide leaching to maintain the alkaline pH. Lime will be delivered in bulk by 20-ton trucks, which will be off-loaded pneumatically into an 85 tonne storage silo with a variable speed feeder that will meter lime directly onto the ore being carried by haul trucks to the heap leach pad and will be added in proportion to the tonnage of ore in each truck. A 5 tonne lime silo will be added with the crushing expansion to allow for convenient lime addition to haul trucks from either Pinion or Dark Star pits as well as a 50 tonne lime silo for lime addition to crushed Pinion ore.

Lime will be consumed at an estimated 1.0 kg/tonne ore for the Pinion and Dark Star ROM ore and 0.5 kg/tonne ore for Pinion crushed ore.

17.2.3 Cement

Portland Type II cement will be used to agglomerate the crushed ore prior to cyanide leaching to maintain the permeability and alkaline pH. Cement will be delivered in bulk by 20-ton trucks, which will be off-loaded pneumatically into two-200 tonne storage silos with variable speed feeders that will meter cement directly onto the conveyor belt which feeds the agglomeration drum in proportion to the tonnage of ore.

Cement will be consumed at an estimated 2.0 kg/tonne ore for the Pinion ore and 7.0 kg/tonne for the Dark Star ROM ore. Based on this dosage, the approximate maximum heap height is 60 m.

17.2.4 Activated Carbon

Activated carbon will be used to adsorb precious metals from the leach solution in the adsorption columns. Make-up carbon will be 6 x 12 mesh and will be delivered in 1,000 kg supersacks. It is estimated that approximately 4% of the carbon stripped will have to be replaced due to carbon fines losses.

17.2.5 Sodium Hydroxide (Caustic)

Sodium hydroxide (caustic) will be delivered to site as a liquid at 50% caustic by weight (1.53 s.g.). Liquid caustic will be stored in a 100 m³steel tank and metered to the strip solution tank and acid wash circuits by a caustic metering pump. Caustic has been estimated to be consumed at 340 kg per strip.

17.2.6 Hydrochloric Acid

Hydrochloric acid (32%) will be used in the acid wash section of the elution circuit prior to desorption. Hydrochloric acid (32% by weight, 1.16 s.g.) will be delivered in 1 m³tote bins. Acid washing consists of circulating a dilute acid solution through the bed of carbon to dissolve and remove scale from the carbon. Carbon acid washing will be done before each desorption cycle. Hydrochloric acid has been estimated to be consumed at 200 L/tonne carbon stripped.

17.2.7 Fluxes

Various fluxes will be used in the smelting process to remove impurities from the bullion in the form of a glass slag. The normal flux components are a mix of silica sand, borax, and sodium carbonate (soda ash). The flux mix composition is variable and will be adjusted to meet individual project smelting needs: fluorspar and/or potassium nitrate (niter) are


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sometimes added to the mix. Dry fluxes will be delivered in 50 lb bags. Average consumption of fluxes has been estimated to be 1.0 kg per kg of gold and silver produced.

17.2.8 Antiscalant

Antiscalant will be used to prevent the build-up of scale in the process solutions and heap irrigation lines. Antiscalant will be added directly into pipelines or tanks, and consumption will vary depending on the concentration of scale-forming species in the process stream. Delivery will be in liquid form in 1 m³tote bins.

Antiscalant will be added directly from the supplier tote bins into the pregnant, barren, and desorption pumping systems using variable speed chemical-metering pumps. On average, antiscalant consumption is expected to be about 6 ppm for leach solutions and 10 ppm for strip solutions to be treated.

17.3 PROCESS FLOWSHEET

An overall process flowsheet for the project is presented in Figure 17-1.


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Figure 17-1: Process Flowsheet for the Pinion-Dark Star Project


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17.4 ROM TRUCK STACKING

Excavation, loading, hauling, and dumping of ROM material will be conducted by the mining fleet. ROM ore will be loaded into 100-tonne haul trucks and transported to the active stacking face at an average rate of 12,500 tonnes/day. ROM production and stacking will vary based on the low-grade ore availability from the mine pits.

Quicklime (CaO) will be used for pH control of the process with an estimated consumption of 1.0 kg/tonne for both Pinion and Dark Star based on metallurgical test work. Pebble quicklime will be stored in 85-tonne silos which will be equipped with a variable speed feed system that will feed a clam gate for lime addition to the trucks. Once the haul trucks have been loaded, the lime will be metered directly into the loaded trucks which will then deliver the ore to the active stacking area. One lime silo will be installed at the haul road for both the Pinion and Dark Star mine pits with the Pinion lime silo being purchased as part of the crushing addition. Lime will be added in proportion to the tonnage of ore being hauled.

The ore haul trucks will operate on top of the lift being constructed. A ramp, or ramps, will be constructed to reach the top of each current lift. The trucks will direct-dump the ore on the current lift and a dozer will push the ore over the edge of the lift to form the expanding heap. The stacked ore will be deep-shank cross-ripped with the dozer prior to leaching. Ore will be stacked in 9 m high lifts with a maximum ore heap height of 100 m.

Prior to stacking a new lift over the top of an old one, the top of the old lift will be cross-ripped to break up any cemented/compacted sections and to redistribute any fines that may have been stratified by the irrigation solution or rainfall.

Following stacking, the ore will be drip irrigated with dilute cyanide leach solution and the resulting gold-bearing solutions collected in the pregnant solution tank. The leach pad will be a multiple-lift, single-use type pad.

17.5 CRUSHING

High grade ore from the Pinion and Dark Star pits (approximately 60% of the total mineral reserves) will be crushed beginning in Year 2 of operations in a three-stage crushing circuit. Ore will be transported from the mine in surface haul trucks and dumped either directly into the crusher dump hopper, or at the crusher area stockpile where it will be reclaimed and fed to the crusher hopper by a loader. The dump hopper will be equipped with a static grizzly with 762 mm openings to prevent oversize rocks from being fed to the crushing circuit. Blending will be required for some ore types, especially for higher clay areas. The crushing plant will process an average of 10,000 tonnes of ore per day. A preliminary crushing general arrangement is presented in Figure 17-2.


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Figure 17-2: Crushing General Arrangement


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Ore will be fed from the ROM dump hopper to a vibrating grizzly feeder via an apron feeder. The vibrating grizzly feeder will have parallel bars spaced 100 mm apart with grizzly oversize being fed to the primary jaw crusher and the grizzly undersize being recombined with the jaw crusher product on the primary crusher discharge conveyor. The primary jaw crusher will operate with a 102 mm discharge setting. The primary crusher discharge conveyor transfers primary crushed ore to the primary crushed ore stockpile feed conveyor. The primary crushed ore stockpile has a total capacity of 16,000 tonnes with a live capacity of 4,100 tonnes (approximately 9 hours of operation).

Ore from the primary crushed ore stockpile will be reclaimed by two each vibrating pan feeders onto a secondary screen feed conveyor. A tramp metal electromagnet and metal detector will be installed on the secondary screen feed conveyor to protect the secondary crushers. The secondary screening circuit includes a single double-deck vibrating screen with 76 mm and 50 mm top and bottom deck openings, respectively. Oversize material (+50 mm) will be transferred to the secondary feed conveyor by the screen oversize conveyor and undersize (-50 mm) will be transferred to the secondary product stockpile feed conveyor by the secondary screen undersize conveyor. Oversize material will be crushed by a secondary standard cone crusher which will operate with a 50 mm closed side setting and will discharge onto the secondary crusher product conveyor. The secondary crushing circuit will be operated in closed circuit with the secondary crusher product conveyor feeding the secondary screen feed conveyor; a splitter chute with a flop gate is positioned at the discharge of the secondary crusher product conveyor to bypass secondary cone product from the secondary screen and tertiary crusher circuit by feeding material directly to the tertiary crusher product stacking conveyor. Due to the high abrasiveness of the ore, a second standby cone crusher is included in the design to meet plant availability requirements.

The secondary screen undersize (secondary product) will be stockpiled by the secondary stockpile feed conveyor. The secondary product stockpile will have a total capacity of 16,000 tonnes with a live capacity of 4,100 tonnes. Ore from the secondary product stockpile will be reclaimed by two vibrating pan feeders each onto the HPGR feed conveyor. A tramp metal electromagnet and metal detector will be installed on the HPGR feed conveyor to protect the HPGR (high pressure grinding roll) crusher.

The tertiary crushing circuit will consist of an HPGR crusher with partial recycle. Ore will be fed to the HPGR feed bin which choke feeds the HPGR unit. Approximately 20% of the HPGR product is recycled back to the HPGR feed conveyor. Crushed product from the HPGR, 100% passing 14 mm, is stockpiled by the tertiary crusher product stacking conveyor.

Modular motor control centers will be located in containers near the primary, secondary, and tertiary crushing circuits. A PLC control unit will be located in a central control room which will control and monitor all crushing equipment, as well as monitor the conveyor stacking equipment. All of the conveyors will be interlocked so that if one conveyor trips out, all upstream conveyors and the vibrating grizzly feeder will also trip. This interlocking is designed to prevent large spills and equipment damage. Both of these features are considered necessary to meet the design utilization for the system.

Water sprays will be located at all material transfer points to reduce dust generation by the crushing circuit.

17.6 AGGLOMERATION AND CONVEYOR STACKING

The high-grade crushed product stockpile is sized to accommodate a total capacity of approximately 16,000 tonnes (live capacity of approximately 4,100 tonnes). Crushed ore will be reclaimed from the stockpile by two vibrating pan feeders to a reclaim conveyor in a tunnel below the stockpile. Portland Type II cement (cement) will be added to the reclaim tunnel conveyor at an average rate of 2 kg per tonne of ore for Pinion material and 7 kg/tonne of ore for Dark Star material from one of two 200-tonne silos, each equipped with a bin activator, variable speed rotary feeder, screw conveyor, and dust collector. Pebble lime from a 50-tonne silo will be metered at a rate of 0.5 kg/tonne ore for Pinion ore for additional pH control. The reclaim conveyor feeds a 3.6 x 10 m agglomeration drum where ore and cement are


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mixed with process solution to form agglomerates. Agglomerated ore discharges the agglomeration drum onto a transfer conveyor which feeds the heap leach overland conveyor and conveyor stacking system.

The heap stacking system consists of an overland conveyor (864 mm x 470 m), 19 grasshopper transfer conveyors (864 mm x 35 m), nine ramp grasshopper conveyors (864 mm x 35 m), an index feed conveyor (864 mm x 24 m), horizontal index conveyor (864 mm x 35 m), and a radial stacker (864 mm x 37 m). The overland conveyor feeds material to the grasshopper conveyors in the active stacking zone, which transfer the material to the conveyor stacking system. The conveyor stacking system includes the index feed conveyor, horizontal index, and radial stacker conveyors. The horizontal index and radial stacker are able to retreat and stack ore onto the heap. The number of grasshopper conveyors required varies depending on the area of the heap being stacked with a maximum of 28 grasshopper conveyors being required.

Once an area has finished leaching and is sufficiently drained and dry a new lift can be stacked over the top of the old lift. The old lift will be cross-ripped prior to stacking new material on top of any old heap area or access road/ramp to break up any compacted or cemented sections.

Stacked lifts will progress in a stair-step manner. The maximum planned heap height for crushed and agglomerated ore is 60 m over the composite leach pad liner.

17.7 LEACHING AND SOLUTION HANDLING

After each leach cell has been stacked and dozer ripped, the irrigation system will be installed. Dripline emitters will be used to apply a dilute cyanide solution, at an application rate of 8 L/hr/m²for both ROM and crushed material. A leach cycle of 100 days has been selected for ROM and 50 days for crushed material, based on a review of the leach curves.

Barren leach pH solution will be maintained at a minimum value of 10 and will be controlled by the addition of lime and cement. Barren solution will be delivered from a barren tank located at the recovery plant, by high-flow high-head pumps at the initial nominal flow rate of 650 m³/hr. In Year 1 the nominal flow rate to the heap will increase to 1,000 m³/hr. This solution will be carried by a steel pipeline to the base of the heap and then to a network of sub-headers and risers to the top of the heap where it is finally applied to the material by drip emitters.

Solution passing through the heap will dissolve the values contained therein and be collect in a network of perforated solution collection pipes, which feed to a common discharge point at the base of the heap. The solution will then be carried by gravity to a pregnant solution tank. Excess solution from the heap will overflow from the pregnant tank to a lined process pond. Pregnant solution is pumped from the pregnant tank to the adsorption carbon column circuit at the recovery plant.

The carbon adsorption circuit consists of a series of cascade-style columns. Pregnant solution flows through the columns to load the soluble gold onto the carbon. Barren solution exiting the columns is directed to the barren tank where make up cyanide is added, and the solution returned to the heap for further leaching. Overflow from the barren tank is directed to a barren/event pond.

17.8 LEACH PAD PHASING AND CONSTRUCTION

It is assumed the leach pad will be constructed in four phases with a fifth phase that involves no new construction or liner deployment but consists of post mining gold recovery for a period of up to 2 years. The estimated cumulative lined areas for Phase 1, Phase 2, Phase 3, and Phase 4 are approximately 376,000 m², 558,000 m², 740,000 m², and 923,000 m²respectively, and will contain approximately 47.3 million tonnes of material at the end of Phase 4 with some potential to expand the total tonnage of ore stacking.


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For the initial the first year ROM ore will be stacked with trucks in nominal 9 m thick lifts across the entire eastern toe are of Phase 1 leach pad. Barren solution containing cyanide will be irrigated onto the ore using drip irrigation. Pregnant solution will be collected at the base of the heap by the leach pad liner and collection system, which will route the pregnant solution to the process plant for gold recovery and reagent reconditioning. At the end of year 1, higher grade ore will be crushed and agglomerated and stacked on a dedicated section of the pad closest to the crusher using conveyor stacking system in 7 m thick lifts. The ROM and HPGR ores have different leach cycles and times, hence the need to segregate them on the leach pad. Once an area has been leached for the target time or metal recovery, the next lift will be placed on top of the already leached ore and the process repeated. This will be continued until the heap is stacked to the design elevation of 2100 m for the 47.3 Mt capacity. A limiting total depth for crushed and agglomerated ore is 60m. With the overall side slopes of the heap ore material limited to 3H:1V, the depth of crushed ore limit of 60m is maintained to full build out of the heap facility.

An overliner layer will be provided to protect the geomembrane primary liner from mechanical damage during ore stacking as well as weather conditions before the geomembrane is covered with ore. The overliner will also provide drainage of leach solutions and storm water entering the system both through the permeability of the drainage gravel and a network of drainage pipes installed within the overliner. The overliner material will be 500 mm thick and consist of select, durable crushed ore screened to a P100 of 38 mm.

The primary geosynthetic liner will be a robust, 2.0 mm thick LLDPE material with the bottom side textured to provide an intimate bond with the underlying GCL. The two materials used on nearly every heap leach pad in the industry is high density polyethylene (HDPE) and linear low-density polyethylene (LLDPE). LLDPE has been chosen for its superior resistance to puncturing; overall durability; and proven performance in heap leach applications. LLDPE also maintains flexibility at lower temperatures and has a larger window of ambient temperatures during which it can be installed. Leaks through geomembranes occur through damage rather than inherent permeability or transmissivity. In other words, these materials are impermeable for most practical considerations as they are manufactured. The highest risk period of their service life is not the static load of the heap under leach, but rather construction stresses and especially those caused by installing the liner and placing the overliner. The installation specifications include performance of electrical leak location surveys at the completion of each stage of installation to give the highest assurance of a leak-free installation.

The compacted clay liner (CCL) layer will utilize an on-site clay source to produce a compacted clay liner with identified properties to have a maximum permeability of 10^-6cm/sec.

The leak detection system for the leach pad will consist of wick drains placed directly underneath the primary collection pipes between the prepared subgrade and the CCL liner in each of cells for the leach pad. These wick drains will be extended to and are booted through the perimeter solution collection trench liner system to discharge into the lined solution collection trench 1 m above the trench bottom. This will enable visual monitoring of the individual cells and sampling of the to enable visual monitoring and sampling of the leak detection ports as necessary.

17.8.1 Solution Ponds

Two storage ponds, the process pond and event pond, are planned for the management of solutions. The process pond will collect overflow from the pregnant solution tank and is sized to additionally contain 24 hours of pregnant solution working volume, essentially 24 hours of heap solution drain down in the event of barren pump failure or power loss. The event pond will collect overflow from the barren solution tank during process upsets and is sized to additionally handle storm water collection from a 100 yr., 24-hr storm event, plus the accumulation from a wet year snowpack over the ultimate pad lined area. Based on preliminary assumptions and data, the process and event ponds are sized at approximately 30,023 m³and 97,067 m³respectively for a total storage capacity of 127,090 m³including free board.


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The pond lining system will consist of either a compacted clay soil or GCL underliner, overlain by two HDPE geomembrane liners separated by an HDPE geonet for leak detection and recovery. Solutions collected in these ponds will be pumped back to the corresponding barren or pregnant solution tanks using submersible pond pumps for distribution either to the recovery plant or to the heap.

17.9 ADR PLANT

The recovery plant at South Railroad has been designed to recover gold and silver values using an adsorption-desorption-recovery (“ADR”) process. Pregnant leach solution from the heap leach will be pumped to the carbon in column circuit (“CIC”) and adsorbed onto activated carbon (adsorption). Loaded carbon from the CIC circuit will be desorbed in a high-temperature elution process coupled to an electrowinning circuit (desorption), followed by retorting to remove mercury and smelting of the resulting sludge to produce doré bullion (recovery). Before elution, each batch of carbon will be acid washed to remove any scale and other inorganic contaminants that might inhibit gold adsorption on carbon. All or a portion of the carbon will be thermally reactivated using a rotary kiln.

The ADR plant General Arrangement is presented in Figure 17-3.


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Figure 17-3: ADR Recovery Plant General Arrangement


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17.9.1 Adsorption

Adsorption of gold and silver onto carbon will occur in the carbon adsorption circuit. The adsorption circuit will consist of two trains of five, cascade type open-top up-flow mild-steel CICs each. Each of the carbon columns are nominally 3.5 meters in diameter by 3.9 meters high and are sized to hold 5 tonnes of activated carbon.

For the first year and a half of operation when only ROM ore is being processed on the heap, pregnant solution from the pregnant solution tank will be pumped to the adsorption circuit at a nominal rate of 648 m³/h. Once crushing and agglomeration begins, the nominal flow to the adsorption circuit will increase to 1,000 m³/hr. Barren solution exiting the last carbon adsorption column in the train will flow through a static screen to separate any floating carbon from the solution, then flow by gravity into the barren tank.

Antiscalant will be added at the pregnant solution tank to prevent scaling of carbon and reduction of the carbon loading capability. Magnetic flowmeters equipped with totalizers will measure solution flow to the adsorption circuit. Pregnant solution will flow by gravity through each set of five columns in series, exiting the lowest column as barren solution. Pregnant and barren solution continuous samplers will be installed at the feed and discharge end of each carbon column train, respectively. Solution samples will be used to measure pregnant and barren solution gold and silver concentrations.

Adsorption of gold and silver from pregnant leach solutions from the heap circuit will be a continuous process. Once the carbon in the lead column achieves the desired precious metal load it will be advanced to the elution (desorption) circuit using screw type centrifugal pumps. Carbon in the remaining columns will be advanced counter current to the solution flow to the next column in series. New or acid washed/regenerated carbon will be added to the last column in the train.

Generally, the stripping of carbon will occur each day.

17.9.2 Carbon Acid Wash

Acid washing will consist of circulating a dilute acid solution through the bed of carbon to dissolve and remove scale from the carbon. Acid washing will be performed on a batch basis before every desorption cycle.

After carbon has been transferred into the acid wash column, but before any acid is introduced, fresh water will be circulated through the bed of carbon to remove any entrained caustic/cyanide solution. The rinse solution will be pumped to the carbon safety screen using the acid wash circulation pump. A dilute acid solution will then be prepared in the mix tank, and circulation established between the acid wash vessel and the acid mix tank. Concentrated acid will be injected into the recycle stream to achieve and maintain a pH ranging from 1.0 to 2.0. Completion of the cycle will be indicated when the pH stabilizes between 1.0 and 2.0 without acid addition for a minimum of one full hour of circulation.

After acid washing has been completed, the acid wash pump will pump spent acid solution from the acid mix tank and wash vessel to the carbon safety screen. The carbon will then be rinsed with raw water followed by rinsing with dilute caustic solution to remove any residual acid. Total time required for acid washing a batch of carbon will be four to six hours. After acid washing has been completed, a carbon transfer pump will transfer the carbon to the desorption circuit.

17.9.3 Desorption

A pressure Zadra hot caustic desorption circuit for the stripping of metal values from carbon has been selected for South Railroad, which requires 18 hours or less to complete a cycle. During the elution cycle, gold and silver are continuously extracted by electrowinning from the pregnant eluate concurrently with desorption.


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The desorption circuit is sized to strip gold and silver from carbon in 5-tonne batches and will be equipped with a strip solution tank, strip solution pump, primary (heat up), secondary (heat recovery), and tertiary (cooling) heat exchangers, hot water heater, elution column, and elution column drain pump. After carbon has been transferred to the elution column, barren strip solution (eluant) containing sodium hydroxide and sodium cyanide will be pumped through the heat recovery and primary heat exchangers and introduced to the elution vessel at a nominal temperature of 135°C and a nominal operating pressure of approximately 65 psig.

Under normal operating conditions, barren eluant solution from the solution storage tank will pass through the heat recovery exchanger to be preheated by hot pregnant eluate leaving the elution column. The barren eluant solution then passes through the primary heat exchanger to raise the temperature up to 135°C using pressurized hot water (~180°C) from the hot water heater system.

The elution column will contain internal stainless-steel inlet screens to hold carbon in the column and to distribute incoming stripping solution evenly in the column. Pregnant eluate leaving the elution column will pass through two external stainless-steel screens before passing through the heat recovery exchanger and the cooling heat exchanger to reduce the temperature to about 75°C (to prevent boiling). The cooled pregnant eluate solution will flow to the electrowinning cells.

After desorption is complete, the stripped carbon will be transferred to the carbon regeneration circuit by a carbon transfer pump.

17.9.4 Electrowinning

The electrowinning circuit will be operated in series with the elution circuit. Solution will be pumped continuously from the barren strip solution tank through the elution column, then through the electrowinning cells, and back to the strip solution tank in a continuous closed loop process.

The electrowinning circuit will include two each 2.3 m³electrowinning cells, each equipped with a rectifier. The gold and silver-laden solution exiting the elution column will be filtered to trap any carbon escaping from the column; will pass through the heat recovery exchanger and the cooling exchanger to reduce the solution temperature to 75°C, then will flow to the electrowinning circuit.

Gold and silver will be won from the eluate in the electrowinning cells using stainless steel cathodes using a current density of approximately 50 amperes per square meter of anode surface. Caustic soda (sodium hydroxide) in the eluate solution will act as an electrolyte to encourage free flow of electrons and promote the precious metal winning from solution. To keep the electrical resistance of the solution low during desorption and the electrowinning cycle, make-up caustic soda will sometimes be added to the strip solution tank. Barren eluant solution leaving the electrolytic cells will discharge to the E-cell discharge tank from which it will be pumped back to the eluate storage tank for recycle through the elution column.

Periodically, all or part of the barren eluant will be dumped to the barren solution tank. Typically, about one-third of the barren eluant will be discarded after each elution or strip cycle. Sodium hydroxide and sodium cyanide will be added as required from the reagent handling systems to the barren eluant tank during fresh strip solution make-up.

The precious metal-laden cathodes in the electrolytic cells will be removed about once per week and processed to produce the final doré product. Loaded cathodes will be transferred to a cathode wash box where precipitated precious metals will be removed from the cathodes with a pressure washer. The resulting sludge will be pumped to a plate-and-frame filter press to remove water and the filter cake will be loaded into pans for retorting. After mercury has been removed in the retort, the dried gold and silver sludge will be mixed with fluxes and smelted in a propane-fired furnace to produce doré bullion.


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17.9.5 Carbon Handling & Thermal Regeneration

The carbon handling and regeneration circuit will include all equipment required to store, prepare, transfer, and regenerate carbon.

The carbon preparation and storage system will include a 0.5 tonne agitated carbon attritioning tank, a 5-tonne carbon storage tank, carbon dewatering screen, carbon fines storage tank, carbon fines filter press, and carbon transfer pumps. New and acid washed/regenerated carbon will be stored in the carbon storage tank to be returned to the CIC circuit as makeup carbon. Carbon being transferred to the carbon storage tank will pass to a carbon fines/dewatering screen in order to remove any carbon fines from the system. Carbon fines will be stored in a carbon fines storage tank, which will be periodically pumped through the carbon fines filter press; carbon fines from the filter press will be stored in bulk bags for removal from the system.

New carbon being added to the system will first be attritioned in the carbon attritioning tank before being pumped to the carbon dewatering screen to remove carbon fines and is then transferred to the carbon storage tank.

Thermal regeneration will consist of drying the carbon thoroughly and heating it to approximately 700ºC for ten minutes in order to maintain carbon activity levels. The carbon regeneration circuit has been designed to regenerate 100% of the carbon.

Carbon from the elution circuit to be thermally reactivated will be dewatered on a static screen, transferred to the regeneration kiln feed hopper and fed to the regeneration kiln by a screw feeder. Hot, regenerated carbon leaving the kiln will pass into a water-filled quench tank for cooling before being transferred to the carbon dewatering screen and carbon storage tank. Ultimately, quenched regenerated carbon will be pumped to the CIC tanks to be loaded with precious metals.

17.9.6 Refining & Smelting

Cathode sludge from the filter press will be dried and treated in a mercury retort to remove and recover any mercury that may be present. The sludge will be placed into pans and heated in the retort for a minimum of 6 hours at 480ºC to volatilize the mercury. A vacuum system will remove mercury vapor from the retort and pass the vapor through a water-cooled mercury condenser. Condensed mercury will be collected in a trap, and then transferred and stored in flasks. Cooled, mercury-depleted vapor leaving the trap will be passed through a sulfur-impregnated carbon scrubber to remove any residual mercury.

After mercury removal, fluxes will be mixed with the cathode sludge and then fed to a propane-fired tilting crucible furnace. After melting, slag will be poured off into cast iron molds until the remaining molten furnace charge will be mostly molten metal (doré). Doré will poured off into bar molds, cooled, cleaned, and stored in a vault pending shipment to a third-party refiner. The doré poured from the furnace will represent the final product of the processing circuit.

Periodically, slag produced from the smelting operation will be re-smelted on a batch basis to recover residual metal values.

17.9.6.1 Mercury Abatement System

In addition to the mercury retort, the ADR facility will be fitted with an exhaust gas handling system to treat mercury emissions from the various pieces of equipment. The exhaust system will be designed to combine mercury-containing exhaust streams and treat them in two separate sulfur-impregnated carbon beds prior to discharge to the atmosphere.

The first carbon bed will be dedicated to treat fumes from the smelting furnace. The smelting furnace will be fitted with a hood which will collect fumes and direct them to a scrubber, which will remove suspended particles from the gas and


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cool the gas before passing through the carbon bed. The carbon bed will collect traces of mercury vapor before exhausting the gas to atmosphere.

The second carbon bed will treat the combined exhaust gas streams from the electrowinning cells, eluant solution storage tank, elution vessel, and carbon regeneration kiln. The kiln exhaust gas will be first treated through a wet scrubber to remove particulates and cool the gas, which will then be combined with the remaining exhaust gas streams and pass through the carbon bed.

17.10 ADR REAGENTS AND UTILITIES

Recovery plant reagents will include cyanide, caustic, hydrochloric acid, antiscalant, activated carbon, and various furnace fluxes. Propane will be used to fuel thermal equipment in the plant.

17.11 LABORATORY FACILITIES

Analytical support, including fire assays and metallurgical testing required to support the project operations, will be conducted on-site using a dedicated laboratory. It is assumed that approximately 100 samples per day will be delivered from the mine for fire assay. A small number of fire assays, solutions, and carbon assays will be required for metallurgical control for processing. A metallurgical lab area is also included for running bottle roll and column tests.


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18 PROJECT INFRASTRUCTURE

The infrastructure for South Railroad has been developed to support mining and heap leaching operations. This includes the access road to the facility, power supply, communication, heap leach pad, process plant and ancillary buildings. Water supply to the site including tanks, pipelines, ponds, and diversions are described in Section 18.5. Haul roads within the mining area as well as the mine waste storage facility are described in Section 16.

18.1 ACCESS ROAD

The infrastructure envisioned for South Railroad is shown in Figure 18-1.

Entrance to the site will be located approximately 20 miles south of Carlin, NV along Nevada State Highway 278 (Hwy-278). The main access road to the site will be along an existing 10-mile route east of Hwy-278, which is to be improved to a standard two-way road consisting of a 4-meter wide lane and 2-meter wide shoulder in each direction. The shoulders will provide area for any safety and drainage structures that will be needed along the route. New north-south turning lanes will be added to the highway to allow for safe access onto the new access road. Delivery of all personnel, operating equipment, consumables, and construction equipment will be along this primary access road.

The access road climbs gradually as it heads east and adjacent to an existing wash for approximately 5 miles. The remaining 5 miles to the site encounter mountainous grades and winding alignment of the existing dirt road. This road will be improved to straighten the alignment, where possible, and reduce grades to a maximum of 8-10 percent to allow for easier access to the site and promote safety. As the access road approaches the site all traffic will be required to check in at the security office located in the saddle before heading past Administration and down to the site facilities located between the Pinion and Dark Star pits.

18.2 POWER SUPPLY

Electrical power will be supplied from Wells Rural Electric Company (WREC) in Carlin, NV down their existing line along Hwy-278 to the main access road intersection. This 24-mile WREC leg will include a 397.5 3-Phase conductor upgrade along with a new substation transformer. Power to the site will be provided from the Hwy-278 tap location by NVE and transmitted to the project via a power line constructed along the main 10-mile access road to the onsite substation located in Figure 18-1.

Total connected power requirements are estimated to be 13.1 MW and peak load power requirements are estimated to be 8.8 MW.

18.3 PROJECT BUILDINGS

The proposed heap-leach facility will be located just Northeast of the Pinion pit on the west side of the valley. The crusher will be located just south of the Heap Leach pad. From the HPGR facility the ore will be conveyed to the leach pad from an overland conveyor to a series of grasshopper conveyors that will distribute the ore onto the pad in the prescribed courses. The grasshopper conveyors are not demonstrated on the drawings at this time for clarity as they will be moved throughout the construction of the Heap Leach. Pregnant Leach Solution (PLS) will flow by gravity to the PLS Pond directly east of the Heap Leach Pad. An event pond will be located adjacent to the PLS Pond to allow for passive overflow if an excessive runoff event occurs. Road access is provided just along the west edge of the heap leach facility which will allow access onto the leach pad for ROM material. An access point is also provided at the base of the pad to allow for haul truck ingress for the initial ore placement on the pad.

A truck shop is planned northwest of the Dark Star waste dump. A fuel island will be constructed just west of the truck shop. Safety and training areas will be provided within the shop building. In addition, Mine Services offices are integral


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to the truck shop and a laydown yard is proposed directly east of the facility. The Pinion and Dark Star pits are tied to their respective waste dumps and the primary crusher by haul roads.


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Figure 18-1 Site Plan Drawing


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18.3.1 Crushing Buildings

The process infrastructure includes three buildings for the crushing circuit; Secondary Screen Building, Secondary Crusher Building, and HPGR Building. The crushing buildings will be uninsulated, engineered steel buildings and have been designed to protect the crushing and screening components from harsh weather elements. Each of the crushing buildings will have partial walls on all four sides of the building to facilitate access by maintenance and support equipment.

18.3.2 Security Building at Access Gate

The site Security Building is located at the top of a hill for optimal visibility, approximately 6 kilometers along the main access road from the west property line. The Security Building includes an entry access gate that will control all site ingress egress. From the entry gate a continuous security fence surrounds the active facilities on site.

18.3.3 Administration Building

The site Administration Building is just past the Security Building also on the main road. The building will be comprised of (12) 12’ x 60’ mobile units that will be assembled into a single unit divided for the variety of use. Ten of these units will be used for the Administration Building, while the remaining two will be used to house the Change House Facilities.

