Klondex Mines Ltd. | Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada | Page ii |
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Date and Signature Page
The undersigned prepared this Technical Report (Technical Report) report, titled: Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014, in support of the public disclosure of Mineral Resource and Mineral Reserve estimates for the Fire Creek Project. The format and content of the Technical Report have been prepared in accordance with Form 43-101F1 of National Instrument 43-101 – Standards of Disclosure for Mineral Projects of the Canadian Securities Administrators.
Dated this March 16, 2015
Signed “Mark Odell” | No. 13708, Nevada |
Mark Odell, P.E | SME No. 2402150 |
Practical Mining LLC | Sealed) |
495 Idaho Street, Suite 205 | |
Elko, Nevada 89815, USA | |
(775) 345-3718 ext. 101 | |
Email: markodell@practicalmining.com | |
Signed “Laura Symmes” | SME No. 4196936 |
Laura Symmes | (Sealed) |
Practical Mining LLC | |
495 Idaho Street, Suite 205 | |
Elko, Nevada 89815, USA | |
(775) 345-3718 ext. 102 | |
Email: laurasymmes@practicalmining.com | |
Signed “Sarah Bull” | No. 22797, Nevada |
Sarah Bull, P.E | (Sealed) |
Practical Mining LLC | |
495 Idaho Street, Suite 205 | |
Elko, Nevada 89815, USA | |
775-345-3718 ext. 502 | |
Email:sarahbull@practicalmining.com | |
Signed “Karl T. Swanson” | AusIMM No. 304871 |
Karl T. Swanson, M.Eng., SME, AusIMM | SME No. 4043076 |
PO Box 86 | (Sealed) |
Larkspur, CO 80118, USA | |
Fax: (501) 638-9162 | |
Email: karl.swanson@yahoo.com |
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Table of Contents
Date and Signature Page | iii | ||
Table of Contents | iv | ||
List of Tables | x | ||
List of Figures | xii | ||
List of Abbreviations | xv | ||
1. | Summary | 16 | |
1.1. | Property Description | 16 | |
1.2. | Geology | 17 | |
1.3. | History | 17 | |
1.4. | Mineral Resource Estimate | 18 | |
1.5. | Mineral Reserve Estimate | 21 | |
1.6. | Cash Flow Analysis and Economics | 22 | |
1.7. | Conclusions | 23 | |
1.8. | Recommendations | 24 | |
2. | Introduction | 25 | |
2.1. | Terms of Reference and Purpose of this Technical Report | 25 | |
2.2. | Qualification of the Authors | 25 | |
2.3. | Sources of Information | 26 | |
2.4. | Units of Measure | 26 | |
2.5. | Coordinate Datum | 27 | |
3. | Reliance on Other Experts | 28 | |
4. | Property Descriptionand Location | 29 | |
4.1. | Property Description | 29 | |
4.2. | Property Location | 29 | |
4.3. | Status of Mineral Titles | 31 | |
4.4. | Location of Mineralization | 39 | |
5. | Accessibility, Climate, Vegetation, Physiography, Local Resources and Infrastructure | 41 | |
5.1. | Access to Project | 41 | |
5.2. | Climate | 41 |
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5.3. | Vegetation | 41 | ||
5.4. | Physiography | 41 | ||
5.5. | Local Resources and Infrastructure | 42 | ||
6. | History | 43 | ||
6.1. | Exploration History | 43 | ||
6.2. | Production History | 44 | ||
7. | Geological Setting and Mineralization | 45 | ||
7.1. | Regional Geology | 45 | ||
7.2. | Local Geology | 50 | ||
7.2.1. | Rock Units | 50 | ||
7.2.2. | Structure | 55 | ||
7.2.3. | Veins | 59 | ||
7.2.4. | Alteration | 60 | ||
7.2.5. | Mineralization | 63 | ||
8. | Deposit Types | 66 | ||
9. | Exploration | 68 | ||
9.1. | Historical Exploration | 68 | ||
9.2. | 2011 Drilling | 68 | ||
9.3. | 2012 Drilling | 70 | ||
9.4. | 2013 Drilling | 72 | ||
9.5. | 2014 Drilling | 74 | ||
10. | Drilling and Sampling Methodology | 77 | ||
10.1. | Collar Surveying | 79 | ||
10.1.1. | Surveying Surface Drill Collar Locations | 80 | ||
10.1.2. | Surveying Underground Drill Collar Locations | 83 | ||
10.1.3. | Locating Channel Samples | 83 | ||
10.2. | Downhole Surveying | 84 | ||
10.3. | Core Recovery | 84 | ||
10.4. | Security Procedures | 85 | ||
10.5. | Logging Drilled Core Observations | 85 |
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10.5.1. | Current Logging Protocol | 85 | ||
10.5.2. | Historic Logging Protocol | 86 | ||
10.5.3. | Re-logging Protocol for 2012- 2013 | 87 | ||
10.6. | Core Sampling Methodology | 89 | ||
10.7. | RC Sampling Methodology | 90 | ||
10.8. | Channel Sampling Methodology | 90 | ||
10.8.1. | Channel Sampling | 91 | ||
11. | Sample Preparation, Analysis, and Security | 93 | ||
11.1. | Historic Sample Preparation | 93 | ||
11.2. | Current Sample Preparation | 94 | ||
11.2.1. | Core Sample Preparation | 94 | ||
11.2.2. | Channel Sample Preparation | 94 | ||
11.3. | Sample Analysis Protocol | 95 | ||
11.3.1. | Historic Drill Sample Analysis | 95 | ||
11.3.2. | Drill Sample Analysis from 2012 through April 30, 2014 | 96 | ||
11.3.3. | Current Drill Sample Analysis | 96 | ||
11.3.4. | Channel Sample Analysis | 98 | ||
11.3.5. | Handling Analyses Results | 99 | ||
11.4. | Sample Security Measures | 99 | ||
11.5. | Quality Control Measures | 100 | ||
11.5.1. | QAQC Prior to 2012 | 101 | ||
11.5.2. | Current QAQC Procedures | 103 | ||
11.6. | QAQC Analysis | 104 | ||
11.6.1. | Duplicates Performance | 104 | ||
11.6.2. | Blank Assay Performance | 106 | ||
11.6.3. | Standards Performance | 114 | ||
11.7. | Opinion on the Adequacy of the Sampling Methodologies | 123 | ||
11.7.1. | Sampling Protocol Issues | 123 | ||
11.7.2. | Standards and Blanks Performance Issues | 124 | ||
12. | Data Verification | 125 |
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12.1. | Results of Drill Data Review | 125 | ||
12.1.1. | Collar Location Checks | 126 | ||
12.1.2. | Downhole Survey Checks | 126 | ||
12.1.3. | Geology Checks | 127 | ||
12.2. | Results of Channel Sample Data Review | 127 | ||
12.2.1. | Location Measurement Check | 127 | ||
12.2.2. | Geology Check | 127 | ||
12.2.3. | Assay Check | 127 | ||
12.3. | Summary of Database Verification | 128 | ||
13. | Mineral Processing and Metallurgical Testing | 129 | ||
13.1. | Early Test Work | 129 | ||
13.2. | 2013 Test Work | 129 | ||
13.3. | 2014 Test Work | 130 | ||
14. | Mineral Resource Estimates | 133 | ||
14.1. | Introduction | 133 | ||
14.2. | Database and Compositing | 133 | ||
14.2.1. | Assays | 133 | ||
14.2.2. | Lithology | 136 | ||
14.2.3. | Compositing | 138 | ||
14.3. | Geology and Vein Modelling | 138 | ||
14.4. | Density | 140 | ||
14.5. | Statistics | 140 | ||
14.6. | Grade Capping | 146 | ||
14.7. | Variography | 148 | ||
14.8. | Block Model | 151 | ||
14.9. | Grade Estimation | 151 | ||
14.10. | Mined Depletion and Sterilization | 153 | ||
14.11. | Model Validation | 155 | ||
14.12. | Mineral Resource Statement | 161 | ||
15. | Mineral Reserve Estimates | 165 |
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16. | Mining Methods | 168 | ||
16.1. | Mine Development | 168 | ||
16.1.1. | Access Development | 168 | ||
16.1.2. | Ground Support | 168 | ||
16.1.3. | Ventilation and Secondary Egress | 169 | ||
16.2. | Mining Methods | 169 | ||
16.2.1. | End Slice Stoping | 169 | ||
16.2.2. | Drift and Fill Stoping | 172 | ||
16.3. | Underground Labor | 173 | ||
16.4. | Mobile Equipment Fleet | 173 | ||
16.5. | Mine Plan | 174 | ||
17. | Recovery Methods | 178 | ||
17.1. | Mill Capacity and Process Facility Flow Diagram | 178 | ||
17.2. | Physical Mill Equipment | 182 | ||
17.3. | Operation and Recoveries | 185 | ||
17.4. | Tailings Storage Capacity | 185 | ||
17.5. | Processing Costs | 186 | ||
17.6. | Production | 186 | ||
17.7. | Midas Mill Operating Permits | 187 | ||
18. Project Infrastructure | 188 | |||
18.1. | Road Access | 188 | ||
18.2. | Power and Electrical Infrastructure | 188 | ||
18.3. | Water Management and Water Treatment | 188 | ||
18.4. | Communication Infrastructure | 189 | ||
18.5. | Site Infrastructure | 189 | ||
19. | Market Studies and Contracts | 191 | ||
19.1. | Precious Metal Markets | 191 | ||
19.2. | Contracts | 191 | ||
19.3. | Project Financing | 192 | ||
20. | Environmental Studies, Permitting and Social or Community Impact | 193 |
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20.1. | Environmental Compliance and Monitoring | 193 | ||
20.1.1. | Waste Rock Disposal Facility | 193 | ||
20.1.2. | Other Environmental Issues | 193 | ||
20.2. | Reclamation Bond Estimate | 193 | ||
20.3. | Major Permitting and Approvals | 194 | ||
20.4. | Future Permitting | 196 | ||
21. | Capital and Operating Costs | 197 | ||
21.1. | Capital Costs | 197 | ||
21.2. | Operating Costs and Cutoff Grade | 197 | ||
22. | Economic Analysis | 200 | ||
22.1. | Life of Mine Plan and Economics | 200 | ||
22.2. | Sensitivity Analysis | 202 | ||
22.3. | Adjusted Plan at $1,200 Gold | 205 | ||
23. | Other Relevant Data and Information | 206 | ||
24. | Interpretation and Conclusions | 207 | ||
24.1. | Conclusions | 207 | ||
24.2. | Project Risks | 207 | ||
25. | Recommendations | 208 | ||
26. | Bibliography | 210 | ||
27. | Glossary | 214 | ||
28. | Appendix A: Certification of Authors and Consent Forms | 222 |
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List of Tables
Table 1-1 | Chronology of Ownership of the Fire Creek Project | 17 |
Table 1-2 | Mineral Resource Statement as of December 31, 2014 | 19 |
Table 1-3 | Fire Creek Mineral Reserves as of December 31, 2014 | 22 |
Table 1-4 | Key Operating and After Tax Financial Statistics | 23 |
Table 2-1 | Qualified Professionals | 26 |
Table 2-2 | Units of Measure. | 27 |
Table 4-1 | Summary of Klondex Owned Unpatented Mining Claims | 33 |
Table 4-2 | Summary of Fee Land Holdings | 35 |
Table 4-3 | Summary of Leased Fee Land Holdings | 36 |
Table 4-4 | Summary of Fire Creek Project Holding Costs | 37 |
Table 6-1 | Exploration History | 43 |
Table 10-1 | Surface Drill Collars Re-surveyed by Klondex | 82 |
Table 11-1 | ALS In-house QAQC Datasets Reviewed | 101 |
Table 11-2 | Blank Assay Set Performance | 106 |
Table 11-3 | Standard Assay Performance | 114 |
Table 12-1 | Data Review Summary Drilled Material | 126 |
Table 13-1 | Summary of Cyanidation Test Results from 2011 Technical Report | 129 |
Table 13-2 | Combined Metallurgical Results, Gravity/Cyanidation Tests, 80% -212 νm Feed (Grav.), Reground to 80% - 75 νm (CN) | 130 |
Table 13-3 | Summary Metallurgical Results, Bottle Roll Tests, Fire Creek West Zone Drill CoreComposites | 131 |
Table 13-4 | Gold Metallurgical Results, Whole Mineralized Material Gravity Concentration with Cyanidation of the Gravity Cleaner and Rougher Tailings | 132 |
Table 13-5 | Silver Metallurgical Results, Whole Mineralized Material Gravity Concentration with Cyanidation of the Gravity Cleaner and Rougher Tailings | 132 |
Table 14-1 | Summary of Drill Hole and Channel Samples | 134 |
Table 14-2 | Lithology Codes | 137 |
Table 14-3 | Vein Drill Hole Statistics | 140 |
Table 14-4 | Vein Channel Sample Statistics | 141 |
Table 14-5 | Low Grade Drill Hole Statistics | 141 |
Table 14-6 | Low Grade Channel Sample Statistics | 141 |
Table 14-7 | Cap Grades for Measured Mineral Resources | 147 |
Table 14-8 | Cap Grades for Indicated and Inferred Mineral Resources | 147 |
Table 14-9 | Variogram Parameters for the Vonnie Vein | 149 |
Table 14-10 | Estimation Search Parameters by Resource Category | 152 |
Table 14-11 | Estimation Search Ellipsoids | 152 |
Table 14-12 | Estimation Method Mean Grade Comparison | 155 |
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Table 14-13 | Mineral Resource Cutoff Grade Parameters | 161 |
Table 14-14 | Mineral Resource Statement as of December 31, 2015 | 161 |
Table 15-1 | Mineral Reserves Cut Off Grade Calculation | 165 |
Table 15-2 | Fire Creek Mineral Reserves as of December 31, 2014 | 166 |
Table 16-1 | Underground Workforce 2015 and 2016 | 173 |
Table 16-2 | Underground Mobile Equipment | 173 |
Table 16-3 | Heading Productivity | 174 |
Table 16-4 | Annual Production and Development Plan | 176 |
Table 17-1 | Process Equipment Itemization by Area | 182 |
Table 17-2 | Midas Mill Operating Costs | 186 |
Table 17-3 | 2014 Fire Creek Mineralized Material Processed at the Midas Mill | 187 |
Table 19-1 | FNC Gold Delivery Schedule | 192 |
Table 20-1 | Fire Creek Project Significant Permits | 195 |
Table 21-1 | Capital Costs | 197 |
Table 21-2 | Underground Development Unit Costs | 197 |
Table 21-3 | Operating Costs | 197 |
Table 21-4 | Cut-off Grade Calculation | 198 |
Table 22-1 | Income Statement 2015 – 2018 ($000’s) | 201 |
Table 22-2 | Cash Flow Statement 2015 – 2019 ($000’s) | 201 |
Table 22-3 | Key Operating and After Tax Financial Statistics | 202 |
Table 22-4 | Key Operating and Financial Statistics for $1,200 Plan | 205 |
Table 24-1 | Potential Project Risks | 207 |
Table 25-1 | Recommendation Estimated Costs | 209 |
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List of Figures
Figure 1-1 | Fire Creek Project Overview Showing Vein Locations and Underground Development | 21 |
Figure 4-1 | Project Location Map | 30 |
Figure 4-2 | Klondex Land Position | 32 |
Figure 4-3 | Location of Fire Creek Project Relative to the Northern Nevada Rift System | 40 |
Figure 7-1 | Northern Nevada Rift in North-Central Nevada | 47 |
Figure 7-2 | Regional Geologic Map of the Northern Shoshone Range with the Project Underground Workings (purple) | 48 |
Figure 7-3 | Stratigraphic Sections for the Project and the Mule Canyon Mine with Tie Lines for Volcanic Packages | 49 |
Figure 7-4 | Geologic Map of the Fire Creek District | 50 |
Figure 7-5 | Example of Tbeq Basalt | 52 |
Figure 7-6 | Example of Tbma/Tlat Lithological Contact | 53 |
Figure 7-7 | Example of Tim Lithology | 55 |
Figure 7-8 | Fault Locations | 57 |
Figure 7-9 | Fault Block Model | 58 |
Figure 7-10 | Alteration Progression | 62 |
Figure 7-11 | Typical Argillic to Propylitic Alteration Progression Adjacent to the Karen Vein | 63 |
Figure 7-12 | Banded Vein Sample from the Vonnie Vein that Contains Large Clots of Native Gold | 64 |
Figure 7-13 | Picture of Split Core Sample Containing Dendritic Gold | 65 |
Figure 8-1 | Schematic Diagram of Low-Sulfidation Au, Ag Solutions in Relationship with Magma at Depth | 67 |
Figure 9-1 | Surface and Underground Holes Completed in 2011 | 69 |
Figure 9-2 | Surface and Underground Holes Completed in 2012 | 71 |
Figure 10-1 | Placing Core (January 2013) | 79 |
Figure 10-2 | Core Logging Facility, (January 2013) | 86 |
Figure 11-1 | Core Logging Facility (January 2013) | 100 |
Figure 11-2 | ALS Au Duplicates | 105 |
Figure 11-3 | AAL Au Dupicates | 105 |
Figure 11-4 | AAL Ag Duplicates | 106 |
Figure 11-5 | ALS AuBlank54 | 107 |
Figure 11-6 | AAL AuBlank54 | 108 |
Figure 11-7 | ALS FCBlank02 Au | 108 |
Figure 11-8 | AAL FCBlank02 Au | 109 |
Figure 11-9 | ALS FCBlank02 Ag | 109 |
Figure 11-10 | AAL FCBlank02 Ag | 110 |
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Figure 11-11 | AAL FCBlank03 Au | 110 |
Figure 11-12 | AAL FCBlank03 Ag | 111 |
Figure 11-13 | AAL FCBlan04 Au | 111 |
Figure 11-14 | AAL FCBlank04 Ag | 112 |
Figure 11-15 | ALS FCRDBLNK01 Au | 112 |
Figure 11-16 | ALS FCRDBLNK01 Ag | 113 |
Figure 11-17 | ALS FCOXBLNK01Au | 113 |
Figure 11-18 | ALS FCOXBLNK01 Ag | 114 |
Figure 11-19 | ALS FCRDLOW01 | 115 |
Figure 11-20 | AAL FCRDLOW01 | 116 |
Figure 11-21 | ALS OXN92 | 116 |
Figure 11-22 | AAL OXN92 | 117 |
Figure 11-23 | ALS OXP91 | 117 |
Figure 11-24 | ALS OXQ90 | 118 |
Figure 11-25 | AAL OXQ90 | 118 |
Figure 11-26 | ALS SG56 | 119 |
Figure 11-27 | ALS SN60 | 119 |
Figure 11-28 | AAL SN60 | 120 |
Figure 11-29 | ALS SP59 | 120 |
Figure 11-30 | AAL SP59 | 121 |
Figure 11-31 | ALS SQ48 | 121 |
Figure 11-32 | ALS SQ70 Au | 122 |
Figure 11-33 | ALS SQ70 Ag | 122 |
Figure 11-34 | ALS SQ83 | 123 |
Figure 14-1 | Drill Hole and Vein Locations | 135 |
Figure 14-2 | Channel Sample Locations Relative to the Veins | 136 |
Figure 14-3 | Long Section with Lithology | 137 |
Figure 14-4 | Long Section with Lithology and Tuff Models | 138 |
Figure 14-5 | Long Section with Lithology and Vein Models | 138 |
Figure 14-6 | Typical N75E Vein Model Cross Section | 140 |
Figure 14-7 | Gold Histogram and Cumulative Frequency for Vein Drill Hole Composites | 142 |
Figure 14-8 | Silver Histogram and Cumulative Frequency for Vein Drill Hole Composites | 143 |
Figure 14-9 | Gold Histogram and Cumulative Frequency for Vein Channel Sample Composites | 144 |
Figure 14-10 | Silver Histogram and Cumulative Frequency for Vein Channel Sample Composites | 145 |
Figure 14-11 | Vonnie Vein Gold Grade Cap | 148 |
Figure 14-12 | Vonnie Vein Silver Grade Cap | 148 |
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Figure 14-13 | Vonnie Vein Variogram (0,0) | 150 |
Figure 14-14 | Vonnie Vein Variogram (0,-90) | 150 |
Figure 14-15 | Vonnie Vein Mining Extent | 154 |
Figure 14-16 | Joyce Vein Mining Extent | 154 |
Figure 14-17 | Karen Vein Mining Extent | 155 |
Figure 14-18 | Legend Gold or Silver | 156 |
Figure 14-19 | Vonnie Vein Comparison of Composite and Estimated Block Gold Grade | 156 |
Figure 14-20 | Vonnie Vein Comparison of Composite and Estimated Block SilverGrade | 156 |
Figure 14-21 | Joyce Vein Comparison of Composite and Estimated Block Gold Grade | 157 |
Figure 14-22 | Joyce Vein Comparison of Composite and Estimated Block Silver Grade | 157 |
Figure 14-23 | Karen Vein Comparison of Composite and Estimated Block Gold Grade | 158 |
Figure 14-24 | Karen Vein Comparison of Composite and Estimated Block Silver Grade | 158 |
Figure 14-25 | Gold Swath Plot of the Vonnie Vein Along the North Axis | 159 |
Figure 14-26 | Gold Swath Plot of the Vonnie Vein Along the Z Axis | 159 |
Figure 14-27 | Silver Swath Plot of the Vonnie Vein Along the North Axis | 160 |
Figure 14-28 | Silver Swath Plot of the Vonnie Vein Along the Z Axis | 160 |
Figure 15-1 | Existing Workings, Reserve Excavations and Joyce, Vonnie and Karen Veins | 167 |
Figure 16-1 | Existing Development and Vein Traces at the 5400 Elevation | 168 |
Figure 16-2 | Long Section View of a Typical End Slice Stope | 170 |
Figure 16-3 | Cross Section Looking North Through the Joyce Vein and Vonnie Vein ShowingDrift and Fill Mining, Stope Development Drifting and Designed Stopes | 172 |
Figure 16-4 | Joyce Vein Long Section Looking East Showing Existing Mine Workings and Reserves Mine Plan | 175 |
Figure 16-5 | Vonnie Vein Long Section Looking East Showing Existing Mine Workings and Reserves Mine Plan | 175 |
Figure 16-6 | Karen Vein Long Section Looking East Showing Existing Mine Workings and Reserves Mine Plan | 176 |
Figure 17-1 | Process Facility Flow Sheet (Klondex, 2015) | 181 |
Figure 18-1 | Site Facilities | 190 |
Figure 19-1 | Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average | 191 |
Figure 21-1 | Cutoff Grade Sensitivity to Gold Price | 199 |
Figure 22-1 | 5% NPV Sensitivity | 203 |
Figure 22-2 | 10% NPV Sensitivity | 204 |
Figure 22-3 | Profitability Index Sensitivity | 204 |
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List of Abbreviations
A | Ampere | kA | kiloamperes |
AA | atomic absorption | kCFM | thousand cubic feet per minute |
A/m2 | amperes per square meter | Kg | Kilograms |
AGP | Acid Generation Potential | km | kilometer |
Ag | Silver | km2 | square kilometer |
ANFO | ammonium nitrate fuel oil | kWh/t | kilowatt-hour per ton |
ANP | Acid Neutralization Potential | LOI | Loss On Ignition |
Au | Gold | LoM | Life-of-Mine |
AuEq | gold equivalent | m | meter |
btu | British Thermal Unit | m2 | square meter |
°C | degrees Celsius | m3 | cubic meter |
CCD | counter-current decantation | masl | meters above sea level |
CIL | carbon-in-leach | mg/L | milligrams/liter |
CoG | cut-off grade | mm | millimeter |
cm | centimeter | mm2 | square millimeter |
cm2 | square centimeter | mm3 | cubic millimeter |
cm3 | cubic centimeter | MME | Mine & Mill Engineering |
cfm | cubic feet per minute | Moz | million troy ounces |
ConfC | confidence code | Mt | million tonnes |
CRec | core recovery | MTW | measured true width |
CSS | closed-side setting | MW | million watts |
CTW | calculated true width | m.y. | million years |
° | degree (degrees) | NGO | non-governmental organization |
dia. | diameter | NI 43-101 | Canadian National Instrument 43-101 |
EIS | Environmental Impact Statement | oz | Troy Ounce |
EMP | Environmental Management Plan | opt | Troy Ounce per short ton |
FA | fire assay | % | percent |
Ft | Foot | PLC | Programmable Logic Controller |
Ft2 | Square foot | PLS | Pregnant Leach Solution |
Ft3 | Cubic foot | PMF | probable maximum flood |
g | Gram | POO | Plan of Operations |
g/L | gram per liter | ppb | parts per billion |
g-mol | gram-mole | ppm | parts per million |
g/t | grams per tonne | QAQC | Quality Assurance/Quality Control |
ha | hectares | RC | reverse circulation drilling |
HDPE | Height Density Polyethylene | ROM | Run-of-Mine |
HTW | horizontal true width | RQD | Rock Quality Description |
ICP | induced couple plasma | SEC | U.S. Securities & Exchange Commission |
ID2 | inverse-distance squared | Sec | second |
ID3 | inverse-distance cubed | SG | specific gravity |
ILS | Intermediate Leach Solution | SPT | Standard penetration test |
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1. | Summary |
Practical Mining LLC was engaged by Klondex Gold & Silver Mining Company (KGS), U.S. subsidiary of Canadian based Klondex Mines Ltd. (Klondex or the Company), to prepare a Prefeasibility Study (PFS) in accordance with National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. Practical Mining’s evaluation of the Fire Creek Project (Fire Creek or the Project), located in Lander County, Nevada, is presented herein. This report, dated the 16th day of March, 2015, with an effective date of December 31, 2014, updates the previous mineral resource estimate (Odell et al., 2014) and sets out the first mineral reserve estimate for the Project. The lineage of technical reports on Fire Creek as is continued in this Technical Report includes:
1. | 2006 NI 43-101 Technical Report: Ullmer, Edwin, and Hawthorn, Gary W., 2006: Fire Creek Gold Property, Lander Co., Nevada, September 15, 2006; |
2. | 2009 NI 43-101 Technical Report: Updated Report on the Fire Creek Gold Property, Lander Co., Nevada, March 30, 2009; |
3. | 2011 NI 43-101 Technical Report: Raven, Wesley, Ullmer, Edwin, Hawthorn, Gary W., 2011: Updated Technical Report and Resource Estimation on the Fire Creek Gold Property, Lander County, Nevada; |
4. | 2013 NI 43-101 Technical Report: Odell, Mark, Swanson, Karl and White, Michele, 2013: Technical Report (Amended), Fire Creek Exploration Project, Lander County, Nevada; and |
5. | 2014 NI 43-101 Technical Report: Odell, Mark, Swanson, Karl, Symmes, Laura and Bull, Sarah, 2014: Preliminary Economic Assessment of the Fire Creek Project, Lander County, Nevada, Amended. |
The Project is located in Lander County in north-central Nevada about 16 miles south of Interstate Highway I-80. The Project is centered on latitude 46.4627° North (N) and longitude 116.6518° West (W), (529639 E, 4478991 N - Universal Transverse Mercator (meters), North American Datum of 1983). Most of the current exploration is located within Sections 15 and 22 of Township 30 North, Range 47 East. Klondex controls approximately 17,000 acres (26.5 square miles) which includes 831 unpatented lode mining claims; 1,114 acres of private fee land; and 229 acres of leased fee lands.
1.1. Property Description
The Project is located primarily in Lander County, Nevada and to a lesser extent in Eureka County, Nevada, approximately 63 miles west of Elko, Nevada. The Project comprises private fee lands (both leased and owned) and unpatented lode mining claims. The land position includes approximately 15,420 acres of unpatented federal lode mining claims, 1,110 acres of private fee land and 230 acres of mineral leases. Overall, the Fire Creek land package is approximately 17,000 acres.
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1.2. Geology
The deposit is an epithermal deposit vertically-zoned within high-angle northwest striking structures, low-sulfidation, hosted in a mid-Miocene basalt package. Gold mineralization occurs in two habits: shallow structurally-controlled gold in variably altered Tertiary basalt and primarily native gold steeply dipping quartz-calcite veins or structures. A package of middle-Miocene basalt and basaltic andesite flows package has been cut by high-angle normal faults related to both Northern Nevada Rift (NNR) and Basin and Range extension that form grabens and half-grabens which are the structural controls for in the district.
High-grade gold mineralization has been delineated between approximately 4,900 feet and 5,700 feet AMSL and is open up and down dip as well as on strike. Lower-grade gold mineralization occurs from the surface and mineralization is open at depth. Vein textures, gangue minerals, and alteration ore typical of low-sulfidation systems. Widespread propylitic alteration grades to argillic alteration proximal to veins and/or other structural fluid conduits. Elevated content is often spatially associated with the argillic alteration zone. Gold mineralization often occurs along discrete horizons within veins. An opaline silica cap is discontinuously preserved above the deeper mineralization. Mineralized faults near the opaline silica were targeted by early prospecting and later shallow drilling by previous operators in the 1980’s.
1.3. History
Table 1-1 Chronology of Ownership of the Fire Creek Project
Dates | Company | Details |
1967 | Union Pacific Resources | Drilled two core holes. |
1974 to 1975 | Placer Development Ltd. | Drilled 22 rotary holes. |
1975 | Klondex Mines Ltd. | Acquired the property. 1980- 1983 drilled 64 rotary holes. 1981 gold test production. |
1984 | Minex Resources, Inc. | Leased the property from Klondex, drilled 13 rotary holes. |
1986 to 1987 | Alma American Mining Company (“Alma”) | Leased the property from Klondex, drilled 64 rotary holes. |
1988 | Aurenco Joint Venture (“Aurenco JV”) | Aurenco JV formed between Black Beauty Mining and Covenanter Mining. |
1988 to 1990 | Aurenco JV | Leased the property from Klondex. |
1990 to 1995 | Klondex Mines Ltd. | No activity. |
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Dates | Company | Details |
1995 to 1996 | North Mining Inc. (“North Mining”) | Leased the property from Klondex. Drilled 67 holes, performed IP and HEM surveys. |
1996 to 2004 | Klondex Mines Ltd. | No activity. |
2004 to 2012 | Klondex Mines Ltd. | Began a deep exploration program. Development commenced in 2011. |
2012 to Present | Klondex Mines Ltd. | New Management and Board of Directors in 2012, ongoing exploration. |
Recent drill programs conducted by Klondex have defined two major north-northwest striking vein arrays, each comprised of several en echelon veins. Several new target areas outside of the known vein arrays have been defined by both gradient-array and dipole-dipole induced polarization surveys.
This Technical Report (Report) updates the Project mineral resource estimate and provides an initial estimate of mineral reserves. This report incorporates the technical information available through December 31, 2014, which is the effective date of this Report.
1.4. Mineral Resource Estimate
The mineral resource estimate is based on data from 475 surface and underground drill holes, through December 31, 2014. This estimate also includes 1,457 independently assayed rib, back, and face channel samples from underground drifting on the Joyce Vein, Vonnie Vein, and Karen Vein.
Wire frame models were constructed for 47 vein shoots that strike approximately N15°W and dip steeply to both the east and west. The vein models were compiled from cross sections constructed on 25-foot intervals. Assay values were composited into 10-foot lengths and truncated at the vein hanging wall and footwall. Only composites flagged as representing vein material were used in the grade estimation. A grade capping scheme based on resource category and vein was employed. Grades were assigned to individual blocks using the Inverse Distance Cubed method (ID3).
The 47 veins were each assigned a specific variography based on their respective approximate orientation. Measured blocks require a minimum of four drill hole intercepts or channel samples within an average anisotropic search radius of 25 feet. Indicated blocks required three drill hole intercepts within 100 feet. Inferred blocks required two composite samples within 300 feet. Grades were estimated only for blocks contained within the modeled veins. Block extents are five feet long along strike and oriented vertically. Perpendicular to strike, the block extents were limited to the width of the vein with 0.2 to five-foot resolution. This method allows veins as narrow as 0.2 foot to be modeled precisely.
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Mineral resources were estimated for only blocks within the modeled vein wireframes. Low grade mineralization immediately adjacent to the veins was also modeled from the vein hanging wall or footwall out to a cut off of 0.1 Au ounces per ton (opt). In all cases, the vein boundary with the low grade mineralization was treated as a “hard” boundary, and composite assay data from the vein was not used to estimate the low grade mineralization.
The mineralized vein arrays extend over 5,000 feet along strike and from near surface to 1,000 feet in depth. These vein arrays are open both along strike and down depth.
A density of 0.0774 tons per cubic foot was used for all veins. This value was derived from 15 samples collected from the Joyce Vein and Vonnie Vein and analyzed by SGS North America, Inc. (SGS) of Elko, Nevada; an independent laboratory. The SGS (Elko) laboratory forms part of the SGS Minerals' global group of laboratories. The SGS (Elko) laboratory is not independently certified by a standards association but is associated with the SGS (Vancouver) laboratory, which is an ISO 9001:2008 accredited facility.
Table 1-2 Mineral Resource Statement as of December 31, 2014
Mass | Grade (opt) | Contained Metal (koz.) | |||||
(kton) | Au | Ag | AuEq | Au | Ag | AuEq | |
Measured | |||||||
Main | 26.2 | 2.83 | 2.04 | 2.87 | 74.2 | 53.3 | 75.0 |
West | 11.9 | 1.32 | 1.03 | 1.34 | 15.7 | 12.3 | 15.9 |
North | 53.3 | 2.28 | 1.77 | 2.31 | 121.6 | 94.4 | 123.1 |
South | 1.6 | 0.24 | 0.46 | 0.25 | 0.4 | 0.7 | 0.4 |
Far North | |||||||
Total Measured | 93.0 | 2.28 | 1.73 | 2.31 | 212.0 | 160.8 | 214.5 |
Indicated | |||||||
Main | 101.9 | 0.95 | 0.79 | 0.97 | 97.2 | 80.8 | 98.5 |
West | 10.8 | 0.38 | 0.23 | 0.39 | 4.2 | 2.5 | 4.2 |
North | 109.2 | 0.74 | 0.46 | 0.74 | 80.3 | 50.5 | 81.0 |
South | 43.7 | 0.36 | 0.57 | 0.37 | 15.9 | 24.9 | 16.3 |
Far North | 18.8 | 0.32 | 0.25 | 0.32 | 6.0 | 4.6 | 6.1 |
Total Indicated | 284.4 | 0.72 | 0.57 | 0.72 | 203.5 | 163.4 | 206.1 |
Measured and Indicated | |||||||
Main | 128.0 | 1.34 | 1.05 | 1.36 | 171.4 | 134.2 | 173.5 |
West | 22.7 | 0.88 | 0.65 | 0.89 | 19.9 | 14.8 | 20.1 |
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Mass | Grade (opt) | Contained Metal (koz.) | |||||
(kton) | Au | Ag | AuEq | Au | Ag | AuEq | |
North | 162.5 | 1.24 | 0.89 | 1.26 | 201.9 | 144.9 | 204.1 |
South | 45.3 | 0.36 | 0.57 | 0.37 | 16.3 | 25.7 | 16.7 |
Far North | 18.8 | 0.32 | 0.25 | 0.32 | 6.0 | 4.6 | 6.1 |
Total Measured & Indicated | 377.4 | 1.10 | 0.86 | 1.11 | 415.5 | 324.2 | 420.5 |
Inferred | |||||||
Main | 88.4 | 0.44 | 0.43 | 0.44 | 38.5 | 37.8 | 39.1 |
West | 93.5 | 0.39 | 0.29 | 0.39 | 36.5 | 27.4 | 36.9 |
North | 276.5 | 0.53 | 0.53 | 0.54 | 146.3 | 147.6 | 148.6 |
South | 11.6 | 0.56 | 0.27 | 0.56 | 6.5 | 3.2 | 6.5 |
Far North | 370.1 | 0.35 | 0.28 | 0.36 | 130.5 | 104.9 | 132.2 |
Total Inferred | 840.0 | 0.43 | 0.38 | 0.43 | 358.3 | 320.8 | 363.3 |
Notes: | ||
1. | Mineral resources have been calculated at a gold price of $1,200/troy ounce and a silver price of $19.00 per troy ounce; | |
2. | Mineral resources are calculated at a grade thickness cut-off grade of 1.256 Au equivalent opt-feet and a diluted Au equivalent cut-off grade of 0.256 opt; | |
3. | Gold equivalent ounces were calculated based on one ounce of gold being equivalent to 64.53 ounces of silver; | |
4. | The minimum mining width is defined as four feet or the vein true thickness plus one foot, whichever is greater; | |
5. | Mineral resources include dilution to achieve mining widths and an additional 10% unplanned dilution; | |
6. | Mineral resources include allowance for 5% mining losses; | |
7. | Mineral resources are inclusive of mineral reserves; | |
8. | Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant issues; | |
9. | The quantity and grade of reported inferred mineral resources in this estimation are uncertain in nature and there is insufficient exploration to define these inferred mineral resources as an indicated or measured mineral resource and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category, and;. | |
10. | Mineral resource estimates can be materially affected by environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other factors. |
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1.5. Mineral Reserve Estimate
Excavation designs for stopes, stope development drifting, and access development were created using Vulcan software. Stope designs were aided by the Vulcan Stope Optimizer Module. The stope optimizer produces the stope cross section which maximizes value within given geometric and constraints.
Design constraints included a four-foot minimum mining width for long-hole stopes with development drifts spaced at 40-foot vertical intervals. Stope development drift dimensions maintained a constant height of 10 feet and a minimum width of six feet. Drift and fill dimensions are the same as those for stope development.
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Table 1-3 Fire Creek Mineral Reserves as of December 31, 2014
Au | Ag | Au Equiv. | |||||
Tons | Au Eq | Ounces | Ounces | Ounces | |||
Vein Designation | (000's) | Au opt | Ag opt | opt | (000's) | (000's) | (000's) |
Proven Reserves | |||||||
Joyce | 31 | 0.914 | 0.678 | 0.924 | 27.9 | 20.7 | 28.2 |
Vonnie | 9.3 | 3.301 | 2.151 | 3.335 | 30.6 | 19.9 | 30.9 |
Karen | 41 | 1.454 | 1.192 | 1.473 | 59.7 | 49.0 | 60.5 |
Proven Reserves | 80.9 | 1.462 | 1.108 | 1.479 | 118.2 | 89.6 | 119.6 |
Probable Reserves | |||||||
Joyce | 60 | 0.779 | 0.357 | 0.784 | 47.0 | 21.5 | 47.3 |
Vonnie | 34 | 1.920 | 1.626 | 1.945 | 66.1 | 56.0 | 67.0 |
Karen | 10 | 0.733 | 0.500 | 0.741 | 7.4 | 5.0 | 7.5 |
Probable Reserves | 104.9 | 1.149 | 0.787 | 1.161 | 120.5 | 82.6 | 121.8 |
Proven + Probable Reserves | |||||||
Joyce | 91 | 0.824 | 0.464 | 0.831 | 74.9 | 42.2 | 75.6 |
Vonnie | 44 | 2.213 | 1.738 | 2.240 | 96.7 | 75.9 | 97.9 |
Karen | 51 | 1.312 | 1.056 | 1.328 | 67.1 | 54.0 | 68.0 |
Proven + Probable Reserves | 185.8 | 1.285 | 0.927 | 1.300 | 238.7 | 172.2 | 241.4 |
Notes: | ||
1. | Reserves have been estimated with a gold price of $1,000/ounce and a silver price of $15.83/ounce; | |
2. | Metallurgical recoveries for gold and silver are 94% and 92% respectively; | |
3. | Gold equivalent ounces are calculated based on one ounce of gold being equivalent to 64.53 ounces of silver; | |
4. | Mineral Reserves are estimated at a cutoff grade of 0.494 Au opt and an incremental cutoff grade of 0.259 Au opt, and; | |
5. | Mine losses of 5% and unplanned mining dilution of 10% have been applied to the designed mine excavations. |
1.6. Cash Flow Analysis and Economics
The reserves mine plan was evaluated using constant dollar cash flow analysis, and the results are summarized in Table 1-4. The grade value of the resources and the low capital requirements facilitated with the addition of the Midas Mine and Mill to Klondex’s project portfolio, combine to produce a short 0.5 year capital payback period and a relatively high 5.0 profitability index (PI) calculated at a 10% discount rate.
