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NI 43-101 Technical Report Preliminary Economic Assessment Crescent Silver Project Kellogg, Idaho, USA Effective Date: July 22, 2013 Report Date: September 24, 2013

Report Prepared for

United Silver Corp. 1220 Big Creek Road Kellogg, Idaho 83837

Report Prepared by

SRK Consulting (U.S.), Inc. 5250 Neil Road, Suite 300 Reno, NV 89502 SRK Project Number: 202500.030

Signed by Qualified Persons: J. B. Pennington, CPG Kent Hartley, BSc, PE Mining Peer Reviewed by: Bruno Serra, BS (Economics) Peter Clarke, BSc Mining, MBA, PE (Mining) Matthew Hastings, MSc (Geology and Resources) Mark Willow, CEM (Environmental)

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JBP/KH/BJM CrescentPEA_NI43-101_TR_202500.030_Rev19-AK September 24, 2013

Table of Contents 1 Summary ....................................................................................................................... 1

1.1 Property Description and Ownership .................................................................................................. 1

1.2 Geology and Mineralization ................................................................................................................ 2

1.3 Status of Exploration Development and Operations ........................................................................... 3

1.4 Mineral Processing and Metallurgical Testing .................................................................................... 4

1.5 Mineral Resource Estimate ................................................................................................................. 5

1.6 Mineral Reserve Estimate ................................................................................................................... 6

1.7 Mining Methods ................................................................................................................................... 6

1.8 Recovery Methods .............................................................................................................................. 7

1.9 Project Infrastructure ........................................................................................................................... 7

1.10 Environmental Studies and Permitting ................................................................................................ 8

1.11 Capital and Operating Costs ............................................................................................................... 9

1.12 Economic Analysis ............................................................................................................................ 10

1.13 Conclusions and Recommendations ................................................................................................ 12

1.13.1 Recommended Work Programs and Costs ........................................................................... 12

2 Introduction ................................................................................................................ 14

2.1 Terms of Reference and Purpose of the Report ............................................................................... 14

2.2 Qualifications of Consultants (SRK) .................................................................................................. 14

2.3 Details of Inspection .......................................................................................................................... 15

2.4 Sources of Information ...................................................................................................................... 15

2.5 Effective Date .................................................................................................................................... 15

2.6 Units of Measure ............................................................................................................................... 15

3 Reliance on Other Experts ........................................................................................ 16

3.1 Economic Analysis ............................................................................................................................ 16

3.2 Mining Methods ................................................................................................................................. 16

4 Property Description and Location .......................................................................... 17

4.1 Property Location .............................................................................................................................. 17

4.2 Mineral Titles and Surface Rights ..................................................................................................... 17

4.2.1 Nature and Extent of Issuer’s Interest ................................................................................... 20

4.3 Royalties, Agreements and Encumbrances ...................................................................................... 20

4.4 Environmental Liabilities and Permitting ........................................................................................... 21

4.4.1 Required Permits and Status ................................................................................................ 22

4.5 Other Significant Factors and Risks .................................................................................................. 23

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ........ 29

5.1 Topography, Elevation and Vegetation ............................................................................................. 29

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5.2 Accessibility and Transportation to the Property .............................................................................. 29

5.3 Climate and Length of Operating Season ......................................................................................... 30

5.4 Sufficiency of Surface Rights ............................................................................................................ 30

5.5 Infrastructure Availability and Sources .............................................................................................. 30

5.5.1 Power .................................................................................................................................... 30

5.5.2 Water ..................................................................................................................................... 31

5.5.3 Mining Personnel ................................................................................................................... 31

5.5.4 Potential Tailings Storage Areas ........................................................................................... 31

5.5.5 Potential Waste Disposal Areas ............................................................................................ 31

5.5.6 Potential Heap Leach Pad Areas .......................................................................................... 32

5.5.7 Potential Processing Plant Sites ........................................................................................... 32

6 History ......................................................................................................................... 33

6.1 Prior Ownership and Ownership Changes ........................................................................................ 33

6.2 Previous Exploration and Development Results ............................................................................... 34

6.2.1 Historical Development ......................................................................................................... 34

6.2.2 Bunker Hill 1942-1985 ........................................................................................................... 34

6.2.3 GFN 2007-2008 ..................................................................................................................... 35

6.3 Historic Mineral Resource and Reserve Estimates .......................................................................... 35

6.3.1 Bunker Hill Mineral Resource and Reserve Estimates ......................................................... 35

6.3.2 SRK Mineral Resource Estimate ........................................................................................... 35

6.3.3 Historic Production ................................................................................................................ 36

6.3.4 Historic Metallurgy ................................................................................................................. 37

7 Geological Setting and Mineralization ..................................................................... 40

7.1 Regional Geology .............................................................................................................................. 40

7.2 Local Geology ................................................................................................................................... 40

7.3 Property Geology .............................................................................................................................. 43

7.4 Significant Mineralized Zones ........................................................................................................... 44

8 Deposit Type .............................................................................................................. 48

8.1 Mineral Deposit ................................................................................................................................. 48

8.2 Geological Model .............................................................................................................................. 48

9 Exploration ................................................................................................................. 50

9.1 Relevant Exploration Work ............................................................................................................... 50

9.2 Sampling Methods and Sample Quality ............................................................................................ 50

9.3 Significant Results and Interpretation ............................................................................................... 53

10 Drilling ......................................................................................................................... 56

10.1 Type and Extent ................................................................................................................................ 56

10.1.1 GFN 2007-2008 Programs .................................................................................................... 56

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10.1.2 USC 2011 Programs ............................................................................................................. 57

10.2 Procedures ........................................................................................................................................ 57

10.3 Interpretation and Relevant Results .................................................................................................. 59

10.3.1 GFN 2007-2008 Programs .................................................................................................... 59

10.3.2 USC 2011 Programs ............................................................................................................. 59

11 Sample Preparation, Analysis and Security ............................................................ 64

11.1 Security Measures ............................................................................................................................ 64

11.1.1 GFN 2007-2008 Drill Programs ............................................................................................. 64

11.1.2 USC 2011 Drill Programs ...................................................................................................... 64

11.2 Sample Preparation for Analysis ....................................................................................................... 64

11.2.1 GFN 2007-2008 Drill Programs ............................................................................................. 64


11.3 Sample Analysis ................................................................................................................................ 65

11.3.1 GFN Sample Analysis ........................................................................................................... 65


11.4 Quality Assurance/Quality Control Procedures ................................................................................ 66

11.4.1 Standard Reference Materials............................................................................................... 66

11.4.2 Blank Samples ...................................................................................................................... 67

11.4.3 Duplicate Samples ................................................................................................................ 67

11.4.4 Actions ................................................................................................................................... 67

11.4.5 Results ................................................................................................................................... 68

11.5 Opinion on Adequacy ........................................................................................................................ 68

12 Data Verification ......................................................................................................... 76

12.1 Procedures ........................................................................................................................................ 76

12.1.1 GFN Check Sampling ............................................................................................................ 76

12.1.2 USC Check Sampling ............................................................................................................ 76

12.1.3 Comparison of Assay Certificates and Drillhole Database ................................................... 76

12.2 Limitations ......................................................................................................................................... 77

12.3 Opinion on Data Adequacy ............................................................................................................... 77

13 Mineral Processing and Metallurgical Testing ........................................................ 80

13.1 Testing and Procedures .................................................................................................................... 80

13.1.1 Sample Representativeness ................................................................................................. 80

13.2 Relevant Results ............................................................................................................................... 80

13.2.1 Mill Feed Hardness ............................................................................................................... 81

13.2.2 Rougher Flotation Tests ........................................................................................................ 81

13.2.3 Batch Cleaner Flotation Tests ............................................................................................... 82

13.2.4 Bulk Sample Milling ............................................................................................................... 82

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13.2.5 Results of Bulk Sample Milling .............................................................................................. 82

13.3 Recovery Estimate Assumptions ...................................................................................................... 83

13.4 Significant Factors ............................................................................................................................. 83

14 Mineral Resource Estimate ....................................................................................... 85

14.1 Introduction ....................................................................................................................................... 85

14.2 Block Model ....................................................................................................................................... 85

14.3 Density .............................................................................................................................................. 86

14.4 Geology and Mineral Domains .......................................................................................................... 86

14.5 Drillhole Database ............................................................................................................................. 87

14.6 Silver Assays Analysis ...................................................................................................................... 88

14.7 Outlier Treatment .............................................................................................................................. 90

14.8 Compositing ...................................................................................................................................... 90

14.9 Variogram Analysis ........................................................................................................................... 91

14.10 Estimation Methodology .................................................................................................................... 91

14.11 Model Validation ................................................................................................................................ 92

14.12 Resource Classification..................................................................................................................... 93

14.13 Mineral Resource Statement ............................................................................................................ 94

14.14 Mineral Resource Sensitivity ............................................................................................................. 95

14.15 Relevant Factors ............................................................................................................................... 95

15 Mineral Reserve Estimate ........................................................................................ 116

16 Mining Methods ........................................................................................................ 117

16.1 Current or Proposed Mining Methods ............................................................................................. 117

16.1.1 Current Mining Method ........................................................................................................ 117

16.1.2 Proposed Mining Method .................................................................................................... 117

16.1.3 Selection of Mining Method ................................................................................................. 117

16.1.4 Mine Development .............................................................................................................. 118

16.1.5 South Vein Development and Mining .................................................................................. 118

16.1.6 Alhambra Vein Development and Mining ............................................................................ 118

16.2 Parameters Relevant to Mine Designs and Plans .......................................................................... 119

16.2.1 Geotechnical ....................................................................................................................... 119

16.2.2 Hydrological ......................................................................................................................... 119

16.3 Mine Optimization ........................................................................................................................... 119

16.3.1 Development and Operations.............................................................................................. 119

16.3.2 Ground Support ................................................................................................................... 119

16.3.3 Drilling .................................................................................................................................. 120

16.3.4 Blasting ................................................................................................................................ 120

16.3.5 Backfill ................................................................................................................................. 120

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16.3.6 Ventilation ............................................................................................................................ 120

16.3.7 Utilities: Compressed Air, Water and Electric Power .......................................................... 120

16.3.8 Health and Safety ................................................................................................................ 120

16.3.9 Organization and Staffing .................................................................................................... 121

16.4 Mine Production Schedule .............................................................................................................. 121

16.4.1 Mine Production .................................................................................................................. 122

16.5 Waste and Stockpile Design ........................................................................................................... 125

16.6 Mining Fleet and Requirements ...................................................................................................... 126

16.7 Mine Dewatering ............................................................................................................................. 126

17 Recovery Methods ................................................................................................... 133

17.1 Operation Results ........................................................................................................................... 133

17.1.1 Concentrate Sales ............................................................................................................... 134

17.1.2 Mill Reconciliation with Mine Production ............................................................................. 134

17.2 Processing Methods ........................................................................................................................ 135

17.3 Plant Design and Equipment Characteristics.................................................................................. 135

17.3.1 Crushing and Grinding Circuits ........................................................................................... 136

17.3.2 Flotation Circuit ................................................................................................................... 136

17.3.3 Paste Tailings Disposal ....................................................................................................... 137

17.4 Consumable Requirements ............................................................................................................ 137

18 Project Infrastructure ............................................................................................... 140

18.1 Infrastructure and Logistic Requirements ....................................................................................... 140

18.1.1 On-Site Infrastructure .......................................................................................................... 140

18.1.2 Site Water Management ...................................................................................................... 140

18.1.3 Service Roads and Bridges ................................................................................................. 140

18.1.4 Mine Operations and Support Facilities .............................................................................. 140

18.1.5 Process Support Facilities ................................................................................................... 141

18.1.6 Additional Support Facilities ................................................................................................ 142

18.1.7 Power Supply and Distribution ............................................................................................ 142

18.1.8 Water Supply ....................................................................................................................... 142

18.1.9 Rail ...................................................................................................................................... 142

18.1.10 Port .................................................................................................................................. 143

18.2 Tailings Management Area ............................................................................................................. 143

18.3 Off-Site Infrastructure and Logistics Requirements ........................................................................ 143

19 Market Studies and Contracts ................................................................................ 144

19.1 Summary of Information .................................................................................................................. 144

19.1.1 Nature of Material Terms ..................................................................................................... 144

19.1.2 Commodity Price Projections .............................................................................................. 144

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19.2 Contracts and Status ....................................................................................................................... 144

19.2.1 Terms .................................................................................................................................. 144

20 Environmental Studies, Permitting and Social or Community Impact ................ 146

20.1 Required Permits and Status .......................................................................................................... 146

20.1.1 Summary of Operating and Environmental Permits ............................................................ 147

20.2 Environmental Study Results .......................................................................................................... 147

20.3 Environmental Issues ...................................................................................................................... 147

20.4 Operating and Post Closure Requirements and Plans ................................................................... 148

20.5 Post-Performance or Reclamations Bonds ..................................................................................... 148

20.6 Social and Community .................................................................................................................... 148

20.7 Mine Closure ................................................................................................................................... 148

20.8 Reclamation Measures During Operations and Project Closure .................................................... 148

20.9 Reclamation and Closure Cost Estimate ........................................................................................ 148

21 Capital and Operating Costs ................................................................................... 149

21.1 Capital Cost Estimates .................................................................................................................... 149

21.1.1 Preproduction Capital .......................................................................................................... 149

21.1.2 Post Development Capital ................................................................................................... 150

21.2 Operating Cost Estimates ............................................................................................................... 151

21.2.1 Basis for Mining Operating Cost Estimates ......................................................................... 151

21.2.2 Mill Feed Transportation and Processing ............................................................................ 152

21.2.3 General and Administrative Cost ........................................................................................ 152

22 Economic Analysis .................................................................................................. 154

22.1 Principal Assumptions and Input Parameters ................................................................................. 154

22.2 Cashflow Forecasts and Annual Production Forecasts .................................................................. 154

22.3 Taxes, Royalties and Other Interests .............................................................................................. 155

22.4 Sensitivity Analysis .......................................................................................................................... 156

23 Adjacent Properties ................................................................................................. 158

24 Other Relevant Data and Information ..................................................................... 159

25 Interpretation and Conclusions .............................................................................. 160

26 Recommendations ................................................................................................... 163

26.1 Recommended Work Programs and Costs .................................................................................... 163

27 References ................................................................................................................ 165

28 Glossary .................................................................................................................... 168

28.1 Mineral Resources .......................................................................................................................... 168

28.2 Mineral Reserves ............................................................................................................................ 168

28.3 Definition of Terms .......................................................................................................................... 169

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28.4 Abbreviations .................................................................................................................................. 170

List of Tables Table 1.5.1: Mineral Resource Statement for the Crescent Silver Project, SRK Consulting (U.S.) Inc., July 22,

2013 ....................................................................................................................................................... 6

Table 1.11.1: Preproduction Mine Development Cost ........................................................................................ 9

Table 1.11.2: Post Development Capital ............................................................................................................ 9

Table 1.11.3: Mine Productivities and Unit Costs ............................................................................................. 10

Table 1.12.2: Market Parameters ..................................................................................................................... 10

Table 1.12.3: Production Parameters ............................................................................................................... 11

Table 1.12.4: Economic Results ....................................................................................................................... 11

Table 1.12.5: Sensitivity Analysis of NPV @ 8% (US$000) ............................................................................. 11

Table 1.13.1.1: Recommended Pre-Production Work Program Costs ............................................................ 13

Table 2.3.1: Site Visit Participants .................................................................................................................... 15

Table 4.2.1: Patented Mining Claims and Surface Ownership ......................................................................... 19

Table 5.3.1: Summary of Meteorological Data ................................................................................................. 30

Table 5.5.1.1: Underground Electrical Power Estimate .................................................................................... 31

Table 6.3.2.1: 2009 Crescent Mine Mineral Resource Statement (at 11.0 oz/t CoG) ...................................... 36

Table 6.3.2.2: 2009 Crescent Mine Indicated Mineral Resource Sensitivity .................................................... 36

Table 6.3.2.3: 2009 Crescent Mine Inferred Mineral Resource Sensitivity ...................................................... 36

Table 10.2.1: 2011 Drilling Program ................................................................................................................. 59

Table 10.3.2.1: Significant Intercepts in 2011 Drilling Programs ...................................................................... 60

Table 13.2.1: Head Assay Data for G&T Sample ............................................................................................. 81

Table 13.2.2: Head Assay Data for South Vein Bulk Sample .......................................................................... 81

Table 13.2.1.1: Bond Ball Mill Work Index Data ............................................................................................... 81

Table 13.2.2.1: G&T Rougher Flotation Test Results ...................................................................................... 82

Table 13.2.3.1: G&T Cleaner Flotation Test Result ......................................................................................... 82

Table 13.2.5.1: Mass Balance for South Vein Bulk Sample ............................................................................. 83

Table 14.2.1: Crescent Model Origin and Extents ............................................................................................ 85

Table 14.3.1: Density Assignment by Material Type ........................................................................................ 86

Table 14.6.2: Silver Assay Statistics by Sample Type ..................................................................................... 90

Table 14.8.1: Crescent Composite Statistics by Domain ................................................................................. 90

Table 14.8.2: Declustered Composite Statistics by Domain ............................................................................ 91

Table 14.10.1: Estimation Parameters for the SRK Crescent Block Model ..................................................... 92

Table 14.11.1: Comparison of IDW and NN Tonnage and Grade at a Zero CoG ............................................ 93

Table 14.12.1: Crescent Resource Classification Criteria ................................................................................ 94

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Table 14.13.1: Mineral Resource Statement for the Crescent Silver Project, SRK Consulting (U.S.) Inc., July 22, 2013 ............................................................................................................................................... 94

Table 14.14.1: Crescent Measured and Indicated Resource Sensitivity .......................................................... 95

Table 14.14.2: Crescent Inferred Resource Sensitivity .................................................................................... 95

Table 16.2.1.1: Dimensions of Working Areas ............................................................................................... 119

Table 16.3.9.1: Staffing at Full Production .................................................................................................... 121

Table 16.4.1.1: Production Schedule ............................................................................................................. 122

Table 16.4.1.2: Production Summary ............................................................................................................. 123

Table 16.6.1: Existing Mining Equipment ....................................................................................................... 126

Table 16.6.2: Cost of Additional Mining Equipment ....................................................................................... 126

Table 17.1: Head Data for South Vein Mining ................................................................................................ 133

Table 17.1.1: Mass Balance for the South Vein Milling Campaign 2012 & 2013 ........................................... 133

Table 17.1.2: Assay and Recovery Data for the South Vein by Sublevel ...................................................... 134

Table 17.1.1.1: Selected Elements in Concentrate ........................................................................................ 134

Table 17.1.2.1: Mine to Mill Reconciliation ..................................................................................................... 134

Table 17.3.2.1: Flotation Reagent Scheme .................................................................................................... 136

Table 17.4.1: Consumable Requirements ...................................................................................................... 137

Table 19.2.1.1: Smelting and Refining Costs ................................................................................................. 144

Table 20.1.1: Summary of Existing Permits ................................................................................................... 147

Table 20.9.1: Closure Cost Estimates ............................................................................................................ 148

Table 21.1.1: Capital Cost Summary .............................................................................................................. 149

Table 21.1.1.1: Preproduction Mine Development Cost ................................................................................. 149

Table 21.1.1.2: Other Preproduction Capital Requirements .......................................................................... 150

Table 21.1.1.3: Cost of Additional Mining Equipment .................................................................................... 150

Table 21.1.2.1: Post Production Development ............................................................................................... 150

Table 21.1.2.2: Other Post Development Capital ........................................................................................... 150

Table 21.2.1.1: Equipment Operating Rates .................................................................................................. 151

Table 21.2.1.2: Hourly Labor Rates (US$) ..................................................................................................... 151

Table 21.2.1.3: Mine Operating Productivities and Unit Cost ......................................................................... 152

Table 21.2.2.1: Surface Mill Feed Transportation and Process Cost ............................................................. 152

Table 21.2.3.1: G&A Labor Rates .................................................................................................................. 152

Table 21.2.3.2: G&A operating Cost Summary .............................................................................................. 153

Table 22.1.1: Market Parameters ................................................................................................................... 154

Table 22.1.2: Production Parameters ............................................................................................................. 154

Table 22.2.1: Annual Mine Production and Cashflow Summary .................................................................... 154

Table 22.2.2: Economic Results ..................................................................................................................... 155

Table 22.4.1: Sensitivity Analysis of NPV @ 8% (US$000) ........................................................................... 156

Table 25.1: Relevant Risks and Opportunities ............................................................................................... 160

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Table 26.1: Recommended Pre-Production Work Program Costs................................................................. 163

Table 28.3.1: Definition of Terms ................................................................................................................... 169

Table 28.4.1: Abbreviations ............................................................................................................................ 170

List of Figures Figure 4.1.1: Project Location Map (large-scale) ............................................................................................. 24

Figure 4.1.2: Crescent Mine Map (small scale) ................................................................................................ 25

Figure 4.2.1: Mineral Rights Map .................................................................................................................... 26

Figure 4.2.2: Atypical Mineral Claim Boundaries Map .................................................................................... 27

Figure 4.2.3: Surface Rights Map .................................................................................................................... 28

Figure 6.2.1.1: Long Section Showing Historic Development .......................................................................... 39

Figure 7.1.1: Geologic Map of the Coeur d’Alene District ............................................................................... 45

Figure 7.2.1: Geologic Map of the Crescent Mine Area ................................................................................... 46

Figure 7.2.2: Geologic Cross-Section of Crescent Mine Area .......................................................................... 47

Figure 9.3.1: Development on the South Vein .................................................................................................. 54

Figure 9.3.2: Underground Advance by USC 2010-2012 ................................................................................. 55

Figure 10.1.1.1: Map of GFN and USC Drillhole Traces .................................................................................. 61

Figure 10.1.2.1: Alhambra Long Section with 2011 Drillhole Intercepts .......................................................... 62

Figure 10.3.2.1: South Vein Long Section with 2011 Drill Intercepts ............................................................... 63

Figure 11.4.5.1: GFN Standards Analyses ....................................................................................................... 70

Figure 11.4.5.2: USC Standards Analyses ....................................................................................................... 71

Figure 11.4.5.3: GFN and USC Blanks ............................................................................................................ 72

Figure 11.4.5.4: GFN Duplicates ...................................................................................................................... 73

Figure 11.4.5.5: GFN Quarter Core Samples ................................................................................................... 74

Figure 11.4.5.6: USC Duplicates ...................................................................................................................... 75

Figure 12.1.1.1: GFN Check Samples .............................................................................................................. 78

Figure 12.1.2.1: USC Check Samples .............................................................................................................. 79

Figure 13.2.5.1: Particle Distribution of Bulk Sample Flotation Feed ............................................................... 84

Figure 14.4.1: Plan View of Drilling and Underground Workings .................................................................... 96

Figure 14.4.2: Plan View of Mineral Domains at the 3940 Elevation .............................................................. 97

Figure 14.4.3: Cross-Section View of Mineralized Veins and Model Blocks ................................................... 98

Figure 14.4.4: Modeled Oxidation on the South Vein ...................................................................................... 99

Figure 14.6.1: Box Plot of Silver by Vein (grades in oz/t) .............................................................................. 100

Figure 14.6.2: Box Plot of Silver in the South Vein by Sample Type (grades in oz/t) ................................... 101

Figure 14.7.1: Cumulative Probability Plot for Silver in the Alhambra Vein .................................................. 102

Figure 14.7.2: Cumulative Probability Plot for Silver in the South Vein ........................................................ 103

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Figure 14.7.3: Cumulative Probability Plot for Silver in the Jackson Vein ..................................................... 104

Figure 14.11.1: Declustered Composites on the South Vein (Looking N35E, -12) ....................................... 105

Figure 14.11.2: Modeled Silver on the South Vein (Looking N35E, -12) ...................................................... 106

Figure 14.11.3: Declustered Composites on the Alhambra Vein (Looking N35E, -12) ................................. 107

Figure 14.11.4: Modeled Silver on the Alhambra Vein (Looking N35E, -12) ................................................ 108

Figure 14.11.5: Distribution of Model Blocks and Composites ...................................................................... 109

Figure 14.11.6: Swath Plot – Alhambra East-West ....................................................................................... 110

Figure 14.11.7: Swath Plot – Alhambra North-South .................................................................................... 111

Figure 14.11.8: Swath Plot – South Vein East-West ..................................................................................... 112

Figure 14.11.9: Swath Plot – South Vein North-South .................................................................................. 113

Figure 14.12.1: Classification for the South Vein .......................................................................................... 114

Figure 14.12.2: Classification for the Alhambra Vein .................................................................................... 115

Figure 16.1.4.1: Proposed Development Long Section .................................................................................. 127

Figure 16.1.4.2: Countess Decline and BD#4 Development Cross Section .................................................. 128

Figure 16.1.5.1: South Vein I-Drift Cross Section........................................................................................... 129

Figure 16.1.5.2: South Vein Stope Cut ........................................................................................................... 130

Figure 16.1.6.1: Typical Alhambra Stope Development ................................................................................. 131

Figure 16.3.9.1: Mine Organization ................................................................................................................ 132

Figure 17.2.1: Flotation Flowsheet ................................................................................................................. 138

Figure 17.3.1.1: Particle Size Distribution – Grinding Circuit ......................................................................... 139

Figure 19.1.2.1: Silver Price, January – August 2013 .................................................................................... 145

Figure 22.4.1: Sensitivities .............................................................................................................................. 157

Appendices Appendix A: Certificates of Qualified Persons

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1 Summary This report was prepared as a National Instrument 43-101 (NI 43-101) Technical Report, Preliminary Economic Assessment (PEA) for United Silver Corp. (USC) on their Crescent Silver Project (Crescent, or the Project) located in Kellogg, Idaho, USA. Crescent is an advanced underground silver project located in the “Silver Belt” of the Coeur d’Alene Mining District of Idaho with a favorable economic projection based on PEA-level capital and operating costs.