18.3.4 Truck Shop Building

As the road continues from the Administration Building to the northeast the Truck shop is located just past the Primary Mine Substation and Fueling Station. The Truck shop is a 260’ x 100’ facility that has 6 bays with 2 of them embedded rail to receive tracked vehicles or loaders with tire chains. The Mine Warehouse Facility is included within the footprint of the Truck Shop at the ground floor at the opposite of the bay side. The Mine Services Office and Training Space is designed to be included above the warehouse space.

18.3.5 ADR Plant

The ADR Plant is located directly to the north and west of the Truck Shop. PLS from the Heap Leach Pad will be processed in an ADR (adsorption, desorption and recovery) plant where gold and silver will be adsorbed onto activated carbon and recovered by stripping the carbon and eventually recovering the precipitate by electrowinning. The ADR facility includes an open CIC circuit consisting of two carbon column trains operated in parallel as well as an 807 m²insulated, engineered steel walled building with an overall height of 14 meters. The building will contain the desorption, acid wash, and carbon handling and regeneration circuits, as well an office, break/lunch room, and men’s and women’s locker/bathroom facilities. The ADR facility also includes an attached refinery building which will be a 485 m²insulated, engineered steel walled building with an overall height of 7.7 meters and will contain the electrowinning, mercury recovery, and smelting furnace. The ADR building includes two roll-up doors for forklift and maintenance vehicle access as well as man doors around building. The Refinery includes a secure man-door access as well as access for armored trucks via a roll-up door. The facility will include all necessary eyewash/safety shower water and fire protection systems.

18.3.6 Laboratory

The Laboratory building will be comprised of a series of mobile buildings that will be assembled into a single unit to allow for a more conventional layout. The layout will include (6) 12’ x 72’ buildings (60’ x72’ building footprint) and accommodate proper scrubbers, acid containment system, dust collection, and necessary sample processing equipment. Offices, restrooms, and change facilities for the Lab are incorporated into the layout.


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18.4 SITEWIDE WATER MANAGEMENT STRATEGY

This section presents the overall strategy for managing the water produced from the mine as well as meet the demands of mine processes and supporting facilities. A process flow diagram illustrating how water will be managed at the site is presented in Figure 18-2 and locations of water management infrastructure, excluding stormwater controls, is shown in Figure 18-3. Further details, as well as the supporting studies and model results used to develop the strategy and cost estimate presented herein can be found in the Mine Water Management Plan South Railroad Project (in progress; Stantec, 2019).

18.4.1 Source of Mine Water

18.4.1.1 Dark Star Groundwater Dewatering System

The main source of water generated from the mine will be from the groundwater dewatering systems required to support the mining operation of the Dark Star North Pit (phase I). Pit dewatering wells located around the Dark Star North Pit will be connected to a 12-inch pipeline through a common header that will convey water to the 350,000-gallon Rapid Infiltration Basins (RIBs) Raw Water Tank (Tank 1). Water will be pumped from Tank 1 to either to the RIBs via a 12-inch pipeline or to the main 250,000-gallon Mine Raw Water Tank (Tank 2).

Based on current modeling, this system will consist of ten wells, each pumping between 55 and 275 gallons per minute (gpm) and will produce a total peak and sustained flow rate of approximately 1,430 gpm and 550 gpm, respectively. Refer to Table 18-1 below for pumping rates by year. Note that the required pumping rate for years 1 – 3 is determined by the pit dewatering schedule. The pumping rates for years -2 and -1 and years 4 – 9 are provided to meet mine processes needs, rather than the dewatering schedule. Thus, pumping rates during these years will be variable and based on demand, and therefore, may be less on average than the listed 550 gpm.

The dewatering system will continue to operate for 6 years following cessation of dewatering of the Dark Star North Pit (phase I) to supply water to support the mine. Water generated from the groundwater dewatering system will be beneficially used in the mining, leaching and processing, and to support mine facilities as discussed below. Based on the groundwater modeling conducted and water demands that have been identified, the mine should have enough water to meet all water demands throughout the life cycle of the mine.

A portion of water collected in Tank 1 will be transferred to Tank 2 via an 8-inch HDPE pipeline at a rate maximum rate of 700 gpm. Water from Tank 2 will be:

Transferred to a 428,000-gallon Raw Water Tank (Tank 3) to supply make-up and fire water for the ADR, HLP, and crushing areas;
Diverted for non-potable mine water uses; and
Routed to the potable water treatment and distribution system.

All tanks will be fitted with a level sensor that will control the flow to the tanks.


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Table 18-1: Current Modeled Pumping Rates for Dark Star Pit Dewatering System


		*Year -2*	*Year -1*	*Year 1*	*Year 2*	*Year 3*	*Year 4*	*Year 5*	*Year 6*	*Year 7*	*Year 8*	*Year 9*
*Well ID*	*Depth (ft)*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*	*Pumping*
		*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*	*(gpm)*
Well 1	1200	220	275	220	220	220	275	275	275	275	275	275
Well 2	1000	110	0	110	110	110	0	0	0	0
Well 3	1200	0	275	220	220	220	275	275	275	275	275	275
Well 4	1200	0	0	220	220	220	0	0	0	0
Well 5	1200	0	0	220	220	220	0	0	0	0
Well 6	1000	0	0	110	110	110	0	0	0	0
Well 7	1200	0	0	0	110	110	0	0	0	0
Well 8	1000	0	0	0	55	55	0	0	0	0
Well 9	1000	0	0	0	55	55	0	0	0	0
Well 10 (mine bottom)	200	0	0	0	0	110	0	0	0	0
Total Pumping		330	550	1,100	1,320	1,430	550	550	550	550	550	550


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Figure 18-2: Water Management Process Flow Diagram


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Figure 18-3: Pipeline Plan General Arrangement


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18.4.1.2 Stormwater Conveyance Facilities

Stormwater from the site will be managed as contact and non-contact stormwater. Non-contact stormwater are the flows that do not come in contact with ore or mine processing facilities. Non-contact flows will be collected and conveyed around the site and directly discharged to existing stream channels. Contact stormwater will be routed to the WRDF seepage ponds, the HLP Event Pond, and ponds near the Crusher Pad and Ore Stack Pad. Excluding the HLP Event Pond, contact water will be pumped and blended with other water sources in Tank 2. The operation of the WRDF collection ponds and the Crusher and Stacker Ponds are discussed in the following section. The HLP operations are discussed separately by others.

The collection and conveyance of stormwater runoff will be managed by the construction of stormwater channels, culverts, and energy dissipation structures. A total of 15 stormwater channels and 27 culverts will be used to convey stormwater around the site (non-contact water) or to lined storage ponds (contact water). The stormwater conveyance systems and collection ponds are shown on Figure 18-4.


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Figure 18-4: Stormwater Controls General Arrangement


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18.4.1.3 Seepage and Stormwater Collection Facilities

During operation, the WRDFs at Pinion and Dark Star will generate seepage water from precipitation migrating through the waste rock. Based on the water balance modeling conducted to date, maximum seepage rates of 115, 210, and 270 gpm are anticipated from the Pinion, Dark Star West, and Dark Star East WRDFs, respectively. The estimated seepage rates are influenced by the timing of the waste rock development and the anticipated concurrent reclamation of the facilities. The seepage from the WRDFs will be collected in lined collection ponds and pumped to Tank 2.

Due to the space limitations at the site, management of the WRDF seepage during operations by simple storage and evaporation alone is not practical. Therefore, the seepage collected from the ponds during operations will be blended with the groundwater in Tank 2.

In addition, potential runoff from the HLP, the Crushing Pad, and the Stacker Pad areas will also be collected as part the zero-discharge operating requirement. The 100-year, 24-hour stormwater volume reporting to the Crusher Pad Pond and the Stacker Pad Pond are 2.3 and 2.4 acre-feet, respectively. Stormwater from the Crushing Pad and Stacker Pad Areas are also routed to Tank 2 and recirculated in mine operations. The HLP water handling is discussed separately and is largely confined to the HLP and mineral processing areas in a self-contained system.

It is anticipated that stormwater and seepage from the WRDF ponds, the Crusher Pad Pond, and the Stacker Pad Pond (i.e., the ponds supplying Tank 2) will episodically supply all mine water needs with potential excess. During these events, all pit dewatering discharges will be sent to the RIB and any excess operating water will be sent to Tank 3. From Tank 3, it is assumed that excess water will be stored the HLP Event Pond and reused in mine operations. The excess water is generated due to pond size and pumping constraints to prevent overtopping and losing water to the environment.

These excess water periods would typically occur during winter or spring months, when operation water requirements are low. In early years of operation, these events would typically occur in 1- or 2-day periods, with average excess volumes being less than 1 acre-feet. This would be a small percentage of the total HLP event pond volume.

The greatest volume of excess water is currently expected in operating year 8. This would be when operational water requirements decrease, thereby increasing the likelihood of generating excess water from Pinion WRDF seepage or stormwater runoff into the Crusher and Stacker Ponds. The operations in year 8 would occur prior to full closure of the Pinion WRDFs and the Crusher and Stacker Ponds as described by the current mine schedule. Excess water generated in schedule year 8 would be expected throughout the winter and spring runoff periods, with projected excess water reaching approximately 2 acre-feet per day. The projected future excess water rate is dependent on a number of conditions during the period of operation including the actual weather conditions at the site, the closure and construction of WRDFs, and the Crusher and Stacker Ponds, and the timing of mine water needs at the HLP and other facilities.

18.4.2 Beneficial Reuse

The main water demands at the site are associated with:

The ADR Plant (KCA, 2019),
The HLP (Mines Group, 2019),
The Crushing and Screening Plant (KCA, 2019),
Mine facilities such as water for dust suppression, operational drilling water, and the truck wash (GSV, 2018), and
Potable water for the administrative building (GSV, 2018).

Water from Tank 1 will be transferred at a rate of 700 gpm to Tank 2 to provide enough water for the mine facilities. Water in Tank 2 will either be conveyed to the mine facilities area via gravity or pumped to Tank 3 for make-up and fire


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water for the ADR Plant, HLP, and Crushing Screening Plant. A description of the principal beneficial reuses for the site are presented below.

18.4.2.1 ADR Plant

The ADR Plant will require make-up water at a nominal rate of approximately 18 gpm (KCA, 2019). Water will be conveyed via gravity from Tank 3. Note that this water demand was incorporated into the overall HLP water demand estimate (Mines Group, 2019).

18.4.2.2 Heap Leach Pad

Based on water balance modeling for the HLP (Mines Group, 2019), water demands for the HLP will fluctuate significantly up to a maximum rate of 529 gpm, though a more typical make water demand is in the range of 100 to 300 gpm.

18.4.2.3 Crushing and Screening Plant

The crushing and screening plant will require water for dust suppression at a nominal flow rate of 20.4 gpm (KCA, 2019). Water will be conveyed from Tank 2 via gravity line to either a separate tank or directly into the water distribution plant for the area. Note that this water demand was incorporated into the overall HLP water demand estimate (Mines Group, 2019).

18.4.2.4 Mine Facilities

The mine facilities non-potable water demands (GSV, 2018) will consist of the following:

Dust Suppression – average and peak of 208 and 417 gpm, respectively;
Drilling and Construction – average and peak of 81 and 292 gpm, respectively; and
Vehicle Washdown – 10 gpm.

A series of distribution piping from Tank 2 will supply water to the mine facilities. Tank 2 has been located to supply water via gravity for these uses.

18.4.2.5 Potable Water

Potable water demands have been estimated at 10 gpm (GSV, 2018). Potable water will be required at both the mine facilities area, administration building, and ADR plant. Given that the source of potable water in Tank 2 will be blended from the WRDF seepage water and the mine dewatering system, treatment for potable use is required to meet potable water quality standards.

Treatment will be achieved by installation and operation of a 10 gpm packaged reverse osmosis (RO) system. Treated water from the RO system will be stored in the Potable Water Tank (Tank 4). From Tank 4 potable water will be pumped to the mine facilities, administration, or ADR buildings where water will be stored in a 1,500-gallon storage tank located at each area.

18.4.3 Water Disposal

The methods for disposal of excess water are discussed below.


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18.4.3.1 Excess Mine Dewatering Disposal

Excess water generated from the dewatering system will be pumped from Tank 1 through a 12-inch pipeline into a series of RIBs for disposal. Based on the minimum permeability requirements for RIBs of 2 in/hr., it is estimated that a total surface area of 2 acres will be required to dispose of the estimated peak excess water flow rate of 1,430 gpm and incidental rainfall. To allow for maintenance of the RIBs, a minimum total area of 4 acres will be required to allow for maintenance of one 2-acre area while the other 2-acre area is available for operation.

18.4.3.2 Domestic Wastewater Disposal

Domestic wastewater will be disposed of in three septic systems: one located at the mine facilities area, another at the Administration Building and the third at the ADR plant. Based on the estimated domestic wastewater flow rate, each septic system will be 1,500 gallons.

18.5 WATER MANAGEMENT INFRASTRUCTURE

This section discusses the infrastructure required to manage mine water at the site.

18.5.1 Dark Star Groundwater Dewatering System

Infrastructure associated with the Dark Star dewatering system is described in the below section.

18.5.1.1 Wells

Groundwater modeling has indicated that ten wells installed to varying depths between 200 ft to 1,200 ft will be required to provide sufficient dewatering capacity. Threaded carbon steel piping will be used to connect the pump to the surface completion piping. Well locations are shown on Figure 18-5. Typical well construction details are shown on Figure 18-6 and Figure 18-7.


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Figure 18-5: Map Showing Locations of Dewatering Wells


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Figure 18-6: Typical Dewatering Well Head Plan General Arrangement


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Figure 18-7: Typical Dewatering Well Head Section


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18.5.1.2 Well Pumps

A total of 10 well pumps will be procured for the project. Based on the maximum flow rate, each well pump will be required to pump at a maximum rate between 55 to 275 gpm and will be installed to depths between 200 and 1,200 ft bgs.

18.5.1.3 Pipelines

Each well will be connected with a threaded, carbon-steel discharge pipe to a 12-inch header that will be connected to Tank 1 via 12-inch HDPE pipeline. The header network will be divided into two sections, one on the north side of the Dark Star North Pit and the other on the south side of the Dark Star North Pit. Each section will consist of 5 wells. Water from Tank 1 will be pumped to Tank 2 via an 8-inch pipeline for further use and the remaining water will be pumped to the RIBs via a 12-inch HDPE pipeline for disposal.

18.5.1.4 Tanks

Tank 1 will serve as a buffer tank. Water from Tank 1 will be pumped to Tank 2 for further use at the mine. The Tank 1 will be a carbon steel tank having capacity of 350,000 gallons.

18.5.1.5 Well Pumps

A submersible pump will be installed in each well. Due to the variations in pumping rates from each well, 55, 110, or 275 gpm pumps will be used in an effort to standardize.

18.5.1.6 Distribution Pump

There will be two sets of distribution pumps installed at Tank 1 that will be used to transfer water to either the RIBs or to Tank 2. The pumps have been sized to provide adequate pumping capacity to meet the expected peak flow rate to the RIBs or Tank 2.

18.5.1.7 Instrumentation and Controls

Each well will have a level sensor installed to control the pumps that will operate the pump between high and low level to maintain the groundwater level below the bottom of the pit.

Tank 1 will be installed with a level sensor that will control the flow and operate the pumps. The pumps will maintain designated operating levels in the tank by adjusting the flow rate to the RIB’s with a variable frequency drive (VFD) motor on the pump. The distribution pumps transferring water to the Tank 2 will be turned off at low water level in Tank 1.

18.5.1.8 Electrical

Electrical supply will be required to each of the well locations and to Tank 1 to power the actuating valves and two pumps. The total installed power will be 1285 HP including spares, whereas the energy demand for all conveyance systems will be 828 KW.

18.5.2 Seepage and Stormwater Management System

Infrastructure associated with the seepage and stormwater management system is described in the below section.


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18.5.2.1 Ponds

Five ponds will be used to manage contact stormwater and seepage from the WRDFs and crushing and stacking area during operations. Pumping systems will be installed in each pond to pump water when the pond levels reach a predetermined level. During operations, the water pumped from the ponds will be discharged to the Tank 2. It should be noted that during certain times, water pumped from the ponds will make up 100% of the mine facility make-up water in Tank 2. As such, an assessment of the predicted water quality from the ponds will need to be performed to determine if any additional treatment is required.

18.5.2.2 Pipelines

The following pipeline will be used to transfer water from the ponds to Tank 2:

A 6-inch HDPE pipeline from Pinion WRDF Pond to Tank 2;
An 8-inch HDPE pipeline from Dark Star West and East WRDF Ponds to a common 10-inch HDPE connected to Tank 2;
An 8-inch HDPE pipeline from Crusher Pond to Tank 2; and
An 8-inch HDPE pipeline from Stacker Pond to Tank 2.

18.5.2.3 Pumps

Each pumping system will include two submersible pumps. The expected nominal and maximum flow rate for each system is shown in Table 18-2. The wide range of expected flow rates will be covered with VFDs that will control the speeds of the pumps. The pumps were standardized to reduce the number of spares and parts required.

Table 18-2: Expected Pumping Rates for Contact Water Ponds


*Pond*	*Nominal (gpm)*	*Maximum* *(gpm)*
Dark Star East WRDF Pond	50	400
Dark Star West WRDF Pond	50	400
Pinion WRDF Pond	50	400
Crusher Pond	200	400
Stacker Pond	150	350

18.5.2.4 Instrumentation and Controls

The pond pumping system will be controlled using level sensors that will be used to turn on and off the pumps at preset high and low levels.

18.5.2.5 Electrical

Electrical power supply will be required to each of the pond locations to power the pressure transducers and two pumps. The total installed power will be 450 HP including spares, whereas the energy demand for all conveyance systems will be 216 KW.

18.5.3 Beneficial Reuse System

Infrastructure associated with the beneficial reuse system is described in the below section.


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18.5.3.1 Pipelines

The following water distribution pipelines will be required to convey water for mining process and facilities:

An 8-inch pipeline to convey water from Tank 1 to Tank 2;
A 10-inch pipeline from Tank 2 to Tank 3;
A 6-inch pipeline to convey water from Tank 2 to the mine facilities; and
Ancillary smaller diameter distribution pipelines for potable water and the various mine facility uses.

18.5.3.2 Tanks

Three tanks, Tank 2, Tank 3, and Tank 4 will be installed to manage water discharged from the Tank 1 and flows from the ponds. Tank 2 will be the main mine facilities water tank and will serve as the primary storage for water for the mine. Tank 2 will be a 250,000-gallon carbon steel tank. Tank 3 will be the main make-up and fire water storage tank for the ADR, crushing and screening, and HLP areas. Tank 3 will be a 428,000-gallon carbon steel tank. Tank 4 will store treated potable water and will be a 15,000-gallon carbon steel tank.

18.5.3.3 Pumps

Two pumps will be located downstream of Tank 2 to pump water to Tank 3 at a maximum rate of 1200 gpm.

A third pump will be located downstream of Tank 4 to pump potable water to the other smaller potable water holding tanks located at the mine facilities area, administration building, and ADR plant.

18.5.3.4 Instrumentation and Controls

All tanks will include low-level, high-level, and high-high level sensors. These sensors will be used to control pumps and valves downstream of the various tanks feeding each system.

18.5.3.5 Treatment

A 10 gpm prefabricated RO system will be installed near the mine facilities area to treat water from Tank 2 at a rate of 10 gpm to supply potable water to the mine facilities, administration, and ADR Plant.

18.5.3.6 Electrical

Electrical power supply will be required at each of the distribution system to power the pumps. The total installed power will be 620 HP including spares, whereas the energy demand for all conveyance systems will be 238 KW.

18.6 HEAP LEACH PAD FACILITY

The heap leach facility consists of a conventional lined leach pad to support a multi-lift, free-draining heap, event and pregnant ponds, access roads, solution distribution piping (barren solution to the heap) and heap drainage solution collection piping (pregnant solution to the ponds and plant).

For the first year ROM ore will be stacked with trucks in nominal 9 m thick lifts. Barren solution containing cyanide will be irrigated onto the ore using drip irrigation. Pregnant solution will be collected at the base of the heap by the leach pad liner and collection system, which will route the pregnant solution to the process plant for gold recovery and reagent reconditioning. At the end of year 1, higher grade ore will be crushed and agglomerated and stacked on a dedicated section of the pad closest to the crusher using conveyor stacking system in 7 m thick lifts. The ROM and Crushed ore (which includes HPGR ores have different leach cycles and times, hence the need to segregate them on the leach pad. Once an area has been leached for the target time or metal recovery, the next lift will be placed on top of the already


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leached ore and the process repeated. This will be continued until the heap is stacked to the design elevation of 2100 m for the 47.3 Mt capacity.

The leach pad will consist of a graded area to the west of the ADR process plant and northwest of the Dark Star open pit. The leach pad will be constructed in phases, with each phase large enough to provide ore leaching capacity for 2 to 3 years. For each phase, topsoil will be removed and stockpiled for use in reclamation.

After removal of topsoil, the site will be graded by cutting and filling to achieve targeted slopes, elevations and grades. Cell separation berms will be constructed as part of the subgrade preparation. The resulting subgrade will then be prepared to make it suitable for supporting the liner system. Prior to laying of the synthetic liner system, a leak detection collection system consisting of a wick drain will be laid on the composite clay liner directly below the primary solution collection pipes for each cell and will report to the perimeter solution collection trench. The pad area will then be lined with a layer of primary geosynthetic liner (LLDPE) to form a composite liner system, consisting of a 2.0 mm-thick single-side textured LLDPE geomembrane overlying and in direct contact with the 300 mm-thick compacted clay liner (CCL). A network of drainage pipes and drainage gravel will be placed on top of the primary LLDPE to protect the liner and piping from damage, to limit the maximum hydraulic head over the liner system to an average of 0.3 m, and to collect the pregnant solution and direct it to ADR facility for processing.

The event and pregnant ponds will be located near and adjacent to the ADR process plant. A total of two ponds are planned for the HLF. The initial event pond will be single lined with the primary liner being a 1.5 mm thick smooth HDPE (or other suitable material). The pregnant pond will be double lined with both the primary and the secondary liner being a 1.5 mm HDPE. Drainage between the two liners for the pregnant pond can be provided by installing a geonet between the liners or one of the liners can be a drain liner (manufactured liner with separation knobs or buttons) to separate the two liners providing a drainage path to the leak detection sump. These two geomembranes will be underlain by a composite clay liner (CCL) to form a composite bottom system liner. A leak detection sump will be installed in the low corner of the pregnant pond. Operational solution will be routed via in tanks located at the process plant. There will be two tanks for pregnant solution, and one for the barren solution. The second pregnant tank is for maintenance which can also be used for maintenance of the barren tank. The solution tank sizes are included in process plant design report. The pregnant pond is designed to have the storage capacity for 24 hours of drain-down from the leach pad in the event of any issues with processing of operational solutions .The event pond will be sized for storage of the runoff from the 100-year, 24-hour storm event as well as the larger of the associated storm surge or the pond inflow from the wettest month timestep as determined from the deterministic water balance model for the leach pad (which would include snowmelt runoff). The ponds will have a dedicated generator and pump back system for moving solutions as needed during a “power outage.”

18.7 HEAP LEACH FACILITY WATER BALANCE ANALYSIS

Heap leaching involves the dissolving of precious metals contained in a low-grade ore using the application and circulation of a weak cyanide solution through the ore. An operational water balance model has been developed for the proposed HLF at the project site. The model provides output to evaluate meteoric (weather) impacts on the facility design and to predict the freshwater demand during operations and subsequent post mining freshwater circulation. The water balance model for a heap leach pad operation is essentially a water budget that tracks all of the water entering and leaving the lined containment system. Sources of water entering the system include pore water delivered with the ore, precipitation falling as rain or snow, and any fresh water (makeup water) added to the system from outside the lined limits of the pad. System losses are a bit more complicated and include three basic categories of loss.

Evaporative losses
Losses due to surface tension
Extraction losses


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In the case of an operating heap leach pad, the area under active leach is assumed to be continuously wetted by sprinklers or emitters with a limitless supply of water. Therefore, the full potential depth of evapotranspiration is applied to that area. Outside of the area under active leach, the ore surface is assumed to be dry, except for that fraction of the month’s rainfall events that coated the soil particles or infiltrated into the soil and did not run off. This volume of water is assumed to be available during that month for evapotranspiration. Any portion of the infiltrated water volume that is not lost to evapotranspiration during the same month it falls is assumed to be beyond the reach of evapotranspiration in the following month and is routed into the solution collection system along with the other applied solution. Therefore, during months where evaporation/evapotranspiration greatly exceeds rainfall, rain events add nothing to the water volume stored in the system. However, during months where rainfall greatly exceeds evaporation/ evapotranspiration, a significant volume of water may be added to storage.

Environments like the S RR Project site where snowfall is a substantial part of the precipitation regime create a special case. During much of the year, a snowpack will exist on the surface of the HLF which will significantly hinder evaporative loss but create a new opportunity for “sublimation” loss (which is a phase change where water goes directly from the solid phase to the gas phase without passing through a liquid state).

Losses to surface tension involve changes in the water content of the ore during operations. The ore is not delivered to the heap leach pad in a truly dry condition, but rather contains some relatively small amount of moisture in the pore spaces that is held in place by surface tension. This delivered water content is typically less than the “specific retention” of the ore. The specific retention is a threshold moisture content that marks the position on the soil water characteristic curve where the soil begins refusing to release its water to gravity (i.e., below that moisture content it simply will not readily drain). Therefore, for ore to release the applied solution carrying the dissolved precious metals to the solution collection system, it is necessary to raise the moisture content of the soil to a level above the specific retention. The moisture content of the ore must be increased to a level that allows the water to be passed through the ore at the same rate that it is being applied so that the system is in equilibrium or in balance. Once an area is no longer actively being leached (i.e., no new solution is being applied), then the ore would drain back down to its specific retention moisture content and release the difference back into the solution collection system. The water balance model tracks these changes in moisture content in the ore and accounts for the addition and subtraction of water volume in the system. Once all additions and losses to the volume of water stored in the system have been estimated and accounted for at the end of the month, the model evaluates whether or not there is sufficient water available in storage to maintain the solution application rate for the next month. Heap leach pads are designed as fully lined containment systems that in theory release nothing back into the environment. Solutions that are not stored within the ore itself are routed through the system and stored in various lined ponds. However, should extreme events exceed the storage capacity of the system, then the excess must be extracted from the system.

Precipitation was studied by Stantec and utilized multiple regional sources of data including the site-specific Dark Star climate station. The site-specific data has a record length of only about two (2) years. Available regional meteoric records included data sets as long as 130 years. Details on the development of a representative meteoric record for the project site can be found in a report from Stantec dated April 19, 2019.

Given the location of the site in mountainous terrain at elevations well above 6000 ft above mean sea level (amsl) and the existence of sub-freezing temperatures from late October through April each year, a significant percentage of the precipitation at site occurs as snow. The accumulation of water as the snow water equivalent (SWE) in a growing snowpack over the winter months has an impact on the hydrology of the site by storing water from November through March or early April, then rapidly releasing that stored water over the months of April and May. The water balance model controls the accumulation of SWE in the snowpack as a function of precipitation and temperature using a monthly series of snowpack factors. The monthly snowpack factors were selected to mimic as closely as possible the behavior observed at Snotel sites in the region (the snowpack growing rapidly from November through February, leveling out from March through early April, and declining rapidly from April through May. The snowpack algorithms affect the routing and the timing of the winter precipitation and spring melt, but they have no impact on the net water balance.


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Results of the deterministic modeling are as follows. In general, outside makeup water is required from startup through the end of the facility life which is anticipated to be on the order of 9 years (7 years of ore stacking and leaching and up to two years of additional leaching and gold production after ore stacking operations end). Modeling disclosed no significant trend toward accumulation of water in the system over time during normal operations. However, water does begin to accumulate in the pond system after ore stacking ends due to the elimination of the ore wetting component of total system loss. The model assumes that concurrent reclamation consisting of the placement of a low permeability cover will begin by Operating Year 5 as areas become available on the pad where no further irrigation will occur and that cover placement will continue from May through October each year at an average rate of 20,000 m²/month. Reclamation sheds clean runoff to the environment and helps reduce the risk of the accumulation of water in the system as the lined footprint grows over time.

The water balance model covers the period of leach pad operation that includes startup and Phase 1 through Phase 5 which includes up to two (2) years of post-mining leaching. Operations effectively end with the termination of gold production late in Year 9. Upon completion of active leaching operations, solution management will be required until such time as the closure cover is established and clean runoff is diverted off the facility. Once the solution draindown rate falls to a level that can be safely and passively contained in the post-closure Event Pond(s), active solution management can cease (i.e., no pumping). The current water balance model does not address these post-closure conditions (which will need to be addressed in a separate draindown model at some later time).

Detailed phasing and scheduling of the liner deployment is beyond the scope of this pre-feasibility study. However, the following assumptions were made to approximate a reasonable schedule of liner deployment over time. Phase 1 is assumed to begin with an initial lined footprint of 376,238 m²accommodating approximately 10 million tonnes (MT) of ore. Three (3) additional phases (Phase 2 through 4) are approximately equally spaced in time reaching a final, maximum lined footprint of 922,511 m²during Phase 4 (see Table 18-3). Phase 5 consists of post-operations gold production with no additional liner deployment.

Table 18-3: Summary of Phased Liner Deployment


Phase	Lined Surface Area (m²)
1	376,238
2	558,329
3	740,420
4	922,511
5	922,511

Table 18-4 summarizes results from the deterministic modeling using the typical/average range cycle of the meteoric record. Mean values reported in the table are averaged over each stage of operations. Simple volumes reported as m³are the totals over the duration of a one (1) month timestep. It should be noted that modeling of water storage in system ponds assumes no specific day to day management/mitigation. Facilities are commonly operated with ponds in an empty or near empty state and maintained that way through modest adjustments to pumping rate and/or area under leach to move water in and out of dynamic storage. Reported pond levels do not reflect that level of day to day management.


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Table 18-4: Results Summary from the Deterministic Model – Typical/Average Range Cycle


Parameter	Phase	Max	Mean	Min
Water Stored in Event Ponds (m3)	1	35,371	3531	0
	2	29,608	2428	0
	3	43,189	3805	0
	4	41,719	9753	0
	5	51,738	9710	0

Runoff from Reclaimed Areas (m3/month)	1	0	0	0
	2	0	0	0
	3	26,134	3704	0
	4	30,332	9941	0
	5	42,079	10,483	0

Outside Makeup Water (m3/month)	1	54,671	25,477	0
	2	55,439	28,778	0
	3	55,130	22,934	0
	4	52,257	26,590	0
	5	41,524	12,208	0

Outside Makeup Water (liters/s)	1	20.80	9.69	0
	2	21.09	10.95	0
	3	20.98	8.73	0
	4	19.88	10.12	0
	5	15.80	4.64	0

Outside Makeup Water (liters/tonne of ore)	1	98.8	46.2	0
	2	81.0	42.0	0
	3	80.5	33.5	0
	4	76.3	38.8	0
	5	0	0	0

% of Time Makeup Water Demand is Zero	1	---	13.0%	---
	2	---	8.3%	---
	3	---	16.7%	---
	4	---	28.6%	---
	5	---	40.5%	---

Preliminary pond sizing is based on a hydrologic analysis and the results of the deterministic water balance modeling.

The HLF design team for the GSV S RR Project has adopted the following minimum pond sizing design criteria:

The immediate runoff from the 100-yr 24-hr storm event over the area of the ore stack and any additional exposed liner (with or without liner cover) over the full lined footprint of the HLF.
The average expected volume in storage in the pond system plus the larger volume of the storm surge (the infiltration portion from the 100-yr 24-hr design storm) or the maximum expected increase in pond volume for any single time step (i.e., 1 month) based on the deterministic HLF water balance model.
24 hrs. of draindown at the full pumping rate (backup power with backup pumps are to be provided).
2 ft of pond freeboard.


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Given that the lined footprint of the HLF will change over time, the volume requirements of the design criteria will also change over time. Table 18-5 will summarize the development of the required emergency storage volume for pond design by phase (although the constructed pond volume does not necessarily have to be phased).