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Table 1-4 Key Operating and After Tax Financial Statistics
Material Mined and Processed (kt) | 186 |
Avg. Gold Grade (opt) | 1.327 |
Avg. Silver Grade (opt) | 0.96 |
Contained Gold (koz) | 237.0 |
Contained Silver (koz) | 171 |
Avg. Gold Metallurgical Recovery | 94% |
Avg. Silver Metallurgical Recovery | 92% |
Recovered Gold (koz) | 224.4 |
Recovered Silver (koz) | 158 |
Reserve Life (years) | 3.8 |
Operating Cost ($/ton) | $460 |
Cash Cost ($/oz) 1. | $410 |
Total Cost ($/oz) 1. | $492 |
Gold Price ($/oz) | $1,000.00 |
Silver Price ($/oz) | $15.83 |
Capital Costs ($ Millions) | $18.4 |
Payback Period (Years) | 0.5 |
Cash Flow ($ Millions) | $85.80 |
5% Discounted Cash Flow ($ Millions) | $78.10 |
10% Discounted Cash Flow ($ Millions) | $71.40 |
Profitability Index (10%) 2. | 5.0 |
Internal Rate of Return | NA |
Notes; | ||
1. | Net of Byproduct Sales, and; | |
2. | Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates break even. |
1.7. Conclusions
The Project is a modern, mechanized narrow vein mine. Only the mineralized veins accessible from main development have been defined to a sufficient level of detail to categorize as reserves. In the opinion of the authors of this Technical Report, additional potential exists to extend reserves along strike in both directions as underground access is developed. As the footprint of the mine grows and the number of available mining areas grows with it, the mining rate can be increased, and cost reductions realized through economies of scale.
The conventional Merrill Crowe mill facility of the Midas Mine is an efficient well maintained modern mineral processing plant capable of processing 1,200 tons per day (tpd). The plant is able to operate with a minimum crew which results in cost reductions when operated at capacity. The underutilized processing capacity can accept increased mine production from the Project or the Midas Mine as well as third party processing agreements.
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Capital requirements for the Project are minimal. Ongoing mine development comprises the majority of capital costs, and the ability to access multiple veins from common development greatly reduces the unit cost per ounce.
In the opinion of the authors of this Technical Report, the high grade reserves in the mine plan provide a high return and will sustain profitable operations with up to 40% adverse variations in metal prices, operating or capital costs. The total cost per ounce, including capital expenditures and net of byproduct sales, is less than $500 per ounce.
1.8. Recommendations
Exploration: Underground drilling should continue in the veins identified near the current development workings to increase the level of confidence in these veins to an indicated classification. The decline should be advanced to provide an underground drill platform from which to drill the veins in the North and Far North Zones. While the decline is being advanced, additional drilling in this area can be completed from surface to refine the vein targets.
Delineation: Rib sampling has limited value and should continue to be supplemented by drilling shallow ten to 20-foot deep holes into the rib with the “Termite” drill or hand held drills and sample the drill cuttings. This sampling method will add a third dimension to the potential wall rock mineralization.
Stope Planning: Complete the drift and fill stopes currently underway. Set up new areas for long hole stoping. The use of short probe holes (i.e. “Termite” holes) discussed above should provide the planning engineers enough detail to efficiently design stopes with minimal loss of mineralization.
Rapid Infiltration Basin Commissioning: In order to reduce delays caused by intercepting perched water, the Rapid Infiltration Basins (RIBs) and water handling systems need to be functioning at capacity.
Geologic Database Administration: All of the Project data collected to date, including drill samples, channel samples and Quality Assurance and Quality Control (QA/QC) samples, need to be stored and archived in a permanent and indelible manner. The system software for this system has been procured, but a full time database administrator has not been selected.
Quality Assurance and Quality Control: Timely follow-up of QAQC assay deviations and re-assay requests needs to be aggressively pursued and should become an automated process once the database software is fully implemented.
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2. | Introduction |
2.1. Terms of Reference and Purpose of this Technical Report
This Report provides Klondex a preliminary feasibility study of the Project. This evaluation includes measured, indicated, and inferred resources, as well as proven and probable mineral reserves. This Report was prepared in accordance with the requirements of NI 43-101 and Form 43-101F1 (43-101F1) for technical reports.
Mineral resource and mineral reserve definitions are set forth in Section 27 of this Report in accordance with the companion policy to NI 43-101 (43-101CP) of the Canadian Securities Administrators and “Canadian Institute of Mining, Metallurgy and Petroleum (CIM) – Definition Standards for Mineral Resources and Mineral Reserves adopted by CIM Council on May 10, 2014.”
2.2. Qualification of the Authors
This Report includes technical evaluations from four independent consultants. The consultants are specialists in the fields of geology, geological engineering, exploration, and open pit and underground mining.
None of the authors has any beneficial interest in Klondex or any of its subsidiaries or in the assets of Klondex or any of its subsidiaries. The authors will be paid a fee for this work in accordance with normal professional consulting practices.
The individuals who have provided input to the current Report are cited as “author” and are listed below with the dates on which they visited the Project. These authors have extensive experience in the mining industry and are members in good standing of appropriate professional institutions.
Mr. Odell has visited the Project on several occasions, the most recent on January 12, 2015;
Ms. Symmes visited the Project on several occasions, the most recent on September 18, 2014;
Ms. Bull visited the Project on September 18, 2014, and;
Mr. Swanson visited the Project on January 19 through 22, 2015.
Mr. Odell is the qualified person (QP) for this Technical Report and is cited as “primary author.”
Mr. Odell, Ms. Symmes and Mr. Swanson inspected the underground mining operations and reviewed the site geology. Ms. Symmes was responsible for reviewing the core drilling and sampling procedures, core handling and security procedures, data management, and the geology.
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Ms. Bull was responsible for reviewing the underground mine design. In addition, Ms. Symmes and Mr. Odell reviewed reports detailing the Company’s land position at the Project. The key personnel contributing to this report are listed in Table 2-1. The Certificate and Consent Forms are provided in Appendix A: Certification of Authors and Consent Forms.
Table 2-1 Qualified Professionals
Company | Name | Title | Discipline |
Practical Mining, LLC | Mark Odell | Primary Author & Mining Engineer | Mining and mineral resources |
Practical Mining, LLC | Laura Symmes | Sr. Geologist | Geology |
Practical Mining, LLC | Sarah Bull | Mining Engineer | Mining |
Independent Consultant | Karl Swanson | Consulting Geologist | Resource model and geology |
2.3. Sources of Information
The sources of information include data and reports supplied by Klondex staff.
Additional information is included in the Report which is based on discussions with Klondex staff as it relates to their field of expertise at the Project. The required financial data and operating statistics were also provided by Klondex staff.
Information sources are documented either within the text and cited in references, or are cited in references only. The primary author believes the information provided by Klondex staff to be accurate based on their work at the Project. The authors asked detailed questions of specific Klondex staff to help verify contributions included in this document. These contributions are clearly stated within the text.
2.4. Units of Measure
The units of measure used in this report are shown in Table 2-2 below. U.S. Imperial units of measure are used throughout this document unless otherwise noted. The glossary of geological and mining related terms is also provided in Section 27 of this Technical Report.
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Table 2-2 Units of Measure.
US Imperial to Metric conversions |
Linear Measure |
1 inch = 2.54 cm |
1 foot = 0.3048 m |
1 yard = 0.9144 m |
1 mile = 1.6 km |
Area Measure |
1 acre = 0.4047 ha |
1 square mile = 640 acres = 259 ha |
Weight |
1 short ton (st) = 2,000 lbs = 0.9071 metric tons |
1 lb = 0.454 kg = 14.5833 troy oz |
Assay Values |
1 oz per short ton = 34.2857 g/t |
1 troy oz = 31.1036 g |
1 part per billion = 0.0000292 oz/ton |
1 part per million = 0.0292 oz/ton = 1g/t |
2.5. Coordinate Datum
Spatial data utilized in analysis presented in this Technical Report are projected to Nevada State Plane Central Zone North American Datum 1983 (NV SPCS) feet truncated to the last six digits. All spatial measurement units used in the mineral resource estimate are U.S. Imperial feet, and currency is expressed in United States dollars unless stated otherwise.
Survey data was collected and reported using several coordinate systems. Historically, survey data was originally collected in North American Datum of 1983 (NAD83) meters as a default in the instrumentation settings, and then the data was converted to NAD83 feet for reports as requested by Klondex staff in Nevada. Klondex’s Nevada staff further converted the data from NAD83 feet to UTM NAD27 Zone 11N feet. Early in 2014, all the Project data was again converted to NV SPCS NAD83 coordinates.
In addition, downhole surveys were collected without compensating for magnetic declination. Klondex staff applied corrections to raw downhole survey data to compensate for the local declination at the Project, which is 13.35 degrees according to the National Oceanic and Atmospheric Administration (NOAA) calculator.
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3. | Reliance on Other Experts |
The technical status for the claims and land holding is reliant on information provided by Klondex’s legal counsel. The status of the Klondex environmental program and the permitting process were provided by Ms. Lucy Hill, Klondex Environmental Manager. These contributions are presented based on the assumption that they are accurate portrayals of the environmental condition of the Project at the time of writing this Report.
Observations made at the Project by the authors include overseeing all aspects of mining activities including: underground core drilling, labeling core boxes, moving core, splitting core, safety procedures, haulage and equipment maintenance, water treatment, security, road maintenance, general geology, and character of mineralization. A review of historical databases and sampling protocol was performed by author Laura Symmes for the purpose of validating data integrity.
The authors reviewed land tenure to verify the nature of the good standing with regulatory authorities and the Bureau of Land Management (BLM) of Klondex’s unpatented lode mining claims and a title opinion report dated July 30, 2014, written by Erwin & Thompson LLP. The legal status or ownership of the fee properties and/or any agreements that pertain to the mineral Fire Creek mineral deposit as described in Section 4 were provided by Klondex legal counsel for all relevant mining claims. Assumptions made as to accuracy of land tenure are based on the Erwin & Thompson LLP legal opinion.
The opinions expressed in this Report are based on the authors’ field observations and assessment of the technical data supplied by Klondex.
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4. | Property Description and Location |
4.1. Property Description
The Project is located primarily in Lander County, Nevada and to a lesser extent in Eureka County, Nevada, approximately 63 miles west of the major city of Elko, Nevada, USA in a sage and grass covered weathered basalt hillside overlooking Crescent Valley. There are multiple small towns along paved highways within a short commute of the Project, and the northern edge of the residential area of the town of Crescent Valley abuts the main access road. The Project’s land coverage is approximately 17,000 acres.
4.2. Property Location
The Project is located in Lander County, Nevada, approximately 34 miles west of Carlin (63 miles west of Elko) and 16 miles south of Interstate Highway I-80. Figure 4-1 shows the location of the Project. The closest town to the Project is Crescent Valley on Nevada State Highway 306. Access from Elko takes approximately one hour.
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4.3. Status of Mineral Titles
The Project comprises private fee lands (both leased and owned) and unpatented lode mining claims. Figure 4-2 depicts the current land status. The land position shown on Figure 4-2 includes approximately 15,421 acres of unpatented federal lode mining claims, 1,114 acres of private fee land, 229 acres of mineral leases. Overall, the Fire Creek land package is approximately 17,000 acres.
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Table 4-1 lists the 831 unpatented lode mining claims held by Klondex for the Project. Table 4-2 itemizes fee lands owned by KGS, and Table 4-3 itemizes fee lands leased by KGS. Unpatented claims are in current good standing through September, 2015. Leases are in good standing until the lease payment is due.
Table 4-1 Summary of Klondex Owned Unpatented Mining Claims
Number | |||||
Claim Name | Section | Township | Range | Location | of |
Date | Claims | ||||
Wood Tick 2, 4, 6, 8, 10, 12, 14, 16, 18, 20, 22 | 2 | 30N | 47E | 18-Jul-87 | 11 |
Wood Tick 24, 26, 28, 30, 32, 34, 36 | 2 | 30N | 47E | 18-Jul-87 | 7 |
Wood Tick 38, 40, 42, 44, 46, 48, 50, 52 | 36 | 31N | 47E | 21-Jul-87 | 8 |
G 1 - G 16 | 26 | 30N | 47E | 23-Jan-90 | 16 |
Deb #2, #4 | 34 | 30N | 47E | 13-Dec-91 | 2 |
Revenge 2, 20 | 34 | 30N | 47E | 16-Dec-91 | 2 |
Revenge 4, 6, 8 | 34 | 30N | 47E | 17-Dec-91 | 3 |
Revenge 10, 12, 14 | 34 | 30N | 47E | 18-Dec-91 | 3 |
Revenge 22 | 34 | 30N | 47E | 9-Jan-92 | 1 |
Revenge 28 | 34 | 30N | 47E | 26- Jan-92 | 1 |
Revenge 16, 18 | 34 | 30N | 47E | 6-Feb-92 | 2 |
Revenge 24, 26 | 34 | 30N | 47E | 13- Feb-92 | 2 |
K 1 - 20 1 | 16 | 30N | 47E | 25-Jun-92 | 20 |
K 21 - 27 2 | 16 | 30N | 47E | 26-Jun-92 | 7 |
Alan 1-14 | 31 | 30N | 47E | 15- Feb-93 | 14 |
N 2, 4, 6, 8, 10,12, 14, 16, 18 | 32 | 30N | 47E | 17-Nov-93 | 9 |
N 20, 22, 24, 26, 28, 30 | 32 | 30N | 47E | 18- Nov-93 | 6 |
TL 2, 4, 6 | 20 | 30N | 47E | 8-Nov-93 | 3 |
TL 8, 10, 12, 14, 16, 18 | 20 | 30N | 47E | 10-Nov-93 | 6 |
TL 20, 22, 24, 26 | 20 | 30N | 47E | 21-Jun-94 | 4 |
FCRA 1 - 20 | 26 | 30N | 47E | 28-Sep-95 | 20 |
T 1 - 10 | 14 | 30N | 47E | 13-Oct-91 | 10 |
T 11 - 18, 27 - 36 | 14 | 30N | 47E | 24-Sep-03 | 18 |
T 19, 21 - 26 | 14 | 30N | 47E | 23-Sep-03 | 7 |
T 20 | 10, 14 | 30N | 47E | 23- Sep-03 | 1 |
Hondo, 1, 3, 5, 7, 9, 11, 13, 15, 18, 20, 22, 24, 26, 28, 30, 32, 157, 158 | 24 | 30N | 47E | 20- Sep-03 | 18 |
Deb 1, 3, 5 | 34 | 30N | 47E | 22-Sep-03 | 3 |
Revenge 1, 11, 13, 15, 17, 19, 21, 23, 25, 27 | 34 | 30N | 47E | 22-Sep-03 | 10 |
Revenge 3, 5, 7, 9, 29, 30, 31 | 34 | 30N | 47E | 23-Sep-03 | 7 |
FC 1-16, 18 | 36 | 30N | 47E | 21- Sep-03 | 17 |
FC 17 | 26, 36 | 30N | 47E | 21-Sep-03 | 1 |
What If 29 - 37 | 36 | 30N | 47E | 21-Sep-03 | 9 |
FC 38 - 46 3 | 36 | 30N | 47E | 21-Sep-03 | 9 |
T 38 - 60 | 10 | 30N | 47E | 5-Oct-06 | 23 |
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Number | |||||
Claim Name | Section | Township | Range | Location | of |
Date | Claims | ||||
T 61 - 71 | 10 | 30N | 47E | 6-Oct-06 | 11 |
T 72 | 2, 10 | 30N | 47E | 6-Oct-03 | 1 |
FCXX 1, 2 | 15, 22 | 30N | 47E | 24-Nov-03 | 2 |
FCXX 3- 40 | 22 | 30N | 47E | 24-Nov-03 | 38 |
CH 1 - 18 | 30 | 30N | 47E | 19- Sep-06 | 18 |
Hondo 2, 4, 6, 8, 10, 12, 14, 16, 17, 19, 21, 23, | 24 | 30N | 47E | 4-Oct-06 | 17 |
25, 27, 29, 31, 156 Hondo 155 | 24 | 30N | 47E | 4-Oct-06 | 1 |
N 1, 3, 11, 13, 19, 21, 23, 25, 27 | 32 | 30N | 47E | 11-Sep-06 | 9 |
N 5, 7, 9, 15,17, 29, 31 | 32 | 30N | 47E | 12-Sep-06 | 7 |
TL 1, 3, 5, 7, 9, 11, 13, 15, 17 | 20 | 30N | 47E | 13- Sep-06 | 9 |
TL 19, 21, 23, 25, 27, 28, 29, 30, 31 | 20 | 30N | 47E | 14-Sep-06 | 9 |
TWE 1 - 16, 18 | 28 | 30N | 47E | 10-Oct-06 | 17 |
TWE 17 | 22, 28 | 30N | 47E | 10-Oct-06 | 1 |
TWE 19 - 36 | 28 | 30N | 47E | 20-Sep-06 | 18 |
WT 1, 3, 5, 7, 9, 11, 13, 15, 17, 29, 31, 33, 35 | 2 | 30N | 47E | 31-Oct-06 | 13 |
WT 19, 21, 23, 25, 27 | 2 | 30N | 47E | 7-Nov-06 | 5 |
WT 37, 39, 41, 43, 45, 47, 49, 51, 53 - 55 | 36 | 31N | 47E | 1-Nov-06 | 11 |
WT 56 - 72 | 36 | 31N | 47E | 8-Nov-06 | 17 |
MALPAIS 1-24 | 4 | 29N | 47E | 4-Oct-14 | 24 |
MALPAIS 25-30 | 16 | 29N | 47E | 4-Oct-14 | 6 |
MALPAIS 210-220 | 18 | 30N | 47E | 4-Oct-14 | 11 |
MALPAIS 221- 260 | 30 | 30N | 47E | 4-Oct-14 | 40 |
MALPAIS 261-265 | 31 | 30N | 47E | 4-Oct-14 | 5 |
MALPAIS 31-48 | 4 | 30N | 47E | 5-Oct-14 | 18 |
MALPAIS 87-92, 111-128 | 6 | 30N | 47E | 5-Oct-14 | 24 |
MALPAIS 129-164 | 8 | 30N | 47E | 5-Oct-14 | 36 |
MALPAIS 201- 209 | 16 | 30N | 47E | 5-Oct-14 | 9 |
MALPAIS 316-347 | 32 | 31N | 47E | 5-Oct-14 | 32 |
MALPAIS 49-66 | 4 | 30N | 47E | 6-Oct-14 | 18 |
MALPAIS 67-86, 93-110 | 6 | 30N | 47E | 6-Oct-14 | 38 |
MALPAIS 302-315 | 18 | 31N | 47E | 6-Oct-14 | 14 |
MALPAIS 165- 200 | 12 | 30N | 47E | 7-Oct-14 | 36 |
MALPAIS 266-301 | 16 | 31N | 48E | 7-Oct-14 | 36 |
Unpatented Mining Claims | 831 | ||||
Notes | |||||
1. Amended K17 17-Aug-1992, K 18, K20 14-Aug- 1992 | |||||
2. Amended K22, K 24, K25, K26, K 27 17-Aug-1992 | |||||
3. Amended map 8/31/2006 |
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Table 4-2 Summary of Fee Land Holdings
APN | Legal Description | Royalty | Acres |
Section 9 T30N R47E MDB&M | |||
007-110-01 | NW1/4 | N/A | 160 |
007-110-13 | E1/2 NE 1/4 NE1/4, SE1/4 NE1/4, SE1/4 SW1/4 NE 1/4 | N/A | 70 |
Section 15 T30N R47E MDB&M | |||
007-140-01 | N1/2 NW1/4 | N/A | 80 |
007-140-03 | SW1/4 NW1/4 | N/A | 40 |
007-140-05 | SW1/4 NE1/4 | N/A | 40 |
007-140-12 | SE1/4 SW1/4 | N/A | 40 |
007-140-14 | Lots 1 & 2 | N/A | 65.39 |
007-140-15 | SE1/4 NE1/4 SW1/4 | N/A | 10 |
007-140-19 | S1/2 NW1/4 NE 1/4 | N/A | 20 |
007-140-20 | N1/2 NW1/4 NE1/4 | N/A | 20 |
007-140-21 | NW1/4 NE1/4 SW1/4 | N/A | 10 |
007-140-22 | NE1/4 NE 1/4 SW1/4 | N/A | 10 |
007-140-23 | SW1/4 NE1/4 SW1/4 | N/A | 10 |
007-140-25 | NW1/4 NE1/4 NE1/4 | N/A | 10 |
Section 23 T30N R47E MDB&M | |||
007-160-06 | E1/2 SE1/4 NE1/2 | N/A | 20 |
007-160-08 | N1/2 NE1/4 SE1/4 | N/A | 20 |
007-160-09 | SE1/4 NE1/4 SE1/4 | N/A | 10 |
007-160-16 | N1/2 SE1/4 NW1/4 | 5% NSR | 20 |
007-160-17 | N1/2 NW1/4 SW1/4 | N/A | 20 |
007-160-18 | NW1/4 NW1/4 | N/A | 40 |
007-160-19 | NE1/4 NW1/4 | N/A | 40 |
007-160-20 | NE1/4 SW1/4 NW1/4 | N/A | 10 |
007-160-21 | S1/2 SE1/4 NW1/4 | N/A | 20 |
007-160-22 | NE1/4 NE/1/4 SW1/4 | N/A | 10 |
007-160-25 | W1/2 SW1/4 NW1/4, SE1/4 SW1/4 NW1/4 | 5% NSR | 30 |
007-160-26 | NW1/4 NE1/4 SW1/4 | N/A | 10 |
007-160-27 | NE1/4, SW1/4 SE1/4, SE1/4 NW1/4SE1/4 | N/A | 20 |
007-160-28 | SW1/4 NE1/4 SE1/4, NW1/4 SE1/4 SE1/4 | N/A | 20 |
Section 21 T30N R47E MDB&M | |||
007-610-01 | NW1/4 | N/A | 160 |
Section 33 T30N R47E MDB&M | |||
007-640-06 | S1/2 NW1/4 | N/A | 80 |
30 | Fee Parcels | 1115.39 |
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Table 4-3 Summary of Leased Fee Land Holdings
APN | Legal Description | Lessor | Royalty | Expiration | Acres |
Section 15 T30N R47E MDB&M | |||||
007- 140-04 | SE1/4 NW1/4 | Third Party Lessor | 4% NSR | (2) | 40 |
007- 140-06 | SE1/4 NE1/4 | Third Party Lessor | 4% NSR | (2) | 40 |
007- 140-010 | NE1/4 SE1/4, E1/2 NW1/4 SE1/4 | Third Party Lessor | 2.5% NSR | (2) | 60 |
007- 140-07 | N2NW4SW4 | Third Party Lessor | 3% NSR & 0.5% wheelage royalty (1) | 31-July-33 | 20 |
007- 140-09 | W2NW4SE4 | Third Party Lessor | 3% NSR & 0.5% wheelage royalty (1) | 31-July-33 | 20 |
Section 23 T30N R47E MDB&M | |||||
007- 160-04 | SW4NE4 | Third Party Lessor | 3% NSR & 0.5% wheelage royalty (1) | 31-July-33 | 40 |
007- 160-24 | NE4NW4SE4 | Third Party Lessor | 3% NSR & 0.5% wheelage royalty (1) | 31-July-33 | 10 |
Section 19 T30N R48E MDB&M | |||||
007- 060-69 | Parcel 1 of the Sharp Hospital Map recorded in the Office of the Lander County Recorder in Book 375, Official Records, Page 170 | Third Party Lessor | 3% NSR & 0.5% wheelage royalty (1) | 31-July-33 | 9.28 |
8 Leased Fee Parcels | 239.28 |
Notes: | ||
1. | Wheelage royalty is calculated on mineralization mined from other properties which is transported underground through the leased property, and; | |
2. | The lease agreement remains in full force and effect for so long as any mining operations (as defined in the lease agreement) are being conducted on the relevant property on a continuing basis. |
Unpatented lode mining claims grant mineral rights and access to the surface within the boundaries of the claim. These rights are maintained by paying a maintenance fee of $155 per claim to the BLM prior to September 1st of each year. Failure to timely pay the maintenance fees will deem the claims “closed” by the BLM. The unpatented lode mining claims held by Klondex are currently in good standing through September 1, 2015, at which time Klondex must pay $128,805 to the BLM in order to maintain the claims for the following assessment year.
In addition to BLM maintenance fees, Klondex must record a Notice of Intent to Hold and pay a fee to the county in which the unpatented lode mining claims are situated. These fees and Notices of Intent to Hold must be paid and recorded in the applicable county by November 1st of each year. The unpatented lode mining claims held by Klondex are currently in good standing through November 1, 2015, by which time it must record the applicable Notices of Intent to Hold and pay fees in the amount of $8,351.50 to Lander County and $382 to Eureka County.
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Page 37 | Property Description and Location | Klondex Mines Ltd. |
The private fee lands and leases are subject to differing cash payments, net smelter return royalties (NSR), and wheelage royalties.
Royalties affect the following parcels owned and / or leased by Klondex, as listed in Table 4-2 and Table 4-3. Royalties applicable to the unpatented mining claims are discussed below. Property agreement obligations are listed in Table 4-4.
Table 4-4 Summary of Fire Creek Project Holding Costs
Due Date | Proj # | File # | Commitment/Obligation | $ Oblig | Payable/Due to | Notes |
9/1/2005 | L1010 | 2.4-6 | 3 Leased Parcels - Extended Term | Third Party Lessors | 1. 1987 Leases extended for 10 years from 9/1/2005 | |
8/18/2015 | L1010 | 2.4-6 | Property Taxes - 3 Leased Parcels | $ 146.78 | Lander County Treasurer | Lessee to pay property taxes |
8/18/2015 | L1010 | 8.3 | Property Taxes 29 - Klondex Owned Parcels | $ 1,032.33 | Lander County Treasurer | Real Property Taxes Due 3rd Monday of August annually |
8/18/2015 | L1010 | 8.3 | Property Taxes 2 - Klondex Owned Parcels | $84.08 | Eureka County Treasurer | Real Property Taxes Due 3rd Monday of August annually |
8/31/2015 | L1010 | 10 | BLM Claim Fees - 831 Claims | $ 128,805.00 | Bureau of Land Management | 831 Klondex Owned Claims x $155/Claim |
9/1/2015 | L1010 | 2.4-6 | 3 Leased Parcels - Annual AMR Payment | $ 24,000.00 | 7 Third Party Lessors | Annual AMR payment due on lease anniversary |
9/1/2015 | L1010 | 2.4-6 | Insurance Certificates | 7 Third Party Lessors | Insurance certificates required under terms of leases | |
11/1/2015 | L1010 | 10 | Lander County NOI to hold - 795 Claims | $ 8,351.50 | Lander County Recorder | 795 Klondex Owned Claims x $10.50/claim + $4 |
11/1/2015 | L1010 | Eureka County NOI to hold - 36 Claims | $ 382.00 | Eureka County Recorder | 36 Klondex Owned Claims x $10.50/claim + $4 | |
9/1/2015 | L1010 | 2.4-6 | 3 Leased Parcels - Expire | 7 Third Party Lessors | Leases expire - Renew | |
Total | $ 162,801.69 |
Notes: | ||
1. | The lease agreement remains in full force and effect for so long as any mining operations (as defined in the lease agreement) are being conducted on the relevant property on a continuing basis. | |
Source: Erwin and Thompson Title Report |
In addition, pursuant to a mining lease agreement effective July 31, 2013, with respect to five leased fee parcels, Klondex is required to pay minimum rental payments of $50,000 per year for the first ten years of the lease, which increase by $10,000 for each subsequent ten year period (including any renewal period). This lease also includes provisions that subject Klondex to an additional increase under certain circumstances. Pursuant to such mining lease, Klondex paid a minimum rental payment of $49,000.
On February 12, 2014, the Company entered into a royalty agreement (the “FC Royalty Agreement”) with Franco-Nevada US, a subsidiary of FNC, and KGS. Pursuant to the FC Royalty Agreement, KGS raised proceeds of US $1,018,050 from the grant to Franco-Nevada US of a 2.5% NSR royalty for Fire Creek. The royalty applies to all production from Fire Creek beginning in 2019.
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KGS entered into a gold supply agreement with Waterton Global Value, L.P. (Waterton) dated March 31, 2011, as amended and restated October 4, 2011 (the Gold Supply Agreement). Pursuant to the Gold Supply Agreement, the Company granted Waterton the right to purchase refined bullion (as defined in the Gold Supply Agreement) produced from the Project for the period commencing February 28, 2013 and ending February 28, 2018, subject to adjustment (the Term). If the Company has not delivered an aggregate minimum of 150,000 ounces of refined bullion during the first four years prior to the end of the Term, the Term will be extended until an aggregate of 185,000 ounces of refined bullion has been delivered (including any refined bullion delivered during the original Term) to Waterton. Under the Gold Supply Agreement, in the event that Waterton exercised its right to purchase refined bullion during the period of February 28, 2013 to May 31, 2013, the purchase price per ounce payable by Waterton was to be the purchase price per ounce of the last settlement price of gold on the London Bullion Market Association (the LMBA) PM Fix on the last trading day prior to the date Waterton provides notice to the Company that it intended to exercise its purchase right (the Pricing Date) less a 1% discount (which discount is only applicable if such price is more than US$900 per ounce). In the event that Waterton exercises its right to purchase refined bullion during the period following May 31, 2013 and before February 28, 2016, the purchase price per ounce payable by Waterton is the average settlement price of gold on the LMBA PM Fix for the 30 trading days immediately preceding the applicable Pricing Date (the Average Price) less a 1% discount; provided that in each case, if such price per ounce is less than US$900 the discount will be nil. In addition, in the event that Waterton exercises its right to purchase refined bullion after February 28, 2016, the purchase price per ounce will be the Average Price immediately preceding the applicable Pricing Date, without any discount.
The claim locations are based on location of monuments and their dimensions cited to the BLM. The claims’ boundaries are not surveyed, and the boundaries’ exact locations depend on physical positions of the location posts in the field. The authors are not aware of any conflicting surface rights in this area or near Fire Creek. Other considerations that might affect accessing claim status include grazing rights and protected habitats. Grazing rights may exist in the area, and conflicts with local ranchers are not common in this region. Newly established protected habitat for sage grouse has not been defined in this area at the time of this Report. There are archaeological considerations in the immediate area of the Project; however, all new surface disturbance proposed by Klondex is reviewed and permitted by the BLM prior to construction. Land information regarding claim status and fee lands was provided by Klondex, and to the authors’ knowledge at the time this document was authored, there were no environmental or social factors that would affect land title.
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Page 39 | Property Description and Location | Klondex Mines Ltd. |
4.4. Location of Mineralization
Gold mineralization at the Project occurs in steeply dipping epithermal veins within Tertiary basalt flows and intrusive rocks. The mineralized basalt rocks are a suite of mafic, extrusive rocks associated with the regional north-northwest-trending NNR structural zone. The NNR system has been documented in multiple geophysical and geological studies (e.g. John et al., 2000; Ponce, D.A. et al., 2008; Watt, J.T. et al., 2007) and is distinguished as a linear magnetic anomaly approximately 30 miles wide that extends 190 miles south-southeast from the Oregon-Nevada border to central Nevada. The NNR originates from the McDermitt Caldera in northwest Nevada and is likely related to impingement of the Yellowstone hot-spot on continental crust (Zoback et al., 1994). Figure 4-3 shows the location of the Project relative to the NNR.
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Klondex has an approved plan of operations with the BLM covering the current exploration activities at the Project as well as an approved bulk sampling permit from the State of Nevada. There are no environmental permitting issues known to the authors which are related to proposed Project activities.
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Page 41 | Accessibility, Climate, Vegetation, Physiography, Local Resources and Infrastructure | Klondex Mines Ltd. |
5. | Accessibility, Climate, Vegetation, Physiography, Local Resources and Infrastructure |
5.1. Access to Project
The Project is easily reached from the town of Elko by driving west on Highway I-80 for 40 miles to the Beowawe and Crescent Valley Exit #261. From Exit #261, proceed south on Nevada State Highway 306 for 16 miles (passing through Beowawe) to 10th Street (there is a sign on the right). On 10th Street, there is a Company sign at the turn that indicates, “Klondex Gold & Silver Mines, Limited”. 10th Street is the Project access road. The Project is located five miles west on 10th Street in Lander County, Nevada.
The state and county roads leading to the Project are mostly paved and maintained in order to service the ranches and mines in Crescent Valley such as Barrick Gold Corporation’s Cortez Mine. In this part of Nevada, it is common practice for mine staff to commute long distances for work on a daily basis. The average commute for Klondex staff is one hour each way.
5.2. Climate
Project climate is typical for northern Nevada with hot, dry summers and cold winters. Average daily summer temperatures range from 80° Fahrenheit (°F) to 90°F, and average winter low temperatures range from the low 40s°F to 20°F. Summer temperature extremes may reach 100°F for short periods, and winter extreme temperatures may drop below 0°F for short periods. Fieldwork, including exploration drilling, is commonly conducted throughout the year in this area. Mines in the Crescent Valley typically operate all year without experiencing any major weather-related problems.
5.3. Vegetation
Fire Creek vegetation is mainly limited to sagebrush, other species of low vegetation and some grasses. There are no trees at the Project. Due to the low amount of rainfall, the vegetation is low and sparse. There is a small marsh associated with the Fire Creek drainage that provides some wetland vegetation.
5.4. Physiography
The Project lies in elevation between 4,900 feet and 7,200 feet. The United States Geological Survey (USGS) published a base-relief map, which covers the Project area titled, “Mud Spring Gulch Quadrangle Nevada-Lander Co. 7.5 Minute Series (Topographic)”. The topographic relief is moderate with mature topography consisting mostly of rounded hills with steeper grades along more competent strata. The stream down-gradient from the Project are ephemeral and are sourced by up-gradient springs.
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5.5. Local Resources and Infrastructure
The nearest rail siding is located in the town of Beowawe, a small community of about 50 people, approximately 15 miles north of the Project. Crescent Valley, a small town with a population of approximately 200 people, is about seven miles south of the Project.
The towns of Battle Mountain and Elko, about 52 miles northwest and 63 miles northeast of the Project, respectively, are the nearest larger towns and supply most of the labor force. These towns are the only locations with amenities and services such as motels, fuel, grocery stores, and restaurants. The nearest commercial retail stores for fuel and groceries are located in Battle Mountain, 52 miles to the northwest.
Klondex’s land holdings at Fire Creek have adequate acreage to support future exploration and mining activities. Fire Creek mineralization will be transported to the Company’s Midas Mill for processing.
Electrical power is provided to the Project by NV Energy, Inc. (“NV Energy”) through a transmission line and substation located near the eastern Project boundary. The substation was connected to the NV Energy electrical grid in 2013.
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Page 43 | History | Klondex Mines Ltd. |
6. | History |
6.1. Exploration History
The first recorded lode claim dates to 1933, but no other activity is known prior to 1967. Table 6-1 below itemizes exploration performed since 1967.
Table 6-1 Exploration History
Dates | Company | Details |
1967 | Union Pacific Resources | Drilled two core holes. |
1974 to 1975 | Placer Development Ltd. | Drilled 22 rotary holes. |
1975 | Klondex Mines Ltd. | Acquired the Project . 1980-1983 drilled 64 rotary holes. 1981 gold test production. |
1984 | Minex Resources, Inc. | Leased the Project from Klondex, drilled 13 rotary holes. |
1986 to 1987 | Alma American Mining Company (“Alma”) | Leased the Project from Klondex, drilled 64 rotary holes. |
1988 | Aurenco Joint Venture (“Aurenco JV”) | Aurenco JV formed between Black Beauty Mining and Covenanter Mining. |
1988 to 1990 | Aurenco JV | Leased the Project from Klondex. |
1990 to 1995 | Klondex Mines Ltd. | No activity. |
1995 to 1996 | North Mining Inc. (“North Mining”) | Leased the Project from Klondex. Drilled 67 holes, performed IP and HEM surveys. |
1996 to 2004 | Klondex Mines Ltd. | No activity. |
2004 to 2012 | Klondex Mines Ltd. | Began a deep exploration program. Development commenced in 2011. |
2012 to Present | Klondex Mines Ltd. | New Management and Board of Directors in 2012, ongoing exploration. |
Prior to 1994, exploration focused on near-surface oxide mineralization most likely for bulk-mineable targets. Klondex acquired Fire Creek in 1975 and subsequently performed rotary drilling and a small test heap leach operation that produced 67 oz. Au. Minex leased the Project in 1984-1985, performed a small amount of drilling and conducted a larger test heap leach operation using approximately 30,000 tons of material. Due to the use of only the exploration drilling and no ore control, the material was primarily waste and ultimately produced less than 1,000 oz. Au. Alma American Mining Company, a division of Coors Brewery, leased the Project from 1986-1987 and performed rotary drilling and other exploration work. The Aurenco Joint Venture, formed between Black Beauty Mining and Covenanter Mining, leased the Project from 1988-1999. From 1988 to 1990, the Aurenco JV completed 51,476 feet of rotary drilling, 500 soil samples, and 750 surface rock chip samples. The Project was ventured with Coeur Mining from 1993 to 1994. The Fire Creek Joint Venture was formed between Aurenco and North Mining in 1995. During 1995 and 1996, North Mining commenced the first technical exploration drilling program to examine deeper targets. North Mining drilled 67 rotary and core holes for a total of 39,570 feet. This program successfully drilled the first high-grade gold intercept at depth at Fire Creek. In 1995, North Mining conducted an IP-Resistivity survey along ten east-west lines. Much of North Mining’s drill locations from 1995 and 1996 targeted results from these geophysical tests; however, the wide point and line spacing did not detect the narrow vein anomalies. Details of this earlier geophysical survey were itemized in the Fritz Geophysics report for Klondex (Fritz, 2006) and in an unpublished report for North Mining (Edmondo, 1996). North Mining dropped the Project in 1996 after determining that the Project was not likely to meet their minimum contained gold requirement for continued exploration. Aurenco dropped the Project in 1999 without conducting further work, and the Project reverted to 100% Klondex control.
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No work took place until 2004, when Klondex began systematically and aggressively drilling deep targets to define the mineralization potential recognized by North Mining. In 2004, Klondex based its initial drilling targets on the results of North Mining’s drilling program carried out from 1995 to 1996 in combination with information including integrated geologic mapping, surface geochemistry, airborne helicopter electromagnetic (HEM) surveys and IP dipole-dipole surveys. Klondex focused its exploration drilling on targets ranging from 500 to 1,700 feet below the surface, yielding grades up to 1.0 ounces per ton (opt).
Klondex conducted another IP survey in 2004 that used tighter line spacing and dipole points and, which identified north-northwest trending alterated zones, coincident with the general strike of veins identified by Klondex drilling and coincident with the general trend of NNR faults (see Regional Geology, Section 7.1 of this Report). From 2004 to 2010, Klondex drilled 231 surface holes for a total of 297,586 feet.