This Technical Report provides Mineral Resource estimates, and a classification of resources in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum Standards on Mineral Resources and Reserves: Definitions and Guidelines, November 27, 2010 (CIM). This report follows and supports a press release issued by USC on July 29, 2013 entitled “United Silver Corp. Releases Positive Preliminary Economic Assessment and Reports Updated Estimated Resources” in which the Company provided an updated resource estimate and a preliminary projection of Project economics.

1.1 Property Description and Ownership The Project is located in Shoshone County, Idaho at 116.074349° West longitude and 47.505195° North latitude, in the Idaho Panhandle. The Project is in Idaho Township 48 North, Range 03 East, Sections 15, 16, 17, 18, 19 and 20. It is located about four miles southeast of Kellogg, Idaho, and about 75 miles east of Spokane, Washington.

The property is located in the northern Rockies and has average summer high temperatures of 82° F and average winter low temperatures of 22° F. The average rainfall in Kellogg, Idaho (elevation approximately 2,300 ft amsl) is about 31 inches per year and the mean average snowfall is 54 inches per year (WRCC, 2009). Surface drilling has been conducted from May through November. Historically, the Project operated year-round.

On December 12, 2006, Gold Finder Explorations, Ltd. (GFN) acquired the right to purchase the Project from Shoshone County, Idaho for US$650,000. The transaction was completed on January 2, 2007. On July 13, 2007, two claim fractions named Queen Lode, lying within section 16 were purchased from The New Bunker Hill Mining Company (NBH). These were fractions lying within the contiguous property boundaries. On November 28, 2007, GFN acquired the surface rights to a 25.69 acre (ac) parcel north of, and contiguous with, the surface rights already owned on the Diana and Hiawatha claims.

On December 30, 2009, United Silver Corp (USC) signed an Earn-In Agreement with GFN to earn up to an 80% interest in the Crescent Project. On June 1, 2011, USC and GFN jointly announced that USC had earned an 80% interest in the Crescent Project through the expenditure of US$9 million in exploration work (2011). The Project, at the time the Earn-In Agreement was signed, consisted of 25 patented mining claims or portions thereof totaling 346.190 ac, and surface rights totaling 40.639 ac. A 25.69 ac portion of the surface rights was inadvertently omitted from the Earn-In Agreement and so the surface rights detailed in the Agreement consist only of the 14.949 ac parcel that comprises the surface rights of the Diana and Hiawatha claims. USC is working to correct this deficiency.

In June 2010, during the Earn-In Option period, USC purchased the surface rights to a 10.01 ac parcel of the Countess patented mining claim, in order to construct a portal to begin underground



development. This surface parcel was contributed to the Project property package per the Earn-In Agreement. In December 2010, USC purchased the mineral rights for an additional 42 patented mining claims or portions thereof from NBH. These claims consisted of two blocks, one block contiguous on the northeast of the original Project, and the other block contiguous on the west. The purchase from NBH brought the total number of patented mining claims or portions thereof to their current number, 64, comprising 897.581 ac.

The mineral rights of the Project are all in the form of patented claims. These claims have been modified over the decades by agreements between previous owners and the adjoining property owners that established inclined or horizontal boundaries in some places. These atypical boundaries do not affect the area of the current resource.

Patented claims typically convey full surface ownership, but the surface rights of the Project claims were severed from the underlying mineral rights in 1991 by a previous owner, and sold to a timber company with the exception of two claims that cover the area of the Hooper portal and associated mine buildings.

The current surface rights of the Project form two separate parcels. The first is a contiguous, irregularly-shaped parcel of ground totaling 14.949 ac. This parcel consists of surface rights to the patented Hiawatha claim (11.331 ac) and Diana claim (3.618 ac); it covers the area of the Hooper portal and dump. The second parcel consists of a parallelogram-shaped block of ground measuring 10.01 ac within the Countess claim, and covering the area of the Countess portal and dump.

The Hiawatha and Diana claims have never had their surface rights severed from the mineral rights, but they are part of an area affected by a 1958 agreement with the Sunshine Mine which conveys the mineral rights below 2,691 ft amsl to Sunshine. The 10.01 acre surface rights block on the Countess claim overlies mineral rights controlled by the Project.

The remaining surface rights on the patented mineral claims that comprise the Project are largely within sections 15, 16, 17 and 18, and are controlled by Stimson Lumber Co., while an area of about 92 ac within sections 19 and 20 is controlled by Silver Mountain Corp. and by Silver Mountain Ventures LLC. The surface rights extend 100 ft vertically below the current topography, below which, mineral rights are controlled by USC. The sub-surface mineral claims controlled by USC fully encompass the resource discussed below.

1.2 Geology and Mineralization Regional geology is dominated by Precambrian sedimentary rocks of the Belt-Purcell Supergroup (Belt), which have been strongly deformed during the Cretaceous age Sevier Orogeny. This regional deformation has resulted in large-scale folds cut by numerous west to northwest striking faults and veins. Early compressional tectonics dominated the area forming large-scale folds, reverse and thrust faults. Many of these structures were focused along the west-northwest trending Lewis and Clark Line. This is a regional, deep-seated lineament believed to represent an intra-plate boundary, which has been recurrently active since the Proterozoic. During the late Cretaceous, the Bitterroot Lobe of the Idaho Batholith was emplaced to the south, accompanied by dike emplacement in this area and normal movement along earlier reverse faults. The major mineralizing event is believed to have occurred during the compressional phase of the Sevier Orogeny. The most recent tectonic activity is believed to have occurred during the Tertiary when the Lewis and Clark lineament was reactivated along the Osburn Fault. This event resulted in 16 mi of right lateral, strike slip movement



that has dissected and displaced many of the deposits in the region (Hobbs et al. 1965; Lewis et al., 2002). The Belt rocks host all of the mineralized veins at the Project. The three Belt units present are Wallace, St. Regis and Revett Formations. Detailed lithologic mapping has not yet been completed to know exactly where the St. Regis/Revett formation boundary lies with respect to the mineralized veins. The interfingered facies character of the boundary has complicated efforts to pinpoint its location.

At Crescent, silver mineralization has been mined on the Alhambra Fault from near surface to 1,500 ft below mean sea level, a vertical distance of about 5,800 ft. In the upper workings of the mine, above the 1,500 ft elevation, the higher-grade mineralization occurs on the immediate hangingwall of the fault (Julihn and Horton, 1936), where the vein cuts Revett and St. Regis Formations. The zones mined between mean sea level and -1,500 ft elevation occur on the immediate footwall in the Revett Formation. The other significant production from the mine has been from a set of veins found within the footwall of the Alhambra Fault. These include the East Footwall, the Hook, and the BJ veins.

The mineralized veins of the Project are typical “Silver Belt” veins, and are composed of siderite, quartz, and various sulfides including pyrite, tetrahedrite, chalcopyrite, arsenopyrite and galena. Most of the silver is found within the tetrahedrite, which is argentiferous (silver bearing) throughout the district. It generally contains between 2% and 6% silver by weight. Substantial amounts of silver are also recovered from galena. In some silver mines of the district, chalcopyrite has contributed recoverable copper. Hershey (1916) believed supergene processes have enhanced the silver grades of some of the oxidized mill feed. Secondary oxide minerals that have been noted historically include cerargyrite, native silver, cerrusite, malachite, cuprite, argentite, chalcocite, and pyromorphite.

Primary hydrothermal zoning within veins has not been demonstrated in the district, with the exception of two veins in which the iron content of the sphalerite varied with depth. In one vein, the iron content increased with depth, and in the other, it decreased with depth (Fryklund, 1964). Mineral zoning has been described in historical reports. This observation was later explained by the superimposition of mineral concentrations formed by several mineralizing pulses, rather than the result of a single hydrothermal fluid evolving in composition as it traveled upward.

Both the Alhambra and South Veins are partly oxidized. The Alhambra fault and vein zone displays a normal oxidation pattern from surface to a depth of 250 to 300 ft below surface. The South vein has a 200 to 400 ft wide zone of oxidation plunging down-dip, parallel to the mineralization. The oxidation fluids appear to be following the mineralized portion of the structure, or may be an indication of an intersecting fault.

Mineralized zones in the district generally have more vertical than lateral extent. Historic stoping in the Project suggests that the higher-grade mineralization has vertical: lateral ratios between 2.5:1 and 4:1. The mineralized zones plunge down-dip, parallel to the shear lineation developed during the same deformation that brackets the mineralization.

1.3 Status of Exploration Development and Operations At the time of publication, no exploration or mining activities were in progress. Test mining and underground development ceased in September 2012. The most recent exploration drilling program was completed in 2011.



1.4 Mineral Processing and Metallurgical Testing In June of 2011, USC commissioned G&T Metallurgical Services, Ltd. (G&T) of Kamloops, British Columbia, to complete metallurgical test work on an 84 kg sample of South Vein mill feed obtained from the 3950 sublevel of the Crescent mine. Nearby I-drift face samples reported grades between about 25 and 100 oz/t silver in the South vein. The test work program consisted of the following:

Chemical content; Material hardness; Rougher flotation tests; and Cleaner flotation tests.

G&T provided a report of the test results (G&T, 2011).

In April of 2011, a bulk sample of about 123 t of South vein material from the 3950 sublevel was processed in the old (pre-expansion) 100 t/d circuit at the New Jersey mill in Kellogg, Idaho. The bulk sample was mined soon after development of the first I-drift (3950 Level) on the South Vein began. Face sample assay results encountered assays from 25 to over 100 oz/t. The material was processed over a period of four days, and the objectives of the bulk sampling program were:

Determine the silver head grade and the quantity of other metals present; Obtain metallurgical data for expanded mill design; Obtain metallurgical data for economic assessment; Evaluate different flotation reagents; and Determine if a marketable concentrate could be produced.

A report by William C. Rust, consulting metallurgist, summarized the results of the bulk sample mill test (Rust, 2011).

The 84 kg sample sent to G&T averaged 24.5 oz/t silver; the bulk sample, 10.55 oz/t. G&T completed a bond ball mill work index test to assess hardness of the Crescent South vein, and classified the South vein material as moderately hard with a Bond work index test result of 13.3 kWh/t. The actual Bond Work Index during bulk sample milling average 9.65 kWh/t (Rust, 2012). Therefore, the actual work index is about 75% of what was predicted from G&T’s test work.

The primary grind of G&T’s rougher flotation tests varied from 80% passing 65 µm to about 211 µm. G&T reported that a primary grind of about 80% passing 97 µm (97 µm P80) had the best metallurgical performance with 92% silver recovery into a rougher concentrate with 7% of the mass balance reporting to the rougher concentrate. G&T performed a single batch cleaner test using the rougher concentrate from the material that had a primary grind of 97 µm P80 and additional Z-55 collector was added in the cleaner test. The rougher concentrate was cleaned in three stages of sequential cleaning without regrinding. This cleaner test produced a final concentrate of 691 oz/t (23,700 g/t) silver with 83% silver recovery.

During test milling of the 123 t bulk sample, several flotation reagents were evaluated and particle size from the grinding circuit was monitored. The recovery of silver in the final concentrate from test milling was 89.5%, and 1.6% of the total mass, including gangue minerals, went to the concentrate. The ratio of concentration for silver was approximately 63 to 1.



The results of these two testing programs were used to design the NJMC mill expansion. As of September 2013, the expanded mill has processed over 12,000 t. After the commissioning period, from August 1st through September 3rd when the mill was operating 24 hours per day, recoveries averaged 94%. This recovery indicates that the expanded mill is achieving the design goals.

Based on the metallurgical and recovery results obtained milling more than 12,000 t Crescent mill feed, silver recoveries in sulfide mineralization are expected to exceed 94%. However, recoveries decline to between 80% and 84% when the mineralized material is oxidized. This is typical for other mines in the district. Mine plans call for the mining below the known elevation oxidization during the early years of production to ensure excellent silver recoveries. This will provide time for metallurgical testing to be completed on the oxidized material to see if recoveries and, by extension, the economics of oxidized material can be improved. Controlling dilution and maintaining the grade milled in the 15 to 16 oz/t range will ensure that economic extraction will continue.

Penalties for arsenic and antimony in the concentrates are expected to stay within acceptable limits.

Based on results from milling 12,000 t of material to date, no significant factors are known or expected to affect silver recovery.

1.5 Mineral Resource Estimate SRK estimated silver grades using inverse distance weighted (IDW) to the second power for each of the three geologically controlled individual vein wireframes. Estimation was carried out on declustered full-vein-with composites using a two-pass search, with increasingly expanded search distances. In addition to IDW metal grades, the estimation runs stored average distance to composites, number of composites and number of drillholes used for the block estimate for classification. A second grade estimation routine was conducted to store nearest neighbour (NN) grades for use in model validation.

This 2013 resource estimate is informed by a larger drill database and the addition of over 2,300 ft of production data from recent development drifting compared to the 2012 resource estimate. USC also provided a more extensive database of density determinations, which resulted in a minor increase in the average density.

The updated mineral resource for the Project is presented in Table 1.5.1. Resources stated in this Technical Report are in situ and are not constrained to a mining configuration. A cut-off grade (CoG) of 8 oz/t was applied based on the economic evaluation of the Project in this study. The CoG for the resource was determined using a silver price of US$20.00/oz, a recovery of 92%, combined mining and processing costs of US$133.00 per run-of-mine (RoM) ton and a 2% net smelter return (NSR) royalty.



Table 1.5.1: Mineral Resource Statement for the Crescent Silver Project, SRK Consulting (U.S.) Inc., July 22, 2013

Alhambra South Vein Total Ag Cut-off

Mass Grade Cont. Metal



8.0 oz/t (t) (Ag oz/t) (Ag Moz) (t) (Ag oz/t) (Ag Moz) (t) (Ag oz/t) (Ag Moz)

Measured 9,000 13.2 0.1 51,000 17.2 0.9 60,000 16.6 1.0 Indicated 143,000 13.4 1.9 317,000 14.7 4.7 461,000 14.3 6.6 Measured and Indicated 152,000 13.2 2,0 368,000 15.0 4.5 520,000 14.4 7.5

Inferred 118,000 10.2 1.2 152,000 18.4 2.8 530,000 16.2 8.6 Source: SRK Notes:

Mineral resources that are not mineral reserves do not have demonstrated economic viability. No mineral reserves have been defined. The CoG for mineralized zone interpretation was 4 oz/t. The block CoG for defining Mineral Resources was 8 oz/t. A silver price used was US$20/oz, mining and processing costs of US$100/mill feed ton, and 92% mill recovery

were used to define the 8 oz/t cut-off. A 2% NSR royalty associated with the Project was applied in metal value calculations; The resources reported above are non-diluted. Measured Resources required blocks to be informed by a minimum of 8 composites and those blocks must be less

than 120 ft from previous production. Indicated Resources required blocks to be informed by composites from a minimum of two drillholes and distance

from data less than 300 ft The resources mined from the intermediate drifts have been deleted from the 2013 updated resources. Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate and

numbers may not add due to rounding.

1.6 Mineral Reserve Estimate Mineral Reserves have not been estimated for the Project, and are not appropriate at this stage of Project development. No formal engineering or economic work that would enable identification of reserves has yet been carried out.

1.7 Mining Methods The Project currently consists of an existing office and shop area. There are numerous old mine working on the property. There are two main development drifts that will be used in the mining operation:

The Big Creek #4 Cross Cut (BC#4) is a trackless drift that is been completed to a length of 3,500 ft. The drift will be extended by 4,050 ft. Once the drift is completed, rail will be installed and the BC#4 will be the main access to the mine workings and used for mill feed haulage to the surface.

The Countess Decline is presently 3,710 ft long at a grade of minus 15% to 17%. It will be extended as a spiral ramp 3,330 ft, when it breaks into the BC#4. The Countess is situated between the South and Alhambra Veins and will be the primary access to these veins.

The Alhambra and South Veins are steeply dipping, narrow and have hanging and footwalls that require support subsequent to mining. A modified cut and fill method has been chosen for mining in the Crescent Mine. In this method, as the mill feed is mined it is replaced with a waste fill. The fill provides support for the hanging and foot walls of the stope.



The modification to the cut and fill method that has been chosen for use in the Crescent Mine is to use a mining technique called resuing that separates mill feed and waste by individually blasting them and also mucking them separately. At the Crescent Mine the resuing practice will be modified by leaving the waste in the stope to use as a fill material and to provide a work platform from which to mine the next cut. An added benefit of this system is that the bulk of the waste rock, which is required to be mined to allow working room, does not have to be transported to surface. The Alhambra Vein, being narrower, will be mined using a “captive” stoping method where the broken mill feed will be transported to chutes with a slusher, than hauled by rail or rubber tire equipment to the main mill feed pass.

The South Vein true thickness averages 3.5 ft. Additional waste taken from the footwall will provide additional working room that will allow the use of underground loaders. The use of such equipment will allow for a higher production rate than at the Alhambra Vein. After the initial drift on the mill feed is complete, the mill feed above the drift will be drilled with up-holes and blasted and removed separately from the waste. The waste will then be blasted and allowed to remain on the floor as backfill and as a working platform.

1.8 Recovery Methods In early 2011, USC and New Jersey Mining Company (NJMC) formed a joint venture (NJMC 65.2% and USC 34.8%) to expand and operate the New Jersey mill capacity to 440 t/d (18.4 t/h). Construction started in the spring of 2011 and was completed in June of 2012. The mill feed is crushed to -3 inches in a jaw crusher. Material over 2 inches is screened off and sent to a cone crusher in a closed loop. Screen undersize drops into a 385 t fine mill feed bin. The fine mill feed bin discharges onto the bin discharge belt conveyor which, in turn, discharges onto the ball mill feed belt conveyor. Lime is added to the mill feed to adjust pH. The ball mill is a 250 kW Marcy-type ball mill that is 8 ft in diameter and 13 ft long. The ball mill discharge is pumped to a hydro-cyclone where cyclone underflow reports back to the mill for additional grinding, and cyclone overflow is conveyed via gravity to the rougher flotation cells.

Flotation reagents are fed into the rougher flotation circuit via metering pumps. The rougher circuit consists of one Wemco 144 and five Wemco 66D flotation cells. Rougher concentrate is fed to two banks of three Wemco 36 cleaner flotation cells operated in series for two stages of cleaning. Concentrate from the second stage of cleaner cells forms the final concentrate which is piped to a 5 ft diameter thickener where it is then transferred to agitated tanks for storage.

The final concentrate is either pumped into a concrete mixing truck for delivery to a local refinery or filtered with a plate and frame pressure filter and placed in one-ton supersacks for delivery to a foreign smelter.

Tailings from the rougher flotation circuit are pumped into two Deep Cone Thickeners (DCT) operated in parallel to produce paste tailings. Thickener overflow is piped to a process water storage tank. Underflow from the paste thickeners is pumped to a paste tailings stack east about 500 ft east of the mill building.

1.9 Project Infrastructure The Crescent is an old, well established mine that was operated between approximately 1910 and 1986. Following GFN’s purchase of the Project in late 2006, surface exploration and intermittent



underground work resumed in 2007 and continued under USC’s direction from December 2009, when USC entered into an option agreement to acquire an 80% interest in the Project. There is currently a small staff in place which handles the care and maintenance of the mine. The existing infrastructure, mining equipment, and related support facilities are in good condition. The mine is accessed on a paved road maintained by the county.

The existing buildings are dedicated for offices, shops, and storage. They are in good condition and can be used for future mine operations. Power and water provided by public utilities are available on site and are adequate for the planned operation. Compressed air and electric power at both the Countess Decline and Big Creek #4 are currently supplied by diesel generators and compressors. Before resuming operations, power and air lines will be run from the mine shop area to the Big Creek #4 portal, replacing the diesel powered units. Once the Countess and BC#4 connect, utilities for the whole mine will be through the BC#4 feed. One additional electric powered compressor will be added at the mine shop.

Industrial water supplies are obtained from the capture of naturally occurring ground water inside the mine. Excess water produced from the mine is pumped to the Sunshine Mine, where it is treated and discharged under an agreement with Sunshine.

The mine office and ancillary facilities are adjacent to the Hooper Tunnel, and located just above the valley bottom. Access from the office area to the Big Creek # 4 tunnel is along approximately 5,600 ft of two lane dirt road to reach a location about 500 ft above the Hooper Tunnel. This dirt road continues approximately 17,850 ft to the Countess decline located about 1,500 ft above the Hooper Tunnel. The company maintains the dirt road year around.

1.10 Environmental Studies and Permitting The mine and mill are fully permitted and all permits are in compliance. Because the mine is on patented mining claims (privately-owned land), only a limited number of environmental permits are required for mining and milling operations.

The Crescent Mine and New Jersey Mill are located within the Bunker Hill Superfund site (EPA National Priorities Listing IDD048340921). While cleanup activities are proceeding within the Superfund area, the EPA has decided that active mining sites operating under state and federal permits are not to be targets for investigation or cleanup as long as they remain in compliance with their permits. The EPA has, however, targeted mining companies to pay for remedial cleanup of sites with arbitration presently ongoing. The state of Idaho’s position is that unless the owner was an active participant in the release, they will be held harmless for remediation (Schuld, 2010). USC was not a participant in the release of contaminated material. The mine and mill are fully permitted and all permits are in compliance. Because the mine is on patented mining claims (private land) only environmental permits are required for mining and milling operations.