Table 18-5: Summary of Total Required Emergency Storage Volume for Pond Sizing by Phase


Phase	100-Yr 24-Hr Runoff (m3)	Larger of Storm Surge or Snowmelt (m3)	Mean Volume in Pond Storage (m3)	24 Hrs of Draindown (m3)	2 ft of Freeboard (m3)*	Total Required Emergency Storage Volume (m3)
1	20,539	20,061	5254	24,000	12,000	81,854
2	25,000	24,129	5254	24,000	13,000	91,383
3	33,154	29,653	5254	24,000	13,500	105,561
4	41,307	36,946	5254	24,000	14,000	121,507
5	41,307	38,718	5254	24,000	14,500	123,779

*- Approximate estimate in the absence of an actual pond design

18.8 SEISMIC HAZARD ANALYSIS

The site resides in the Basin and Range physiographic province which consists of a region of crustal extension (spreading) that began approximately 17 million years ago during the Miocene Epoch. The province extends from southern Oregon and Idaho southeastward penetrating well into Mexico. Its westernmost extent is the range front fault(s) of the eastern Sierra Nevada Range and its easternmost extent the range front fault(s) of the Wasatch Range. The southern projection of the province is bounded on the west by the gulf of California and the Baja Peninsula and on the east by the Laramide aged thrust front of the Sierra Madre Occidental Range. The spreading and thinning of the crust in the Nevada portion of the province is dominated by listric normal faulting that bounds the mountain ranges and flattens out with depth, even joining opposing faults at times. This pattern has resulted in what is described as “horst and graben topography” where the horsts are the uplifted areas (mountain ranges) and the grabens are the down-dropped blocks (alluvial valley floors) between ranges.

The identification of representative seismic source zones for a project of this type requires a review of the patterns revealed in a plot of the mapped earthquake epicenter locations classed by magnitude, and a review of the patterns revealed in a plot of the mapped young, potentially active fault locations. We have identified eight (8) seismic source zones which (proceeding from southwest to northeast) are as follows:

1.	Sierra Range-front Zone
2.	Walker Lane Zone
3.	Shoshone Mountains Zone
4.	Southern Nevada Zone
5.	Nevada Great Basin Zone
6.	Idaho Mountains Zone
7.	Salt Lake Zone
8.	Wasatch Front – Hurricane Fault Zone

The purpose of identifying discrete seismic source zones is to characterize and quantify the nature of the largest earthquake that is likely to occur within the zone. This information can be utilized in either a deterministic seismic hazard analysis (DSHA) or a probabilistic seismic hazard analysis (PSHA). Although different in approach, they probably have more in common than they have differences. Of interest is the largest earthquake that could reasonably be expected to occur within the zone, the mean rate of occurrence or recurrence interval, and the location of the earthquake. Seismic source zones come in two (2) varieties:


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1.	An Aerial Seismic Source where earthquakes are uniformly distributed throughout the area and assumed to be equally likely to occur anywhere within the area.
2.	A Linear Seismic Source where earthquakes occur along a narrow linear band (fault) but are again assumed to be equally likely to occur anywhere along the fault line.

Figure 18-8: Plot of Historic Earthquake Events and Selected Seismic Source Zones within a 500 km Radius


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Ground motion response to earthquakes depends not just on the character of the earthquake, but also the character of the subsurface conditions at the site. Of concern is the nature of the soil/rock in the upper 30 m of soil/rock profile. Test pits and drilling at the site indicate that the soil cover is of moderate thickness (typically 6 m to 14 m thick) and the underlying rock moderately to highly weathered. Therefore, for the purpose of this investigation, the site was assigned to Site Class D (stiff soil) with an assumed representative shear velocity (Vs30) of 365 m/s (1200 ft/s) consistent with the recommendations in the ASCE 7-16 design standard.

The steps involved in a DSHA analysis are as follows:

1.	Using information derived from geologic maps, fault maps, and plots of historic earthquake events, identify discrete seismic source zone polygons.
2.	Extract “clipped” data sets lying within each seismic source zone that represent the nature of the seismicity within the zone.
3.	Estimate the Maximum Considered Earthquake (“MCE”) associated with each seismic source.
4.	Estimate the closest point of approach to the site of interest for the selected MCE in each seismic source zone.
5.	Estimate site specific ground motions by attenuating motions over the distance between the earthquake epicenter and the site.

A review of the mapped USGS faults revealed a maximum surface rupture length within the Nevada Great Basin Zone on the order of 29 km at a location approximately 42.5 km south of the site. Using the criteria of Wells and Coppersmith (1994), and assuming the maximum surface fault rupture for a single event to be half of the mapped length, the maximum expected event magnitude would be 6.4.The closest location of a mapped active fault is 5.3 km from the site and the fault has a total surface rupture length of 5.55 km. Conservatively assuming this fault rupture to represent a single event, the 5.55 km length corresponds an event magnitude of 5.9. Therefore, for the Nevada Great Basin Zone containing the site, three (3) ground motion attenuation profiles were developed; one for the MCE of magnitude 6.4 at a distance of 42.5 km, one for a magnitude 5.9 event at a distance of 5.3 km, and a magnitude 4.5 event at a distance of 1 km.

Results of analyses for all seismic source zones are summarized in Table 18-6 and Table 18-7. Most building codes (including ASCE 7-16 and IBC 2018) allow for either a site-specific deterministic design approach or a probabilistic approach. For the site specific DSHA procedure the estimated spectral acceleration values at the various natural periods are used to develop a mean spectral acceleration response spectrum and an 84^thpercentile response spectrum. These accelerations are then used to develop design response spectra for estimating seismic loads used for structural design (which will also vary as a function of occupancy and use) and for geotechnical analysis.

The PSHA analysis can be performed using the same seismic source zones and by replacing the maximum credible earthquake with the probability distribution of earthquake events, the site distance with the probability distribution of site distances and adding a random component to the attenuated spectral acceleration values, then using a Monte Carlo type sampling model to compile a new distribution of spectral accelerations associated with an exceedance probability. However, some developed countries, including the U.S. and Canada, have their own web-based PSHA programs that use regionally based maps of seismic source zones (similar, but not the same as those used in our DSHA analysis), and compute site distances by asking you to enter a specific geographic site location using latitude and longitude.

A deterministic approach to seismic hazard analysis or DSHA and a probabilistic approach or PSHA have produced similar design pseudo-acceleration response spectra with the DSHA results being the larger of the two (see Figure 18-9). Per ASCE 7-16 guidelines, the lesser of the two or the PSHA results for a maximum considered earthquake having a 2% probability of exceedance in 50 yrs. was selected as the Site-Specific Design Response Spectra (see Figure 18-10).


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Figure 18-9: Plot of PSHA Results and Comparison with DSHA Results

Geotechnical design procedures often involve estimates of the Peak Ground Acceleration (“PGAm”) or a reduced/scaled version of the ground acceleration referred to as the pseudostatic acceleration coefficient. The design PGAm value for the site is 0.294 g.

For seismic slope stability analyses in soil and rock, a hierarchy of analysis methods should be implemented with progression to the next method in the hierarchy required only in the event of failure to satisfy the requirements of the previous method. Recommended methods in the order of their application are as follows:

Pseudostatic stability analysis using a pseudostatic acceleration coefficient of 0.06 g (to be used only at sites with no liquefaction potential).
Seismic displacement analysis using the procedures of Newmark, 1965, Makdisi and Seed, 1978, or Bray and Travasarou, 2007 showing acceptably small displacements.
Full dynamic analysis of soil-structure interaction coupled with continuum modeling showing acceptably small displacements.


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Table 18-6: Mean Deterministic Pseudo-Acceleration Response Spectrum by Seismic Source Zone

Table 18-7: 84^thPercentile Deterministic Pseudo-Acceleration Response Spectrum by Seismic Source Zone


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Figure 18-10: PSHA Results and Design Response Spectra


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19 MARKET STUDIES AND CONTRACTS

Gold doré bullions will be the commercial product from the South Railroad operation. Gold doré is readily sold on the global market to commercial smelters and refineries, and it is reasonable to assume that doré from the South Railroad property will also be saleable.

19.1 METAL PRICING

In determining the appropriate and reasonable gold price to use for the economic analysis, under the current estimated project schedule, long term lead items are expected to be committed to in 2020 and construction is intended to commence in 2021, with potential commercial production in 2022. Reviewing the Bloomberg Consensus price for gold in the years 2020, 2021, and 2022, which was respectively $1,500, $1,500, and $1,600, the average of the gold price for these three years was calculated and amounted to $1,533. The trailing three average gold price is $1,276. Then averaging the trailing three-year average gold price and the three-year average Bloomberg consensus price noted above of $1,533, the average gold price is $1,404.50 per gold ounce. For purposes of this study, the base case gold price was rounded to $1,400/oz with sensitivity from $1,250-$1,550/oz also evaluated.”

Table 19-1: Bloomberg Consensus Pricing for Gold


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Table 19-2 Bloomberg Consensus Pricing for Silver


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20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

EM Strategies, Inc. (“EMS”), a permit acquisition strategy and government relations consulting firm, provided the following information on environmental considerations, permitting, and social and community impacts.

20.1 INTRODUCTION

As environmental consultants to Gold Standard, and at the request of Gold Standard, EMS has completed the following assessment of environmental studies, permitting, and social or community impacts for the proposed Gold Standard’s South Railroad Mine Project (“SRMP”), which is located within South Railroad portion of the Railroad-Pinion property. The SRMP has been defined for permitting purposes and is currently approximately 2,693.3 hectares in size. The SRMP is a hard rock precious-metal development project. Gold Standard is planning to submit a PoO (under 43 Code of Federal Regulations [CFR] 3809) and an RP Application (under Nevada Administrative Code [NAC] 519A) (Plan Application) to the BLM Tuscarora Field Office and the NDEP’s Bureau of Mining Regulation and Reclamation (“BMRR”) in the near future. In addition, a right-of-way (“ROW”) application will be submitted for a 46 km powerline that would be located on public and private lands.

The SRMP is located on public lands administered by the BLM and private lands controlled by Gold Standard in Sections 13 through 16, and 19 through 29, Township 30 North, Range 53 East (T30N, R53E), and Sections 24 through 28, T30N, R52E, Mount Diablo Base and Meridian. In general, the proposed mine operations will consist of two open pit mines and rock storage areas, and the processing of the ore will use a heap leaching method. Gold Standard plans the construction, operation, reclamation, and closing of this mining operation. Major components include:

Two areas of open pits (Pinion and Dark Star deposits);
Three rock storage areas;
Crushing and conveying system;
One heap leach processing facility;
Reagent area;
Exploration;
Laydown areas;
A water delivery and distribution system;
A power delivery and distribution system;
Excess water management system;
Storm water diversion ditches and storm water sediment basins; and
Haul and access roads.

Gold Standard proposes to mine approximately 47.3 million tonnes of heap-leach mineralized material and 147.3 million tonnes of waste rock (total of 194.6 million tonnes). The mineralized material and waste would be extracted from the open pits using conventional surface mining methods of drilling, blasting, loading, and hauling. Gold Standard would use hydraulic shovels or front-end loaders to load the blasted mineralized material and waste into the haul trucks. The haul trucks would transport the waste rock to the rock disposal area near the open pit and transport the mineralized material either directly to the heap leach pad as ROM ore, or to the crushing system where the mineralized material would be crushed to a nominal size between ¼ and ¾ inch. The crushed mineralized material would be conveyed to the heap leach pad. The heap leach would use a dilute NaCN solution to liberate the precious metals. A carbon absorption desorption process would be used to precipitate the precious metals. The precipitate would then be refined in a furnace to produce doré bars for shipment off site. The project facilities would disturb approximately 642.8 hectares. There is an existing PoO which covers planned mining facilities and authorizes up to 101.2 hectares of exploration surface disturbance within the SRMP. Exploration activities, estimated to disturb up to 20.2 additional hectares, would also occur within the SRMP and incorporate the existing Plan level disturbance. The current exploration Plan would continue to be used for exploration outside of the Plan Application boundary. The project would therefore include 663


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hectares of authorized and proposed surface disturbance associated with exploration and mining activities in the SRMP. The exploration activities would be based on work plans submitted to the BLM for review and concurrence that the activities are consistent with the Plan.

The review and approval process for the Plan Application by the BLM constitutes a federal action under the NEPA and BLM regulations. Thus, for the BLM to process the Plan Application the BLM is required to comply with the NEPA and prepare either an EA, or an EIS. Gold Standard anticipates that the BLM will require an EIS, due to the mine dewatering and potential pit lake.

The Department of Interior (“DOI”) issued an order (Secretarial Order 3355) on August 31, 2017. This order directed the bureaus under DOI to implement page and time limits on EISs. The order set a one-year time limit. Subsequently, DOI issued directives on April 27, 2018, and August 6, 2018, that provide additional guidance. The BLM in Nevada is implementing a new permitting and NEPA process to further address the requirements of the order and directives. The process commences with the submittal of a brief project description and map and then a meeting to discuss the scope of necessary baseline data collection. Following the completion of the baseline reports the BLM reviews and approves them. The Plan Application is then completed and submitted to the BLM for review and a determination by the BLM that it is complete. If the Plan Application is determined to be complete, the BLM will make a decision whether an EA or EIS will need to be prepared to comply with NEPA. Prior to initiating the NEPA document (EA or EIS), the NEPA contractor will prepare Resource Reports for each environmental resource, which will evaluate the potential effect of the project on each environmental resource. Each Resource Report is then reviewed and approved by the BLM. The NEPA contractor then uses the Resource Reports to complete the NEPA document.

The following sections provide additional detailed information on the principal permits necessary to develop each phase of the project and the NEPA process, as well as the status relative to each permit process.

20.2 ENVIRONMENTAL BASELINE STUDIES

Gold Standard has been conducting environmental baseline studies over the past several years as part of their ongoing permitting efforts and in preparation for the submittal of permit applications for conduct mining operations. The main portion for the Project Area has been surveyed for surface water resources, including WOTUS, biological resources, and cultural resources. The SRMO access road, the powerline route, and the water management area remain to be surveyed. In 2018, Gold Standard commenced material characterization testing of the mineralized material and waste rock to determine the metal leaching and acid generation potential. In addition, an evaluation of the groundwater resources was commenced to determine groundwater supply potential, as well as the potential impacts from groundwater pumping and pit lake development. Gold Standard had a meeting with the BLM in January 2019 to determine any additional baseline data collection needs for the permitting process.

Within and adjacent to the Project Area there are Greater Sage Grouse and Golden Eagles. These species will have an effect on how the SRMP is permitted and what mitigation in required or proposed.

20.3 BUREAU OF LAND MANAGEMENT PLAN OF OPERATIONS / NEVADA BUREAU OF MINING REGULATION AND RECLAMATION, NEVADA RECLAMATION PERMIT

The BLM and the BMRR have implemented a process for the Plan Application that commences prior to the submittal and continues through the review and approval process for the Plan Application. Gold Standard will submit a Plan Application for the project and BLM approval of this Plan Application will likely occur in the fourth quarter of 2020, assuming the BLM uses the EIS process.


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20.3.1 Bureau of Land Management Pre-Application Planning

As part of the pre-Plan Application planning process with the BLM, an initial meeting is scheduled between the proponent and the BLM to discuss the anticipated scope of the mining operation and review the likely environmental resource baseline data needs required for the processing of the Plan Application by the BLM. This initial meeting generally occurs some time prior to the submittal of the Plan Application, depending on the anticipated complexity of the mining operations and baseline data needs, which varies for each project. Several meetings between Gold Standard and the BLM Tuscarora Field Office have occurred over the last year.

The process for collecting baseline data generally includes the development of baseline data collection work plans, which are submitted to the BLM for review and approval prior to initiating the baseline data collection. Following approval, field surveys are carried out to collect relevant baseline data. Depending on the environmental resource to be evaluated, desktop studies may be utilized in lieu of field surveys. Findings of the field surveys are then summarized in a report that documents the data collected. This Technical Report is then submitted to the BLM for review and approval. In some cases, the baseline data collection process will also involve the State of Nevada, depending on the resource being assessed, particularly for geochemical and hydrological surveys. Baseline data for the project is being collected and the reports have yet to be submitted and reviewed or accepted by the BLM. The required environmental baseline data include the following: mineralized material and waste rock geochemical characterization; hydrogeological characterization; a pit lake evaluation; an assessment of ecological risk; air quality modeling; and cultural and biological resources.

Cultural resource and biology surveys have been completed over the SRMP and will be completed over the powerline and access road routes in 2019. The cultural and biology reports are currently under preparation and will be completed in the fourth quarter of 2018. Sample collection for the characterization of the mineralized material and waste has been completed and analysis of those samples is underway. The characterization report will be completed in the first half of 2019. The hydrogeologic evaluation commenced in the third quarter of 2018 and the report will be completed in the third quarter of 2019.

20.3.2 Plan of Operations Processing

The Plan Application is submitted to the BLM and the BMRR for any surface disturbance in excess of five acres. The single application utilizes the format of the Plan Application document accepted by the BLM and the BMRR. The Plan Application describes the operational procedures for the construction, operation, and closure of the project. As required by the BLM and BMRR, the Plan Application includes a waste rock management plan, quality assurance plan, a storm water plan, a spill prevention plan, reclamation plan, a monitoring plan, and an interim management plan. In addition, a reclamation report with a Reclamation Cost Estimate (“RCE”) for the closure of the project is required. The content of the Plan Application is based on the mine plan design and the data gathered as part of the environmental baseline studies. The Plan Application includes all mine and processing design information and mining methods. The BLM determines the completeness of the Plan Application and, when the completeness letter is submitted to the proponent, the NEPA process begins. The RCE is reviewed by both agencies and the bond is determined prior to the BLM issuing a decision record on the Plan Application and BMRR issuing the RP.

The Plan Application will be submitted for the project when operational and baseline surveys are complete and operations and design for the project are at a level where a Plan Application can be developed to the necessary level of detail. Submittal of the Plan Application is likely to occur in the third quarter of 2019. Key baseline reports for the project will be included in the Plan Application submittal to the BLM and NDEP/BMRR. These reports have yet to be reviewed by the agencies.

The BLM will need to complete their review of the baseline reports in the Plan Application and approve the final version of the reports prior to moving on to the NEPA process.


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20.4 UNITED STATES ARMY CORPS OF ENGINEERS SECTION 404 PERMIT

Gold Standard has delineated and the United States Army Corps of Engineers (“USACE”) has determined that there are WOTUS, including wetlands, within the Project Area. Based on the current design of the SRMP, the SRMP will likely have impacts to WOTUS, and at a level to will require an individual permit under Section 404 of the Clean Water Act. As part of their Section 404 permit application review process, the USACE looks at an avoid, minimize, mitigate process as part of their assessment. GSV is unable to avoid all the WOTUS in the SRMP design; however, Gold Standard has designed the SRMP to avoid as much of the WOTUS as is reasonably possible. Gold Standard will need to then mitigate for the WOTUS that is affected by the SRMP design.

20.5 NATIONAL ENVIRONMENTAL POLICY ACT

The NEPA process is triggered by a federal action. In this case, the issuance of a completeness letter for the Plan Application and the submittal of the Section 404 permit application triggers the federal action. The NEPA review process is completed with either an EA or an EIS. GSV anticipates that the BLM and the USACE will require an EIS for this project. In addition, Gold Standard anticipates that the BLM will be the lead federal agency for the completion of the NEPA process and the USACE will a cooperating agency under NEPA.

The EIS process is conducted in accordance with NEPA regulations (40 CFR 1500 et. seq.), BLM, as lead federal agency, guidelines for implementing the NEPA in BLM Handbook H-1790-1 (updated January 2008), and BLM Washington Office Bulletin 94-310. The intent of the EIS is to assess the direct, indirect, residual, and cumulative effects of the project and to determine the significance of those effects. Scoping is conducted by the BLM and includes a determination of the environmental resources to be analyzed in the EIS, as well as the degree of analysis for each environmental resource. The scope of the cumulative analysis is also addressed during the scoping process. Following scoping and baseline information collection, the Draft EIS is prepared for the BLM by a third-party contractor. When the BLM determines the Draft EIS is complete, it would be submitted to the public for review. Comments received from the public would be incorporated into a Final EIS, which would in turn be reviewed by the BLM and the public prior to a record of decision (“ROD”). Under an EIS there can be significant impacts. The preparation of an EIS is a lengthier and more expensive process than an EA. The project proponent pays for the third-party contractor to prepare the EIS, and also pays recovery costs to the BLM for any work on the project by BLM specialists.

If the BLM requires the preparation of an EIS to comply with the NEPA for the SRMP, then under the new Secretarial Order 3355 the EIS has to be completed in 365 days (from the Notice of Intent publication in the Federal Register to the signing of the Record of Decision) and must be less than 150 pages (unless a DOI waiver is obtained, which then allows for 300 pages).

20.6 STATE OF NEVADA PERMITS

There are a number of environmental permits issued by the NDEP that are necessary to develop the SRMP and which Gold Standard needs to permit the SRMP ect. The NDEP issues permits that address water and air pollution, as well as land reclamation. The Nevada Division of Water Resources (“NDWR”) issues water rights for the use and management of water.

20.6.1 Water Pollution Control Permit

A WPCP from the BMRR is needed to construct, operate, and close a mining facility in the State of Nevada. The contents of the application are prescribed in the NAC Section 445A.394 through 445A.399. A WPCP application for the project will be prepared and will be based on the following:

Open pit mining, with an anticipated post-mining pit lake formation;
Storage of non-acid and acid generating waste rock;


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Exploration;
Dewatering and water management;
Heap leach and process plant management; and
Ancillary facilities that include storm water diversions, and sediment control basin.

WPCP applications will include an engineering design for waste rock storage areas and mill/tailings facilities, waste rock characterization reports, hydrogeological summary reports, engineering design for process components including methods for the control of storm water runoff, and containment reports detailing specifications for containment of process fluids. Applications will also contain the appropriate WPCP plans, including a process fluid management plan, a monitoring plan, an emergency response plan, a temporary closure plan, and a tentative plan for permanent closure of the mine.

20.6.2 Air Quality Operating Permit

Gold Standard will need an air quality operating permit from the Nevada Bureau of Air Pollution Control (“BAPC”). The permit will likely be a Class II permit, where the emissions of each criteria pollutant would be less than 90.7 tonnes per year. The application would include specifics on each process component that could emit air pollutants and a detailed emissions inventory, as well as air quality modeling. The application preparation and processing time frame would be approximately three months.

20.6.3 Water Rights

Gold Standard will need to obtain water rights from the NDWR. Water and water rights will have to come from either Pine Valley or the Dixie Creek - Ten Mile Creek designated hydrologic basins. These basins are essentially over appropriated. The likelihood of obtaining new water rights is low and the purchase or leasing of existing rights will likely be necessary.

20.6.4 BLM Right-of-Way

A ROW for the powerline from the substation at the Town of Carlin to the project substation will be required. Two ROWs will be issued to NV Energy and Wells Rural Electric, as the powerline will traverse both provider’s service territory. The ROW Applications include a standard BLM form and a Plan of Development (“POD”) that outlines how the powerline would be constructed and operated. To process the application the BLM would have to comply with the NEPA and would analyze this ROW action with the project as described in the Plan Application.

20.7 ELKO COUNTY

Gold Standard will need a Special Use Permit issued by Elko County. This permit will need to include a road maintenance agreement for any county road to be used to access the project.

20.8 OTHER PERMITS

In addition to the principal environmental permits outlined above, Table 20-1 lists other notifications or ministerial permits that may likely be necessary to operate the project.


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Table 20-1: Ministerial Permits, Plans, and Notifications


Notification/Permit	Agency	Timeframe	Comments
Plan of Operations	Bureau of Land Management	Dependent on NEPA
Nevada Reclamation Permit	Nevada Bureau of Mining Regulation and Reclamation	Four Months
Water Pollution Control Permit	Nevada Bureau of Mining Regulation and Reclamation	Eight Months
Air Quality Operating Permit	Nevada Bureau of Air Pollution Control	Three Months
Industrial Artificial Pond Permit	Nevada Department of Wildlife	Three weeks
Water Rights	Nevada Division of Water Resources
Mine Registry	Nevada Division of Minerals	30 days after mine operations begin
Mine Opening Notification	State Inspector of Mines	Before mine operations begin
Solid Waste Landfill	Nevada Bureau of Waste Management	180 days prior to landfill operations
Hazardous Waste Management Permit	Nevada Bureau of Waste Management	Prior to the management or recycling of hazardous waste
General Storm Water Permit	Nevada Bureau of Water Pollution Control	Prior to construction activities
Hazardous Materials Permit	State Fire Marshall	30 days after the start of operations
Fire and Life Safety	State Fire Marshall	Prior to construction
Explosives Permit	Bureau of Alcohol, Tobacco, and Firearms	Prior to purchasing explosives	Mining contractor may be responsible for permit
Mine Identification Number	MSHA	Prior to start-up
Notification of Commencement of Operation	MSHA	Prior to start-up
Radio License	Federal Communications Commission	Prior to radio use


20.9	ENVIRONMENTAL STUDY RESULTS AND KNOWN ISSUES

As previously outlined, the SRMP is a previously explored minerals property with exploration related disturbance. However, there have been very long periods of non-operation. There are no known ongoing environmental issues with any of the regulatory agencies. Gold Standard has been conducting baseline data collection for a couple of years for environmental studies required to support the Plan Application and permitting process. The waste and mineralized material characterization and the hydrogeologic evaluation are currently in their latter stages of development. Material characterization indicates the need to manage a significant portion of the waste rock as potentially acid generating in engineered facilities. Additional results to date indicate limited cultural issues, air quality impacts appear to be within State of Nevada standards, traffic and noise issues are present but at low levels, and socioeconomic impacts are positive. There are golden eagle and Greater sage-grouse in the SRMP and the vicinity, which will need to be addressed in the permitting of the project. Gold Standard is working with the BLM on the management of these species.


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20.10	WASTE DISPOSAL AND MONITORING

Waste rock characterization is being conducted and it is anticipated that the results will indicate that a portion of the waste rock and mineralized material are likely to be reactive, acid generating, and would leach metals. As a result, a detailed waste rock management plan and waste rock management strategy will need to be developed. This strategy and plan will be developed once the characterization work is completed.


20.11	SOCIAL AND COMMUNITY ISSUES

Social and community impacts have been and are being considered and evaluated for the Plan Amendment and Plan Application performed for the project in accordance with the NEPA and other federal laws. Potentially affected Native American tribes, tribal organizations, and/or individuals are consulted during the preparation of all plan amendments to advise on the proposed projects that may have an effect on cultural sites, resources, and traditional activities.

Potential community impacts to existing population and demographics, income, employment, economy, public finance, housing, community facilities, and community services are evaluated for potential impacts as part of the NEPA process. There are no known social or community issues that would have a material impact on the project’s ability to extract mineral resources. Identified socioeconomic issues (employment, payroll, services and supply purchases, and state and local tax payments) are anticipated to be positive.


20.12	MINE CLOSURE

A Tentative Plan for Permanent Closure (“TPPC”) for the project would be submitted to the BMRR with the WPCP application. In the TPPC, the proposed heap leach closure approach would consist of fluid management through evaporation, covering the heap leach growth media, and then revegetating. The design of the process components is not sufficiently advanced to determine the closure costs. Any residual heap leach drainage will be managed with evaporation cells.

The current bond for the SRMP is approximately $1,135,693 to reclaim the exploration related disturbance.


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21	CAPITAL AND OPERATING COSTS

Capital and operating costs were estimated for the PFS by MDA (mining and infrastructure) and KCA (process plant), Stantec (waters systems and storm water diversion), The MINES Group (heap leach facilities) and M3 (site development and ancillaries. Table 21-1 shows the estimated capital costs for the project. This includes $194 million dollars in Year -1, $88.3 million in Year 1 and $20.4 million for sustaining capital. Total capital costs are estimated at $302.7 million.

Table 21-1: Capital Cost Summary


Cateogry	Units		Initial		Expansion		Sustaining		Total
Site General (inc Roads, Water Systems)	K USD	$	16,226	$	4,915	$	8,167	$	29,308
Crush, Agglomeration, & Stacking	K USD	$	2,046	$	32,191	$	1,400	$	35,638
Heap Leach Pad (including Liners)	K USD	$	10,775	$	7,343	$	4,196	$	22,315
Process Plant (ADR, Refinery, Reagents)	K USD	$	16,798	$	-	$	-	$	16,798
Power Supply & Distribution	K USD	$	5,172	$	-	$	-	$	5,172
ADR & Refinery Building	K USD	$	4,553	$	730	$	-	$	5,283
Ancillaries (Warehouse, Maint, Admin, Fuel)	K USD	$	7,615	$	-	$	-	$	7,615
Freight	K USD	$	1,823	$	2,002	$	-	$	3,825
Sub-Total Direct Cost (Process Plant)	K USD	$	65,009	$	47,181	$	13,763	$	125,954
Construction Support (inc. Mobilization)	K USD	$	1,801	$	1,101	$	-	$	2,903
Engineering, Procurement, Const. Mgmt.	K USD	$	7,771	$	7,363	$	-	$	15,134
Vendor Support	K USD	$	1,880	$	381	$	-	$	2,261
Spare Parts	K USD	$	579	$	838	$	-	$	1,417
Contingency	K USD	$	11,556	$	8,530	$	-	$	20,086
Owner's Cost	K USD	$	6,420	$	4,649	$	-	$	11,069
Taxes (County)	K USD	$	1,618	$	1,777	$	-	$	3,395
Sub-Total Indirect Cost	K USD	$	31,625	$	24,638	$	-	$	56,263
Mine Capital Equipment	K USD	$	69,143	$	16,485	$	6,617	$	92,246
Preproduction Costs	K USD	$	28,226	$	-	$	-	$	28,226
Sub-Total Mine Capital	K USD	$	97,369	$	16,485	$	6,617	$	120,472
TOTAL CAPITAL COST	K USD	$	194,004	$	88,304	$	20,380	$	302,688

Table 21-2 shows the estimated operating costs for the LOM project. Operating costs were estimated at $560 million for the LOM. This is $11.83 per tonne processed or $601 per ounce of gold produced.

Table 21-2: Operating Cost Summary


			Production Cost
Category		K USD		$ / ton		$ / Au oz		$ / Au oz*
Mining Costs	$	348,505	$	7.36	$	374.25	$	-
Process Plant	$	156,936	$	3.31	$	168.53	$	-
G&A	$	33,637	$	0.71	$	36.12	$	-
Refining	$	4,679	$	0.10	$	5.03	$	-
Royalty	$	16,282	$	0.34	$	17.48	$	-
TOTAL OPERATING COST	$	560,039	$	11.83	$	601.41	$	582.30

* Including Silver Credit as a Reduction to Total Operating Cost


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21.1	MINING CAPITAL

Mining capital estimates for this PFS assume owner operations of mining equipment and were based on the equipment and facilities required to achieve the production schedule shown in Table 16-4. Capital costs are based on vendor quotations, estimation guides, and recent costs for similar projects. The mining capital estimate is summarized by year in Table 21-3.

Table 21-3: Mining Capital Cost by Year


*Total Mining Capital*	Units		Pre-Prod		Yr 1		Yr 2		Yr 3		Yr 4		Yr 5		Yr 6		Yr 7		Yr 8		Total
Primary Equipment	KUSD	$	49,202	$	16,475	$	3,505	$	537	$	570	$	605	$	642	$	682	$	-	$	72,218
Support Equipment	KUSD	$	12,731	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	12,731
Blasting Equipment	KUSD	$	409	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	409
Mine Maintenance Equipment	KUSD	$	1,695	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	1,695
Other Mine Capital	KUSD	$	5,107	$	10	$	2	$	63	$	13	$	-	$	-	$	-	$	-	$	5,193
Mine Preproduction	KUSD	$	28,226	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	28,226
Total Mine Capital	KUSD	$	97,369	$	16,485	$	3,506	$	600	$	583	$	605	$	642	$	682	$	-	$	120,472


21.1.1	Primary Equipment

Primary equipment purchases refer to the purchase of drills, loading equipment, and haul trucks. The total LOM primary equipment cost estimate is $72.2 million which includes:

$1.4 million for pioneering drills;
$7.1 million for production drills;
$5.1 million for a large loader;
$13.7 million for hydraulic shovels; and

$45.0 million for 136-tonne capacity haul trucks. Note that the large loader is assumed to be leased from year 1 through year 7. The lease terms are assumed to be 20% down and 6% per year with quarterly payments. The down payment and principal payments are capitalized; thus, the full capital is realized, but it is spread out over the 7-year period. The 6% interest payments are added to the capital costs.