6.2. Production History
Historic production, as itemized previously (Raven et al., 2011), is limited to marginal mining of oxidized siliceous cap material from a pit and the construction of a small test heap leach operation from 1988 to 1990. A summary of the Raven report follows:
“In the early 1980’s a joint venture with Aurenco/Black Beauty Coal mined a small amount of material from a pit and constructed a small test heap leach operation. This work focused on the siliceous cap material that overlies the deeper, epithermal vein systems; mining of the siliceous cap did not prove to be economic as the grades were too erratic and high clay content hampered the heap leaching.” (Page 11)
With the exceptions of current operations, there has been no other production at the Project since 1990.
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Page 45 | Geological Setting and Mineralization | Klondex Mines Ltd. |
7. | Geological Setting and Mineralization |
7.1. Regional Geology
The Project is located on the northeast flank of the Shoshone Range in Lander County Nevada, and in the western half of the NNR (Figure 7-1). The surface and near-surface NNR is composed of an alignment of middle-Miocene basaltic (and lesser rhyolitic) dikes and up to 42,000 feet of basin-filling lava flows, pyroclastic units and lacustrine sedimentary units (Zoback et al., 1994; John et al., 2000) that are distinguishable regionally as a prominent, north-northwest trending aeromagnetic anomaly that extends some 300 miles south-southeastward from the Oregon-Nevada border. The NNR is likely related to a pre-Cenozoic, deep-crustal fault reactivated between 16.5 and 14.7 million annum (Ma) (Zoback et al., 1994; Theodore et al., 1998; John et al., 2000) and reflects west-southwest – east-northeast regional extension (Wallace & John, 1998; John & Wallace, 2000). Some workers (Zoback & Thompson, 1978; Pierce & Morgan, 1992) postulate that impingement of the Yellowstone hot spot on this area at approximately 17 Ma is related to Cenozoic NNR activity.
Basement rocks of the northern Shoshone Range are comprised of lower Paleozoic primarily siliciclastic sedimentary units of the Roberts Mountain Allochthon upper plate (John & Wrucke, 2003; Figures 7-2 and 7-3). In this area, the upper plate is 1,000 to 2,000 feet thick, and the Roberts Mountain Thrust dips west-northwest (Kiska Metals Corp., 2014). The primary upper plate units in the Fire Creek area are imbricate thrust stacks of Ordovician Valmy Formation, which is comprised of sandstone, shale, chert, and quartzite and the Devonian Slaven Chert (Gilluly & Gates, 1965; John & Wrucke, 2003).
Overlying the Paleozoic sedimentary rocks is a discontinuous tuff layer. John et al. (2003) and John & Wrucke (2003) assigned this unit as the Caetano Tuff (33.87 Ma) in the vicinity of Mule Canyon. However, Colgan et al. (2014) documents the Tuff of Cove Mine (34.4 Ma) and the Nine Hill Tuff (25.4 Ma) in the northern Shoshone Range in this stratigraphic position. The origin and continuity of this unit remains enigmatic.
A middle-Miocene package of intercalated basalt and basaltic andesite flows and associated pyroclastic units intrudes and unconformably overlies the lower sedimentary and tuffaceous rocks. As these rocks represent local paleotopography, their presence and thickness are highly variable. Competent flow units in this package form the dominant host for gold mineralization both at the Fire Creek Project and the nearby Mule Canyon Mine. As such, local expressions of this package have been informally named the Mule Canyon Sequence (John et al., 2003 and references therein) and the Fire Creek Sequence (McMillin & Milliard, 2013).
The Andesite of Horse Heaven, a sparsely porphyritic andesite to basaltic andesite, conformably overlies the basalt flow package (John & Wrucke, 2003). This unit covers an extensive area of the Northern Shoshone Range (Gilluly & Gates, 1965) and ranges from less than 130 feet to greater than 800 feet thick (John & Wrucke, 2003). Samples from this unit collected near the Mule Canyon Mine yielded whole-rock ages of 15.86±0.12 Ma and 15.2±0.8 Ma (John & Wrucke, 2003). Another sample collected near Corral Canyon, south of the Project, yielded a whole-rock age of 15.76±0.80 Ma (John et al., 2000). The Andesite of Horse Heaven is currently recognized as the youngest unit preserved at the Project.
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Thick flows of dacite and trachydacite uncomformably overly younger mafic units. John & Wrucke (2003) describe these as occurring mainly to the east of the Muleshoe Fault and represent rift-filling lavas that were sourced from the Sheep Creek Range. They report 40Ar/39Ar plagioclase age dates of 15.33±0.09 Ma and 15.34±0.10 Ma for samples collected near the Mule Canyon Mine and in the Sheep Creek Range, respectively.
Numerous steeply dipping, north-northwest- to north-striking mafic dikes are evident at the Project from drill data and mining operations (Edmondo, 1996; McMillin & Milliard, 2013) and are exposed in the open pits at the Mule Canyon Mine (John et al., 2003 and references therein), however, few mafic dikes have been mapped at the surface. These are interpreted as feeder dikes for the upper Mule Canyon Sequence and lower Andesite of Horse Heaven (Edmondo, 1996; John & Wrucke, 2003). Field and core observations at the Project support this interpretation.
The western margin of the NNR in the Northern Shoshone Range is marked by two high-angle fault sets. The dominant set is parallel to the rift axis striking north-northwest (N15-30°W) and exhibits dip-slip movement. The most prominent of these is the Muleshoe Fault, which is less than a mile east of both the Mule Canyon Mine and the Fire Creek Project (John et al., 2003). Faults in this orientation commonly host mafic dikes and provided structural control on eruption and volcanic rock deposition. A second high-angle fault set oriented east-northeast (N60-80°E) was active during NNR formation, most notably the Malpais and Argenta Faults (John et al., 2000; John et al., 2003). These display left-lateral oblique-slip, however, some of these were reactivated in the late Miocene after a clockwise rotation of extension direction (Zoback et al., 1981, 1994).
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Project, Lander County, Nevada |
7.2. Local Geology
7.2.1. | Rock Units |
Basement rocks beneath the Fire Creek deposit have not been drilled sufficiently for positive unit identification. Imbricate stacks of Ordovician Valmy Fm. and Devonian Slaven Chert, part of the Roberts Mountain Thrust upper plate, are mapped to the west of the deposit and are presumed to lie beneath the local Miocene volcanic package (Figures 7-2 and 7-3). Thickness of the upper plate rocks in this region is unconstrained. Lower plate rocks are thought to be Roberts Mountain Formation, but this has not been drill-tested, and no outcrops of this unit occur nearby.
Overlying the Paleozoic sedimentary package is a 0 to 300-foot thick, discontinuous tuff unit, tentatively identified as the tuff of Cove Mine (C. Henry, pers. comm., 2013; D. John, pers. comm., 2014). The discontinuous nature of this unit is thought to be a function of paleo-topography.
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Progressing upwards, unconformably overlying the tuff of Cove Mine, is approximately 500-foot thick section of interbedded lithic tuff beds, basalt flows and sills, and thin, laminated lacustrine sedimentary beds. These are grouped together under the Ttb (Tertiary tuff and basalt) moniker. Tuff layers are commonly intensely argillically altered. Alteration in basalts varies from unaltered to moderate propylitization.
The informal Fire Creek Sequence comprises three volcanic/volcaniclastic units that overlie the Ttb series. These are presented in ascending order. Descriptions are after Edmondo (1996), Anderson (2013), and Milliard et al. (in prep).
Tbeq (Tertiary basalt equigranular; Figure 7-5) is a 400- to 700-foot thick, black to dark green, aphanitic and equigranular basalt flow package. The dominant textural characteristic of this unit are randomly oriented, curvilinear, interconnected hackly or tortoise-shell joints that develop in response to cooling and are thus a primary textural feature (McPhie et al., 1993). Poorly-formed columnar jointing is also present locally. Hyaloclastite is common at the unit base. Thin, discontinuous, and volumetrically minor tuff layers can be present. This unit is the primary ore host. It is thought that Tbeq possessed the bulk strength to hold open space during faulting/fracturing and was present at the correct elevation with respect to the paleo-water table to allow fluid boiling and vein deposition. In the vicinity of the Fire Creek deposit, a large percentage of this unit is altered. Propylitic alteration volumetrically dominates the alteration package and ranges from thin selvages along tortoise-shell joints to pervasive. Argillic alteration is proximal to veins and dikes. Silicification is intermittent and, when present, is immediately adjacent to veins and dikes.
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Tbma (Figure 7-6) discontinuously overlies Tbeq and is a 0 to 500-foot thick series of black, aphanitic, vitreous, and peperitic basalt flows that may be intercalated with thin tuff layers of the overlying Tlat. No gold mineralization is known in this unit. Alteration is non-existent to weakly propylitic.
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Tlat (Tertiary lapilli ash tuff; Figure 7-6) also discontinuously overlies Tbeq, at the same or higher stratigraphic level as Tbma. Commonly, the contact between Tlat and Tbma, when both are present, is difficult to determine. Tlat is a 0 to 200-foot thick, tan to buff, non-welded lithic lapilli tuff with 10 to 40% heterolithic basalt and scoria fragments. Groundmass comprises shard and pumice fragments with 10 to 15% lapilli component. Although discontinuous, it is regionally extensive. In the vicinity of the Fire Creek deposit, this unit is commonly intensely argillized.
The Andesite of Horse Heaven is the youngest package preserved at the Project. Locally, this package is broken into five units. Tb1, Tb2, and Tb3 directly overlie the Fire Creek Sequence and the Fire Creek deposit. Tb4 and Tb5 are only present to the east and northeast of the current mine area and may reflect compartmentalized lava fill into a fault-bounded basin. Descriptions are after Edmondo (1996).
Tb3 is the youngest unit present at the Fire Creek deposit. It consists of interbedded andesite and basalt flows. Typically very fine grained with rare plagioclase and biotite phenocrysts up to 0.1 millimeters in diameter. Individual flows display features characteristic of subaerial emplacement including autobreccia at flow tops and bases, pahoehoe textures, dense flow interiors and increasing vesiculation density near flow tops. Above the known deposit, Tb3 is argillized and hosts gold mineralization.
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Tb2 is a black, aphanitic to sugary, weakly glassy basalt that contains trace to 10% plagioclase phenocrysts and five to seven percent magnetite as needles. Emplacement as subaerial flows, similar to Tb3, is indicated by autobrecciation along flow tops and bottoms, dense flow interiors, and strong vesiculation. Thicker flows may weather spheroidally. The base of Tb2 is weakly altered, and this alteration rapidly decreases to zero vertically.
Tb1 shares many similarities to Tb2, specifically that it is a black, aphanitic to sugary, weakly glassy basalt with trace to 10% plagioclase phenocrysts. However, instead of magnetite needles this unit can be distinguished by the presence of three to five percent magnetite as crystals. The sugary groundmass is slightly coarser grained than Tb2. Flow textures are the same as Tb2. Tb1, and Tb2 are commonly separated by a thin volcaniclastic unit and, in outcrop, may be marked by an angular flow foliation discordance of less than 10 degrees. Hypogene alteration in this unit has not been observed.
Tb4 is light red-grey to grey, platy to massive andesite interbedded with black, glassy, perlitic, porphyritic andesite. Phenocrysts of plagioclase and pyroxene volumetrically compose up to 25% and range from two to five millimeters in length. In contrast to the consistent foliation displayed by Tb1, 2, and 3, Tb4 has highly variably flow foliations and forms gently rolling antiforms and synforms with an overall west dip. In the Project area, Tb4 is present to the east of a range-front-parallel fault located to the east of the deposit. Here, it underlies Tb3 and is in fault contact with Tb1 and 2.
Tb5 is a series of fine grained to aphanitic, brown to black basalt flows with one to three percent magnetite and pyroxene phenocrysts. Individual flows have flaggy to platy bases and highly vesicular tops. It appears to underlie Tb4 although exposure is limited to the northeast corner of the Project area.
Units underlying Tb1 are cut by numerous black to dark green mafic dikes referred to as Tim (Figure 7-7). Textures include aphanitic, fine-grained phaneritic and weakly porphyritic. Dikes generally strike north-northeast and many exploited north-northeast-striking (Muleshoe-parallel, see below) faults. Contacts between dikes and wall rocks range from knife-edge sharp to brecciated zones up to one foot. Both acted as conduits for mineralizing fluids, and vein emplace along these contacts (e.g. Vonnie Vein). Dikes can be altered along with wall rock, but often comparatively pristine dikes cut through intensely argillized wall rock, suggesting dikes were emplaced late relative to the bulk of fluid migration.
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7.2.2. | Structure |
The current Fire Creek deposit is fault-bounded to the north, east, and south. The west remains structurally open, although data for this area is sparse. Drilling from underground has roughly defined the Alimak Fault, a north-northwest striking structure that intersects the westernmost extent of the underground workings. It is unknown if this is a system-bounding fault; however, ground conditions change sharply across it. The bounding structures are described below.
North: John et al. (2000) documents the northeast to east-northeast striking Malpais and Argenta Rims (Figure 7-1), the result of late Miocene to Quaternary (i.e. post-mineralization), north-down normal faulting and associated sets of steeply north-dipping normal faults clustered in the north sides of the respective fault blocks. Geophysical and drill data indicate these subsidiary north-dipping normal faults truncate and may offset the Fire Creek deposit to the north.
East: The Muleshoe Fault is a regionally important structure, forming, along with the Dunphy Pass Fault, a graben with over 1,200 feet of volcanic fill and forming the eastern edge of the Mule Canyon deposit (John et al., 2000; John et al., 2003). The Fire Creek deposit is bound to the east by a N15°W, steeply east-dipping fault interpreted as a paleo-scarp. Volcanic rock texture and composition changes abruptly across an undeformed flow contact. Gold assay values also abruptly change from an average of 0.0X opt to below detection limit across this contact. 3D modeling of these criteria supports the hypothesis that this is a high-angle fault plane. Throughout the Project history, this fault surface has been referred to as the Muleshoe Fault, and this report will continue to do so. It should be noted that this fault has never been positively linked to the Muleshoe Fault proper documented at Mule Canyon, and geophysical evidence suggest that the true Muleshoe Fault lies farther to the east.
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South: Fire Creek itself runs east-west and lies just south of the known deposit. Surface mapping indicates that the Tb2 unit on the south side of the creek is significantly thicker than Tb2 on the north side. This relationship suggests that Fire Creek follows the surface trace of a south-block-down normal fault (the Fire Creek Fault) that either predated emplacement of Tb2 or was synchronous with Tb2 emplacement, forming a volcanic growth fault. Geophysics and limited drill data support the hypothesis that volcanic stratigraphy is displaced across the Fire Creek Fault.
There are currently three known fault sets within the area described above (Figure 7-8). The most recent set comprises the NE1 and NE2 faults. Both are northeast-striking and dip steeply to the north, sub-parallel to the Malpais Rim and subsidiary structures. Apparent displacement across the NE1 and NE2 faults is small and of variable motion sense, probably more reflective of local perturbations in the volcanic stratigraphy than of true offset. These are interpreted to be a continuation of the late Miocene to Quaternary Malpais Rim structure set and thus postdate and offset mineralization. Displacement across these structures likely increases with proximity towards the Malpais Rim.
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The other two fault sets are cut by and thus predate the NE1 and NE2 faults. The N15°W set comprises the vertical to steeply east-dipping Muleshoe Fault and Alimak Fault and several other parallel, smaller-displacement faults (not shown for clarity) that dip steeply to the east and west. All show apparent normal displacement. Displacement across the Muleshoe Fault and Alimak Fault is east-block-down based on offset volcanic stratigraphy. Direct evidence for an oblique component does not exist, but these are thought to contain a subordinate right-lateral component based on overall NNR development patterns. North of Fire Creek proper, where Tb2 is very thin and Tb1 is either thin or eroded, the N15°W fault orientation is strongly reflected in current topography. South of Fire Creek, Tb2 is significantly thicker, and the N15°W fault set is not topographically expressed. This implies that the relative age of Muleshoe-parallel faulting can be bracketed between Tb1 and Tb2 emplacement. The N45°W fault set comprises the NW1 and NW2 faults. These represent breached relay ramps (Crider, 2001; Trudgill & Cartwright, 1994; Figure 7-9) and formed contemporaneously with Muleshoe-parallel faults. Both of these fault sets are thought to result from NNR development.
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7.2.3. | Veins |
The vein system reflects extensional structural fabrics generated during NNR development. Veins were emplaced primarily along faults and dike contacts, both striking approximately N15°W and with variable but steep dips, and north-south-striking, moderately east-dipping extensional structures. North-northwest-striking veins are typically thin, less than three feet, subvertical and are parallel to the Muleshoe Fault set. North-south striking veins are thicker, approximately 10 feet, than north-northeast striking veins. Host rocks are usually restricted to the more competent members of the volcanic sequence; in the known deposit this is primarily Tbeq. Tuffaceous units are less favorable for vein formation due to poor fracturing characteristics.
The following description of Fire Creek veins is abstracted from Raven et al. (2011) and includes relevant updates.
The veins consist of colloidal silica, crystallize chalcedony and coarser crystallize quartz, calcite, pyrite, chlorite, arsenopyrite, adularia, and clays including kaolinite, smectite and illite. Crustiform/colloform-banded and brecciated quartz, stockwork texture and calcite-replacement textures including bladed quartz are common. Drusy and cockscomb calcite and quartz often coat open spaces. Vein composition ranges from quartz-dominant to calcite-dominant, even within the same vein.
As of this writing, 47 individual veins or mineralized structures have been identified. Of these, five have been sufficiently characterized to warrant individual descriptions.
Joyce Vein
The Joyce Vein has been defined for 1,750 feet along strike and 1,135 feet of dip extent. It is dominated by coarse, bladed calcite (60 to 70%) with quartz as the remainder.
Vonnie Vein
The Vonnie Vein has been defined for 1,910 feet along strike and 550 feet of dip extent. Textures are dominantly crustiform/colloform quartz banding with lesser carbonate. This vein formed predominately along a dike contact.
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Karen Vein
The Karen Vein has been defined for 1,035 feet along strike and 450 feet of dip extent. Average vein width is approximately 0.5 foot although mineralized widths can reach up to approximately 12 feet and can include fault-related breccias and discrete veins. The vein is predominately calcite with lesser quartz and commonly has open space vugs. The Karen Vein exploited a north-south striking extensional linking structure rather than a Muleshoe-parallel fault or dike contact.
Hui Wu Vein
The Hui Wu (pronounced Whey-Woo) structure has been defined for 650 feet along strike and 500 feet of dip extent. This structure is primarily mineralized tectonic breccia that is punctuated by a moderately developed discrete vein system.
Honeyrunner Structure
The Honeyrunner structure has been defined for 1,515 feet along strike and 525 feet of dip extent. Geologic data suggest this structure may be a locally important fault parallel to the Muleshoe Fault system. Instead of a typical vein, this structure is a combination of tectonic breccias and a large basalt dike; however, current drill piercements do not preclude the presence of a vein either along strike or at depth.
7.2.4. | Alteration |
Alteration is zoned laterally and vertically with respect to paleo-fluid conduits and is dependent on rock type. Conduits include high-angle structures such as faults (either with or without vein fill) and dike contacts and to a lesser extent low-angle structures such as lithologic contacts and highly vesiculated flow tops. Zonation is well-developed in Tbeq basalt. Alteration in tuffaceous units tends to be pervasive rather than zoned.
Idealized lateral distal-to-proximal alteration zonation around a single fluid conduit or vein within Tbeq or Ttb basalt typically follows the progression outlined below (Figures 7-10 and 7-11). Not all stages may be present and overprinting is common.
1. | Distal, widespread, propylitic alteration characterized by pyritiferous and chloritic selvages along hackly or tortoise -shell joints; | |
2. | Pervasive propylitic alteration characterized by chlorite ± calcite replacement of plagioclase and pyroxene and abundant formation of both disseminated and selvage pyrite and | |
3. | Pervasive argillic alteration characterized by montmorillonite ± nontronite ± illite replacement of plagioclase and pyroxene (or their chloritized equivalents). |
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Selvage and/or pervasive silicification through addition of silica.
Acid-leach silicification resulting from preferential removal of mobile, non-silica constituents. This alteration style is more common in the upper portion of the hydrothermal system.
Argillic alteration in tuffaceous units and interbeds is characterized by near-complete replacement by illite ± kaolinite ± smectite ± montmorillonite ± nontronite. It is widespread and is not zoned. The typical propylitic outer halo is either non-existent or has been completely overprinted.
Alteration in Ttb basalt units is generally weak to moderate, pervasive propylitic alteration characterized by chlorite replacement of plagioclase and pyroxene.
A discontinuous, 15 to 65 feet thick, white to reddish-brown, amorphous to opaline silica cap is present between Tb1 and Tb2. Although specific fluid pathways have not been identified in Tb1, an elongate zone of moderate to intense, vertically zoned argillic alteration directly overlies the Joyce Vein in Tb1 and is exposed at the surface. This alteration is characterized by alunite + kaolinite beneath the silica cap and gives way to smectite + kaolinite with depth. Nontronite-alteration as vein, vug-fill and pervasive basalt alteration appears to overprint other alteration events.
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7.2.5. | Mineralization |
Gold is primarily present in its native state along discrete layers within veins. Native gold can occur as large clots or bands, less than ¼ inch (Figure 7-12) dendritic growths and (Figure 7-13) and fine-grained disseminations. Other less common habits include encapsulations in quartz, pyrite replacements and coatings on pyrite or arsenopyrite (Thompson, 2014). Silver occurs encapsulated in quartz and locally in naumannite or ruby silver encapsulations in quartz (Thompson, 2014). Dark grey ginguro bands of an unidentified silver-bearing mineral is present along vein banding as well. The silver:gold ratio is approximately one to one.
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8. | Deposit Types |
The Fire Creek deposit is considered to be a low-sulfidation, epithermal deposit.
A composite description for low-sulfidation epithermal deposits, abstracted from Simmons et al. (2005), Cooke & Simmons (2000), White & Hedenquist (1995), Kamenov et al. (2007), and Hedenquist et al. (2000) is shown below in Figure 8-1.
Low-sulfidation epithermal systems are also referred to as quartz ± calcite ± adularia ± illite or adularia-sericite epithermal systems. These nomenclatures refer to the oxidation state of the ore fluid sulfur component, gangue mineralogy and hydrothermal fluid pH, respectively. Ore-fluids in a low-sulfidation hydrothermal system are reduced, have a near-neutral pH and are dominated by deeply-circulated meteoric water. These deposits form in the shallow crust, 0.5 to 1.5 miles at temperatures of greater than 300°C in subaerial volcanic settings. Steeply-dipping, open-space veins are common. Quartz is the principal gangue mineral and can be accompanied by chalcedony, adularia, illite, pyrite, calcite, and rhodochrosite. Boiling is the dominant metal deposition mechanism and commonly results in vein textures including crustiform-colloform bands and platy calcite and/or quartz-after-calcite pseudomorphs. Ore metals are usually Au-Ag, Ag-Au or Ag-Pb-Zn and, contrary to the ore-fluid source, metals in NNR-related epithermal deposits are sourced from mantle-derived basaltic magmas (Kamenov et al., 2007).
Zoned hydrothermal alteration comprises widespread and deep propylitization that grades upwards to clay, carbonate and zeolite formation. Proximal alteration comprises quartz, adularia, and pyrite. High-level advanced argillic alteration characterized by clay-carbonate-pyrite or kaolinite-alunite-opal ± pyrite alteration can be present above the ore-grade zone and is the result of steam-heated, acidic, ascending fluids generated during boiling.
Features that classify the Project as a low-sulfidation epithermal deposit include:
• | Precious metal mineralization occurs primarily within steeply dipping veins; | |
• | Extensional, open-space forming tectonic environment active during vein emplacement; | |
• | Vein gangue is composed of quartz and calcite and exhibits boiling textures; | |
• | Mineralization is gold -silver; | |
• | Alteration halo comprises distal propylitization that grades to argillic and proximal silicification; | |
• | Presence of a high-level, advanced argillic alteration zone capped with opaline silica; and | |
• | Altered host rock indicates a reduced ore fluid. |
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9. | Exploration |
9.1. Historical Exploration
An itemized summary of exploration activities at the Project is below.
• 1933: First recorded lode claim at Fire Creek;
• 1967: Union Pacific drilled two diamond holes;
• 1974 – 1975: Placer Development Ltd. acquired an exploration lease and drilled 22 rotary holes;
• 1980: Klondex acquired the Project from Placer Development, Ltd;
• 1981/1982: Klondex conducted a 2,000-ton test heap leach that produced 67 ounces of gold;
• 1980 – 1983: Klondex drilled 64 rotary holes;
• 1984: Klondex leased the Project to Minex Resources, Inc. who drilled 13 holes and heap leached approximately 30,000 tons of mixed ore and waste which produced approximately 1,000 ounces of gold;
• 1986 – 1987: Klondex leased the Project to Alma American Mining Co. who drilled 64 holes;
• 1988 – 1999: Klondex leased the Project to the Aurenco Joint Venture which composed of Black Beauty Gold Co. and Covenanter Mining, who drilled 51,463 feet of reverse circulation,
• 1993 – 1994: The Aurenco JV ventured the Project with Coeur Exploration. Coeur conducted a gradient-array resistivity survey and drilled seven reverse circulation and two diamond holes;
• 1995 – 1996: The Aurenco JV and North Mining form the Fire Creek Joint Venture.
North Mining conducted a dipole-dipole IP/Resistivity survey and drilled 39,593 feet of reverse circulation and diamond core;
• 1999: The Aurenco JV relinquished their lease;
• 2004: Klondex began an exploration program for deep vein-hosted gold mineralization;
• 2005: Newmont Mining Corp. performed a gravity survey;
• 2006: Klondex conducted a gradient-array IP/Resistivity survey; and
• 2004 – 2010: Klondex drilled 231 holes, primarily core with RC pre-collars, for a total length of 297,586 feet.
9.2. 2011 Drilling
Fifty-five drill holes comprising 37 surface holes and 18 underground holes with a length of 65,225 feet were completed (Figure 9-1). Surface drilling focused on identifying mineralization on the north end of defined veins. Underground drilling focused on identifying mineralization on the southern extent.
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9.3. 2012 Drilling
Sixty-one drill holes comprising of 25 surface holes and 36 underground holes with a total length of 54,969 feet were completed (Figure 9-2). Four of the surface holes were geotechnical holes drilled to gather data near the planned vent raise. Three holes were drilled to test IP anomalies south of the Project. These did not encounter significant gold mineralization; however, the holes were terminated prior to encountering the target horizon and may have been located too far to the east. The remainder were drilled to define a bulk sample area that encompassed the Joyce Vein and the Vonnie Vein between the 5370 and 5400 crosscuts. One of these holes (FC1211) returned a result of 2,910 parts per million (ppm) Au (85 opt Au) assay from the Vonnie Vein.
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9.4. 2013 Drilling
Sixty-one drill holes comprising five surface holes and 56 underground holes with a total length of 33,501 feet were completed in 2013 (Figure 9-3). This drilling identified several new veins west of the decline and identified probable southern extensions of the Joyce Vein and the Vonnie Vein.
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9.5. 2014 Drilling
Two hundred eighty-three holes comprising nine reverse-circulation surface holes with a total length of 2,385 feet (Figure 9-4), two HQ diamond surface holes with a total length of 2,943 feet (Figure 9-4) and, 272 AQ, BQ and HQ diamond underground holes with a total length of 73,339 feet (Figure 9-5) were completed in 2014. Five of the surface reverse circulation (RC) holes were converted into groundwater monitoring wells GW-4 through GW-8. The remaining five surface RC holes had piezometers installed. Two HQ diamond holes were drilled for condemnation purposes. Underground drilling in 2014 primarily focused on infilling and extending the Joyce Vein, Vonnie Vein, Karen Vein, and Hui Wu Vein. Underground exploration targeted zones to the east and west of the decline and yielded positive results including discovery of the ore-grade Honeyrunner structure.
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10. | Drilling and Sampling Methodology |
Drilling protocols from 2004 through 2010 are documented in Raven et al. (2011):
“Most core holes were pre-collared with a reverse circulation rotary (RC) drill that advanced to a planned depth well short of the intended target intercept. The RC holes were then cased and core drilled to completion with HQ (2.5 inch diameter core)-sized core. Two of the borings, 410 and 411, were only rotary holes drilled to completion. RC drilling was done by O’Keefe Drilling of Butte, Montana. Core drilling was carried out primarily by Boart-Longyear out of their Salt Lake office, Ruen Drilling from Clark Fork, Idaho and Major Drilling from Salt Lake City.”
“The directions and angles of the drill holes were spotted to intercept the veins as close to perpendicular as practicable within the limitations of the equipment. Most holes were drilled at azimuths of 75° or 255° and located as close as practical on the surveyed grid lines with azimuths of 75° … The line spacings are 50 metres. The deep holes have established that veins or vein systems have a general azimuth strike of 345° with varying dips ranging from steep westward dips of about 75° to steep eastward dips of about 80°. Most holes were inclined at an angle of -45°. Holes were drilled both ENE and WSW; sometimes the ideal direction/declination had to be compromised because of drill location setup problems.”
“The Klondex holes are all surveyed for vertical and horizontal deviation by International Directional Services LLC, whose local office is in Elko, Nevada. Plotting the boring deviations permit accurate vein and other gold anomaly intercept locations leading to reliable geologic mineralization locations, interpretations of vein trends, structure dips, zone widths, reserve estimates, and polygon locations.” (Page 21)
The 2013 surface drilling procedures for RC are summarized below:
1. | Klondex contracted Rimrock Drilling Services from Elko, Nevada to drill 15, 600-foot RC pre-collar holes; | |
2. | Surface collar locations were based on the location of previously drilled and surveyed geotechnical holes. The azimuth and dip were set using a Brunton compass and measured with a tape, the designations were written on a flagged lathe; | |
3. | The drill rig set up to drill a fan pattern, the mast was checked for correct azimuth and dip prior to drilling; | |
4. | Five-inch surface casing was installed for the upper 20 feet; | |
5. | Drilling advances were paused at the end of each sample run to flush out the cuttings; |
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6. | Upon completion of the hole, International Directional Services (IDS) of Elko surveyed the collar azimuth, dip, and downhole inclinations using a gyroscopic downhole survey tool; | |
7. | The completed RC hole was cased to 600-foot with five-inch casing; and | |
8. | Variations of azimuth and dip for subsequent drilling within the fan array were based on the results of the pre-collar survey and adjusted to account for any deviation, which may have occurred. |
The 2013 surface drilling procedures for core are as summarized below:
1. | Klondex contracted American Drilling Corporation from Winnemucca, Nevada to drill surface core holes; | |
2. | Prior to drilling, the pre-collars were cemented at the bottom of the casing to seal the core hole; | |
3. | HQ diameter core was drilled with five-foot core barrels and ten-foot rod lengths; | |
4. | Core material was retrieved using a triple- tube and placed in cardboard boxes; | |
5. | Downhole surveys of the entire hole (RC and core -tail) were taken by IDS. A digital copy of the IDS report was emailed to Klondex; | |
6. | Boxes of core were transported to the logging facility for photography and collecting geological observations before being sent to the splitter for sample preparation; and | |
7. | Surface drill collar surveys were taken by Alidade, Inc. (“Alidade”) when drilling the fan was completed. |
In January 2013, authors of this report observed a sequence of handling underground drilled core as follows:
1) | Drill hole status is tracked on a dry erase board as well as in MS Excel spreadsheets for the following information: (drilled status, logging status, sampling status, dispatching a hole, and receiving assay results); | |
2) | Handling of the drilled core from the station includes: drilling with a Diamec U8 core rig (other types of drill rigs have been used in the past for drilling both surface and underground). Drillers label core box lids with a unique Bore Hole Identification number (BHID), which includes the year), box number, and drilled interval, drillers put the core in boxes (Figure 10-1) with top of drilled sequence leading the run in the box and end of drilled interval ending the run in the box. Drillers label the end of the run to the nearest one tenth of a foot and measure and record the recovery in feet on wooden blocks, which are put at the end of the drilled interval; |
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3) | Note: In 2014, the hole naming convention was changed. The final hole with the old naming convention was FC14125U. The first hole with the new naming convention was FCU-0001; |
4) | Drillers stack full core boxes on a pallet in numerical order; | |
5) | Drillers or drill helpers either deliver the pallet to surface or take a partial delivery of core boxes in the back of their motorized underground personnel vehicle. They leave the pallet of core boxes (or individual core boxes on a spare pallet) at the core logging facility; and | |
6) | Current surface drilling protocols remain similar to the Raven summary, but the protocol has evolved somewhat to allow for more precise drill targeting and data tracking. The 2014 drilling program focused primarily on additional delineation of targets and expanding the resource. These targets are more accessible and more efficiently drilled from underground drill platforms. |
10.1. | Collar Surveying |
Currently, surface hole collars are surveyed by Alidade following completion of the hole. Underground hole collars are surveyed by the mine surveyor after the drill has been removed from the drill station. When an underground collar survey is required and the surveyor is not available, the geologist triangulates the collar location using distance measurements between surveyed reference points in the drill station relative to the drill rig.
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The channel sample locations are stored as “synthetic drill holes” in the database in order to utilize them spatially with software (northings, eastings, elevations, azimuth, dip, and length). The northing, easting, and elevation of the samples are derived from geologists’ face distance measurements in relation to underground survey mapping (asbuilt).
10.1.1. Surveying Surface Drill Collar Locations
Historic surface drill collar survey data was kept in Reno, Nevada by Mr. Richard Kern of MinQuest, Inc. (“MinQuest”), as he was the Project Manager and responsible person for the database on behalf of Klondex. Klondex received the historic data in spreadsheets from Mr. Kern in May 2012. All collar northings and eastings drilled prior to 2012 came from MinQuest at that time. The elevation of the drill hole locations in the MinQuest dataset were adjusted by Mr. Steve McMillin, former Chief Geologist for Klondex, by assigning elevations from topographic contours generated from 2012 photogrammetry.
Methods used to locate collars drilled from March 2004 through December 2010 were inadequately documented, and raw data were not archived. The (non-documented) method for locating early collars was to locate the drill pad along a surveyed grid of lines (lines spaced 50 feet apart) to intercept veins as close to perpendicular as possible within the limitations of the equipment and topography.
In August of 2008, Alidade surveyed and located some of the drill pads and collars for Small Mine Development, LLC. (“SMD”). Historical survey reports for that period have not survived though Alidade’s methodology for ground control is documented in a Company memo from Alidade (Klondex, 2006):
“On our first day on the project we set a 5/8 rebar with a plastic “Alidade Control” cap on a hillside above and about a 1,000 feet north of the Project. We set up our GPS receiver on this point called “AL1”, and recorded two plus hours of static GPS data at one second intervals. This data was subsequently sent to the National Geodetic Survey (NGS) Online User Positioning Service (OPUS) and processed”.
“OPUS provided both the NAD83 Nevada Central Zone and UTM Zone 11 North coordinate values for the new point. The grid coordinates provided were expressed in meters for both systems as is standard for OPUS. We (Alidade) converted the NAD83 coordinates from meters to US Survey feet and established a coordinate system and projection for our GPS software”.
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From 2010 to the beginning 2012 (up to drill hole FC1207S), surface collar survey information was recorded by the site geologist reading a hand-held GPS device on the drill rig. Using a hand held device requires the geologist to allow the device to sit for approximately 20 minutes before a reading can be taken. The coordinates were hand-entered on a log form. The original datum is unknown. It is also not known if any conversion between datum was made as a part of this process.
All surface holes drilled since January 2012 have been surveyed by Alidade with a Trimble Real Time Kinematic (“RTK”) unit in conjunction with Global Positioning System (GPS) with a base station of a known survey point and rover unit. There are no early surviving survey reports from this methodology. The original datum is not known. It is also not known if any conversion between datum was made as a part of this process.
In June 2013, Klondex undertook to re-survey all locatable surface collar locations drilled prior to January 2012. Mr. McMillin located historically drilled holes using a ground magnetometer and a track excavator to search for buried collar casing. A total of 29 surface holes (approximately 10% of the surface drill hole population from that era) were located and resurveyed by Alidade using the current protocols. Average northing and easting errors were 5.39 and 5.71 feet, respectively. Table 10-1 contains the collar location data obtained in the re-survey.
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Table 10-1 Surface Drill Collars Re-surveyed by Klondex
BHID | Eastings | Northings | Elevation | PROJECTION SYSTEM |
FC1207S | 14696485.18 | 1737337.22 | 6055.043 | NAD 83 |
FC1208S | 14696484.23 | 1737332.01 | 6055.256 | NAD 83 |
FC1209S | 14696489.35 | 1737337.59 | 6055.136 | NAD 83 |
FC1210S | 14696480.66 | 1737331.12 | 6054.994 | NAD 83 |
FC1211S | 14696478.28 | 1737335.75 | 6055.341 | NAD 83 |
FC1212S | 14696478.45 | 1737331.66 | 6055.134 | NAD 83 |
FC1213S | 14696470.69 | 1737334.5 | 6054.263 | NAD 83 |
FC1214S | 14696471.37 | 1737331.63 | 6054.913 | NAD 83 |
FC1215S | 14696472.21 | 1737327.84 | 6055.272 | NAD 83 |
FC1216S | 14696465.29 | 1737334.07 | 6054.913 | NAD 83 |
FC1217S | 14696467.25 | 1737329.53 | 6054.438 | NAD 83 |
FC1218S | 14696464.22 | 1737333.05 | 6054.379 | NAD 83 |
FC1219S | 14696457.89 | 1737328.95 | 6054.24 | NAD 83 |
FC1220S | 14696460.78 | 1737325.61 | 6055.07 | NAD 83 |
FC1221S | 14696463.23 | 1737322.85 | 6054.835 | NAD 83 |
FC1207S | 14695832.79 | 1737596.9 | 6055.043 | NAD27 |
FC1208S | 14695831.84 | 1737591.69 | 6055.256 | NAD27 |
FC1209S | 14695836.96 | 1737597.27 | 6055.136 | NAD27 |
FC1210S | 14695828.27 | 1737590.8 | 6054.994 | NAD27 |
FC1211S | 14695825.89 | 1737595.43 | 6055.341 | NAD27 |
FC1212S | 14695826.06 | 1737591.34 | 6055.134 | NAD27 |
FC1213S | 14695818.3 | 1737594.18 | 6054.263 | NAD27 |
FC1214S | 14695818.98 | 1737591.31 | 6054.913 | NAD27 |
FC1215S | 14695819.82 | 1737587.52 | 6055.272 | NAD27 |
FC1216S | 14695812.9 | 1737593.75 | 6054.913 | NAD27 |
FC1217S | 14695814.86 | 1737589.21 | 6054.438 | NAD27 |
FC1218S | 14695811.83 | 1737592.73 | 6054.379 | NAD27 |
FC1219S | 14695805.5 | 1737588.63 | 6054.24 | NAD27 |
FC1220S | 14695808.39 | 1737585.29 | 6055.07 | NAD27 |
FC1221S | 14695810.84 | 1737582.53 | 6054.835 | NAD27 |
Additionally, surface drill hole FC1222S, drilled in late November 2012, was surveyed by Mr. McMillin using this same methodology. Currently, surface collar surveys are taken by Carl C. de Baca, of Alidade, using a Trimble RTK roving unit and base-station set on a known survey point based on projections described above.
The result of locating the 29 drill hole collars verified the historic collar coordinates for the surface holes as being accurate and within acceptable means. The authors consider the results of this study as validating the historic surface collar locations.
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10.1.2. Surveying Underground Drill Collar Locations
Underground drill hole collars are surveyed by the mine surveyor. The first phase of underground drilling began in September 2011 and continued into August 2012. Fifty-two holes were drilled during this period, all but two of from Drill Station 1. Drill collar locations were originally derived from drill station planned coordinates. Collar surveys for phase one holes were finalized in August 2012 when the drill was moved and collars were accessible to the surveyor. SMD engineer Paul Joggerst surveyed the collars (2012 Joggerst), utilizing North American Datum (NAD) 27 UTM US feet. A geologist assisted in locating each collar and identifying the borehole ID.