A Modified Phase I Environmental Assessment Report dated February 9, 2007 was prepared for USC (LFR Inc., 2007). LFR notes five “Recognized Environmental Conditions” as follows:

1. Location within a Superfund-Designated Area. 2. Listing within the EPA “remedial investigation/feasibility study” for the presence of adit

drainage, upland waste rock, and surface disturbance with potential for erosion. 3. Surface Water Discharge Contaminant Contributions and NPDES Permitting.



4. Upland Soil and Waste Rock Contamination. 5. Underground Contaminant Sources and Ground Water Impacts.

The consensus of this evaluation is that USC faces the possibility of being required to remediate features related to historic mining activity on the property. The EPA has modified its approach to assessing active mining sites for in the scope of remedial actions based on input from the public and the State of Idaho. Active mining sites operating under state and federal permits are not to be targets for investigation or clean up as long as they remain in compliance with their permits.

1.11 Capital and Operating Costs Capital costs used in this study were based on both vendor quotes and cost estimates for underground development. SRK has added 25% contingency to these quotes and estimates for omissions. Similarly, operating costs, which are driven by consumables and labor rates, were supported by recent, relevant vendor information, recent labor rates at the mine or public domain mining services costs providers, typically Infomine.

Capital costs in the preproduction period and after production begins are outlined in Tables 1.11.1 and 1.11.2.

Table 1.11.1: Preproduction Mine Development Cost

Description Cost US$(000s)

Underground Mine Development $4,120 Mine Equipment $2,102 Underground Communications $111 Utilities/Electrical $132 Exploration Cost $650 Contingency 25% $1,779

Preproduction Capital $8,894

Table 1.11.2: Post Development Capital

Description Cost US$(000s)

Underground Mine Development $2,668 Laboratory $209 Rail for BC#4 $157 Sustaining Capital $160 Mine Closure $405 Other $481 Contingency 25% $1,020

Total Post Development Capital $5,099

Mine operating cost and productivities, based on tons of mill feed produced, including waste removal, are outlined in Table 1.11.3.



Table 1.11.3: Mine Productivities and Unit Costs

Description Productivity Unit Rate Alhambra Vein Development 7.5 ft/day $252 US$/ft Raise Development 39 t/d $45,000 US$/ea Alhambra Mill Feed Production 52 t/d $54 US$/t I drift South Vein 43 t/d $217 US$/ft Attack Ramp Wedge Allowance $3.50 US$/t South Vein Mill Feed Production 50 t/d $ 61 US$/t Underground Mill Feed Transport Allowance $2.00 US$/t Mine G&A Allowance $500,000 US$/yr

Other cost factors include:

Mill feed haul from mine to mill – during development: US$7.00/t processed; Mill feed haul from mine to mill – during production: US$3.00/t processed; Mill feed processing cost: US$23.50/t processed; and G&A Cost: US$2,198,000/year.

1.12 Economic Analysis The indicative economic results are shown on Table 1.12.1. The following provide the basis of the SRK LoM plan and economics:

A mine life of 6 years; An overall average process recovery rate of 92%; Post Development Capital of US$5.1 million will be required; A site operation cash cost of US$146.43/t processed; Royalty applied at 2.0% of Net Smeter Return; Mining Tax applied at a 1% rate; State Tax applied at a 7.6% rate; Federal Tax applied at a 35% rate; Depletion applied at a 15% rate of adjusted Gross Income: and A tax loss carry forward of US$2.67 million was included.

Financial assumptions used are summarized in Table 1.12.2 and 1.12.3.

Table 1.12.2: Market Parameters

Description Value Unit Market Price: Silver $20.00 US$/oz Smelter Payment 95.0%Refining $2.75 US$/oz Smelting Charge $750.00 US$/t-conc Transportation $0.00 Included Insurance 0.05%Royalty 2.00%



Table 1.12.3: Production Parameters

Description Value Unit Mine Life 6 years RoM Mined (undiluted) 451 kt RoM Processed (diluted) 601 kt Recovered Ag 6,108 koz Payable Ag 5,803 koz

Project economic results and estimated cash costs are summarized in Table 1.12.4.

Table 1.12.4: Economic Results

Description Value units Production Summary

RoM Mined (undiluted) 451 kt RoM Processed (diluted) 601 kt

Estimate of Cash Flow Gross Income $122,165 US$/t-process

Refining ($26,775) $44.58 Gross Revenue $95,390

Royalty ($1,157) $1.93 Net Revenue $93,722

Operating Costs Mining $33,084 $55.08

Processing $16,074 $26.76 G&A $10,864 $18.09

Total Operating $60,022 $99.93 Site Operation Cash Cost $87,954 $146.43

Operating Margin $34,211 Initial Capital $8,894

Post Development Capital $5,099 Income Tax $0

Cash Flow Available for Debt Service $13,865 NPV 8% $8,575

IRR 32%

Table 1.12.5 provides sensitivity analysis of Project economics using alternative metal prices.

Table 1.12.5: Sensitivity Analysis of NPV @ 8% (US$000)

Description -20% -10% Base 10% 20%Silver Price 16.00 18.00 20.00 22.00 24.00 Revenues (9,859) (633) 8,575 17,188 23,324Capital Costs 11,222 9,898 8,575 7,251 5,927 Operating Costs 17,092 13,484 8,575 3,665 1,245



1.13 Conclusions and Recommendations SRK finds no material impediments to the development of the Project. The State of Idaho has permitting and closure regulations that are amenable to mining, and the Silver Valley, where the Project is located, has over 100 years of profitable mining production history. The USC staff is experienced and capable of managing the re-start of underground operations.

At PEA level, inferred resources have been used to evaluate potential mineral economics. Reserves have not yet been stated for the Project. SRK has supported USC in preparing the necessary drilling and sampling programs designed to upgrade Inferred resources. There is a strong potential to expand existing resources with additional drilling, especially at structural intersections where the newly identified Jackson Vein intersects the Alhambra and the South Vein.

From an economic perspective, the Project benefits from previous test mining in 2011/12, which provided essential information about the nature of mineralization, mining methods, processing and costs. These data underpin the economic evaluation of this PEA. Additionally, the Company benefits from a loss carry forward, which offsets most of the tax burden. The areas of notable project risk include: 1) some drill gaps in the resource; 2) proving out the mining methods proposed to minimize mining dilution; and, 3) securing surface rights for waste placement.

The resource gap is the part of the resource on the South Vein between the Countess decline and the Hooper adit. With some additional planned drilling, the location grade and thickness of the vein can be confirmed.

USC has proposed a modified cut and fill mining method, with resuing of mill feed and waste. Dilution has been built into the mine plans at rate of 25% for the South Vein and 50% for the Alhambra Vein. Achieving these dilution targets is crucial for the profitability of the Project. Proposed mining and backfill methods are innovative. It will be critical, during initial re-start, to closely track production metrics to ensure that the proposed mining throughput (400 t/d) can be met.

At present, USC has not secured the surface rights to add waste rock to the North American Dump. SRK’s understanding is that negotiations toward this end are in progress and there are alternative locations for waste placement that could be pursued.

1.13.1 Recommended Work Programs and Costs

Additional work recommended to advance the Crescent Project consists of the following key elements:

Confirmation drilling on the South Vein; Metallurgical testing of mineralization in oxide; Underground grade control laboratory; Test mining to confirm mining methods and economics; and Hydrogeochemical Studies.

The total cost estimate for additional work is US$1,735,000 and is expected to be completed by the end of the third quarter of 2014 assuming test mining can be financed and initiated by the end of 2013. At that point the mine would go into full production. The work program is summarized in Table 1.13.1.1



Table 1.13.1.1: Recommended Pre-Production Work Program Costs

Work Element Estimated Cost US$ Assumptions/Comments Underground Confirmation Core Drilling 650,000 20 holes to 540 ft @ US$60/ft Metallurgical Testwork of Oxide 100,000 Underground Grade Control Lab 60,000 Test mining (Year 1) 825,000 Does not include cost recovery from productionHydrogeochemical Studies 100,000 Total US$ 1,735,000



2 Introduction

2.1 Terms of Reference and Purpose of the Report This report was prepared as a National Instrument 43-101 (NI 43-101) Technical Report, Preliminary Economic Assessment (PEA) for United Silver Corp. (USC) on their Crescent Silver Project (Crescent, or the Project) located in Kellogg, Idaho, USA. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in SRK’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by USC subject to the terms and conditions of its contract with SRK and relevant securities legislation. The contract permits USC to file this report as a Technical Report with Canadian securities regulatory authorities pursuant to NI 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. The responsibility for this disclosure remains with USC. The user of this document should ensure that this is the most recent Technical Report for the property as it is not valid if a new Technical Report has been issued.

This report provides mineral resource estimates, and a classification of resources prepared in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves: Definitions and Guidelines, November 27, 2010.

This report follows and supports a press release issued by USC on July 29, 2013 entitled “United Silver Corp. Releases Positive Preliminary Economic Assessment and Reports Updated Estimated Resources” in which the Company provided an updated resource estimate and a preliminary projection of project economics.

2.2 Qualifications of Consultants (SRK) The Consultants preparing this technical report are specialists in the fields of geology, exploration, mineral resource and mineral reserve estimation and classification, underground mining, geotechnical, environmental, permitting, metallurgical testing, mineral processing, processing design, capital and operating cost estimation, and mineral economics.

None of the Consultants or any associates employed in the preparation of this report has any beneficial interest in USC. The Consultants are not insiders, associates, or affiliates of USC. The results of this Technical Report are not dependent upon any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings between USC and the Consultants. The Consultants are being paid a fee for their work in accordance with normal professional consulting practice.

The following individuals, by virtue of their education, experience and professional association, are considered Qualified Persons (QP) as defined in the NI 43-101 standard, for this report, and are members in good standing of appropriate professional institutions. The QP’s are responsible for specific sections as follows:



J. B. Pennington, MSc, C.P.G., Nevada Mining Practice Leader, is the QP responsible for geology and resources Sections 2-12, 14, and portions of Sections 1, 25 and 26 summarized therefrom, of this Technical Report.

Kent Hartley, P.E., Principal Mining Engineer, is the QP responsible for metallurgy and engineering Sections 13, and 15-24, and portions of Sections 1, 25 and 26 summarized therefrom, of this Technical Report.

2.3 Details of Inspection Table 2.3.1 details the site visits conducted by the QPs.

Table 2.3.1: Site Visit Participants

Personnel Company Expertise Date(s) of Visit Details of Inspection Kent Hartley SRK Mining Engineering August 8-9, 2012 Review of mining methods Jay Pennington SRK Geology April 22-23, 2013 Review of sample material and QA/QC data Kent Hartley SRK Mining Engineering April 22-23, 2013 Mining cost validation

2.4 Sources of Information The sources of information include data and reports supplied by USC personnel as well as documents cited throughout the report and referenced in Section 27.

2.5 Effective Date The effective date of this report is July 22, 2013.

2.6 Units of Measure The US System for weights and units has been used throughout this report, unless otherwise stated. Tons are reported in short tons of 2,000 lb. All currency is in U.S. dollars (US$) unless otherwise stated.



3 Reliance on Other Experts USC has employed a technical team of resource scientists, mining engineers and cost engineers to develop the Crescent Project. SRK’s opinion contained herein is based on information provided by USC throughout the course of the investigations.

SRK used their experience to determine if the information from previous reports was suitable for inclusion in this technical report and adjusted information that required amending. This report includes technical information, which required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, SRK does not consider them to be material.

The QPs have fully relied upon USC for information pertaining to mineral tenure, surface rights, property ownership including water rights, permitting and environmental studies, social issues, taxation and marketing. USC’s sources of information are cited in text and listed in the References section. These items have not been independently reviewed by SRK and SRK did not seek an independent legal opinion of these items.

3.1 Economic Analysis The QPs have relied on Valerie Obie, Principal Mineral Economist, of SRK Consulting, for information presented in Section 22, Economic Analysis.

3.2 Mining Methods The QPs have sought the opinion of an independent underground mining consultant, William Walker of Mine Development Analysis, Inc., for information presented in Section 16, Mining Methods. Mr. Walker has made several visits to the Project with the QP for mining engineering, and they have worked together to develop the mining methods presented in this report.



4 Property Description and Location

4.1 Property Location The Project is located in Shoshone County, Idaho at 116.074349° West longitude and 47.505195° North latitude, in the Idaho Panhandle. The Project is in Idaho Township 48 North, Range 03 East, Sections 15, 16, 17, 18, 19 and 20. It is located about 4 miles southeast of Kellogg, Idaho, and about 75 miles east of Spokane, Washington. The general Project location is shown in Figure 4.1.1, and a map of the Project’s land position is presented in Figure 4.1.2.

4.2 Mineral Titles and Surface Rights On December 12, 2006, Gold Finder Explorations, Ltd. (GFN, formerly SNS Silver Corp.) acquired the right to purchase the Project from Shoshone County, Idaho for US$650,000. The transaction was completed on January 2, 2007. On July 13, 2007, two claim fractions named Queen Lode, lying within section 16 were purchased from The New Bunker Hill Mining Company (NBH). These were fractions lying within the contiguous property boundaries. On November 28, 2007, GFN acquired the surface rights to a 25.69 ac parcel north of, and contiguous with, the surface rights already owned on the Diana and Hiawatha claims.

On December 30, 2009, USC signed an Earn-In Agreement with GFN to earn up to an 80% interest in the Crescent Project. On June 1, 2011, USC and GFN jointly announced (USC, 2011) that USC had earned an 80% interest in the Crescent Project through the expenditure of US$9 million in exploration work. The Project, at the time the Earn-In Agreement was signed, consisted of 25 patented mining claims or portions thereof totaling 346.190ac, and surface rights totaling 40.639 ac. A 25.69 ac portion of the surface rights was inadvertently omitted from the Earn-In Agreement and so the surface rights detailed in the Agreement consist only of the 14.949 ac parcel that comprises the surface rights of the Diana and Hiawatha claims. The surface rights for the North American Dump area are in the parcel omitted. USC is working to correct this deficiency, to use the North American Dump area for waste rock storage.

In June 2010, during the Earn-In Option period, USC purchased the surface rights to a 10.01 ac parcel of the Countess patented mining claim, in order to construct a portal to begin underground development. This surface parcel was contributed to the Project property package per the Earn-In Agreement. In December 2010, USC purchased the mineral rights for an additional 42 patented mining claims or portions thereof from NBH. These claims consisted of two blocks, one block contiguous on the northeast of the original Project, and the other block contiguous on the west. The purchase from NBH brought the total number of patented mining claims or portions thereof to their current number, 64, comprising 897.581 ac (Figure 4.2.1).

The mineral rights of the Project are all in the form of patented claims. These claims have been modified over the decades by agreements between previous owners and the adjoining property owners that established inclined or horizontal boundaries in some places. These atypical boundaries do not affect the area of the current resource (Figure 4.2.2).

Patented claims typically convey full surface ownership, but the surface rights of the Project claims were severed from the underlying mineral rights in 1991 by a previous owner, and sold to Idaho



Forest Industries, Inc., with the exception of two claims that cover the area of the Hooper portal and associated mine buildings.

The current surface rights of the Project form two separate parcels. The first is a contiguous, irregularly-shaped parcel of ground totaling 14.949 ac. This parcel consists of surface rights to the patented Hiawatha claim (11.331 ac) and Diana claim (3.618 ac); it covers the area of the Hooper portal and dump. The second parcel consists of a parallelogram-shaped block of ground measuring 10.01 ac within the Countess claim, and covering the area of the Countess portal and dump (Figure 4.2.3).

The Hiawatha and Diana claims have never had their surface rights severed from the mineral rights, but they are part of an area affected by a 1958 agreement with the Sunshine Mine which conveys the mineral rights below 2,691 ft amsl to Sunshine. The 10.01 acre surface rights block on the Countess claim overlies mineral rights controlled by the Project.

The remaining surface rights on the patented mineral claims that comprise the Project are largely with sections 15, 16, 17 and 18, and are controlled by Stimson Lumber Co., while an area of about 92 ac within sections 19 and 20 is controlled by Silver Mountain Corp. and by Silver Mountain Ventures LLC. The surface rights extend 100 ft vertically below the current topography, below which, mineral rights are controlled by USC.

The claims and surface rights parcels are listed in Table 4.2.1.



Table 4.2.1: Patented Mining Claims and Surface Ownership

Parcel Number Patent Number

Claim Names Instrument/Deed # Ownership Type

48N03E-17-7400 MS 3013 Countess 470526 Surface MC 0265 MS 3475 A-1, A-2 and A-3 470526 Mineral Interest MC 0274 MS 2129 Rebel, Redeemer, Ruby, Skyline 470526 Mineral Interest MC 0274 MS 2204 Grant and McArthur 470526 Mineral Interest MC 0274 MS 2207 Blue Jay, Crescent, Empire, Jackson and Monte Christo 470526 Mineral Interest MC 0274 MS 2274 Homestake, Jupiter, Old Sol and Yellow Jacket 470526 Mineral Interest MC 0274 MS 3013 Countess 470526 Mineral Interest MC 0274 MS 3014 King 470526 Mineral Interest MC 0274 MS 3015 Queen 470526 Mineral Interest MC 0274 MS 3185 Sumner 470526 Mineral Interest MC 0274 MS 3217 Duke 470526 Mineral Interest MC 0499 MS 2274 Hornet 470526 Mineral Interest MC 0529 MS 2274 Artic, Chicago, New York, Surprise 470526 Mineral Interest MC 0533 MS 2129 Hiawatha 470526 Surface and Mineral Interest MC 0533 MS 3185 Diana 470526 Surface and Mineral Interest MC 0588 MS 2129 Hiawatha (surface and mineral), Rebel (mineral only) 470526 Surface and Mineral Interest MC 0588 MS 3185 Diana (surface and mineral), Sumner (mineral only) 470526 Surface and Mineral Interest MC 0629 MS 1035 Milo 470526 Mineral Interest MC 0629 MS 1347 Alhambra 470526 Mineral Interest MC 0629 MS 1348 Fanny May 470526 Mineral Interest MC 0629 MS 1349 Bonnie Jean, Dawn Fraction, Iuka, Lucky Chance, Midday, No. 1, Protection, Tough Nut 470526 Mineral Interest MC 0629 MS 1359 McKinley 470526 Mineral Interest MC 0629 MS 2203 Bonanza King 470526 Mineral Interest MC 0629 MS 2204 Funston, Get There, Junction, Last Chance, Lucky Chance/Lucky 470526 Mineral Interest MC 0629 MS 2274 Atlas, Globe, Hornet, Long Green, OK, Timothy, Yellow Jacket 470526 Mineral Interest MC 0629 MS 2611 A1, A2 470526 Mineral Interest MC 0629 MS 2870 Standard 470526 Mineral Interest MC 0629 MS 3013 Count 470526 Mineral Interest MC 0629 MS 3503 Badger, Governor, Number 2, Scott, Winfield 470526 Mineral Interest MC 0630 MS 2274 Artic, Banner Hill, Surprise, Surprise Fraction 470526 Mineral Interest MC 0631 MS 2274 Banner Hill, Surprise Fraction 470526 Mineral Interest



4.2.1 Nature and Extent of Issuer’s Interest

Surface rights: USC owns 10.01 ac at the Countess portal as surface rights, and 14.949 ac at the Hooper portal as patented claims.

Mineral Rights: USC controls the mineral rights of 64 patented mining claims. On all but two of these claims, the surface rights were severed and sold off by a former owner in 1991 (see section 4.2 above). These mineral rights form a contiguous block with two small internal areas controlled by other entities. The first of these is a 7.3 ac wedge-shaped parcel in the eastern half of the property, owned by the state of Idaho. The second is a 4.3 ac rectangular-shaped parcel in the western half of the property controlled by the BLM (Figure 4.2.1). Neither of these two internal parcels affects the current resource area. The sub-surface mineral claims controlled by USC fully encompass the resource model area.

The property is accessed via public roads, or over land controlled by USC. As USC maintains control of surface rights discussed above, it will also maintain legal surface access to the mine infrastructure.

4.3 Royalties, Agreements and Encumbrances There are underlying agreements made by The Bunker Hill Company (BH), former owner of the Project, with owners of adjoining properties. The rights conferred by these agreements were transferred with the changes in ownership, and are considered to still be in effect as noted in a legal opinion dated May 14, 2007 written by Michael E. Regan, Coeur d’Alene, Idaho (Stoel Rives LLP 2007).

The first of these are three different 1958 Agreements that re-define property boundaries between the Crescent Property and adjoining properties. These agreements re-defined originally vertical boundaries to become inclined boundaries to the Crescent property (and the adjoining properties) such that the mineralization within the Alhambra, East Footwall, Hook and BJ veins below the 3,000 ft elevation remain under the ownership of the owner of the Crescent Property. These agreements also convey extralateral rights to the owners of adjoining properties for other veins, not considered important to the Crescent Property. None of these Agreements affects the area of the current resource.

Another Agreement concerns a 1973 mine water discharge agreement with Sunshine Mining Company (Sunshine). By this Agreement, mine water effluent from the Crescent Mine is discharged into the Sunshine water discharge line and from there is discharged to Sunshine’s tailings impoundment and water treatment facility. The Agreement, dated April 16, 1973, provides that effluent up to 150 gpm from the Crescent Property will be transported and treated for a period of fifty (50) years, until 2023, through Sunshine’s discharge lines to Sunshine mine’s tailings impoundment area. The terms of this agreement contain the normal successors and assigns clause such that the Agreement passes through to any new owners on either side of the Agreement during the 50-year term of the Agreement.

On December 14, 2010, USC announced in a press release the acquisition of the mineral rights to 42 patented mining claims contiguous to the Crescent Property. The claims were purchased from the New Bunker Hill Company (NBH) for US$1,250,000 plus a 1% NSR. The 1% NSR can be bought



back at any time prior to December 2014. As noted above, GFN is also entitled to a 2% NSR on these 42 patented mineral claims.

In a joint venture agreement with New Jersey Mining Company (NJMC) dated January 7, 2011, USC agreed to pay for the expansion of New Jersey’s mill to 400 tonnes (441 t) per day and in return acquire a one-third interest in the mill. The NJMC mill is fully permitted, including tailings storage, and located less than four miles from the Crescent Mine Property. USC is guaranteed 300 tonnes (330 t) per day capacity for the life of the mill, any excess capacity if available, and access to the tailings to use for backfill. The mill expansion was completed in June 2012.

In May 2012, USC acquired GFN’s 20% interest in the Project (USC, 2012). In exchange for GFN receiving a 2% NSR on both the original 25 patented mineral claims plus two claim fractions that USC acquired from GFN plus the 42 patented mineral claims USC purchased from NBH, GFN paid USC US$400,000. Half of this was in cash and the other half to be deducted from future royalty payments, if any should be made. As defined in the Settlement Agreement, royalty payments, if any, will be paid after commercial production begins, commercial production being defined therein as beginning after the first 100,000 tonnes (110,230 t) of material has been mined and milled.

On June 26, 2012, USC formed the Crescent Mine LLC (CM LLC) with Hale Capital Partners (Hale). USC then put the Project, including the 42 adjacent patented mining claims, into CM LLC. Under the terms of the LLC, Hale made a net initial contribution of US$2,462,704 in exchange for a 20% interest in CM LLC. At any time after USC has satisfied its obligation to Hale under a previous convertible debt financing, USC may acquire one-half of Hale’s interest in CM LLC for a payment equal to the initial contribution.

4.4 Environmental Liabilities and Permitting Environmental Liabilities

The Crescent Mine and New Jersey Mill are located within the Bunker Hill Superfund site (EPA National Priorities Listing IDD048340921). While cleanup activities are proceeding within the Superfund area, the EPA has indicated that active mining sites, operating under state and federal permits, are not to be targets for investigation or cleanup so long as they remain in compliance with their permits. The State of Idaho maintains the position that, unless the current owner of a mining operation was an active participant in the Superfund-designated release(s), they will be held harmless for remediation (Schuld, 2010). USC was not a participant in the release of contaminated material.