21.1.2	Support Equipment

Support equipment includes the equipment required to support the primary mining equipment. This includes dozers to manage dumping locations and cleanup of benches for drilling and loading equipment. This also includes road maintenance equipment such as water trucks and graders. The total estimated capital for support equipment is $12.7 million and includes:

$5.6 million for dozers of various sizes;
$2.8 million for motor graders;
$3.6 million for water trucks;
$61,000 for in-pit pumps to control runoff water;
$505,000 for a 50-ton capacity crane (to be shared between mining and process); and
$116,000 for a flatbed truck used for moving maintenance items within the mine.


21.1.3	Blasting Equipment

Blasting equipment includes explosives trucks for use in loading blast holes and a skid loader to be used for stemming holes. The cost estimate for blasting equipment is $409,000 which includes $280,000 for one explosives truck and $129,000 for a skid loader.


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21.1.4	Mine Maintenance Capital

Mine maintenance capital includes one large lubrication truck at $1,017,000; two mechanic’s trucks totaling $321,000, and two tire trucks totaling $356,000.


21.1.5	Other Capital

Other capital includes an assortment of equipment and facilities totaling $5.1 million. This includes:

$81,000 for light plants;
$80,000 for ANFO storage bins;
$12,000 for powder magazines to store boosters;
$7,000 for a cap magazine;
$61,000 for mobile radios in equipment and assorted handheld radios;
$750,000 for general shop equipment including hoists and other tooling;
$105,000 for engineering computers, plotters, and other office equipment;
$20,000 for dust suppression storage bladders;
$150,000 for surveying equipment and GPS base stations;
$53,000 for fuel island facilities;
$3.5 million for shop and office facilities;
$225,000 in access roads to each deposit and site preparation; and
$150,000 for ambulance and firefighting equipment.

Note that the access roads to each deposit and site preparations are estimated for each deposit with $150,000 applied to the development of Dark Star; $75,000 applied for the preparation of Pinion. These amounts do not include the estimate for the main access road from Highway 278.


21.1.6	Mine Pre-production

Mine pre-production is considered as the cost of all mining prior to the start of gold production from the ROM leach pad. For the PFS, this is a 3-month period. The total mining costs during pre-production total $28.3 million.


21.2	PROCESS CAPITAL

21.2.1	Process Capital Cost Summary

The process plant costs are comprised of costs for the process facilities, as well as costs for site access, heap leach pad and ponds construction, infrastructure development, power supply and distribution, and ancillaries. These direct costs are developed from labor, materials, plant equipment, construction equipment, and freight. Indirect costs are applied to the direct costs to account for items such as: construction support; engineering, procurement, and construction management (EPCM); vendor support during specialty construction and commissioning; spare parts; contingency; owner’s costs; and taxes. Together, the direct and indirect costs form the capital costs.

The direct process plant cost for this PFS has multiple contributors. Stantec developed the direct costs for the water systems and storm water diversion system. KCA developed the costs for the process plant, including crushing, agglomeration, stacking, the adsorption, desorption, recovery plant, refinery, and reagents. The Mines Group developed the costs for the heap leach facility. M3 developed the costs for site access, site layout, and several ancillaries, such as the laboratory, warehouse and maintenance, including the truck shop, administration building, and the fuel station.


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Indirect costs were then calculated following industry accepted methodologies, including contingency at 15% of total contracted cost. Total contracted costs include all process plant direct costs, plus construction support costs, EPCM costs, vendor support costs, and spare parts costs. First fills were calculated by KCA. Owner’s Costs were added at 8% of total contracted costs. Elko County Sales taxes are included at 7.10% of plant equipment and material costs.

Process plant capital costs were independently developed for the initial phase of the project and well as for the expansion phase. All capital cost estimates are based on the purchase of new equipment.

The total evaluated project cost is projected to be in the accuracy range of +/-20%.

Table 21-4: Initial Capital Process Plant Cost Summary


Category (all costs are in USD 1,000)	Labor	Plant Equip.	Material	Sub Contract	Const. Equip.	Total
Site General (inc Roads, Water Systems)	8,473	-	3,910		3,843	16,226
Crush, Agglomeration, & Stacking	793	87	524	-	642	2,046
Heap Leach Pad (including Liners)	3,902	-	3,007	-	3,866	10,775
Process Plant (ADR, Refinery, Reagents)	6,375	7,421	1,542	-	1,460	16,798
Power Supply & Distribution	560	350	531	3,643	88	5,172
ADR & Refinery Building	1,409	1,157	1,584	90	314	4,553
Ancillaries (Warehouse, Maint, Admin, Fuel)	1,693	1,510	1,167	3,071	175	7,615
Freight	-	842	981	-	-	1,823
Sub-Total Direct Cost (Process Plant)	23,206	11,367	13,246	6,803	10,388	65,009
Construction Support (including Mobilization)						1,801
Engineering, Procurement, Construction Management						7,771
Vendor Support						1,880
Spare Parts						579
Contingency						11,556
Owner's Cost						6,420
Taxes (County)						1,618
Sub-Total Indirect Cost						31,625
TOTAL CAPITAL COST						96,634


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Table 21-5: Expansion Capital Process Plant Cost Summary


Category (all costs are in USD 1,000)	Labor	Plant Equip.	Material	Sub Contract	Const. Equip.	Total
Site General (inc Roads, Water Systems)	3,045	-	1,169	-	701	4,915
Crush, Agglomeration, & Stacking	9,171	15,233	6,149	161	1,478	32,191
Heap Leach Pad (including Liners)	2,775	-	1,957	-	2,611	7,343
Process Plant (ADR, Refinery, Reagents)	-	-	-	-	-	-
Power Supply & Distribution	-	-	-	-	-	-
ADR & Refinery Building	214	-	515	-	-	730
Ancillaries (Warehouse, Maint, Admin, Fuel)	-	-	-	-	-	-
Freight	-	1,219	783	-	-	2,002
Sub-Total Direct Cost (Process Plant)	15,206	16,451	10,573	161	4,790	47,181
Construction Support (including Mobilization)						1,101
Engineering, Procurement, Construction Management						7,363
Vendor Support						381
Spare Parts						838
Contingency						8,530
Owner's Cost						4,649
Taxes (County)						1,777
Sub-Total Indirect Cost						24,638
TOTAL CAPITAL COST						71,819


21.2.2	Freight

Estimates for equipment and material freight costs are based on bulk freight loads and have been estimated at 8% of the equipment cost.


21.2.3	Construction Support

Mobilization is included as an indirect cost at 3.5% of civil and concrete total commodity direct costs and at 1% of the balance of the other commodity direct costs.

Temporary construction facilities are included at 0.5% of total direct field cost (TDFC). Temporary construction power is included at 0.1% of TDFC.


21.2.4	EPCM

Engineering is included at 6.0% of total constructed cost (TCC), excluding KCA turn-key scope items, which are detailed in the vendor support section. Project services are included at 1.0% of TCC, excluding KCA turn-key scope items. Project controls are included at 0.75% of TCC, excluding KCA turn-key scope items. Construction Management is included at 6.5% of TCC, excluding KCA turn-key scope items.

An EPCM Fee has been excluded.

EPCM construction trailers are included at 0.25% of TCC, excluding KCA turn-key scope items.


21.2.5	Vendor Support

Supplier engineering for the ADR and associated facilities is quoted at $1,617,000 as part of the turn-key ADR proposal. The turn-key scope includes design and supply of the pregnant and barren solution handling system, the ADR facility, the refinery, and the plant reagents systems.


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Vendor supervision of specialty construction is included at 1.5% of plant equipment supply costs. Vendor pre-commissioning is included at 0.5% of plant equipment supply costs. Vendor commissioning is included at 0.5% of plant equipment supply costs.


21.2.6	Spare Parts

Capital spare parts are included at 5.0% of plant equipment supply costs. Commissioning spare parts are included at 0.5% of plant equipment supply costs. Two-year operating spare parts are excluded.


21.3	OWNER’S COSTS

Owner’s costs were added at 8% of total contracted costs. Total contracted costs include all process plant direct costs, plus construction support costs, EPCM costs, vendor support costs, and spare parts costs. The initial phase capital and the expansion phase capital each includes an allocation for the owner’s costs.


21.3.1	Land Purchases

At total of $3,860,000 is estimated for surface estate purchase obligations based on amounts provided by Gold Standard. This includes costs for mining leases and surface use agreements.


21.4	MINE OPERATING COST

Mine operating costs were estimated using first principals. This is done using estimated hourly costs of equipment and personnel against the anticipated hours of work for each. The equipment hourly costs are estimated for fuel, oil and lubrication, tires, under-carriage, repair and maintenance costs, and special wear items.

The largest consumable costs are those in tires, fuel, and explosives. Tire costs vary by equipment and assumed hours per tire. Fuel cost were assumed to be $0.66 per liter ($2.50 per gallon). ANFO and emulsion blend is assumed to be $700 per tonne with an additional charge of $35.00 per tonne for transportation.

Personnel costs include supervision, operating labor, and maintenance labor. The mine operating costs are summarized by year in Table 21-6. Note that while the costs for pre-production are shown in the cost tables below, these costs are capitalized as pre-production costs. The average LOM mine operating cost with pre-production is estimated to be $376.7 million or $1.94 per tonne mined.


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Table 21-6: Yearly Mine Operating Cost Estimate


Mine Op Cost Summary	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Mine General Service	K USD	$740	$844	$844	$844	$844	$844	$844	$844	$282	$6,929
Mine Maintenance	K USD	$1,868	$2,491	$2,491	$2,491	$2,492	$2,491	$2,491	$2,491	$1,120	$20,426
Engineering	K USD	$565	$716	$716	$716	$716	$716	$716	$716	$239	$5,819
Geology	K USD	$331	$414	$414	$414	$414	$414	$414	$414	$138	$3,367
Drilling	K USD	$3,369	$6,735	$6,980	$7,098	$7,632	$7,471	$6,109	$3,496	$977	$49,868
Blasting	K USD	$3,512	$6,673	$5,272	$5,004	$6,029	$6,078	$5,775	$2,265	$627	$41,235
Loading	K USD	$3,894	$8,338	$6,800	$6,564	$7,599	$7,802	$7,415	$3,034	$828	$52,275
Hauling	K USD	$9,469	$21,885	$24,828	$23,818	$22,099	$20,360	$17,010	$7,332	$1,796	$148,596
Mine Support	K USD	$4,478	$5,956	$5,956	$5,956	$5,965	$5,956	$5,956	$5,956	$1,979	$48,158
Total Mining Cost	K USD	$28,226	$54,111	$54,301	$52,905	$53,790	$52,132	$46,731	$26,548	$7,987	$376,731
*Cost per Ton*
Mine General Service	$/t	$0.05	$0.03	$0.03	$0.04	$0.03	$0.03	$0.03	$0.09	$0.12	$0.04
Mine Maintenance	$/t	$0.12	$0.08	$0.10	$0.11	$0.09	$0.09	$0.09	$0.27	$0.46	$0.10
Engineering	$/t	$0.04	$0.02	$0.03	$0.03	$0.03	$0.02	$0.03	$0.08	$0.10	$0.03
Geology	$/t	$0.02	$0.01	$0.02	$0.02	$0.01	$0.01	$0.01	$0.04	$0.06	$0.02
Drilling	$/t	$0.21	$0.21	$0.28	$0.30	$0.27	$0.26	$0.22	$0.37	$0.40	$0.26
Blasting	$/t	$0.22	$0.20	$0.21	$0.21	$0.21	$0.21	$0.21	$0.24	$0.26	$0.21
Loading	$/t	$0.24	$0.26	$0.27	$0.28	$0.27	$0.27	$0.27	$0.33	$0.34	$0.27
Hauling	$/t	$0.59	$0.67	$0.99	$1.01	$0.78	$0.70	$0.61	$0.79	$0.74	$0.76
Mine Support	$/t	$0.28	$0.18	$0.24	$0.25	$0.21	$0.20	$0.21	$0.64	$0.81	$0.25
Total Mining Cost	$/t	$1.76	$1.66	$2.17	$2.24	$1.90	$1.79	$1.67	$2.84	$3.27	$1.94


21.4.1	Mine General Services

Mine general services includes mining supervision along with engineering and geology services. Supervision allows for a mine superintendent, mine general foreman and mine shift foremen. Engineering personnel include a chief engineer along with engineers and surveying crew to support mine planning and operations. Geology is intended to support ore control, geological mapping, and sampling requirements.

Table 21-7 shows the yearly cost estimate for the mine general services.


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Table 21-7: Mine General Services Costs


21.4.2	Mine Maintenance

Mine maintenance costs include the cost of personnel for maintenance, supervision, and planning, along with shop support personnel, including light vehicle mechanics, welders, servicemen, tire men, and maintenance labor.

The estimated mine maintenance costs are shown in Table 21-8. Note that these costs do not include the maintenance labor directly allocated to the various equipment, which is accounted for in the other mining cost categories.

Table 21-8: Yearly Mine Maintenance Costs


*Wages & Salaries*	Units		Pre-Prod		Yr 1		Yr 2		Yr 3		Yr 4		Yr 5		Yr 6		Yr 7		Yr 8		Yr 9		Total
Supervision	K USD	$	383	$	511	$	511	$	511	$	511	$	511	$	511	$	511	$	235	$	-	$	4,192
Planners	K USD	$	116	$	155	$	155	$	155	$	155	$	155	$	155	$	155	$	77	$	-	$	1,276
Hourly Personnel	K USD	$	886	$	1,181	$	1,181	$	1,181	$	1,181	$	1,181	$	1,181	$	1,181	$	591	$	-	$	9,744
Total	K USD	$	1,385	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	903	$	-	$	15,212
*Other Costs*
Supplies	K USD	$	108	$	144	$	144	$	144	$	144	$	144	$	144	$	144	$	48	$	-	$	1,164
Light Vehicles	K USD	$	14	$	21	$	21	$	21	$	21	$	21	$	21	$	21	$	11	$	-	$	175
Total	K USD	$	122	$	165	$	165	$	165	$	165	$	165	$	165	$	165	$	59	$	-	$	1,339

Consumables & Other Costs	K USD	$	407	$	543	$	543	$	543	$	544	$	543	$	543	$	543	$	184	$	-	$	4,394
Parts / MARC Cost	K USD	$	76	$	101	$	101	$	101	$	102	$	101	$	101	$	101	$	34	$	-	$	820
Wages & Salaries	K USD	$	1,385	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	1,846	$	903	$	-	$	15,212
Total	K USD	$	1,868	$	2,491	$	2,491	$	2,491	$	2,492	$	2,491	$	2,491	$	2,491	$	1,120	$	-	$	20,426

Consumables	$/t	$	0.03	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.06	$	0.08	$	-	$	0.02
Parts / MARC Cost	$/t	$	0.00	$	0.00	$	0.00	$	0.00	$	0.00	$	0.00	$	0.00	$	0.01	$	0.01	$	-	$	0.00
Maintenance Labor	$/t	$	0.09	$	0.06	$	0.07	$	0.08	$	0.07	$	0.06	$	0.07	$	0.20	$	0.37	$	-	$	0.08
Total	$/t	$	0.12	$	0.08	$	0.10	$	0.11	$	0.09	$	0.09	$	0.09	$	0.27	$	0.46	$	-	$	0.10


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21.4.3	Drilling

Drilling cost estimates are shown in Table 21-9. The LOM drilling costs are estimated to be $49.9 million or $0.26 per tonne including pre-production.

Table 21-9: Yearly Drilling Costs


*Drilling Operating Costs*	Units		Pre-Prod		Yr 1		Yr 2		Yr 3		Yr 4		Yr 5		Yr 6		Yr 7		Yr 8		Yr 9		Total
Prod Drill Fuel Consumption	K Liters		629		1,457		1,473		1,525		1,610		1,627		1,322		700		200		-		10,542
Prod Drill Fuel Cost	K USD	$	415	$	962	$	973	$	1,007	$	1,063	$	1,074	$	873	$	462	$	132	$	-	$	6,962
Prod Drill Lube & Oil	K USD	$	214	$	496	$	502	$	519	$	548	$	554	$	450	$	239	$	68	$	-	$	3,591
Prod Drill Undercarriage	K USD	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-
Prod Drill Drill Bits & Steel	K USD	$	689	$	1,595	$	1,613	$	1,670	$	1,763	$	1,781	$	1,447	$	767	$	219	$	-	$	11,544
Prod Drill Total Consumables	K USD	$	1,318	$	3,054	$	3,087	$	3,196	$	3,375	$	3,410	$	2,770	$	1,468	$	419	$	-	$	22,097
Prod Drill Parts	K USD	$	689	$	1,595	$	1,613	$	1,670	$	1,763	$	1,781	$	1,447	$	767	$	219	$	-	$	11,544
Prod Drill Maintenance Labor	K USD	$	332	$	677	$	739	$	724	$	743	$	739	$	614	$	409	$	110	$	-	$	5,087
Pioneer Drill Fuel Consumption	K Liters		80		-		-		-		37		-		-		-		-		-		117
Pioneer Drill Fuel Cost	K USD	$	53	$	-	$	-	$	-	$	24	$	-	$	-	$	-	$	-	$	-	$	77
Pioneer Drill Lube & Oil	K USD	$	15	$	-	$	-	$	-	$	7	$	-	$	-	$	-	$	-	$	-	$	22
Pioneer Drill Undercarriage	K USD	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-
Pioneer Drill Drill Bits & Steel	K USD	$	40	$	-	$	-	$	-	$	19	$	-	$	-	$	-	$	-	$	-	$	59
Pioneer Drill Total Consumables	K USD	$	108	$	-	$	-	$	-	$	50	$	-	$	-	$	-	$	-	$	-	$	158
Pioneer Drill Parts / MARC Cost	K USD	$	40	$	-	$	-	$	-	$	19	$	-	$	-	$	-	$	-	$	-	$	59
Pioneer Drill Maintenance Labor	K USD	$	61	$	-	$	-	$	-	$	44	$	-	$	-	$	-	$	-	$	-	$	105
Total Drill Fuel Consumption	K Liters		709		1,457		1,473		1,525		1,647		1,627		1,322		700		200		-		10,659
Total Drill Fuel Cost	K USD	$	468	$	962	$	973	$	1,007	$	1,088	$	1,074	$	873	$	462	$	132	$	-	$	7,040
Total Drill Lube & Oil	K USD	$	229	$	496	$	502	$	519	$	555	$	554	$	450	$	239	$	68	$	-	$	3,613
Total Drill Undercarriage	K USD	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-
Total Drill Drill Bits & Steel	K USD	$	729	$	1,595	$	1,613	$	1,670	$	1,782	$	1,781	$	1,447	$	767	$	219	$	-	$	11,603
Total Drill Total Consumables	K USD	$	1,427	$	3,054	$	3,087	$	3,196	$	3,425	$	3,410	$	2,770	$	1,468	$	419	$	-	$	22,256
Total Drill Parts / MARC Cost	K USD	$	729	$	1,595	$	1,613	$	1,670	$	1,782	$	1,781	$	1,447	$	767	$	219	$	-	$	11,603
Total Drill Maintenance Labor	K USD	$	393	$	677	$	739	$	724	$	787	$	739	$	614	$	409	$	110	$	-	$	5,192
Total Drill Maintenance Allocation	K USD	$	1,123	$	2,272	$	2,352	$	2,394	$	2,568	$	2,521	$	2,061	$	1,176	$	329	$	-	$	16,795
Total Operator Wages & Burden	K USD	$	820	$	1,410	$	1,541	$	1,508	$	1,639	$	1,541	$	1,278	$	852	$	229	$	-	$	10,818
Total Drilling Cost	K USD	$	3,369	$	6,735	$	6,980	$	7,098	$	7,632	$	7,471	$	6,109	$	3,496	$	977	$	-	$	49,868
*Drilling Cost per Tonne Mined by Item*
Fuel Cost	$/t	$	0.03	$	0.03	$	0.04	$	0.04	$	0.04	$	0.04	$	0.03	$	0.05	$	0.05	$	-	$	0.04
Lube & Oil	$/t	$	0.01	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.03	$	0.03	$	-	$	0.02
Undercarriage	$/t	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-	$	-
Drill Bits & Steel	$/t	$	0.05	$	0.05	$	0.06	$	0.07	$	0.06	$	0.06	$	0.05	$	0.08	$	0.09	$	-	$	0.06
Total Consumables	$/t	$	0.09	$	0.09	$	0.12	$	0.14	$	0.12	$	0.12	$	0.10	$	0.16	$	0.17	$	-	$	0.11
Parts / MARC Cost	$/t	$	0.05	$	0.05	$	0.06	$	0.07	$	0.06	$	0.06	$	0.05	$	0.08	$	0.09	$	-	$	0.06
Maintenance Labor	$/t	$	0.02	$	0.02	$	0.03	$	0.03	$	0.03	$	0.03	$	0.02	$	0.04	$	0.05	$	-	$	0.03
Total Maintenance Allocation	$/t	$	0.07	$	0.07	$	0.09	$	0.10	$	0.09	$	0.09	$	0.07	$	0.13	$	0.13	$	-	$	0.09
Operator Wages & Burden	$/t	$	0.05	$	0.04	$	0.06	$	0.06	$	0.06	$	0.05	$	0.05	$	0.09	$	0.09	$	-	$	0.06
Total Drilling Cost	$/t	$	0.21	$	0.21	$	0.28	$	0.30	$	0.27	$	0.26	$	0.22	$	0.37	$	0.40	$	-	$	0.26


21.4.4	Blasting

LOM blasting costs, including pre-production, are shown in Table 21-10. These costs are based on owner operations for blasting and assume heavy ANFO costs of $700/tonne with transportation costs for blasting agents at $35/tonne. A blasting accessories cost of $22.00 per hole was included.


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Table 21-10: Yearly Blasting Costs


Blasting Costs	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Fuel	K Liters	70	93	93	93	93	93	93	93	31	751
Blasting Consumables	K USD	$3,175	$6,225	$4,823	$4,555	$5,580	$5,630	$5,327	$1,817	$478	$37,609
Equipment Consumables	K USD	$53	$71	$71	$71	$71	$71	$71	$71	$23	$572
Equipment Maintenance
Allocations	K USD	$12	$15	$15	$15	$15	$15	$15	$15	$5	$125
Personnel	K USD	$254	$338	$338	$338	$338	$338	$338	$338	$113	$2,734
Supplies	K USD	$9	$12	$12	$12	$12	$12	$12	$12	$4	$97
Outside Services	K USD	$9	$12	$12	$12	$12	$12	$12	$12	$4	$97
Total Blasting Costs	K USD	$3,512	$6,673	$5,272	$5,004	$6,029	$6,078	$5,775	$2,265	$627	$41,235
Cost per Ton
Blasting Consumables	$/t	$0.20	$0.19	$0.19	$0.19	$0.20	$0.19	$0.19	$0.19	$0.20	$0.19
Equipment Consumables	$/t	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.01	$0.01	$0.00
Equipment Maintenance	$/t	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00
Allocations
Personnel	$/t	$0.02	$0.01	$0.01	$0.01	$0.01	$0.01	$0.01	$0.04	$0.05	$0.01
Supplies	$/t	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00
Outside Services	$/t	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00
Total	$/t	$0.22	$0.20	$0.21	$0.21	$0.21	$0.21	$0.21	$0.24	$0.26	$0.21


21.4.5	Loading

Loading costs are based on operation of two hydraulic shovels with 22 cubic meter buckets for all primary production. In addition, a 21 cubic yard front-end-loader is assumed to be used for stockpile management and re-handling as well as backup for production during shovel maintenance. The yearly loading cost estimate is shown in Table 21-11.


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Table 21-11: Yearly Loading Costs


*Shovel Cost*	Units		Pre-Prod		Yr 1		Yr 2		Yr 3		Yr 4		Yr 5		Yr 6		Yr 7		Yr 8		Yr 9		Total
Fuel Consumption	K Liters		610		1,130		858		807		970		999		956		319		84		-		6,734
Fuel Cost	K USD	$	403	$	746	$	567	$	533	$	641	$	660	$	631	$	211	$	55	$	-	$	4,447
Lube & Oil	K USD	$	372	$	689	$	523	$	492	$	592	$	610	$	583	$	195	$	51	$	-	$	4,108
Tires / Under Carriage	K USD	$	259	$	480	$	364	$	343	$	412	$	424	$	406	$	136	$	35	$	-	$	2,859
Wear Items & GET	K USD	$	299	$	553	$	420	$	395	$	475	$	490	$	468	$	156	$	41	$	-	$	3,298
Total Consumables	K USD	$	1,333	$	2,468	$	1,875	$	1,763	$	2,120	$	2,184	$	2,088	$	698	$	183	$	-	$	14,712
Parts / MARC Cost	K USD	$	1,703	$	3,151	$	2,394	$	2,251	$	2,707	$	2,788	$	2,666	$	891	$	233	$	-	$	18,784
Total Equip. Allocation (no labor)	K USD	$	3,036	$	5,620	$	4,269	$	4,014	$	4,827	$	4,972	$	4,754	$	1,589	$	416	$	-	$	33,495
*Loader Cost*
Fuel Consumption	K Liters		-		465		362		373		466		574		570		273		87		-		3,170
Fuel Cost	K USD	$	-	$	307	$	239	$	246	$	308	$	379	$	376	$	180	$	57	$	-	$	2,094
Lube & Oil	K USD	$	-	$	96	$	75	$	77	$	96	$	118	$	117	$	56	$	18	$	-	$	653
Tires / Under Carriage	K USD	$	-	$	231	$	201	$	170	$	137	$	102	$	65	$	26	$	-	$	-	$	932
Wear Items & GET	K USD	$	-	$	30	$	24	$	24	$	30	$	38	$	37	$	18	$	6	$	-	$	207
Total Consumables	K USD	$	-	$	664	$	539	$	518	$	572	$	637	$	596	$	280	$	81	$	-	$	3,887
Parts / MARC Cost	K USD	$	-	$	174	$	135	$	139	$	174	$	214	$	213	$	102	$	32	$	-	$	1,183
Total Equip. Allocation (no labor)	K USD	$	-	$	838	$	674	$	657	$	746	$	852	$	809	$	382	$	113	$	-	$	5,070
*Total Loading Cost*
Fuel Consumption	K Liters		610		1,595		1,220		1,180		1,437		1,574		1,525		592		171		-		9,904
Fuel Cost	K USD	$	403	$	1,053	$	806	$	779	$	949	$	1,039	$	1,007	$	391	$	113	$	-	$	6,541
Lube & Oil	K USD	$	372	$	785	$	598	$	569	$	688	$	728	$	700	$	251	$	69	$	-	$	4,761
Tires / Under Carriage	K USD	$	259	$	710	$	566	$	513	$	549	$	527	$	471	$	161	$	35	$	-	$	3,791
Wear Items & GET	K USD	$	299	$	584	$	444	$	420	$	506	$	527	$	505	$	174	$	47	$	-	$	3,506
Total Consumables	K USD	$	1,333	$	3,132	$	2,413	$	2,281	$	2,692	$	2,821	$	2,684	$	978	$	264	$	-	$	18,598
Parts / MARC Cost	K USD	$	1,703	$	3,325	$	2,529	$	2,390	$	2,881	$	3,002	$	2,879	$	993	$	265	$	-	$	19,967
Total Equip. Allocation (no labor)	K USD	$	3,036	$	6,457	$	4,942	$	4,671	$	5,573	$	5,823	$	5,563	$	1,971	$	529	$	-	$	38,565
Maintenance Labor	K USD	$	393	$	677	$	739	$	724	$	787	$	739	$	614	$	409	$	110	$	-	$	5,192
Operator Wages & Burden	K USD	$	465	$	1,204	$	1,118	$	1,170	$	1,239	$	1,239	$	1,239	$	654	$	189	$	-	$	8,517
Total Loading Costs	K USD	$	3,894	$	8,338	$	6,800	$	6,564	$	7,599	$	7,802	$	7,415	$	3,034	$	828	$	-	$	52,275
*Cost per Ton*
Fuel Cost	$/t	$	0.03	$	0.03	$	0.03	$	0.03	$	0.03	$	0.04	$	0.04	$	0.04	$	0.05	$	-	$	0.03
Lube & Oil	$/t	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.03	$	0.03	$	0.03	$	-	$	0.02
Tires / Under Carriage	$/t	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.01	$	-	$	0.02
Wear Items & GET	$/t	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	0.02	$	-	$	0.02
Total Consumables	$/t	$	0.08	$	0.10	$	0.10	$	0.10	$	0.09	$	0.10	$	0.10	$	0.10	$	0.11	$	-	$	0.10
Parts / MARC Cost	$/t	$	0.11	$	0.10	$	0.10	$	0.10	$	0.10	$	0.10	$	0.10	$	0.11	$	0.11	$	-	$	0.10
Total Equip. Allocation (no labor)	$/t	$	0.19	$	0.20	$	0.20	$	0.20	$	0.20	$	0.20	$	0.20	$	0.21	$	0.22	$	-	$	0.20
Maintenance Labor	$/t	$	0.02	$	0.02	$	0.03	$	0.03	$	0.03	$	0.03	$	0.02	$	0.04	$	0.05	$	-	$	0.03
Operator Wages & Burden	$/t	$	0.03	$	0.04	$	0.04	$	0.05	$	0.04	$	0.04	$	0.04	$	0.07	$	0.08	$	-	$	0.04
Total Loading Cost	$/t	$	0.24	$	0.26	$	0.27	$	0.28	$	0.27	$	0.27	$	0.27	$	0.33	$	0.34	$	-	$	0.27


21.4.6	Hauling

Haulage cost was estimated using the truck hour estimates discussed in Section 16.5.3. The yearly haulage cost estimate is shown in Table 21-12.