The 2012 Joggerst methodology included use of a robotic total station set by plumb-bob using a known survey location as datum. A survey prism was used to define each drill collar location to be recorded by the total station. 2012 Joggerst provided survey reports to Klondex in the form of electronic spreadsheets. All underground surveys were conducted in NAD 27 UTM, US feet.
Since drilling resumed in 2013, collar locations have been surveyed by the Klondex mine surveyor using Company-owned survey equipment. The Project survey equipment is a Trimble S6 DR Plus total station device used in conjunction with Leica prisms. The 2013 surveys were in NAD27 UTM US feet, and in 2014 Klondex began using NV SPCS feet.
When an underground collar survey is required and the surveyor is not available, the geologist triangulates the collar location using string and surveyed reference points. This method requires the drill rig to be in the station. The geologist ties and pulls the string between the surveyed spad points so that the string crosses the drill rig twice. The string is then measured, and the lengths are recorded relative to the drill steel. The measurements are then plotted in Vulcan using the same survey spad points. A line is drawn between the two generated points and produces an azimuth. The inclination is measured by the geologist with a Brunton compass.
In addition, the location of the collar is calculated by measuring the distances on the face or rib from the front site spad to where the drill steel enters the ground. Those measurements are also plotted into Vulcan and the Easting, Northing, and Elevation are recorded.
10.1.3. Locating Channel Samples
The coordinates of the channel samples are calculated using measurements taken by geologists. For each mining face, the geologist measures the distance along the left rib from a known reference point to the face. This distance is recorded on a daily face sheet. The channel sample is collected across the face from left to right, so the measured distance corresponds with the start of the channel. The distance recorded on the face sheet is measured on the asbuilt to find the X and Y coordinates of the sample. Because the channel samples are collected at chest height, the elevation of the channel is calculated by adding 5 feet to the sill elevation of the asbuilt.
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10.2. | Downhole Surveying |
Downhole surveys were performed for all holes drilled from 2004 to 2011 as noted by Raven et al., 2011:
“The Klondex holes are all surveyed for vertical and horizontal deviation by International Directional Services LLC, whose local office is in Elko, Nevada”. (Page 21)
IDS has continued to survey all surface holes at the Project from 2011 to present. IDS is a reputable borehole survey company with a well-established history of performing downhole surveys in accordance with industry standards.
When underground drilling began at the Project in the fourth quarter of 2011, Klondex leased a PeeWee downhole survey tool from Minex in Minnesota. The PeeWee has the option of being manually set for local declination or collecting data relative to magnetic north. Klondex collected raw uncorrected data and then applied corrections to compensate for the local declination of 13.35 degrees according to the NOAA calculator. Readings were taken by the PeeWee every 50 feet. Occasionally the raw data reflected excessive fluctuation between adjacent points, and the unreasonable point was deleted before finalizing the survey. In that case, reliable points above and below the erroneous point are used for projecting the drill hole, which is acceptable industry practice. Occasionally, the surveyor will collect “collar and quill” surveys by positioning the survey rod in the collar and recording multiple survey shots along the survey rod to measure azimuth and dip. The results can be compared to the data collected by the downhole survey tool as a rough check of the tool’s accuracy.
Since the beginning of 2014, all underground downhole surveys have been performed by International Directional Services (IDS) using a Maxibor tool.
10.3. | Core Recovery |
Core recovery has previously been described (Raven et al., 2011) and is summarized below: “Core recovery was excellent; 100% in most instances. The high-grade intervals were logged as having near or 100% recovery in nearly all cases, whether the intercept was a vein or a breccia zone. Core recovery was typically very good throughout the Klondex program.” (Page 21)
Since 2012, the percent core recovery has been calculated by measuring the material between blocks per drilled interval, then dividing the measured recovery by the run footage and multiplying that value by one hundred. The average current recovery for underground core at the Project is 95%. Drilling from underground is a more cost effective and efficient way to drill high angle veins and faults. Drill intercepts of these zones are designed to be as orthogonal as possible to best reflect the true thickness of the zones. The costs attributable to a given hole are also reduced because a single hole can be utilized to test multiple targets at a preferred elevation.
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10.4. | Security Procedures |
From early 2004 until March 2012, material from split core, rejects, RC chips, and pulps were stored in multiple storage units at the business of Security Storage, 355 East Greg Street, Sparks, Nevada. RC chip and rejects were transported directly to these storage units either from the Project or from the ALS Minerals (ALS) Lab in Sparks, Nevada. Core material was first logged at the Project by a MinQuest geologist and then transported to Sparks for cutting and sampling by a MinQuest geotechnician. After cutting and sampling, the remaining core was archived in one of the storage units.
For the 2013 core re-logging program, core was retrieved from storage units in the Sparks warehouse and moved down the street to a rented logging warehouse. Once the re-logging was complete, the core was palletized, banded, wrapped, and transported back to the Project. All rejects, RC chips, and pulps were also removed from the storage units and transported to the Project. Since March 2012, sampled materials have been handled and stored on site. Rejects and pulps are periodically returned to the Project from assay labs.
Currently, all archived sampled material is stored at the Project in a fenced area at the Rapid Infiltration Basin (RIB) yard.
10.5. | Logging Drilled Core Observations |
Drill sample logging codes at the Project have evolved over time with an increased understanding of the geology. Interpretive codes were updated, most recently in early 2014, to more accurately describe the lithology, veins, and particularly the alteration typical of an epithermal system. The new codes were adapted from similar observations at the Company’s Midas Mine and exemplify direct observations of the Project’s geology. The new codes allow for Company uniformity at similar deposits.
10.5.1. Current Logging Protocol |
Beginning June 2013, Klondex geologists began a quick log assessment prior to the detailed logging in order to quickly identify important contacts and to verify intersections or expected horizons in the core. The advantage of this additional step is an updated geologic model as soon as the core is available for preliminary review as opposed to waiting until all the logged data is collected. The quick update to the geologic model allows for modifying the drill plan in order to better intersect mineralization and to refine the mine plan.
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Detailed log data is entered into an Excel spreadsheet using standardized interpretive codes to record data such as lithology, alteration, and structure. The interpretive codes were updated in 2012. Data collected prior to the update were manually converted to match the current codes. Records of the conversion were kept.
Core is logged in the Project’s logging facility (Figure 10-2). Core is categorized as Production or Exploration.
o | Production core only receives gold-silver assay analysis; and | |
o | 48-element ICP analysis is performed on each Exploration core sample. |
10.5.2. Historic Logging Protocol
Klondex’s historical lithology database, acquired from MinQuest in 2012, contained simplified data hand-entered into RockWare LogPlot software from detailed paper drill hole data logs. The digital version of the logs lumped the tuffs and basalts into two generalized unit codes, which comprised the lithology portion of the database. The RC pre-collar and core-tail portions of the holes had separate logs.
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Klondex’s logging format was revised in 2012 with a new code system. The new codes allowed tuff and basalt lithologies to be separated into specific units to allow more detailed modeling. The 2013 re-logging program mentioned in Section 10.4 captured the new codes for historically drilled holes. Klondex re-logged approximately 240,000 feet of core to document the details of tuff and basalt units according to the new coding system and to obtain better assay resolution on mineralized intervals. Previous sampling was based strictly on five-foot sample intervals regardless of geology. This was an issue at the Project because mineralized veins typically occur within a restricted portion of a five-foot interval, and samples did not accurately reflect either the size of the vein or the distribution of gold. On occasion, veins were also misrepresented during core splitting, and the result was loss of assay opportunity. In 2013, re-logging included re-sampling of several mineralized intervals that were diluted by either being divided across intervals or represented a fraction of a five-foot interval. New sample interval footages were selected to blend into the previous sample numbering sequence without gaps or overlaps. The new sampling intervals were determined using geological observations. Better density information, multi-element analytical data and core photos were also collected.
The lithological units at the Project which contain the mineralized veins include interbedded basalt and tuff units and dikes. Klondex’s lithology database used for the resource model utilizes the new, more detailed 2014 interpretive lithological codes for these units. The unit codes used in the model were derived from current logging procedures, data converted from 2013 codes, and interpretation of the older RC Log Plot descriptive data for holes which could not be re-logged in 2013.
A direct correlation between the original logs and the current Klondex geology database is complex since the data evolves over time. The current database was converted from the 2013 codes to the 2014 codes. The 2013 codes were either logged directly as part of the re-logging program, converted from historic logging codes or derived from reading the geologists’ detailed descriptions in the comments field rather than from the lithological code.
Each of these geological logging systems was reviewed by the authors, and the results validate the geology in the Klondex database. Lithological source data for 198 channel samples were also reviewed by the authors and found to correlate well with the database.
10.5.3. Re-logging Protocol for 2012-2013
In January 2012, inadequacies in historic logging procedures became apparent. Specifically, sampling intervals were strictly five-foot regardless of interval of mineralization, observations of lithology and alteration were broadly generalized, and no core had been photographed.
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Until April 2012, core was logged at the Project and then shipped to Sparks, Nevada for processing. Split core was shelved in 23 storage units at Secure Storage in Sparks, Nevada.
In October 2012, Klondex began to re-log the core stored in Sparks before relocating it to the Project, the objectives being to:
• | Improve grade definition on veins that were diluted within a five-foot interval or divided by overlapping intervals; and | |
• | Improve detailed observations of alteration, lithology, and the stratigraphic sequence at Fire Creek. |
Two new 4,500 square foot warehouse units were rented within two miles of Secure Storage. One unit was equipped with eight roller-conveyor tables 70-foot long and two camera stands. Suspended fluorescent lighting was added to provide better lighting to compensate for ceilings 20-foot in height. The other unit was used to store the core in progress.
Twelve contract geologists and eight geotechnicians worked the re-logging program to complete the following tasks:
• | Moving core; | |
• | Washing core; | |
• | Photographing core; | |
• | Logging core; | |
• | Sampling core; | |
• | Measuring density and magnetic susceptibility of the core; and | |
• | Palletizing core for long-term storage. |
Logging core included collection of geotechnical data, such as strength, approximate Rock Quality Data (RQD) from split core, lithology, alteration, structure, mineralization, and vein density. Density measurements were taken using a water-immersion densi-meter after sealing samples in wax.
Core selection for re-sampling focused on localized alteration and vein material which were originally poorly represented by the five-foot sampling, as discussed previously. Intervals selected for re-assay were sampled by removing the remainder of the historically split core sample from the core box to be submitted for assay. Lathes marked with the interval information were left in the core box.
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Additionally, composite chip samples were collected for 48-element Inductively Coupled Plasma (ICP) analysis throughout the core on 20-foot intervals. Samples were sent to ALS in Reno and Inspectorate in Sparks for analysis.
In total, 228,814 feet of core was re-logged out of an estimated 240,000 feet. The estimated footage was based on the footage totals in the Klondex database. The difference in footages is a result of discarding core from the upper portions of the holes drilled in unaltered basalt. A Micon International Limited inventory list indicates 14,400 feet of core from 29 holes was discarded. Some of this discarded material was used for blank reference material. There are no surviving records citing how much core was used for this purpose.
10.6. | Core Sampling Methodology |
Once geotechnical and geological data has been logged, sample intervals are determined based on geology. Minimum sample interval is approximately one foot, dependent on core diameter and whether the core is split or whole core samples. Maximum sample interval is five feet. Alteration and lithologic boundaries are not crossed. Sample breaks are marked on the core, tagged on the core boxes and entered into the logging spreadsheet.
1) | Core is quick-logged in the yard to identify expected intercepts and to update the working model for ore control geologists; | |
2) | Geologists or geotechnicians set the core boxes on rolling racks in an illuminated, heated, covered plasticized canvas logging facility; | |
3) | Core is washed and verified for completeness and correct labeling of boxes and core blocks. If errors are found, they are addressed to the drilling company foreman and corrected before proceeding; | |
4) | Geotechnical data including Recovery (all holes) and RQD (even-numbered Production holes and all Exploration holes) is logged. | |
5) | Geological data is logged; | |
6) | Sample breaks are marked on the core, tagged on the core boxes and entered into the logging spreadsheet; | |
7) | The core is photographed. Core is positioned so that sample break markings, geologic features and vein/structure orientations are optimally captured in the photograph. | |
8) | After completion of all logging activities, the core is sampled. |
a. | ‘Termite’ holes (AQ or BQ diameter) are whole- core sampled due to limited material with small diameter core. | |
b. | HQ-diameter core is palletized and queued to be split and sampled in the splitting facility adjacent to the core shed. |
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c. | The geotech moves the core box into the splitting facility and splits the core in half. One half is returned to the core box, and the other half is placed in a sample bag according to the sample interval specified by the geologist. | |
d. | The core boxes are palletized, shrink-wrapped and transported to the core storage area. | |
e. | The sampled core is prepared for shipment to the assay lab. QAQC inserts are selected by the geologist. The geologist then selects the appropriate number of sample IDs from a list. Core samples are assigned sample ID of type FCD123456. The sample bags and QAQC inserts are labeled with the sample IDs and stored until they can be transferred to the assay lab. |
9) | A lab submittal form is filled out by the geologist. When enough samples have accumulated for a shipment, the assay lab driver is summoned to site. Samples are loaded on the lab truck, and the submittal and QAQC samples are handed to the driver. |
10.7. | RC Sampling Methodology |
RC samples are taken on five-foot intervals using a rotating wet splitter. Water-flow and sample size are controlled by adding or removing splitter slot covers. The number of covers are tracked for each sample.
1) | Sample bags are placed in a five-gallon bucket under the wet splitter; | |
2) | Sample buckets are placed inside a 20-inch diameter by six-inch deep rubber pan; | |
3) | If the sample bag in the bucket overflows into the pan before completion of a five-foot sample run, then the run- off is re-poured into the sample bucket to recover any fine material; | |
4) | A population of reference chips are collected in a sieve from each sample run and placed in 20-compartment sample trays; | |
5) | Buckets and pans are washed after each run, and the wet splitter is washed after each rod change; and | |
6) | A sample cut-sheet is populated with sample ID numbers and intervals, including sample IDs for QAQC samples as well. The cut-sheet tracks sample numbers on bags and intervals in the rock chip trays; |
Note: Standards, blanks, and duplicates are inserted every 20 samples. The optimum sample size collected is approximately one quarter to one half of a 17-inch by 22-inch sample bag (about 20 to 30 pounds.)
10.8. | Channel Sampling Methodology |
Channel sampling began in 2013 as underground development progressed. The dataset used for the current mineral resource estimate contains 6,691 samples collected in 1,510 face channels. Channel sampling procedure are summarized below.
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10.8.1. Channel Sampling
An ore control geologist checks the face at each round of advancement. The geologist measures the distance to the face along the left rib from a known reference point. This distance is recorded on a daily face sheet along with the geologist’s name, date and time, location, and heading dimensions. The geologist then sketches the face and records sample ID numbers in a column on the face sheet. Each sample ID has a row where sample length, rock type, unit, alteration and vein characteristics can be recorded. The geologist puts a sample bag labeled with the first sample ID in a bucket. Material is chipped from the face into the bucket, working at chest height. The channel is collected across the face from left to right. Material is collected with the goal of realistically representing mineralogy, alteration, and width of the vein. Typically, the first sample starts in waste at the intersection of the left rib and the face, then progresses from left to right towards the vein. The first sample ends near the vein margin, the sample bag is tied and set aside, and the second sample bag is placed in the bucket. The second sample is taken from the vein material. The third sample is collected from beyond the right margin of the vein to the right rib. In the case of multiple veins or otherwise complex geology, the geologist collects as many samples as necessary to characterize the face.
The channel sampling procedure has evolved over time, but the large majority of samples were collected using the current protocol. Ribs and backs may be channel sampled at the geologist’s discretion, but these samples are not included in the data set used for resource estimation.
All samples have a three letter prefix followed by a six digit number: FCF000000 = face or rib samples; FCM000000 = muck samples; FCG000000 = miscellaneous underground grab samples; muck and grab samples are not used in the resource estimation.
Once the channel samples have been collected, the geologist completes the following tasks:
1) | The geologist marks the vein margins, structures, face heading, and distance with spray paint on the rock; | |
2) | The geologist photographs the face; | |
3) | The geologist takes the bagged samples to the staging area outside the geology office and hand enters data into a central Excel spreadsheet; |
NOTE: The locations of the channels are measured from drift entrance points and recorded on face sheets and plan maps. Face sheets are scanned and filed. Channel locations are digitized with Vulcan Software. Channel collar eastings, northings, and elevations are then exported from Vulcan into CSV (comma-separated values) formatted collar files. Individual sample widths are recorded at the time of sampling. Sample width values are hand entered into CSV formatted sample files with assay results posted from laboratory reports. The channel sample files are then imported into Vulcan Software and modeled as synthetic drill holes using the eastings, northings, elevation, width, and assay values.
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4) | QAQC materials are inserted into the sample batch; | |
NOTE: QAQC samples were not utilized in the channel sample stream until June of 2013, after which blanks and standard reference material were added to each sample batch. As of January 1, 2014, standard material was no longer inserted into the sample stream, but several blank material samples are submitted per sample batch submitted. The samples are sent to the assay lab after every 12-hour shift. | ||
5) | All samples collected within a 12- hour shift are entered into a sample submittal form, which is saved on the company server and transported to the lab; and | |
6) | The Klondex lab provides a three- to four-day turn-around time between receipt of sample and assay results. If there is a delay, Klondex holds advancing the heading pending the assay results. In this event, Klondex will identify the missing sample by using an Excel sample tracker spreadsheet maintained by production geologists. |
Project staff demonstrate adequate knowledge of sampling procedures and the corresponding handling of digital data. Data handling methods implemented at the Project to manage sample data are inadequate in relation to data volume; however, the authors have reviewed the data and find that it is sufficiently accurate to be used in the mineral resource estimate. The anticipated implementation of acQuire software during 2015 will improve data management.
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11. | Sample Preparation, Analysis, and Security |
11.1. | Historic Sample Preparation |
Historical sampling methodology was previously documented (Raven et al., 2011), and is summarized below.
“Rotary cuttings are analyzed in 10-foot (3.05 meters) increments over the entire drilled interval including unmineralized rock above the vein zones. Samples in the rotary holes are collected at 5-foot (1.52 meters) intervals but assayed as 10-foot (3.05 meters) composites. The hole was blown clean between the sample intervals to avoid sample contamination. During the 2004 drilling period, cuttings were collected via a cyclone that dumped into a rotating splitter mounted on the drill. The baffles were adjusted to recover a one-quarter split of the total recovered sample. More recently, the 10-foot (3.05 meters) runs of cuttings have been caught in a large bucket and thoroughly mixed by hand before collecting a sample. The approximately 20-pound (9.1 kilograms) samples are placed in canvas bags and labeled with the hole number and footage. A backup sample remains at the Project until assaying is complete and is then discarded. The samples are picked up by ALS/Chemex for preparation at their Elko facility.
“Below the RC precollar boring, HQ size core is drilled and collected in 10-foot (3.05 meters) paper core boxes. Intervals are marked with wooden blocks every two to three feet (0.6 to 0.9 meters). The core is logged on site by a MinQuest geologist who marks sample intervals not to exceed five feet (1.52 meters). In some vein areas, where possible visible gold is observed, the sample interval is reduced to two feet (0.6 meter). The logged and marked core is transported from the Project by the geologist, to secure storage in Battle Mountain. Under the supervision of a Project geologist, the core is transported to Elko and split in half using a core saw by Klondex employees. One-half of the core is sampled on the intervals marked by the geologist, placed in canvas bags, labeled with the hole number and footage and sent to the lab for preparation and analysis as described below. The remaining one-half core is transported to Klondex’s secure storage in Reno. The sample intervals are listed on the drill logs and assay sheets. Author Raven observed numerous intervals of split core, all of which were cleanly sawn in half and appear to evenly represent the vein systems and the sample intervals are clearly marked within the core boxes. The sample quality is of industry standard, and the methods should not introduce any bias into the results. The sampling intervals are determined mainly by the presence/absence of quartz-calcite-pyrite veins or vein stockworks. The barren, upper portions of many holes are not sampled. When veining is encountered a broad interval, above and below. The veins is sampled, and the vein zone itself is sampled at intervals of two to five feet (0.6 -1.52 meters); discrete veins of reasonable size are sampled over the length of the vein while stockwork zones are generally sampled at five-foot (1.52 meters) core lengths.” (Page 23)
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11.2. | Current Sample Preparation |
11.2.1. Core Sample Preparation |
The core sampling facility is set up in a shipping container adjacent to the core logging facility. It is furnished with industry typical sampling apparatus including roller tables and a hydraulic splitter. The following outlines core sample preparation:
1) | A geotechnician positions the pallet containing the core to be sampled near the shipping container and obtains a copy of the sample intervals from the geologist. The geotechnician labels cloth sample bags according to the sample interval sheet; |
2) | The core boxes are lifted onto a rolling counter to the left of the splitter. A sample bag is placed on the floor at the feet of the geotechnician to hold the sample material; |
NOTE: It is possible for empty pre-labeled sample bags to be out of order prior to being filled or a numeric value to be omitted during hand-writing.
3) | The geotechnician splits core to approximate 50% of the sample bisecting veins equally. Geologists supervise the splitting of samples that contain visible gold (VG); |
4) | The left half of the split is returned to the core box, the right is placed into the sample bag; |
5) | When the sample interval has been bagged, the sample bag is stacked in numeric order on the floor by the door; |
6) | QAQC samples are bagged and labeled by geologists from standards kept in a locked cabinet in the Geology office. The geologists assemble the standards and blanks into corresponding sample bags which are hand-labeled according to the cut sheet; |
7) | When an entire drill hole has been completely split, the bags of sample are stacked inside a large, open, plastic bin outside the core facility; |
8) | The geotechnician notifies the geologist when a hole is ready to be sent to AAL (as defined below). An electronic sample submittal sheet is entered into the computer. Two copies are made, one is the original hand-entered submittal, and the other is a scan of the completed submittal. One copy is filed in a core library, and the other is given to the truck driver for AAL; |
9) | The entire bin of samples is picked up and delivered to AAL by the AAL driver; When the driver from AAL arrives at the core logging facility, he is given the QAQC samples to accompany the samples from the corresponding drill hole; and |
10) | The reserved halves of core are returned to their core boxes and are stored outside on shrink wrapped pallets in a fenced lay down area referred to as the ‘RIB Yard’. |
11.2.2. Channel Sample Preparation
The following outlines the channel and sample preparation methodology.
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1) | Channel samples are bagged on site at the face as described in Section 10.8; | |
2) | Bags are brought to the Geology office; | |
3) | QAQC materials are inserted into the channel sample stream; and | |
4) | Channel samples are delivered to the Klondex assay lab every 12- hour shift. |
11.3. | Sample Analysis Protocol |
11.3.1. Historic Drill Sample Analysis
The sample analysis methods used from 2004 through 2011, as previously described in Raven et al., 2011:
“ALS/Chemex does all sample preparation, including crushing, grinding and preparation of the assay pulps, at the Elko facility. The pulp samples are then shipped to the ALS/Chemex facility in Reno for analysis. The samples are never left unattended or insecure by geologic, drilling, or laboratory staff nor are they handled by officers, directors or associates of Klondex. For the RC pre-collar holes ALS/Chemex picks up the samples at the Project and delivers them to Elko for sample prep and to Reno for analysis. After the core samples are cut and labeled for analysis they are delivered to the lab by Klondex employees”. (Page 25)
“Sample preparation involves crushing the entire sample to minus 10 mesh, splitting, then pulverizing 1,000 grams to 80% passing minus 200 mesh (75 microns). These pulps are shipped to the Reno facility of ALS/Chemex for analysis. Analyses for gold were done using a 50-gram charge through to the end of 2009. In 2010 Klondex changed to a 30-gram charge for gold analysis after reviewing the data. Both gold and silver analyses are determined by fire assay with an AA finish. The ALS/Chemex analyses codes are AA23 for gold values under 10 grams per ton (g/t) and GRA (gravimetric) for gold assays over 10 g/t; silver codes are AA61 with over limits run using AA62”. (Page 25)
“The assay laboratory automatically repeated all gold assays that by fire assay with AA finishing reported under one g/t, using 50 grams prior to late 2010, then 30 grams fire assaying subsequently. Any samples reporting under 10 g/t gold by fire assay with AA finish are automatically subjected to gravimetric analysis.” (Page 25)
“When the lab work is complete, the pulps are stored briefly at the lab then transferred to Klondex’s secure storage facility, the same facility that houses the drill core. Coarse rejects that reported significant gold are stored with the pulps, those reporting minimal gold are stored until check assays can be completed and are then discarded and those reporting insignificant gold are discarded”
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“Until late 2010, Klondex did not employ a QA/QC program. Prior to that time, the only QA reporting was derived from the commercial laboratory’s internal QA programs that included internal blanks and standards, and automatic re-assays of pulps in which the gold grades exceeded one g/t. In addition a significant number of samples were sent to a different laboratory for check analysis. Subsequently Klondex has initiated its own internal quality control procedures. Presently Klondex has prepared blank samples using post-mineral basalt core from well above the mineralized zones. In addition two standards were prepared (low and medium grade) by ALS from assay rejects and there have now been enough analyses of the standards to determine their average grade and standard deviation. Since these “standards” were not subjected to multiple assaying by ALS to report the laboratory’s mean and standard deviations, the determination of the statistical quality of these standards has derived solely from a modest population of standards that have been submitted by Klondex in the ongoing drilling program.” (Page 25).
11.3.2. Drill Sample Analysis from 2012 through April 30, 2014
From 2012 until April 30, 2014, Klondex specified that ALS follow sets of assay procedures based on ranges of assay values. For samples with visible gold, Klondex submitted samples to ALS for a metallic screen fire assay. All other samples were initially run with Atomic Absorption fire assay fusion analytical method (AA23). Samples with AA23 results between one ppm Au and 10 ppm Au were re-run as an AA23 duplicate. Samples with an initial result greater than 10 ppm Au up to 20 ppm Au were re-assayed with gravimetric finish. If the assay results were very high grade (greater than 20 ppm Au), then ALS would re-assay the coarse rejects of the high grade sample and the two samples on either side by metallic screen fire assay.
11.3.3. Current Drill Sample Analysis
The drill sample analysis protocol was amended as of May 1, 2014. Drill samples are currently submitted to American Assay Laboratories Inc. (AAL) of Sparks, Nevada. AAL is an ISO/IEC 17025:2005 accredited laboratory. Five assay procedures have been established, one for RC samples and four for core samples. Core samples are assayed according to the designated purpose of the drill hole (Exploration or Production) and grade of the sample. The drill sample analysis protocols are as follows:
RC sample analysis procedure:
Samples are received and dried in-bag at 85° C. The dry sample is crushed to 70% passing minus 10 mesh. The crusher is cleaned with compressed air between each sample. A 1,000 gram pulp is collected from the crushed sample using a rotary splitter. The remainder of the sample is stored and returned to Klondex. The pulp is then pulverized to 85% passing minus 200 mesh. The pulveriser is cleaned with compressed air between each sample. Thirty grams (g) of pulverized sample is used to perform fire assay with ICP finish for gold, and 0.5 g of sample is used to perform analysis for silver with ICP finish. If the result is greater than 10 ppm Au or greater than 100 ppm Ag, then 50 grams of the pulverized pulp is used to run a fire assay for Au and Ag with gravimetric finish. If the gravimetric result is greater than 10 opt Au, then the remaining pulp is screened at 150 mesh for a metallic screen fire assay for Ag and Au with a gravimetric finish. Pulps are stored and returned to Klondex.
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Core sample analysis procedure:
All core samples are received and dried in-bag at 85° C. Samples are crushed to 80% passing minus 10 mesh with a crusher clean-out between each sample. A 1,000 g pulp is taken from the crushed sample using a rotary splitter. The pulp is pulverized to 85% passing minus 200 mesh with a pulverizer clean-out between each sample. The pulps are then assayed according to the designated purpose of the drill hole (Exploration or Production) and whether a high grade result (Au less than 10 opt) is anticipated. All pulps and rejects are returned to Klondex.
Production samples:
For production hole samples which are not anticipated to be high grade, 50 g of the pulp is used for a fire assay for silver and gold with a gravimetric finish. For any sample with a result less than 10 opt Au or Ag, the remaining pulp is re-run as metallic screen fire assay for silver and gold with a gravimetric finish.
High grade production samples:
For production hole samples with visible gold or other high grade characteristics, the entire pulp is screened at 150 mesh and analysed with metallic screen fire assay for silver and gold with gravimetric finish.
Exploration samples:
For exploration hole samples which are not anticipated to be high grade, 50 g of the pulp is used for a fire assay for gold with ICP finish, and 0.5 g of sample is used to perform analysis for silver with ICP finish. Any sample with a result of less than 10 ppm Au or less than 100 ppm Ag is rerun using 50 g of pulp with fire assay for silver and gold with a gravimetric finish. For a gravimetric result of less than 10 opt Au, the remaining pulp is used for a metallic screen fire assay for silver and gold with gravimetric finish.
For any sample with a result less than 10 opt Au, the remaining pulp is re-run as metallic screen fire assay for silver and gold with a gravimetric finish.
High grade exploration samples:
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The procedure for high grade exploration samples is similar to the procedure for other exploration samples, except when more than trace amounts of gold and silver are expected, the fire assay with ICP finish is skipped and the process starts with a 50 g fire assay for gold and silver.
11.3.4. Channel Sample Analysis
Channel samples were sent to SGS North America, Inc. in Elko, Nevada from June 16, 2013 to April 30, 2014. Analysis followed the following protocol:
• | Sample material is dried. Samples weighing more than three kilograms (kg) are split down to three kg then crushed to 75% passing through a two mm screen. Material is split down to 250 g, pulverized to 85% passing through a 75 micron screen; | |
• | QC is performed at the crush and pulverization stages; | |
• | Silver is analyzed by AA methods after a multi -acid digest at a weight of two grams; | |
• | Gold is analyzed by FA with gravimetric finish at a weight of 30 g (the reported code is F 152); and | |
• | Gold is analyzed by FA and gravimetric finish at a weight of 50 g (the reported code is F 133). |
In June 2013, the split was increased to 1,000 g, and the initial fire assay aliquot was increased to 500 g. Rejects for April through June 2013 were sent to SGS’s Vancouver office for metallic screen assays. Results for these assays were incomplete and are not used in the mineral resource model.
Between May 1, 2014 and July 16, 2014, samples were sent to Dave Francisco lab in Fallon,Nevada. Dave Francisco lab followed the same procedures currently used by the Klondex lab. Both labs follow the 17025 Standard, but neither has official lab certifications. QAQC samples support the results from both labs.
Beginning July 17, 2014, the Project sends channel samples to the Klondex lab for analysis. Sample protocol is as follows:
Samples are dried in pans at 250° F. The dried samples are crushed to 80% passing 10 mesh, with a crusher clean-out between each sample. The crusher is cleaned twice following high grade samples. The crushed sample is homogenized 500 g is collected with a riffle splitter then pulverized to 85% passing 200 mesh. The pulverizer is cleaned after every sample, twice after high-grade samples. For 10% of samples, a second pulp is prepared as a preparation duplicate. Remaining coarse rejects are stored.
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Fifty grams of the pulverized pulp is used to run a fire assay for gold and silver with gravimetric finish. In each batch of assays, the lab inserts a standard and blank. The lab also runs five percent of samples as analytical duplicates. Samples with result less than 2.92 opt Au are run with metallic screen fire assay with gravimetric finish.
11.3.5. Handling Analyses Results
1) | AAL sends the assay results and certificates by email to three people: Chief Geologist, Senior Geologist, and Geology Database Administrator. For channel samples, Klondex Lab emails results to these people as well as the ore control geologists; | |
2) | Assay results from AAL are stored as portable document formats (PDF) and MS Excel files on the Klondex server in a hierarchy of folders with a naming convention based on designation of sampled material. Results from the Klondex lab are stored as MS Excel files. Folders include channel samples, UG core, surface core, surface RC, screen filter sampling, truck load samples, rib sampling, muck piles, waste piles, and resamples of these same sources. This folder system is rudimentary and not user-protected; | |
3) | The PDF and Excel files from AAL are renamed to add the BHID for identification and for ease in referencing; | |
4) | Excel files for use in Project modeling software are updated as assay results are finalized by the lab by means of copy and paste from the lab Excel files into the user Excel files; |
NOTE: There are no conversions made, such as from parts per billion [ppb] to ppm.
5) | Excel files undergo extensive manual manipulation and editing. The editing progress is tracked with removal of the dashes from the BHID. These are cumbersome, labor- intensive edits. An acQuire database is currently being built to manage these datasets more efficiently; and | |
6) | The Project has an internet connection that is unreliable. There is no Systems Administrator to plan for or implement an enterprise system for data management, or virtual private network (VPN) connection between the Project and Elko office servers. An enterprise system would create a secure, user-password protected connection for immediate communication and for automated backups. At this time, data is transported on external hard drives between the Project |
11.4. | Sample Security Measures |
Drilled materials are stored under a moderate level of security during the multiple stages of sample handling. Core is handled and stored at the Project, which is staffed by security personnel. Core boxes are stored in the vicinity of the logging facility during the logging and sampling process. The logging facility is not fenced, but geology staff is present on-site during most shifts and are frequently within view of the core facility. Sampling of core with visible gold is supervised by geologists. When sampling is complete, retained core samples are returned to boxes, stacked on pallets and shrink wrapped. The wrapped pallets are moved to a fenced facility at the “RIB yard”. Coarse rejects and pulps returned by the laboratories are also shrink wrapped on pallets and stored at the RIB yard. The authors conclude that sample security measures at the Project are adequate.
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11.5. | Quality Control Measures |
Historically, QAQC measures used to check the consistency in assay reporting were either lacking or not included in any surviving reports. Beginning in March 2004 through the second quarter of 2012, Klondex samples were submitted to ALS and were reliant solely on the laboratory’s in-house QAQC to monitor the sampling results. The current practice of inserting blanks and standards and specifying prep duplicates began in the second quarter of 2013 when Klondex began processing core on site. Prior to this time, core was transported to Reno for cutting and sampling, and any QAQC measures were directed by MinQuest in Reno.
From March 2004 through February 2012, ALS’s QAQC checks on the Project samples included 12,465 in-house standard samples inserted into the Klondex sample runs and 11,201 re-assays of the immediately previous sample as part of their protocols. Also, beginning in August 2010 through February 2013, ALS completed 1,264 in-house check duplicates derived from pulp of the sample prepared for Project sample runs. Recently, ALS sent a summary of their in-house QAQC sample results to Klondex as part of recording QAQC documentation. Their report combines sample results from both surface and underground drilling.
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The populations of datasets for ALS in-house QAQC sampling are itemized in Table 11-1 below.
Table 11-1 ALS In-house QAQC Datasets Reviewed
ALS internal | ALS | ALS internal | SRM Au | |||
QC | internal QC | QC prep- | and Ag | |||
Datasets: | standards | dups (March | dups (Aug. | standards | Klondex | Klondex |
(March 2004 | 2004 - Feb. | 2010 - Feb. | (Nov. | standards*1 | duplicates*2 | |
- Feb. 2013) | 2013) | 2013) | 2010) | |||
mixed | mixed | mixed | ||||
UG Core | surf+ug | surf+ug | surf+ug | 0 | 193 | 77 |
Surface | mixed | mixed | mixed | |||
RC/core | surf+ug | surf+ug | surf+ug | 94 | 152 | 39 |
Totals | 12465 | 11201 | 1264 | 94 | 345 | 116 |
*Surface standards and dups dates: June 2012 - Jan. 2013 | ||||||
*UG standards and dups dates: August 2012 - May 2013 |
11.5.1. | QAQC Prior to 2012 |
Historic data validation has previously been addressed (Raven et al., 2011). A summary of their work includes:
“…Until late 2010 Klondex did not employ any submitted sample based QAQC program. Prior to that time, the only QA reporting was derived from the commercial laboratory’s internal QA programs that included internal blanks and standards, and automatic re-assays of pulps in which the gold grades exceeded one g/t. In addition a significant number of samples were sent to a different laboratory for check analysis. Subsequently Klondex has initiated its own internal quality control procedures. Presently (2011) Klondex has prepared blank samples using post-mineral basalt core from well above the mineralized zones. In addition two standards were prepared (low and medium grade) by ALS from Fire Creek assay rejects and there have now been enough analyses of the standards to determine their average grade and standard deviation.” (Page 25)
“…A blank and two standards are now included in each drill hole as standard practice.” (Page 25)
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“… A review of the data from the 2010 drilling campaign that made use of the new QAQC procedures did not outline any difficulties with the new standards and blanks that would indicate an error at the lab. The check assays performed on drill core samples that assayed under one g/t gold show good agreement between the original assay and the check assay.” (Page 27).
“…The ALS/Chemex facility at Elko is certified to ISO 9001:2008 standards and only handles sample receiving and preparation. The ALS/Chemex facility in Reno provides a broader range of analytical services and is also certified to ISO 9001:2008 Standards; in addition it has received accreditation to ISO/IEC 17025:2005 from the Standards Council of Canada (SSC) for Fire Assay gold by Atomic Absorption, which is the analytical method Klondex utilizes for its gold analyses.” (Page 27)
“…All gold assays in excess of one g/t are rerun at least once. A large number of gold reruns are also carried out where values are less than one g/t. These were either on samples adjacent to intervals with elevated gold assays, on samples with elevated silver values and low gold, or at the discretion of the geologist when lithologic characteristics were suspect.: (Page 29)
“…samples with greater than 10 g/t gold were rerun using a 50 gram fire assay with gravimetric finish (ALS-Chemex Au-GRA22 procedure) to late 2010 then a 30 gram charge subsequently.” (Page 29)
“…The checked assays are usually in good agreement with the original assay indicating no significant nugget effect.” (Page 29)
“…Additional check assays have been received from the 2009 and 2010 drilling campaigns and they show a similarly good correlation between the original assay and the duplicate, or check assays.” (Page 29)
“…There have been approximately 4,000 duplicate samples submitted for check analyses as part of the QAQC program.” (Page 31)
“…Klondex undertook some umpire assays at different laboratories to verify a portion of the higher grade results and compared analytical methods for gold by fire assay with an AA finish vs. a gravimetric finish. Silver was also included in the analysis between the two labs.” (Page 32)
“…The authors (Raven et al., 2011) verified a portion of the drill core data by re-assaying sample pulps sent to SGS Mineral Services in Vancouver, British Columbia. The SGS laboratory is an ISO 9001:2008 accredited facility. Coarse reject material for all the samples selected was not available so sample pulps were chosen over splitting the remaining core. The samples selected for verification were from a broad range of drill holes and designed to test various grades of mineralization from low- to –high grade.” (Page 34)
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“…There is a good agreement between the original values vs. the check assays as noted in the charts above for nearly 4,000 check samples and it is felt that this correlation is sufficient and demonstrates that while there are spurious values indicating some nugget effect, in most cases the nugget effect is minimal.” (Page 36)
“…Author Raven did note in the drill core and corresponding assay results for those intervals that the better gold grades are confined to intervals containing quartz +/- carbonate veining, either larger (less than 1.5 feet) discrete veins or stockwork systems of veining. Klondex has assayed numerous intervals of visually barren mafic volcanics (no veining, fracturing or faulting) and those intervals do not return anomalous gold assay.” (Page 36)
11.5.2. | Current QAQC Procedures |
From 2012 through March 2014, Klondex’s QAQC protocol at the Project was to submit a blank as the first sample of each drill hole, followed by one of three types of QAQC standards every 20th sample in the sample stream. Beginning April 2014, geologists insert QAQC standards as five percent of the sample stream. The type and location of each standard is at the geologist’s discretion. At least one QAQC sample is inserted per hole. The three standard types are 1) blank, 2) standard, or 3) duplicate;
1) | Blanks are crushed, homogenous barren material. Their IDs and values are as follows: |
FCRDBLNK01= <0.005 ppm Au (reduced);
FCOXBLNK01= <0.005 ppm Au (oxidized) and;
AUBLANK54 = <0.002 ppm Au.