The first cleanup actions under the Superfund designation occurred in 1987 and then resumed in 1989, continuing to the present. Contaminants, including lead, arsenic, zinc, and cadmium, from historical mining in the Silver Valley area were found in Coeur d’Alene Lake and the Spokane River system, extending as far as 60 miles downstream. An estimated 62 Mt of mine wastes were dumped directly into the South Fork of the Coeur d’Alene River and its tributaries between 1884 and 1968. Cleanup actions are determined by the Records of Decision (ROD) issued by the EPA in a public involvement process. Currently, the RODs are divided into three cleanup areas called “Operable Units”:



Operable Unit 1 - populated areas of the Bunker Hill Box in a one-square-mile area (addressed first because of the high levels of contaminants in soil and in blood lead levels in humans that were tested);

Operable Unit 2 - non-populated, non-residential areas; and Operable Unit 3 - areas of mining-related contamination in the South Fork and main Coeur

d’Alene River watersheds, including Coeur d'Alene Lake and the Spokane River to Upriver Dam in Washington State.

The Crescent Mine and New Jersey Mill are located within Operable Unit 3 (OU3).

In 2010, the EPA issued a draft Focused Feasibility Study (FFS) for the Upper Basin of OU3. This Upper Basin FFS focused on developing remedies for identified areas of concern in the drainage of the South Fork of the Coeur d’Alene River stretching from Mullan to Kingston, within the Silver Valley. The draft FFS identified, by name and location, both active and historic mining sites or features and allocated specific remedies to these sites/features. When the draft FFS was shared with the public, it was met with widespread criticism for assigning expensive remedies to active mining sites that are currently operating under active state and federal permits, and to many historical sites with little or no factual information on site conditions.

As a result of this overwhelming public response, EPA has modified its approach to assessing mining sites for inclusion in the scope of remedial actions. Active mining sites operating under state and federal permits are now considered as contingency sites and are not to be targets for investigation or clean up, so long as they remain in compliance with their permits (Hydrometrics, Inc., 2011).

Separate from EPA’s Bunker Hill Superfund cleanup, a Modified Phase I Environmental Assessment Report for the site, dated February 9, 2007, was prepared for GFN (LFR Inc., 2007). In that report, LFR noted five “Recognized Environmental Conditions,” including:

1. Location within a Superfund-Designated Area. 2. Listing within the EPA “remedial investigation/feasibility study” for the presence of adit

drainage, upland waste rock, and surface disturbance with potential for erosion. 3. Surface Water Discharge Contaminant Contributions and NPDES Permitting. 4. Upland Soil and Waste Rock Contamination. 5. Underground Contaminant Sources and Ground Water Impacts.

In the end, the area-wide encumbrances of the EPA Superfund cleanup will not affect the ability of USC to mine and process material from the Project, but may remain a longer-term risk of future liability should the EPA reverse course and again pursue existing operations to shoulder the burden of the cleanup efforts.

4.4.1 Required Permits and Status

The mine and mill are fully permitted and all permits are in compliance with state and federal regulations. The permits are listed in Table 4.4.1.1. Because the mine is on patented mining claims (privately-owned land), only a limited number of environmental permits and authorizations are required for mining and milling operations. For the most part, these permits are all issued by the State of Idaho.



Table 4.4.1.1: Summary of Permits

Permit CommentsMine Water Rights License - UG No Water Rights License is required because the mine is located on

Patented Mining Claims (Private Property) Water Rights License - Surface No Current Permit – These permits are obtained on an annual and as

needed basis. Water Discharge Permit Water is discharged under Sunshine Mine NPDES Permit ID#-000006-0.

Permitted under a 1973 Agreement with Sunshine Silver Mines Corp. Stormwater Discharge Permit #-IDR050000 (Permit Tracking # -IDR05CA75). Explosives Permit ATF # 9-ID-079-33-4K-00329 Reclamation Bond None Required Mill Cyanidation Permit Idaho Permit # CN-000027 This permit includes the reclamation plan Reclamation Bond No bond required Permit to Appropriate Water Ground water permit # 94-07509 Stormwater Discharge Permit # IDR05A383

4.5 Other Significant Factors and Risks SRK is not aware of any other significant factors or risks associated with the proposed mine development at this site.

Crescent Silver Project,

Kellogg, Idaho

Figure 4.1.1

Project Location Map (large scale) Source: USC, 2013


Kellogg, Idaho

Figure 4.1.2

Crescent Mine Map (small scale) Source: USC, 2013


Kellogg, Idaho

Figure 4.2.1

Mineral Rights Map Source: SRK, 2013


Kellogg, Idaho

Figure 4.2.2

Atypical Mineral Claim Boundaries Map Source: SRK, 2013


Kellogg, Idaho

Figure 4.2.3

Surface Rights Map Source: SRK, 2013



5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Topography, Elevation and Vegetation The property topography is characterized as foothills of rugged mountains cut by the narrow valley of the Big Creek drainage. The property elevations range from 2,500 to 5,600 ft above sea level. Most of the property is covered with a mixed conifer forest including hemlock and Douglas fir, with moderate to thick underbrush where most recently logged. Outcrop on the property is limited because of extensive vegetation and the development of a soil horizon.

The elevation at the portal of the Hooper Tunnel, historically the main haulage level, is 2,690 ft. The Big Creek #4 Tunnel, which will become another main haulage level for the Project, is at an elevation of 3,220 ft at the Portal.

5.2 Accessibility and Transportation to the Property Access from Spokane, Washington follows Interstate 90 east for 75 mi to the Big Creek, Idaho exit, #54, and then proceeds south on Big Creek Road for 2 mi to the Project. Spokane has an international airport. The nearest port is Seattle, Washington, which is 360 mi away by Federal interstate highway. Rail service is no longer provided to the Valley.

From the Big Creek exit, the Big Creek Road, a paved county highway, extends south-southwest for 2.3mi. This road also accesses the Sunshine Mine, adjacent to the Project, and is the main access to several forest and lumber roads in the area. The road is considered a good all-weather condition road, and in winter is regularly plowed by the county.

The executive offices for USC are located on Big Creek Road at 1.2 miles from Interstate 90, and the turn-off for the Hooper Tunnel dump and mine offices is 2.1 miles from the interstate. A short, 0.15 mi paved driveway connects the Big Creek Road to the gravel/dirt parking lot adjacent to the mine offices, shops, warehouse and core shed of the Project.

Also at the 2.1 mi mark is the turn-off for a dirt road that serves as the main access to the portals of the Big Creek #4 (1.1 mi from Big Creek Road) and the Countess (3.8 mi from Big Creek Road). Most of the length of this access is owned by the state and by Stimson Lumber Co., with only one short segment occurring on Project surface rights ground, but USC has right-of-way for its entire length.

The road to the Countess portal will see light duty traffic. With completion of the secondary entrance to the mine, the Big Creek #4 will become the main haulage level, and thus the road to the Big Creek #4 portal will become a haul road. If deemed necessary, year-round road maintenance will be the responsibility of the mine operation.

In addition to the access to mine workings, USC has right-of-way permission to use the extensive network of logging roads above the Project’s mineral rights for exploration purposes, as deeded when a prior owner of the patented claims sold the surface rights to Idaho Forest Industries, Inc., a predecessor of Stimson Lumber Co.



A network of four-wheel drive gravel surface roads in good condition accesses the surface exploration area.

The underground workings are accessed via three development levels, namely the Hooper Tunnel, the Countess decline and the Big Creek #4 Tunnel.

5.3 Climate and Length of Operating Season The property is located in the northern Rockies and has average summer high temperatures of 82° F and average winter low temperatures of 22° F (Table 5.3.1). The average rainfall in Kellogg, Idaho (elevation approximately 2,300 ft amsl) is about 31 inches per year and the mean average snowfall is 54 inches per year (Western Regional Climate Center, 2009). Surface drilling has been conducted from May through November. Historically, the Project operated year-round.

Table 5.3.1: Summary of Meteorological Data

Heading Heading Average Summer High 82° F (28° C) Average Winter Low 22° F (-6° C) Average rain fall (Kellogg) 31 inches per year Average snow fall (Kellogg) 54 inches per year

5.4 Sufficiency of Surface Rights The surface rights held by USC at the time of this report were sufficient for underground mining, but insufficient for waste storage and mill feed stockpiles. USC was negotiating to acquire the necessary surface rights.

5.5 Infrastructure Availability and Sources The Coeur d’Alene mining district has had continuous mine production for the past 130 years, and supports local businesses that cater to the mining industry, including fabricators, suppliers, contractors and technical services. The towns within the Silver Valley have services including restaurants, hotels, hardware stores, grocery stores, schools, and other facilities. A community hospital with emergency care is located in Kellogg, Idaho.

5.5.1 Power

The net energy demand for the Project is estimated to be less than one MW. The main power users are the ventilation system, fresh and operational water pumping station(s), dewatering, compressed air station, shops, and lighting (underground and surface). Power is supplied off the regional grid. There is an 11kVA line, part of Avista’s transmission system, near the Project.

Table 5.5.1.1 presents the estimated underground electrical requirements for the mine at maximum production rate.



Table 5.5.1.1: Underground Electrical Power Estimate

Item Number of Units

Unit power (kW)

Connected power (kW)

Load Factor

Load (kW)

Utilization Factor

Energy/month (kW hrs/month)

Vein Runner II drill 1 55 55 80% 44 60% 19,008 Main fan 2 100 200 80% 160 100% 115,200 Aux fans 6 30 180 70% 126 100% 90,720 Shop 1 25 25 50% 13 50% 4,500 Lighting 1 25 25 90% 23 100% 16,200 Diamond drill 2 56 112 90% 101 70% 50,803 Subtotal 13 597 467 296,431 Misc. allowance (20%) 120 59,286 Total 717 355,717

The Project site has a 2,000 kVA substation which is located near the portal of the Hooper Tunnel. The capacity of this substation is sufficient for all current and estimated future surface and underground activities. A new power line will need to be installed from the substation to the Big Creek #4 Portal. Costs are included in the capital cost.

5.5.2 Water

The main source of fresh potable water to the mine offices is from the Central Shoshone County Water District well at Enaville, from which it is pumped up the valley of the South Fork Coeur d’Alene River, and thence up Big Creek.

Water sufficient for development mining will be sourced from a sump in the Countess decline. The minimum natural in-flow to the Countess workings is about 40 gpm, with higher amounts during spring run-off. This source is expected to meet the needs of planned exploration drilling, development, and production.

Water for processing has been established and permitted for use in the New Jersey Mill.

On the Hooper level, water for development and exploration drilling will be sourced from a diamond drillhole.

5.5.3 Mining Personnel

The Silver Valley is an established mining community with two producing mines at present. A sufficient workforce of experienced underground miners is locally available from the nearby communities of Kellogg (population 2,395) and Wallace (population 960).

5.5.4 Potential Tailings Storage Areas

There are no tailings storage areas located on the Project site and no potential locations for tailings disposal within the current surface ownership. The New Jersey mill site, of which USC holds a 34.8% interest, has sufficient area to contain tailings for approximately 5 Mt of production, before potential utilization of tailings for backfill purposes, and expansion of mill infrastructure.

5.5.5 Potential Waste Disposal Areas

The mine currently has two waste dumps, the Countess and the North American dumps. Waste from exploration drifting by both GFN and USC has been placed on these dumps. There is ample capacity



to accommodate near term underground development waste material. The surface rights to the North American dump were inadvertently omitted from the Earn-In Agreement and subsequent Settlement Agreement with GFN. The North American dump is planned to be used as a waste dump for material from the Big Creek #4 Tunnel. USC is working to resolve the issue, and if an agreement with GFN does not occur, surface rights to a suitable portion of ground overlying the Project’s mineral rights will be purchased from the Stimson Lumber Co. USC plans to initiate contingency discussions with Stimson should additional surface rights be required.

5.5.6 Potential Heap Leach Pad Areas

The Crescent Mine will not require a heap leach pad for mineral processing.

5.5.7 Potential Processing Plant Sites

There are no potential locations for construction of a concentrator within the current surface ownership. The New Jersey Mill, of which USC holds a 34.8% interest, has sufficient capacity for processing material from the Project.



6 History

6.1 Prior Ownership and Ownership Changes Three historic claim groups located sometime prior to 1908 underlie the current Project. These are the Alhambra group, the Crescent group, and the Big Creek group.

The Big Creek group was located around the turn of the 20th Century, and was the subject of a property dispute in 1908 between Big Creek Mining Company and North American Mining Company. This dispute was resolved in 1911 with Big Creek holding ownership.

"During the period prior to 1922, the claims were owned by the Big Creek Mining Co. Ltd. In 1922,

the mine workings from Big Creek No. 3 level to surface were leased to George Kinmouth and

Associates. In the 1920’s the Big Creek Mining Company went into receivership and the Bunker Hill

and Sullivan Mining Company purchased the property at Sheriffs auction.” (Radford, 1985)

BH purchased the adjacent Crescent group in 1926 and the Alhambra group in 1937, thus consolidating the three historic claim groups under one owner (Baldry, 1981). By the time the Bunker Hill Mine closed in 1981, BH had amassed some 620 mining claims on and around the Bunker Hill Mine in a large contiguous area, which included on its eastern end all of the mining claims that underlie the current Project.

In November 1982, the Bunker Hill Limited Partnership (BHLP) purchased the Kellogg mining property of the bankrupt BH, including the large mining claim block. The upper workings in the Crescent, from the Bud Level to the surface, were leased from 1984 to 1990 to Intermountain Mineral Engineers, Inc. In 1991, most of the surface rights of the claims in the current Project were severed from the mineral rights, and sold to Idaho Forest Industries, Inc. Shortly thereafter, BHLP defaulted on debt to Fausett International, Inc. (Fausett), and a block of claims that Fausett had secured as collateral for the debt, were deeded to Fausett in June, 1992. This block, comprised of the old Crescent and Big Creek claim groups, became the core property of the Project as acquired by GFN in January 2007. Meanwhile, the Bunker Hill Mine and the large remaining block of claims within which it sits, including the Alhambra group, were acquired from a bankrupt BHLP by NBH in 1992.

The Crescent Mine was leased by Fausett for a period to Royal Silver Mines, Inc., and then in September 2001, Shoshone County took possession of the mine from Fausett in lieu of payment of back taxes.

At some point between 1991 and 2007, the surface rights of the claims in the current Project were acquired by Stimson Lumber Co. from Idaho Forest Industries, Inc.

On January 2, 2007, GFN purchased the Project from Shoshone County, Idaho for US$650,000. On July 13, 2007, two claim fractions named Queen Lode, lying within section 16, were purchased from NBH. These were fractions lying within the otherwise contiguous property boundaries, and had inadvertently been omitted from the 1992 Fausett deed. In November 2007, a 25.69 ac parcel of surface rights was purchased by GFN from Stimson Lumber Co.; this parcel covers the area of the North American dump. The North American parcel is contiguous with the Hiawatha/Diana surface rights parcel that covers the area of the Hooper and Big Creek #4 portals, and that was part of the



original GFN purchase in January. Together, the three purchases comprised the Project at the time of the Earn-In Agreement with USC dated 30 December 2009.

6.2 Previous Exploration and Development Results

6.2.1 Historical Development

Julihn and Horton (1936) reported that "Development work was begun on the Alhambra claim group by 1893”. This group was explored by several workings from the Elk Creek drainage, including upper and lower Alhambra tunnels. The lower tunnel drifted east on the Alhambra fault to the property line with the Crescent group sometime prior to 1913.

Development on the Big Creek group began in about 1916 by four main tunnels from the Big Creek drainage. By 1922, mill feed was being produced from these tunnels, known as the "Anderson Mine". The (lower) Alhambra tunnel was connected to the Big Creek No. 4 tunnel prior to 1928.

Bunker Hill embarked on major development projects after acquisition of the Crescent and Big Creek claims in 1926. Between January 1929 and February 1930, the 5,000 ft long Hooper tunnel was driven from surface on the Big Creek drainage to, and then along, the Alhambra fault. By the time WWII temporarily closed the mine in 1942, the Ellis Shaft had been sunk from the Hooper level to the 1200 Level (1,200 ft below the Hooper Level), and the Hooper raise was driven from the Hooper level to the Alhambra Tunnel level 550 ft above.

After the mine reopened in 1951, the Ellis Shaft was deepened to the 3100 Level. The Yreka United crosscut was driven over 2 mi from the 2300 Level in the Bunker Hill Mine to connect with the Crescent 3100 Level. The Crescent No. 2 shaft was sunk vertically from the 3100 Level to below the 4100 Level. The 4300 Level, the bottom level in the mine, was developed by a decline with rubber-tired equipment from the 4100 Level. Figure 6.2.1.1 is a longitudinal section showing the historic development within the Project with respect to the nearby Bunker Hill and Sunshine Mines.

Documentation about development work between 1982 when BHLP acquired the property and 1991 when BHLP filed for bankruptcy is very poor but it appears that little development work was completed by BHLP. No exploration was conducted from the time BHLP filed for bankruptcy until after GFN acquired the Project in 2007.

GFN first rehabilitated the Hooper Tunnel, which was in generally good condition, but required ventilating plus timber and track repair. The rehab work was followed by extension of the track drift westward along the Alhambra Fault for a distance of 1,000 ft. The rehab and the new drifting were completed during 2008.

GFN began rehabilitation of the Big Creek #4 Tunnel in early 2008. A 10 ft culvert was installed at the portal for ground support, and several caved sections of the tunnel in the area of old workings required spiling to rehabilitate. By August 2008, the level had been rehabilitated from the portal to the top of the Hooper Raise.

6.2.2 Bunker Hill 1942-1985

Existing drill logs indicate that Bunker Hill, owner of the Crescent Mine, drilled 167 holes totaling 54,734 ft during the period 1942 to 1985. All of the holes were core holes and none of the core is extant. Records consist of drill logs and maps showing drillhole traces. The logs are annotated with



assay results and the bearing and azimuth at the collar; some of the holes have downhole survey information. Most of the drilling was done in the lower mine, below 1,000 ft elevation, and distant from current operations. The assay data is historical in nature, and has not been used in resource estimates by SRK, GFN or USC.

6.2.3 GFN 2007-2008

During 2007 and 2008, GFN drilled a total of 104,084 ft in 37 underground and 70 surface diamond drillholes, nearly all targeting the Alhambra and South Veins. Most of this drilling is described in the 2010 Technical Report (SRK, 2010). After the cut-off date for the 2010 report, five holes were drilled from the Hooper level to test for the up-dip projection of the East Footwall mill feed shoot, and two holes were drilled from the Hooper to test the Alhambra and its immediate footwall between the 400 and 800 levels. All of the GFN drilling is described in Section 8 of this report because it remains relevant to the current resource model.

6.3 Historic Mineral Resource and Reserve Estimates

6.3.1 Bunker Hill Mineral Resource and Reserve Estimates

Bunker Hill calculated reserves on an annual basis, using industry-accepted standards of the period, through 1981. Between 1982 and 1985 under the BHLP ownership, reserves were not calculated regularly, but a formal, independent reserve estimation was completed in 1983 (Springer, 1983), and an in-house “informal review” was done in 1986 (Dalton, 1986). Dalton estimates a 1986 “proven and probable” reserve of 112,756 t at 25.3 oz/t Ag containing a total of 2.8 Moz. In addition, three other categories of mineralization (possible, contingent and projected) are also reported for an additional 356,722 t at 16.8 oz/t Ag containing 6.0 Moz. This historical reserve is located primarily in the lower levels of the mine, which are currently flooded. USC has no intentions at this time to dewater these levels and establish access to these areas. These historical reserve estimates are not NI 43-101

compliant and are presented for historical information only and as such, should not be relied

upon. These historic reserves are presented solely as an indication of the potential

magnitude of exploration targets within the Project. A qualified person has not done

sufficient work to classify the historical estimate as a current resource estimate or mineral

reserve.

6.3.2 SRK Mineral Resource Estimate

In 2008, USC retained SRK to complete a mineral resource estimate. This was reported most recently in “Amended NI 43-101 Technical Report on Resources, United Silver Corp., Crescent Mine, Kellogg, Idaho”, by SRK Consulting, originally effective May 7, 2010, and amended September 30, 2011. The mineral resource estimate, as summarized below, was endorsed by Bart Stryhas, PhD, CPG, a Qualified Person according to the tests in Section 1.4 of the NI 43-101. This is a NI 43-101 compliant resource.

Mineral Resource Statement and Sensitivity

The 2009 Crescent Project mineral resource statement is presented in Table 6.3.2.1. An 11.0 oz/t CoG was chosen for resource reporting based on cost data and CoG used at the nearby Sunshine Silver Mine. The results reported in the resource statement have been rounded to reflect the approximation of grade and quantity, which can be achieved at this level of resource estimation.



Table 6.3.2.1: 2009 Crescent Mine Mineral Resource Statement (at 11.0 oz/t CoG)

Resource Category Total Tons (kt) Ag oz/t Average Grade Contained Ag (Moz)

Indicated 324 18.7 6.1 Inferred 211 19.5 4.1

The grade tonnage distributions of the Indicated and Inferred Mineral Resources at the Project are presented in Tables 6.3.2.2 and 6.3.2.3, and reflect the resource sensitivity to variation in CoG.

Table 6.3.2.2: 2009 Crescent Mine Indicated Mineral Resource Sensitivity

Ag oz/t Cut-off Total Tons(kt) Ag oz/t Grade Contained Ag (Moz)8 408 16.9 6.9 9 390 17.2 6.7

10 358 17.9 6.4 11 324 18.7 6.1 12 295 19.4 5.7 13 268 20.1 5.4 14 247 20.7 5.1 15 215 21.6 4.6

Table 6.3.2.3: 2009 Crescent Mine Inferred Mineral Resource Sensitivity

Ag oz/t Cut-off Total Tons(kt) Ag oz/t Grade Contained Ag (Moz)8 498 13.6 6.7 9 403 14.8 6.0

10 247 18.1 4.5 11 211 19.5 4.1 12 182 20.7 3.8 13 153 22.3 3.4 14 133 23.6 3.1 15 123 24.4 3.0

The mineral resources described above constitute contained metal in the ground and had not been included in any formal plan of exploitation at the time of the report. There were no known material issues related to environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues which could have affected the mineral resources. Additionally, there were no known material issues related to mining, metallurgy, infrastructure and other relevant issues, which may have affected the mineral resources.

A prefeasibility study is required to demonstrate the economic merit of mineral resources in order for their conversion to reserves. At the time, no such study had been completed and therefore, the Project had no Reserves when the 2010 SRK report was completed.

6.3.3 Historic Production

Production in the Big Creek, or “Anderson” Mine, appears to have been from several veins, one along the Alhambra fault (the "fault vein"), and two in the hangingwall of the Alhambra. A 1922 internal memo refers to on-going production from two stopes, the "silver stope", and the "lead stope". This document notes that 1,199 t of material containing 33,488 lb of lead and 128,078 oz silver (1.4% Pb and 107 oz/t Ag) was produced from above the No. 3 tunnel level (Anderson, 1922). A series of "Crescent Mine" memos from H. M. Childs to Stanley Easton, dated 1928 and 1929, reports



progress on stoping and drifting, including the East Bud, West Bud, and Brooks stopes. These reports also mention the "Old Taylor Stope”.

After stoping on the Brooks and Bud stopes was exhausted, four levels were developed from the Ellis Shaft below the Hooper level. Because of narrow widths and low grades, only a limited amount of production came from these lower levels before the mine was closed in 1942.