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Table 21-12: Yearly Haulage Costs


*Total Truck Hours*	Units		Pre-Prod			Yr 1			Yr 2			Yr 3			Yr 4			Yr 5			Yr 6			Yr 7			Yr 8			Yr 9			Total
Productive Hours	Prod Hrs		30,172			67,729			77,746			72,799			66,299			60,815			51,764			20,172			4,869			-			452,363
Operating Efficiency	%		83	%		83	%		83	%		83	%		83	%		83	%		83	%		83	%		83	%		83	%		83	%
Operating Hours	Op Hrs		36,352			81,601			93,670			87,709			79,878			73,271			62,366			24,303			5,866			-			545,015
Equipment Hours	Eq Hrs		41,545			93,258			107,051			100,239			91,289			83,738			71,275			27,775			6,704			-			622,874
Number of Trucks	#		9			11			12			12			12			12			14			14			14			14			14
Truck Availability	%		90	%		90	%		90	%		90	%		90	%		90	%		90	%		90	%		90	%		90	%		90	%
Available Equipment Hours	Op Hrs		37,006			89,822			98,622			98,155			93,427			85,528			68,619			34,704			8,675			-			614,558
Use of Available Hours	%		98	%		91	%		95	%		89	%		85	%		86	%		91	%		70	%		68	%		0	%		89	%
*Haulage Cost*
Fuel Consumption	K Liters		4,359			9,784			11,231			10,516			9,577			8,785			7,478			2,914			703			-			65,347
Fuel Cost	K USD	$	2,879		$	6,462		$	7,417		$	6,945		$	6,325		$	5,802		$	4,938		$	1,924		$	464		$	-		$	43,157
Lube & Oil	K USD	$	972		$	2,182		$	2,505		$	2,346		$	2,136		$	1,959		$	1,668		$	650		$	157		$	-		$	14,575
Tires	K USD	$	1,329		$	2,984		$	3,426		$	3,208		$	2,921		$	2,680		$	2,281		$	889		$	215		$	-		$	19,932
Wear Items & GET	K USD	$	104		$	233		$	268		$	251		$	228		$	209		$	178		$	69		$	17		$	-		$	1,557
Total Consumables	K USD	$	5,284		$	11,861		$	13,616		$	12,749		$	11,611		$	10,650		$	9,065		$	3,533		$	853		$	-		$	79,222
Parts / MARC Cost	K USD	$	1,164		$	2,613		$	3,000		$	2,809		$	2,558		$	2,346		$	1,997		$	778		$	188		$	-		$	17,453
Total Equip. Allocation (no labor)	K USD	$	6,448		$	14,474		$	16,615		$	15,558		$	14,169		$	12,997		$	11,062		$	4,311		$	1,040		$	-		$	96,675
Maintenance Labor	K USD	$	1,007		$	2,470		$	2,738		$	2,753		$	2,643		$	2,454		$	1,982		$	1,007		$	252		$	-		$	17,307
Operator Wages & Burden	K USD	$	2,014		$	4,940		$	5,475		$	5,507		$	5,287		$	4,909		$	3,965		$	2,014		$	503		$	-		$	34,614
Total Haulage Costs	K USD	$	9,469		$	21,885		$	24,828		$	23,818		$	22,099		$	20,360		$	17,010		$	7,332		$	1,796		$	-		$	148,596
*Cost per Tonne Moved*
Fuel Cost	$/t	$	0.18		$	0.20		$	0.30		$	0.29		$	0.22		$	0.20		$	0.18		$	0.21		$	0.19		$	-		$	0.22
Lube & Oil	$/t	$	0.06		$	0.07		$	0.10		$	0.10		$	0.08		$	0.07		$	0.06		$	0.07		$	0.06		$	-		$	0.07
Tires	$/t	$	0.08		$	0.09		$	0.14		$	0.14		$	0.10		$	0.09		$	0.08		$	0.10		$	0.09		$	-		$	0.10
Wear Items & GET	$/t	$	0.01		$	0.01		$	0.01		$	0.01		$	0.01		$	0.01		$	0.01		$	0.01		$	0.01		$	-		$	0.01
Total Consumables	$/t	$	0.33		$	0.36		$	0.54		$	0.54		$	0.41		$	0.36		$	0.32		$	0.38		$	0.35		$	-		$	0.41
Parts / MARC Cost	$/t	$	0.07		$	0.08		$	0.12		$	0.12		$	0.09		$	0.08		$	0.07		$	0.08		$	0.08		$	-		$	0.09
Total Equip. Allocation (no labor)	$/t	$	0.40		$	0.44		$	0.66		$	0.66		$	0.50		$	0.45		$	0.40		$	0.46		$	0.43		$	-		$	0.50
Maintenance Labor	$/t	$	0.06		$	0.08		$	0.11		$	0.12		$	0.09		$	0.08		$	0.07		$	0.11		$	0.10		$	-		$	0.09
Operator Wages & Burden	$/t	$	0.13		$	0.15		$	0.22		$	0.23		$	0.19		$	0.17		$	0.14		$	0.22		$	0.21		$	-		$	0.18
Total Haulage Costs	$/t	$	0.59		$	0.67		$	0.99		$	1.01		$	0.78		$	0.70		$	0.61		$	0.79		$	0.74		$	-		$	0.76


21.4.7	Mine Support

Yearly mine support cost estimates are shown in Table 21-13 including pre-production costs. These costs assume the hourly costs for required support equipment and personnel as discussed in Sections 16.5 and 16.6 respectively.

Table 21-13: Yearly Mine Support Costs


*Total Mine Support Costs*	Units		Pre-Prod		Yr 1		Yr 2		Yr 3		Yr 4		Yr 5		Yr 6		Yr 7		Yr 8		Yr 9		Total
Consumables	K USD	$	1,846	$	2,451	$	2,451	$	2,451	$	2,457	$	2,451	$	2,451	$	2,451	$	812	$	-	$	19,820
Parts / MARC Cost	K USD	$	614	$	815	$	815	$	815	$	817	$	815	$	815	$	815	$	270	$	-	$	6,590
Maintenance Labor	K USD	$	673	$	897	$	897	$	897	$	897	$	897	$	897	$	897	$	299	$	-	$	7,249
Operating Labor	K USD	$	1,345	$	1,794	$	1,794	$	1,794	$	1,794	$	1,794	$	1,794	$	1,794	$	598	$	-	$	14,499
Total	K USD	$	4,478	$	5,956	$	5,956	$	5,956	$	5,965	$	5,956	$	5,956	$	5,956	$	1,979	$	-	$	48,158
*Cost per Tonne Mined*
Consumables	$/t	$	0.12	$	0.07	$	0.10	$	0.10	$	0.09	$	0.08	$	0.09	$	0.26	$	0.33	$	-	$	0.10
Maintenance Allocations	$/t	$	0.04	$	0.02	$	0.03	$	0.03	$	0.03	$	0.03	$	0.03	$	0.09	$	0.11	$	-	$	0.03
Maintenance Labor	$/t	$	0.04	$	0.03	$	0.04	$	0.04	$	0.03	$	0.03	$	0.03	$	0.10	$	0.12	$	-	$	0.04
Operating Labor	$/t	$	0.08	$	0.05	$	0.07	$	0.08	$	0.06	$	0.06	$	0.06	$	0.19	$	0.24	$	-	$	0.07
Total Costs	$/t	$	0.28	$	0.18	$	0.24	$	0.25	$	0.21	$	0.20	$	0.21	$	0.64	$	0.81	$	-	$	0.25


21.5	PROCESS OPERATING COST SUMMARY

Process operating costs have been estimated by KCA from first principles. Labor costs were estimated using project specific staffing, salary and wage, and benefit requirements. Unit consumptions of materials, supplies, power, and delivered supply costs were also estimated. LOM overall average processing costs are estimated at $3.31 per tonne ore with average costs of $1.83 per tonne for ROM ore and $4.87 per tonne for crushed ore. Process operating costs by pit and process type are detailed in Table 21-14.


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Table 21-14: LOM Operating Costs by Pit and Process Type, US$/tonne ore


Type	Pinion	Dark Star	Total
ROM	2.04	1.79	$1.83
Crushed	4.70	5.08	$4.87
Total	3.94	2.92	$3.31

Operating costs were estimated based on 3rd quarter 2019 US dollars and are presented with no added contingency based upon the design and operating criteria present in this Technical Report. Operating costs are considered to have an accuracy of +/- 25%.

The process operating costs presented are based upon the ownership of all process production equipment and site facilities. The owner will employ and direct all operating maintenance and support personnel for all site activities.

Operating costs estimates have been based upon information obtained from the following sources:

Project metallurgical test work and process engineering;
Supplier quotes for reagents and fuel
Recent KCA project file data; and
Experience of KCA staff with other similar operations.

Where specific data do not exist, cost allowances have been based upon consumption and operating requirements from other similar properties for which reliable data exist. Freight costs have been estimated where delivered prices were not available. Overall LOM operating costs by year and process type are presented in Table 21-15.


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Table 21-15: Life of Mine Average Process Operating Cost by Year


Category	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	LOM Total
Total Tonnes
ROM Ore (000's)	4,013	4,563	4,562	4,575	3,722	841	1,435	486	24,197
Crushed Ore (000's)	0	2,993	3,650	3,660	3,638	3,650	3,650	1,534	22,775
TOTAL Ore	4,013	7,556	8,212	8,235	7,360	4,491	5,085	2,019	46,972
ROM Ore (US$ 000's)
Labor (All Process Areas)	$2,775	$2,347	$2,160	$2,228	$1,908	$708	$1,077	$936	$14,138
Area 250 - Heap Leach Systems	$476	$574	$574	$575	$468	$106	$181	$61	$3,014
Area 420 - Gold Recovery	$488	$446	$480	$408	$330	$105	$174	$76	$2,507
Area 510 - Smelting	$113	$110	$112	$109	$90	$23	$38	$14	$610
Area 800 - Reagents	$4,001	$4,520	$4,531	$4,505	$3,622	$816	$1,391	$475	$23,862
Area 900 - Ancillaries and Buildings	$43	$30	$30	$30	$30	$30	$22	$22	$238
TOTAL ROM Ore	$7,896	$8,027	$7,887	$7,856	$6,488	$1,788	$2,883	$1,584	$44,369
Crushed Ore (US$ 000's)
Labor (All Process Areas)	$0	$3,237	$3,424	$3,355	$3,676	$4,876	$4,506	$4,647	$27,721
Area 113 - Primary Crushing	$0	$578	$705	$707	$703	$705	$705	$296	$4,401
Area 114 - Secondary Crushing	$0	$1,694	$2,066	$2,072	$2,059	$2,066	$2,066	$868	$12,891
Area 115 - Tertiary Crushing	$0	$1,487	$1,813	$1,818	$1,807	$1,813	$1,813	$647	$11,199
Area 116 - Agglomeration	$0	$260	$318	$318	$317	$318	$318	$133	$1,981
Area 120 - Stacking	$0	$523	$638	$639	$636	$638	$638	$268	$3,979
Area 250 - Heap Leach Systems	$0	$376	$459	$460	$458	$459	$459	$193	$2,864
Area 420 - Gold Recovery	$0	$817	$1,022	$570	$1,175	$1,300	$1,165	$503	$6,552
Area 510 - Smelting	$0	$111	$135	$108	$155	$165	$155	$69	$897
Area 800 - Reagents	$0	$6,704	$8,156	$7,422	$4,931	$4,556	$4,532	$1,898	$38,199
Area 900 - Ancillaries and Buildings	$0	$30	$30	$30	$30	$30	$30	$30	$212
TOTAL Crushed Ore	$0	$15,818	$18,766	$17,501	$15,947	$16,925	$16,388	$9,553	$110,897
GRAND TOTAL (US$ 000's)	$7,896	$23,845	$26,652	$25,357	$22,395	$18,714	$19,270	$11,137	$155,265
Fixed Costs (US$ 000's)	$2,818	$5,644	$5,644	$5,644	$5,644	$5,644	$5,635	$5,635	$42,309
Variable Costs (US$ 000's)	$5,078	$18,201	$21,008	$19,713	$16,751	$13,070	$13,635	$5,501	$112,957


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21.5.1	Personnel and Staffing

Staffing requirements for process personnel have been estimated by KCA based on experience with similar sized operations. Total process personnel is estimated at 32 persons for the initial ROM operation and 67 persons once crushing and conveyor stacking begins. Personnel requirements and costs are estimated at $2.8 million per year for the initial ROM operation and $5.6 million for the ROM and Crushing operations.


21.5.2	Power

Power usage for the process and process-facilities was derived from estimated connected loads assigned to powered equipment from the mechanical equipment list. Equipment power demands under normal operation were assigned and coupled with estimated on-stream times to determine the average energy usage and cost. Power requirements for the project are presented in Table 21-16.

Table 21-16: Power Requirements Summary


	ROM Process Only			ROM & Crushing Expansion
Area Description	Attached Power (kW)	Demand (kW)	Annual (kW)	Attached Power (kW)	Demand (kW)	Period kWh (kW)
Area 113 - Primary Crushing				445	254	2,224,263
Area 114 - Secondary Crushing	0	0	0	2,260	1,151	10,080,364
Area 115 - Tertiary Crushing	0	0	0	2,244	1,596	13,977,869
Area 116 - Agglomeration	11	6	55,065	420	217	1,902,877
Area 120 - Conveying & Stacking	0	0	0	954	644	5,643,042
Area 250 - Heap Leach System	2,224	503	4,407,112	2,210	1,001	8,766,920
Area 420 - Recovery	374	99	422,845	444	139	694,098
Area 510 - Electrowinning & Smelting	146	49	238,156	146	49	238,156
Area 800 - Reagents & Air	27	2	13,057	3	2	21,891
Area 900 - Ancillaries and Buildings	0	0	0	100	56	246,375
Total	2,782	659	5,136,234	9,226	5,109	43,795,855

The total attached power for the process and process infrastructure is estimated at 2.8 MW for the initial ROM operation, with an average draw of 0.7 MW at start up increasing to 9.2 MW attached with a demand of 5.1 MW once crushing and conveyor stacking begins. Line power will be supplied to the project site at an estimated $0.07/ kW h.


21.5.2	Consumable Items

Operating supplies have been estimated based upon unit costs and consumption rates predicted by metallurgical tests and have been broken down by area. Freight costs are included in all operating supply and reagent estimates. Reagent consumptions have been derived from test work and from design criteria considerations. Other consumable items have been estimated by KCA based on KCA’s experience with other similar operations. Table 21-17 presents average consumptions for major consumables.


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Table 21-17: Process Consumables Average Annual Consumptions


Item	Form	Average Annual Consumption Y1	Average Annual Consumption Y2+
Sodium Cyanide	Liquid at 33% NaCN by Weight	925 tonnes	1,900 tonnes
Lime	Bulk Delivery (20 tonne)	4,500 tonnes	6,300 tonnes
Cement	Bulk Delivery (20 tonne)	N/A	20,800 tonnes
Antiscalant	Liquid Tote 1 m³Bins	39 tonnes	60 tonnes
Carbon	500 kg Supersacks	20 tonnes	72 tonnes
Hydrochloric Acid	Liquid at 32% HCl by weight	99 m³	360 m³
Caustic	Liquid at 50% NaOH by Weight	33 tonnes	120 tonnes
Silica	Dry Solid Sacks	0.6 tonnes	2.8 tonnes
Borax	Dry Solid Sacks	1.0 tonnes	4.4 tonnes
Niter	Dry Solid Sacks	0.5 tonnes	2.2 tonnes
Soda Ash	Dry Solid Sacks	0.4 tonnes	1.7 tonnes

Operating costs for consumable items have been distributed based on tonnage and gold/silver production or smelting batches, as appropriate.


21.5.3.1	Heap Leach Consumables

Pipes, Fittings and Emitters – The heap pipe costs include expenses for broken pipe, fittings and valves, and abandoned tubing. The heap pipe costs are estimated to be $0.03/t ore and are based on previous detailed studies conducted by KCA on similar projects.

Sodium Cyanide (NaCN) – Delivered sodium cyanide is estimated $2.90/kg based on recent supplier quotes. Cyanide is primarily consumed in the heap leach at 0.22 to 0.25 kg/t ore depending on the ore type.

Pebble Lime (CaO) – Pebble lime is consumed at an average rate of 1.00 kg/t ROM ore and 0.5 kg/t for Pinion crushed ore for pH control at the heap. A delivered price of $239/t has been quoted for the project.

Cement – Portland Type II cement is used as a binder for agglomeration to maintain the heap permeability as well as for pH control at the heap. Cement is consumed at an average rate of 42.0 kg/t ore for crushed Pinion ore and 7.0 kg/t for crushed Dark Star ore. A delivered price of $198/t has been estimated based on recent supplier quotes.

Antiscale Agent (Scale Inhibitor) – Antiscalant consumption is based on a dosage range of 6 ppm to the suctions of the barren and pregnant pumps. A delivered price of $6.42 per kg has been quoted for this project.


21.5.3.2	Recovery Plant Consumables

Carbon – Carbon is used for the adsorption of gold and silver from pregnant solution for the heap circuit. Carbon consumption is estimated at 4% per strip batch due to attrition. Carbon supply cost is quoted at $2.31 per kg.

HCl – Hydrochloric acid is used in the acid wash circuit to remove scale from the carbon which inhibits the adsorption of gold and silver. Hydrochloric acid consumption is estimated at 200 liters per tonne of carbon stripped with a quoted supply cost of $0.51 per kg.

Caustic – Caustic is delivered to site as a liquid at 50% concentration by weight. Caustic is used in the ADR and is consumed in the strip and acid wash circuits. Caustic consumption is based on a 2% caustic strip solution with approximately one third of the solution being discarded each strip. Caustic supply cost is quoted at $0.54 per kg.


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Smelting Fluxes - It has been estimated that 1.0 kg of mixed fluxes per kg of precious metal produced will be required. The estimated delivered cost of these fluxes, which includes borax, silica, niter, and soda ash, is $2.45 per kg, which is based on recent quotes in KCA’s database.


21.5.3.3	Laboratory

Operating costs in the laboratory area include power, facilities support, operating and maintenance supplies, and processing of an average of 100 fire assays per day, plus supporting solution assays.


21.5.3.4	Propane

Propane is consumed by the boiler, kiln and smelting furnace during operations. Propane costs are estimated at $1.23 per gallon ($0.33 per L) based on recent project costs.


21.5.3.5	Miscellaneous Operating & Maintenance Supplies

Overhaul and maintenance of equipment along with miscellaneous operating supplies for each area have been estimated as allowances based on tonnes of ore processed. The allowances for each area were developed based on published data as well as KCA’s experience with similar operations.


21.5.4	General Facilities (Process Equipment Costs)

Operating costs in the general facilities area include the operating costs of two D7 dozers, two forklifts, one skid-steer, and a maintenance truck. These are assigned to crushed material. No costs are assigned to ROM material for this area.

Operating costs for other facilities and support, such as site buildings and site maintenance, are excluded from KCA’s costs.


21.5.5	Process Operating Cost Exclusions

The following operating costs are excluded from KCA’s estimate:

G&A costs;
Site maintenance including access roads and internal roads;
Maintenance and operation costs of site utilities and infrastructure including power supply and distribution, water supply and distribution, emergency backup power, administrative facilities, warehousing and storage, septic, waste storage and disposal, and communications;
Any operating costs related to mining and the truck stacking fleet;
Operating cost contingency;
Escalation costs;
Currency exchange fluctuations.


21.6	G&A COSTS

G&A costs were estimated based on personnel requirements for administrative, accounting, safety and security, and environmental departments to support mining and processing activities. Costs are also included for legal, land, permit bonding, power, etc. Table 21-18 shows the yearly G&A cost estimate.


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Table 21-18: Yearly G&A Costs


*Personnel Costs*	Units	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Yr 9	Yr 10	Total
Construction Management Personnel	K USD	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ 250
Admin Salaried Personnel	K USD	$ 504	$ 504	$ 504	$ 504	$ 504	$ 504	$ 504	$ 504	$ 504	$ 84	$ 4,796
Admin Hourly Personnel	K USD	$ 276	$ 276	$ 276	$ 276	$ 276	$ 276	$ 276	$ 254	$ 210	$ 35	$ 2,585
Safety & Security Salaried Personnel	K USD	$ 90	$ 90	$ 90	$ 90	$ 90	$ 90	$ 90	$ 90	$ 90	$ 15	$ 897
Safety & Security Hourly Personnel	K USD	$ 162	$ 162	$ 162	$ 162	$ 162	$ 162	$ 162	$ 162	$ 162	$ 27	$ 1,622
Evironmental Salaried Personnel	K USD	$ 110	$ 110	$ 110	$ 110	$ 110	$ 110	$ 110	$ 110	$ 110	$ 18	$ 1,104
Recruitment Costs	K USD	$36	$ 24	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ 84
Total Personnel Costs	K USD	$ 1,178	$ 1,166	$ 1,142	$ 1,142	$ 1,142	$ 1,142	$ 1,142	$ 1,120	$ 1,076	$ 179	$ 11,337
General G&A Costs
Construction Management Expenses	K USD	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ -	$ 333
Supplies & General Maintenance	K USD	$ 288	$ 288	$ 288	$ 288	$ 288	$ 288	$ 288	$ 288	$ 288	$ 48	$ 2,712
Land Holdings	K USD	$ 240	$ 240	$ 240	$ 240	$ 240	$ 240	$ 240	$ 240	$ 240	$ 40	$ 2,260
Off Site Overhead	K USD	$ 120	$ 120	$ 120	$ 120	$ 120	$ 120	$ 120	$ 120	$ 120	$ 20	$ 1,200
Legal, Audits, Consulting, MSHA	K USD	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 42	$ 2,500
Computers, IT, Internet, Software, Hardware	K USD	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 13	$ 750
Environmental, Monitoring Wells, Reporting	K USD	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 250	$ 42	$ 2,500
Bond Carry Cost – Pinion	K USD	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 16	$ 960
Bond Carry Cost – Dark Star	K USD	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 96	$ 16	$ 960
Donations, Dues, PR	K USD	$ 30	$ 30	$ 30	$ 30	$ 30	$ 30	$ 30	$ 30	$ 30	$ 5	$ 300
Fees, Licenses, Misc Taxes, Insurance	K USD	$ 480	$ 480	$ 480	$ 480	$ 480	$ 480	$ 480	$ 480	$ 480	$ 80	$ 4,800
Travel, Lodging, Meals, Entertainment	K USD	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 17	$ 1,000
Telephones, Computers, Cell Phones	K USD	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 75	$ 13	$ 750
Light Vehicle Maintenance, Fuel	K USD	$ 43	$ 43	$ 43	$ 43	$ 43	$ 43	$ 43	$ 36	$ -	$ -	$ 343
Small Tools, Janitorial, Safety Supplies	K USD	$ 85	$ 85	$ 85	$ 85	$ 85	$ 85	$ 85	$ 85	$ 85	$ 14	$ 850
Equipment Rentals	K USD	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 100	$ 17	$ 1,000
Access Road Maintenance	K USD	$ 150	$ 150	$ 150	$ 150	$ 150	$ 150	$ 150	$ 150	$ 150	$ 25	$ 1,500
Office Power	K USD	$ 60	$ 60	$ 60	$ 60	$ 60	$ 60	$ 60	$ 60	$ 60	$ 10	$ 600
Total General G&A Costs	K USD	$ 2,538	$ 2,538	$ 2,538	$ 2,538	$ 2,538	$ 2,538	$ 2,538	$ 2,531	$ 2,495	$ 416	$ 25,318

Total G&A	K USD	$ 3,716	$ 3,704	$ 3,680	$ 3,680	$ 3,680	$ 3,680	$ 3,680	$ 3,651	$ 3,571	$ 595	$ 36,656


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22	ECONOMIC ANALYSIS

The economic analysis in this study includes a preliminary feasibility study-compliant modeling of the annual cash flows based on projected production volume, sales revenue, initial capital, operating cost, and sustaining capital with resulting evaluation of the key economic indicators such as the internal rate of return (IRR), the net present value (NPV), and payback period (time in years to recapture the initial capital investment) for the Project. The sales revenue is based on the production of gold and silver in doré bullion. The estimates of the capital expenditures and site production costs have been developed specifically for this project and have been presented in the previous section of this Technical Report.


22.1	MINING PHYSICALS

The cash-flow model uses the mining and production schedules as discussed in Section 14 and summarized in Table 22-1.

Table 22-1: Yearly Mine & Process Physicals


	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total

Material Mined
Total Ore	K Tonnes	789	3,449	7,775	9,328	7,671	7,614	4,061	5,085	1,571	47,344
	Au g/t	0.62	0.71	1.13	1.28	0.50	0.65	0.64	0.56	0.63	0.82
	Ag g/t	-	-	-	-	0.55	3.81	5.49	4.63	3.25	1.78
	K oz Au	15.75	78.18	282.53	383.63	122.60	158.07	83.55	91.26	31.94	1,248
	K oz Ag	-	-	-	-	135.62	932.67	716.18	756.54	164.38	2,705
Total Waste	K Tonnes	15,259	29,246	17,295	14,246	20,679	21,586	23,862	4,247	870	147,289
Total Mined	K Tonnes	16,048	32,695	25,070	23,574	28,350	29,200	27,923	9,333	2,441	194,633
Strip Ratio	W : O	19.34	8.48	2.22	1.53	2.70	2.83	5.88	0.84	0.55	3.11
ROM Process	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Total ROM	K Tonnes	-	4,013	4,563	4,562	4,575	3,722	841	1,435	486	24,197
	Au g/t	-	0.67	0.51	0.57	0.36	0.39	0.31	0.31	0.34	0.25
	Ag g/t	-	-	-	-	0.24	1.23	3.22	2.94	2.08	0.29
Total Placed	K oz Au	-	86	75	84	54	47	9	14	5	373
Total Recoverable	K oz Au	-	61	52	61	37	31	5	8	3	259
Total Recovered	K oz Au	-	32	42	58	46	40	18	12	6	259
Total Placed	K oz Ag	-	-	-	-	36	148	87	136	32	439
Total Recoverable	K oz Ag	-	-	-	-	10	43	16	20	7	96
Total Recovered	K oz Ag	-	-	-	-	5	25	21	22	15	96
Cumulative Recovery	% Au	-	37%	46%	54%	60%	63%	67%	68%	68%	69%
Cumulative Recovery	% Ag	-	-	-	-	14%	16%	19%	18%	20%	22%
HPGR Process	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Total HPGR Ore	K Tonnes	-	-	2,993	3,650	3,660	3,638	3,650	3,650	1,534	22,775
	Au g/t	-	-	2.10	2.22	0.70	0.91	0.70	0.66	0.66	0.55
	Ag g/t	-	-	-	-	0.82	5.16	6.27	5.48	3.78	1.49
Total Placed	K oz Au	-	-	202	260	83	106	82	77	32	842
Total Recoverable	K oz Au	-	-	168	215	60	75	56	52	23	650
Total Recovered	K oz Au	-	-	124	230	77	72	61	52	32	650
Total Placed	K oz Ag	-	-	-	-	97	604	736	643	186	2,267
Total Recoverable	K oz Ag	-	-	-	-	37	257	318	274	78	965
Total Recovered	K oz Ag	-	-	-	-	27	204	313	287	128	965
Cumulative Recovery	% Au	-	-	61%	77%	79%	77%	77%	76%	77%	77%
Cumulative Recovery	% Ag	-	-	-	-	28%	33%	38%	40%	42%	43%
Sulfide Toll Process	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Total Toll Ore	K Tonnes	-	-	23	91	175	83	-	-	-	372
	Au g/t	-	-	1.72	2.70	2.80	2.76	-	-	-	0.02
	Ag g/t	-	-	-	-	-	-	-	-	-	-
Total Placed	K oz Au	-	-	1	8	16	7	-	-	-	32
Total Recoverable	K oz Au	-	-	1	7	13	6	-	-	-	27
Total Recovered	K oz Au	-	-	0	6	12	9	-	-	-	27
Total Placed	K oz Au	-	-	-	-	-	-	-	-	-	-


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	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total

Total Recoverable	K oz Au	-	-	-	-	-	-	-	-	-	-
Total Recovered	K oz Au	-	-	-	-	-	-	-	-	-	-
Cumulative Recovery	% Au	-	-	27%	72%	76%	85%	85%	85%	85%	85%
Cumulative Recovery	% Ag	-	-	-	-	-	-	-	-	-	-
Total Ore Processed	Units	Pre-Prod	Yr 1	Yr 2	Yr 3	Yr 4	Yr 5	Yr 6	Yr 7	Yr 8	Total
Total Ore Processed	K Tonnes	-	4,013	7,579	8,304	8,410	7,442	4,491	5,085	2,019	47,344
	Au g/t	-	0.67	1.14	1.32	0.56	0.67	0.63	0.56	0.58	0.82
	Ag g/t	-	-	-	-	0.49	3.14	5.70	4.77	3.37	1.78
Total Placed	K oz Au	-	86	278	352	152	160	91	91	38	1,248
Total Recoverable	K oz Au	-	61	222	283	110	112	61	61	26	936
Total Recovered	K oz Au	-	32	166	294	136	121	80	64	38	936
Total Placed	K oz Ag	-	-	-	-	132	752	823	779	219	2,705
Total Recoverable	K oz Ag	-	-	-	-	48	300	334	294	85	1,061
Total Recovered	K oz Ag	-	-	-	-	32	229	334	309	144	1,061
Cumulative Recovery	% Au	-	37%	54%	69%	72%	73%	74%	74%	75%	75%
Cumulative Recovery	% Ag	-	-	-	-	24%	30%	35%	36%	39%	39%


22.3	PROCESS PLANT PRODUCTION STATISTICS

Ore will be processed by cyanide heap leaching as ROM or after HPGR crushing and agglomeration. A small fraction of sulfide ore will be sent out to a third-party operation for toll milling. Overall production over the life of mine is summarized in Table 22-2.

Table 22-2: Life of Mine Process Statistics


Total Ore (kt)	47,344
Gold (g/t)	0.82
Silver (g/t), Pinion only	4.70
Contained Gold (kozs)	1,248
Contained Silver (kozs)	2,705
Gold Recovery %	75.0%
Silver Recovery %	39.2%
Recovered Gold (kozs)	936
Recovered Silver (kozs)	1,061


22.4	SMELTER RETURN FACTORS

No contractual payable metal rates have yet been negotiated with smelters. M3 used typical rates based on industry experience or published guidelines. Payable rates for metals used were 99.5% for gold and 98.0% for silver. A bullion refining, transportation and insurance charge of $5 per troy ounce of gold was applied.


22.5	CAPITAL EXPENDITURE

The capital expenditure schedule for the life of mine is shown in Table 22-3 below.


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Table 22-3: Capital Expenditure Schedule


Capital Expenditure, $000	Initial	Expansion	Sustaining
Capital Expenditure, $000	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7
Pre-stripping	$28,226
Mine	$69,143	$16,485	$3,506	$600	$583	$605	$642	$682
Process	$90,214	$67,171	$6,044	$5,621	$2,098
Owner's Cost	$6,420	$4,649
Total	$194,004	$88,304	$9,551	$6,220	$2,681	$605	$642	$682


22.5	REVENUE

Annual revenue is determined by applying metal prices to the annual payable metal estimated for each operating year. Sales prices have been applied to all life-of-mine production without escalation or hedging. Gold bullion revenue is based on the gross value of the payable metals sold before refining and transportation charges. Gold and silver metal pricing is based on a market study by the Owner as presented in Section 19:

Gold $1,400 per troy ounce
Silver $17.11 per troy ounce


22.6	TOTAL PRODUCTION COST

The total production cost includes mine operations, process plant operations, general administration, reclamation and closure, and government fees. Table 22-4 shows the estimated operating costs by area based on payable metals for the life of mine.

Table 22-4: LOM Operating Costs


LOM Operating Cost ($000)
Mining	$348,505
Process Plant	$156,936
G&A	$33,637
Refining	$4,679
Total Operating Cost	$543,757
Royalty	$16,282
Reclamation/Closure	$49,094
Total Production Cost	$609,133


22.7	DEPRECIATION

The depreciation cost was calculated using a 7-year modified accelerated cost recovery system (MACRS) depreciation method following both initial and sustaining capital.


22.8	ROYALTIES

As discussed in Section 4.2 to this Technical Report, portions of the unpatented and private lands are encumbered with royalties predominantly in the form of standard NSR or GSR and MP royalty agreements, or NPI agreements. GSV intends to buy down certain existing NSR royalties prior to production. As a result of the buy downs, the NSR royalties upon production are assumed to be approximately 1.24% and this has been reflected in the PFS.


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22.9	GOVERNMENT FEES

No government fees have been applied to the financial model.


22.10	INCOME TAX

A net proceeds tax of 5% is applied to revenue minus operating cost and depreciation. Regular corporate tax of 21% is applied to taxable corporation income after adjustments for state tax, if any, and net proceeds tax. No state income tax was applied.


22.11	NET INCOME AFTER TAX

The net income after taxes is $337.1 Million.


22.12	PROJECT FINANCING

It is assumed that the project will be all equity financed.


22.13	ECONOMIC INDICATORS

The economic analyses for the project are summarized in Table 22-5 below.

Table 22-5: Key Economic Results


Indicators ($000)	Before Tax	After Tax
LOM Cash Flow	$409,665	$337,113
NPV @ 5%	$302,081	$241,474
NPV @ 10%	$217,392	$166,153
IRR	32.4%	27.8%
Payback (years)	2.6	2.7


22.14	SENSITIVITY ANALYSIS

Table 22-6 below shows the sensitivity analysis of the key economic indicators (cash flow, NPV, IRR, and payback) to changes in the gold and silver prices.