2) | Klondex uses several QAQC standards. Some were produced in- house from locally derived low-grade basalt. Others were purchased from ROCKLABS, a reputable supplier of reference material. Standard IDs and values are as follows: |
FCRDLOW01 1.246 ppm Au
OXQ90 24.88 ppm Au
OXP91 14.82 ppm Au
OXN92 7.643 ppm Au
SG56 1.027 ppm Au
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SN60 8.596 ppm Au
SP59 18.12 ppm Au
SQ48 30.25 ppm Au
SQ83 30.64 ppm Au
SQ70 39.62 ppm Au, 159.5 ppm Ag
3) | For duplicate sampling, Klondex submits an empty bag labeled with the required sample ID in sequence. The lab takes a split from the pulp of the previous sample to run as a duplicate. |
11.6. | QAQC Analysis |
11.6.1. | Duplicates Performance |
Three sets of duplicate assays are available for review. The first set is Au assays from ALS and is shown in Figure 11-2. These values agree quite well with the 95% confidence intervals bracketing the ideal one to one trend line.
The second and third set of duplicate assays are gold and silver values from AAL. The gold assays exhibit a slightly low bias with the upper 95% CI plotting below the ideal trend (Figure 11-3). The silver assays show a larger high bias with the lower 95% confidence limit plotting above the ideal trend line. (Figure 11-4)
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11.6.2. | Blank Assay Performance |
Table 11-2 shows the results from both ALS and AAL of all blank assays from July 2012 to December 2014. These results are displayed graphically in Figure 11-5 through Figure 11-18. Most of the values reported are at one half the detection limit for the method used, and sample dilution or assay errors at either lab occurs infrequently.
Table 11-2 Blank Assay Set Performance
Designation | Count | Mean | Std. dev. |
ALS-AUBlank54 | 67 | 0.009 | 0.011 |
AAL-AUBlank54 | 28 | 0.052 | 0 |
ALS-FCBlank02 Au | 10 | 0.038 | 0.029 |
AAL-FCBlank02 Au | 144 | 0.022 | 0.061 |
ALS-FCBlank02 Ag | 10 | 1.185 | 1.137 |
AAL-FCBlank02 Ag | 144 | 3.344 | 0.525 |
AAL FCBlank03 Au | 117 | 0.034 | 0.025 |
AAL FCBlank03 Ag | 116 | 2.403 | 1.537 |
AAL FCBlank04 Au | 90 | 0.056 | 0.074 |
AAL FCBlank04 Ag | 90 | 0.270 | 1.420 |
ALS FCRDBLNK01 | |||
Au | 144 | 0.037 | 0.145 |
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Designation | Count | Mean | Std. dev. | |
ALS | FCRDBLNK01 | |||
Ag | 140 | 0.296 | 0.287 | |
ALS | FCOXBLNK01 | |||
Au | 55 | 0.007 | 0.01 | |
ALS | FCOXBLNK01 | |||
Ag | 55 | 0.271 | 0.089 |
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11.6.3. | Standards Performance |
Table 11-3 shows the results of all standard assay sets analyzed from July 2012 through December 2014. Results from AAL were slightly better than ALS with AAL test groups returning lower standard deviations. The results of all standard assays are presented in Figure 11-19 through Figure 11-34. Failure of the t-test to accept the hypothesis does not always indicate the negative, but rather that the results are inconclusive one way or the other.
Table 11-3 Standard Assay Performance
Standard | Standard Value | Count | Mean | Std. dev. | T- statistic | T 0.95 | Comment |
ALS-FCRDLOW1 | 1.246 | 100 | 1.285 | 0.127 | 3.095 | -1.984 | Reject |
AAL-FCRDLOW1 | 1.243 | 109 | 1.234 | 0.069 | -1.742 | -1.982 | Accept |
ALS-OXN92 | 7.643 | 28 | 7.517 | 0.215 | -3.105 | -2.052 | Reject |
AAL-OXN92 | 7.643 | 21 | 7.619 | 0.156 | -0.709 | -2.086 | Accept |
ALS-OXP91 | 14.82 | 18 | 14.622 | 0.332 | -2.525 | -2.110 | Reject |
AAL-OXP91 | 2 | ||||||
ALS-OXQ90 | 24.88 | 16 | 24.825 | 0.326 | -0.676 | -2.131 | Accept |
AAL OXQ90 | 24.88 | 4 | 24.949 | 0.477 | 0.288 | -3.182 | Accept |
ALS-SG56 | 1.027 | 41 | 1.018 | 0.032 | -1.764 | -2.021 | Accept |
ALS-SN60 | 8.596 | 55 | 8.419 | 0.222 | -5.900 | -2.005 | Reject |
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Standard | Standard Value | Count | Mean | Std. dev. | T- statistic | T 0.95 | Comment |
AAL-SN60 | 8.596 | 149 | 4.977 | 4.18 | -10.569 | -1.976 | Reject |
ALS SP59 | 18.12 | 38 | 17.646 | 0.468 | -6.238 | -2.026 | Reject |
AAL-SP59 | 18.12 | 7 | 18.092 | 0.266 | -0.274 | -2.447 | Accept |
ALS-SQ48 | 30.25 | 16 | 29.963 | 0.401 | -2.865 | -2.131 | Reject |
ALS-SQ70 Au | 39.62 | 9 | 41.378 | 6.693 | 0.788 | -2.306 | Accept (one outlier) |
ALS-SQ70 Ag | 159.5 | 10 | 160.2 | 4.59 | 0.482 | -2.262 | Accept |
AAL-SQ70 Au | 39.62 | 2 | |||||
AAL-SQ70 Ag | 2 | ||||||
ALS-SQ83 | 30.64 | 15 | 29.947 | 0.426 | -6.308 | -2.145 | Reject |
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11.7. | Opinion on the Adequacy of the Sampling Methodologies |
Project staff have shown a solid understanding with regard to management of the sampled material and associated digital data. The methods of handling the drilled material, both physically and electronically, are acceptable for use in an analysis of the potential mineral resource; however, system improvements exist that should be implemented in the future, as the Project develops.
11.7.1. | Sampling Protocol Issues |
The electronic data is rudimentarily compiled in a series of Excel spreadsheets. This method of storing and editing data is susceptible to multiple errors throughout the data management process. In particular, the FROM-TO intervals contain overlap errors, and raw data is cut and pasted by hand into non-secure spreadsheets. An AcQuire database is currently under construction and near completion, which will abate this issue.
Klondex has plans to implement AcQuire in 2015. This software system will address all data management issues.
There are Systems Administrators currently planning and implementing an enterprise system solution for data management linking servers through VPN connection and implementing a user-protected access to folders.
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As of the date of this Report, a direct assay certificate-to-spreadsheet-values audit has upheld the integrity of the spreadsheet data for a reliable mineral resource estimate for preparation of this Report.
11.7.2. | Standards and Blanks Performance Issues |
Duplicate assay checks performed by AAL showed no bias, while those of ALS had minor biasing. This may be the result of too few duplicate check assays available for review, and the database needs to be enlarged. Additionally, the samples chosen for duplicates are all below the cutoff grade for the deposit, and this program should be expanded to include higher grade samples as well. A third lab should be engaged to perform duplicate assays on samples originally assayed at one of the other labs as another means of quality assurance.
The blank data collected and used by Klondex does not present any underlying problems with sample handling, assay methods or laboratories. As a matter of routine, whenever a blank assay outside of acceptable limits is received, the entire assay set should be re-assayed, and the initial results replaced with the succeeding results.
Review of assay standards sets shows AAL to have smaller deviations than ALS, however, the results do not show any problems with the underlying data.
The authors’ opinion is that Klondex’s current QAQC program, for sampling protocols, is managed in an acceptable manner. QAQC verification does not indicate any underlying deficiencies in the database.
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12. | Data Verification |
The authors analyzed the sample data used in the mineral resource estimation to verify its suitability for use in this Report. The dataset includes records of drilled and channel-sampled material collected from 2004 Through December 2014 and compiled by Klondex into centralized master spreadsheets. Ms. Hazel Reynolds, Klondex Geology Data Administrator, provided the authors with a copy of the master spreadsheets. The authors formatted the data and loaded it into Vulcan ISIS databases. The authors chose a representative semi-random subset of the ISIS data, representing at least five percent, and requested the corresponding raw data source files, which were provided by Ms. Reynolds. The accuracy of the data was verified by comparing the values in the ISIS databases to the values in the original source files. The raw assay data contained in the source files has been determined adequate for use in the mineral resource estimation as discussed in Section 11.5.
Two ISIS databases were used to estimate the mineral resource: one database was compiled from drilled material and the other from channel-sampled material. The drilled material dataset contains data from surface holes drilled from March 2004 through December 2014 and from underground holes drilled from September 2011 through December 2014. The channel sample dataset contains data collected from April 2013 through December 2014.
12.1. | Results of Drill Data Review |
The four categories of data reviewed for the drill dataset are collar location surveys, down-hole surveys, assays and geology.
• | Collar location surveys reviewed: 68 surveys of underground hole collars and one surface collar survey were reviewed, representing about 10% of the holes in the dataset; |
• | Downhole surveys reviewed: 42 downhole surveys of underground holes and 24 downhole surveys of surface holes were reviewed, representing about nine percent of the holes in the dataset; |
• | Geology review: geology logs were checked for 80 underground holes and 212 surface holes, representing about 42% of the holes in the dataset; and |
• | Assay review: original assay result certificates were reviewed for 109 underground holes and 149 surface holes, representing about 37% of holes in the database. |
Table 12 1, below, summarizes the numbers and percent of drill holes reviewed for this report:
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Table 12-1 Data Review Summary Drilled Material
Datasets for: | Total drill | Holes: | Holes: | Holes: | Holes: Assay |
holes | Collar | Downhole | Geology | certificates | |
Survey | Survey | reviewed | reviewed | ||
reviewed | reviewed | ||||
UG Core | 382 | 68 | 42 | 80 | 109 |
Surface | 307 | 1 | 24 | 212 | 149 |
Core/RC | |||||
Totals | 689 | 69 | 66 | 292 | 258 |
Percent of population reviewed: | 10% | 9% | 42% | 37% |
12.1.1. | Collar Location Checks |
The authors compared 13 underground collar survey reports to collar easting, northing, elevation and TD values in the database and found 100% correlation for holes drilled since August 2012. Collar locations of underground holes drilled prior to August 2012 are considered reliable as discussed in Section 10.1.2.
The authors compared one surface collar survey report to the collar easting, northing, elevation and TD values in the database and found 100% correlation for holes drilled since 2012. A majority of the surface holes were drilled prior to the 2013 surface collar re-survey; surface collar locations for holes drilled before 2012 are considered reliable as discussed in Section 10.1.1.
12.1.2. | Downhole Survey Checks |
The authors compared 42 downhole survey reports for underground holes with the depth, azimuth and dip values in the database. Some data mismatches exist between the raw azimuth data and the azimuth column of the database because the downhole survey apparatus used prior to 2014 did not automatically adjust for local declination. Geologists adjusted the declination before entering the data in the master spreadsheet. Declination was adjusted correctly for all reviewed holes, yielding a 100% correlation.
The authors compared 24 downhole survey reports for surface holes with the depth, azimuth and dip values in the database. A total of 595 records were checked and 24 mismatches were found, yielding a 96% correlation. The authors consider the data to be reliable.
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12.1.3. | Geology Checks |
The authors compared geology logs for 80 underground holes and 212 surface holes to the database. A direct correlation between the original logs and the current Klondex database is complex. The current database was converted from the 2013 codes to the 2014 codes. The 2013 codes were either logged directly as part of the re-logging program or derived from historic RC Log Plot data or from reading the geologists’ detailed descriptions in the comments field rather than from the lithological code. Each of these geological logging systems was reviewed by the authors, and the results validate the geology in the Klondex database. The vein flag, which is the component of the database which directly affects the resource model, was found to have 100% correlation for holes reviewed.
12.2. | Results of Channel Sample Data Review |
The authors reviewed 198 channels, representing about 12% of the 1,554 channels in the ISIS channel sample database. The channels were chosen at random while generally attempting to select a representative subset. The authors requested the raw data, which is in the form of the geologist’s daily face sheets, for the 198 selected channel samples. Ms. Hazel Reynolds provided scans of the face sheets. The three categories of data reviewed for the channel sample dataset are location, assays, and geology.
12.2.1. | Location Measurement Check |
The authors compared the location of the channel in Vulcan software with the distance measured by the geologist in the mine heading and recorded on the face sheet. No channels were found out of place. The authors also viewed all channels relative to the asbuilt in 3-D in Vulcan as described in Section 10.1.3 to check for consistency. The authors consider the channel locations to be acceptable for use in the mineral resource estimation.
12.2.2. | Geology Check |
The authors compared geology data recorded on the face sheets to geology data in the ISIS database and found the data to be congruent. No errors were found in the vein flag portion of the data. The authors consider the geology data in the channel database to be acceptable for use in the mineral resource estimation.
12.2.3. | Assay Check |
The authors reviewed assays for 198 channels, about 12% of the channels in the channel database.
A total of 874 samples in the channel database, comprising about 13% of the samples, lack silver values. Samples with missing silver values are from the earlier part of the channel sampling program and are generally restricted to the first cuts on the Joyce Vein and Vonnie Vein. Channel samples from more recent mining in areas adjacent to those with missing silver values provided near by data points for use in the silver mineral resource estimation. The volume and spacing of data, coupled with the fact that the silver-gold ratio is less than one, minimizes the effect of the missing silver values. Silver analysis was completed but not uploaded to the database for most of the samples with missing values, and it is expected that the data will be incorporated when Klondex launches their AcQuire data management system.
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Table 12-2, below, summarizes the numbers of channels reviewed for this report:
Table 12-2 Data Review Summary Channel Sampled Material
Datasets for: | Total channels | Channels: Location measurements reviewed | Channels: | Channels: Assay certificates reviewed |
Channels | 1,553 | 198 | 198 | 198 |
Percent of population reviewed: | 12% | 12% | 12% |
The authors consider the assay data in the channel database to be acceptable for use in the mineral resource estimation.
12.3. | Summary of Database Verification |
For each data set used in the mineral resource estimate, at least five percent of the data was verified against original source data. The data review verified that historic and current drill, channel and control samples are acceptable. In particular, the accuracy of the assay data has been quantified by independent review of 37% of drill holes and 12% of channels by direct correlation with assay certificates from accredited laboratories (drill samples) and accredited and local production laboratories (channel samples).
The drilling (fc_5sept2014.ddh.isis) and sampling (fc_5sept2014.chn.isis) ISIS databases, which contain data compiled by Klondex between March 2004 and December 2014, comply with standards prescribed by CIM protocol for use in mineral reserve estimates.
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13. | Mineral Processing and Metallurgical Testing |
13.1. | Early Test Work |
A summary of the cyanidation test work conducted on twelve samples discussed in the 2011 NI 43-101 Technical Report by W. Raven, E. Ullmer, and G. Hawthorn is shown below in Table 13-1.
Table 13-1 Summary of Cyanidation Test Results from 2011 Technical Report
Sample | Drill | Head Grade | Test | Grind | Duration | Au | |||
ID | Zone | Hole | Interval | Au (g/t) | Au (opt) | Type | Size | (hrs) | Recovery |
1 | North Main | FC0401 | 2.0 | 0.058 | CIL | 75.9% | |||
2 | North Main | FC0403 | 14.5 | 0.423 | CIL | 80.0% | |||
3 | North Main | FC0405 | 34.6 | 1.009 | CIL | 60.1% | |||
5 | North Main | FC0402 | 905-910 | 37.1 | 1.082 | STD | 25%-200M | 33.2% | |
5 | North Main | FC0402 | 905-910 | 37.1 | 1.082 | STD | 90%-200M | 81.6% | |
C4 | North Main | FC0528 | 1450-1470 | 7.8 | 0.227 | STD | 80%-60M | 48 | 72.6% |
7 | Main | FC0413 | 850-855 | 109.0 | 3.178 | STD | 25%-200M | 74.4% | |
7 | Main | FC0413 | 850-855 | 109.0 | 3.178 | STD | 90%-200M | 98.7% | |
C1 | Main | FC0419 | 777-780 | 37.4 | 1.091 | STD | 80%-70M | 48 | 88.2% |
C3 | West Main | FC0515 | 925-935 | 116.4 | 3.394 | STD | 80%-65M | 48 | 86.8% |
4 | Far North-New North | FC0415 | 850-855 | 10.0 | 0.292 | STD | 25%-200M | 14.0% | |
4 | Far North-New North | FC0415 | 850-855 | 10.0 | 0.292 | STD | 90%-200M | 15.8% | |
6 | Far North-New North | FC0415 | 830-835 | 10.8 | 0.315 | STD | 25%-200M | 29.5% | |
6 | Far North-New North | FC0415 | 830-835 | 10.8 | 0.315 | STD | 90%-200M | 54.5% | |
C5 | Far North-New North | FC0418 | 895-915 | 6.1 | 0.178 | STD | 80%-65M | 48 | 45.4% |
C6 | Far North-New North | FC0522 | 1040-1050 | 20.1 | 0.586 | STD | 80%-80M | 48 | 77.2% |
13.2. | 2013 Test Work |
Metallurgical test work was conducted by McClelland Laboratories (MLI Job #3834) on two samples taken from the underground development to determine the amenability of the Project material to gravity and/or cyanidation treatment. Composite sample FCM1 was taken from material stockpiled during the development of the 5400 and 5370 crosscuts. Sample 3834-01 was generated by compositing coarse assay rejects from the face sampling on the Joyce 5400 N.
Each sample was milled to 80% minus 212 micrometers (µm) and processed through a laboratory Knelson concentrator to determine precious metal recovery via gravity concentration. The tailings from the Knelson concentrator were reground to 80% minus 75µm. Direct cyanidation tests (96-hour bottle roll tests) were then conducted on the gravity tailings to determine precious metal recovery and reagent consumption. Results of the test work are shown in the Table 13-2 below.
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Table 13-2 Combined Metallurgical Results, Gravity/Cyanidation Tests, 80% -212 υm Feed (Grav.), Reground to 80% -75 υm (CN)
g/tonne | Reaget Consumption | ||||||||||
Recovery % of Total | Extracted | Head Grade | kg / tonne | ||||||||
Grav. | CN | Combine | Grav. | CN | |||||||
Composite | Conc. | (Grav. | d | Conc. | Leach | Tail | Calculated | Assayed | NaCN | Lime | |
3771 Composite FCM1 | Au | 19.6% | 75.3% | 94.8% | 2.24 | 8.61 | 0.59 | 11.44 | 15.00 | 0.16 | 5.0 |
Sample 3834-91 | Au | 54.4% | 44.7% | 99.0% | 80.80 | 66.33 | 1.42 | 148.55 | 157.07 | 0.24 | 3.1 |
3771 Composite FCM1 | Ag | 14.4% | 67.8% | 82.2% | 1.30 | 6.10 | 1.60 | 9.00 | 6.00 | ||
Sample 3834-91 | Ag | 44.6% | 44.8% | 89.4% | 44.40 | 44.60 | 10.5 | 99.50 | 115.00 |
Results indicate that both samples were readily amenable to gravity and/or cyanidation treatment. Gold and silver recoveries achieved from composite sample FCM1 were 94.8% and 82.2%, respectively. Gold and silver recoveries achieved from sample 3834-01 were 99.0% and 89.4%, respectively. Cyanide consumptions were low, averaging 0.20 kg/million tons (Mt) material.
13.3. | 2014 Test Work |
In early 2014, nine drill core composite samples from the West Zone were submitted to McClelland Laboratories (MLI Job #3870) for metallurgical testing to determine the amenability of the Fire Creek West Zone material to direct cyanidation and gravity/cyanidation treatment.
Each composite was milled to 80% minus 75µm, and direct cyanidation tests (bottle roll tests) were then conducted to determine precious metal recovery and reagent consumption. Results from the test work are shown in the Table 13-3 below.
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Table 13-3 Summary Metallurgical Results, Bottle Roll Tests, Fire Creek West Zone Drill Core Composites
Au | g Au/mt ore | Ag | g Ag/mt ore | Reagent Requirements | |||||||||
Test | Recovery, | Calculated | Head | Recovery, | Calculated | Head | kg/mt mineralized material | ||||||
Number | Composite | % | Extracted | Tail | Head | Assay | % | Extracted | Tail | Head | Assay | NaCN Cons. | Lime Added |
CY-1 | 3870-1 | 96.0 | 34.88 | 1.46 | 36.34 | 46.10 | 94.1 | 17.4 | 1.1 | 18.5 | 30.3 | 0.17 | 0.8 |
CY-2 | 3870-2 | 94.9 | 20.23 | 1.08 | 21.31 | 26.18 | 74.9 | 12.8 | 4.3 | 17.1 | 26.2 | 0.39 | 5.6 |
CY-3 | 3870-3 | 89.8 | 6.66 | 0.76 | 7.42 | 10.28 | 67.4 | 6.4 | 3.1 | 9.5 | 15.3 | 0.33 | 6.9 |
CY-4 | 3870-4 | 96.9 | 14.51 | 0.46 | 14.97 | 12.51 | 76.9 | 1.0 | 0.3 | 1.3 | 1.9 | 0.17 | 7.6 |
CY-5 | 3870-5 | 93.2 | 38.28 | 2.80 | 41.08 | 30.30 | 56.3 | 57.4 | 44.6 | 102.0 | 92.5 | 0.28 | 3.7 |
CY-6 | 3870-61) | 66.9 | 3.92 | 1.94 | 5.86 | 7.67 | 81.2 | 22.9 | 5.3 | 28.2 | 36.8 | 12.16 | 20.5 |
CY-7 | 3870-7 | 84.0 | 22.32 | 4.24 | 26.56 | 30.33 | 57.8 | 17.0 | 12.4 | 29.4 | 35.7 | 0.38 | 3.6 |
CY-8 | 3870-8 | 82.1 | 60.94 | 13.30 | 74.24 | 63.33 | 71.7 | 34.0 | 13.4 | 47.4 | 36.9 | 0.31 | 2.4 |
CY-9 | 3870-9 | 98.7 | 48.41 | 0.62 | 49.03 | 73.87 | 83.5 | 27.8 | 5.5 | 33.3 | 50.3 | 0.34 | 4.2 |
Notes: | ||
1. | Problems encountered with high viscosity, low D.O. and low free cyanide levels. Switched to mechanically agitated leach @ 2.0 g NaCN/L, 25% Solids at 20 hours, initiated are sparge at 24 hours. |
Results indicate that all but one (Composite #3870-6) of the samples were readily amenable to direct cyanidation treatment. Gold recoveries achieved from the eight composite samples ranged from 82.1% to 98.7% . Silver recoveries achieved from the eight composite samples ranged from 56.3% to 94.1% . Cyanide consumptions were low, averaging 0.30 kg/Mt material.
Problems were encountered during direct cyanidation testing of composite #3870-6 due to high viscosity, low dissolved oxygen content and low free cyanide levels. This composite was transferred to a mechanically agitated leach apparatus to complete the test. Gold and silver recoveries achieved from composite #3870-6 were 66.9% and 81.2%, respectively. Cyanide and lime requirements for this sample were very high.
After direct cyanidation testing was complete, two master composites were prepared for gravity/cyanidation testing. A high-grade master composite (HG master comp) was prepared by combining the coarse rejects from Composites 3870-5 and 3879-6. A mid-grade master composite (MG master comp) was prepared by combining coarse rejects from Composites 3870-2, 3870-3 and 3870-4.
Each master composite was milled to 80% minus 300µm and processed through a laboratory Knelson concentrator to determine precious metal recovery via gravity concentration. The tailings from the Knelson concentrator were reground to 80% minus 75µm. Direct cyanidation tests (96-hour bottle roll tests), with and without lead nitrate addition, were then conducted on the gravity tailings to determine precious metal recovery and reagent consumption. Results of the test work are shown in Table 13-4 and Table 13-5 below.
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Table 13-4 Gold Metallurgical Results, Whole Mineralized Material Gravity Concentration with Cyanidation of the Gravity Cleaner and Rougher Tailings
Weight , % of Total | g Au/mt mineralized material | |||||||||
Lead | Combined | |||||||||
Nitrate | Gravity | Cl. & Ro. | Ball Mill | Gravity | Extracted | Calc. | Predicted | |||
Composite | Added | Cl. Conc | Tail | Total | Clean Out | Cl. Conc | (CN) | Tail | Head | Head |
3870-29 (HG Master Comp.) | No | 0.21 | 99.79 | 100.0 | 0.14 | 10.416 | 19.79 | 2.70 | 33.05 | 30.32. |
Yes | 0.21 | 99.79 | 100.0 | 0.14 | 10.416 | 17.60 | 2.54 | 30.70 | ||
3879-30 (MG Master Comp.) | No | 0.26 | 99.74 | 100.0 | 0.02 | 4.68 | 10.45 | 0.69 | 15.84 | 12.54 |
Yes | 0.26 | 99.74 | 100.0 | 0.02 | 4.68 | 8.50 | 0.73 | 13.93 | ||
Au Distribution % of Total | kg/mt ore | |||||||||
Ball Mill | Cl. | Extracted | NaCN | Lime | ||||||
Composite | Clean Out | Conc | (CN) | Combined | Tail | Total | Cons. | Added | ||
3870-29 (HG Master Comp.) | 0.4 | 31.5 | 59.9 | 91.4 | 8.2 | 100.0 | 0.31 | 3.5 | ||
0.5 | 33.9 | 57.3 | 91.2 | 8.3 | 100.0 | 0.31 | 3.5 | |||
3879-30 (MG Master Comp.) | 0.1 | 29.5 | 66.0 | 95.5 | 4.4 | 100.0 | 0.09 | 6.5 | ||
0.1 | 33.6 | 61.0 | 94.6 | 5.3 | 100.0 | 0.15 | 6.7 |
Table 13-5 Silver Metallurgical Results, Whole Mineralized Material Gravity Concentration with Cyanidation of the Gravity Cleaner and Rougher Tailings
Weight , % of Total | g Ag/mt mineralized material | |||||||||
Lead | Combined | |||||||||
Nitrate | Gravity | Cl. & Ro. | Ball Mill | Gravity Cl. | Extracted | Calc. | Predicted | |||
Composite | Added | Cl. Conc | Tail | Total | Clean Out | Conc | (CN) | Tail | Head | Head |
3870-29 (HG Master Comp.) | No | 0.21 | 99.79 | 100.0 | 0.12 | 7.056 | 31.43 | 25.45 | 64.06 | |
Yes | 0.21 | 99.79 | 100.0 | 0.12 | 7.056 | 48.00 | 11.28 | 66.45 | ||
3879-30 (MG Master Comp.) | No | 0.26 | 99.74 | 100.0 | 0.06 | 2.184 | 7.48 | 3.29 | 13.02 | |
Yes | 0.26 | 99.74 | 100.0 | 0.06 | 2.184 | 6.48 | 3.39 | 12.12 | ||
Au Distribution % of Total | ||||||||||
Ball Mill | Cl. | Extracted | ||||||||
Composite | Clean Out | Conc | (CN) | Combined | Tail | Total | ||||
3870-29 (HG Master Comp.) | 0.2 | 11.0 | 49.1 | 60.1 | 39.7 | 100.0 | ||||
0.2 | 10.6 | 72.2 | 82.8 | 17.0 | 100.0 | |||||
3879-30 (MG Master Comp.) | 0.5 | 16.8 | 57.5 | 74.3 | 25.3 | 100.0 | ||||
0.5 | 18.0 | 53.5 | 71.5 | 28.0 | 100.0 |
Results indicate that both master composites were readily amenable to gravity/cyanidation treatment. Gold and silver recoveries achieved from the HG master composite were 91.4% and 60.0%, respectively, without lead nitrate, and 91.2% and 82.8% with lead nitrate addition. Gold and silver recoveries achieved from the MG master composite were 95.5% and 74.3%, respectively, without lead nitrate, and 94.6% and 71.3% with lead nitrate addition.
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14. | Mineral Resource Estimates |
14.1. | Introduction |
The Fire Creek mineral resource was estimated in accordance with The Canadian Institute of Mining, Metallurgy and Petroleum’s CIM Definitions Standards for Mineral Resources and Mineral Reserves, adopted by CIM Council on May 10, 2014 (CIM 2014). This estimate updates the previous Mineral Resource Estimate effective January 31, 2014 (Odell et al, 2014) and includes all of the new drilling, channel sampling, and underground geological mapping completed since that date.
All data coordinates are measured in the Nevada State Plane Central Zone, NAD83 feet truncated to the last six whole digits. All quantities are given in imperial units unless indicated otherwise.
The gold and silver mineralization at the Project was estimated using the Vulcan modeling software. The estimate was performed by Karl Swanson, an Independent Consultant, with assistance from Klondex geology staff. The vein solids were modeled from both the assay data and the lithology logging from drilling and channel samples. No strict grade cutoff was honored, but care was taken to ensure that only vein material was modeled regardless of the grade.
Vulcan software was used in most aspects of the modeling process, though statistics and variography calculated with third-party software. ID3 and Nearest Neighbor (NN), also known as polygonal, estimation methods were used. The lack of sufficient closely spaced drilling or channel sample composites precludes the use Ordinary Kriging (OK) estimation methods and the calculation of meaningful variograms.
14.2. | Database and Compositing |
Klondex provided drilling data in CSV format to Practical Mining. The gold and silver assays were given in g/t, which were converted to opt by multiplying by 34.2854.
Assays were assigned to a specific vein or “flagged” by their spatial relationship to the digital vein models. Samples were then composited into a single weighted average value spanning the width of the vein or ten feet, whichever is less.
14.2.1. | Assays |
This analysis used 475 surface and underground drill holes and 1,457 channel and rib sample sets. The composites of all flagged assays were used for statistical analysis and estimation. No drill holes or channels were eliminated for any reason. Table 14-1 summarizes the overall quantity of data available by type and the quantity flagged for use in the estimation.
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Table 14-1 Summary of Drill Hole and Channel Samples
Not Flagged by Veins | Flagged by Vein Models | Total | ||||||||
No. | Length | Length | No. | Length | Length | Length | No. | Length | Length | |
Type | Holes | Drilled | Sampled | Holes | Drilled | Sampled | Flagged | Holes | Drilled | Sampled |
Drill | 214 | 203,391 | 199,497 | 475 | 325,917 | 323,718 | 5,603 | 689 | 529,308 | 523,215 |
Channel | 96 | 494 | 411 | 1,457 | 10,785 | 10,653 | 3,646 | 1,553 | 11,279 | 11,064 |
Drill hole and channel sample locations relative to the vein models are shown in Figure 14-1 and Figure 14-2 below.
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14.2.2. | Lithology |
The rock types identified in the lithology logging are shown in Table 14-2. Intervals logged as vein or structure along with assay values were used to identify veins.
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Table 14-2 Lithology Codes
Lithology Code | Description |
OVB | overburden |
SEDS | sedimentary |
OPAL | opalized sinter |
INT | intrusive |
STR | structure |
FLT | fault |
VN | vein |
BAS | basalt |
BX | breccia |
TUFF | tuff |
ND | no data |
The core logging shows that there is an upper and lower tuff unit within the basalt. Figure 14-3 is a long section through the deposit showing the tuff in blue and the basalt in light green. Figure 14-4 and Figure 14-5 show the same section with the tuff and Vein Models.
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14.2.3 | Compositing |
The assays were composited on ten-foot downhole interval lengths honoring the vein intersections. Therefore, the assays within the veins are separated from the lower grade values outside of the veins. This compositing method usually calculates a single composite across the vein interval as most vein intercepts are less than ten feet in length. Where the interval within the vein was longer than ten feet, more than one composite was created.
Vein intercepts from 1,932 channels and drill holes were flagged. These vein intercepts consisted of 4,070 samples with gold assays greater than zero for a total intercept length of 8,800 feet. These intercepts comprise 3,341 composites with gold grades greater than zero for a total of 8,819 feet. All of these flagged composites were used for statistics and estimation.
14.3. | Geology and Vein Modelling |
The basalt and tuff units were modeled but are not used in the block model. These rock units do not seem to impact vein location or mineralization as the veins cross through basalt and tuff.
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Displacement of the rock units indicates faulting, both pre and post mineralization, which helps in understanding local geology and guide exploration.
Forty seven veins were modeled on two main northwest linear trends separated approximately 1,300 feet east to west. Both trends generally strike N15°W, but the east trend ends at about 767,400 N where the west trend begins. This may be the result of a northeast trending fault but is not clear at this time. Mining and channel sampling has occurred on the Joyce Vein, Vonnie Vein, and Karen Vein at the center of the east trend.
A low grade halo is sometimes present immediately adjacent to the veins. This material is usually a stock work of quartz veining or porous basalt or intrusive rock. The low grade mineralization adjacent to the veins was modeled using a cutoff grade of 0.1 opt and extends from the cutoff value to the vein. The thickness of this low grade material around vein varies and can occur either in the footwall, hanging wall, or both.
Vein models were constructed on N75°E cross sections spaced 25 feet apart using the gold assay values from drill and channel intercepts and lithology logging. A strict cutoff grade was not enforced as portions of the vein are very low grade, but regardless of the grade, an assay was chosen to represent the vein if the drill hole intercepted the modeled vein. Figure 14-6 shows a N75°E cross section of the veins with drilling color-coded by gold grade.
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Where channel samples are present, the vein model produced from the channel samples replaces the vein model produced from drill composites, and the channel samples take precedence over drilling.
14.4. | Density |
A density value of 0.0774 tons per cubic foot was assigned to all vein and low grade mineralization. This value is supported by 15 samples collected on the Joyce Vein and Vonnie Vein and analyzed by SGS Laboratories in Elko, Nevada.
14.5. | Statistics |
Drill hole and channel samples were grouped according to vein or low grade designations and univariate statistics calculated for each sample type and group. The summary statistics are shown in Table 14-3 through Table 14-6. Histogram and cumulative frequency plots are shown in Figure 14-7 through Figure 14-10.
Table 14-3 Vein Drill Hole Statistics
Field | # Comps | Min | Max | Mean | Std Dev | CV |
au opt | 1008 | 0.0001 | 108.221 | 1.3237 | 5.3416 | 4.0354 |
ag opt | 1005 | 0.0073 | 84.2926 | 1.1 | 4.706 | 4.2782 |
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Table 14-4 Vein Channel Sample Statistics
Field | # Comps | Min | Max | Mean | Std Dev | CV |
au opt | 1008 | 0.0001 | 108.221 | 1.3237 | 5.3416 | 4.0354 |
ag opt | 1005 | 0.0073 | 84.2926 | 1.1 | 4.706 | 4.2782 |
Table 14-5 Low Grade Drill Hole Statistics
Field | # Comps | Min | Max | Mean | Std Dev | CV |
au opt | 212 | 0.0515 | 4.15 | 0.243 | 0.3768 | 1.5505 |
ag opt | 212 | 0.0073 | 6.2132 | 0.3737 | 0.599 | 1.6027 |
Table 14-6 Low Grade Channel Sample Statistics
Field | # Comps | Min | Max | Mean | Std Dev | CV |
au opt | 533 | 0.019 | 4.728 | 0.5758 | 0.7237 | 1.2568 |
ag opt | 447 | 0.05 | 48 | 0.8254 | 3.6097 | 4.3735 |
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14.6. | Grade Capping |
A cap grade was applied to the gold and silver composites used to estimate block grades. For the first (Measured) estimation pass, values greater than the cap are used in the estimation but are restricted by an area of influence based on a search radius of 25 feet by 25 feet. Beyond 25 feet, capped grade value is excluded from the estimation.
The cap grade for measured mineral resources was determined for each of the three main veins (Vonnie, Joyce, Karen) individually and for the remaining veins as a group.
For the indicated and inferred mineral resource estimation passes, only drill hole composites and no channel sample composites, were used in the estimations. The indicated and inferred mineral resources grade cap was not estimated individually for the Vonnie Vein, Karen Vein, and Joyce Vein as there are not enough drill composites to make the determination individually.
For the indicated and inferred mineral resource estimation passes, the grade cap value is used for the estimation of all vein blocks within the normal sphere of influence as shown in Table 14-11.
Table 14-7 shows the cap grade for the measured mineral resource estimation for gold and silver for both the high and low grade composites and the number of composites that are above the cap value.
Table 14-8 shows the cap grade for the indicated and inferred mineral resource estimations for gold and silver composites for all veins and the number of composites that are above the cap value.
The distribution of gold composite grades for the Vonnie Vein is shown in Figure 14-11 and for silver in Figure 14-12.
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Table 14-7 Cap Grades for Measured Mineral Resources
Cap Grade | # Comps | |||
Vein | Group | Variable | (opt) | Affected |
Vonnie | vein | au opt | 70.00 | 41 |
Vonnie | low | au opt | 1.90 | 18 |
Vonnie | vein | ag opt | 70.00 | 24 |
Vonnie | low | ag opt | 1.40 | 15 |
Joyce | vein | au opt | 35.00 | 12 |
Joyce | low | au opt | 1.00 | 19 |
Joyce | vein | ag opt | 20.00 | 11 |
Joyce | low | ag opt | 1.20 | 10 |
Karen | vein | au opt | 31.00 | 7 |
Karen | low | au opt | 0.90 | 13 |
Karen | vein | ag opt | 30.00 | 6 |
Karen | low | ag opt | 0.90 | 11 |
all other | vein | au opt | 14.00 | 6 |
all other | low | au opt | 0.37 | 20 |
all other | vein | ag opt | 14.00 | 8 |
all other | low | ag opt | 0.70 | 22 |
Table 14-8 Cap Grades for Indicated and Inferred Mineral Resources
Cap Grade | # Comps | |||
Vein | Group | Variable | (opt) | Affected |
all | vein | au opt | 11.00 | 19 |
all | low | au opt | 0.37 | 20 |
all | vein | ag opt | 11.00 | 11 |
all | low | ag opt | 0.70 | 22 |
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14.7 | Variography |
Variograms were calculated with SAGE2001 software using the gold channel composites within the Vonnie Vein. The closely spaced channel samples within the Vonnie Vein allow construction of a meaningful variogram. The variogram model use a custom LLL-ZYX output from SAGE. The nugget and structure parameters for the Vonnie Vein at the Project are in Table 14-9. The representative variogram plots are in Figures 14-13 and 14-14. Too few closely spaced samples exist within the other veins to calculate reliable kriging parameters, and therefore an ordinary kriging estimation was not performed at this time.