Most of the documented production occurred in the period 1953 to 1981, with all or nearly all production coming from the lower mine (below an elevation of 1,000 ft). In 1951, the mine was reopened and the Ellis Shaft sunk to the 3,300 level (elevation -560 ft). Drifting on the 3100 level located several mill feed shoots that were mined up- and down-dip, between the 2500 level and the 4300 level.

A 1982 BH Compilation Report states that the Project produced a total of 25,139,655 oz of Ag from 978,750 t of material at an average grade of 27.0 oz/t Ag between the period 1917 and 1981. Approximately 700,000 t of this material was mined from the lower part of the mine.

In 1980, BH excavated 42,564 t from the Big Creek No.3, No.4 and North American dumps. This material averaged 2.05 oz/t Ag, and yielded 67,767 oz of Ag (Radford, 2009; Radford, 1985; Bunker Hill Company, 1980).

Very limited production occurred from 1983 until operations were halted in June 1986. None of the areas of historic production are located within the areas of the current resource estimation of this report.

6.3.4 Historic Metallurgy

A 120 t/d concentrator operated on the property prior to closure in 1942. A typical concentrate analysis from an average feed grade of 25 oz/t Ag is as follows:

250 oz/t Ag; 29% Fe; 32% S; 6% Pb; 0.002% Bi; 4.5% As; 4% Sb; 8% Cu; 0% Zn; and 4% insoluble.

“The sulphide ores from the Crescent are well-adapted to flotation; a recovery of 95 percent of the silver is made when sulphides alone are treated, but with mixed sulphides and oxides the recovery ranges from 80 to 85 percent. As the recovery in milling oxidized ores is relatively poor, high-grade ore from the oxidized zone, averaging 100 oz silver per ton, is sorted out in the stopes and shipped directly to the Bunker Hill and Sullivan smelter.” (Julihn and Horton 1936).

GFN sent a 194 t sample of material stockpiled from the extension drifting along the Alhambra on the Hooper level to the Sunshine mill, operated by Sterling Mining Co., in 2008. The test milling achieved



an 86% recovery of silver. Sterling’s metallurgist noted that because of a low lead content, the lead concentrate from the test contained a high percentage of pyrite.


Kellogg, Idaho

Figure 6.2.1.1

Long Section Showing Historic Development Source: SRK, 2013



7 Geological Setting and Mineralization

7.1 Regional Geology The regional geology is dominated by Precambrian sedimentary rocks of the Belt-Purcell Supergroup (Belt), which have been strongly deformed during the Cretaceous age Sevier Orogeny. This regional deformation has resulted in large-scale folds cut by numerous west to northwest striking faults and veins. A geologic map of the Coeur d’Alene district in Idaho and Montana is shown in Figure 7.1.1.

During the Middle Proterozoic, the area was dominated by a large intra-cratonic basin that was subsiding along syn-sedimentary faults. The basin sediments range in age from about 1,470 Ma to 1,400 Ma and are composed of medium-tofine-grained clastic and carbonate-bearing clastic rocks. The oldest exposed unit is the Prichard Formation, a deep-water argillite/siltite unit up to 12,000 ft thick. The Ravalli Group, consisting of quartzites and siltites up to 8,000 ft thick and deposited in a shallow water environment, overlies this. These are overlain by the Middle Belt Carbonate, comprised of shallow water dolomitic quartzites and arenaceous dolomites up to 6,500 ft thick. The youngest sediments are the Missoula Group, a shallow water sequence of inter-bedded quartzite and argillite up to 1,500 ft thick.

These sediments are believed to have remained relatively stable until approximately 1,350 Ma when portions of the basin were affected by compressional tectonics of the East Kootenay Orogeny. This orogeny was followed by rifting of the basin during the late Proterozoic-early Paleozoic when the western portion of the basin was transported away, and the western margin of North America was developed.

The next major tectonic event occurred during the Cretaceous Sevier Orogeny. Early compressional tectonics dominated the area forming large-scale folds, reverse and thrust faults. Many of these structures were focused along the west-northwest trending Lewis and Clark Line. This is a regional, deep-seated lineament believed to represent an intra-plate boundary, which has been recurrently active since the Proterozoic. During the late Cretaceous, the Bitterroot Lobe of the Idaho Batholith was emplaced to the south, accompanied by dike emplacement in this area and normal movement along earlier reverse faults. The major mineralizing event is believed to have occurred during the compressional phase of the Sevier Orogeny. The most recent tectonic activity is believed to have occurred during the Tertiary when the Lewis and Clark lineament was reactivated along the Osburn Fault. This event resulted in 16 miles of right lateral, strike slip movement that has dissected and displaced many of the deposits in the region (Hobbs et al. 1965; Lewis et al., 2002).

7.2 Local Geology Lithology

The Coeur d'Alene mining district is hosted by lightly-metamorphosed sediments of the Belt-Purcell Supergroup. Figure 7.2.1 is a geologic map of the mine area, and Figure 7.2.2 is a geologic cross-section. West of the district, in eastern Washington, the sediments accumulated to a thickness of 55,000 ft or more at the line along which the Belt basin was pulled apart.

The Prichard Formation is the lowermost exposed unit of the Belt, and is sub-divided into Lower and Upper units. In the Coeur d’Alene mining district, the Lower part is composed of thin to thick bedded,



medium gray argillite and quartzose argillite, laminated in part with some discontinuous quartzite zones. Pyrite is abundant. The Upper part is comprised of inter-bedded medium-gray argillite and quartzose argillite and light gray impure to pure quartzite. The base of the Prichard is not exposed.

Above the Prichard lie the Ravalli Group rocks. These are subdivided into the Burke, Revett, and St. Regis Formations. The Burke and Revett are characterized by thick-bedded quartzites interbedded with siltite-argillites, while the St. Regis is composed of siltite-argillite. The St. Regis is characterized by abundant mud cracks and mud chips, and by its purple color.

The Wallace Formation is the lowermost unit of the Middle Belt Carbonate in the district; it overlies the Ravalli Group and is sub-divided into an Upper and Lower part. The Lower Wallace is composed of gray and greenish gray siltite-argillites and argillites. The Wallace is distinguished from the St. Regis by the presence of dolomitic silt and the lack of purple color, as well as by more subtle differences in bedforms. The Upper Wallace is medium to greenish-gray finely laminated argillite. It contains some arenaceous dolomite and impure quartzite and minor gray dolomite and limestone.

The Missoula Group is locally represented by the Striped Peak Formation. This unit includes inter-bedded quartzite and argillite with some arenaceous quartzite. The Striped Peak Formation is not exposed at the Project area.

Regional igneous rocks are represented by large granitic batholiths of Cretaceous age. The Bitterroot Lobe of the Idaho Batholith is located to the south, and the Kaniksu Batholith is located to the northwest. Within the district, a series of small monzonitic stocks were intruded coeval to batholith development. Igneous dikes of many compositions including lamprophyre, diabase, and diorite, occur in insignificant volumes. Some of the dikes are associated with the monzonitic stocks, but most are of unknown age (Hobbs et al 1965).

Alteration and Mineralization

Wall rock alteration associated with veining consists of changes in carbonate mineralogy plus sulfidation and silicification. Ankerite is the typical disseminated carbonate phase found in unaltered wall rock. Zoning is discernible with siderite adjacent to the vein, grading outward to a zone of mixed calcite and ankerite +/- siderite, to distal ankerite, only. The siderite alteration may be found for tens to hundreds of feet from the veins.

Pyritization of wall rocks is locally strong, and takes the form of fine-grained disseminated grains, and streaks of coarse grains. Silicification occurs as flooding and veining. Pyritization and silicification are favorable indications of higher-grade mineralization but not necessarily correlative.

Bleached halos occur around veins that cut purple sediments because of destruction of the hematite by hydrothermal fluids penetrating the wall rock adjoining the veins. The St. Regis is characterized by purple and maroon colors that result from finely disseminated earthy hematite. Subordinate volumes of the Revett are colored by hematite as well. The bleaching is not evident in rocks that were originally green, but a line of euhedral pyrite is often found at the bleaching front, no matter what the original rock color was. The bleaching front in purple rocks is often “peppered” by very fine octahedral magnetite (White, 1998b; Strand, 2002).

The Silver Belt can be described as a corridor of structural preparation parallel to the Lewis and Clark Line and south of the Osburn Fault. The mineralization of economic value here is confined to



discrete veins and veinlets of siderite, quartz and tetrahedrite. The tetrahedrite typically fills a fine fracture pattern within siderite and very rarely occurs as disseminated grains in wall rock.

Galena occurs within the siderite and quartz veins as isolated grains or fine- to coarse-grained streaks. It also occurs as disseminated zones in quartzose wall rocks described locally as “blue rock” which may contain up to a few percent lead. Most of the galena mineralization exhibits crosscutting relations that dates it later than the siderite-tetrahedrite veining. District wide, the galena contains a baseline content of 0.12% silver present as a coupled substitution with antimony in the galena structure.

Structure

The Coeur d'Alene district hosts structural deformation features related to compressional, extensional and trans-current movement. The district is located at the intersection of two predominant linear belts of deformation. The Lewis and Clark Line (LCL) is a 30mi-wide zone of tectonism that extends more than 200 mi from western Idaho into western Montana. This regional tectonic zone trends west to northwest and is a deep-seated lineament believed to represent an intra-plate boundary, which has been recurrently active since the Proterozoic era. The Noxon Line is a north to northwest trending structural high that is partly defined by thinning of Belt stratigraphy. This feature extends from southern British Columbia south to the LCL.

White (1998a) identified five deformation events within the district listed in chronologic order as:

1. Folds trending west-northwest. 2. Folds trending north. 3. Reverse faulting. 4. Normal faulting along earlier reverse faults. 5. Right lateral strike slip faulting mainly represented by the Osburn Fault.

The west- to northwest-trending folds are interpreted to be related to deformation along the LCL. A younger and separate tectonic event is responsible for the north trending folds. This are generally more pronounced north of the district within the Noxon Line. The structural deformation within the Coeur d’Alene Mining District is dominated by the tectonic fabric of the LCL. The reverse movement is seen predominantly along west-northwest striking faults. White (1998b) postulated that the faults and veins developed along metamorphic shear fabric related to west northwest-trending reverse faults. These same fabrics appear to have been re-activated during an extensional event producing normal movement. The final deformation was right lateral displacement along the Osburn Fault, which displaces the mineralization. This displacement is the only transcurrent movement documented well in the district and resulted in approximately 16 mi of offset (White 1998b).

The large productive veins of the Silver Belt strike west-northwest and dip steeply to the south. They occur along major faults or as "links" between these. A less common style of veining is splay structures that occur at fault bends in the hangingwall or footwall. These zones are typically short in strike length, but very high grade.



7.3 Property Geology Lithology

The Belt rocks host all of the mineralized veins at the Project. The three Belt units present are Wallace, St. Regis and Revett Formations. Detailed lithologic mapping has not yet been completed to know exactly where the St. Regis/Revett formation boundary lies with respect to the mineralized veins. The interfingered facies character of the boundary has complicated efforts to pinpoint its location.

Mineralization

At the Crescent, silver mineralization has been mined on the Alhambra Fault from near surface to 1,500 ft below mean sea level, a vertical distance of about 5,000 ft. In the upper workings of the mine, above the 1,500 ft elevation, the higher grade mineralization occurs on the immediate hangingwall of the fault (Julihn and Horton, 1936), where the vein cuts Revett and St. Regis Formations. The zones mined between mean sea level and -1,500 ft elevation occur on the immediate footwall in the Revett Formation. The other significant production from the mine has been from a set of veins found within the footwall of the Alhambra Fault. These include the East Footwall, the Hook, and the BJ veins.

The mineralized veins of the Project are typical “Silver Belt” veins, and are composed of siderite, quartz, and various sulfides including pyrite, tetrahedrite, chalcopyrite, arsenopyrite and galena. Most of the silver is found within the tetrahedrite, which is argentiferous (silver bearing) throughout the district. It generally contains between 2% and 6% silver by weight. Substantial amounts of silver are also recovered from galena. In some silver mines of the district, chalcopyrite has contributed recoverable copper.

Hershey (1916) believed supergene processes have enhanced the silver grades of some of the oxidized mill feed. Secondary oxide minerals that have been noted historically include cerargyrite, native silver, cerrusite, malachite, cuprite, argentite, chalcocite, and pyromorphite.

Primary hydrothermal zoning within veins has not been demonstrated in the district, with the exception of two veins in which the iron content of the sphalerite varied with depth. In one vein, the iron content increased with depth, and in the other, it decreased with depth (Fryklund and Weis, 1964).

Mineral zoning has been described in historical reports. This observation was later explained by the superimposition of mineral concentrations formed by several mineralizing pulses, rather than the result of a single hydrothermal fluid evolving in composition as it traveled upward.

Both the Alhambra and South Veins are partly oxidized. The Alhambra fault and vein zone displays a normal oxidation pattern from surface to a depth of 250 to 300 ft below surface. The South vein has a 200 to 400 ft wide zone of oxidation plunging down-dip, parallel to the mineralization. The oxidation fluids appear to be following the mineralized portion of the structure, or may be an indication of an intersecting fault.

Mineralized zones in the district generally have more vertical than lateral extent. Historic stoping in the Project suggests that the higher-grade mineralization has vertical: lateral ratios between 2.5:1 and 4:1. The mineralized zones plunge down-dip, parallel to the shear lineation developed during the same deformation that brackets the mineralization.



Structure

The predominant fault at the Project is the Alhambra Fault. This structure shows an estimated 3,000 ft of stratigraphic offset, such that the St. Regis units in the hangingwall are in fault contact with Wallace Formation in the footwall. Movement on the Alhambra appears to be largely post-mineral, with historic production occurring in the immediate hangingwall of the fault in the upper mine, and on the immediate footwall in the deep mine.

The South Fault is a generally east-west fault some 800 to 1,000 ft south of the Alhambra. The offset on this fault is unknown.

7.4 Significant Mineralized Zones The Project has been historically explored and/or exploited along five mineralized structures. These are the Alhambra and associated hangingwall veins, South fault, East Footwall, Hook and BJ. The current resource estimation includes only mineralization above the 2,000 ft elevation within the Alhambra and South Veins. The Alhambra vein is located in the immediate hangingwall of the Alhambra reverse fault above the 1,500 ft elevation, and in the immediate footwall below that elevation. The South fault is located within the hangingwall of the Alhambra Fault. The East Footwall, Hook, and BJ veins are all located in the footwall of the Alhambra Fault.


Kellogg, Idaho

Figure 7.1.1

Geologic Map of the Coeur d’Alene District Source: USGS

U.S. Department of the Interior U.S. Geological Survey Open-File Report 96-299


Kellogg, Idaho

Figure 7.2.1

Geologic Map of the Crescent Mine Area Source: Derkey, et al., 1996

Hobbs, S.W., Griggs, A.B., Wallace, R.E., and Campbell, A.B., 1965, Geology of the Coeur d'Alene district, Shoshone County, Idaho: U.S. Geological Survey Professional Paper 478.


Kellogg, Idaho

Figure 7.2.2

Geologic Cross-Section of Crescent Mine Area Source: USGS, 1965



8 Deposit Type

8.1 Mineral Deposit The mineralization is characterized by steep, narrow, mesothermal veins of metamorphic hydrothermal origin that contain economic concentrations of tetrahedrite in a gangue of quartz and siderite. The tetrahedrite is silver-bearing.

8.2 Geological Model Fleck et al. (2002) have documented that mineralization in the district was related to a complex metamorphic/hydrothermal event and that the source of metals was likely from scavenging within the Belt sediments. Wavra et al. (1994) demonstrated that silver mineralization in the Sunshine Mine occurred during a compressional, reverse slip tectonic/metamorphic event with higher grade shoots plunging steeply southwest. White (1998b) supports this observation and postulates that the faults and veins developed within zones of intensified shearing where mechanically generated heat caused dynamic metamorphism. Most evidence suggests that this mineralizing event occurred in the Late Cretaceous after major folding.

The district has a regional zonation defined by base and precious metals. The mines located north of the Osburn Fault, and in the southwestern part of the district, are typically dominated by galena and/or sphalerite mineralization. The mines in the southeastern part of the district are known as the "Silver Belt" and are dominated by tetrahedrite mineralization.

The Project is the westernmost of the Silver Belt mines, which, from east to west, include the Galena, the Coeur, the Coeur d'Alene, the Silver Summit, the Polaris and the Sunshine Mines. These mines are typified by siderite-tetrahedrite veins, with local galena zones.

By removing the post mineralization displacement along the Osburn Fault and reconstructing the district, a coherent zoning pattern is seen. This shows distinct zoning of siderite-tetrahedrite mineralization in the east, changing to galena and sphalerite mineralization in the west. To date, there is no plausible explanation as to why this zoning occurs.

Relevant Geological Controls

The west-northwest striking, steeply south dipping veins are the principal geologic control on mineralization. Many of these show disseminated sulfides in the adjacent wall rocks and for years, a lively debate was carried on as to whether mineralization within the veins formed from local mobilization of metals from the immediate wall rocks. Recently, the favored hypothesis is that the metals were mobilized from Belt sediments at depth and then transported by metamorphic/hydrothermal fluids to the veins (White 1998b, Fleck et al 2002).

The Prichard Formation regionally carries anomalous metal values, leading some observers to compare it to the Kupferschiefer of Europe. However, this unit is not an important host for Coeur d'Alene-type veins. Most of the production from the district has been from veins hosted within the overlying Ravalli Group rocks. A district wide compilation by Farmin (1975) attributes 75% of all metal production to veins hosted within the Revett and St. Regis Formations.

Historically, district exploration consisted of tracing major faults and shear zones, and exploring them where they cut favorable stratigraphy. Past geologists at the Project have considered the quartzites



of the Revett Formation to be the most favorable vein host, while geologists at the neighboring Sunshine Mine have considered the St. Regis Formation to be the most favorable vein host (Colson, 1958). At the nearby Bunker Hill lead/zinc mine, a systematic geologic research program in the 1970’s identified the thick, clean quartzite of the Upper Revett Formation as the most favorable vein host (White, 1977; Juras, 1977). At present, USC geologists at the Project consider both St. Regis and Revett to be favorable lithologies to host silver bearing veins.

Approximately 98% of the Silver Belt production has been from veins hosted within the St. Regis and Revett Formations. Defining the split between these two units is problematic for two reasons. First, lithologic labels have evolved through time. Older mapping, prior to the 1970’s, described all of the rocks of both formations as "quartzites", distinguishing them as "thick-bedded" or "thin-bedded". Second, the location of the Revett/St. Regis boundary, as determined by modern mapping, does not always agree with that identified in historic mapping. Much of the discrepancy in correlation is due to rapid facies changes particularly the rapid pinching out of individual quartzite beds. White and Winston (1977) point out that historically, all siltite-argillites were assumed to be St. Regis Formation; however, the quartzite dominated Revett Formation does contain significant amounts of siltite-argillite, which may be misinterpreted.

Determining the exact formation name of a given interval is of less importance than identifying the lithologic character of the rock. Examples of stratigraphic control to vein development are observed throughout the district. Near the Sunshine Mine, these relations suggest that larger, higher-grade veins favor either quartzite-dominant intervals or siltite-argillite-dominant intervals, but not inter-mixed intervals of both lithologies. Two other examples of stratigraphic control of vein development are found in the Lucky Friday Mine, in Mullan. The Lucky Friday Vein contains argentiferous galena and sphalerite over 5,000 ft of vertical extent. The vein is thickest and richest in the quartzites of the Upper and Lower Revett, but thins where it cuts the intervening siltite-argillite dominated Middle Revett. The nearby Silver Vein contains siderite and tetrahedrite, and cuts a similar stratigraphic interval, but displays opposite characteristics. It is thicker in the siltite-argillite and thinner in the quartzite. These observations suggest that rheological properties of the rocks may have varied during the differing pulses of mineralization.



9 Exploration

9.1 Relevant Exploration Work Development of underground workings comprised the recent exploration work completed at the Project. Exploration borehole drilling is described separately, in Section 10.

Drifts to access existing workings and for test mining along vein structures provided means for mapping and sampling. By accessing the mineral resource directly, the true geometry of mineralization was evaluated in a way that drillhole samples do not allow.

A mine geologist examined all underground development and sampled any potential mineralization. The mine surveyor located points along all drifts at about 4 ft above the floor to create as-built line work of the underground workings. USC geologists used the survey data to locate samples in space, and added the coordinates of each sample to the chip sample database.

9.2 Sampling Methods and Sample Quality Chip samples were routinely collected by the shift geologist from most working faces in the test I-drifts. As of January 1, 2013, there were 1,485 chip samples in the database. Most of these (1,370) were collected from the working faces of the I-drifts and were used for qualitative vein modeling. The remainder of this dataset is from rib or grab samples, or from structures encountered during development of ramps and laterals. Chip samples were intended to test for mineralization, and not to represent the entire rock mass shipped to the mill. In this respect, the samples collected are representative of recognized vein zones. Intervals between samples were assumed barren, based on visual assessment, and were assigned zero grade.

Chip samples were taken across each face of the test I-drifts, unless the face could not be accessed before the round was shot. In those instances, back samples were taken at the approximate face location. The faces were 4 to 10 ft apart along the drifts. The chip samples from working faces were taken at chest height if the mineralization at that height was representative of the entire face. Unmineralized intervals (i.e., barren of sulfide, and unoxidized) were often not sampled, and included in calculations as zero grade. This practice is a source of bias in the chip sample assay database, because intervals assigned zero grade may have had at least trace amounts of silver. If all intervals had been sampled and analyzed, grades of composited intervals may have been slightly higher. Conversely, samples with analytical values are higher grade than the average rock mass mined and milled.

All locations in the test drifts are referenced by distance from a centerline (CL), which is established as the middle of the access slot where it intersects the vein in the case of rubber-tired drifts, and from the center of the raise in the case of slusher I-drifts. A measure point (MP), is established along the rib as soon as possible after the floor is started, and the face is measured by pulling a tape from the MP to the face, and adding the distance from the CL to the MP. All width measurements are made in the horizontal plane, and height measurements are vertical.

The general sampling procedure was:

The sampler measured the distance to the face by pulling a tape from a measure point; The face was washed;



Spray paint was used to mark the sampling intervals so that like is sampled with like; The tickets were filled out with the date, heading identification, distance, sample width, drift

width, and description of sampled material; The samples were chipped, with sample numbering proceeding from footwall (north) to

hangingwall (south); Cloth sample bags were marked with the sample number and a numbered sample ticket was

placed in the bag with the sample; One or more photographs were taken of the face; and A sketch of the face was made on the first sample ticket of the series.

The sampler took the samples back to the geology office where the numbering was checked, commercial blanks and standards or coarse reject duplicates were inserted into the series, and a submittal form was filled out. Chain of custody (COC) signatures were collected on the form, and the mine received a copy of the form once it was completed. The samples remained in the office until they were picked up by an employee of the assay lab. The office was locked if unattended.

The sampler entered the data from the tickets into a spreadsheet. When assay results were received, the results were copied and pasted into the spreadsheet. The sampler downloaded the photos and labeled them.

Sample width was determined by material changes observed in the face, and are highly variable. Samplers used several factors to determine sample breaks: differences in sulfide abundance, rock quality, color or other physical properties related to mineralization. The average sample width in the chip sample data set is 1.0 ft. Four percent of the samples exceeded 4.0 ft in width. About 39% of the samples were between 2.0 and 4.0 ft wide, and 57% were less than 2.0 ft wide. Only 4% of chip samples were greater than 4.0 ft wide.

Samples were taken to the surface where they were picked up by American Analytical Labs., a certified assay laboratory, and transported to the laboratory for analysis. Sample chain of custody was maintained.

Detailed descriptions of sampling completed in several mine areas are included below.