Table 22-6: Sensitivity Analysis


Financial Indicators	Base +150	Base +$50	Base Case	Base -50	Base -150
Gold Price (per troy oz)	$1,550	$1,450	$1,400	$1,350	$1,250
Silver Price (per troy oz)	$18.94	$17.72	$17.11	$16.50	$15.28
Pre-tax Cash Flow, $M	$549.50	$456.28	$409.66	$363.05	$269.83
Pre-tax Net Present Value (5%) in $M	$417.64	$340.60	$302.08	$263.56	$186.52
Pre-tax Internal Rate of Return (IRR)	40.49%	35.2%	32.4%	29.51%	23.4%
Pre-tax Payback (Years)	2.4	2.5	2.6	2.6	2.8
After-tax Cash Flow, $M	$448.12	$374.18	$337.11	$299.76	$222.85
After-tax Net Present Value (5%) in $M	$333.23	$272.11	$241.47	$210.61	$147.05
After-tax Internal Rate of Return (IRR)	34.70%	30.14%	27.77%	25.30%	19.95%
After-tax Payback (Years)	2.5	2.6	2.7	2.7	2.9


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22.15	DETAILED FINANCIAL MODEL

The detailed financial model, shown in Table 22-7 below, was developed in compliance with the PFS requirement. This model has captured all the parameters of the mine production volume, annual sales revenue, and all the associated costs. This model was used calculate to the economics of the project, as well as for the sensitivity analysis.


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Table 22-7: Detailed Financial Model


GSV South Railroad PFS-Financial Model
M3-PN185074	LOM	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	Year 13	Year 14	Year 15
Mine
Ore (kt)	47,344	789	3,449	7,775	9,328	7,671	7,614	4,061	5,085	1,571	-	-	-	-	-	-	-
Gold (g/t)	0.82	0.62	0.71	1.13	1.28	0.50	0.65	0.64	0.56	0.63	-	-	-	-	-	-	-
Silver (g/t)	1.78	-	-	-	-	0.55	3.81	5.49	4.63	3.25	-	-	-	-	-	-	-
Contained Gold (kozs)	1,248	15.8	78.2	282.5	383.6	122.6	158.1	83.6	91.3	31.9	-	-	-	-	-	-	-
Contained Silver (kozs)	2,705	-	-	-	-	135.62	932.67	716.18	756.54	164.38	-	-	-	-	-	-	-
Waste (kt)	147,289	15,259	29,246	17,295	14,246	20,679	21,586	23,862	4,247	870	-	-	-	-	-	-	-
Total Material Mined	194,633	16,048	32,695	25,070	23,574	28,350	29,200	27,923	9,333	2,441	-	-	-	-	-	-	-
Process Plant
ROM Processing
Dark Star (kt)	19,061		4,013	4,563	4,562	4,109	1,814	-	-	-	-	-	-	-	-	-	-
Gold (g/t)	0.51		0.67	0.51	0.57	0.36	0.37	-	-	-	-	-	-	-	-	-	-
Silver (g/t)			-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Contained Gold (kozs)	314		86	75	84	48	21	-	-	-	-	-	-	-	-	-	-
Contained Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Gold Recovery %	70.7%
Silver Recovery %	0.0%
Recovered Gold (kozs)	222		32	42	58	45	31	11	4	0	-	-	-	-	-	-	-
Recovered Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Pinion (kt)	5,136		-	-	-	466	1,908	841	1,435	486	-	-
Gold (g/t)	0.36		-	-	-	0.37	0.41	0.31	0.31	0.34	-	-
Silver (g/t)			-	-	-	2.39	2.41	3.22	2.94	2.08	-	-	-	-	-	-	-
Contained Gold (kozs)	59		-	-	-	6	25	9	14	5	-	-	-	-	-	-	-
Contained Silver (kozs)	439		-	-	-	36	148	87	136	32	-	-	-	-	-	-	-
Gold Recovery%	62.3%
Silver Recovery%	21.9%
Recovered Gold (kozs)	37		-	-	-	2	10	7	9	6	2	1	-	-	-	-	-
Recovered Silver (kozs)	96		-	-	-	5	25	21	22	15	6	2	-	-	-	-	-
Total ROM (kt)	24,197		4,013	4,563	4,562	4,575	3,722	841	1,435	486	-	-	-	-	-	-	-
Gold (g/t)	0.48		0.67	0.51	0.57	0.36	0.39	0.31	0.31	0.34	-	-	-	-	-	-	-
Silver (g/t)	0.56		-	-	-	0.24	1.23	3.22	2.94	2.08	-	-	-	-	-	-	-
Contained Gold (kozs)	373		86	75	84	54	47	9	14	5	-	-	-	-	-	-	-
Contained Silver (kozs)	439		-	-	-	36	148	87	136	32	-	-	-	-	-	-	-
Gold Recovery %	69.3%
Silver Recovery %	21.9%
Recovered Gold (kozs)	259		32	42	58	46	40	18	12	6	2	1	-	-	-	-	-
Recovered Silver (kozs)	96		-	-	-	5	25	21	22	15	6	2	-	-	-	-	-
HPGR Processing
Dark Star (kt)	10,023		-	2,993	3,650	3,033	347	-	-	-	-	-	-	-	-	-	-
Gold (g/t)	1.67		-	2.10	2.22	0.69	0.78	-	-	-	-	-	-	-	-	-	-
Silver (g/t)			-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Contained Gold (kozs)	537		-	202	260	67	9	-	-	-	-	-	-	-	-	-	-
Contained Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Gold Recovery %	81.6%
Silver Recovery %	0.0%
Recovered Gold (kozs)	439		-	124	230	69	15	0	-	-	-	-	-	-	-	-	-
Recovered Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Pinion(kt)	12,752		-	-	-	627	3,291	3,650	3,650	1,534	-	-	-	-	-	-	-
Gold (g/t)	0.74		-	-	-	0.78	0.92	0.70	0.66	0.66	-	-	-	-	-	-
Silver (g/t)	5.53		-	-	-	4.79	5.71	6.27	5.48	3.78	-	-	-	-	-	-	-
Contained Gold (kozs)	305		-	-	-	16	97	82	77	32	-	-	-	-	-	-	-
Contained Silve r(kozs)	2,267		-	-	-	97	604	736	643	186	-	-	-	-	-	-	-
Gold Recovery %	69.3%
Silver Recovery %	42.6%
Recovered Gold (kozs)	211		-	-	-	8	56	61	52	32	1	-	-	-	-	-	-
Recovered Silver (kozs)	965		-	-	-	27	204	313	287	128	6	-	-	-	-	-	-
Total HPGR (kt)	22,775		-	2,993	3,650	3,660	3,638	3,650	3,650	1,534	-	-	-	-	-	-	-
Gold (g/t)	1.15		-	2.10	2.22	0.70	0.91	0.70	0.66	0.66	-	-	-	-	-	-	-
Silver (g/t)	3.10		-	-	-	0.82	5.16	6.27	5.48	3.78	-	-	-	-	-	-	-
Contained Gold (kozs)	842		-	202	260	83	106	82	77	32	-	-	-	-	-	-	-


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GSV South Railroad PFS-Financial Model
M3-PN185074	LOM	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	Year 13	Year 14	Year 15
Contained Silver (kozs)	2,267		-	-	-	97	604	736	643	186	-	-	-	-	-	-	-
Gold Recovery %	77.1%
Silver Recovery %	42.6%
Recovered Gold (kozs)	650		-	124	230	77	72	61	52	32	1	-	-	-	-	-	-
Recovered Silver (kozs)	965		-	-	-	27	204	313	287	128	6	-	-	-	-	-	-
Toll Processing
Dark Star (kt)	372		-	23	91	175	83	-	-	-	-	-	-	-	-	-	-
Gold (g/t)	2.70		-	1.72	2.70	2.80	2.76	-	-	-	-	-	-	-	-	-	-
Silver (g/t)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Contained Gold (kozs)	32		-	1	8	16	7	-	-	-	-	-	-	-	-	-	-
Contained Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Gold Recovery %	85.0%
Silver Recovery %	0.0%
Recovered Gold (kozs)	27		-	0	6	12	9	-	-	-	-	-	-	-	-	-	-
Recovered Silver (kozs)	-		-	-	-	-	-	-	-	-	-	-	-	-	-	-	-
Tota lProcessing
Total Ore (kt)	47,344		4,013	7,579	8,304	8,410	7,442	4,491	5,085	2,019	-	-	-	-	-	-	-
Gold (g/t)	0.82		0.67	1.14	1.32	0.56	0.67	0.63	0.56	0.58	-	-	-	-	-	-	-
Silver (g/t)	1.78		-	-	-	0.49	3.14	5.70	4.77	3.37	-	-	-	-	-	-	-
Contained Gold (kozs)	1,248		86	278	352	152	160	91	91	38	-	-	-	-	-	-	-
Contained Silver (kozs)	2,705		-	-	-	132	752	823	779	219	-	-	-	-	-	-	-
Gold Recovery %	75.0%		36.6%	59.8%	83.6%	89.4%	75.5%	87.8%	70.6%	101.7%	0.0%	0.0%	0.0%	0.0%	0.0%	0.0%	0.0%
Silver Recovery %	39.2%		0.0%	0.0%	0.0%	24.3%	30.5%	40.5%	39.7%	65.6%	0.0%	0.0%	0.0%	0.0%	0.0%	0.0%	0.0%
Recovered Gold (kozs)	936		32	166	294	136	121	80	64	38	4	1	-	-	-	-	-
Recovered Silver (kozs)	1,061		-	-	-	32	229	334	309	144	12	2	-	-	-	-	-
Payable Metals
Gold (kozs)	931		31	165	293	135	120	79	64	38	4	1	-	-	-	-	-
Silver (kozs)	1,040		-	-	-	31	224	327	303	141	11	2	-	-	-	-	-
Metal Prices
Gold ($/oz)	$1,400.00		$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$1,400.00	$0.00	$0.00	$0.00	$0.00	$0.00
Silver ($/oz)	$17.11		$17.11	$17.11	$17.11	$17.11	$17.11	$17.11	$17.11	$17.11	$17.11	$17.11	$0.00	$0.00	$0.00	$0.00	$0.00
Revenues ($000)
Gold	$1,303,691		$43,880	$231,385	$409,983	$189,341	$168,111	$111,150	$89,830	$53,376	$5,455	$1,180	$0	$0	$0	$0	$0
Silver	$17,796		$0	$0	$0	$539	$3,837	$5,595	$5,186	$2,409	$194	$35	$0	$0	$0	$0	$0
Total Revenues	$1,321,487		$43,880	$231,385	$409,983	$189,880	$171,948	$116,745	$95,017	$55,784	$5,649	$1,216	$0	$0	$0	$0	$0
OperatingCost($000)
Mining	$348,505		$54,111	$54,301	$52,905	$53,790	$52,132	$46,731	$26,548	$7,987	$0	$0	$0	$0	$0	$0	$0
Process Plant	$156,936		$7,896	$23,845	$26,652	$25,357	$22,395	$18,714	$19,270	$11,137	$1,114	$557	$0	$0	$0	$0	$0
G&A	$33,637		$3,716	$3,704	$3,680	$3,680	$3,680	$3,680	$3,680	$3,651	$3,571	$595	$0	$0	$0	$0	$0
Refining	$4,679		$158	$831	$1,472	$680	$603	$399	$322	$192	$20	$4	$0	$0	$0	$0	$0
Total Operating Cost	$543,757		$65,880	$82,681	$84,709	$83,506	$78,811	$69,523	$49,821	$22,966	$4,704	$1,156	$0	$0	$0	$0	$0
Royalty	$16,282		$541	$2,851	$5,051	$2,339	$2,119	$1,439	$1,171	$687	$70	$15	$0	$0	$0	$0	$0
Salvage Value	$0		$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Reclamation/Closure	$49,094		$2,183	$2,183	$3,876	$3,876	$3,422	$3,422	$1,729	$1,729	$0	$26,674	$0	$0	$0	$0	$0
Total Production Cost	$609,133		$68,604	$87,714	$93,636	$89,722	$84,351	$74,384	$52,721	$25,382	$4,774	$27,845	$0	$0	$0	$0	$0
Operating Income	$712,353		-$24,725	$143,671	$316,347	$100,158	$87,597	$42,361	$42,296	$30,402	$875	-$26,629	$0	$0	$0	$0	$0
Depreciation($000)
Initial Capital	$194,004		$27,723	$47,511	$33,931	$24,231	$17,325	$17,305	$17,325	$8,653	$0	$0	$0	$0	$0	$0	$0
Sustaining Capital	$108,685		$12,619	$22,991	$18,672	$14,606	$10,909	$10,215	$9,988	$5,940	$1,474	$713	$292	$145	$90	$30	$0
Total Depreciation	$302,688		$40,342	$70,502	$52,604	$38,837	$28,234	$27,520	$27,313	$14,593	$1,474	$713	$292	$145	$90	$30	$0
Net Income after Depreciation	$409,665		-$65,066	$73,169	$263,743	$61,321	$59,364	$14,841	$14,983	$15,809	-$599	-$27,343	-$292	-$145	-$90	-$30	$0
Taxes ($000)
Net Proceeds Tax	$26,956	$0	$0	$3,910	$13,634	$3,377	$3,245	$985	$894	$911	$0	$0	$0	$0	$0	$0	$0
Income Tax	$45,595	$0	$0	$0	$28,834	$6,187	$6,368	$1,351	$1,385	$1,469	$0	$0	$0	$0	$0	$0	$0
Total Taxes	$72,552		$0	$3,910	$42,468	$9,564	$9,614	$2,336	$2,280	$2,380	$0	$0	$0	$0	$0	$0	$0
Net Income after Taxes ($000)	$337,113		-$65,066	$69,259	$221,275	$51,757	$49,750	$12,504	$12,703	$13,429	-$599	-$27,343	-$292	-$145	-$90	-$30	$0
Cash Flow ($000)
Net Income before Taxes	$409,665	$0	-$65,066	$73,169	$263,743	$61,321	$59,364	$14,841	$14,983	$15,809	-$599	-$27,343	-$292	-$145	-$90	-$30	$0
Add back Depreciation	$302,688	$0	$40,342	$70,502	$52,604	$38,837	$28,234	$27,520	$27,313	$14,593	$1,474	$713	$292	$145	$90	$30	$0
Operating Cash Flow	$712,353	$0	-$24,725	$143,671	$316,347	$100,158	$87,597	$42,361	$42,296	$30,402	$875	-$26,629	$0	$0	$0	$0	$0
Working Capital ($000)


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GSV South Railroad PFS-Financial Model
M3-PN185074	LOM	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	Year 13	Year 14	Year 15
Accounts Receivable	$0		-$1,202	-$5,137	-$4,893	$6,030	$491	$1,512	$595	$1,075	$1,374	$121	$33	$0	$0	$0	$0
Accounts Payable	$0		$5,415	$1,381	$167	-$99	-$386	-$763	-$1,619	-$2,207	-$1,501	-$292	-$95	$0	$0	$0	$0
Inventory (parts)	$0	$0
Total Working Capital	$0	$0	$4,213	-$3,756	-$4,726	$5,931	$105	$749	-$1,024	-$1,132	-$127	-$170	-$62	$0	$0	$0	$0
Initial Capital Expenditures ($000)
Pre-stripping	$28,226	$28,226	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Mining	$69,143	$69,143	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Process	$90,214	$90,214	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Owner's Cost	$6,420	$6,420	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Expansion Capital Expenditures ($000)
Mining	$16,485		$16,485	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Process	$67,171		$67,171	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Owner'sCost	$4,649		$4,649	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Sustaining Capital Expenditures ($000)
Mining	$6,617		$0	$3,506	$600	$583	$605	$642	$682	$0	$0	$0	$0	$0	$0	$0	$0
Process	$13,763		$0	$6,044	$5,621	$2,098	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Owner's Cost	$0		$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Total Capital	$302,688	$194,004	$88,304	$9,551	$6,220	$2,681	$605	$642	$682	$0	$0	$0	$0	$0	$0	$0	$0
Cash Flow before Taxes ($000)	$409,665	-$194,004	-$108,816	$130,364	$305,400	$103,409	$87,098	$42,468	$40,590	$29,269	$748	-$26,800	-$62	$0	$0	$0	$0
Cumulative Cash Flow before Taxes ($000)		-$194,004	-$302,820	-$172,456	$132,944	$236,353	$323,451	$365,919	$406,509	$435,779	$436,526	$409,727	$409,665	$409,665	$409,665	$409,665	$409,665
Taxes	$72,552	$0	$0	$3,910	$42,468	$9,564	$9,614	$2,336	$2,280	$2,380	$0	$0	$0	$0	$0	$0	$0
Cash Flow after Taxes ($000)	$337,113	-$194,004	-$108,816	$126,453	$262,932	$93,845	$77,484	$40,132	$38,311	$26,890	$748	-$26,800	-$62	$0	$0	$0	$0
Cumulative Cash Flow after Taxes ($000)		-$194,004	-$302,820	-$176,366	$86,566	$180,411	$257,895	$298,027	$336,337	$363,227	$363,975	$337,175	$337,113	$337,113	$337,113	$337,113	$337,113
NPV @0%	$409,665
NPV @5%	$302,081
NPV @10%	$217,392
IRR	32.4%
Payback (years)	2.6		1.0	1.0	0.6	-	-	-	-	-	-	-	-	-	-	-	-
NPV @0%	$337,113
NPV @5%	$241,474
NPV @10%	$166,153
IRR	27.8%
Payback (years)	2.7		1.0	1.0	0.7	-	-	-	-	-	-	-	-	-	-	-	-
Payable Au (kozs)	931	-	31	165	293	135	120	79	64	38	4	1	-	-	-	-	-
Mining	$348,505	$0	$54,111	$54,301	$52,905	$53,790	$52,132	$46,731	$26,548	$7,987	$0	$0	$0	$0	$0	$0	$0
Process Plant	$156,936	$0	$7,896	$23,845	$26,652	$25,357	$22,395	$18,714	$19,270	$11,137	$1,114	$557	$0	$0	$0	$0	$0
G&A	$33,637	$0	$3,716	$3,704	$3,680	$3,680	$3,680	$3,680	$3,680	$3,651	$3,571	$595	$0	$0	$0	$0	$0
Refining	$4,679	$0	$158	$831	$1,472	$680	$603	$399	$322	$192	$20	$4	$0	$0	$0	$0	$0
Royalty	$16,282	$0	$541	$2,851	$5,051	$2,339	$2,119	$1,439	$1,171	$687	$70	$15	$0	$0	$0	$0	$0
Cash Cost before By-Product Credit	$560,039	$0	$66,421	$85,531	$89,760	$85,846	$80,929	$70,962	$50,992	$23,654	$4,774	$1,171	$0	$0	$0	$0	$0
$/Au oz	$601	$0	$2,119	$518	$307	$635	$674	$894	$795	$620	$1,225	$1,389	$0	$0	$0	$0	$0
Silver Credit	$17,796	$0	$0	$0	$0	$539	$3,837	$5,595	$5,186	$2,409	$194	$35	$0	$0	$0	$0	$0
Cash Cost after By-Product Credit	$542,244	$0	$66,421	$85,531	$89,760	$85,307	$77,092	$65,367	$45,806	$21,245	$4,579	$1,136	$0	$0	$0	$0	$0
$/oz Au	$582	$0	$2,119	$518	$307	$631	$642	$823	$714	$557	$1,175	$1,347	$0	$0	$0	$0	$0
Sustaining Capital Expenditures
Mining	$6,617	$0	$0	$3,506	$600	$583	$605	$642	$682	$0	$0	$0	$0	$0	$0	$0	$0
Process	$13,763	$0	$0	$6,044	$5,621	$2,098	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Owner's Cost	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Salvage Value	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0	$0
Reclamation/Closure	$22,420	$0	$2,183	$2,183	$3,876	$3,876	$3,422	$3,422	$1,729	$1,729	$0	$0	$0	$0	$0	$0	$0
Net Proceeds Tax	$26,956	$0	$0	$3,910	$13,634	$3,377	$3,245	$985	$894	$911	$0	$0	$0	$0	$0	$0	$0
AISC	$612,000	$0	$68,604	$101,175	$113,490	$95,240	$84,364	$70,416	$49,110	$23,885	$4,579	$1,136	$0	$0	$0	$0	$0
$/oz Au	$657	$0	$2,189	$612	$388	$704	$703	$887	$765	$626	$1,175	$1,347	$0	$0	$0	$0	$0


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23	ADJACENT PROPERTIES

The Railroad- Pinion property is situated along the southeastern portion of the Carlin Gold Trend. The Rain Mining District, which is largely controlled by Nevada Gold Mines, is located 2 to 3 km (1.2 to 2 miles) north of the Railroad-Pinion property. The Rain District has been an active exploration and mining area for several decades and is the location for current and past mining activities by Nevada Gold Mines and Newmont Mining at the Rain open pit and underground mine and Emigrant open pit mine. To the south of the Railroad-Pinion property, several exploration areas have received sporadic exploration over the past three to four decades including Pony Creek. Adjacent properties with bearing or influence on the Railroad-Pinion property are described below. The authors of this Technical Report have not visited or worked at any of these projects and where references are made to past production and/or historic or current mineral resources the authors have not verified the information.


23.1	RAIN

Rain is a Carlin-style, sedimentary rock-hosted gold deposit that is located approximately four miles (seven kilometers) north of Gold Standard’s North Bullion mineral resource. Newmont operated the Rain open pit mine, the Rain underground mine and the SMZ open pit mine from 1988 to 2000; and produced approximately 1.24 million ounces (Ressel et al., 2015. Longo et al. (2002) summarized a number of mineral resources for the three deposits as follows: Rain open pit 15.5 million tons (14.1 million tonnes) at 0.066 opt (2.3 g/t) Au for a total of 1,017,300 ounces of gold; Rain Underground 1.154 million tons (1.04 million tonnes) at 0.23 opt (7.9 g/t) Au for a total of 265,000 ounces of gold and the SMZ open pit 1.5 million tons (1.4 million tonnes) at 0.019 opt (0.65 g/t) Au for a total of 30,000 ounces of gold. The mineral resources pre-date NI 43-101 and little or no detailed information such as potential mineral resource category or number of drill holes is presented for the estimates or how the mineral resources were arrived at. Therefore, the estimates are considered historic in nature and should not be relied upon. The authors of this Technical Report have been unable to verify this and this information is not necessarily indicative of the mineralization of the Railroad-Pinion property.

Along strike to the northwest of the Rain Project and likely on the same structure are the Saddle and Tess gold deposits. The mineralized zones are roughly 3.5 km (2 miles) north of the Railroad-Pinion Project and 10 km (6 miles) northwest of the North Bullion mineral resource. Longo et al. (2002) states that Newmont identified a primarily underground high sulphide mineral resource of 1.37 million tons (1.23 million tonnes) at 0.572 opt (19.6 g/t) Au for a total of 782,000 ounces of gold at Saddle and 3.99 million tons (3.59 million tonnes) at 0.37 opt (12.7 g/t) Au for a total of 1,475,000 ounces of gold at Tess. The project was part of the Newmont South Area of operations but has recently been consolidated under the Newmont/Barrick Joint Venture (Nevada Gold Mines). No mining has been conducted at the two deposits. The mineral resources pre-date NI 43-101 and little or no detailed information such as potential mineral resource category or number of drill holes etc. is presented for the estimates or how the mineral resources were arrived at, therefore, the estimates are considered historic in nature and should not be relied upon. The authors of this Technical Report have not visited the Rain property, nor have they verified the historic estimates provided by Longo et al. (2002).

The Rain trend of mineralization is characterized by disseminated gold mineralization hosted in dominantly oxidized, silicified, dolomitized, and barite rich collapse breccia with rare sulfides, developed along the Webb Formation mudstone/Devils Gate Formation calcarenite contact and along the Rain Fault. Ore-controlling features at Rain include the west-northwest striking Rain fault, the Webb/Devils Gate contact, collapse breccia and northeast striking cross faults. Shallow oxide zones at the Rain deposit give way along the west-northwest trend to deeper sulphide- and carbon-bearing zones of substantial size and grade at the Saddle and Tess deposits.


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23.2	EMIGRANT

Emigrant is a Carlin-style, sedimentary rock-hosted gold deposit that is located approximately four miles (seven kilometers) north-northeast of the North Bullion mineral deposits. Until recently Newmont/Nevada Gold Mines has been mining the deposit through open pit methods and processing the ore at an onsite, run of mine heap leach operation with some crushing. The operation currently appears to be shut down. Disseminated gold mineralization is hosted in oxidized, silicified, dolomitized, and barite rich collapse breccia developed within the Webb Formation mudstone. Important ore-controlling features at Emigrant include the north-south-striking Emigrant Fault, collapse breccia and the Northeast Fault.

Open pit, oxide mineral resource and mineral reserve calculations for Newmont’s Carlin Trend operations are typically commingled into a single heading of “Carlin open pits, Nevada” category. In 2003, mineral reserves at Emigrant were published at 1,220,000 ounces (Newmont, 2012). No details were provided by Newmont as to the quality of the mineral reserves. The mine is expected to produce roughly 800,000 ounces of gold over a ten plus year mine life and has recently commenced production (Harding, 2012). The authors of this Technical Report have been unable to verify this and this information is not necessarily indicative of the mineralization on the Railroad-Pinion property.


23.3	PONY CREEK PROPERTY

Pony Creek is located approximately six miles (10 kilometers) south of the Pinion deposit. Gold mineralization is hosted in north to northeast-trending shears in rhyolite intrusive and Mississippian to Permian age sediments proximal to the intrusive (Russell, 2006).


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24	OTHER RELEVANT DATA AND INFORMATION

There are no additional data for the Railroad Pinion property beyond that discussed in the preceding sections..


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25	INTERPRETATION AND CONCLUSIONS

The authors of this Technical Report believe that South Railroad is a project of merit and warrants advancing the study to feasibility level.

The authors have reviewed the project data, including the drill-hole database and available metallurgical information, and have visited the project site. The authors believe that the data provided by Gold Standard, as well as the geological interpretations Gold Standard has derived from the data, are generally an accurate and reasonable representation of the Railroad-Pinion property. Based on the positive results of this PFS, the project should continue on a path to a production decision.

Results of historical metallurgical tests and those commissioned by Gold Standard indicate there are multiple metallurgical material types within the Pinion and Dark Star gold deposits. Due to the multiple material types and the dependence of gold recoveries on head grades, 54 different gold recovery equations are used to project the processing and gold and silver production estimates presented in this Technical Report.

The process selected for recovery of gold and silver from the Pinion and Dark Star mineralized material is a conventional heap-leach recovery circuit. The material will be mined by standard open-pit mining methods and trucked from each deposit to a centralized area of heap-leach pads and processing facilities. Based on material grades within the Pinion and Dark Star mineable mineral resources, approximately 24.3 million tonnes will be selected for ROM processing and 28.5 million tonnes will be selected for HPGR crushing. The nominal processing rate through the crushing circuit will be 10,000 tonnes per day, and the design ROM processing rate (for the leaching and adsorption design basis) is 12,500 tonnes per day.

Over the LOM of 8 years, the PFS indicates gold production of about 116,986 ounces per year, with peak production in Year 3 of 294,316 ounces of gold. Cash costs are estimated to be $582 per ounce of gold and total costs are estimated to be $657 per ounce of gold. The resulting after-tax cash flow is $337.1 million, for an after-tax NPV (5%) of $241.5 million and an estimated IRR of 27.8% and payback period of 2.7 years.


25.1	PROJECT RISKS


	1.	Initial power supply baseline costing is from NVE. The PFS analysis assumes that power distribution will come from the Carlin substation via a new line constructed from Carlin, NV, along highway 278 for approximately 25 miles, and then follow the access road to the substation located on site. NV is currently working on the engineering and power distribution system for the project, so current costs in the PFS financials are from previous projects, and not actuals to be supplied by NV in the future.

	2.	RIB location yet to be determined for water disposal from dewatering activities. The PFS assumes that water disposal from dewatering activities will be disposed using RIBs, which is typical methodology at Nevada mining sites. At the writing of this report, a RIB location has not been identified and tested for infiltration rates required by the State of Nevada. That work is ongoing. Alternatives to RIB disposal are injection wells and release to surface waters.


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	3.	Water treatment is not being considered in the PFS with the assumption that water can be discharged to a RIB. Should water treatment be required, 3.5 MW will have to be added to the connected load, which increases the demand load by 2.8 MW.

	4.	Facility geotechnical work need completion for major foundations. At the date of this Technical Report, surface facility geotechnical work has not been completed for major foundations. Costs in the PFS assume no major engineering will be required. Surface geotechnical work is currently ongoing for proposed surface dumps and heap leach facilities with Golder.


25.2	PROJECT OPPORTUNITIES


	1.	Project economics currently apply capital in the form of purchasing mining equipment in Year 0, with expected life of mine performance. Project should evaluate opportunity for 7-8-year capital leasing of major mine equipment to reduce initial capital and maintain lower cost owner operated fleet for expansion opportunities beyond.

	2.	ADR and HPGR plants were single-source quoted, full installation costed for the project. Both plants should be competitively bid to reduce initial and expansion capital requirements for the project.

	3.	Project mining fleet capital cost is single source quoted. Major pieces of equipment should be competitively bid and/or used/leased equipment considered to reduce initial capital for the project.

	4.	Pinion deposit currently contains a large increase in gold ounces within the quoted mineral resource ($1,500 Au) compared to the mineral reserve ($1.275 Au). Drilling, metallurgy, and hydrology should be evaluated to determine if additional work can bring these mineral resources into the mine plan.

	5.	Oxide mineral resources ($1.500 Au) are currently drilled to Inferred (Jasperoid Wash and Pod) and Indicated (Pod) status. These mineral resources are not included in the current mine plan and should be evaluated for impacts to the project net present value, mine life, and work required to bring forward.

	6.	Complete trade off studies to optimize current mine plan for project.


	a.	Evaluate feed rate to HPGR from current proposed 10,000 tpd to optimize throughput, capital spend, and processing operating costs.

	b.	Evaluate lower capital cost options for the project, including standard crushing (compared to proposed HPGR), Run of mine without crushing.

	c.	Optimize timing for installation of HPGR circuit.


	7.	Pit designs should undergo various iterations to


	a.	Determine ultimate pit while minimizing strip and capitalized pre-strip. Current Dark Star North Pit contains a west extension ramp system outside the LG cones, to accommodate environmental concerns. Opportunities may exist for alternative pit designs to reduce stripping requirements, especially in year 0.

	b.	Evaluate opportunity for utilizing surface exposed ore at Dark Star Main and Pinion for placement as crushed over liner.

	c.	Investigate opportunities to shorten waste hauls by backfilling existing pits, especially at Pinion, to reduce operating costs.


25.3	EXPLORATION AND MINERAL RESOURCE EXPANSION

Dark Star remains open to exploration and expansion in all directions and requires additional drilling to define the economic edges to mineralization. The Pinion deposit contains a significant mineral resource of Measured and


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Indicated ounces (180,000 oz Au) and Inferred (224,000 oz) that is exclusive of the current mineral reserves. This material floats cones at the $1,400 gold price. The potential expansion of Pinion to the southwest with the known mineral resources should be evaluated in a rising gold price environment. Jasperoid Wash and oxide mineralization at North Bullion contain oxide mineral resources at $1,500 gold price. Current classification of mineral resources as Inferred at Jasperoid Wash and portions of North Bullion prevented this material from consideration within the PFS and mineral reserves. Additionally, Jasperoid Wash is open along strike and mineral resources can potentially be increased with expansion drilling. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

Gold Standard’s Railroad-Pinion property is centered on another window of the Carlin trend. The property has all the promising geologic characteristics of other productive districts of the Carlin trend, including carbonate host rocks, older thrust faults and folds, younger extensional faults and an Eocene (Carlin age) magmato-thermal event. Deposits at Railroad-Pinion are hosted both in collapse breccia developed along the Devonian Devils Gate limestone/Mississippian Tripon Pass micrite contact and within highly permeable Pennsylvanian-Permian carbonate units. These units are common hosts for Carlin-type gold deposits throughout north-central Nevada. The structural setting with north-, northeast- and northwest-striking Tertiary extensional faults overprinted on earlier compressional structures is a classic Carlin framework. There are numerous un-drilled targets along prospective structural corridors.


26	RECOMMENDATIONS

To advance the study to feasibility level, the authors recommend a study program and level of expenditures outlined below, focused on the gold deposits in the South Railroad portion of the property.

Based on the results of the preliminary feasibility work the authors believe that the Railroad-Pinion property is a project of merit and warrants the proposed program and level of expenditures outlined below, focused on the gold deposits in the South Railroad portion of the property.