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Table 14-9 Variogram Parameters for the Vonnie Vein
Nugget ==> 0.312 | ||||
C1 ==> 0.317 | ||||
C2 ==> 0.371 | ||||
First Structure -- Exponential with Practical Range | ||||
1st rotation about the Z axis ==> 9 | ||||
2nd rotation about the Y' axis ==> -15 | ||||
3rd rotation about the X' axis ==> -47 | ||||
Range along the Z' axis 102.0 | Azimuth 176 | Dip 41 | ||
Range along the Y' axis 24.1 | Azimuth 25 | Dip 45 | ||
Range along the X' axis 31.5 | Azimuth 99 | Dip -15 | ||
Second Structure -- Exponential with Practical Range | ||||
1st rotation about the Z axis ==> 18 | ||||
2nd rotation about the Y' axis ==> 19 | ||||
3rd rotation about the X' axis ==> -21 | ||||
Range along the Z' axis 134.1 | Azimuth 238 | Dip 62 | ||
Range along the Y' axis 30.9 | Azimuth 11 | Dip 20 | ||
Range along the X' axis 78.7 | Azimuth 108 | Dip 19 |
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14.8. | Block Model |
The block model was constructed using a 3,500-foot by five-foot by five-foot parent block size (XYZ), with sub-blocking in the veins and low grade mineralization as small as 0.2 feet by five feet by five feet. This modeling method creates a single block across the vein in the X direction with a tolerance of 0.2 feet. Therefore, the block width across the vein is within 0.2 feet of the actual width of the vein solid. Because low grade material was modeled around the high grade center vein, one to three blocks may exist across the vein in the X direction depending on whether low grade exists on one or both sides of the vein. If no low grade exists, then only one block defines the vein.
The model is rotated with the X-axis oriented N75W. The block model origin (lower left corner) is 643000E, 763200N, 4600EL. The X length is 3,500 feet, Y length is 7,500 feet and the Z length is 1,900 feet.
The unique vein name is assigned to a block variable. The vein name is superseded by “l” to indicate a low grade block.
14.9. | Grade Estimation |
Gold and silver values were estimated using ID3 and NN methods. The ID3 method was applied in multiple passes defining the extents of the measured, indicated and inferred classifications.
The channel composites were only used for the measured pass, which has a search ellipsoid of 40 feet by 40 feet by 20 feet. This was done to ensure that high grade was not extended further than is supported by the data. This along with the capping strategy limits the range of influence of the high grade channel sample composites.
Anisotropic search parameters for gold were set to the average orientation of the veins. Search distances were selected based on the spacing of drill composites intercepting the digital vein models and on the general orientation and shape of the interpreted veins. The vein’s gold and silver grades were estimated only using composites from within the vein, and low grade blocks were estimated using only low grade composites. The boundary separating the veins and low grade blocks is regarded as a hard boundary with the data in each isolated from the other.
The estimation search parameters are shown in Table 14-10. The search ellipse orientations for each vein is shown in Table 14-11.
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Table 14-10 Estimation Search Parameters by Resource Category
Parent | Major | Semi | Minor | Min | Max | Max | |||
Pass | X | Y | Z | (ft) | (ft) | (ft) | Samp | Samp | /DH |
Measured | 10 | 10 | 10 | 40 | 40 | 20 | 4 | 12 | 2 |
Indicated | 10 | 10 | 10 | 100 | 100 | 50 | 3 | 12 | 2 |
Inferred | 10 | 10 | 10 | 300 | 300 | 150 | 2 | 12 | 2 |
Table 14-11 Estimation Search Ellipsoids
Vein | Est ID | Bearing | Plunge | Dip |
Vonnie | vn_v | 173 | 0 | -80 |
Vonnie 1 | vn_v1 | 167 | 0 | -90 |
Joyce | vn_j | 335 | 0 | -86 |
Joyce 1 | vn_j1 | 338 | 0 | -90 |
Karen | vn_k | 355 | 0 | -80 |
Karen 1 | vn_k1 | 355 | 0 | -75 |
Karen 2 | vn_k2 | 355 | 0 | -75 |
Karen 3 | vn_k3 | 340 | 0 | -90 |
03 | vn03 | 171 | 0 | -80 |
04 | vn04 | 168 | 0 | -80 |
05 | vn05 | 168 | 0 | -65 |
06 | vn06 | 160 | 0 | -84 |
07 | vn07 | 183 | 0 | -70 |
08 | vn08 | 163 | 0 | -80 |
09 | vn09 | 330 | 0 | -84 |
10 | vn10 | 167 | 0 | -50 |
11 | vn11 | 167 | 0 | -80 |
12 | vn12 | 165 | 0 | -76 |
13 | vn13 | 340 | 0 | -90 |
14 | vn14 | 331 | 0 | -90 |
15 | vn15 | 345 | 0 | -90 |
16 | vn16 | 137 | 0 | -87 |
18 | vn18 | 292 | 0 | -87 |
19 | vn19 | 340 | 0 | -90 |
Honey Runner - 20 | vn20 | 339 | 0 | -85 |
21 | vn21 | 345 | 0 | -85 |
22 | vn22 | 340 | 0 | -66 |
23 | vn23 | 170 | 0 | -80 |
24 | vn24 | 168 | 0 | -80 |
25 | vn25 | 158 | 0 | -77 |
26 | vn26 | 163 | 0 | -75 |
27 | vn27 | 163 | 0 | -81 |
28 | vn28 | 335 | 0 | -74 |
29 | vn29 | 345 | 0 | -70 |
30 | vn30 | 340 | 0 | -72 |
31 | vn31 | 359 | 0 | -90 |
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Vein | Est ID | Bearing | Plunge | Dip |
32 | vn32 | 6 | 0 | -90 |
33 | vn33 | 165 | 0 | -67 |
34 | vn34 | 115 | 0 | -87 |
35 | vn35 | 163 | 0 | -70 |
Hui Wu - 36 | vn36 | 349 | 0 | -73 |
37 | vn37 | 308 | 0 | -88 |
39 | vn39 | 356 | 0 | -80 |
40 | vn40 | 349 | 0 | -77 |
41 | vn41 | 345 | 0 | -85 |
42 | vn42 | 359 | 0 | -84 |
43 | vn43 | 353 | 0 | -86 |
44 | vn44 | 179 | 0 | -80 |
45 | vn45 | 345 | 0 | -75 |
46 | vn46 | 343 | 0 | -90 |
47 | vn47 | 348 | 0 | -90 |
99 | vn99 | 345 | 0 | -90 |
Project
Significant parameters used in the gold and silver estimations include:
1. | Assigning of parent block values to sub- blocks. Estimates are only calculated at the center of each ten-foot by ten-foot by ten-foot block, and those values are assigned to all sub- blocks existing within the parent block space; | |
2. | Only composites with a value greater than zero were used; | |
3. | A minimum of four and maximum of 12 samples were used to estimate measured blocks, a minimum of three and maximum of 12 to estimate indicated blocks, and minimum of two and maximum of 12 to estimate inferred blocks; | |
4. | A maximum of two composites were used per drill hole; | |
5. | Composites were selected using anisotropic distances; | |
6. | Only composites within the veins were used to estimate blocks within the veins; | |
7. | Grades were capped (search restricted) for measured material; | |
8. | Grades were capped with a top cut for indicated and inferred material; and | |
9. | Gold and silver for blocks outside of the low and high grade vein solids were not estimated. |
14.10. | Mined Depletion and Sterilization |
The block model is depleted by the as-built survey of the underground workings. Blocks within the survey were flagged as “mined”. The grades and the density within the flagged blocks remain intact in order reconcile with mining. Remnant blocks within the hanging wall or footwall of the veins which are not inside the mine survey but immediately adjacent to it were also flagged as “mined”.
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The Joyce Vein, Vonnie Vein, and Karen Vein are the only three that have been mined as of the effective date of this report. The vein models and as-built surveys for each are shown Figure 14-15 through Figure 14-17.
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14.11. | Model Validation |
The mean gold and silver grades for each estimation method are shown in Table 14-12.The values correlate well with Method ID3 returning the lowest value as would be expected.
Table 14 -12 Estimation the Mean Grade Comparison
Cut Off | Average | |
Estimation Method | Grade (opt) | Grade (opt) |
Gold | ||
Inverse Distance Cubed | 0.01 | 0.764 |
Nearest Neighbor | 0.01 | 0.813 |
Composite | 0 | 1.037 |
Silver | ||
Inverse Distance Cubed | 0.01 | 0.639 |
Nearest Neighbor | 0.01 | 0.764 |
Composite | 0 | 0.806 |
On a local scale, model validation is confirmed by the comparison of block grades to composite grades for each vein in the long sections shown in Figure 14-19 through Figure 14-24. The color legend of Figure 14-18 is applied to all block and composite grade values for comparative purposes. Examination indicates good agreement of block grade estimations and distribution with the composite date.
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Further model validation is provided by the swath plots of Figure 14-25 through Figure 14-28. These plots compare the average grade from composites, NN and ID3 estimations from within swaths or 25-foot thick two dimensional slices through the vein. Examination of the swath plots shows a reasonable agreement among the gold and silver estimation values.
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14.12 | Mineral Resource Statement |
The narrow vein mining methods practiced at the Project require a minimum stope width of four feet. The veins can vary in thickness from a few inches to over ten feet. Potentially economic mineralization must meet standard cut-off grade criteria as well as a grade thickness criterion before it is included in a Mineral Resource estimate. Grade thickness is calculated by multiplying the block true width by its equivalent grade. The parameters used in determining the cut-off grade and grade thickness cut-off are listed in Table 14-13.
Table 14-13 Mineral Resource Cutoff Grade Parameters
Gold | Silver | ||
Sales Price | $/Ounce | $1,200 | $19.00 |
Refining and Sales Expense | $/Ounce | Included in Milling | |
Royalty | 1% | ||
Metallurgical Recovery | 94% | 92% | |
Operating Costs | |||
Ore Haulage (Portal to Mill) | $/ton | $ 32.66 | |
Direct Processing | $/ton | $ 47.10 | |
Administration and Overhead | $/ton | $ 47.32 | |
Mining | $/ton | $ 158.75 | |
Total | $/ton | $ 285.82 | |
Gold Equivalent | 1 | 64.54 | |
Unplanned Dilution | 10% | ||
Cut-off Grade | Eq. opt | 0.256 | |
Minimum Mining Width | feet | 4 | |
Grade Thickness cut-off | Eq. opt-ft. | 1.126 |
Mineral resources meeting the dual constraints of cut-off grade and grade-thickness cut-off for each vein are listed in Table 14-14 below.
Table 14-14 Mineral Resource Statement as of December 31, 2015
AuEq | AuEq | ||||||
Vein Name | kton | Au opt | Ag opt | opt | Au koz | Ag koz | koz |
Measured | |||||||
Joyce | 30.1 | 1.863 | 1.217 | 1.882 | 56.0 | 36.6 | 56.6 |
Vonnie | 9.8 | 4.584 | 3.556 | 4.640 | 44.9 | 34.9 | 45.5 |
Karen | 43.6 | 2.272 | 1.888 | 2.301 | 99.0 | 82.2 | 100.2 |
Honey Runner | 0.5 | 1.786 | 0.166 | 1.788 | 0.9 | 0.1 | 0.9 |
Hui Wu | 2.2 | 0.530 | 0.866 | 0.543 | 1.2 | 1.9 | 1.2 |
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AuEq | AuEq | ||||||
Vein Name | kton | Au opt | Ag opt | opt | Au koz | Ag koz | koz |
08 | 1.0 | 0.466 | 0.460 | 0.473 | 0.5 | 0.5 | 0.5 |
11 | 0.1 | 0.365 | 1.298 | 0.385 | 0.0 | 0.1 | 0.0 |
13 | 1.8 | 0.668 | 0.096 | 0.669 | 1.2 | 0.2 | 1.2 |
14 | 3.3 | 1.988 | 1.075 | 2.004 | 6.5 | 3.5 | 6.6 |
19 | 0.1 | 0.247 | 0.005 | 0.247 | 0.0 | 0.0 | 0.0 |
40 | 0.6 | 3.030 | 1.322 | 3.051 | 1.7 | 0.7 | 1.7 |
Total Measured | 93.0 | 2.279 | 1.728 | 2.306 | 212.0 | 160.8 | 214.5 |
Indicated | |||||||
Joyce | 78.9 | 0.900 | 0.677 | 0.911 | 71.0 | 53.5 | 71.9 |
Vonnie | 80.2 | 0.795 | 0.737 | 0.807 | 63.8 | 59.1 | 64.7 |
Karen | 13.9 | 1.109 | 0.877 | 1.123 | 15.4 | 12.2 | 15.6 |
Honey Runner | 33.4 | 0.589 | 0.223 | 0.592 | 19.7 | 7.4 | 19.8 |
Hui Wu | 7.0 | 0.441 | 0.260 | 0.445 | 3.1 | 1.8 | 3.1 |
03 | 3.1 | 0.335 | 0.524 | 0.343 | 1.0 | 1.6 | 1.1 |
04 | 2.5 | 0.208 | 0.373 | 0.213 | 0.5 | 0.9 | 0.5 |
05 | 1.9 | 0.318 | 1.256 | 0.337 | 0.6 | 2.3 | 0.6 |
06 | 6.0 | 0.441 | 0.899 | 0.454 | 2.6 | 5.4 | 2.7 |
08 | 4.4 | 0.360 | 0.243 | 0.363 | 1.6 | 1.1 | 1.6 |
11 | 7.2 | 0.358 | 0.875 | 0.372 | 2.6 | 6.3 | 2.7 |
13 | 5.6 | 0.564 | 0.138 | 0.566 | 3.2 | 0.8 | 3.2 |
14 | 12.4 | 0.622 | 0.317 | 0.627 | 7.7 | 3.9 | 7.8 |
19 | 2.6 | 0.342 | 0.130 | 0.344 | 0.9 | 0.3 | 0.9 |
21 | 0.7 | 0.327 | 0.571 | 0.336 | 0.2 | 0.4 | 0.2 |
27 | 8.8 | 0.315 | 0.261 | 0.319 | 2.8 | 2.3 | 2.8 |
29 | 0.6 | 0.146 | 0.374 | 0.152 | 0.1 | 0.2 | 0.1 |
30 | 9.5 | 0.330 | 0.224 | 0.333 | 3.1 | 2.1 | 3.2 |
34 | 0.2 | 0.232 | 0.371 | 0.238 | 0.0 | 0.1 | 0.0 |
40 | 3.0 | 0.468 | 0.379 | 0.474 | 1.4 | 1.1 | 1.4 |
43 | 1.0 | 1.327 | 0.023 | 1.328 | 1.3 | 0.0 | 1.3 |
44 | 1.7 | 0.529 | 0.267 | 0.533 | 0.9 | 0.5 | 0.9 |
Total Indicated | 284.4 | 0.716 | 0.575 | 0.725 | 203.5 | 163.4 | 206.1 |
Measured and Indicated | |||||||
Joyce | 109.0 | 1.166 | 0.826 | 1.179 | 127.1 | 90.1 | 128.5 |
Vonnie | 90.0 | 1.208 | 1.044 | 1.224 | 108.7 | 94.0 | 110.2 |
Karen | 57.5 | 1.990 | 1.643 | 2.015 | 114.4 | 94.5 | 115.9 |
Honey Runner | 33.9 | 0.607 | 0.222 | 0.610 | 20.6 | 7.5 | 20.7 |
Hui Wu | 9.2 | 0.463 | 0.406 | 0.469 | 4.3 | 3.8 | 4.3 |
03 | 3.1 | 0.335 | 0.524 | 0.343 | 1.0 | 1.6 | 1.1 |
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AuEq | AuEq | ||||||
Vein Name | kton | Au opt | Ag opt | opt | Au koz | Ag koz | koz |
04 | 2.5 | 0.208 | 0.373 | 0.213 | 0.5 | 0.9 | 0.5 |
05 | 1.9 | 0.318 | 1.256 | 0.337 | 0.6 | 2.3 | 0.6 |
06 | 6.0 | 0.441 | 0.899 | 0.454 | 2.6 | 5.4 | 2.7 |
08 | 5.4 | 0.380 | 0.284 | 0.384 | 2.0 | 1.5 | 2.1 |
11 | 7.3 | 0.359 | 0.880 | 0.372 | 2.6 | 6.4 | 2.7 |
13 | 7.4 | 0.589 | 0.128 | 0.591 | 4.4 | 0.9 | 4.4 |
14 | 15.7 | 0.909 | 0.476 | 0.916 | 14.2 | 7.4 | 14.3 |
19 | 2.7 | 0.339 | 0.126 | 0.341 | 0.9 | 0.3 | 0.9 |
21 | 0.7 | 0.327 | 0.571 | 0.336 | 0.2 | 0.4 | 0.2 |
27 | 8.8 | 0.315 | 0.261 | 0.319 | 2.8 | 2.3 | 2.8 |
29 | 0.6 | 0.146 | 0.374 | 0.152 | 0.1 | 0.2 | 0.1 |
30 | 9.5 | 0.330 | 0.224 | 0.333 | 3.1 | 2.1 | 3.2 |
34 | 0.2 | 0.232 | 0.371 | 0.238 | 0.0 | 0.1 | 0.0 |
40 | 3.5 | 0.879 | 0.530 | 0.887 | 3.1 | 1.9 | 3.1 |
43 | 1.0 | 1.327 | 0.023 | 1.328 | 1.3 | 0.0 | 1.3 |
44 | 1.7 | 0.529 | 0.267 | 0.533 | 0.9 | 0.5 | 0.9 |
Total Meas. and Ind. | 377.4 | 1.101 | 0.859 | 1.114 | 415.5 | 324.2 | 420.5 |
Inferred | |||||||
Joyce | 65.4 | 0.424 | 0.375 | 0.430 | 27.7 | 24.5 | 28.1 |
Vonnie | 29.0 | 0.648 | 0.414 | 0.654 | 18.8 | 12.0 | 19.0 |
Karen | 16.5 | 0.609 | 0.425 | 0.616 | 10.0 | 7.0 | 10.1 |
Honey Runner | 63.0 | 0.860 | 0.607 | 0.870 | 54.1 | 38.2 | 54.7 |
Hui Wu | 0.4 | 0.288 | 0.145 | 0.290 | 0.1 | 0.1 | 0.1 |
04 | 39.3 | 0.283 | 0.338 | 0.288 | 11.1 | 13.3 | 11.3 |
05 | 1.1 | 0.336 | 1.081 | 0.352 | 0.4 | 1.2 | 0.4 |
06 | 1.9 | 0.272 | 0.594 | 0.282 | 0.5 | 1.1 | 0.5 |
08 | 1.4 | 0.294 | 0.179 | 0.296 | 0.4 | 0.2 | 0.4 |
09 | 41.4 | 0.490 | 0.170 | 0.493 | 20.3 | 7.0 | 20.4 |
11 | 11.7 | 0.327 | 0.337 | 0.332 | 3.8 | 3.9 | 3.9 |
14 | 62.4 | 0.468 | 0.347 | 0.473 | 29.2 | 21.7 | 29.5 |
15 | 24.7 | 0.423 | 2.172 | 0.456 | 10.4 | 53.6 | 11.3 |
19 | 6.0 | 0.204 | 0.141 | 0.207 | 1.2 | 0.8 | 1.2 |
21 | 15.1 | 0.498 | 0.085 | 0.499 | 7.5 | 1.3 | 7.5 |
22 | 42.0 | 0.443 | 0.383 | 0.448 | 18.6 | 16.1 | 18.8 |
23 | 128.5 | 0.162 | 0.118 | 0.164 | 20.9 | 15.1 | 21.1 |
24 | 90.9 | 0.520 | 0.664 | 0.530 | 47.3 | 60.3 | 48.2 |
25 | 37.1 | 0.599 | 0.106 | 0.601 | 22.2 | 3.9 | 22.3 |
26 | 17.7 | 0.235 | 0.084 | 0.236 | 4.1 | 1.5 | 4.2 |
27 | 27.1 | 0.225 | 0.133 | 0.227 | 6.1 | 3.6 | 6.2 |
28 | 35.4 | 0.490 | 0.308 | 0.495 | 17.3 | 10.9 | 17.5 |
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AuEq | AuEq | ||||||
Vein Name | kton | Au opt | Ag opt | opt | Au koz | Ag koz | koz |
29 | 9.9 | 0.165 | 0.291 | 0.170 | 1.6 | 2.9 | 1.7 |
30 | 23.4 | 0.468 | 0.281 | 0.473 | 11.0 | 6.6 | 11.1 |
34 | 17.6 | 0.312 | 0.388 | 0.318 | 5.5 | 6.8 | 5.6 |
41 | 15.6 | 0.301 | 0.216 | 0.304 | 4.7 | 3.4 | 4.7 |
43 | 0.4 | 0.322 | 0.023 | 0.323 | 0.1 | 0.0 | 0.1 |
44 | 0.1 | 0.406 | 0.227 | 0.410 | 0.1 | 0.0 | 0.1 |
45 | 15.1 | 0.210 | 0.239 | 0.213 | 3.2 | 3.6 | 3.2 |
Total Inferred | 840.0 | 0.427 | 0.382 | 0.432 | 358.3 | 320.8 | 363.3 |
Notes: | ||
1. | Mineral resources have been calculated at a gold price of $1,200/troy ounce and a silver price of $19.00 per troy ounce; | |
2. | Mineral resources are calculated at a grade thickness cut-off grade of 1.126 Au equivalent opt-feet and a diluted Au equivalent cut-off grade of 0.256 opt; | |
3. | Gold equivalent ounces were calculated based on one ounce of gold being equivalent to 64.53 ounces of silver; | |
4. | The minimum Mining width is defined as four feet or the vein true thickness plus one foot, whichever is greater; | |
5. | Mineral resources include dilution to achieve mining widths and an additional 10% unplanned dilution. | |
6. | Mineral resources include allowance for 5% mining losses; | |
7. | Mineral resources are inclusive of mineral reserves; | |
8. | Mineral resources, which are not mineral reserves, do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, title, socio-political, marketing, or other relevant issues, and; | |
9. | The quantity and grade of reported inferred mineral resources in this estimation are uncertain in nature and there is insufficient exploration to define these inferred mineral resources as an indicated or measured mineral resource and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category. | |
10. | Mineral resource estimates can be materially affected by environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other factors. |
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15. | Mineral Reserve Estimates |
Excavation designs for stopes, stope development drifting and access development were created using Vulcan software. Stope designs were aided by the Vulcan Stope Optimizer Module. The stope optimizer produces the stope cross section which maximizes value within given geometric and constraints.
Design constraints included four feet minimum width for long hole stopes with development drifts spaced at 40-foot vertical intervals. Stope development drift dimensions maintained a constant height of ten feet and a minimum width of six feet. Drift and fill dimensions are the same as stope development.
Mining and backfill tasks were created from all designed excavations. These tasks were assigned costs and productivities specific to the excavation or backfill task type. Additionally, the undiscounted cash flow for each task was calculated. All tasks were then ordered in the correct sequence for mining and backfilling. Any task sequence or subsequence that did not achieve a positive cumulative undiscounted cash flow was removed from consideration for mineral reserves. Stope development, necessary to reach reserve excavations and exceeding the incremental cut-off grade shown in Table 15-1, are also included in mineral reserves.
Table 15-1 Mineral Reserves Cut Off Grade Calculation
Gold | Silver | ||
Sales Price | $/Ounce | $1,000 | $15.83 |
Refining and Sales Expense | $/Ounce | Included in Milling | |
Royalty | 1% | ||
Metallurgical Recovery | 94% | 92% | |
Operating Costs | |||
Ore Haulage (Portal to Mill) | $/ton | $32.66 | |
Direct Processing | $/ton | $93.10 | |
Administration and Overhead | $/ton | $115.49 | |
Mining | $/ton | $218.37 | |
Total | $/ton | $459.62 | |
Gold Equivalent | 1 | 64.54 | |
Unplanned Dilution | 10% | ||
Incremental Cut Off Grade | 0.259 | ||
Cut-off Grade | Eq. opt | 0.494 | |
Minimum Mining Width | feet | 4 | |
Grade Thickness cut-off | Eq. opt-ft. | 2.173 |
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Table 15-2 Fire Creek Mineral Reserves as of December 31, 2014
Au | Ag | Au Equiv. | |||||
Tons | Au Eq | Ounces | Ounces | Ounces | |||
Vein Designation | (000's) | Au opt | Ag opt | opt | (000's) | (000's) | (000's) |
Proven Reserves | |||||||
Joyce | 31 | 0.914 | 0.678 | 0.924 | 27.9 | 20.7 | 28.2 |
Vonnie | 9.3 | 3.301 | 2.151 | 3.335 | 30.6 | 19.9 | 30.9 |
Karen | 41 | 1.454 | 1.192 | 1.473 | 59.7 | 49.0 | 60.5 |
Proven Reserves | 80.9 | 1.462 | 1.108 | 1.479 | 118.2 | 89.6 | 119.6 |
Probable Reserves | |||||||
Joyce | 60 | 0.779 | 0.357 | 0.784 | 47.0 | 21.5 | 47.3 |
Vonnie | 34 | 1.920 | 1.626 | 1.945 | 66.1 | 56.0 | 67.0 |
Karen | 10 | 0.733 | 0.500 | 0.741 | 7.4 | 5.0 | 7.5 |
Probable Reserves | 104.9 | 1.149 | 0.787 | 1.161 | 120.5 | 82.6 | 121.8 |
Proven + Probable Reserves | |||||||
Joyce | 91 | 0.824 | 0.464 | 0.831 | 74.9 | 42.2 | 75.6 |
Vonnie | 44 | 2.213 | 1.738 | 2.240 | 96.7 | 75.9 | 97.9 |
Karen | 51 | 1.312 | 1.056 | 1.328 | 67.1 | 54.0 | 68.0 |
Proven + Probable Reserves | 185.8 | 1.285 | 0.927 | 1.300 | 238.7 | 172.2 | 241.4 |
Notes: | ||
1. | Mineral Reserves have been estimated with a gold price of $1,000/ounce and a silver price of $15.83/ounce; | |
2. | Metallurgical recoveries for gold and silver are 94% and 92% respectively; | |
3. | Gold equivalent ounces are calculated on the basis of one ounce of gold being equivalent to 64.53 ounces of silver, and; | |
4. | Mine losses of 5% and unplanned mining dilution of 10% have been applied to the designed mine excavations. |
Fire Creek mineral reserves could be materially affected by economic, geotechnical, permitting, metallurgical or other relevant factors. Mining and processing costs are sensitive to production rates. A decline in the production rate can cause an increase in costs and cutoff grades resulting in a reduction in mineral reserves. Geotechnical conditions requiring additional ground support or more expensive mining methods will also result in higher cutoff grades and reduced mineral reserves.
The Project has the necessary permits to continue exploration and current operations. Failure to maintain permit requirements may result in the loss of critical permits necessary for continued operations.
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16. | Mining Methods |
16.1. | Mine Development |
16.1.1. Access Development |
Access to the mining areas will be from haulage drifts, up to 15 feet wide and between 15 to 17 feet high. Drift gradients will vary from – 15% to + 15% to reach the desired elevation. Secondary drifts, spiral ramps and vertical raises will connect the haulage drifts to provide a pathway for ventilation to the surface and serve as a secondary escape way. (Figure 16-1)
16.1.2. Ground Support |
The ground conditions at the Project are typical of the northern Nevada extensional tectonic environment. Joint spacing varies from a few inches to a foot or more. To date, split sets and Swellex rock bolts along with welded wire mesh have been successfully employed to control all conditions encountered during decline development and stoping. Shotcrete has also been liberally applied to prevent long-term deterioration of the rock mass.
All major access drifts require a minimum of wire mesh and rock bolts for support. Under more extreme conditions, resin anchor bolts, cable bolts, and shotcrete can be used to supplement the primary support. Steel sets and spiling may also be used to support areas with the most severe ground conditions.
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16.1.3. Ventilation and Secondary Egress |
Underground mining relies heavily on diesel equipment to extract the mineralized material and waste rock and to transport backfill to the stopes. Diesel combustion emissions will require large amounts of fresh ventilation air to remove the diesel exhaust and maintain a healthy working environment. A combination of the main access drifts and vertical raises to the surface are arranged in a manner to provide a complete ventilation circuit. The mine portal can be used as either an intake or an exhaust. Air movement will be facilitated by primary ventilation fans placed at the surface or underground in strategic locations. Small auxiliary fans and ducting will draw primary ventilation air directly into the working faces.
The ventilation raise connecting the main decline to the surface is approximately 690 feet in length and is entirely lined with corrugated metal pipe to support the ribs and maintain a uniform cross sectional area. Since the vertical extent of the raise exceeds the maximum 300 feet permitted for a continuous ladder way, it has been equipped with an automatic hoist and personnel capsule for evacuating the mine in the event of an emergency.
16.2. | Mining Methods |
Mining may be completed using end slice stoping with delayed backfill, also referred to as long hole stoping, and drift and fill stoping. The final choice of mining method will depend upon the geometry of the stope block, proximity to main access ramps, ventilation and escape routes, the relative strength or weakness of the mineralized material and adjacent wall rock, and finally the value or grade of the mineralized material. The choice of mining method will not be made until after the stope delineation and definition drilling is completed. Each method will be discussed briefly in 16.2.1. the following paragraphs.
16.2.1. End Slice Stoping |
End slice, or long hole, stoping has the highest degree of mechanization of the three expected mining methods at the Project, is the lowest cost method and generally provides the lowest total cost per ounce. End slice stoping requires the greatest amount of waste development and can be mined to a minimum width of four feet. The potential for unplanned wall dilution with this method is the greatest. The current reserve mine plan incorporates 67% end slice stoping for exploitation of the reserve. Figure 16-2 shows a typical end slice stoping arrangement.
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To prepare an area for end slicing, access for the mobile equipment must be developed to each level. Mine utilities for communication, water, electrical power, and compressed air must also be provided through the access development. Level spacing is limited to 40 feet to control dilution and may be increased if vein geometry, ground conditions, and vein thickness are favorable. The minimum level spacing achievable with this method is 30 to 35 feet and is limited by the stability of the intervening pillar between levels. Mining will progress upwards from the lowest level of the stope block. Drilling and blasting will be carried out from the drift above the active stope while the broken mineralized material will be removed from the bottom drift. The loader used for excavation is equipped with line of sight remote control to allow the removal of all blasted rock without exposing the operator to the open stope and the potential risk of ground falls.
The amount of mineralization that can be removed prior to backfilling will be constrained by the strength of the gangue material and jointing present immediately adjacent to the stope. Backfill, consisting of either waste rock or cemented rock fill, will be transported from the surface using the same haulage equipment used to remove mineralized material and waste rock from the mine. Where possible, waste rock will be retained within the mine and placed directly into a stope requiring backfill. The stope will be backfilled from the drift used for drilling and blasting.
Cemented rock fill (CRF), which consists of screened mine waste, fly ash, and cement will be mixed on the surface and transported underground in the same trucks used to haul blasted rock to the surface. CRF will be placed to create an artificial pillar where additional mining is planned adjacent to or underneath the stope being filled. Normal backfill unconfined compressive strengths (UCS) of 400 to 600 pounds per square inch (psi) will be achieved by blending a mixture containing up to four percent cement and fly ash. When mining is anticipated to occur below the backfilled stope, the UCS of the fill will be increased up to 1,000 psi by adding up to eight percent cementatious binder.
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A typical end slice stoping arrangement at Fire Creek is presented in Figure 16-3. Stope development drifting is planned at six feet wide and ten feet high to accommodate the production drill. Levels are located at 40-foot vertical intervals to control dilution and may be increased as experience is gained in mining the Fire Creek veins. Stope widths have been designed at either four feet or the horizontal vein thickness plus two feet, whichever is greater.
All stope cross sections were calculated using Vulcan stope optimizer software. This software calculates the optimal stope dimensions and orientation within the dip constraints on the hanging wall and footwall as well as the given level spacing.
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16.2.2. Drift and Fill Stoping |
Drift and fill stoping will be used to extract 33% of the Project mineral reserves. This method can be employed where the wall rock is too weak for end slice stoping, the vein dip is less than 50° or where there is variable vein geometry. Cut and fill stoping is the the highest cost mining method of the two considered. A typical cut and fill stope arrangement is shown in Figure 16-3.
A drift in fill stope is initiated by driving a waste crosscut from the access ramp to the vein. The cross cut is driven at a negative gradient up to minus 15% in order to reach the lowest elevation of the stope. Drifting along the vein strike progresses in both directions from the cross cut. Drift dimensions are a minimum of six feet in width and 10 feet high. The width can be increase to accommodate wider sections of the vein.
Once the end of the stope is reached, the drift can be backfilled with CRF if there is unmined ore below or with unconsolidated waste backfill (GOB) if mining below is not planned. Once filled, breasting down the waste above the back of the cross cut begins at a gradient sufficient that the sill of the crosscut is now at the same elevation as the back of the preceding drift. This process will be repeated until vein within reach from the cross cut has been mined out, and mining will proceed from the next level above.
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16.3. | Underground Labor |
Peak underground work force requirements for the Project are presented in Table 16-1. This estimate was prepared using productivity rates typical for small-scale mechanized mining in North America. The Project will operate 24 hours per day seven days per week. Project operations workforce will be divided into four crews scheduled to work 14 out of every 28 days.
Table 16-1 Underground Workforce 2015 and 2016
Job Classification | Count |
Miners | 45 |
Mechanics | 15 |
Supervision | 6 |
Technical Staff | 12 |
Manager | 1 |
Total | 79 |
16.4. | Mobile Equipment Fleet |
Development drifting averages 12 feet per day in 2015 and six feet per day in 2016. All capital development is completed by the end of 2016. Production rates of 150 tons per day (tpd) are planned in the first year and will peak in year two at 180 tpd. Table 16-2 lists the mining fleet necessary to achieve the development and production goals outlined in the mine plan. The majority of this equipment was surplus equipment from the Midas Mine purchase by Klondex in 2014 and is on site. The stope production drill is the only major piece of equipment that has not yet been acquired.
Table 16-2 Underground Mobile Equipment
Units on | Additional | |
Description | Site | Units Req’d |
2 boom jumbo | 1 | |
Single boom jumbo | 1 | |
Bolter | 1 | |
6 Yard LHD | 1 | |
4 Yard LHD | 1 | |
2 Yard LHD | 2 | |
15 - 20 ton trucks | 2 | |
30 Ton Trucks | 2 | |
Stope ring drill | 1 |
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Units on | Additional | |
Description | Site | Units Req’d |
Shotcrete pump, robotic spray arm and cariier | 1 | |
Shotcrete 16.5. remix trucks | 2 |
16.5. | Mine Plan |
The productivities of Table 16-3 represent typical values achieved in northern Nevada for the type and size of excavations planned at the Project. These productivities were used to develop the production plan shown in Figure 16-4 through Figure 16-6 and Table 16-4.
The production plan is dictated by the number of available stoping areas. Increasing the available labor force or the equipment fleet will not have significant impact on the production rate or anticipated mine life of 3.8 years. The decline in production rate in the second half of the mine plan is the result of exhausting the available stoping areas.
Table 16-3 Heading Productivity
Heading Type | Units | Daily Rate |
Capital Development Drift | Feet/day | 16 |
Drop Raise | Feet/Day | 5 |
Stope Development (6 x 10) | Feet/day | 21 |
End Slice (Long Hole) Stoping | Ton/day | 160 |
Drift and Fill Stoping | Ton/Day | 100 |
Backfill | Ton/Day | 200 |
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Table 16-4 Annual Production and Development Plan
Calendar Year | 2015 | 2016 | 2017 | 2018 | Total |
Reserves Mined | |||||
Proven Ore Mined (000's Tons) | 21.8 | 19.7 | 16.6 | 22.7 | 80.9 |
Gold Grade (Ounce/Ton) | 1.066 | 1.235 | 1.581 | 1.952 | 1.462 |
Silver Grade (Ounce/Ton) | 0.810 | 0.889 | 1.317 | 1.431 | 1.108 |
Contained Gold (000's Ounces) | 23.3 | 24.3 | 26.3 | 44.4 | 118.2 |
Contained Silver (000's Ounces) | 17.7 | 17.5 | 21.9 | 32.5 | 89.6 |
Probable Ore Mined (000's Tons) | 31.0 | 46.7 | 22.1 | 5.2 | 104.9 |
Gold Grade (Ounce/Ton) | 0.810 | 1.203 | 1.460 | 1.361 | 1.149 |
Silver Grade (Ounce/Ton) | 0.488 | 0.886 | 0.967 | 0.915 | 0.787 |
Contained Gold (000's Ounces) | 25.1 | 56.1 | 32.2 | 7.1 | 120.5 |
Contained Silver (000's Ounces) | 15.1 | 41.4 | 21.3 | 4.8 | 82.6 |
Total Reserves Mined (000's Tons) | 52.8 | 66.4 | 38.7 | 27.9 | 185.8 |
Gold Grade (Ounce/Ton) | 0.916 | 1.212 | 1.512 | 1.842 | 1.285 |
Silver Grade (Ounce/Ton) | 0.621 | 0.887 | 1.117 | 1.335 | 0.927 |
Contained Gold (000's Ounces) | 48.4 | 80.5 | 58.5 | 51.4 | 238.7 |
Contained Silver (000's Ounces) | 32.8 | 58.9 | 43.2 | 37.3 | 172.2 |
Contained Gold Equiv. (000's Ounces) | 48.9 | 81.4 | 59.1 | 52.0 | 241.4 |
Production Mining | |||||
Stope Development and Drift and Fill Mining (000's Tons) | 38.9 | 23.4 | 62.2 | ||
Longhole Stope Mining (000's Tons) | 13.9 | 43.0 | 38.7 | 27.9 | 123.5 |
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Calendar Year | 2015 | 2016 | 2017 | 2018 | Total |
Reserves Mined (000's Tons) | 52.8 | 66.4 | 38.7 | 27.9 | 185.8 |
Reserves Mining Rate (tpd) | 144.6 | 181.3 | 106.0 | 102.3 | 135.7 |
Backfill | |||||
Cellular Backfill (000's Tons) | 5.2 | 8.5 | 13.7 | ||
CRF and GOB Backfill(000's Tons) | 13.3 | 63.7 | 42.9 | 15.9 | 135.8 |
Total Backfill (000's Tons) | 18.5 | 72.2 | 42.9 | 19.8 | 153.4 |
Waste Mining | |||||
Expensed Drift Waste (000's Tons) | 6.8 | 4.9 | 11.8 | ||
Bench Waste (000's Tons) | 0.4 | 0.4 | |||
Expensed Waste (000's Tons) | 1.1 | 0.1 | 1.2 | ||
Primary Capital Drifting (Feet) | 4,420 | 2,296 | 6,716 | ||
Secondary Capital Drifting (Feet) | 578 | 689 | 1,267 | ||
Capital Raising (Feet) | 156 | 103 | 259 | ||
Capitalized Mining (000's Tons) | 79.2 | 47.3 | 126.5 | ||
Total Tons Mined (000's Tons) | 139.2 | 118.7 | 38.7 | 27.9 | 324.5 |
Mining Rate (tpd) | 381 | 324 | 106 | 102 | 237 |
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17. | Recovery Methods |
A local contractor transports mineralized material from the bulk sampling program at the Fire Creek Project to the Midas Mill on public roadways which is a distance of approximately 131 miles. Mineralized material from each mine is segregated through the crushing circuit. The mill, has two 500-ton fine ore bins located between the secondary crusher and the ball mill and one bin is dedicated to each mine. Head samples are taken on each reclaim conveyor at regular intervals, and tonnage measured by a belt scale prior to comingling the mineralization streams.
The Midas Mill was constructed in 1997 and has a nameplate capacity of 1,200 tpd. The mill uses conventional leach technology with Counter Current Decantation (CCD) followed by Merrill Crowe precipitation. Doré refining is finalized by Johnson Matthey refineries in Salt Lake City, Utah. Midas has performed toll milling periodically since 2008.
17.1 | Mill Capacity and Process Facility Flow Diagram |
A process facility flow sheet is shown in Figure 17-1. Underground mineralized material extracted under the bulk sampling permit is delivered from the Fire Creek Project and the Midas Mine to the run of mine (ROM) pad where it is placed on short term ROM mineralized material stockpiles. Typical mineralized material classifications are: low grade less than 0.3 opt Au or less than six opt silver; high grade (0.3 to 0.5 opt gold or six to 20 opt silver); and ultra-high grade less than 0.5 opt gold or less than 20 opt silver). Separate stockpiles are maintained for each mine. Underground mineralized material is hand-picked on the pad for scrap wire mesh and rock bolts before being fed to the crusher.