GFN Hooper Level Drifting

During 2007-2008, GFN extended the Hooper Tunnel westward for 1,000 ft along the Alhambra Fault. GFN geologists collected chip samples from the face and back, plus car samples, at intervals of 5 to 20 ft during the drifting. The face samples yielded an average, undiluted grade of 7.8 oz/t Ag over a strike length of 358 ft, with an average vein width of 4.0 ft. This sampling was not reported in the 2010 Technical Report.

USC Hooper Level Lateral

During 2011-2011, USC extended the Hooper Tunnel west for an additional 1170 ft in order to establish a diamond drill platform to test the down-dip extents of the South Vein mineralization. This lateral was run in the hangingwall of the Alhambra Fault rather than along it, to take advantage of better ground conditions. The lateral was also driven on a bearing south of west to close the distance to the South Fault. The drill station was established at coordinates 9210E, -2300N.



USC Countess Decline Drifting

Most of the estimated resources identified by the several GFN drilling programs occur above an elevation of 3,260 ft above mean sea level (amsl). This places the estimated resources above accessible historic mine workings. Resources were identified on both the historically mined Alhambra Vein, and on the South Vein. To further develop and quantify the drill indicated resources, USC collared a decline portal on the Countess mining claim at the 4,220 ft elevation to access both veins. The decline was designed to access both the Alhambra and South Veins to enable exploration development drifting to more fully characterize and quantify mineralization within each vein.

Construction of the Countess decline began in the summer of 2010. The decline was driven southerly from the north side of the property at a grade of minus 13%. The decline was designed to cross-cut or intersect the veins near where drill intercepts encountered silver grades higher than 15 oz/t. The Alhambra vein was cross-cut at an elevation of 4,120 ft amsl and by June 2011, 450 ft of exploration drifting was completed on the Alhambra Vein. This exploration drifting identified two sections of mineralization: an eastern block 80.0 ft long with an average diluted width of 4.5 ft and an average Ag grade of 10.0 oz/t., and a western block 106.8 ft long with an average diluted width of 6.0 ft and an average Ag grade of 9.7 oz/t.

The decline reached the South Vein in March 2011 at an elevation of 3,940 ft amsl. An exploration development drift was completed by December 31, 2011. The exploration drifting identified three mineralized shoots having a combined strike length of 776.5 ft. The eastern block is 325.0 ft long with an average diluted mining width of 4.0 ft and an average Ag grade of 17.8 oz/t. The central block is 292.0 ft long with an average diluted mining width of 5.6 ft and an average Ag grade of 21.3 oz/t. The western block is 157.8 ft long with an average diluted mining width of 5.8 ft and an average Ag grade of 15.4 oz/t.

In March 2012 work was resumed on the Countess decline and a spiral ramp at minus 13% was started in the footwall of the South Vein. The decline was advanced to the 3740 elevation by August 2012 when decline development was stopped. Two additional exploration drifts were mined, on the 3890 elevation and 3840 elevations.

On the 3890 elevation, 585.4 ft of mineralized shoots were identified. The eastern block is 223.8 ft long at an average diluted mining width of 4.6 ft and an average Ag grade of 10.5 oz/t. The central block is 281.8 ft long with an average diluted mining width of 6.0 ft and an average Ag grade of 13.3 oz/t. The western block is 79.8 ft long with an average diluted mining width of 4.0 ft and an average Ag grade of 21.9 oz/t.

On the 3840 elevation only 33.1 ft of the eastern block was developed (there remains more than 200 ft yet to be developed) with an average diluted mining width of 4.5 ft at an average Ag grade of 4.9 oz/t. The central block is 153.0 ft long with an average diluted mining width of 6.5 ft and an average Ag grade of 15.1 oz/t. Associated with both the above blocks but splaying into the footwall and identified as the Jackson Vein is a splay of mineralization 38.2 ft long with an average diluted mining width of 4.8 ft and an average grade of 8.2 oz/t. The western block is 86.7 ft long with an average diluted mining width of 5.2 ft and an average Ag grade of 26.7 oz/t.



9.3 Significant Results and Interpretation Figure 9.3.1 shows the underground workings along the South Vein completed during test mining. Figure 9.3.2 shows the phases of advance in the Crescent Mine.

The test drifts provided information on the relationship between the mineralization and the fault zones, the geometry and continuity of the mineralization, and the characteristics of host lithologies. Test mining provided information on ground conditions, water inflows and operational costs.

Both the Alhambra and South Faults have evidence of post-mineral movement on the order of hundreds of feet. Movement on the South Fault has redistributed lenses of mill feed to an unknown extent, but appears to be significant. Minor faults, not including the South and Alhambra Faults, have less post-mineral offset than the Alhambra and South Faults. The greatest apparent offset observed on a minor fault was about 4 ft.

In the Alhambra test drift, about 4,110 ft amsl, all of the mineralization occurs on the hangingwall side of the fault. In the South Vein test drifts, between 3,850 and 3,950 ft amsl, silver mineralization occurs on both the hangingwall and footwall of the South Fault.

Other important geologic insights gained from the test mining are:

The orientations of mineralized structures to enable reinterpretation of drillhole data; A better understanding of the geometry of oxidation within the South vein system; Identification of an area of convergence of the mineralized South Fault with a mineralized

siderite vein, newly named the Jackson vein, that coincides with the highest grade drill intercepts; and

Identification of a major syncline that presents a new exploration target along its projection.

The diluted mining width used to calculate the average grades and mining widths described above was determined following a review of parameters used to estimate historic reserves at the Crescent mine. Historic reserve estimations added one half foot of dilution on each side of the mineralization or a minimum mining width of 4.0 ft, whichever was the greater. However, some reserves were also estimated using a 4.5 ft minimum mining width. For the average grades and mining widths reported in this document, a 4.5 ft minimum mining width was used to because it was more conservative.

One of USC’s applications of chip sample data was to estimate the grade of material sent to the mill. Three separate stockpiles were made at the mill for each of the three South test drifts. USC compared tonnage and grades calculated from the chip samples and the tons and grade measured at the mill during throughput, for each test drift. The mill reported fewer contained ounces than the mine calculated, 39% less for the first test drift, and 13% less each for the second and third test drifts. The difference between the first and second/third test drifts is believed due to different material handling procedures. If the sampling method were to be used for future test drifts, it would be expected to continue to yield results 13% lower than the mill. The explanation for the discrepancy is not known, but the most likely causes are sampling bias by the samplers, loss of fines or incorrect assignment of tonnage factors. Results of this comparison are reported in Section 13, Metallurgy and Mineral Processing.


Kellogg, Idaho

Figure 9.3.1

Development on the South Vein Source: USC, 2013


Kellogg, Idaho

Figure 9.3.2

Underground Advance by USC 2010-2012 Source: USC, 2013



10 Drilling GFN conducted surface and underground diamond drilling programs at the Project in 2007 and 2008. USC conducted surface and underground drilling programs in 2011. Descriptions of the GFN and USC surface and underground drilling programs are included below.

10.1 Type and Extent

10.1.1 GFN 2007-2008 Programs

One hundred and seven diamond drillholes were completed by GFN from surface and underground locations during 2007 and 2008, using various drill contractors. A fan-type drilling pattern was used from the drill stations in order to target the Alhambra Fault and South Veins on nominal 200 ft spacing. Figure 10.1.1.1 is a plan view of the drillhole traces.

Summer 2007 Surface Program

The summer 2007 surface drilling program was contracted to Atlas Fausett Contracting and Kettle Drilling, Inc. A total of 41,081 ft was drilled in 39 NQ diameter holes from 7 drill pads, using a Longyear 38, a Hagby 1500, a Hagby 1000, and two different Atlas Copco U8 drilling machines.

December 2007 to July 2008 Underground Program

Between December 2007 and July 2008, Kettle Drilling Inc. completed an underground drilling program using a U-8 drill. A total of 35,018 ft was drilled in 29 NQ diameter holes and 1 HQ diameter hole from 3 drill stations on the Hooper Tunnel level at elevations of 2,710 to 2,735 ft. These holes targeted the South Vein, but some of them also passed through the Alhambra Fault.

Summer 2008 Surface Program

In the summer of 2008, Kettle Drilling Inc. completed another surface program. A total of 27,026 ft in 31 NQ diameter holes were completed from four drill pads. From June to August 2008, up to three drill rigs were running around the clock in order to take advantage of water supplied by seasonal runoff. A lightweight, Zinex A-5 drill began drilling on Alhambra Fault targets in June and completed its program in July. The total footage from this machine was 13,713 ft in 12 holes. Two of these were not completed to the target depth. A second lightweight drill, a Longyear LF-70, began drilling South Vein targets in June and completed its program in August. The total footage from this machine was 7,754 ft in six holes. A Hagby 1000 began drilling Alhambra vein targets in June and completed its program in July. The total footage from this machine was 5,550 ft in nine holes. One of the holes was not completed to the target depth. The U-8 drill was moved from underground onto the surface in August. This rig completed 3,717 ft in four holes during a one month period.

October to November 2008 underground program

In the fall of 2008, seven holes were drilled by American Drilling in an underground program from a single drill station on the Hooper level, targeting the footwall of the Alhambra fault.

Most of the GFN drillholes were between 500 and 2,000 ft in length, targeting one of two structures, either the Alhambra Fault or the South Vein. The holes targeting the Alhambra Fault tested an area about 3,500 ft along strike, from the surface down-dip approximately 2,300 ft. The holes targeting the South Vein tested an area about 3,500 ft along strike from surface down-dip approximately 2,500 ft.



10.1.2 USC 2011 Programs

Between April and July 2011, USC completed 11 NQ-diameter core holes. Targets were the South Vein and Alhambra Vein, west of the known resource area. Total drilled footage for this program was 12,908 ft. Additional drillholes were planned, but due to budget constraints, were not completed. Seven of the completed drillholes targeted the South Vein structure; the other four, the Alhambra Vein structure. Figure 10.1.2.1 shows the surface locations and traces of the drillholes in the 2011 program.

2011 Surface Drilling Program

The surface drilling was contracted to American Drilling of Spokane, WA. An Atlas-Copco CS-10 drill rig was mobilized to the property, and began to drill on April 9, 2011. The CS-10 is a surface rig that is not capable of drilling the shallowly-inclined holes that comprised most of the proposed program, and because an underground-style rig was unavailable, USC selected steeply-inclined holes as the first four holes of the program, and drilled them with the CS-10.

The four holes were drilled to test the Alhambra fault, and intersect the Alhambra in the area of the Dawn adit. Drill core recoveries were excellent.

Subsequently, a Diamec U-8 drill was mobilized to replace the CS-10, and commenced drilling on May 6. Three holes were drilled from the S-2 pad, followed by 2 holes from the S-3 pad, at which point the drill program was terminated on July 9. All five holes were drilled from the surface to the south at shallow inclinations to intersect the South fault.

2011 Underground Drilling Program

The underground drilling was also contracted to American Drilling of Spokane, WA. The holes were drilled from a station cut at the present west end of the west Hooper lateral, using a Diamec U-6 drill rig.

Two holes drilled to test the down-dip projection of the South vein were completed before drilling was terminated on July 1.

10.2 Procedures All holes were NQ in size, and surveyed after completion with a Flexit Multishot® downhole survey instrument. Drill collar locations were surveyed by the mine surveyor. The 2011 drilling program is summarized in Table 10.2.1.

During the drilling operation, the core was retrieved from the core barrel and laid sequentially into cardboard core boxes. Interval blocks were placed at all run breaks. Once the box contained approximately 10 ft of core, the ends and sides were labeled with drillhole identification, from and to intervals and the sequential box number. The box was then covered by a cardboard lid and stacked at the rig to assure that the core was not exposed to any potential contamination or mix-ups.

At the end of each drilling shift, the boxes of core were transported by the drilling contractor in a pickup truck or underground cart to the USC core shed on site. The drill contractors delivered core, at the end of each shift, to the core shed if geology staff were present. If geology staff were not present, the core was placed in an annex building and the building was locked by the drillers, transferring custody of the core to USC.



Upon receipt at the core shed, the core boxes were arranged in order and the core was washed, photographed, and examined to identify mineralized zones. The boxes containing identified mineralization were marked and set aside to be logged and sampled immediately; the remaining boxes were stacked on pallets and stored within one of two locked buildings, to be logged as staff time allowed.

Geologic logging was done on paper log forms using a graphic log format. The log form contains columns which record: interval drill depths, core recovery, carbonate alteration, sulfide alteration, vein intensity, lithology, and color. After logging the zones of alteration, the sample intervals were determined and marked on the core and the core boxes. Most samples represent core intervals of 1.0 to 3.0 ft; about 10% of the sampling is of intervals from 3.0 to 4.3 ft in length. Intervals were chosen on the basis of visual determinations of grade in an attempt to break out separate sample material that varied from adjacent material by orders of magnitude. Samples assaying greater than 4 oz/t Ag are “bracketed” by intervals that assay < 1 oz/t Ag. The sample intervals are recorded in three places; the sample log sheet, the sample tag booklets and by the placement of cards or flagging delimiting each interval in the boxes. A brief description of the sample was noted on the sample sheet which is kept with the drill log. Data from paper logs was tabulated to include in the drillhole database.

After logging, core samples were bisected along the core axis, to create representative halves. Competent samples were generally sawn, though some samples, especially well-fractured ones, were split with a hydraulic splitter. Incompetent material was split by hand using a putty knife. Half of the cut core was then placed into a pre-labeled polyethylene bag along with a sample identification tag with a blind sample number. Each bag was closed with a zip-tie. A digital database is maintained, which records the drillhole identification and from-to intervals of all sample tags. Assay results were added to the database after quality assurance review.

After splitting, the unsampled half of the core was returned to the box for archive. The archive boxes of half core were then stacked in one of the two locked core buildings to maintain sample security and protect them from weather.

The individual sample bags containing the core samples were stored within the locked core building until transported by USC staff to the assay lab or picked up by an employee of the assay lab. Commercial blanks and standards or coarse reject duplicates were inserted into the series, and a submittal form filled out. Part of the submittal form includes Chain of Custody (COC) documentation- the assay lab returned a copy of each signed COC form to the mine. All samples were sent to American Analytical Services, Inc. (AAS) in Osburn, ID, an ISO-accredited laboratory located about 6.5 mi from the Project.



Table 10.2.1: 2011 Drilling Program

DDH azimuth dip TD (ft) Target Depth of Target

(ft) C-1 329 -61 867 Alhambra 783’ C-2 327 -74 970 Alhambra 902’ C-3 17 -69 804 Alhambra 653’ C-4 335 -66 713.5 Alhambra 631’ C-5 159 -16 1358 South vein 1336’ C-6 182 -35 1815 South vein 1672’ C-7 188 -21 1357 South vein Unknown C-8 177 -23 1686 South vein 1400’ C-9 192 -31 1687 South vein 1618’ H-6 210 6 855 South vein 722’ H-7 216 28 778 South vein 727’ Source: USC, 2013

Several factors could affect the accuracy and reliability of the location surveys and analytical data. Drillhole collar location survey methods appear consistent with the underground workings, but the precision of these measurements is unknown. The drillhole collars were located by the Crescent Mine survey professionals, so this factor is minor. The reliability of the downhole surveys for about 4 recent drillholes is suspect, because the mineralized intercepts do not fall within the modeled vein envelope defined by the rest of the drillhole data. This discrepancy has been addressed in the modeling process. The last noted risk factor for the reliability of the results is less than full recovery in mineralized fault zones. Most reported core recovery was good for the recent drilling programs. This issue should be discussed with any future drilling contractors, who should make every effort to get full core recovery.

10.3 Interpretation and Relevant Results

10.3.1 GFN 2007-2008 Programs

One hundred and seven diamond drillholes were completed by GFN from surface and underground locations during 2007 and 2008, for a total of 112,276 ft.

The GFN drilling identified a new mineralized area on the South structure between 3,000 ft and 4,500 ft amsl. The South structure had not been explored previously above an elevation of -350 ft amsl with the exception of a few trenches along the strongly weathered surface trace nearly a mile above. Additionally, the GFN drilling confirmed the presence of mineralization along the Alhambra Fault west of the historic Brooks stoping and above the Alhambra Level.

10.3.2 USC 2011 Programs

A total of 12,908 ft was drilled in 9 surface and 2 underground diamond drillholes during 2011. H-7 showed a 6 ft core loss through the target. Table 10.3.2.1 summarizes the significant intercepts of the 2011 drill program.



Table 10.3.2.1: Significant Intercepts in 2011 Drilling Programs

Hole ID From (ft) To (ft) Length (ft) Approximate true width (ft) Ag (oz/t) C-1 783.30 785.50 2.2 1.6 4.84 C-3 652.7 654.7 2.0 1.3 4.50 C-5 992.3 995.2 2.9 2.5 4.58 H-6 721.40 722.40 1.0 0.8 96.4 H-7 727.00 727.65 0.65 0.5 8.80

The four holes drilled to test the Alhambra fault show a consistent, low-grade mineralized zone in the immediate hangingwall of the fault. The mineralization is characterized by pyrite as stockworks that locally comprise more than 50% of the interval, with subordinate siderite veining. The intercepts are shown on Figure 10.1.2.1.

Two of the five holes drilled to the South structure identified a hitherto unknown zone of mineralization in the footwall of the South structure. The zone, now named the Jackson Vein, is a stockwork of quartz, siderite and pyrite veining some tens of feet thick, of a type that has historically shown promise of finding mill feed shoots within it by following it along strike or dip. The geometry of this zone is not well known, but the available drillhole data suggests that it may run between the South and Alhambra structures.

DDH C-7 was drilled through the projected location of the South Fault at the time the hole was drilled, but subsequent mining indicates that the strike of the South Fault changes to the west, and that the C-7 ended short of the new projected location of the fault.

Holes C-8 and C-9 intercepted igneous dike at the approximate projection of the South Fault.

The two underground holes intersected the South structure on projection down-dip of the existing resource, and suggest that potential exists to extend the existing resource to greater depth, though additional drill testing is needed. The intercepts from the 2011 drilling program are shown in Figure 10.3.2.1.


Kellogg, Idaho

Figure 10.1.1.1

Map of GFN and USC Drillhole Traces Source: SRK, 2013


Kellogg, Idaho

Figure 10.1.2.1

Alhambra Long Section with 2011 Drillhole Intercepts Source: SRK, 2013


Kellogg, Idaho

Figure 10.3.2.1

South Vein Long Section with 2011 Drillhole Intercepts Source: SRK, 2013



11 Sample Preparation, Analysis and Security GFN collected 1,465 drill core samples during their drill programs, and shipped them to four different laboratories. A multi-element ICP analysis was first run and then all samples testing above the ICP threshold for silver were reanalyzed by fire assay.

GFN began to generate core and chip samples in July 2007. These early samples were submitted to two local labs for analysis, American Analytical in Osburn, Idaho, and Chris Christopherson in Smelterville, Idaho. Neither lab was accredited at the time. Following SRK’s recommendation, in April of 2008 GFN began sending their samples only to accredited labs. One batch of samples was sent to American Assay of Sparks, Nevada and from May 2008 on, all samples were sent to ALS Chemex. All samples sent to ALS Chemex were prepared at the Sparks, Nevada facility and then analyzed at the North Vancouver Laboratory. ALS Chemex located at 212 Brooksbank Ave in North Vancouver Canada was certified under ISO 9001:2000 for the provision of assay and geochemical services according to QMI Management Systems Registration.

11.1 Security Measures

11.1.1 GFN 2007-2008 Drill Programs

The individual canvas sample bags containing the core samples were placed in 5 gallon plastic buckets stored within the locked core building. Once a bucket was full its lid was snapped closed and taped shut so that no tampering can occur. A sample transmittal list was then compiled. The buckets of samples were transported by Federal Express or United Parcel Service to ALS Chemex in Reno, Nevada by standard transport truck.

11.1.2 USC 2011 Drill Programs

The core samples were stored within the locked core building until transported by staff to the assay lab or picked up by an employee of the assay lab. The underground chip samples were taken to the geology office where they remained until picked up by an employee of the assay lab. The office was locked if unattended. Chain of custody transfer from USC to the assay lab was documented.

11.2 Sample Preparation for Analysis

11.2.1 GFN 2007-2008 Drill Programs

At ALS Chemex, the samples were unpacked upon arrival and arranged in order, then logged into the system by sample identification number. Each sample bag was emptied into a clean metal sample tray and placed into a drying oven at 60°C for approximately four hours. The samples were then run through a primary jaw crusher and then a secondary cone crusher to produce a product with specifications of 70% less than 2 mm in size. The sample was then blended and run through a Jones riffle splitter to produce a 250 g subsample. The reject material was returned to the original sample bag and archived. The 250 g subsample was next run through a ring pulverizer to produce a product with specification of 85% less than 200 mesh. The crushers, splitter and pulverizers were blown clean with an air hose after every sample and the sample preparation room was equipped with a dust collection system. All samples sent to ALS Chemex were prepared at the Sparks, Nevada facility and then analyzed at the North Vancouver Laboratory. ALS Chemex located at 212



Brooksbank Ave in North Vancouver Canada was an independent laboratory certified under ISO 9001:2000 for the provision of assay and geochemical services according to QMI Management Systems Registration.


The samples from the 2010-2012 chip sampling and from the 2011 drill program were prepared and assayed by an independent laboratory, American Analytical Services (AAS) in Osburn, ID, certified under ISO/IEC 17025:2005.

At AAS, samples were jaw and cone crushed to approximately 10 mesh, then a 250 g split was taken. The split was ring pulverized to 140 mesh and the rejected portion of sample was returned to the original sample bag, saved and returned to USC. Barren rock was run through the crushing equipment between samples to clean the crushers, and one sample each day per client was provided as a sample preparation blank to be run with the normal samples. During pulverization, silica sand was used between each sample to clean bowl and rings to prevent cross-contamination. Each 140 mesh pulp sample was placed in a sample envelope and sent to the digestion room. Samples were delivered to the instrument lab for ICP analysis and fire assay.

11.3 Sample Analysis

11.3.1 GFN Sample Analysis

The samples were analyzed by ALS Chemex using Inductively Coupled Plasma Atomic Emission Spectrometry (ICP-AES) for a suite of 35 elements (ME-ICP41). Any samples returning over the maximum detection limit for Ag, 100 ppm, were then analyzed by mill feed ICP (Ag-OG46), and any samples returning over the mill feed grade maximum detection limit for Ag, 1,500 ppm, were analyzed by fire assay (Ag-GRA21).

During the 35-element ICP-AES, a 0.2 g portion of the pulp sample was first placed into a test tube and dissolved using an aqua regia digestion. A typical atomic absorption spectrometer consists of an appropriate light source (usually a hollow cathode lamp containing the element to be measured), an absorption path (usually a flame, but occasionally an absorption cell), a monochromator (to isolate the light of appropriate wavelength) and a detector. The most common form of atomic absorption spectroscopy is called flame atomic absorption. In this technique, a solution of the element of interest is drawn through a flame in order to generate the element in its atomic form. At the same time, light from a hollow cathode lamp is passed through the flame and atomic absorption occurs. The flame temperature can be varied by using different fuel and oxidant combinations; for example, a hotter flame is required for those elements which resist atomization by tending to form refractory oxides (ALS Chemex 2009).

The primary limitation of IPC-AES is that all measurements are made following chemical dissolution of the element of interest. Therefore, the measurement can only be as good as the quality of the sample digestion. A second limitation is that occasionally, interferences from other elements or chemical species can impact atomic emission and depress absorbance, thereby reducing sensitivity. For these reasons, most reputable laboratories (ALS Chemex included) recommend that ICP-AES not be used for reserve estimations or bankable feasibility studies.