The total cost of recommendations is expected to reach $21 million for a multi-faceted program including exploration, permitting, development, metallurgical testing and engineering. The subsection describe the recommended program.


26.1	EXPLORATION

The Railroad-Pinion property is large, and merits continued exploration outside of the immediate deposit areas. Recommended exploration includes mapping and sampling within under-explored portions of the property. Exploration would include mapping, sampling, and 2D seismic to help define faults, and exploration/reconnaissance initial drilling of 10,000 m in about 20 drill holes ($1,700,000).


26.2	INFILL DRILLING

Infill drilling is justified to potentially upgrade the Inferred mineral resources within the Pinion mineral resource to Measured or Indicated mineral resource classification. This mineral resource is currently within $1400 Au ounce shells based on LG Cones. This should be further evaluated at current gold prices >$1,450/ounce. This drilling will total 16,150 m and can be done with RC and core methods at an estimated cost of $3.0 million.

Infill drilling (10,800 m) and metallurgy at Jasperoid Wash is warranted because of the potential to contribute to the property mineral resources and mineral reserves. This can be done with RC methods at an estimated cost of $2.5 million.

Oxide mineralization and mineral resources at POD and Sweet Hollow are generally known from historical drilling.

These should be evaluated metallurgically with core drilling and metallurgical testing. Estimated costs are $1.5 million.


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26.3	EXPLORATION AND EXPANSION DRILLING

Dark Star mineralization is not closed off around/near the current pit designs, and additional exploration and expansion drilling is merited to determine the ultimate footprint of this deposit. Additional core and RC drilling is recommended. The total for this task is $1.5 million.


26.4	CONDEMNATION DRILLING

Condemnation drilling of 3,300 m in ten drill holes is recommended within and near the footprints of the planned facilities. The estimated total cost is $500,000.


26.5	METALLURGICAL TEST WORK

Metallurgical work requirements for advancing the Pinion and Dark Star deposits to feasibility study level are estimated at $2.05 million, which includes additional metallurgical test work, engineering and trade-off studies, material-fracture characterization, metallurgical consulting and a 10% contingency as described below.

HPGR Design and Costing: Metallurgical testwork has demonstrated that HPGR crushing in the tertiary circuit provides the best recovery option for the Dark Star and Pinion material. Additional testwork to finalize HPGR design, evaluate wear and materials use, and ultimately operating costs should be conducted through Thyssen-Krupp and KCA. Cost for this include

o PQ Core drilling to provide adequate material for test work: 17 holes, 1,137m, $600,000
o Material preparation and analysis by KCA $200,000
o Material testing by Thyssen Krupp $500,000

Material Flow Characterization: To facilitate design of primary and secondary crusher, conveyor drop boxes, chutes, and stockpile reclaim equipment selections. $100,000
Metallurgical testwork on additional mineral resources identified at Pinion, Jasperoid Wash, and POD. Mineral resources have an opportunity to impact future mine plans. Preliminary metallurgical testwork to establish recoveries and material characterization should be undertaken during next stages of testing. $500,000
Contingency: $146K; 10% on vendor testing/characterization, data gap filling identified by the preliminary feasibility engineering contractors and any missing scope of work items.


26.6	PERMITTING AND BASELINE STUDIES

It is recommended that Gold Standard initiate permitting and NEPA activities in support of open-pit mining at Dark Star and Pinion. The estimated cost is estimated at $2.0 million, including a EIS contractor.


26.7	ENGINEERING STUDIES AT FEASIBILITY LEVEL

The current PFS considers the Pinion and Dark Star mineral reserves as potentially minable. Additional work is required to optimize the mineral resources and mineral reserves, including trade off studies, engineering to feasibility level, and geotechnical work in support of facilities. It is expected that the work to progress to a feasibility level will be approximately $3.0 million.


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26.8	HYDROLOGY

Additional testing and monitoring wells maybe required by the state and federal agencies as part of the EIS process. This is anticipated in the form of drilling and installation of additional piezometer wells, water sampling and analysis, and pump testing. Projected costs are $2.0 million.


26.9	GEOTECHNICAL SURVEY

Geotechnical studies are recommended for assessing pit designs at a feasibility level. All-in costs, including drilling, field work, and reporting are estimated to be $1.1 million.

Geotechnical study of the mill site is also recommended to determine geotechnical parameters to be used for structural design. Estimated cost is $110,000.


26.10	TOTAL COST OF RECOMMENDED STUDY PROGRAM

Table 26-1 is a summary of the costs of the recommended work to advance the study to feasibility level.

Table 26-1: Cost Estimate for the Recommended Study Program


Exploration	Cost	Sub-total
Mapping and Sampling	$50,000
Seismic	$500,000
Drilling	$1,150,000	$1,700,000

Infill and Expansion Drilling
Pinion mineral resource Infill	$3,000,000
Jasperoid Wash Infill	$2,500,000
POD Metallurgy and Infill Drilling	$1,500,000
Dark Star Development/Step-out Drilling	$1,500,000
Facility Condemnation Drilling	$500,000	$9,000,000

Metallurgical Test work
HPGR Design Work (Thyssen-Krupp	$1,300,000
Material Flow	$100,000
Pinion, POD, JW Metallurgy (columns and BR)	$500,000
Contingency (10%)	$146,000	$2,046,000

Permitting and Baseline Studies		$2,000,000

Engineering and Feasibility Study		$3,000,000

Hydrology		$2,000,000

Geotechnical		$1,210,000
Grand Total (rounded to x,000s)		$20,956,000


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27	REFERENCES

American Selco Incorporated, 1970, Review of Field Investigations, 1968, and Proposed Exploration Program for 1969 in the Railroad Mining District Elko County, Nevada: document prepared for the Directors of Aladdin Sweepstake Consolidated Mining Company.

AMTEL, 2018 (September), Abbreviated Gold Deportment Analysis of Milled Column Residues: unpublished report to Gold Standard Ventures Inc., AMTEL Ltd. Report 18/46, no author, 13p.

Arthur, B., 2013, Roaster Tests on Rail Road Ore 2013-OP-003, Newmont Mining Corporation memo.

Barr Engineering, 2019, Fragmentation Study Dark Star & Pinion Deposits, Design Basis, Prepared for Gold Standard Ventures.

Bartels, E., 1999, Railroad Project, Pod Area Sectional Resource Calculation: internal company memorandum prepared for Kinross Gold U.S.A, Inc.

Bettles, K., 2002, Exploration and Geology, 1962-2002, at the Goldstrike Property: in Thompson, T.B., Teal, L., and Meeuwig, R., eds., Gold Deposits of the Carlin Trend: Nevada Bureau of Mines and Geology Bulletin 111, p.54-75.

Bharti Engineering, 1996, Preliminary Assessment of the South Bullion Deposit: confidential internal report, 5p.

Bloomberg Consensus Precious Metals Pricing, 2019 (September 30), www.bloomberg.com.

Calloway, V., 1992, Exploration Summary, Results and Recommendations for the Cord Ranch Lease, Elko County, Nevada: Crown Resources internal report, 78p.

Cline, J.L., 2004. Controversies on the origin of world-class gold deposit, Part 1: Carlin-type gold deposits in Nevada. Introduction to Carlin-type deposits. SEG Newsletter no. 59, pp. 1,11-12.

Cline, J.L., 2005. Carlin-type gold deposits in Nevada: critical geologic characteristics and viable models: Economic Geology 100^thAnniversary Volume, pp. 451-484.

Clode, C.H., Grusing, S.R., Johnston, I.M., and Heitt, D.G., 2002, Geology of the Deep Star Gold Deposit: in Thompson, T.B., Teal, L., and Meeuwig, R., eds., Gold Deposits of the Carlin Trend: Nevada Bureau of Mines and Geology Bulletin 111, p. 76-90.

Crafford, A.E.J., 2007, Geologic Map of Nevada: U.S. Geological Survey Data Series 249.

DeMatties, T.A., 2003 (January), NI 43-101 Technical Report: An Evaluation of the Pinion Gold Property, Elko County, Nevada, USA: prepared for Royal Standard Minerals Inc.

Dufresne, M.B. and Koehler, S.R., 2016 (March), Technical Report on the Railroad – Pinion Project, Elko County, Nevada USA: unpublished technical report (NI43-101 compliant) prepared for Gold Standard Ventures Corp., 471 p.

Dufresne, M.B., and Nicholls, S.J., 2016 (April), Technical Report Resource Estimate Update Pinion Project Elko County, Nevada USA: NI 43-101 technical report prepared by APEX Geoscience Ltd., 220 p.


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FORM 43-101F1 TECHNICAL REPORT

Dufresne, M.B., and Nicholls, S.J., 2017a (August), Technical Report Resource Estimate Update Dark Star Project, Elko County, Nevada USA: NI 43-101 technical report prepared by APEX Geoscience Ltd., 220 p.

Dufresne, M.B., and Nicholls, S.J., 2017b (November), Technical Report Maiden Resource Estimate North Bullion and Railroad Project, Elko County, Nevada, USA: unpublished technical report (NI43-101 compliant) prepared for Gold Standard Ventures Corp., 258 p.

Dufresne, M.B., and Nicholls, S.J., 2018 (February), Technical Report Maiden Resource Estimate North Bullion and Railroad Project, Elko County, Nevada, USA Amended and Restated: unpublished technical report (NI43-101 compliant) prepared for Gold Standard Ventures Corp., 292 p.

Dufresne, M.B., Nicholls, S.J., and Turner, A, 2014 (October), Technical Report Maiden Resource Estimate Pinion Project, Elko County, Nevada USA: unpublished technical report (NI43-101 compliant) prepared for Gold Standard Ventures Corp., 141 p.

Dufresne, M.B., Nicholls, S.J., and Turner, A.J., 2015 (April), Technical Report Maiden Resource Estimate Dark Star Deposit Elko County, Nevada USA: NI 43-101 technical report prepared by APEX Geoscience Ltd., 179 p.

Dufresne, M.B., Koehler, S.R., and Jackson, M.R., 2017 (March), Technical Report on the Railroad – Pinion Project, Elko County, Nevada USA: unpublished technical report (NI43-101 compliant) prepared for Gold Standard Ventures Corp., 259p.

Emmons, W.H., 1910, A Reconnaissance of Some Mining Camps in Elko, Lander and Eureka Counties, Nevada: U.S. Geological Survey Bulletin 408.

Galey, J.T., 1983, Appraisal, Railroad Project, Elko County, Nevada, internal company report of AMAX Exploration, Inc.

Golder Associates, Inc, 2019 (July 10, by R. Browne), Geotechnical Testing of TIW Deposists from Corehole DS17-02, South Railroad Project, Nevada.

Harp, M.T., Edie, R.J., Jackson, M.R., Koehler, S.R., Norby, J.W., Whitmer, N.E., Moore, S., and Wright, J.L., 2016, New Discovery within the Dark Star Corridor, Railroad-Pinion District, Elko County, Nevada: Association for Mineral Exploration British Columbia, Mineral Exploration Roundup 2016, Vancouver, abstract volume, p. 44.

Henry, C.D., Jackson, M.R., Mathewson, D.C., Koehler, S.R., and Moore, S.C., 2015, Eocene Igneous Geology and Relation to Mineralization: Railroad District, Southern Carlin Trend, Nevada: in Pennell, W.M., and Garside, L.J., eds., New Concepts and Discoveries, Geological Society of Nevada 2015 Symposium, Reno, Nevada, p. 939-965.

Hofstra, A.H., and Cline, J.S., 2000. Characteristics and models for Carlin-type gold deposits: in Society of Economic Geologists Reviews vol. 13, pp. 163-220.

Hunsaker III, E.L., 2010 (May), Technical Report on the Railroad Project, Elko County, Nevada, USA: unpublished Technical Report prepared for Gold Standard Ventures Corp (NI43-101 compliant), 67p.

Hunsaker III, E.L., 2012a (February), Technical Report on the Railroad Project, Elko County, Nevada, USA: unpublished Technical Report prepared for Gold Standard Ventures Corp (NI43-101 compliant), 78p.


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FORM 43-101F1 TECHNICAL REPORT

Hunsaker III, E.L., 2012b (April), Amended and Restated Technical Report on the Railroad Project, Elko County, Nevada, USA: unpublished Technical Report prepared for Gold Standard Ventures Corp (NI43-101 compliant), 82p.

Jackson, M.R., Lane, M., and Leach, B., 2002, Geology of the West Leeville Deposit: in Thompson, T.B., Teal, L., and Meeuwig, R., eds., Gold Deposits of the Carlin Trend: Nevada Bureau of Mines and Geology Bulletin 111, p. 106-114.

Jackson, M.R. and Koehler, S.R., 2014, Carlin-style Gold and Polymetallic Targets within a Large Eocene, Magmato-thermal System on the Carlin Trend, Nevada: Abstract for the AMEBC Mineral Exploration Roundup 2014, Vancouver, British Columbia.

Jackson, M.R., Mathewson, D.C., Koehler, S.R., Harp, M.T., Edie, R.J., Whitmer, N.E., Norby, J.W., and Newton, M.N., 2015, Geology of the North Bullion Gold Deposit: Eocene Extension, Intrusion and Carlin-style Mineralization, the Railroad District, Carlin Trend, Nevada: in Pennell, W.M., and Garside, L.J., eds., New Concepts and Discoveries, Geological Society of Nevada 2015 Symposium, Reno, Nevada, p. 313-331.

Jones, W.C., and Postlethwaite, C., 1992, Pieretti Land Block, Pieretti Ranch and Pinon Projects, Elko and Eureka Counties, Nevada: internal company report dated April, 1991, 148p.

Jones, M., Thomas, D., Shabestari, P. and Babcock, J., 1999 (October), Railroad Project 1999 Summary Report: ed. Cupp, B.L., unpublished report prepared for Kinross Gold U.S.A., Inc., 55p.

Kappes, Cassiday & Associates, 2004 (March), Pinion & Railroad Project - Report on Metallurgical Testwork, Project No. 126C, File 7355, unpublished report prepared for Royal Standard Minerals, by Kappes, Cassiday and Associates, 130p.

Kappes, Cassiday & Associates, 2006, Pinion Railroad Project Report of Metallurgical Test Work, March 2006, Project 126C, File 7355. Unpublished confidential report prepared on behalf of Royal Standard Minerals, 146p.

Kappes, Cassiday & Associates, 2016a (April), Pinion Project Bottle Roll Leach Testing on Sample Received 08 March 2016, Report on Metallurgical Testwork, April 2016, Project No. 7355C, File 7355, Report I.D.:KCA0160038_PIN04_01. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 99p.

Kappes, Cassiday & Associates, 2016b (June), Pinion Project Bottle Roll Leach Check Test Work, Report on Metallurgical Testwork, June 2016, Project No. 7355C, File 7355, Report I.D.: KCA0160059_PIN06_02, unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 45p.

Kappes, Cassiday & Associates, 2016c (June), Pinion Project Coarse Crushed and Milled Bottle Roll Leach Testing, Report on Metallurgical Testwork, June 2016, Project No. 7355C, File 7355, Report I.D.:KCA0150075_PIN01_04, unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 283p.

Kappes, Cassiday & Associates, 2017a (June), Pinion Project Column Leach Tests, Report on Metallurgical Testwork, June 2017, Project No. 7355C, File 7355, Report I.D.: KCA0160137 _PIN07_02. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 855p.


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FORM 43-101F1 TECHNICAL REPORT

Kappes, Cassiday & Associates, 2017b (November), Dark Star Project, Bottle Roll and Column Leach Tests, Report of Metallurgical Test Work, November 2017. Project No. 9108C, File 9108, Report I.D.: KCA0170014_DKST_01. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes Cassiday and Associates, 1180p.

Kappes, Cassiday & Associates, 2018a (June), Pinion Project HPGR Test Work, Report of Metallurgical Test Work, June 2018. Project No.7355C, File 7355, Report I.D.: KCA0170101_PIN08-01. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes Cassiday and Associates, 115p.

Kappes, Cassiday & Associates, 2018b (June), Dark Star Project HPGR Test Work – Report of Metallurgical Test Work, June 2018. Project No. 9108C, File 9108, Report ID: KCA0170085_DKST02_01. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates.

Kappes, Cassiday & Associates, 2018c (October), Jasperoid Wash Project, Report of Metallurgical Test Work: Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates.

Kappes, Cassiday & Associates, 2019a (July), Pinion Project Phase 3 Metallurgical Program – Report of Metallurgical Test Work, July 2019. Project No. 7355 C, File 7355, Report ID: KCA0180061_PIN09_03. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 1182p.

Kappes, Cassiday & Associates, 2019b, (July), Dark Star Project Phase 3 Metallurgical Program – Report of Metallurgical Test Work, July 2019. Project No. 9108 C, File 9108, Report ID: KCA0180065_DKST03_01. Unpublished report prepared for Gold Standard Ventures Corp., by Kappes, Cassiday and Associates, 2047p.

Ketner, K.B. and Smith Jr., J.F., 1963, Geology of the Railroad mining district, Elko County, Nevada: U.S. Geological Survey Bulletin 106.

Koehler, S.R., Dufresne, M.B. and Turner, A., 2014 (March), Technical Report on the Railroad and Pinion Projects, Elko County, Nevada USA: unpublished Technical Report prepared for Gold Standard Ventures Corp. (NI43-101 compliant), 158p.

Koehler, S.R., Edie, R.J., Harp, M.T., Henry, C., Jackson, M.R., Mathewson, D.C., Norby, J.W., and Whitmer, N.E., 2015, Precious and Base Metal Mineralization Within a Large Eocene, Magmato-Thermal System, Railroad District, Carlin Trend, Nevada: in Pennell, W.M., and Garside, L.J., eds., New Concepts and Discoveries, Geological Society of Nevada 2015 Symposium, Reno, Nevada, p. 1229-1242.

Kuehn, C.A. and Rose, A.W., 1992, Geology and geochemistry of wall-rock alteration at the Carlin gold deposit, Nevada: Economic Geology, v. 87, p. 1697–1721.

Kuhl, T.F., 1985, Geological Ore Reserves, Railroad Project, Elko County, Nevada: internal company memorandum prepared for NICOR Mineral Ventures.

LaPointe, Daphne, D., Tingley, Joseph V., and Jones, Richard B., 1991, Mineral Resources of Elko County, Nevada: Nevada Bureau of Mines and Geology Bulletin 106.

Longo, A.A., Thompson, T.B., and Harlan J. B., 2002, Geologic Overview of the Rain Subdistrict: in Thompson, T.B, Teal, L., and Meeuwig, R.O., eds., Gold Deposits of the Carlin Trend, edited by, Nevada Bureau of Mines and Geology Bulletin 111.


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FORM 43-101F1 TECHNICAL REPORT

Lund, K., 2008. Geometry of the Neoproterozoic and Paleozoic rift margin of western Laurentia: Implications for mineral deposit settings: Geosphere, v. 4, pp. 429-4440

Macy, F.A., 1991 (September), Report on Preliminary Cyanidation Testwork – Cord Ranch Cuttings Composites, MLI Job No. 1666, September 5, 1991, unpublished report by McClelland Laboratories Inc. to Crown Resources Corporation, 15p.

Masters, T.D., 2003a (February), Gold mineral resource Calculations of the POD Gold Zone at Railroad Bullion: unpublished report prepared for Royal Standard Minerals Inc., 8p.

Masters, T.D., 2003b (July), History, Geology, Resources, Discovery Potential and proposed Exploration Drilling Program for the Pinion – Railroad Project, Nevada: unpublished report prepared for Royal Standard Minerals Inc., 38p.

Mathewson, D.C., 2002, Carlin Gold Trend, Nevada Longitudinal Section – The “Four Windows”: unpublished cross section, files of Gold Standard Ventures Corp.

McClelland Laboratories Inc., 1995, Report on Cyanidation Testwork – South Bullion Samples. MLI Job No. 2136 -June 9, 1995: unpublished report prepared for Cypress Metals Company,

McComb, M., 2016 (May), Petrographic Examination of Fourteen Core Samples from the Dark Star Project, Nevada, internal company report prepared for Gold Standard Ventures.

McCusker, R. and Drobeck, P., 2012, Piñon Project Summary: unpublished report prepared for Royal Standard Minerals Inc.

Muntean, J.L., Coward, M.P. and Tarnocai, C.A., 2011. Reactivated Paleozoic normal faults: controls on the formation of Carlin-type gold deposits in north-central Nevada: in Ries, A.C., Butler, R.W.H., and Graham, R.R. (eds), Deformation of the Continental Crust: The Legacy of Mike Coward. Geological Society, London, Special Publications, no. 272, pp. 571-587.

Muntean, J.L., and Cline, J.S., 2018, Introduction, Diversity of Carlin-style Gold Deposits: in Muntean, J.L., ed., Diversity of Carlin-style Gold Deposits, Reviews in Economic Geology, v. 20, Society of Economic Geologists, p. 1-5.

Norby, J.W. and Orobona, M.J.T., 2002, Geology and Mineral Systems of the Mike Deposit: in Thompson, T.B., Teal, L., and Meeuwig, R., eds., Gold Deposits of the Carlin Trend: Nevada Bureau of Mines and Geology Bulletin 111, p. 143-167.

Norby, J.W., Edie, R.J., Harp, M.T., Jackson, M.R., Koehler, S.R., Mathewson, D.C., Moore, S., Whitmer, N.E., and Wright, J.L., 2015, Pinion Gold Deposit, Elko County, Nevada: in Pennell, W.M., and Garside, L.J., eds., New Concepts and Discoveries, Geological Society of Nevada 2015 Symposium, Reno, Nevada, p. 169-189.

Nordquist, W.A., 1992, Railroad Project, Elko County Nevada 1991 Annual Report: Westmont Gold Inc., internal company report.

Oversby, B., 1973, New Mississippian formation in northeast Nevada, and its possible significance: American Association of Petroleum Geologists Bulletin, v. 57, p. 1779-1783.


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FORM 43-101F1 TECHNICAL REPORT

Parr, A.J., 1998, Pinon Project, Elko County Nevada, 1997 Exploration Report: internal company report prepared for Cameco (US) Inc.

Parr, A.J., 1999, Pinon and Jasperoid Wash Projects, Elko County Nevada, 1999 Exploration Report: internal company report prepared for Cameco (US) Inc.

Peek, B.C., 1994 (February), Dark Star Project, Elko County, Nevada, Geologic Resource Estimate: internal report prepared for Crown Resources Corp., 4 p.

Rayias, A.C., 1999, Stratigraphy, Structural Geology, Alteration, and Geochemistry of the Northeastern Railroad District, Elko County, Nevada: unpublished M.Sc. thesis, University of Nevada, Reno.

Redfern, R. R., 2002, Geological Report on the Dixie Creek Property: unpublished report for Frontier Pacific Mining Corporation, 34p.

Ressel, M.W., 2000, Summary of Research on Igneous Rocks and Gold Deposits on the Carlin Trend, Nevada: Ralph J. Roberts Center for Research in Economic Geology Annual Research Meeting 1999, Program and Reports, 38p.

Shaddrick, D.R., 2012, Technical Report on the Railroad Project, Elko County, Nevada, USA: Technical Report prepared for Gold Standard Ventures Corp (NI43-101 compliant).

Smith, J.F., and Ketner, K.B., 1975, Stratigraphy of Paleozoic Rocks, Carlin-Piñon Range area, Nevada: U.S. Geological Survey Professional Paper 867-A.

Smith, J.F., and Ketner, K.B., 1978, Geologic Map of the Carlin-Pinon Range Area, Elko and Eureka Counties, Nevada: United States Geological Survey, Map I-1028, 1:62,500.

Steffen Robertson and Kirsten (B.C.) Inc., 1989 (August), Draft Acid Rock Drainage Technical Guide Volume 1: Norecol Environmental Consultants, and Gormely Process Engineering, British Columbia Acid Mine Drainage Task Force Report, 274p.

Stepperud, 2017a, Pinion Project Comminution Testing, Hazen Project 12352 Report – February 1, 2017. Unpublished letter report to Kappes, Cassiday & Associates.

Stepperud, 2017b, Comminution Testing, Hazen Project 12391 Report and Appendices A and B – July 5, 2017. Unpublished letter report to Kappes, Cassiday & Associates.

Stepperud, 2017c, Comminution Testing, Hazen Project 12514, Report and Appendices A and B – April 9, 2017. Unpublished letter report to Kappes, Cassiday & Associates.

Steperud, 2019a, Comminution Testing, Hazen Project 12635 Report and Appendices A and B, March 6, 2019, 73p. Unpublished letter report to Kappes, Cassiday & Associates.

Stepperud, 2019b, Comminution Testing, Hazen Project 12620, Report and Appendices A and B – February 2019, 63p. Unpublished letter report to Kappes, Cassiday & Associates.

Stewart, J.H., 1980, Geology of Nevada: Nevada Bureau of Mines and Geology Special Publication 4, 136p.


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FORM 43-101F1 TECHNICAL REPORT

Teal, L., and Jackson, M., 1997, Geologic Overview of the Carlin Trend Gold Deposits and Description of Recent Deep Discoveries: in Vikre, P., Thompson, T.B., Bettles, K., Christensen, O., and Parratt, R., eds., Carlin-Type Gold Deposits Field Conference, Society of Economic Geologists Guidebook Series, Volume 28, p. 3-37.

Teal, L. and Jackson, M., 2002: Geologic Overview of the Carlin Trend Gold Deposits: in Thompson, T.B., Teal, L., and Meeuwig, R., eds., Gold Deposits of the Carlin Trend: Nevada Bureau of Mines and Geology Bulletin 111, p. 9-19.

Turner, A., Dufresne, M.B. and Koehler, S.R., 2015 (March), Technical Report on the Railroad and Pinion Projects, Elko County, Nevada USA: unpublished Technical Report prepared for Gold Standard Ventures Corp. (NI43-101 compliant), 196p.

US Climate Data, https://www.usclimatedata.com/climate/carlin/nevada/united-states/usnv0107.

Wells, R.A., 1995, Pinion/South Bullion Deposit Resource Estimates: Cyprus Metals Exploration Corporation internal memorandum.

Western Regional Climate Center Historic Climate Information, Carlin Newmont Mine, Nevada, 2011: Administered by N.O.A.A. http://www.wrcc.dri.edu/Climsumetershtml.

Wood, J., 1995, South Bullion/Dark Star Resource Estimates: Cyprus Metals Exploration Corporation internal memorandum.

Wright, J.L., 2013, Railroad Property Gravity Survey – IV. internal company report prepared for Gold Standard Ventures.

Wright, J.L., 2016a, Railroad Property CSAMT Survey – Phase VI GIS Compilation: internal company report prepared for Gold Standard Ventures.

Wright, J.L., 2016b, Railroad Property Pearson, Deritter and Johnson Airborne Magnetic Survey GIS Database: internal company report prepared for Gold Standard Ventures.


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FORM 43-101F1 TECHNICAL REPORT

Appendix A – Preliminary Feasibility Study Contributors and Professional Qualifications – Certificates of Qualified Persons


	M3-PN185074 24 October 2019 Revision 0	433

CERTIFICATE OF QUALIFIED PERSON

Art S. Ibrado

I, Art S. Ibrado, PhD, PE, do hereby certify that:

1.	I am employed as a process engineer at M3 Engineering & Technology Corp., 2051 W Sunset Rd, Suite 101, Tucson, AZ 85704, USA

2.	I hold the following academic degrees:

	Bachelor of Science in Metallurgical Engineering, University of the Philippines, 1980 Master of Science (Metallurgy), University of California at Berkeley, 1986 Doctor of Philosophy (Metallurgy), University of California at Berkeley, 1993

3.	I am a registered professional engineer in the State of Arizona (No. 58140) and a Qualified Professional (QP) member of the Mining and Metallurgical Society of America (MMSA).

4.	I have worked as a metallurgist in the academic and research setting for five years, excluding graduate school research, and in the mining industry for 13 years, before joining M3 Engineering in July 2009.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am the principal author and contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of Septermber 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.1 to 1.3, 1.15, 21.5.3.3, 2 to 5, 21.5.4, 23 to 26. I have visited the project site on September 25, 2019, for a day.

7.	I have no prior involvement with the project or the property that is the subject of the Technical Report.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October 2019.

(Signed) “Art S. Ibrado”
Signature of Qualified Person

Art S. Ibrado
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Matthew Sletten

I, Matthew Sletten, P.E., do hereby certify that:

1.	I am a Project Manager of:

M3 Engineering & Technology Corp.
2175 W. Pecos Rd. Suite 3
Chandler, AZ 85224

2.	I graduated with a BS in Civil Engineering and an MS in Civil Engineering from the South Dakota School of Mines and Technology in 2004 and 2006, respectively.

3.	I am a Professional Engineer in good standing in the State of Arizona in the area of Civil Engineering.

4.	I have worked as an engineer and project manager in the base metals and precious metals industry for a total of 15 years. My experience includes detailed engineering, engineering management, project management, corporate management, capital and operating cost development and report development for major mining projects throughout the world.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of September 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.10, 1.12 to 1.14, 18.1 to 18.5, 19, 21.3, 21.5.4, & 22.

7.	I have not had prior involvement with the property that is the subject of the Technical Report.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October 2019.

(Signed) “Matthew Sletten”
Signature of Qualified Person

Matthew Sletten, PE

CERTIFICATE OF QUALIFIED PERSON

Steven J. Ristorcelli, C.P.G.

I, Steven J. Ristorcelli, C.P.G., do hereby certify that:

1.	I am a Principal Geologist of Mine Development Associates, Inc. (a Division of RESPEC), 210 South Rock Blvd., Reno, Nevada, 89502.

2.	I graduated with a Bachelor of Science degree in Geology from Colorado State University in 1977 and a Master of Science degree in Geology from the University of New Mexico in 1980.

3.	I am a Certified Professional Geologist (#10257) in good standing with the American Institute of Professional Geologists. I am also registered as Professional Geologist in the state of California (#3964).

4.	I have worked as geologist for over 40 years. I have conducted exploration, definition, modeling, and estimation of sediment-hosted epithermal gold-silver deposits in the Western US and Canada.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA”, dated October 24, 2019, with an effective date of September 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.4, 1.5, 1.8, 6, 7, 8, 9, 10, 11, 12, 14.1, 14.3, and 14.4. I have visited the project site on November 18, 2016 for a period of two days.

7.	I have not have had prior involvement with the property that is the subject of the Technical Report.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “S Ristorcelli”
Signature of Qualified Person

Steven J. Ristorcelli
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

MICHAEL B. DUFRESNE, M.SC., P.GEOL., P.GEO.

I, Michael B. Dufresne, M.Sc., P.Geol., P.Geo., do hereby certify that:

1.	I am President and Senior Partner of:

APEX Geoscience Ltd. (APEX)
Suite 110, 8429 – 24th Street NW
Edmonton, Alberta T6P 1L3
Phone: 780-467-3532


2.	I graduated with a B.Sc. in Geology from the University of North Carolina at Wilmington in 1983 and with a M.Sc. in Economic Geology from the University of Alberta in 1987.

3.	I am and have been registered as a Professional Geologist with the Association of Professional Engineers and Geoscientists of Alberta since 1989. I have been registered as a Professional Geoscientist with the Association of Professional Engineers and Geoscientists of BC since 2011, and the Association of Professional Engineers and Geoscientists of Northwest Territories and Nunavut since 2016.

4.	I have worked as a geologist for more than 30 years. My experience includes exploration for, and the evaluation of, gold deposits of various types, including sediment-hosted (Carlin-type) mineralization. I have conducted and supervised numerous resource estimates, including disseminated sediment hosted gold deposits, over the last 17 years

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of Septermber 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am solely responsible for the preparation of sections 10.2, 10.6.1, 10.6.3, 11.1, 11.2, 12.3, 12.8 and 14.5. I am jointly responsible for the preparation of sections 1.4, 1.5, 10.8, 11.5 and 12.9. I have visited the project site on a number of occasions with my most recent visit June 7^thto 9^th, 2017 for a period of 3 days.