Mineralized material is crushed in two stages through a 30-inch by 40-inch primary jaw crusher and 53-inch secondary cone crusher. Both jaw and secondary crusher products are fed to a six feet by 20 feet Nordberg double deck vibrating screen fitted with two-inch top deck and one-half inch bottom deck screen panels to produce a 95% passing one-half inch product. Magnetic material is removed from the crusher screen feed by a continuous self-cleaning belt magnet to protect the cone crusher from damage. Screen undersize is conveyed to one of two 500-ton fine mineralized material bins.
Crushed and screened material is transported from the fine material bins by individual belt feeders into the 10.5 feet by 15 feet rubber lined Nordberg ball mill. The ball mill is charged with a blend of three-inch and two-inch grinding balls to maintain an operating power draw of 800 horse power (HP). Mill discharge pulp is pumped to a nest of four-inch by ten-inch Krebs cyclones (three duty, one standby) for classification. Cyclone overflow, at 85% passing 200 mesh, reports to the trash screen. Cyclone underflow reports to a two millimeters (mm) aperture scalping screen, with the screen undersize being distributed by three-way splitter to the ball mill, verti-mill, and gravity circuit. Lead nitrate solution is added to the ball mill feed chute to enhance silver leach kinetics.
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A split of the screened cyclone underflow reports to the 250 HP verti-mill for open circuit grinding with the verti-mill discharge overflowing back to the primary ball mill discharge pump box. The verti-mill is charged with one inch grinding balls. A split of the screened cyclone underflow also reports to the 20-inch Knelson concentrator for gravity gold recovery. The Knelson operates on a 30-minute cycle providing concentrate for cyanidation in the CS500 Acacia Leach Reactor which conducts three 750 to 1,000 kg batch leaches each week. Pregnant solution from the leach reactor reports to the CCD circuit pregnant solution tank.
Cyclone overflow is screened to remove any plastic debris before reporting to a 42 feet diameter pre-leach thickener. Thickener underflow at 50% solids is pumped to the leach circuit consisting of eight 28 feet by 30 feet air sparged leach tanks, providing a leach residence time of approximately 90 hours at 600 tons per hour (tph) feed rate. The pH in the first leach tank is maintained at 10.4 to 11.0 through the addition of hydrated lime, produced from the on-site slaking of pebble lime. Sodium cyanide concentration in the second leach tank is maintained at 1.25 grams per liter (gpl).
The leach circuit discharge is pumped to the first of five 42.5 feet diameter CCD thickeners, where the pulp is counter-current washed with barren Merrill Crowe liquor at a wash ratio of approximately 3.2:1 . CCD thickener underflow at each stage is maintained at between 50 and 54% solids to maximize wash efficiency.
Pregnant CCD solution at a pH of 11.0 and 400 gallons per minute flow rate is fed to one of two disc filters operating in duty/standby mode utilizing diatomaceous earth for clarification. The clarified pregnant solution is then pumped to a packed bed vacuum de-aeration tower, prior to the addition of zinc dust and lead nitrate to precipitate precious metals from solution. The Merrill Crowe solution is then pumped to one of two plate and frame filter presses for sludge recovery. The sludge is collected from a filter press weekly and smelted to produce 5,500 ounce silver and gold doré bars.
Tailings pulp from the last CCD thickener is pumped to the Inco SO2/Air circuit for cyanide destruction. Cyanide destruction is performed in a single 20 feet by 20 feet agitated, air sparged tank providing approximately one-hour reaction time. Ammonium bi-sulphate, lime, and copper sulphate as a catalyst are added to the tank on a ratio control basis to achieve target weak acid dissociable (WAD) cyanide levels below five ppm. Routine picric acid analyses are used by operating personnel to maintain WAD cyanide in the INCO cyanide destruction tank discharge pulp at target levels.
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Following cyanide destruction, the plant tailings pulp is discharged to one of two lined TSF for consolidation and water recovery. Clarified decant pond solution is evaporated or returned to the mill process water tank for reuse in the plant.
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17.2. | Physical Mill Equipment |
The Midas Mill equipment list is shown in Table 17-1.
Table 17-1 Process Equipment Itemization by Area
Description | Number | Spare | Note | Description | Number | Spare | Note |
AREA 350 GRINDING | |||||||
BIN, MILL TROMMEL | 1 | HEATER, MILL FEED | 1 | 5 kW | |||
REJECTS | CONVEYOR GALLERY | ||||||
CS Construction, w/lift | w/fan | ||||||
lugs, 6.5' x 6.5' x 4' | |||||||
CHUTE, BALL | 1 | LAUNDER, MILL | 1 | ||||
TRANSFER | DISCHARGE | ||||||
CS, Rubber Lined | |||||||
CHUTE. FINE ORE BIN | 1 | PUMP BOX, CYCLONE FEED | 1 | ||||
DISCHARGE | 6' x 6' x 6', 1200 gal, CS, Rubber | ||||||
CS Plate Construction, | Lined | ||||||
AR Plate Lined | |||||||
CHUTE, FINE ORE | 1 | PUMP, CYLCONE FEED | 1 | 1 | 50 HP | ||
FEEDER DISCHARGE | 550 gpm, 4 x 3, Centrifugal | ||||||
CS Plate Construction, | Slurry, VFD, Rubber Lined CS | ||||||
AR Lined | |||||||
CHUTE, MILL FEED | 1 | SAMPLER, CYCLONE | 1 | 0.5 HP | |||
Includes ball charge | OVERFLOW | ||||||
attachment, CS | 223 gpm, single stage slurry | ||||||
Construction, AR Lined | cutter, CS Rubber Lined | ||||||
CHUTE, BALL | 1 | BELT SCALE, MILL FEED | 1 | ||||
DISCHARGE | 30 tph, 24", 4 idler weigh bridge | ||||||
CS Plate Construction, | |||||||
AR Plate Lined | |||||||
CHUTE, MILL | 1 | CYCLONE PACKAGE | 2 | ||||
TROMMEL COVER | 2 - DS15LB-1826 Cyclones, | ||||||
CS Plate Construction | radial manifold, w/ launders | ||||||
CHUTE, MILL | 1 | DUST COLLECTOR | 1 | 20 HP | |||
TROMMEL REJECTS | PACKAGE | ||||||
CS Plate Construction | PULSE Air, induction, 5000 | ||||||
cfm, 0.5 psi | |||||||
CONVEYOR, MILL | 1 | 7.5 HP | FEEDER, FINE ORE | 1 | 5 HP | ||
FEED | DISCHARGE | ||||||
30 tph, troughed rubber | Rotary Valve | ||||||
type, 36" width, 116' | |||||||
Length, 12' lift, 50 fpm | |||||||
FAN, FINE ORE | 2 | 1.0 HP | LUBE SYSTEM, BALL MILL | 2 | 5 kW | ||
LOWER BUILDING | Air operated, w/heater | ||||||
VENT 4000 cfm, | |||||||
Wall exhaust | |||||||
FEEDER, FINE ORE | 1 | 5.0 HP | MILL, BALL | 1 | 800 HP | ||
30 tph, 30" width, 29' | 10.5' Diameter, 14' Length, | ||||||
length, VFD | Rubber Lined | ||||||
AREA 410 LEACH |
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Description | Number | Spare | Note | Description | Number | Spare | Note |
Knelson gravity | CS 500 Acacia leach reactor | ||||||
concentrator, 20 inch | |||||||
AGITATOR, LEACH | 8 | 40 HP | SAMPLER, LEACH TAILS | 1 | 0.5 HP | ||
109" Diam., Dual | 330 gpm, Slurry Cutter | ||||||
Impellers, 8' sch 80 Shaft, | |||||||
292" Length, CS | |||||||
Construction, Rubber | |||||||
Lined | |||||||
FAN, PRE-LEACH | 1 | 0.5 HP | SCREEN, TRASH | 2 | 2.5 HP | ||
THICKENER VENT | 4' X 5', Vibrating | ||||||
3000 CFM @ 0.25 WG | |||||||
HEATER, PRE-LEACH | 1 | 35 HP | STANDPIPE, PRE-LEACH | 1 | |||
THICKENER VENT | THICKENER O/F 2.5' | ||||||
40,000 BTU, propane | Diam., 20' high, Open Top, CS | ||||||
Construction | |||||||
LAUNDER, LEACH, | 8 | PUMP BOX, CCD FEED | 1 | ||||
INTERTANK | SPLIT TO #1 AND #2 600 | ||||||
CS Construction, w/Gate | gal, 4X4X6' w/weirs, CS | ||||||
Construction, Rubber Lined | |||||||
LAUNDER, LEACH, | 7 | PUMP, PRE-LEACH | 1 | 7.5 HP | |||
INTERTANK bypass | THICKENER AREA SUMP | ||||||
CS Construction, w/Gate | 200 gpm, 2.5" Diam. Vertical | ||||||
Slurry, Rubber Lined | |||||||
AREA 410 LEACH | |||||||
PUMP BOX, LEACH | 1 | PUMP, LEACH THICKENER | 1 | 7.5 HP | |||
TAILS | AREA SUMP 200 gpm, 2.5" | ||||||
6' x 6' x 6', 1200 gal, CS, | Diam. Vertical Slurry, Rubber | ||||||
Rubber Lined | Lined | ||||||
PUMP, LEACH TAILS | 2 | 1 | 7.5 HP | TANK, LEACH | 8 | ||
327 gpm, 4X3, | 28' x 30', Open top, CS | ||||||
Centrifugal, CS Rubber | Construction | ||||||
Lined | |||||||
PUMP, PRE-LEACH | 1 | 1 | 15 HP | THICKENER, PRE-LEACH | 1 | 15 HP | |
THICKENER O/F | 59.5' Diameter, 19.5' Height, | ||||||
533 gpm, 3X4, | Feed well, All Gear, CS | ||||||
Centrifugal, CS | Construction | ||||||
Construction, Packed Seal | |||||||
PUMP, PRE-LEACH | 1 | 10 HP | |||||
THICKENER U/F | |||||||
330 gpm, 3X4, | |||||||
Centrifugal, CS | |||||||
Construction, Rubber | |||||||
Lined | |||||||
AREA 430 CCD THICKENING | |||||||
FAN, CCD ARE VENT | 4 | 1 HP | PUMP, CCD THICKENER U/F | 5 | 5 | 4.5 HP | |
6000 cfm, Wall Exhaust | ADVANCE | ||||||
160 gpm, 3X4, Centrifugal, CS | |||||||
Construction, Packed Seal |
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Description | Number | Spare | Note | Description | Number | Spare | Note |
HEATER, CCD ARE | 4 | 1 HP | SAMPLER, LEACH TAILS | 1 | 0.5 HP | ||
VENT | 330 gpm, Slurry Cutter | ||||||
20 MBH, Propane | |||||||
w/motor | |||||||
PUMP, LEACH CCD | 1 | 7.5 HP | STANDPIPE, CCD thickener | 5 | |||
AREA SUMP | 2.5' Diam., 20' high, Open Top, | ||||||
200 gpm, 2.5" Diam. | CS Construction | ||||||
Vertical Slurry, Rubber | |||||||
Lined | |||||||
PUMP, CCD | 5 | 1 | 7.5 HP | THICKENER, CCD | 5 | ||
THICKENER O/F | 42.5' Diam. 19.5' high, feed | ||||||
ADVANCE | well, all gear | ||||||
300 gpm, 3X4, | |||||||
Centrifugal, CS | |||||||
Construction, Packed Seal | |||||||
AREA 450 CYANIDE DESTRUCTION | |||||||
AGITATOR, CYANIDE | 1 | 125 HP | TANK, CYANIDE | 1 | |||
DESTRUCTION | DESTRUCTION | ||||||
121" Diam., Dual | 20' X 20', Open Top, CS | ||||||
Impellers, 10' sch 160 | Construction | ||||||
Shaft, 292" Length, CS | |||||||
Const., Rubber Lined | |||||||
SAMPLER, CYANIDE | 1 | 0.5 HP | |||||
DESTRUCTION | |||||||
200 gpm, Slurry Cutter | |||||||
AREA 470 TAILING HANDLING | |||||||
PUMP, TAILINGS | 1 | 10 HP | PIPE, TAILINGS | 800 ft | |||
DISTRUBUTION | 8" HDPE, SDR 11 | ||||||
420 gpm, 3X4, | |||||||
Centrifugal, CS | |||||||
Construction, Rubber | |||||||
Lined | |||||||
PUMP, CCD | 5 | 1 | 7.5 HP | PIPE, TAILINGS | 800 ft | ||
THICKENER U/F | 12" HDPE, SDR 11 | ||||||
ADVANCE | |||||||
160 gpm, 3X4, | |||||||
Centrifugal, CS | |||||||
Construction, Rubber | |||||||
Lined | |||||||
AREA 510 MERRILL CROWE | |||||||
FILTER, CLARIFYING | 1 | 1 HP | PUMP, PREGNANT | 1 | 1 | 30 HP | |
400 ft2, 210 gpm, 25 ppm | SOLUTION | ||||||
solids, 54" diam. X 8', | 600 gpm, 3X4, CS Construction | ||||||
flushing | |||||||
PUMP, BARREN | 1 | 1 | 15 HP | PUMP, FILTER FEED | 1 | 1 | 15 HP |
SOLUTION | 600 gpm, 3X4, CS Construction, | ||||||
600 gpm, 4X8, | flooded mechanical seal | ||||||
Centrifugal, CS | |||||||
Construction |
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Description | Number | Spare | Note | Description | Number | Spare | Note |
FEEDER, ZINC | 1 | TANK, DEAERATION | 1 | ||||
50 lb./hr | 3' Diam. X 20' high, 22 in. water | ||||||
vacuum | |||||||
AREA 550 REAGENTS | |||||||
PUMP, FLOCCULANT | 1 | 1.5 HP | PUMP, ABS METERING | 1 | 1 | ||
METERING | 75 gpm, Metering Type, | ||||||
2 gpm, Progressive Cavity | Mechanical Seal | ||||||
PUMP, FLOCCULANT | 5 | 1 HP | TANK, COPPER SULFATE | 1 | |||
METERING | STORAGE 2900 | ||||||
0.5 gpm, Progressive | gal, 8' Diameter X 9' high, | ||||||
Cavity | closed, SS Construction | ||||||
PUMP, REAGENT | 3 | 1 | 1 HP | FLOCCULANT PACKAGE, | 1 | 3 HP | |
METERING | SELF CONTAINED Includes | ||||||
25 gpm, Metering Type | Agitator, Blower, Bin Feeder, | ||||||
Mixer, Tanks, SS Construction | |||||||
AREA 650 UTILITIES | |||||||
PUMP, PROCESS | 1 | 125 HP | BLOWER, CYANIDE | 1 | 75 HP | ||
WATER | DETOXIFICATION | ||||||
1200 gpm, 6X8, CS | 1000 cfm, Rotary, Two Stage, | ||||||
Construction, Packed Seal | Intercooler, Filter Intake | ||||||
BLOWER, LEACH | 1 | 30 HP | |||||
TANK | |||||||
320 cfm @ 20 psig, | |||||||
Rotary, Two Stage, | |||||||
Intercooler, Filter Intake |
17.3. | Operation and Recoveries |
Fire Creek mineralization performs quite well under direct cyanidation with daily recoveries as high as 95.1% for gold and up to 95% for silver. The process performance is consistent with gold recovery having a standard deviation of less than two percent. Variances in gold recovery are due to the head grade and grind size, and do not appear to be associated with mineralized material type. The standard deviation of silver recovery is less than four percent with variance due to head grade, grind size, and clay content. Clay enriched mineralization often has higher silver to gold ratios and tend to present recovery difficulties. Recoveries occasionally fall outside the expected distribution because of plant or operating issues. The current grind is 85% passing through 200 mesh. The feasibility of producing a finer grind product to improve gold and silver recovery is currently under analysis by Klondex.
17.4. | Tailings Storage Capacity |
Newmont and Klondex have evaluated the potential for increasing TSF capacity. The remaining capacity in the existing Midas Phase 4/5 TSF is estimated to be 600,000 tons, as of year-end 2014. Two alternatives are available for increasing TSF capacity. The first would raise the existing embankment approximately four feet using an engineered retaining wall. This option would add approximately 400,000 tons capacity and is estimated to cost $1M. This option has the advantage of staying inside the existing TSF footprint and can be permitted with a minor modification to the existing plan of operations. The second option would involve new construction outside the existing TSF footprint. Permit modifications would likely take two to three years to secure. New TSF construction could be completed by early 2018. Klondex is proceeding with construction of the four-foot embankment raise.
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The 2012 performance of the TSF included evaporation of over 90 M gallons of water utilizing ten evaporator units.
Currently, 14 evaporators operate 24 hours per day, seven days per week during the peak evaporation season to assist in the evaporation of excess water in the TSF.
17.5. | Processing Costs |
Midas Mill operating costs for 2012 and to 2014 are summarized in Table 17-2.
Table 17-2 Midas Mill Operating Costs
$/ton | Total Tonnage | |||||
Year | Budget | Actual | Variance | Budget | Actual | Variance |
2012 | $33.12 | $35.02 | $1.90 | 373,000 | 330,000 | -43,000 |
2013 | $35.49 | $39.05 | $3.56 | 255,600 | 207,600 | -48,000 |
2014 1 | $62.53 | $57.49 | -$5.04 | 174,425 | 171,818 | -2,607 |
Notes: | ||
1. | Klondex has only been the operator of the Midas Mill since February 19, 2014. Newmont was the prior operator. |
The elevated cost per ton for 2013 is likely the result of the inflexibility of fixed costs versus diminished throughput. If the total cash costs are divided by the budgeted tonnage, the average cash cost per ton would be US$32 per ton, which is more in line with the projected costs.
Future processing cost projections reflect 2014 consumption rates and pricing levels for reagents, and electrical power. Adequate water is available from onsite supply wells and the Midas Underground Mine.
17.6. | Production |
Doré is shipped to the refinery as 5,500-ounce bars that average approximately 3.94% gold and 90.1% silver plus minor constituents, including less than two percent selenium. Table 17-3 provides a monthly summary of the processing at the Midas Mill of mineralized material extracted at Fire Creek under the bulk sampling permit.
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Table 17-3 2014 Fire Creek Mineralized Material Processed at the Midas Mill
Feb. 1 | March | April | May | June | July | Aug | Sept. | Oct. | Nov. | Dec. | 2014 | |
Tons (000's) | 2.3 | 5.5 | 2.7 | 5.7 | 6.0 | 5.7 | 6.1 | 5.7 | 6.8 | 4.6 | 4.0 | 55.0 |
Au grade | 0.765 | 1.512 | 0.724 | 1.030 | 0.959 | 1.227 | 2.061 | 1.566 | 1.184 | 1.215 | 0.789 | 1.252 |
Ag grade | 0.77 | 1.51 | 0.56 | 0.75 | 0.69 | 1.04 | 1.82 | 1.77 | 1.37 | 1.44 | 0.95 | 1.21 |
feed Au oz (000's) | 1.8 | 8.3 | 1.9 | 5.9 | 5.7 | 7.0 | 12.6 | 8.9 | 8.0 | 5.6 | 3.2 | 68.8 |
feed Ag oz (000's) | 1.8 | 8.3 | 1.5 | 4.3 | 4.1 | 6.0 | 11.1 | 10.0 | 9.3 | 6.6 | 3.8 | 66.7 |
% Au Rec. | 94.7% | 94.5% | 99.9% | 89.9% | 92.7% | 94.8% | 95.7% | 93.1% | 93.8% | 95.1% | 91.8% | 94.1% |
% Ag Rec | 95.7% | 94.0% | 97.6% | 95.8% | 92.6% | 95.4% | 94.5% | 97.6% | 94.8% | 96.2% | 94.1% | 95.4% |
Au oz Rec (000's) | 1.7 | 7.8 | 2.0 | 5.3 | 5.3 | 6.6 | 12.0 | 8.3 | 7.5 | 5.3 | 2.9 | 64.7 |
Ag oz Rec (000's) | 1.7 | 7.8 | 1.6 | 4.1 | 3.8 | 5.7 | 10.5 | 9.8 | 8.8 | 6.4 | 3.6 | 63.7 |
Note: | ||
1. | Includes only production following the completion of the Midas purchase from Newmont on February 11, 2014. |
17.7. | Midas Mill Operating Permits |
The Midas Mill is currently operating under three Air Quality Operating Permits administered by the Nevada Department of Environmental Protection (NDEP) Bureau of Air Pollution Control and one Water Pollution Control Permit administered by the Nevada NDEP Bureau of Water Pollution Control. The permits are discussed in detail in Section 20.
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18. | Project Infrastructure |
18.1. | Road Access |
The Project is easily accessible from paved state highways and from a graded gravel mine access road. The main access passes through a small residential area for about two miles where the speed limit is reduced to minimize any potential impacts on the community. The gravel road can be occasionally impeded by mud in wet or snowy weather.
The state and county roads are well maintained in order to service the ranches and mines in Crescent Valley. Klondex provides some road maintenance assistance to Lander County.
18.2. | Power and Electrical Infrastructure |
A regional electrical transmission line runs two miles east of the Fire Creek Project. A substation was constructed in 2012 to service the Fire Creek Project. The power line joining the Fire Creek Project to the substation was completed in August 2013, eliminating the need to use generators to supply power for mine operation.
18.3. | Water Management and Water Treatment |
Klondex manages surface and underground water using a pond system, drainage ditches, and a water treatment plant (WTP). Surface water from precipitation events is diverted away from the Project infrastructure with a series of drainage ditches. Surface water within the disturbance areas is diverted to one of two ponds: the Stormwater Pond and the Dewatering Storage Pond. The ponds have a combined volume of approximately 4.7 million gallons (Figure 18-1). Klondex is commissioning two RIBs, which will also be included in the water management system. When completed, the RIBs are expected to have the capacity to infiltrate up to 3,000 gpm of water meeting the Profile I standard. Permit approval was obtained in June 2014.
Water from underground exploration operations that does not meet NDEP Profile I standards (Profile I) is pumped to the Dewatering Storage Pond, which holds a total capacity of approximately 2.8 million gallons. This water is treated through the WTP to meet the Profile I requirement. Brine reject solution from the WTP is stored in the Stormwater Pond, where it is evaporated or shipped off-site for disposal. Treated water from the WTP and water from underground that meets the Profile I standard can be managed in several ways: used for dust suppression on roads and during construction events; infiltrated in the RIBs; or used underground for mining activities.
Klondex has permitted and constructed an artesian well, PW-1, which can provide up to three gpm of fresh water to the Project. Klondex currently holds annual water rights for 283 acre-feet of water. A fire water tank is located above the facilities and gravity flows to hydrants located near the Project buildings.
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18.4. | Communication Infrastructure |
Internet connectivity is provided by WesNet, via 11GHz licensed Microwave frequency, with a 20Mbps Direct Internet Access (DIA) connection. Cell phone coverage is provided by Verizon Wireless, and the signal is boosted by a Klondex provided network extender.18.5.
18.5. | Site Infrastructure |
Project infrastructure is comprised of two large tented structures, heavy equipment parking areas, several mobile office units, several Conex mobile containers, and lay-down areas. The two-tented structures are used for maintenance of the mobile fleet and other production related equipment. The east bay is designated the mechanical shop. The west bay is divided into an area for lubrication and a wash bay. Several Conex containers and outbuildings are used for storing parts and tools near the maintenance buildings. The electric storage area and diesel storage area are also located near the maintenance building Figure 18-1.
The engineering and geology offices, security, and staff dry area are in mobile office units with light vehicle parking areas in front. These buildings are connected to non-potable water pipelines and septic system. The core logging facility is a 1,000 square foot (sf) plasticized-canvas covered outbuilding with overhead lights and propane heater located beside a core lay down area of about 500 sf. The core splitting facility is housed next to the core logging facility in a mobile Conex container.
In addition to the offices, there are areas designated for septic leach field, waste rock dump, WTP, sediment control ditches, and re-vegetated stockpiles.
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19. | Market Studies and Contracts |
19.1. | Precious Metal Markets |
Gold and silver markets are mature with reputable smelters and refiners located throughout the world. Following several years of increases, gold and silver prices began declining in 2012. As of December 2014, the 36-month trailing average gold price was $1,449 per ounce, the 24-month trailing average price was $1,339 while the monthly average had dropped to $1,202. The silver price trend shows similar behavior and both are shown in Figure 19-1.
19.2. | Contracts |
As part of normal mining activities, Klondex has entered into contracts with several mining industry suppliers and contractors. The terms of these agreements are customary for mines in the area.
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19.3. | Project Financing |
On February 11, 2014, the Company entered into the Gold Purchase Agreement with Franco-Nevada GLW Holdings Corp., a subsidiary of FNC, pursuant to which the Company raised proceeds of $33,763,640 in consideration for the delivery of an aggregate of 38,250 ounces of gold on a monthly basis over a five-year period ending on December 31, 2018. Under the terms of the Gold Purchase Agreement, the Company is required to make gold deliveries at the end of each month, with the first delivery due on June 30, 2014. Gold deliveries will cease when the delivery of 38,250 ounces of gold is completed on December 31, 2018. The annualized delivery schedule is shown in Table 19-1.
Table 19-1 FNC Gold Delivery Schedule
Year | Gold Ounces |
2014 | 6,750 |
2015 | 7,500 |
2016 | 8,000 |
2017 | 8,000 |
2018 | 8,000 |
Total | 38,250 |
The Company's obligations under each of the Gold Purchase Arrangement and the Company’s concurrent debt financing (the “2014 Debt Financing”) are secured against all of the assets and property of the Company and its subsidiaries. The security granted for the performance of the Company's obligations under the 2014 Debt Financing and the Gold Purchase Arrangement rank pari-passu.
On February 12, 2014, the Company entered into a royalty agreement (the FC Royalty Agreement) with Franco-Nevada US, a subsidiary of FNC, and KGS, pursuant to which KGS raised proceeds of US$1,018,050 from the grant to Franco-Nevada US of a 2.5% NSR royalty for all Fire Creek production beginning January 1, 2019.
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20. | Environmental Studies, Permitting and Social or Community Impact |
Klondex conducts mineral exploration activities in compliance with all applicable environmental protection legislation. Klondex is unaware of any existing environmental issues or compliance problems that have the potential to impede production at the Fire Creek Project. Klondex is working closely with both state and federal regulators to ensure that the permitting and compliance strategies are acceptable and will not cause delays in production or mine development. Klondex has a strong cultural resource preservation program, which allows a third party archeologist time to review potential areas of new disturbance. At this time, there are no community or social impact issues regarding work being completed at the Project.
20.1. | Environmental Compliance and Monitoring |
As required by the environmental operational permits (see Table 19-1), Klondex prepares quarterly and annual reports which are submitted to regulators. Compliance information included in these reports is based primarily on permit requirements and limitations. Permit limits and associated monitoring requirements are specified as a part of each permit. 20.1.1.
20.1.1 Waste Rock Disposal Facility |
Initial humidity cell test work results indicate that a portion of the waste rock removed from the mine will have the potential to degrade waters of the State of Nevada. As a result of mine expansion, Klondex is relocating the waste rock disposal facility, and the new facility design will include waste rock staging to ensure that any material that has potential to degrade water of the state is segregated from non-acid generating materials and engineered to alleviate any acid drainage problem.
20.1.2. Other Environmental Issues |
At this time, Klondex does not anticipate construction or operation of any processing facilities. Heap leaching, tailings management, or other processing components are not included as part of the permitting strategy and not part of the resource.
20.2. | Reclamation Bond Estimate |
Klondex’s last amendment to the Reclamation Bond Estimate (RCE) to include construction and operation of the RIBs was received in February 2014. The total of the RCE is calculated using the Standard Reclamation Cost Estimator (SRCE), which is adjusted annually for inflation. The SRCE was developed in a cooperative effort between the NDEP, Bureau of Mining Regulation and Reclamation, BLM, and the Nevada Mining Association to facilitate accuracy, completeness, and consistency in the calculation of costs for mine site reclamation. Klondex is required to update the total RCE for Fire Creek every three years. The next RCE update is scheduled for 2016.
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RCE costs for reclamation currently include the following categories: roads; exploration roads and drill pads; waste rock repository; RIBs; ponds; electrical infrastructure; building and equipment; adit and vent raise plugging; re-vegetation; and contractor management. The total RCE was approved by BLM and NDEP in the first quarter of 2014 for a total cost to construct of approximately $1.7 million dollars.
Klondex also maintains a second statewide bond for archeological conservation. This bond, totaling $119,284, was implemented to facilitate a treatment plan as required by Nevada State Historic Preservation Office (SHPO) pursuant to 36CFR 800, regulations implementing Section 106 of the National Historic Preservation ACT (NHPA), 16 U.S.C. § 470f. The treatment plan will include, but not be limited to, mitigation of archaeological sites, artifact processing, writing final report of findings, and curation of artifacts.
20.3. | Major Permitting and Approvals |
The major operational permits and a brief summary of the requirement for each permit are outlined in Table 19-1 below.
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Table 20-1 Fire Creek Project Significant Permits
Permit | Permit Number | Agency | Permit Type and Explanation |
Environmental Assessment and Plan of Operations | NVN-079769 | BLM | Plan of Operations is required for all mining and processing activities and exploration exceeding 5 acres of disturbance. BLM approves plan and determines the required environmental studies, usually an environmental assessment or an environmental impact study based on the requirements outline in the National Environmental Policy Act. |
Record of Decision | BLM | A Record of Decision (ROD) in the United States is the formal decision document which is recorded for the public. | |
Water Pollution Control Permit (Operations) | NEV2007104 | NDEP, BMRR | Mines operating in the State of Nevada are generally required to meet a zero discharge performance standard. A WPCP is required for the extraction of mineralized material. A separate permit may be issued for certain activities at a specific facility, such as rapid infiltration. |
Water Pollution Control Permit (Infiltration) | NEV2013102 | NDEP, BMRR | Water Pollution Control Permit for infiltration of water from the underground mine operations. This permit is still in the approval process. |
Water Rights | 28637, 77002, 77003, 75129 | NDWR | Water rights are issued by the Nevada Division of Water Resources based on Nevada water law which issues permits based on prior appropriation and beneficial use. Prior appropriation (also known as "first in time, first in right") allows for the orderly use of the state's water resources by granting priority to parties with senior water rights. This concept ensures the senior uses are protected, even as new uses for water are allocated. |
Reclamation Permit | #0241 | NDEP, BMRR | Summarizes reclamation activities and associated costs. Ensures land disturbed by mining activities are reclaimed to safe and stable conditions to promote safe and stable post-mining land use. A permit is required for any disturbance over 5 acres. The RCE is financially secured with a posted security. The posted surety amount provides assurance that reclamation will be pursuant to the approved reclamation plan. |
Air Quality Permit | AP1041-2774 | NDEP, BAPC | An owner or operator of any proposed stationary source must submit an application for and obtain an appropriate operating permit before commencing construction or operation. Class II Air Permit - Typically for facilities that emit less than 100 tons per year for any one regulated pollutant and emit less than 25 tons per year total HAP and emit less than 10 tons per year of any one HAP. |
Storm Water Permit | NVR300000 | NDEP, BWPC | General storm water discharges associated with activities from metal mining activities. Regulates storm water runoff from waste rock storage piles, roads, and cleared areas. Typical pollutants include suspended solids and minerals eroded from exposed surfaces. |
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20.4. | Future Permitting |
Klondex has an approved plan of operations with the BLM covering their current exploration activities at the Project. Klondex received approval to an amendment to the Plan of Operations and Reclamation permits (NVN-07976 and Reclamation Permit 0028) which allows Klondex to construct and operate several RIBs. Klondex received approval for the Water Pollution Control Permit for the RIBs (WPCP2013102), issued in February 2014. In addition, Klondex has initiated a baseline data collection program to ensure that enough data is collected to be sufficient for additional permitting necessities.
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21. | Capital and Operating Costs |
21.1. | Capital Costs |
Life of Mine (LOM) constant dollar capital expenditures are detailed in Table 21-1. Project development comprises over 61% of total capital requirements; mine equipment 28%; site facilities six percent, and environmental projects five percent. Owner operated mine development unit costs, for similarly sized excavations in north Nevada, are shown in Table 21-2. The remainder of the capital costs are from Klondex’s 2014 capital budget.
Table 21-1 Capital Costs
Cost (000's) | |||||
2015 | 2016 | 2017 | 2018 | Total | |
Mine Development | $7,059 | $4,236 | $11,295 | ||
Rapid Infiltration Basin | $368 | $368 | |||
Site Facilities | $1,127 | $1,127 | |||
Environmental Assessment | $469 | $469 | |||
Mining Equipment | $2,200 | $2,000 | $932 | $5,132 | |
Total | $11,223 | $6,236 | $932 | $0 | $18,391 |
Table 21-2 Underground Development Unit Costs
Unit | |||
Width | Height | Cost | |
Description | (ft) | (ft) | ($/ft) |
Primary Capital Drifting | 14 - 15 | 15 - 17 | $1,350 |
Secondary Capital Drifting | 14 | 14 | $1,350 |
Raising | 10 | 10 | $2,000 |
21.2. | Operating Costs and Cutoff Grade |
LOM operating costs are presented in Table 21-3 below. Unit mining costs are based on actual costs incurred at Fire Creek in 2014. These costs have been adjusted to the planned mining rate where appropriate. The weighted average cost is based on the LOM quantities in each category. Haulage costs to Midas are based on actual costs incurred by the Company and paid to a local contractor during 2014.
Table 21-3 Operating Costs
Description | Unit Cost | Unit |
Mining | ||
Production Stoping | $140.00 | /ton |
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Description | Unit Cost | Unit | |
5 x 11 Stope Development Drift | $215.00 | /ton | |
Backfill | $30.00 | /ton | |
Cellular Backfill | $235.00 | ||
Average Mining Cost | $218.36 | /ton | |
Transportation, Processing and G&A | |||
Haulage Fire Creek to Midas | $32.66 | /ton | |
Processing - Fixed Cost | $16,733 / Process Rate (tpd) | $66.92 | /ton |
- Variable Cost | $26.18 | /ton | |
Site Administration and G&A | $15 900 / Mineral Reserves Mining Rate (tpd) | $115.49 | /ton |
Total | $459.61 | /ton |
Processing costs include fixed and variable components. Appling these to the 2014 actual tonnage processed predicts a total cost of $61.24 per ton. Actual costs for 2014 averaged $57.49 per ton or six percent below the predicted cost. The processing rate used for the cost estimate and cash flow estimate is based only on the reserves mine plans for the Fire Creek Project and the Midas Mine and does not include any toll milling or other sources of plant feed.
Site administration costs are based on actual Fire Creek cost reporting for the later part of 2014. These costs include surface support, environmental, land, legal and other costs allocated to the Project. These costs are treated as 100% fixed and amount to $477,000 per month.
Using the operating costs and parameters above, cut-off grades were calculated at varying gold prices. These are shown in Table 21-4 and Figure 21-1 Cutoff Grade Sensitivity to Gold Price. The incremental cut-off represents the required minimum grade of mineralization to be profitable to process after it has been mined and transported to the surface. Mineralization from development excavations is included in the LOM plan if it exceeds the incremental cut off since processing the incremental material improves the Project cash flow over the alternative of sending this material to the waste dump.
Table 21-4 Cut-off Grade Calculation
Gold | Silver | ||
Metal Sales Price | $/Ounce | $1,000 | $15.83 |
Refining and Sales Expense | $/Ounce | Included in Milling | |
Royalty | 1% | ||
Metallurgical Recovery | 94% | 92% | |
Operating Costs | |||
Ore Haulage (Portal to Mill) | $/ton | $32.66 |
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Direct Processing | $/ton | $93.10 | |
Administration and Overhead | $/ton | 115.49 | |
Mining | $/ton | $218.37 | |
Total | $/ton | $459.62 | |
Gold Equivalent | 1 | 64.54 | |
Unplanned Dilution | 10% | ||
Incremental Cut Off Grade | 0.259 | ||
Cut-off Grade | Eq. opt | 0.494 | |
Minimum Mining Width | feet | 4 | |
Grade Thickness cut-off | Eq. opt-ft. | 2.173 |
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22. | Economic Analysis |
The LOM plan and technical and economic projections in the LOM plan model include forward looking statements that are not historical facts and are required in accordance with the reporting requirements of the Canadian Securities Administrators. These forward-looking statements are estimates and involve risks and uncertainties that could cause actual results to differ materially.
The estimates of capital and operating costs have been developed specifically for the Project and are summarized in Section 21. These costs are derived from actual mine and process operating experience for the Project during 2014 and where appropriate include adjustments applicable to the planned production rates.
The cash flow estimate includes only costs, taxes and other factors applicable to the project and corporate obligations, financing costs, and taxes are excluded. The cash flow estimate includes 35% Federal income tax after appropriate deductions for depreciation and depletion. No consideration has been given for carry forward losses incurred prior to 2015. Nevada does not impose an income tax but does levy a net proceeds tax equal to 5% of the net operating income with some allowances for depreciation of property plant and equipment. The net proceeds tax does not allow a depletion deduction.
Future reclamation costs have been prepaid through reclamation bonding requirements of the BLM and NDEP. These bonds are considered adequate to fund future reclamation liabilities.
22.1. | Life of Mine Plan and Economics |
Constant dollar cash flow analysis of the reserves production and development plan shown in Table 16-4 is presented in the income and cash flow statements of Table 22-1 and Table 22-2, respectively. Table 22-3 lists the life of mine key operating and financial indicators. The grade of the Fire Creek resources and the low capital requirements facilitated with the addition of the Midas Mine and Mill to Klondex’s project portfolio combine to produce a short 0.5 -year capital payback period and an impressive 5.0 profitability index (PI) calculated with a 10% discount rate. PI is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates break even. Calculation of an internal rate of return (IRR) is indeterminate due to the positive cash flow projected to be achieved in each year of the project. Royalties incurred during the LOM plan include the advance minimum royalty payments to third party lessors and the one percent royalty to Waterton under the Gold Supply Agreement as discussed in Section 4.3. The mine plan ends prior to the 2 ½% royalty taking effect as specified in the FC Royalty Agreement with Franco-Nevada US. None of the planned production is from holdings subject to NSR royalties nor will it transit through holdings subject to wheelage royalties.