In general, an aqua regia digestion of relatively high silver-grade samples in the presence of lead can result in precipitation of silver chloride complexes which results in a solution underreporting true silver content. Additionally, the ICP-AES instrument is designed for very low detection limits and solutions must be highly diluted to accurately measure higher concentrations of metals. This can also have a negative impact on the accuracy of the results. GFN made an effort to overcome this limitation by ensuring that all samples with silver mineralization in excess of 1,500 ppm were re-analyzed by standard fire assay methods and they conducted numerous duplicate fire assay analysis of samples originally analyzed by ICP. The maximum detection limit for fire assay was 10,000 ppm Ag; only one sample in the data set exceeded this 1% limit. The fire assay technique used a 30 g charge with a gravimetric finish. This fire assay method is appropriate for the silver grades defining the anomalous mineralization.


During 2010 and 2011, all samples were analyzed for silver by fire assay and for a suite of 10 elements by ICP (As, Bi, Cu, Cd, Fe, Mn, Pb, S, Sb, Zn). All samples collected during 2012 were analyzed by fire assay for silver only. For fire assay, one assay ton (29.16 g) of sample is weighed with approximately 100 g of standard flux mixture into 30 g crucibles. Silver inquarts are added when necessary. Lead buttons are cupeled in bone ash cupels. Duplicates and controls are included in each batch of 20 samples.

11.4 Quality Assurance/Quality Control Procedures Following SRK’s recommendations, GFN began to use standard reference material, blank samples and duplicate samples in April 2008 to ensure that reliable assay data was being obtained.

For the 2011 program, USC incorporated coarse reject duplicates, and commercial standards and blanks in the chip and core sampling streams at a rate of two QA/QC samples in every twenty shipped samples. On average, every 60 samples included two standards, two blanks, and two coarse reject duplicates.

11.4.1 Standard Reference Materials

GFN Standards

Commercial, standard reference material samples were inserted into the sample stream at 1 in 20 intervals, and were considered to have “failed” if the lab returned an assay outside three standard deviations of the certified value.

USC Standards

A total of 77 commercial standard pulps were included and analyzed in batches of samples sent to AAS between December 2010 and September 2012; 66 standards were included with batches of chip samples and the remaining 11 standards were included in the drill core sampling stream.

Eleven different commercial standards were used, all of them silver-bearing standards from WGM Minerals, Inc.



11.4.2 Blank Samples

GFN Blanks

Commercial blanks were also inserted into the sample stream at 1 in 20 intervals, and were considered to have “failed” if the lab returned an assay greater than 5 ppm. Because the commercial blanks were pulverized material, and the blanks do not act as a check of possible contamination in the pulverizing stage, a small number of "field blanks" comprised of barren core were also inserted in the sample stream.

USC Blanks

A total of 76 commercial blank pulps were included and analyzed in batches of samples sent to AAS between December 2010 and September 2012; 66 blanks were included with batches of chip samples and the remaining 10 blanks were included in the drill core sampling stream.

11.4.3 Duplicate Samples

GFN Duplicates

Two types of duplicate analysis were run by GFN: 1) coarse rejects prepared and analyzed at the primary lab and then shipped to a second lab where they were split, pulverized and analyzed; 2) field duplicates created from quarter core samples of the original half core intervals.

A total of 119 coarse reject duplicate samples were analyzed. These consisted of 58 samples initially analyzed by the non-accredited labs and 54 samples by ALS Chemex. All of these were re-analyzed by ALS Chemex.

Thirty-six field duplicate samples were collected to provide a further check on the 2007 analytical data. These were generated by quarter core samples taken from previously sampled vein intervals. The field duplicates were all prepared and analyzed by ALS Chemex. These resamples were also selected from higher-grade intervals that would figure prominently in the resource estimation and were chosen to replicate the intervals of the earlier sampling. This was sometimes difficult, as the original intervals from the early sampling were not always well preserved in the core boxes.

USC Duplicates

A total of 42 coarse rejects (40 chip samples and two drill core samples) were re-assayed by inserting them at intervals within the chip and core sampling stream. An additional 30 coarse rejects (17 chip samples and 13 drill core samples) were analyzed for Au by fire assay; this method yields Ag results, and thus provided further duplicate analyses.

11.4.4 Actions

GFN Actions

Failed standards were handled by rerunning the entire sample batch containing the failure. In most cases, the rerun standard passed and the rerun data was used in the final database. The failures and subsequent corrective action were individually documented by GFN. In one case, ALS Chemex confirmed the switch, and the data was corrected. In another case, a rerun was made from the coarse rejects, and the rerun showed additional switched samples. This data was rejected, and not replaced.



USC Actions

There were no QA/QC actions taken. Analytical results of QA/QC samples were within acceptable value ranges for each sample type.

11.4.5 Results

GFN Standards

The initial failure rate of standards in the ALS Chemex analyses was 13% but this did improve over time. The sample failures typically consisted of values a few percent outside the tolerance range, but there were two instances of obvious sample switching. The results of the various silver standard analyses are shown in Figures 11.4.5.1.

USC Standards

All results were within one standard deviation or less of the published standard mean value, except for one result. The 77 standard sample analyses average 4.3% lower than the COA grade provided by the commercial standards supplier. The results of the various silver standard analyses are shown in Figures 11.4.5.2.

GFN Blanks

Only one failure occurred in 80 blanks, apparently because of a switched sample (Figure 11.4.5.3). There were no failures of field blanks.

USC Blanks

All blanks returned below detection (<0.100 oz/t Ag) except two, which returned 0.105 and 0.128 oz/t Ag; this is 11.4.5.3.

GFN Duplicates

For the coarse reject duplicates, the results of the 58 samples from the non-accredited lab showed good correlation to the duplicate analysis generated by ALS Chemex as shown in Figure 11.4.5.4. The ALS Chemex duplicates also showed very good correlation (Figure 11.4.5.4).

For the quarter core duplicates, the results shown in Figure 11.4.5.5 indicate a reasonable correlation between the two sample results.

USC Duplicates

The duplicate assays compare well; there was one failure in which the original assay was 0.772, and the rerun returned 27.0 oz/t Ag. Figure 11.4.5.6 shows the comparison between original and rerun assays.

11.5 Opinion on Adequacy Security, Sample Preparation, Analytical Procedures

Standard sample results suggest a systematic tendency for AAS to return numbers lower than the standard COA average, especially for two standards, CU 112 and PM 1135, in the 7 to 10 oz/t Ag range. Another standard, PM 1120, however, with an average COA grade of 11 oz/t, compared well. Despite a slight apparent systematic bias, the standards results indicate that AAS performed adequately to keep samples properly labeled. The blanks results suggest that AAS performed



adequately to prevent sample contamination and keep samples properly labeled. There was one failure in coarse reject duplicate pair results. Analytical results from quality control samples indicate that AAS performed adequately to meet resource reporting standards.


Kellogg, Idaho

Figure 11.4.5.1

GFN Standards Analyses Source: SRK, 2013


Kellogg, Idaho

Figure 11.4.5.2

USC Standards Analyses Source: SRK, 2013


Kellogg, Idaho

Figure 11.4.5.3

GFN and USC Blanks Source: SRK, 2013


Kellogg, Idaho

Figure 11.4.5.4

GFN Duplicates Source: SRK, 2013


Kellogg, Idaho

Figure 11.4.5.5

GFN Quarter Core Samples Source: SRK, 2013


Kellogg, Idaho

Figure 11.4.5.6

USC Duplicates Source: SRK, 2013



12 Data Verification Analytical data was verified with check assays sent to a secondary lab. The accuracy of the drillhole and chip sample database was verified by comparison with assay certificates.

12.1 Procedures

12.1.1 GFN Check Sampling

Thirty-five pulp samples prepared and analyzed by ALS Chemex were selected for shipment to SGS Mineral Services of Lakefield, Ontario, where they were analyzed by fire assay. The samples were predominantly chosen from the higher-grade samples because of their significance in the resource estimation.

The SGS results agree reasonably well with the ALS Chemex data (Figure 12.1.1.1). The SGS numbers range from 42% higher to 20% lower, and average 4% lower than the ALS Chemex results. While the SGS numbers were all fire assays, the ALS Chemex numbers were a combination of both fire assay and ICP.

12.1.2 USC Check Sampling

Analytical data was verified by sending a selection of check samples to a secondary lab: ALS Minerals in Reno, Nevada, certification ISO/IEC 17025:2005.

Five percent (n=70) of the chip sample set was selected for check assay. Samples were chosen to represent the range in Ag grade, oxidation state, and location within the two represented vein systems, as well as the time range of assay at the primary lab. Check assay samples were equal proportions of pulps and coarse rejects. Four commercial blanks and four commercial standards were inserted into the sample sequence. The samples were prepared according to the same specifications as the AAS samples, if required, and all were assayed for silver by fire assay.

No drill samples were included in the check sample set because only seven drill core samples occur within the resource area.

In general, the check sample results confirmed the original results, as shown in Figure 12.1.2.1. ALS switched one pair of adjoining samples, and two other samples were significantly different. The standards and blanks returned within two standard deviations of each mean value, and below detection, respectively. The highest grade sample was greater than the ALS detection limit for this method.

12.1.3 Comparison of Assay Certificates and Drillhole Database

Lab results in secure Portable Document Format (pdf) or digitally signed Excel spreadsheets were available for the majority of USC drillhole, face and longhole samples collected between 2010 and 2012. SRK checked the silver assay database against the laboratory reported values for 38% of the drillhole samples (263 of 688 with assay data) and 18% of the face chip samples (238 of 1321 with data). No discrepancies were found in any of the mineralized intervals. The only deviation between the assay database and the reported lab results is the treatment of reported silver values below detection limit. Some were assigned to 0.050 oz/t, half of the lower detection limit, but others were



included in the database as reported, <0.0100 oz/t. These values are much less than the resource CoG, and therefore the assigned values are not material to the assay database.

SRK identified three errors in the database in the conversion of ppm to oz/t. SNS-208 at depth 690.8 ft was corrected from an Ag oz/t value of 1853 to 10.558. H-109 at a depth of 291 ft was corrected from an Ag oz/t value of 546 to 11.142. In the same hole at a depth of 416.6, the Ag oz/t value was corrected from 1,196 to 9.78 as reported in assay certificate source documents.

12.2 Limitations The focus of the data validation for this report is the recent exploration and drilling data. Historic drillhole assays were not verified during this phase of the Project.

Not all assay certificates were available for verification; however, those available provided adequate information to validate the database. The check assay samples are representative of the additional resource development, and the quantity selected is considered adequate by current industry standards.

12.3 Opinion on Data Adequacy SRK is of the opinion that the quality and quantity of data in the drillhole database is adequate for resource estimation.


Kellogg, Idaho

Figure 12.1.1.1

GFN Check Samples Source: SRK, 2013


Kellogg, Idaho

Figure 12.1.2.1

USC Check Samples Source: SRK, 2013



13 Mineral Processing and Metallurgical Testing

13.1 Testing and Procedures In June of 2011, USC commissioned G & T Metallurgical Services, Ltd. (G&T) of Kamloops, British Columbia, to complete metallurgical test work on an 84 kg sample of South Vein material obtained from the 3950 sublevel of the Crescent mine. The test work program consisted of the following:

Chemical content; Material hardness; Rougher flotation tests; and Cleaner flotation tests.

G&T provided a report of the test results (G&T, 2011).

In April of 2011, a bulk sample of about 123 t of South vein material from the 3950 sublevel was processed in the old (pre-expansion) 100 t/d circuit at the New Jersey mill in Kellogg, Idaho. The material was processed over a period of four days, and the objectives of the bulk sampling program were:

Determine the silver head grade and the quantity of other metals present; Obtain metallurgical data for expanded mill design; Obtain metallurgical data for economic assessment; Evaluate different flotation reagents; and Determine if a marketable concentrate could be produced.

A report by William C. Rust, consulting metallurgist, summarizes the results of the bulk sample mill test (Rust, 2011).

13.1.1 Sample Representativeness

The bulk sample was mined soon after development of the first I-drift (3950 Level) on the South Vein began. Face sample assay results encountered assays from 25 to over 100 oz/t. Visually, the mineralization was nearly all sulfide and at an average grade of 24.5 oz/t, is significantly higher grade than the resource average. Based on the test work, recovery was forecast at 89.5% which is exactly the average recovery for the 12,000 t milled to date. Subsequent to extracting the sample, visual observations of the vein in another 2,000 plus feet of I-drifts were made and the sampled area appears very representative of sulfide mineralization in the South Vein.

13.2 Relevant Results The sample sent to G&T was collected from the ball mill feed conveyor during a bulk sample milling test at the New Jersey mill in early 2011. The sample was obtained from a raise round on the 3950 sublevel drift from a location where I-drift samples reported grades between about 25 and 100 oz/t silver in the South vein. The results of G&T’s chemical analysis of the sample are presented in Table 13.2.1.



Table 13.2.1: Head Assay Data for G&T Sample

Sample ID Assays (% or oz/t)Cu Ag Pb Zn As Sb Fe Au

South Vein 0.53% 24.5 oz/t 0.09% 0.01% 0.26% 0.26% 14.9% 0.003 oz/t Source: G&T, 2011

The bulk sample of nearly 123 t was mined from an I-drift along the South vein on the 3950 sublevel. Mill process samples were collected at the ball mill feed, the rougher feed (cyclone overflow), and the final tailings. The ball mill feed was collected by hand on an hourly basis while the rougher feed and final tailings were collected on two minute intervals by automatic slurry samplers. The mill head assay data for the bulk sample is presented below in Table 13.2.2.

Table 13.2.2: Head Assay Data for South Vein Bulk Sample

Product Dry Weight Ag Cu Pb As SbFeed tons oz/t % % % %Calculated 123.1 10.31 0.28 0.08 1.7 1.2 Assay 123.1 10.55 0.33 0.06 1.7 1.2 Source: Rust, 2011

13.2.1 Mill Feed Hardness

G&T completed a Bond ball mill work index test to assess the mill feed hardness of the Crescent South vein. The results are presented in Table 13.2.1.1. G&T classified the South vein mill feed as moderately hard with a Bond work index test result of 13.3 kWh/t.

Table 13.2.1.1: Bond Ball Mill Work Index Data

Sample ID Feed Size (µm P80)

Product Size(µm P80)

Ball MillWork Index (kWh/t)

South Vein 2019 86 13.3 Source: G&T, 2011

The actual Bond Work Index during operations was 9.65 kWh/t (Rust, 2012). Therefore, the actual work index is about 75% of what was predicted from G&T’s test work (2011).

13.2.2 Rougher Flotation Tests

G&T completed five rougher flotation tests to evaluate the effect of primary grind and collector type on metallurgical performance of the South vein mill feed. Lime was added in the grinding circuit to maintain a pH of 9 and sodium cyanide (NaCN) was added at a dosage of 5 g/tonne (0.16 lb/t). Both the lime and NaCN were added to depress pyrite in the rougher cells. Two collectors were also tested; Z-55 and Reagent 571, formerly called Aerofloat 242.

The primary grind of G&T’s rougher flotation tests varied from 80% passing 65 µm to about 211 µm. G&T reported that a primary grind of about 80% passing 97 µm (97 µm P80) had the best metallurgical performance with 92% silver recovery into a rougher concentrate with 7% of the mass balance reporting to the rougher concentrate. Copper recovery for the 97 µm P80 test was 77.7% with 7% of the mass reporting to the rougher concentrate. Results also showed that grinding too fine can be detrimental to recovery. G&T reported similar results for copper performance and that there



was no significant difference between the two collectors tested. Results of the rougher flotation tests are presented in Table 13.2.2.1.

Table 13.2.2.1: G&T Rougher Flotation Test Results

Primary Grind (P80) Ag Recovery (%) Cu Recovery (%) Mass Reporting to Rougher Con (%)

211 µm 82.1 71.0 6.6 144 µm 85.7 73.5 6.2 97 µm 92.2 77.7 7.2 65 µm 92.6 76.1 7.7 Source: G&T, 2011

Rougher flotation tests showed that a primary grind of 65 to 100 µm P80 produced the best silver and copper recoveries. Silver recovery was about 92% in that primary grind range.

13.2.3 Batch Cleaner Flotation Tests

G&T performed a single batch cleaner test using the rougher concentrate from the mill feed that had a primary grind of 97 µm P80 and additional Z-55 collector was added in the cleaner test. The rougher concentrate was cleaned in three stages of sequential cleaning without regrinding. This cleaner test produced a final concentrate of 691 oz/t (23,700 g/t) silver with 83% silver recovery.

Table 13.2.3.1: G&T Cleaner Flotation Test Result

Cleaner Product Ag (oz/t) Cu (%)Final Concentrate 691.4 13.1

13.2.4 Bulk Sample Milling

The bulk sample was delivered to the concrete pad at the New Jersey mill where it was loaded to a bin over a feed conveyor belt that fed a 15” x 24” jaw crusher. The mill feed was crushed to less than 3”. The mill feed was then conveyed to a screen with a 0.5” by 2” slotted screen cloth. The material passing through the screen was conveyed to a fine mill feed bin while the coarse material dropped into a 22” cone crusher for a single pass of crushing with a closed-side setting of 0.5 inch. Fine mill feed was then conveyed to a 100 hp 6 ft x 6 ft Marcy-type ball mill. Ball mill discharge was pumped to a 6 inch Warman hydro-cyclone where cyclone overflow reported to the rougher flotation cells and the cyclone underflow reported back to the ball mill for regrinding. Flotation reagents were added at the first rougher cell. The rougher bank consisted of five Warman 66D flotation cells. Rougher concentrate was pumped to a single bank of Warman 36 flotation cells where cleaning was done in a single stage. Cleaner tailings were recycled back to the first rougher cell. Final tailings were placed into a small impoundment adjacent to the mill as dilute slurry at about 30% solids. Concentrate was pumped to a plate and frame filter and after drying were dropped into 1 t super sacks.

13.2.5 Results of Bulk Sample Milling

Process samples were collected by mill personnel, Rust, and automatic slurry samplers during the four day run of the South vein bulk sample. Table 13.2.5.1 summarizes the results.



The recovery of silver in the final concentrate was 89.5%, and 1.6% of the total mass, including gangue minerals, went to the concentrate. Copper recovery was 76.1%. The ratio of concentration for silver was approximately 63 to 1.

Table 13.2.5.1: Mass Balance for South Vein Bulk Sample

Product Dry Weight

Ag (oz/t) Cu % Pb % As % Sb % Distribution by Mass (%)

Tons % of Total Ag CuFeed –Calculated 123.1 100.0 10.31 0.28 0.08 0.17 0.12 100.0 100.0 Feed – Assay 123.1 10.55 0.33 0.06 0.17 0.12 Concentrate 1.96 1.6 579.80 13.3 3.64 3.84 6.16 89.5 76.1 Tails 121.1 98.4 1.10 0.07 0.02 0.11 0.02 10.5 23.9 Source: Rust, 2011

Samples of various process streams were taken for particle size analyses including cyclone overflow in the grinding circuit to determine the grind that was achieved by the ball mill. Only the results of the cyclone overflow (flotation feed) particle size analysis are presented below in Figure 13.2.5.1.

The grinding circuit achieved 80% passing 102 µm. Approximately 67% of the particles passed 200 mesh (75 µm). The grind achieved during the bulk milling test was very similar to the grind that G&T found to be the best for silver recovery.

The two collectors used in the bulk milling test were Prospec Chemical’s TNC 312 and Reagent 571 (previously Aerofloat 242). No significant difference was found between the two collectors. A frothing agent, Methyl Isobutyl Carbine (MIBC), was not necessary with TNC 312, but used sparingly with 571.

The results of these two testing programs were used to design the NJMC mill expansion. As of September 2013, the expanded mill has processed over 12,000 t of Crescent material. After the commissioning period, from August 1st through September 3rd when the mill was operating 24 hours per day, recoveries averaged 94%. This recovery indicates that the expanded mill is achieving the design goals.

13.3 Recovery Estimate Assumptions Based on the metallurgical and recovery results obtained milling more than 12,000 t Crescent mill feed, silver recoveries in sulfide mineralization are expected to exceed 94%. However, recoveries decline to between 80% and 84% when the mineralized material is oxidized. This is typical for other mines in the district. Mine plans call for the mining below the known elevation oxidization during the early years of production to ensure excellent silver recoveries. This will provide time for metallurgical testing to be completed on the oxidized material to see if recoveries and, by extension, the economics of oxidized material can be improved. Controlling dilution and maintaining the grade milled in the 15 to 16 oz/t range will ensure that economic extraction will continue.

Penalties for arsenic and antimony in the concentrates are expected to stay within acceptable limits.

13.4 Significant Factors Based on results from milling 12,000 t of material to date, no significant factors are known or expected to affect silver recovery.


Kellogg, Idaho

Figure 13.2.5.1

Particle Distribution of Bulk Sample Flotation Feed Source: SRK, 2013



14 Mineral Resource Estimate

14.1 Introduction Mineral resources presented in this report for the Project have been estimated in accordance with standards adopted by the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) in 2000, as amended on November 27, 2010, and prescribed by Canadian Securities Administrators’ NI 43-101 (NI 43-101). The modeling and estimate of the mineral resources were performed by and under the supervision of Jay Pennington, a qualified person with respect to Mineral Resource estimation under NI 43-101. Mr. Pennington is a Certified Professional Geologist (CPG) as recognized by the American Institute of Professional Geologists and has over 28 years of experience in mineral exploration and resource geology for multiple commodities with specialization in precious and base metals. Mr. Pennington is independent of United Silver Corp by the definitions and criteria prescribed in NI 43-101; there is no affiliation between Mr. Pennington and United Silver Corp except that of an independent consultant/client relationship. The Crescent resources were modeled, estimated, and classified in June and July 2013.

Cautionary note to investors concerning estimates of Measured and Indicated Resources and

Inferred Resources. “Inferred Resources” have a great amount of uncertainty as to their existence, and great uncertainty as to their economic and legal feasibility. It cannot be assumed that all or any part of an Inferred Mineral Resource will ever be upgraded to a higher category. Under Canadian rules, estimates of Inferred Mineral Resources may not form the basis of Feasibility or Prefeasibility studies, except in rare cases. Investors are cautioned not to assume that part or all of an Inferred Resource exists, or is economically or legally minable.

14.2 Block Model The Crescent block model has the spatial characteristics and limits shown in Table 14.2.1. It was built in the Bunker Hill mine grid coordinate system. Digital topography was provided by United Silver Corp. All project distance units are in feet and metal grades are in troy ounces per short ton. The model was not rotated. It was built using full blocks and block percents that were coded by mineral domain wireframes. The resource block model was constructed using Mintec’s MineSight 3D® mining software.

Table 14.2.1: Crescent Model Origin and Extents

Direction Minimum (ft) Maximum (ft) 25 ft x 25 ft x 50 ft

Blocks Easting 6400 13200 272 Columns Northing -3300 -600 108 Rows Elevation 2000 5200 64 Levels

Source: SRK

The 25 ft x 25 ft x 50 ft (XYZ) block size for the Crescent model was selected to represent approximately one-third of the data spacing downdip (50 ft) on the veins while maintaining higher resolution along strike (25 ft) to account for higher grade variability identified during recent test mining.



The block model was coded with 3D wireframes of lithology and redox using a 50% rule, and block percent for topography and the silver gradeshells for maximum precision. Wireframe solids were also generated for the underground workings (Countess Decline, Alhambra and South Vein I-drifts). These were intersected with the model blocks and coded as voids for resource reporting purposes.

14.3 Density Model blocks were assigned a density based on rock type and oxidation from a dataset of 307 density measurements from 17 drillholes. The density assignments are summarized in Table 14.3.1.