7.	I have have had prior involvement with the property that is the subject of the Technical Report. I have been employed as an independent qualified person through APEX by Gold Standard Ventures since 2014 and have been a contributing author on a number of Technical Reports from 2014 to 2018 that are all available on sedar.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer and the property applying all of the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “Michael B. Dufresne”
Signature of Qualified Person

Michael B. Dufresne, M.Sc., P.Geol., P.Geo.
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Michael S. Lindholm, C.P.G.

I, Michael S. Lindholm, C.P.G., do hereby certify that:

1.	I am a Senior Geologist of Mine Development Associates, Inc. (a Division of RESPEC), 210 South Rock Blvd., Reno, Nevada, 89502.

2.	I graduated with a Bachelor of Science degree in Geology from Stephen F. Austin State University in 1984 and with a Master of Science degree in Geology from Northern Arizona University in 1989.

3.	I am a Certified Professional Geologist (#11477) in good standing with the American Institute of Professional Geologists. I am also registered as Professional Geologist in the state of California (#8152).

4.	I have worked as geologist for 32 years. I have conducted exploration, definition, modeling, and estimation of sediment-hosted epithermal gold-silver deposits in the Western US.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA”, dated October 24, 2019, with an effective date of September 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.4, 1.5, 1.8, 6, 7, 8, 9, 10, 11, 12, 14.1 and 14.2. I have visited the project site on September 18, 2018 for a period of two days.

7.	I have not have had prior involvement with the property that is the subject of the Technical Report.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “Michael S. Lindholm”
Signature of Qualified Person

Michael S. Lindholm
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Thomas L. Dyer, PE

I, Thomas L. Dyer, PE, do hereby certify that:

1.	I am a Principal Engineer of Mine Development Associates, Inc. (a Division of RESPEC), 210 South Rock Blvd., Reno, Nevada, 89502.

2.	I graduated with a Bachelor of Science degree in Mine Engineering from South Dakota School of Mines and Technology in 1996.

3.	I am a Registered Professional Engineer in the state of Nevada (#15729) and a Registered Member (#4029995RM) of the Society of Mining, Metallurgy and Exploration.

4.	I have worked as mining engineer for more than 22 years. Relevant experience includes mine design, reserve estimation and economic analysis of precious-metals deposits in the United States and various countries in the world. I have worked as Chief Engineer of an operating heap leach and mill gold mine in Nevada.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA”, dated October 24, 2019, with an effective date of September 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.8, 1.9, 15, 16, 21.1, and 21.4. I have visited the project site on November 18, 2016 for a period of two days.

7.	I have had prior involvement with the property that is the subject of the Technical Report. Through MDA, I have completed internal mining and economic studies for Gold Standard Ventures Corp since 2016.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “Thomas L. Dyer”
Signature of Qualified Person

Thomas L. Dyer, PE
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Gary L. Simmons

I, Gary L Simmons, Qualified Professional (QP), do hereby certify that:

1.	I am the Principal Owner of:

GL Simmons Consulting, LLC
15293 Shadow Mountain Ranch Road
Larkspur, CO 80118

2.	I graduated with a Bachelor of Science Degree in Metallurgical Engineering from the Colorado School of Mines, Golden, Colorado, USA, in 1973.

3.	I am a Professional Metallurgical Engineer, registered with the Mining and Metallurgical Society of America, Qualified Professional (QP) Member in Metallurgy, Member Number – 01013QP, in good standing in the USA.

4.	I have practiced in my profession since 1973. My relevant experience includes mine site and corporate level process development, project engineering, operations supervision and as a mineral processing project development consultant, in the base metals and gold/silver mining business, for a total of 45 years.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of Septermber 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am a contributing author for Sections 1.6 and 13. I have visited the project site on June 22, 2017 for period of one day.

7.	I have been involved with this project since 2016 as a metallurgical consultant and have authored internal reports and have been a contributing QP for press releases and regulatory filings relating to metallurgy.

8.	I have no other involvement with the subject property prior to work for the current owner.

9.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

11.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October 2019.

(Signed) “Gary L. Simmons”
Signature of Qualified Person

Gary L. Simmons
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

I, Carl E. Defilippi, RM SME, do hereby certify that:

1.	I am a Project Manager of:

Kappes, Cassiday & Associates
7950 Security Circle
Reno, Nevada 89506

2.	I graduated with a BS in Chemical Engineering and an MS in Metallurgical Engineering from the Mackay School of Mines, University of Nevada, in 1978 and 1981, respectively.

3.	I am a Registered Member of SME in good standing.

4.	I have worked in the precious metals industry for a total of 38 years. My experience includes all aspects of cyanide processing: operations, project management, construction, commissioning, recovery plant fabrication, metallurgical audits and reviews of numerous operating heap leach and milling facilities throughout the world.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of September 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.6.5, 1.7, 13.10, 17.0 to 17.7, 17.9, 17.10, 21.2, 21.5 except 21.5.3.3 and 21.5.4. I have visited the project site on 28 August 2019 for a period of one day.

7.	I have prior involvement with the property that is the subject of the Technical Report. I participated in scoping and pre-feasibility level studies for Royal Standard Minerals in 2003 through 2005. The Project was called Pinion-Railroad at that time.

8.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.5 of National Instrument 43-101.

10.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “Carl E. Defilippi”
Signature of Qualified Person

Carl E. Defilippi
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Richard DeLong

I, Richard DeLong, M.S., P.G., do hereby certify that:

1.	I am President of:

EM Strategies, Inc.
1650 Meadow Wood Lane,
Reno, Nevada 89502

2.	I graduated with a Masters Degree in Geology and a Masters Degree in Resource Management from the University of Idaho.

3.	I am a Professional Geologist in good standing in the State of Idaho in the area of Geology (No. 727). I am also recognized as a Qualified Person Member with special expertise in Environmental Permitting and Compliance with the Mining and Metallurgical Society of America (No. 01471QP).

4.	I have worked as an environmental permitting and compliance specialist for a total of 31 years. My experience includes permit acquisition of sate and federal permits and baseline data acquisition programs for mining and exploration operations.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of Septermber 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 1.11 and 20. I have not visited the project site.

7.	I have prior involvement with the property that is the subject of the Technical Report. My involvement with the property is the ongoing work associated with environmental baseline data collection and the acquisition of the necessary state and federal permits for the development of the mining operation.

8.	I have no other involvement with the project.

9.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10.	I am [independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

11.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) “Richard DeLong”
Signature of Qualified Person

Richard DeLong
Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

Kenneth L. Myers

I, Kenneth L. Myers, P.E. do hereby certify that:

1.	I am President/Principal of:

The MINES Group, Inc.
1325 Airmotive Way
Reno, NV 89502

2.	I graduated with a Bachelor of Science Degree in Civil Engineering from the University of Cincinnati, and a Master of Science Degree in Geological Engineering from the University of Missouri-Rolla.

3.	I am a Professional Engineer in good standing in the State of Nevada in the areas of Civil Engineering, No. 10254. I am also registered as a Professional Engineer in the States of Alaska, California, Idaho, New Mexico, South Dakota, and Wyoming .

4.	I have worked as an engineer for a total of 46 years. My experience includes mine facility design, seismic hazard analysis, surface and groundwater hydrology, site investigation and gepotechnical engineering, slope stability analysis, and water balance analysis

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am a contributing author for the preparation of the technical report titled “South Railroad Project NI 43-101 Technical Report, Preliminary Feasibility Study, Carlin Trend, Nevada, USA” dated October 24, 2019, with an effective date of Septermber 9, 2019 (the “Technical Report”), prepared for Gold Standard Ventures Corp. I am responsible for the preparation of Sections 17.8, 18.6, 18.7, and 18.8. I have visited the project site on October 22, 2019 for a period of 1 day.

7.	I have not have had [prior involvement with the property that is the subject of the Technical Report.

8.	Disclose any additional involvement with the project or collaboration with the Client as needed - NA

9.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

10.	I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

11.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12.	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Signed and dated this 24^thday of October, 2019.

(Signed) Kenneth L .Myers
Signature of Qualified Person

Kenneth L. Myers
Print Name of Qualified Person


		12.5.3	Dark Star Drill Program QA/QC 2015	106
		12.5.4	Dark Star Drill Program QA/QC 2016	107
		12.5.5	Dark Star Drill Program QA/QC 2017	110
		12.5.6	Dark Star Drill Program QA/QC 2018	111
		12.5.7	Dark Star Drill Program QA/QC 2019	113
	12.6	GOLD STANDARD’S PINION DRILL PROGRAM QA/QC		114
		12.6.1	Pinion Drill Program QA/QC CRMs	115
		12.6.2	Pinion Drill Program QA/QC Field Duplicates	122
		12.6.3	External Check Assays for Pinion Drilling	124
		12.6.4	Pinion Drill Program QA/QC Blanks	125
		12.6.5	Twin Holes	128
		12.6.6	Pinion Drill Program QA/QC on Barite	129
	12.7	GOLD STANDARD’S JASPEROID WASH DRILL PROGRAM QA/QC		130
	12.8	NORTH BULLION DEPOSITS DRILL PROGRAM QA/QC		131
		12.8.1	Blanks	131
		12.8.2	Certified Reference Materials	131
	12.9	SUMMARY STATEMENT ON DATA VERIFICATION		134
13	MINERAL PROCESSING AND METALLURGICAL TESTING			136
	13.1	2015 – 2016 GOLD STANDARD PINION DEPOSIT CYANIDE BOTTLE-ROLL LEACH		138
		13.1.1	2015 – 2016 Pinion Head Assays	138
		13.1.2	2015 – 2016 Pinion Bottle-Roll Test Results	138
	13.2	2016 - 2017 GOLD STANDARD PINION DEPOSIT METALLURGICAL TESTING		140
		13.2.1	2017 Pinion Head Assays	141
		13.2.2	2016 – 2017 Bottle Roll and Column Leach Testing (KCA)	142
		13.2.3	2017 Pinion Comminution Characterization at HRI	144
		13.2.4	2017 Pinion Load Permeability Test Work on Column Tailings	145
	13.3	GOLD STANDARD 2018 PINION DEPOSIT HIGH PRESSURE GRINDING ROLL (HPGR) TESTING		145
		13.3.1	2018 Head Assays Pinion Main Zone HPGR Composite	146
		13.3.2	2018 Pinion Main Zone HPGR Bottle-Roll and Column-Leach Testing	146
		13.3.3	2018 Pinion Main Zone HPGR Agglomeration and Load Permeability Testing	149
	13.4	2019 GOLD STANDARD PINION DEPOSIT METALLURGICAL TEST WORK		149
		13.4.1	2019 Pinion Head Assays	150
		13.4.2	2019 Pinion Bottle Roll and Column Leach Testing at KCA	151
		13.4.3	2019 Pinion Comminution Characterization at HRI	153
		13.4.4	2019 Pinion Load Permeability Test Work on Column Tailings	154
	13.5	1991 DARK STAR DEPOSIT METALLURGICAL TESTING		154
	13.6	2017 GOLD STANDARD DARK STAR DEPOSIT METALLURGICAL TESTING		155
		13.6.1	2017 Dark Star Head Assays for Bottle-Roll and Column-Leach Tests	155
		13.6.2	2017 Dark Star Bottle-Roll and Column-Leach Tests at KCA	156
		13.6.3	2017 Dark Star Comminution Characterization at HRI	159


		13.6.4	2017 Dark Star Load Permeability Testing	160
	13.7	2018 GOLD STANDARD DARK STAR HPGR METALLURGICAL TEST WORK		160
		13.7.1	2018 Dark Star HPGR Head Assays	161
		13.7.2	2018 Dark Star HPGR Composite Bottle-Roll and Column-Leach Tests	161
		13.7.3	2018 Dark Star Main & North HPGR-Crushed Load Permeability Testing	163
	13.8	2019 GOLD STANDARD DARK STAR DEPOSIT METALLURGICAL TEST WORK		163
		13.8.1	2019 Dark Star Head Assays for Bottle-Roll and Column-Leach Tests	164
		13.8.2	2019 Dark Star Bottle-Roll and Column-Leach Tests at KCA	164
		13.8.3	2019 Dark Star Comminution Characterization at HRI	168
		13.8.4	2019 Dark Star Load Permeability Testing	169
	13.9	GEO-METALLURGY CHARACTERIZATION		169
		13.9.	Au and Ag Recovery Methodology	170
		13.9.	Pinion Deposit Recovery Models	175
		13.9.	Dark Star Main and Dark Star North Deposit Recovery Models	181
		13.9.	ROM and HPGR Dark Star Gold Recovery Equations (Transition)	187
		13.9.	ROM and HPGR Pinion Gold Recovery Equations (Transition)	188
	13.10	REAGENT CONSUMPTIONS SOUTH RAILROAD PROPERTY		188
		13.10.1	Cyanide	188
		13.10.2	Lime	189
		13.10.3	Cement	189
	13.11	METALLURGICAL TESTING ON JASPEROID WASH AND NORTH BULLION SAMPLES		193
		13.11.1	Jasperoid Wash Deposit Metallurgical Testing	193
		13.11.2	North Railroad Deposits Metallurgical Testing	193
14	MINERAL RESOURCE ESTIMATES			194
	14.1	INTRODUCTION		194
	14.2	DARK STAR MINERAL RESOURCES		194
		14.2.1	Dark Star Database	194
		14.2.2	Dark Star Geologic Model	197
		14.2.3	Dark Star Gold Modeling and Estimation	197
		14.2.4	Dark Star Gold Mineral Resources	207
		14.2.5	Dark Star Cyanide-Soluble Gold and Geo-Metallurgical Models	214
		14.2.6	Dark Star Acid-Base Accounting Model and Estimation	215
		14.2.7	Dark Star Density	218
		14.2.8	Discussion of Dark Star Estimated Gold Mineral Resource and Supporting Models	219
	14.3	PINION DEPOSIT MINERAL RESOURCES		222
		14.3.1	Pinion Database	222
		14.3.2	Pinion Geologic Model	225
		14.3.3	Pinion Gold Modeling and Estimation	226
		14.3.4	Pinion Silver Modeling and Estimation	233
		14.3.5	Pinion Density and Silver Resources	238
		14.3.6	Pinion Geo-Metallurgical Model	244


		14.3.7	Pinion Acid-Base Accounting Model and Estimation	250
		14.3.8	Pinion Density	254
		14.3.9	Discussion of Pinion Estimated Mineral Resources and Supporting Models	255
	14.4	JASPEROID WASH MINERAL RESOURCES		257
		14.4.	Jasperoid Wash Database	258
		14.4.	Jasperoid Wash Geologic Model	261
		14.4.	Jasperoid Wash Gold Modeling and Estimation	261
		14.4.	Jasperoid Wash Gold Mineral Resources	268
		14.4.	Jasperoid Wash Geo-Metallurgical Model	271
		14.4.	Jasperoid Wash Clay Model	273
		14.4.	Jasperoid Wash Density	273
		14.4.	Discussion of Jasperoid Wash Estimated Mineral Resources	273
	14.5	NORTH BULLION DEPOSITS MINERAL RESOURCES		275
		14.5.	North Bullion Deposits Data	275
		14.5.	North Bullion Deposits Lithological Setting of Mineralized Zones	276
		14.5.	North Bullion Deposits Geological Models	277
		14.5.	North Bullion Cyanide-Soluble Model	280
		14.5.	North Bullion Deposits Sample and Composite Statistics and Grade Capping	281
		14.5.	North Bullion Deposits Grade Continuity	282
		14.5.	North Bullion Deposits Density	283
		14.5.	North Bullion Deposits Grade Estimation	283
		14.5.	North Bullion Block Model Validation	284
		14.5.	North Bullion Deposits Mineral Resources	286
15	MINERAL RESERVE ESTIMATES			290
	15.1	INTRODUCTION		290
	15.2	PIT OPTIMIZATION		291
		15.2.1	Economic Parameters	292
		15.2.2	Geometric Parameters	294
		15.2.3	Cutoff Grades	297
		15.2.4	Pit Optimization Methods and Results	299
	15.3	PIT DESIGNS		306
		15.3.1	Road and Ramp Design	306
		15.3.2	Dark Star Pit Designs	307
		15.3.3	Pinion Pit Designs	309
	15.4	DILUTION		313
	15.5	PROVEN AND PROBABLE MINERAL RESERVES FOR DARK STAR AND PINION		313
16	MINING METHODS			315
	16.1	WASTE ROCK STORAGE AREAS AND LEACH PADS		315
	16.2	STOCKPILES		316
	16.3	MINE-PRODUCTION SCHEDULE		316


	16.4	EQUIPMENT SELECTION AND PRODUCTIVITIES		334
	16.5	EQUIPMENT REQUIREMENTS		334
		16.5.	Drilling Equipment	335
		16.5.	Loading Equipment	336
		16.5.	Haulage Productivity	336
		16.5.	Support and Maintenance Equipment	336
	16.6	MINING PERSONNEL AND STAFFING		337
17	RECOVERY METHODS			338
	17.1	GOLD RECOVERY		338
	17.2	REAGENTS AND CONSUMPTIONS		338
		17.2.1	Sodium Cyanide	338
		17.2.2	Lime	339
		17.2.3	Cement	339
		17.2.4	Activated Carbon	339
		17.2.5	Sodium Hydroxide (Caustic)	339
		17.2.6	Hydrochloric Acid	339
		17.2.7	Fluxes	339
		17.2.8	Antiscalant	340
	17.3	PROCESS FLOWSHEET		340
	17.4	ROM TRUCK STACKING		342
	17.5	CRUSHING		342
	17.6	AGGLOMERATION AND CONVEYOR STACKING		344
	17.7	LEACHING AND SOLUTION HANDLING		345
	17.8	LEACH PAD PHASING AND CONSTRUCTION		345
		17.8.1	Solution Ponds	346
	17.9	ADR PLANT		347
		17.9.1	Adsorption	349
		17.9.2	Carbon Acid Wash	349
		17.9.3	Desorption	349
		17.9.4	Electrowinning	350
		17.9.5	Carbon Handling & Thermal Regeneration	351
		17.9.6	Refining & Smelting	351
	17.10	ADR REAGENTS AND UTILITIES		352
	17.11	LABORATORY FACILITIES		352
18	PROJECT INFRASTRUCTURE			353
	18.1	ACCESS ROAD		353
	18.2	POWER SUPPLY		353
	18.3	PROJECT BUILDINGS		353
		18.3.1	Crushing Buildings	356


		18.3.2	Security Building at Access Gate	356
		18.3.3	Administration Building	356
		18.3.4	Truck Shop Building	356
		18.3.5	ADR Plant	356
		18.3.6	Laboratory	356
	18.4	SITEWIDE WATER MANAGEMENT STRATEGY		357
		18.4.1	Source of Mine Water	357
		18.4.2	Beneficial Reuse	363
		18.4.3	Water Disposal	364
	18.5	WATER MANAGEMENT INFRASTRUCTURE		365
		18.5.1	Dark Star Groundwater Dewatering System	365
		18.5.2	Seepage and Stormwater Management System	369
		18.5.3	Beneficial Reuse System	370
	18.6	HEAP LEACH PAD FACILITY		371
	18.7	HEAP LEACH FACILITY WATER BALANCE ANALYSIS		372
	18.8	SEISMIC HAZARD ANALYSIS		376
19	MARKET STUDIES AND CONTRACTS			382
	19.1	METAL PRICING		382
20	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT			385
	20.1	INTRODUCTION		385
	20.2	ENVIRONMENTAL BASELINE STUDIES		386
	20.3	BUREAU OF LAND MANAGEMENT PLAN OF OPERATIONS / NEVADA BUREAU OF MINING REGULATION AND RECLAMATION, NEVADA RECLAMATION PERMIT		386
		20.3.1	Bureau of Land Management Pre-Application Planning	387
		20.3.2	Plan of Operations Processing	387
	20.4	UNITED STATES ARMY CORPS OF ENGINEERS SECTION 404 PERMIT		388
	20.5	NATIONAL ENVIRONMENTAL POLICY ACT		388
	20.6	STATE OF NEVADA PERMITS		388
		20.6.	Water Pollution Control Permit	388
		20.6.	Air Quality Operating Permit	389
		20.6.	Water Rights	389
		20.6.	BLM Right-of-Way	389
	20.7	ELKO COUNTY		389
	20.8	OTHER PERMITS		389
	20.9	ENVIRONMENTAL STUDY RESULTS AND KNOWN ISSUES		390
	20.10	WASTE DISPOSAL AND MONITORING		391
	20.11	SOCIAL AND COMMUNITY ISSUES		391
	20.12	MINE CLOSURE		391


(Signed) “Art S Ibrado”	October 24, 2019
Art S. Ibrado, PE	Date

(Signed) “Mathew Sletten”	October 24, 2019
Mathew Sletten, PE	Date

(Signed) “Steven Ristorcelli”	October 24, 2019
Steven Ristorcelli, CPG	Date

(Signed) “Michael B. Dufresne”	October 24, 2019
Michael B. Dufresne, P.Geol., P.Geo.	Date

(Signed) “Michael Lindholm”	October 24, 2019
Michael Lindholm, CPG	Date

(Signed) “Thomas Dyer”	October 24, 2019
Thomas Dyer, PE	Date

(Signed) “Gary L. Simmons”	October 24, 2019
Gary L. Simmons, QP	Date

(Signed) “Carl Defilippi”	October 24, 2019
Carl Defilippi, PE	Date

(Signed) “Richard DeLong”	October 24, 2019
Richard DeLong, QP MMSA, RG, PG	Date

(Signed) “Kenneth L. Myers”	October 24, 2019
Kenneth L. Myers, PE	Date


SECTION				PAGE
DATE AND SIGNATURES PAGE				I
TABLE OF CONTENTS				II
LIST OF FIGURES AND ILLUSTRATIONS				XIII
LIST OF TABLES				XVIII
LIST OF APPENDICES				XXV
1	SUMMARY			1
	1.1	PRINCIPAL FINDINGS		1
	1.2	PROPERTY DESCRIPTION AND OWNERSHIP		3
	1.3	EXPLORATION AND MINING HISTORY		3
	1.4	GEOLOGY AND MINERALIZATION		3
	1.5	DATA VERIFICATION		4
	1.6	MINERAL PROCESSING AND METALLURGICAL TESTING		5
		1.6.1	Pinion Deposit Tests 2016 – 2019	5
		1.6.2	Gold Standard Dark Star Deposit Metallurgical Testing	8
		1.6.3	Comminution Characterization and Load Permeability Tests	11
		1.6.4	Geo-metallurgy Characterization and Recovery Models	11
		1.6.5	Reagent Consumption	12
	1.7	RECOVERY METHODS		13
	1.8	MINERAL RESOURCE ESTIMATE AND MINERAL RESERVE ESTIMATE		13
		1.8.1	Mineral Resource Estimate	13
		1.8.2	Mineral Reserve Estimate	15
	1.9	MINING METHODS		16
	1.10	INFRASTRUCTURE		17
	1.11	ENVIRONMENT AND PERMITTING		17
	1.12	WATER MANAGEMENT		18
	1.13	CAPITAL COST SUMMARY		19
	1.14	OPERATING COST SUMMARY		19
	1.15	CONCLUSIONS AND RECOMMENDATIONS		19
2	INTRODUCTION AND TERMS OF REFERENCE			22
	2.1	PURPOSE OF REPORT		22


	2.2	SOURCES OF INFORMATION		22
	2.3	PROJECT SCOPE AND TERMS OF REFERENCE		23
	2.4	FREQUENTLY USED ACRONYMS, ABBREVIATIONS, DEFINITIONS, AND UNITS OF MEASURE		24
3	RELIANCE ON OTHER EXPERTS			26
4	PROPERTY DESCRIPTION AND LOCATION			27
	4.1	LOCATION AND LAND AREA		27
	4.2	AGREEMENTS AND ENCUMBRANCES		29
	4.3	ENVIRONMENTAL PERMITS		31
		4.3.1	Other Permits	31
		4.3.2	Private Land Disturbance	31
5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			33
	5.1	ACCESS TO PROPERTY		33
	5.2	CLIMATE		33
	5.3	PHYSIOGRAPHY		33
	5.4	LOCAL RESOURCES AND INFRASTRUCTURE		33
6	HISTORY			35
	6.1	NORTH RAILROAD PORTION OF THE PROPERTY		35
	6.2	SOUTH RAILROAD PORTION OF THE PROPERTY		37
		6.2.1	Pinion Area Exploration History	37
		6.2.2	Dark Star Area Exploration History	40
		6.2.3	Jasperoid Wash	40
		6.2.4	Other Prospects in the South Railroad Portion of the Property	41
	6.3	HISTORICAL MINERAL RESOURCE ESTIMATES		42
		6.3.1	Pinion Deposit Historical Estimates	42
		6.3.2	Dark Star Deposit Historical Estimates	44
		6.3.3	POD (Railroad) Deposit Historical Mineral Resources 1985 - 2003	45
	6.4	HISTORICAL MINE PRODUCTION		46
		6.4.1	North Railroad	46
		6.4.2	South Railroad	46
7	GEOLOGICAL SETTING AND MINERALIZATION			47
	7.1	REGIONAL GEOLOGIC SETTING		47
	7.2	LOCAL AND PROPERTY GEOLOGY		50
		7.2.1	North Railroad Portion of the Property	52
		7.2.2	South Railroad Portion of the Property	54
	7.3	MINERALIZATION		57
		7.3.1	North Bullion Deposits	57
		7.3.2	Pinion Deposit	59


		7.3.3	Dark Star Deposit	60
		7.3.4	Jasperoid Wash Deposit	61
8	DEPOSIT TYPES			63
9	EXPLORATION			66
	9.1	2009 – 2019 GEOPHYSICS		66
	9.2	2010 – 2018 GEOCHEMISTRY		68
	9.3	2009 – 2019 GEOLOGIC MAPPING		69
	9.4	2014 – 2016 DARK STAR AND PINION PETROGRAPHY		71
10	DRILLING			72
	10.1	SUMMARY		72
	10.2	HISTORICAL NORTH RAILROAD DRILLING		75
		10.2.1	1969-1974 American Selco, Placer Amex and El Paso Gas Company	75
		10.2.2	1977-1980 AMAX	75
		10.2.3	1980-1981 Homestake	75
		10.2.4	1983 and 1985-1986 NICOR	75
		10.2.5	1987-1992 Westmont	75
		10.2.6	1994 Ramrod	75
		10.2.7	1995 Newmont	75
		10.2.8	1996-1997 Mirandor	75
		10.2.9	1998-1999 Kinross	76
		10.2.10	2005-2008 Royal Standard Minerals	76
	10.3	HISTORICAL SOUTH RAILROAD DRILLING		76
		10.3.1	1980-1981 AMOCO Minerals	76
		10.3.2	1981-1982 Newmont	76
		10.3.3	1983 Freeport	76
		10.3.4	1984 Cyprus-AMAX	76
		10.3.5	1985 Santa Fe Mining	76
		10.3.6	1987-1989 Newmont	76
		10.3.7	1987-1989 Teck Resources	76
		10.3.8	1988 Battle Mountain	76
		10.3.9	1989-1992 Westmont	77
		10.3.10	1988-1989 Freeport	77
		10.3.11	1990-1993 Crown Resources	77
		10.3.12	1994-1995 Cyprus Mining	77
		10.3.13	1997 Mirandor	77
		10.3.14	1997-1999 Cameco	77
		10.3.15	1998-1999 Kinross	77
		10.3.16	2003 and 2007 Royal Standard Minerals	77
	10.4	GOLD STANDARD DRILLING, NORTH RAILROAD AREA 2010 - 2017		77
		10.4.1	North Bullion Deposits Drilling by Gold Standard	80
		10.4.2	Bald Mountain Drilling by Gold Standard	82
	10.5	GOLD STANDARD DRILLING, SOUTH RAILROAD AREA 2012-2019		83


		10.5.1	Dark Star Area Drilling by Gold Standard	83
		10.5.2	Pinion Area Drilling by Gold Standard	84
		10.5.3	Jasperoid Wash Area Drilling by Gold Standard	85
		10.5.4	Irene Area Drilling by Gold Standard	85
		10.5.5	Dixie Area Drilling by Gold Standard	85
		10.5.6	Ski Track Drilling by Gold Standard	86
	10.6	DRILL-HOLE COLLAR SURVEYS		86
		10.6.1	Historical Collar Surveys, North Railroad Portion of the Property	86
		10.6.2	Historical Collar Surveys, South Railroad Portion of the Property	86
		10.6.3	Gold Standard Collar Surveys, North Railroad Portion of the Property	86
		10.6.4	Gold Standard Collar Surveys, South Railroad Portion of the Property	86
	10.7	DOWN-HOLE SURVEYS		87
		10.7.1	Historical Down-Hole Surveys, North and South Railroad Portions of the Property	87
		10.7.2	Gold Standard Down-Hole Surveys, North and South Railroad Portions of the Property	87
	10.8	SUMMARY STATEMENT		87
11	SAMPLE PREPARATION, ANALYSES AND SECURITY			88
	11.1	HISTORICAL OPERATORS’ DRILLING SAMPLES - NORTH RAILROAD PORTION OF THE PROPERTY		88
	11.2	GOLD STANDARD’S DRILLING SAMPLES - NORTH RAILROAD PORTION OF THE PROPERTY		89
	11.3	HISTORICAL OPERATORS - SOUTH RAILROAD PORTION OF THE PROPERTY		90
	11.4	GOLD STANDARD - SOUTH RAILROAD PORTION OF THE PROPERTY		92
		11.4.1	Pinion Deposit Area Drill Samples	92
		11.4.2	Dark Star Deposit Area Drill Samples	93
		11.4.3	Jasperoid Wash Area Drill Samples	94
		11.4.4	Dixie Area Drill Samples	94
		11.4.5	Ski Track Area Drill Samples	94
	11.5	AUTHOR’S OPINION		94
12	DATA VERIFICATION			95
	12.1	DARK STAR AND PINION DATABASE AUDITS		95
		12.1.1	2019 Audit of Dark Star Carbon, CO2 and Sulfur Data	97
		12.1.2	2019 Audit of Pinion Carbon, CO2 and Sulfur Data	98
	12.2	JASPEROID WASH DATABASE AUDIT		98
	12.3	NORTH BULLION DEPOSITS DATABASE AUDIT		99
		12.3.1	Drill-Hole Collar Locations and Down-Hole Survey Data	99
		12.3.2	Drill-Hole Assay Audit	99
	12.4	GOLD STANDARD QA/QC PROCEDURES		100
	12.5	DARK STAR DRILL PROGRAM QA/QC		100
		12.5.1	Dark Star Drill Program QA/QC 1991	101
		12.5.2	Dark Star Drill Program QA/QC 1997	103


21	CAPITAL AND OPERATING COSTS			392
	21.1	MINING CAPITAL		393
		21.1.1	Primary Equipment	393
		21.1.2	Support Equipment	393
		21.1.3	Blasting Equipment	393
		21.1.4	Mine Maintenance Capital	394
		21.1.5	Other Capital	394
		21.1.6	Mine Pre-production	394
	21.2	PROCESS CAPITAL		394
		21.2.1	Process Capital Cost Summary	394
		21.2.2	Freight	396
		21.2.3	Construction Support	396
		21.2.4	EPCM	396
		21.2.5	Vendor Support	396
		21.2.6	Spare Parts	397
	21.3	OWNER’S COSTS		397
		21.3.1	Land Purchases	397
	21.4	MINE OPERATING COST		397
		21.4.1	Mine General Services	398
		21.4.2	Mine Maintenance	399
		21.4.3	Drilling	400
		21.4.4	Blasting	400
		21.4.5	Loading	401
		21.4.6	Hauling	402
		21.4.7	Mine Support	403
	21.5	PROCESS OPERATING COST SUMMARY		403
		21.5.1	Personnel and Staffing	406
		21.5.2	Power	406
		21.5.3	Consumable Items	406
		21.5.4	General Facilities (Process Equipment Costs)	408
		21.5.5	Process Operating Cost Exclusions	408
	21.6	G&A COSTS		408
22	ECONOMIC ANALYSIS			410
	22.1	MINING PHYSICALS		410
	22.2	PROCESS PLANT PRODUCTION STATISTICS		411
	22.3	SMELTER RETURN FACTORS		411
	22.4	CAPITAL EXPENDITURE		411
	22.5	REVENUE		412
	22.6	TOTAL PRODUCTION COST		412
	22.7	DEPRECIATION		412