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Table 22-1 Income Statement 2015 – 2018 ($000’s)
Year | 2015 | 2016 | 2017 | 2018 | Total |
Income Statement (000's) | |||||
Revenue | |||||
Gold Sales | $45,455.3 | $75,625.5 | $54,964.6 | $48,362.3 | $224,407.7 |
Silver Sales | $477.4 | $857.4 | $629.3 | $543.1 | $2,507.3 |
Total Revenue | $45,932.7 | $76,482.9 | $55,593.9 | $48,905.4 | $226,915.0 |
Operating Costs | |||||
Ore Mining | ($10,304.3) | ($11,044.8) | ($5,414.8) | ($3,910.3) | ($30,674.2) |
Backfill | ($1,619.3) | ($3,919.9) | ($1,287.8) | ($476.2) | ($7,303.1) |
Expensed Waste | ($1,522.0) | ($1,064.0) | $0.0 | $0.0 | ($2,586.1) |
Surface Ore Haulage Portal to Mill | ($1,723.9) | ($2,167.5) | ($1,263.2) | ($912.2) | ($6,066.8) |
Processing | ($3,329.0) | ($4,898.3) | ($3,767.0) | ($5,299.3) | ($17,293.6) |
Site General Administration & Overhead | ($5,724.0) | ($5,724.0) | ($5,724.0) | ($4,281.2) | ($21,453.2) |
Total Operating | ($24,222.4) | ($28,818.6) | ($17,456.7) | ($14,879.3) | ($85,377.0) |
General & Administrative | |||||
Refining & Sales (Included with Processing Costs) | $0.0 | $0.0 | $0.0 | $0.0 | $0.0 |
Royalty | ($459.3) | ($764.8) | ($555.9) | ($489.1) | ($2,269.1) |
Nevada Net Proceeds Tax | ($1,034.5) | ($2,301.3) | ($1,833.1) | ($1,630.9) | ($6,799.8) |
Total Cash Cost | ($25,716.3) | ($31,884.7) | ($19,845.8) | ($16,999.2) | ($94,446.0) |
EBITA | $20,216.4 | $44,598.2 | $35,748.2 | $31,906.2 | $132,469.0 |
Reclamation Accrual (UOP) | $0.0 | $0.0 | $0.0 | $0.0 | $0.0 |
Depreciation | ($2,273.2) | ($6,417.3) | ($5,160.1) | ($4,540.2) | ($18,390.8) |
Total Cost | ($27,989.5) | ($38,302.0) | ($25,005.8) | ($21,539.5) | ($112,836.8) |
Pre-Tax Income | $17,943.2 | $38,180.9 | $30,588.1 | $27,366.0 | $114,078.2 |
Income Tax | ($3,917.8) | ($9,433.1) | ($7,849.4) | ($7,064.7) | ($28,265.1) |
Net Income | $14,025.4 | $28,747.8 | $22,738.7 | $20,301.2 | $85,813.1 |
Table 22-2 Cash Flow Statement 2015 – 2019 ($000’s)
Year | 2015 | 2016 | 2017 | 2018 | 2019 | Total |
Net Income | $14,025.4 | $28,747.8 | $22,738.7 | $20,301.2 | $0.0 | $85,813.1 |
Depreciation | $2,273.2 | $6,417.3 | $5,160.1 | $4,540.2 | $0.0 | $18,390.8 |
Reclamation | $0.0 | $0.0 | $0.0 | $0.0 | $0.0 | $0.0 |
Working Capital (6 weeks) | ($2,967.3) | ($711.7) | $1,389.1 | $328.4 | $1,961.4 | $0.0 |
Operating Cash Flow | $13,331.4 | $34,453.3 | $29,287.9 | $25,169.9 | $1,961.4 | $104,203.9 |
Capital Costs |
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Year | 2015 | 2016 | 2017 | 2018 | 2019 | Total |
MACRS Equipment | ($2,200.0) | ($2,000.0) | ($932.3) | $0.0 | $0.0 | ($5,132.3) |
Capitalized Development | ($7,059.3) | ($4,235.8) | $0.0 | $0.0 | $0.0 | ($11,295.1) |
Mine Capital | ($1,963.5) | $0.0 | $0.0 | $0.0 | $0.0 | ($1,963.5) |
Total Capital | ($11,222.8) | ($6,235.8) | ($932.3) | $0.0 | $0.0 | ($18,390.8) |
Net Cash Flow | $2,108.6 | $28,217.6 | $28,355.6 | $25,169.9 | $1,961.4 | $85,813.1 |
Cumulative Cash Flow | $2,108.6 | $30,326.2 | $58,681.7 | $83,851.7 | $85,813.1 |
Table 22-3 Key Operating and After Tax Financial Statistics
Material Mined and Processed (kt) | 186 |
Avg. Gold Grade (opt) | 1.327 |
Avg. Silver Grade (opt) | 0.96 |
Contained Gold (koz) | 237.0 |
Contained Silver (koz) | 171 |
Avg. Gold Metallurgical Recovery | 94% |
Avg. Silver Metallurgical Recovery | 92% |
Recovered Gold (koz) | 224.4 |
Recovered Silver (koz) | 158 |
Reserve Life (years) | 3.8 |
Operating Cost ($/ton) | $460 |
Cash Cost ($/oz) 1. | $410 |
Total Cost ($/oz) 1. | $492 |
Gold Price ($/oz) | $1,000.00 |
Silver Price ($/oz) | $15.83 |
Capital Costs ($ Millions) | $18.4 |
Payback Period (Years) | 0.5 |
Cash Flow ($ Millions) | $85.80 |
5% Discounted Cash Flow ($ Millions) | $78.10 |
10% Discounted Cash Flow ($ Millions) | $71.40 |
Profitability Index (10%) 2. | 5.0 |
Internal Rate of Return | NA |
Notes: | ||
1. | Net of byproduct credits; | |
2. | Profitability index (PI) is the ratio of payoff to investment of a proposed project. It is useful for ranking project as a measure of the amount of value created per unit of investment. A PI of 1 indicates break even. |
22.2. | Sensitivity Analysis |
The Project’s net present value at five percent and 10% (NPV) and profitability index from the cash flow model presented above were analysed for sensitivity to variations in revenue, operating and capital cost assumptions. This analysis is presented graphically in Figure 22-1 through Figure 22-3 below. These graphs demonstrate the economic resilience of the Project by maintaining profitability with up to 40% unfavorable variances of any one of the three categories.
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22.3. | Adjusted Plan at $1,200 Gold |
In addition to the price sensitivities discussed above, an additional case was evaluated which included adjustment of the LOM to reflect the increase in gold price to $1,200 per ounce. Contained gold ounces in the adjusted plan increased by 13,000 ounces and undiscounted cash flow increased by $30.6M. The results of this additional sensitivity are summarized in Table 22-4.
Table 22-4 Key Operating and Financial Statistics for $1,200 Plan
Material Mined and Processed (kt) | 224 |
Avg. Gold Grade (opt) | 1.12 |
Avg. Silver Grade (opt) | 0.80 |
Contained Gold (koz) | 250 |
Contained Silver (koz) | 179 |
Avg. Gold Metallurgical Recovery | 94% |
Avg. Silver Metallurgical Recovery | 92% |
Recovered Gold (koz) | 235 |
Recovered Silver (koz) | 165 |
Reserve Life (years) | 3.8 |
Operating Cost ($/ton) | $416 |
Cash Cost ($/oz) 1. | $435 |
Total Cost ($/oz) 1. | $535 |
Gold Price ($/oz) | $1,200.00 |
Silver Price ($/oz) | $19.00 |
Capital Costs ($ Millions) | $18.4 |
Payback Period (Years) | 0.5 |
Cash Flow ($ Millions) | $116.4 |
5% Discounted Cash Flow ($ Millions) | $105.7 |
10% Discounted Cash Flow ($ Millions) | $96.6 |
Profitability Index (10%) 2. | 5.3 |
Internal Rate of Return | NA |
Notes: | ||
1. | Net of byproduct credits; | |
2. | Profitability index (PI) is the ratio of payoff to investment of a proposed project. It is useful for ranking project as a measure of the amount of value created per unit of investment. A PI of 1 indicates break even. |
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23. | Other Relevant Data and Information |
The authors are not aware of any other relevant data and information having bearing on the Fire Creek mineral resource estimate or mineral reserve estimate or ongoing exploration or operations.
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24. | Interpretation and Conclusions |
24.1. | Conclusions |
The Project is a modern, mechanized narrow vein mine. Only the mineralized veins accessible from main development have been defined to a sufficient level of detail to declare reserves. Additional potential exists to extend reserves along strike in both directions as underground access is developed. As the footprint of the mine grows and the number of available mining areas grows with it, it is expected that the mining rate can be increased and that cost reductions can be realized through economies of scale.
The conventional Merrill Crowe mill facility at Midas is an efficient, well-maintained modern mineral processing plant capable of processing 1,200 tpd. The plant is capable of operating with a minimum crew compliment resulting in cost reductions when operated at capacity. The underutilized capacity can accept increased mine production from the Fire Creek Project or the Midas Mine as well as third party processing agreements.
Capital requirements for the Project are minimal. Ongoing mine development comprises the majority of capital costs and the ability to access multiple veins from common development greatly reduces the unit cost per ounce.
Based on the assumptions described herein, and in the opinion of the authors of this Technical Report, the high grade reserves in the Project mine plan are expected to provide a high return and sustain profitable operations with up to 40% adverse variations in metal prices, operating or capital costs. The total cost per ounce including capital expenditures and net of byproduct sales is expected to be less than $500 per ounce.
24.2. | Project Risks |
Table 24-1 presents the significant risks identified by the Qualified Person that have potential to impact the Fire Creek Project.
Table 24-1 Potential Project Risks
Risk | Potential Impact | Mitigating Measures | Opportunities |
Mine and/or mill operating costs greater than planned | Lower cash flow | Convert Inferred Mineral Resource to Measured or Indicated Mineral Resources near planned mining areas | Additional work areas allow an increase in production rate and achieves economies of scale |
Stope dilution greater than anticipated | Production cost increase and loss of resource | Employ alternative mining methods and/or increase cutoff grade | Dilution can contain mineralization and could aid in obtaining economies of scale. |
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25. | Recommendations |
1. | Exploration: Underground drilling should continue in the veins identified near the current development workings to increase the level of confidence in these veins to an indicated classification. The decline should be advanced to provide an underground drill platform from which to drill the veins in the North and Far North Zones. While the decline is being advanced, additional drilling in this area can be completed from surface to refine the vein targets. | |
2. | Definition Drilling: Rib sampling has limited value and should continue to be supplemented by drilling shallow ten to 20-foot deep holes into the rib with the “Termite” drill or hand held drills and sample the drill cuttings. This sampling method will add a third dimension to the potential wall rock mineralization. These costs are included in mine operation costs estimates. | |
3. | Stope Planning: Compete the drift and fill stopes currently underway, and new areas should be set up for long hole stoping. The use of short probe holes discussed above should provide the planning engineers enough detail to efficiently design stopes with minimal loss of mineralization. | |
4. | Rapid Infiltration Basin Commissioning: In order to reduce delays caused by intercepting perched water, the RIB’s and water handling systems need to be functioning at capacity. | |
5. | Geologic Database Administration: All of the Project data collected to date including drill samples, channel samples and QA/QC samples need to be stored and archived in a permanent and indelible manner. The system software for this system has been procured, but a full time data base administrator has not been selected. | |
6. | QAQC: Timely follow-up for QAQC assay deviations and re-assay requests needs to be aggressively pursued. This should become an automated process once the database is up and running. |
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Table 25-1 Recommendation Estimated Costs
Description | Estimated Costs (000's) |
Phase I | |
Rapid infiltration Basin Commissioning | $368 |
Geologic Database Administration | $50 |
Definition Drilling | Definition and planning |
Stope Planning | costs are included in the |
project operating costs. | |
Phase II | |
Exploration | $11,250 |
QA/QC | Enhanced QA/QC costs |
are included in the | |
sample assaying costs for | |
exploration |
Note: | ||
1. | Phase II recommendations are not contingent on Phase I recommendations and could occur concurrently with Phase I recommendations. |
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26. | Bibliography |
Anderson, R., 2013, Stratigraphy of the Fire Creek low sulfidation Au deposit Preliminary Report: Klondex Gold and Silver Mining Company internal document, 39 p.
Canadian Institute of Mining, Metallurgy and Petroleum, May 10, 2014, “CIM Definition Standards – For Mineral Resources and Mineral Reserves”, 9 p.
Colgan, J., Heny, C. & John, D., 2014, Evidence for large-magnitude, post-Eocene extension in the northern Shoshone Rane, Nevada, and its implications for the structural setting of Carlin-Type gold deposits in the lower plate of the Roberts Mountains allochthon: Economic Geology, v. 109, p. 1843-1862.
Cooke, D. & Simmons, S., 2000, Characteristics and genesis of epithermal gold deposits: Society of Economic Geologists Reviews, v. 13, p. 221-244.
Crider J., 2001, Oblique slip and the geometry of normal-fault linkage: mechanics and a case study from the Basin and Range in Oregon: Journal of Structural Geology, v. 23, p. 1997-2009.
Crowl, W. J. (2011, May 31). NI 43-101 Technical Report, Pinson Project, Humboldt County, Nevada. Edmondo, G., 1996, Fire Creek Project: North Mining, Inc. internal report, 30 p.
Erwin, T. P. (2013, November 27). Mineral Status Report for Klondex Gold and Silver Mining Company - Project King, File NO. 52591.004.
Gilluly, J. & Gates, O., 1965, Tectoic and igneous geology of the northern Shoshone Range, Nevada: Geological Survey Professional Paper 465, 153 p.
Graf, G. (2013, January 13). Midas 2011 - 2012 Surface Exploration Report. Newmont Internal Memorandum. Hedenquist, J., Arribas, A. & Gonzalez-Urien, E., 2000, Exploration for epithermal gold deposits: Society of Economic Geologists Reviews, v. 13, p. 245-277.
Henry, C., 2013, email to R. Anderson.
Hodenquist, J.W., and Lowenstern,J.B., “The Role of Magmas in the Formation of Hydrothermal Ore Deposits”Nature, v 370, p 519-527.
John, D., 2014, discussion with J. Milliard
John, D. A. (2003). Geologic Setting and Genesis of the Mule Canyon Low-Sulfidation Epithermal Gold-Silver Deposit, North-Central Nevada. Economic Geology, 98, 424-463.
John, D. & Wallace, A., 2000, Epithermal gold-silver deposits related to the Northern Nevada Rift, in: Cluer, J., Price, J., Struhsacker, E., Hardyman, R. & Morris, C., eds., Geology and Ore Deposits 2000: The Great Basin and Beyond: Geological Society of Nevada Symposium Proceedings, May 15-18, 2000, p. 155-175.
John, D., Brunner, J., Saderholm, E. & Fleck, R., 2000a, Geology of the Mule Canyon gold-silver deposit, Lander County, Nevada, in: Cluer, J., Price, J., Struhsacker, E., Hardyman, R. & Morris, C., eds., Geology and Ore Deposits 2000: The Great Basin and Beyond: Geological Society of Nevada Symposium Proceedings, May 15-18, 2000, p. 119-134.
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John, D., Wallace, A., Ponce, D., Fleck, R. & Conrad, J., 2000b, New perspectives on the geology and origin of the Northern Nevada Rift, in: Cluer, J., Price, J., Struhsacker, E., Hardyman, R. & Morris, C., eds., Geology and Ore Deposits 2000: The Great Basin and Beyond: Geological Society of Nevada Symposium Proceedings, May 15-18, 2000, p. 127-154.
John, D. & Wrucke, C., 2003, Geologic map of the Mule Canyon Quadrangle, Lander County, Nevada, Nevada Bureau of Mines and Geology Map 144.
Kamenov, G., Saunders, J., Hames, W. & Unger, D., 2007, Mafic magmas as sources for gold in middle Miocene epithermal deposits of the northern Great Basin, United States: Evidence from Pb isotope compositions of native gold: Economic Geology, v. 102, n. 7, p. 1191-1195.
Kassos, G. & Marma, J., in prep., Fire Creek: Nevada’s next high-grade gold project: Geological Society of Nevada 2015 Symposium Program with Abstracts.
Kiska Metals Corporation, 2014, The Colorback and Hilltop Properties: Carlin-style systems in the Battle Mountain-Eureka Trend, Nevada: Executive Summary Report, 19 p.
Klondex Mines Ltd. (2013, December 4). Final Disclosure Schedules to Stock Purchase Agreement.
Leavitt, E. D., Spell, T. L., Goldstrand, P. M., & Arehart, G. B. (2004, December 1). Geochronology of the Midas Low-Sulfidation Epithermal Gold-Silver Deposit, Elko County, Nevada. Econoomic Geology, 99(8), 1665-1686.
Martini, Josepph, SRK Consulting (2014, February). Midas Mine and Mill Reclamation Cost Adequacy, Report for Klondex Mines Ltd.
McPhie, J., Doyle, M. & Allen, R., 1993, Volcanic Textures: A guide to the interpretation of textures in volcanic rocks: University of Tasmania Center for Ore Deposit and Exploration Studies, 196 p.
McMillin, S. & Milliard, J., 2013, Exploration and geology of the Fire Creek deposit, Lander County, Nevada, presented at the November, 2013 Geological Society of Nevada Elko/Winnemucca joint meeting.
Milliard, J., Marma, J. & Kassos, G., in prep., A field trip guide for the Fire Creek Deposit - Nevada’s new high-grade gold project, in: 2015 Geological Society of Nevada Symposium Pre-Meeting Field Trip: Epithermal deposits of northern Nevada.
Newmont Mining Corporation. (2010). Internal Test Parameters Memorandum.
Newmont Mining Corporation. (2013, December). http://www.newmont.com/our-investors/reserves-and-resources.
Odell, M. A., Symmes, L., Bull, S., and Swanson, K., July 24, 2014, “Preliminary Economic Assessment of the Fire Creek Project, Lander County, Nevada, Amended”, NI 43-101 Technical Report, 218 p.
Odell, M. A. (2013). NI 43-101 Technical Report, Fire Creek Exploration Project, Lander County, Nevada. NI 43-101 Technical Report.
Pierce, K. & Morgan, L., 1992, The track of the Yellowstone hotspot: Volcanism, faulting and uplift, in: Link, P., Kuntz, M. & Platt, L., eds., Regional geology of eastern Idaho and western Wyoming: Geological Society of America Memoir 179, p. 1-53.
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Ponce, D. A. (2008, February). A Prominent Geophysical Feature Along the Northern Nevada Rift and its Geologic Implications, North-Central Nevada. Geosphere, 4(1), 207-217.
Postlethwaite, C. (2011, December 19). Progress Report of The 20011 Midas District Mapping and Structural Analysis. Newmont Internal Report.
Raven, W., Ullmer, E. & Hawthorn, G., 2011, Updated technical report and resource estimation on the Fire Creek gold property, Lander Co., Nevada: NI 43-101 technical report filed on SEDAR Sept. 12, 2011.
Rott, E. H. (1931). Ore Deposits of the Gold Cirlce Minng District, Elko County, Nevada. Bulletin of the Nevada Bureau of Mines and Mackay School of Mines.
Saunders, J. A. (2006). Geochronology of Volcanic-Hosted Low-Sulfidation Au-Ag Deposits, Winnemucca-Sleeper Mine Area, Northern Great Basin, USA. US Geological Survey.
Simmons, S., White, N. & John, D., 2005, Geological characteristics of epithermal precious and base metal deposits: Economic Geology 100th Anniversary Volume, p. 485-522.
Theodore, T., Armstrong, A., Harris, A., Stevens, C. & Tosdal, R., 1998, Geology of the terminus of the northern Carlin Trend, in: Tosdal, R., ed., 1998, Contributions to the gold metallogeny of northern Nevada: United States Geological Survey Open-File Report 98-338, p. 69-105.
Thompson, T., 2014, Mineralogy of the MLI3870 composites, Fire Creek, Nevada: McClelland Laboratories, Inc. internal report, 44 p.
Trudgill, B. & Cartwright, J., 1994, Relay-ramp forms and normal-fault linkages, Canyonlands national Park, Utah: Geological Society of America Bulletin, v. 106, p. 1143-1157.
US Department of the Interior (DOI) Bureau of Land Managment (BLM). (2013, March). Midas Underground Support Facilities Newmont Mining Corporation, Environmental Assesment.
Wallace, A. & John, D., 1998, New Studies of Tertiary volcanic rocks and mineral deposits, Northern Nevada Rift, in: Tosdal, R., ed., 1998, Contributions to the gold metallogeny of northern Nevada: United States Geological Survey Open-File Report 98-338, p. 264-278.
Watt, J. T., Glen, J. M., John, D. A., & Ponce, D. A. (2007, December). Three-dimensional Geologic Model of the Northern Nevada Rift and the Beowawe Geothermal System, North-Central Nevada. Geosphere, 3(6), 667-682.
White, N. & Hedenquist, J., 1995, Epithermal gold deposits: Styles, characteristics and exploration: Society of Economic Geologists Newsletter, n. 23.
Zoback, M. & Thompson, G., 1978, Basin and Range rifting in northern Nevada: Clues from a mid-Miocene rift and its subsequent offsets: Geology, v. 6, p. 111-116.
Zoback, M., McKee, E., Blakely, R. & Thompson, G., 1994, The northern Nevada rift: Regional tectono-magmatic relations and middle Miocene stress directions: Geological Society of America Bulletin, v. 106, p. 371-382.
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Zoback, M., Anderson, R. & Thompson, G., 1981, Cainozoic evolution of the state of stress and style of tectonism of the Basin and Range Province of the western United States: Philosophical Transactions of the Royal Society of London, v. 300, p. 407-434.
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27. | Glossary |
Assay: The chemical analysis of mineral samples to determine the metal content.
Asbuilt: (plural asbuilts), a field survey, construction drawing, 3D model, or other descriptive representation of an engineered design for underground workings.
Composite: Combining more than one sample result to give an average result over a larger distance.
Concentrate: A metal-rich product resulting from a mineral enrichment process such as gravity concentration or flotation, in which most of the desired mineral has been separated from the waste material in the ore.
Crushing: Initial process of reducing material size to render it more amenable for further processing.
Cut-off Grade (CoG): The grade of mineralized rock, which determines as to whether or not it is economic to recover its gold content by further concentration.
Dilution: Waste, which is unavoidably mined with ore.
Dip: Angle of inclination of a geological feature/rock from the horizontal.
Fault: The surface of a fracture along which movement has occurred.
Footwall: The underlying side of a mineralized body or stope.
Gangue: Non-valuable components of the ore.
Grade: The measure of concentration of valuable minerals within mineralized rock.
Hanging wall: The overlying side of a mineralized body or stope.
Haulage: A horizontal underground excavation which is used to transport mined rock.
Igneous: Primary crystalline rock formed by the solidification of magma.
Kriging: A weighted, moving average interpolation method in which the set of weights assigned to samples minimizes the estimation variance.
Level: A main underground roadway or passage driven along a level course to afford access to stopes or workings and to provide ventilation and a haulage way for the removal of broken rock.
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Lithological: Geological description pertaining to different rock types.
Milling: A general term used to describe the process in which the ore is crushed, ground and subjected to physical or chemical treatment to extract the valuable minerals in a concentrate or finished product.
Mineral/Mining Lease: A lease area for which mineral rights are held.
Mining Assets: The Material Properties and Significant Exploration Properties.
Sedimentary: Pertaining to rocks formed by the accumulation of sediments, formed by the erosion of other rocks.
Sill1: A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by the injection of magma into planar zones of weakness.
Sill2: The floor of a mine passage way.
Stope: An underground excavation from which ore has been removed.
Stratigraphy: The study of stratified rocks in terms of time and space.
Strike: Direction of line formed by the intersection of strata surfaces with the horizontal plane, always perpendicular to the dip direction.
Sulfide: A sulfur bearing mineral.
Tailings: Finely ground waste rock from which valuable minerals or metals have been extracted.
Thickening: The process of concentrating solid particles in suspension.
Total Expenditure: All expenditures including those of an operating and capital nature.
Variogram: A plot of the variance of paired sample measurements as a function of distance and/or direction.
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Mineral Resources
Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.
A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction.
The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.
Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.
The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently be defined by the consideration and application of Modifying Factors. The phrase ‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. The Qualified Person should consider and clearly state the basis for determining that the material has reasonable prospects for eventual economic extraction. Assumptions should include estimates of cutoff grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments, commodity price or product value, mining and processing method and mining, processing and general and administrative costs. The Qualified Person should state if the assessment is based on any direct evidence and testing.
Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50 years. However, for many gold deposits, application of the concept would normally be restricted to perhaps 10 to 15 years, and frequently to much shorter periods of time.
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Inferred Mineral Resource
An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity.
An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.
An Inferred Mineral Resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred Mineral Resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed Pre-Feasibility or Feasibility Studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred Mineral Resources can only be used in economic studies as provided under NI 43-101.
There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated Mineral Resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an Indicated or Measured Mineral Resource. Under these circumstances, it may be reasonable for the Qualified Person to report an Inferred Mineral Resource if the Qualified Person has taken steps to verify the information meets the requirements of an Inferred Mineral Resource
Indicated Mineral Resource
An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit.
Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing and is sufficient to assume geological and grade or quality continuity between points of observation.
An Indicated Mineral Resource has a lower level of confidence than that applying to a Measured Mineral Resource and may only be converted to a Probable Mineral Reserve.
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Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Pre-Feasibility Study which can serve as the basis for major development decisions.
Measured Mineral Resource
A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit.
Geological evidence is derived from detailed and reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation.
A Measured Mineral Resource has a higher level of confidence than that applying to either an Indicated Mineral Resource or an Inferred Mineral Resource. It may be converted to a Proven Mineral Reserve or to a Probable Mineral Reserve.
Mineralization or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade or quality of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability of the deposit. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.
‘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.
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 Pre-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.
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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 Pre-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, 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, 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
A Probable Mineral Reserve is the economically mineable part of an Indicated, 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 Pre-Feasibility Study.
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Proven Mineral Reserve (Proved 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 Pre-Feasibility Study. Within the CIM Definition standards the term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.
Pre-Feasibility Study (Preliminary Feasibility Study)
The CIM Definition Standards requires the completion of a Pre-Feasibility Study as the minimum prerequisite for the conversion of Mineral Resources to Mineral Reserves.
A Pre-Feasibility Study is a comprehensive study of a range of options for the technical and economic viability of a mineral project that has advanced to a stage where a preferred mining method, in the case of underground mining, or the pit configuration, in the case of an open pit, is established and an effective method of mineral processing is determined. It includes a financial analysis based on reasonable assumptions on the Modifying Factors and the evaluation of any other relevant factors which are sufficient for a Qualified Person, acting reasonably, to determine if all or part of the Mineral Resource may be converted to a Mineral Reserve at the time of reporting. A Pre-Feasibility Study is at a lower confidence level than a Feasibility Study.
Feasibility Study
A Feasibility Study is a comprehensive technical and economic study of the selected development option for a mineral project that includes appropriately detailed assessments of applicable Modifying Factors together with any other relevant operational factors and detailed financial analysis that are necessary to demonstrate, at the time of reporting, that extraction is reasonably justified (economically mineable). The results of the study may reasonably serve as the basis for a final decision by a proponent or financial institution to proceed with, or finance, the development of the project. The confidence level of the study will be higher than that of a Pre-Feasibility Study.
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The term proponent captures issuers who may finance a project without using traditional financial institutions. In these cases, the technical and economic confidence of the Feasibility Study is equivalent to that required by a financial institution.
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28. | Appendix A: Certification of Authors and Consent Forms |
Practical Mining LLC | March 16, 2015 |
CERTIFICATE of QUALIFIED PERSON
Re: Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”):
I, Mark A. Odell, P.E., do hereby certify that:
As of March 16, 2015, I am a consulting mining engineer at: | |
Practical Mining LLC | |
495 Idaho Street, Suite 205 | |
Elko, Nevada 89801 | |
775-345-3718 |
1) | I am a Registered Professional Mining Engineer in the State of Nevada (# 13708), and a Registered Member (#2402150) of the Society for Mining, Metallurgy and Exploration (SME). | |
2) | I graduated from The Colorado School of Mines, Golden, Colorado with a Bachelor of Science Degree in Mining Engineering in 1985. I have practiced my profession continuously since 1985. | |
3) | Since 1985, I have held the positions of mine engineer, chief engineer, mine superintendent, technical services manager and mine manager at underground and surface metal and coal mines in the western United States. The past 9 years, I have worked as a self-employed mining consultant with clients located in North America, Asia and Africa. My responsibilities have included the preparation of detailed mine plans, geotechnical engineering, reserve and resource estimation, preparation of capital and operating budgets and the economic evaluation of mineral deposits. | |
4) | I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my experience and qualifications and good standing with proper designation within a recognized professional organization fully meet the criteria as a Qualified Person as defined under NI 43-101. | |
5) | I am a contract consulting engineer for the Issuer and Project owner: Klondex Mines Ltd. and last inspected the Fire Creek Project on January 12, 2015. | |
6) | I am responsible for preparation of all sections of the Technical Report. | |
7) | I am independent of the Issuer within the meaning of Section 1.5 of NI 43-101. | |
8) | I was paid a daily rate for consulting services performed in evaluation of the Fire Creek Project for Klondex Mines Ltd. and do not have any other interests relating to the project. I do not have any interest in adjoining properties in the Fire Creek area. | |
9) | I have read NI 43-101 and Form 43-101F1, and the sections of the Technical Report for which I am responsible have been prepared in accordance with that instrument and form. |
10) | I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. | |
11) | As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading. |
Dated this16th day of March 2015,
“Signed” Mark A. Odell |
Mark A. Odell, P.E. |
Practical Mining LLC |
markodell@practicalmining.com |
CONSENT OF QUALIFIED PERSON
TO: | British Columbia Securities Commission |
Alberta Securities Commission | |
Saskatchewan Financial Services Commission | |
The Manitoba Securities Commission | |
Ontario Securities Commission | |
New Brunswick Securities Commission | |
Nova Scotia Securities Commission | |
Superintendent of Securities, Prince Edward Island Securities Office | |
Superintendent of Securities, Newfoundland and Labrador |
I, Mark Odell, P.E., do hereby consent to the public filing of the technical report titled Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”) by Klondex Mines Ltd. (the “Company”) with the Canadian securities regulatory authorities listed above and on SEDAR.
The undersigned consents to the use of any extracts from or a summary of the Technical Report in the news releases of the Company dated January 29 and February 23, 2015 (Written Disclosure).
The undersigned certifies that he has read the Written Disclosure being filed by the Company and that it fairly and accurately represents the information in the sections of the Technical Report for which the undersigned is responsible.
Dated this16th day of March 2015,
“Signed” Mark A. Odell |
Mark A, Odell, P.E. |
Practical Mining LLC |
markodell@practicalmining.com |
CERTIFICATE OF AUTHOR
Re: Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”).
I, Laura M. Symmes, SME, do hereby certify that:
As of March 16, 2015, I am a geologist at: |
Practical Mining, LLC |
495 Idaho Street, Suite 205 |
Elko, NV 89801 |
1) | I graduated with a Bachelor of Science degree in Geology from Utah State University in 2003. | |
2) | I am a registered member of the Society for Mining, Metallurgy & Exploration (SME) #4196936. | |
3) | I have worked as a geologist for a total of 11 years since my 2003 graduation from university. My experience has been focused on exploration and production of gold deposits, including planning and supervision of drill projects, generating data from drilled materials and making geologic interpretations, data organization, geologic mapping, building digital models of geologic features and mineral resources, and grade control of deposits in production. | |
4) | I have read the definition of “qualified person” set out in National Instrument 43 -101 (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 on NI 43-101. | |
5) | I am responsible for sections 4 and 6-12 of the Technical Report. I last visited the Fire Creek Project on September 18, 2014. | |
6) | I have not had prior involvement with the property that is the subject of the Technical Report. | |
7) | I am independent of Klondex Mines Ltd. within the meaning of Section 1.5 of National Instrument 43-101. | |
8) | I was paid a daily rate for consulting services performed in evaluation of the Fire Creek Project and do not have any other interests relating to the project. I do not have any interest in adjoining properties in the Fire Creek area. | |
9) | I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form. | |
10) | I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. |
11) | As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading. |
Dated this16th day of March 2015, | ||
“Signed” Laura M. Symmes | ||
Laura M. Symmes, SME | SME No. 4196936 | |
Practical Mining LLC | ||
495 Idaho Street, Suite 205 | ||
Elko, NV 89801 | ||
775-345-3718 | ||
Fax: (501) 638-9162 | ||
laurasymmes@practicalmining.com |
CONSENT OF QUALIFIED PERSON
TO: | British Columbia Securities Commission |
Alberta Securities Commission | |
Saskatchewan Financial Services Commission | |
The Manitoba Securities Commission | |
Ontario Securities Commission | |
New Brunswick Securities Commission | |
Nova Scotia Securities Commission | |
Superintendent of Securities, Prince Edward Island Securities Office | |
Superintendent of Securities, Newfoundland and Labrador |
I, Laura Symmes, SME., do hereby consent to the public filing of the technical report titled Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (Technical Report) by Klondex Mines Ltd. (Company) with the Canadian securities regulatory authorities listed above and on SEDAR.
The undersigned consents to the use of any extracts from or a summary of the Technical Report in the news releases of the Company dated January 29 and February 23, 2015 (Written Disclosure).
The undersigned certifies that she has read the Written Disclosure filed by the Company and that it fairly and accurately represents the information in the sections of the Technical Report for which the undersigned is responsible.
Dated this16th day of March 2015,
“Signed” Laura Symmes |
Laura Symmes, SME |
Practical Mining LLC |
495 Idaho Street, Suite 205 |
Elko, Nevada 89801 |
775-345-3718 |
laurasymmes@practicalmining.com |
CERTIFICATE OF AUTHOR
Re: Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”).
I, Sarah M Bull, P.E., do hereby certify that:
As of March 16, 2015, I am a consulting mining engineer at: | |
Practical Mining LLC | |
495 Idaho Street, Suite 205 | |
Elko, Nevada 89801 | |
775-345-3718 |
1) | I am a Registered Professional Mining Engineer in the State of Nevada (# 22797). | |
2) | I am a graduate of The University of Alaska Fairbanks, Fairbanks, Alaska with a Bachelor of Science Degree in Mining Engineering in 2006. | |
3) | Since my graduation from university I have been employed as a Mine Engineer at an underground gold mining operation and as Senior Mine Engineer for a consulting engineering firm. My responsibilities have included mine ventilation engineering, stope design and mine planning. | |
4) | I have read the definition of “qualified person” set out in National Instrument 43- 101 (NI 43-101) and certify that by reason of my experience and qualifications and good standing with proper designation within a recognized professional organization I fully meet the criteria as a Qualified Person as defined under the terms of NI 43-101. | |
5) | I am a contract consulting engineer for the issuer and Project owner: Klondex Mines Ltd. | |
6) | I am responsible for preparation of section 15 and 16 of the Technical Report. I last visited the Fire Creek Project on September 18, 2014. | |
7) | I am independent of Klondex Mines Ltd. within the meaning of Section 1.5 of NI 43-101. | |
8) | I was paid a daily rate for engineering consulting services performed in evaluation of the Fire Creek Project for Klondex Mines Ltd. and do not have any other interests relating to the project. I do not have any interest in adjoining properties in the Fire Creek Project area. | |
9) | I have read NI 43-101 and Form 43-101F1, and the sections of the Technical Report for which I am responsible have been prepared in accordance with that instrument and form. | |
10) | I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. |
11) | As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading. |
Dated this16th day of March 2015. |
“Signed” Sarah Bull |
Sarah M Bull, P.E. |
Practical Mining LLC |
495 Idaho Street, Suite 205 |
Elko, Nevada 89801 |
775-304-5836 |
sarahbull@practicalmining.com |
CONSENT OF QUALIFIED PERSON
TO: | British Columbia Securities Commission |
Alberta Securities Commission | |
Saskatchewan Financial Services Commission | |
The Manitoba Securities Commission | |
Ontario Securities Commission | |
New Brunswick Securities Commission | |
Nova Scotia Securities Commission | |
Superintendent of Securities, Prince Edward Island Securities Office | |
Superintendent of Securities, Newfoundland and Labrador |
I, Sarah Bull, P.E., do hereby consent to the public filing of the technical report titled Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (Technical Report) by Klondex Mines Ltd. (Company) with the Canadian securities regulatory authorities listed above and on SEDAR.
The undersigned consents to the use of any extracts from or a summary of the Technical Report in the news releases of the Company dated January 29 and February 23, 2015 (Written Disclosure).
The undersigned certifies that she has read the Written Disclosure filed by the Company and that it fairly and accurately represents the information in the sections of the Technical Report for which the undersigned is responsible.
Dated this16th day of March 2015,
“Signed” Sarah Bull |
Sarah Bull, P.E. |
Practical Mining LLC |
495 Idaho Street, Suite 205 |
Elko, Nevada 89801 |
775-304-5836 |
sarahbull@practicalmining.com |
Karl T. Swanson, SME, MAusIMM |
PO Box 86 |
Larkspur, CO 80118, USA |
Fax: (501) 638-9162 |
Email: karl.swanson@yahoo.com |
CERTIFICATE OF AUTHOR
Re: Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”).
I, Karl T. Swanson, SME, MAusIMM, do hereby certify that:
As of March 16, 2015, I am an independent geological and mining engineering consultant at: |
Karl Swanson |
PO Box 86 |
Larkspur, CO 80118, USA |
1) | I graduated with a Bachelor of Science degree in Geological Engineering from Colorado School of Mines in 1990. In addition, I obtained a Master of Engineering degree in Mining Engineering from Colorado School of Mines in 1994. | |
2) | I am a registered member of the Society for Mining, Metallurgy & Exploration (SME) #4043076. I am a member of the Australian Institute of Mining and Metallurgy (AusIMM) #304871. | |
3) | Since my 1990 graduation from university I have been employed as a geologic modeller and resource geologist for metal mining companies and consulting groups. For the past 17 years, I have been a self-employed consulting geologist specializing in digital geologic modelling, geostatistical grade estimation and block modelling for precious metal, base metal and industrial mineral deposits. I have been the principle geostatistician and modeller for several narrow vein gold deposits in the Northern Nevada Rift for over 5 years. | |
4) | I have read the definition of “qualified person” set out in National Instrument 43- 101 (“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 on NI 43-101. | |
5) | I am responsible for Section 14 of the Technical Report. | |
6) | I last visited the Fire Creek Project January 19 – 22, 2015. | |
7) | I have not had prior involvement with the property that is the subject of the Technical Report. | |
8) | I am independent of the Issuer within the meaning of Section 1.5 of NI 43-101. |
9) | I was paid a daily rate for engineering consulting services performed in evaluation of the Fire Creek Project and do not have any other interests relating to the project. I do not have any interest in adjoining properties in the Fire Creek Project area. | |
10) | I have read National Instrument 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 for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report. | |
12) | As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading. |
Dated this16th day of March 2015, | ||
“Signed” Karl T. Swanson | ||
AusIMM No. 304871 | ||
Karl T. Swanson, M.Eng., SME, MAusIMM | SME No. 4043076 | |
PO Box 86 | ||
Larkspur, CO 80118, USA | ||
Fax: (501) 638-9162 | ||
E:mail: karl.swanson@yahoo.com |
Karl T. Swanson, SME |
PO Box 86 |
Larkspur, CO 80118, USA |
Fax: (501) 638-9162 |
Email: karl.swanson@yahoo.com |
CONSENT OF QUALIFIED PERSON |
TO: | British Columbia Securities Commission |
Alberta Securities Commission | |
Saskatchewan Financial Services Commission | |
The Manitoba Securities Commission | |
Ontario Securities Commission | |
New Brunswick Securities Commission | |
Nova Scotia Securities Commission | |
Superintendent of Securities, Prince Edward Island Securities Office | |
Superintendent of Securities, Newfoundland and Labrador |
I, Karl Swanson, SME, MAusIMM, do hereby consent to the public filing of the technical report titled
Preliminary Feasibility Study for the Fire Creek Project, Lander County, Nevada, dated the dated the 16th day of March 2015, with an effective date of December 31, 2014 (the “Technical Report”) by Klondex Mines Ltd. (the “Company”) with the Canadian Securities Regulatory Authorities listed above and on SEDAR.
The undersigned consents to the use of any extracts from or a summary of the Technical Report in the news releases of the Company dated January 29 and February 23, 2015 (Written Disclosure).
The undersigned certifies that he has read the Written Disclosure being filed by the Company and that it fairly and accurately represents the information in the sections of the Technical Report for which the undersigned is responsible.
Dated this16th day of March 2015,
“Signed” Karl Swanson |
Karl Swanson, SME, MAusIMM |
PO Box 86 |
Larkspur, CO 80118, USA |
Fax: (501) 638-9162 |
Email: karl.swanson@yahoo.com |