Table 14.3.1: Density Assignment by Material Type

USC-SG SRK-TF Mat. Type Description n 2.92 10.96 Ox vn-A Alhambra vein oxide 15 3.29 9.73 vn-A Alhambra vein 10 2.94 10.88 ox vn-SV South Vein oxide 61 3.49 9.17 vn-SV South Vein 73 2.62 12.21 ox Wr oxide wall rock (HW) 19 2.72 11.76 wr non-oxide FW wall rock 117 2.54 12.60 wr non-oxide HW wall rock 12

Source: SRK

Specific gravity (SG) data were converted to tonnage factor (TF) for modeling and resource reporting. The units for TF are cubic feet per ton (ft3/t). SRK modeled an oxidation boundary in the South Vein using geologic logs and selective geochemistry. Unoxidized mineralized Alhambra and South Veins have higher density than their oxidized equivalents and their adjacent wall rocks. As expected, density in oxidized material is lower than its unoxidized equivalent.

14.4 Geology and Mineral Domains The predominant fault at the Project is the east-west striking, steeply south dipping Alhambra reverse fault. This structure displaces the stratigraphy of the mine area by approximately 3,000 ft such that the St Regis units in the hangingwall are in fault contact with Wallace Formation in the footwall and Revett within the hangingwall lies adjacent to St Regis in the footwall. The Alhambra mineral domain tracks the fault contact. The South Vein mineral domain lies parallel and 500 to 1,000 ft south of the Alhambra domain in the Alhambra hangingwall. The Jackson Vein lies between the Alhambra and the South Vein, striking slightly northwest relative to the other two veins. All three veins dip consistently from 70° to 75° to the south. A plan view of the drilling and underground development is presented in Figure 14.4.1. Figure 14.4.2 is a plan view of the veins (mineral domains) at the 3940 elevation. A cross section of the veins and modeled blocks in the veins is shown in Figure 14.4.3. The trace of the cross section is indicated on the Figure 14.4.1.

Surface oxidation has penetrated to approximately the 3890 ft level down-dip on the South Vein. SRK modeled oxidation from drill core and drift maps to support density assignments and recovery projections. Oxidation was not modeled on the Alhambra structure due to insufficient data. The modeled oxide in the South Vein is shown on Figure 14.4.4.

Mineralization occurs in narrow, through-going structures that have thousands of feet of continuity both in strike and downdip. SRK used Leapfrog® geologic modeling software and a 4 oz/t CoG to develop 3D wireframes for the three veins. The wireframes were used to code model blocks and



block percentages. The South Vein ranges in thickness from 1 to 20 ft with an average thickness of 3.5 ft. The Jackson Vein is a narrower high grade splay off the South Vein that was identified during test mining. Where the Jackson Vein is coincident with the South Vein, i.e. within the width of a 12 ft I-drift, the two veins were modeled together as the South Vein. The Jackson Vein was modeled separately where it lies more than 12 ft away from the South Vein and would likely require a separate drift or stope for mining. SRK benefited greatly from the recent I-drift mapping and sampling provided by United Silver to define the location of the mineralized veins.

14.5 Drillhole Database The drillhole database used for this resource estimate was compiled by United Silver and provided to SRK in a set of three MS Excel® spreadsheets. The spreadsheets were used to build a MSTorque® database. The data quality is sufficient for resource modeling. The database is comprised of drilling and drift-face or channel sample data. The channel sample data were collected as face and long hole samples while United Silver mined the Countess I-drifts. For resource modeling the channel samples were reformatted as drillholes.

Drillhole database statistics are presented in Table 14.5.1. The database consists of 279 drillholes and 481 channel samples containing 8,876 intervals for a total sampled length of 182,837 ft. Sampling in the core holes at Crescent was selective. Only zones of potential mineralization were sampled. Long intervals of barren rock before and after mineralized zones were not sampled, which is typical in the district. Average assayed sample intervals of both core holes and channels are less than 2 ft. The maximum drillhole length is 2,516 ft and the average hole length is 633 ft, from the database of both surface and underground drilling. All holes were drilled inclined to both the north and south to intersect the veins at high angles producing representative width/thickness intercepts. The data spacing on the mineralized veins is approximately 185 ft.

Table 14.5.1: Drillhole Database Statistics

Sample Types # of

Holes Average

Length

Total Length of

Drilling

# of From/To Intervals

# of Intervals with Ag Grades

Average Length of Ag Grade Intervals

in Gradeshells (n) (ft) (ft) (n) (n) (ft) Drillholes 279 633 176,627.8 6795 5447 1.5 Drift/Face-Channel Samples 481 12.9 6,209.5 N/A 1357 1.9 Totals 760 182,837.3 8876 6804 Source: SRK

Drillhole collar surveys were collected by professional land surveyors. Down-hole deviation surveys were obtained for most holes using a Flexit Multishot® instrument. Channel sample locations were collected by underground surveyors during drift advancement.

There is some uncertainty regarding the down-hole surveys of several of the very deep core holes that pierce the South Vein from the surface. SRK had difficulty reconciling several drill intercepts with the location of the vein as mined in the Countess I-drifts. This can be mitigated by using a gyroscopic survey instrument in future drilling.



14.6 Silver Assays Analysis SRK identified three errors in the database on the conversion of ppm to oz/t. These were corrected prior to modeling. The errors were in holes SNS-208 (690.8 ft) and H-109 (291.0 and 416.6 ft).

Silver assays statistics were analyzed by mineral domain (vein) and by drill type (drill or channel) in Tables 14.6.1 and 14.6.2, respectively. Grade times thickness calculations indicate that silver is fairly well distributed over increasing grade ranges. The South and Jackson Veins have higher mean silver grades than the Alhambra vein as evidenced by the box plot for silver in Figure 14.6.1. The box plot shown in Figure 14.6.2 demonstrates a significant high bias in channel (face/rib) samples compared to drillhole intercepts specific to the South Vein.



Table 14.6.1: Silver Assay Statistics Mineral Domain

Vein Cut-off Valid Length Inc. Pct. Avg. Sample

Length Grd-Thk Inc. Pct. Minimum Maximum Mean Std. Devn. Co. of Variation

(Ag oz/t) (n) (ft) (%) (ft) (oz/t-ft) (%) (Ag oz/t) (Ag oz/t) (Ag oz/t)

Alhambra

>=0.0 424 684.4 50 1.61 4511.2 22 0.00 193.38 6.59 15.92 2.41 >=4.0 136 188.8 14 1.39 3562.6 17 4.05 193.38 18.87 23.83 1.26 >=6.0 110 153.8 11 1.40 3406.8 16 6.01 193.38 22.15 25.43 1.15 >=8.0 94 131.3 10 1.40 3248.8 16 8.05 193.38 24.74 26.67 1.08

>=10.0 79 110.6 8 1.40 3068.2 15 10.10 193.38 27.74 28.12 1.01 >=12.0 67 94.9 7 1.42 2922.6 14 12.05 193.38 30.80 29.53 0.96

South

>=0.0 778 1,402.8 31 1.81 25538.1 18 0.01 441.80 18.21 37.19 2.04 >=4.0 479 846.9 19 1.77 24235.1 17 4.00 441.80 28.62 44.32 1.55 >=6.0 399 706.7 16 1.77 23574.2 17 6.00 441.80 33.36 47.16 1.41 >=8.0 333 587.9 13 1.77 22699.2 16 8.00 441.80 38.61 49.99 1.29

>=10.0 284 502.1 11 1.77 21943.8 16 10.00 441.80 43.70 52.49 1.20 >=12.0 262 460.4 10 1.76 21384.3 15 12.10 441.80 46.45 53.75 1.16

Jackson

>=0.0 193 385.0 35 2.00 6231.9 19 0.01 189.40 16.19 28.04 1.73 >=4.0 104 198.0 18 1.90 5716.1 18 4.05 189.40 28.87 33.36 1.16 >=6.0 86 157.4 14 1.83 5340.1 17 6.09 189.40 33.93 34.62 1.02 >=8.0 76 139.0 13 1.83 5212.7 16 8.04 189.40 37.50 35.32 0.94

>=10.0 67 119.7 11 1.79 4950.5 15 10.30 189.40 41.36 35.91 0.87 >=12.0 61 108.1 10 1.77 4796.0 15 12.10 189.40 44.37 36.27 0.82

Total

>=0.0 1,395 2,472.2 35 1.77 35589.5 19 0.00 441.80 14.40 31.36 2.18 >=4.0 719 1,233.6 18 1.72 33072.0 17 4.00 441.80 26.81 39.86 1.49 >=6.0 595 1,017.9 15 1.71 31930.0 17 6.00 441.80 31.37 42.42 1.35 >=8.0 503 858.2 12 1.71 30767.8 16 8.00 441.80 35.85 44.71 1.25

>=10.0 430 732.4 10 1.70 29593.3 16 10.00 441.80 40.41 46.86 1.16 >=12.0 390 663.4 10 1.70 28813.6 15 12.05 441.80 43.43 48.20 1.11

Source: SRK Length weighted assay values inside mineral domains



Table 14.6.2: Silver Assay Statistics by Sample Type

Sample Type

Cut-off Valid Avg.

Length Minimum Maximum Mean Variance

Co. of Variation

(Ag_oz/t) (n) (ft) (Ag_oz/t) (Ag_oz/t) (Ag_oz/t)

Face/Rib

>=0.0 860 1.97 0.05 441.80 18.85 1330.17 1.93 >=4.0 560 1.82 4.01 441.80 28.07 1799.05 1.51 >=6.0 473 1.82 6.00 441.80 32.33 2013.74 1.39 >=8.0 402 1.81 8.04 441.80 36.83 2234.64 1.28

>=10.0 345 1.81 10.00 441.80 41.42 2456.22 1.20 >=12.0 314 1.80 12.10 441.80 44.43 2598.21 1.15

Drill

>=0.0 535 3.92 0.00 193.38 7.24 343.43 2.56 >=4.0 159 1.57 4.00 193.38 22.37 830.20 1.29 >=6.0 122 1.49 6.13 193.38 27.66 962.88 1.12 >=8.0 101 1.43 8.00 193.38 31.95 1056.79 1.02

>=10.0 85 1.33 10.27 193.38 36.31 1136.82 0.93 >=12.0 76 1.21 12.05 193.38 39.33 1185.96 0.88

Source: SRK

14.7 Outlier Treatment Cumulative probability plots (CPPs) were analyzed for raw assay data within each domain to identify high grade outliers for restriction during grade estimation. The CPPs are shown in Figures 14.7.1 through 14.7.3. Inflection points for the CPP curves were identified at 45, 97 and 57 oz/t for the Alhambra, South and Jackson veins, respectively. SRK applied outlier restrictions during interpolation rather than capping raw assays. The outlier restriction strategy is discussed later in this document.

14.8 Compositing The assay data were composited using the mineral domains as a filter. Full vein width composites were built for drillholes bounded by the entry and exit pierce points of the drillholes through the veins. Composite statistics by domain are provided in Table 14.8.1.

Table 14.8.1: Crescent Composite Statistics by Domain

Vein Valid Minimum Maximum Mean Std. Devn. Co. of Variation (n) (Ag_oz/t) (Ag_oz/t) (Ag_oz/t)/

Alhambra 124 0 128.8 8.6338 14.6047 1.6916 South 366 0.05 224.5 15.9207 22.8833 1.4373 Jackson 114 0.01 73.1857 13.4867 18.5939 1.3787 Total 604 0 224.5 13.9653 20.82 1.4908 Source: SRK

The use of closely space channel samples introduced a change of support relative to the widely spaced drill core intercepts. To mitigate this, SRK declustered the channel samples to a 25 ft x 25 ft x 25 ft cell size and took the arithmetic mean of the composites falling within that new cell to create a single data point. In the exercise, channel samples that were typically spaced 5 ft apart were recomposited to a 25 ft grid spacing. A variety of declustered cell sizes were examined. The 25 ft x 25 ft x 25 ft cell was selected because the mean silver grades of that sized cell most closely matched the mean grade of the drillhole composites. Statistics for the declustered composites are provided in Table 14.8.2.



Table 14.8.2: Declustered Composite Statistics by Domain

Vein Valid Minimum Maximum Mean Std. Deviation Co. of Variation (n) (Ag_oz/t) (Ag_oz/t) (Ag_oz/t) Alhambra 84 0.01 52.6 6.407286 9.52593 1.486734 South 143 0.05 64.875 13.87331 14.13844 1.019111 Jackson 62 0.01 72.926 11.58932 15.90865 1.372699 Total 289 0.01 72.926 11.21327 13.73305 1.224715 Source: SRK

Declustering reduced the number of original vein composites by more than 50% commensurate with a moderate reduction grade and a lower coefficient of variation.

14.9 Variogram Analysis Variography was attempted on the composited data filtered to only the samples within each vein. Due to the highly variable nature of the grade distributions and the relative paucity of data points, no meaningful variograms were constructed. This is common in silver deposits of this type and does not necessary reflect negatively on the continuity of mineralization. It simply indicates that over relatively short distances the grades can vary significantly.

14.10 Estimation Methodology The search parameters for grade estimation were based on the spacing of mineralized intercepts in the plane of the vein and on mineral trends observed in the recent South Vein test mining. While the structures in the Silver Valley are continuous over thousands of feet on strike, neighboring operations report a stronger continuity of mineralization down-dip and commonly with a minor rake. The north-trending rake of mineralization in the plane of the South Vein is prevalent in the adjacent Bunker Hill and Sunshine Mines that flank the Crescent. Adjacent mines report a strong anisotropy of mineralization in 50 to 100 ft-wide shoots with downdip to strike ratios of between 2 and 3:1 (V:H). SRK adopted similar anisotropy for grade estimation.

Grade estimation for the Crescent model used the inverse distance squared algorithm (IDW). Two estimation passes were performed for silver in each mineral domain. The first pass used a short search range with moderate outlier restrictions and results were stored in blocks. The estimate continued with second pass with a longer search range, using composites with stronger outlier restrictions. Blocks were flagged at each estimation pass for subsequent use in the resource classification. Only declustered composites inside the mineral domains were estimated.

The minimum number of drillholes and composites per drillhole was varied by search pass. The short search pass required three composites from a minimum of two drillholes to inform a block. The long search pass allowed a block to be informed by a single drillhole to a maximum distance to data of 450 ft. The search ellipsoid was elongated downdip at an orientation to match that of each mineral domain. The estimated grades were multiplied by the partial percentages of the mineral domains (gradeshells) to enable the calculation of a single weight-averaged, block-diluted grade for each block. Estimation parameters used to define silver grades are provided in Table 14.10.1.



Table 14.10.1: Estimation Parameters for the SRK Crescent Block Model

Mineral Domain

Search Pass

Search Ellipse Range (ft) Search

Orientation (degrees)

No. Composites Outlier1

Major Semi-Major

Minor Z X YMin per

block

Max per

block

Max per

holeAu_oz/t ft

Alhambra Long 400 200 200 1 8 2 10 -25Short 400 200 200 190 -72 0 3 8 2 35 -50

South Long 450 250 250 1 8 2 15 -100Short 200 75 75 226 -62 0 3 8 2 25 -25

Jackson Long 450 250 250 1 8 2 57 -25Short 250 100 100 211 -68 0 3 8 2 57 100

Source: SRK Negative outlier distance indicates maximum distance away from data to apply the original composite grade, beyond which the restricted grade will apply. Positive outlier distance indicates maximum distance away from data to apply the original composite grade, beyond which zero grade will apply.

14.11 Model Validation Various measures were implemented to validate the Crescent resource block model. These measures included the following:

Comparison of drillhole composites with resource block grade estimates from all zones visually in both plan and section;

Statistical comparisons between block and composite data using histogram and cumulative distribution analyses;

Generation of a comparative nearest neighbor (NN) model; and Swath plot analysis (drift analysis) comparing the inverse distance model with the NN model.

Visual Comparison

Visual comparisons between the block grades and the underlying composite grades in plan and section show close agreement. Example model views showing block values and composite silver grades within their respective grade shells are provided in Figures 14.11.1 through 14.11.4.

Block-Composite Statistical Comparison

SRK also conducted statistical comparisons between the IDW blocks contained within vein envelopes and the underlying composite grades. A histogram comparison between block and composite silver grades is provided in Figure 14.11.5. This comparison shows that the model grade distribution for silver is appropriately smoothed when compared with the underlying composite distribution.

Comparison of Interpolation Methods

For comparative purposes, grades were also estimated using NN interpolation methods. The results of the NN model are compared to the IDW model at a zero CoG for the Project in Table 14.11.1 for all blocks. This comparison confirms the conservation of metal at a zero cut-off, and shows an overall agreement on both a grade and total metal basis for the deposit. The model was built intentionally conservative for high grade outlier composites that fall into the inferred classification. Block diluted grades were used in the resource statement for all metals.



Table 14.11.1: Comparison of IDW and NN Tonnage and Grade at a Zero CoG

Classification Model Tons (k) Ag Grade (oz/t) Contained Ag (koz)

Measured & Indicated IDW 2,257 5.25 11,850NN 2,257 5.53 12,491Percent Difference (IDW – NN) 0.00% -5.41% -5.41%

Inferred IDW 4,387 3.65 15,999NN 4,387 4.71 20,676Percent Difference (IDW – NN) 0.00% -29.23% -29.23%

Source: SRK

Swath Plots (Drift Analysis)

A swath plot is a graphical display of the grade distribution derived from a series of bands, or swaths, generated in several directions through the deposit. Using the swath plot, grade variations from the IDW model are compared to the distribution derived from the NN grade model and source composites.

On a local scale, the NN model does not provide reliable estimations of grade, but on a much larger scale it represents an unbiased estimation of the grade distribution based on the underlying data. Therefore, if the IDW model is unbiased, the grade trends may show local fluctuations on a swath plot, but the overall trend should be similar to the NN distribution of grade.

Swath plots were generated for silver along east-west and north-south directions for both the Alhambra and South Veins. The swath plots are shown in Figures 14.11.6 through 14.11.9, inclusive.

There is good correspondence between both veins modeled. The degree of smoothing in the IDW model is evident in the peaks and valleys shown in some swath plots; however, this comparison shows close agreement between the IDW and NN models in terms of overall grade distribution as a function of X and Y especially where there are high tonnages (bars on the plots). The plots also demonstrate the high degree of variance of the input composites and the model smoothing of the composite grades.

14.12 Resource Classification Classification of the resources for Crescent reflects the relative confidence of the grade estimates. Confidence is dependent on several factors including: sample spacing relative to geological and geostatistical observations defining the continuity of mineralization, mining history, SG determinations, accuracy of drill collar locations, and quality of the assay data.

Resources stated in this technical report were classified based on the following criteria and summarized in Table 14.12.1:

Measured Mineral Resources – SRK did not attempt to reconcile reported mine production to the model, primarily due to a lack of completeness of the production records. However, due to the fact that the test mining was a success and mill feed was produced, the resource was classified as Measured in the mining areas and immediately (one block) up and downdip of the I-drifts at Countess and Alhambra. Each block in the mining areas required information from at least eight composites at a maximum distance from source data of 110 ft.



Indicated Mineral Resources – Resource that was estimated with a minimum of two composites within a search radius of 100 to 300 ft (depending on domain). Since the composites are full-vein width, an Indicated classification represents each block being informed by at least two drillholes.

Inferred Mineral Resources – Resource that failed to meet criteria of Measured or Indicated but still fell within the interpreted mineral domain and a maximum search radius of 450 ft was classified as Inferred.

Table 14.12.1: Crescent Resource Classification Criteria

Classification Criteria by Mineral Domain Measured Indicated Inferred min # Comps Avg distance min # Comps Avg distance min # Comps Avg distance Alhambra 8 110 2 300 1 450 South 8 110 2 100 1 450 Jackson 8 100 2 300 1 450 Source: SRK

Classification is presented graphically in Figures 14.12.1 and 14.12.2 for the South Vein and Alhambra, respectively.

14.13 Mineral Resource Statement The SRK mineral resource is stated in Table 14.13.1. Resources stated in this technical report are in situ and are not constrained to a mining configuration. A CoG of 8 oz/t was applied. The CoG for the resource was determined using a silver price of US$20.00/oz, a recovery of 92%, combined mining and processing costs of US$133.00 per RoM ton and a 2% NSR royalty.

Table 14.13.1: Mineral Resource Statement for the Crescent Silver Project, SRK Consulting (U.S.) Inc., July 22, 2013

Alhambra South Vein Total Ag Cut-off




8.0 oz/t (t) (Ag oz/t) (Ag Moz) (t) (Ag oz/t) (Ag Moz) (t) (Ag oz/t) (Ag Moz)

Measured 9,000 13.2 0.1 51,000 17.2 0.9 60,000 16.6 1.0 Indicated 143,000 13.4 1.9 317,000 14.7 4.7 461,000 14.3 6.6 Measured and Indicated 152,000 13.2 2,0 368,000 15.0 4.5 520,000 14.4 7.5

Inferred 118,000 10.2 1.2 152,000 18.4 2.8 530,000 16.2 8.6 Source: SRK Notes:

Mineral resources that are not mineral reserves do not have demonstrated economic viability. No mineral reserves have been defined. The CoG for mineralized zone interpretation was 4 oz/t. The block CoG for defining Mineral Resources was 8 oz/t. A silver price used was US$20/oz, mining and processing costs of US$100/mill feed ton, and 92% mill recovery

were used to define the 8 oz/t cut-off. A 2% NSR royalty associated with the Project was applied in metal value calculations; The resources reported above are non-diluted. Measured Resources required blocks to be informed by a minimum of 8 composites and those blocks must be less

than 120 ft from previous production. Indicated Resources required blocks to be informed by composites from a minimum of two drillholes and distance

from data less than 300 ft The resources mined from the intermediate drifts have been deleted from the 2013 updated resources. Mineral resource tonnage and contained metal have been rounded to reflect the accuracy of the estimate and

numbers may not add due to rounding.



14.14 Mineral Resource Sensitivity In order to assess the sensitivity of the Crescent resource to changes in CoG, SRK summarized tonnage and grade at a series of increasing cut-offs by resource category. The sensitivity analysis for blocks within the resource pit is provided in Table 14.14.1 and 14.14.2 for Measured & Indicated and Inferred, respectively. The base case is bolded at the appropriate CoG to match the resource statement.

Table 14.14.1: Crescent Measured and Indicated Resource Sensitivity

Meas. & Ind. Alhambra South Vein Total

Ag Cut-off




4.0 oz/t 339,000 9.4 3.2 609,000 11.4 6.9 948,000 10.7 10.1 5.0 oz/t 301,000 10.0 3.0 538,000 12.3 6.6 839,000 11.5 9.6 6.0 oz/t 250,000 10.9 2.7 479,000 13.2 6.3 729,000 12.4 9.0 7.0 oz/t 209,000 11.8 2.5 422,000 14.1 5.9 631,000 13.3 8.4 8.0 oz/t 152,000 13.2 2.0 368,000 15.0 4.5 520,000 14.4 7.5 9.0 oz/t 126,000 14.5 1.8 322,000 16.0 5.1 447,000 15.6 7.0

10.0 oz/t 113,000 15.0 1.7 291,000 16.7 4.8 404,000 16.2 6.5 11.0 oz/t 96,000 15.8 1.5 255,000 17.5 4.5 351,000 17.1 6.0 12.0 oz/t 79,000 16.8 1.3 218,000 18.6 4.1 297,000 18.1 5.4

Source: SRK

Table 14.14.2: Crescent Inferred Resource Sensitivity

Inferred Alhambra South Vein Total

Ag Cut-off Mass Grade Cont. Metal



4.0 oz/t 393,000 7.3 2.9 783,000 12.2 9.5 1,176,000 10.5 12.4 5.0 oz/t 335,000 7.8 2.6 688,000 13.2 9.1 1,024,000 11.4 11.7 6.0 oz/t 253000 8.6 2.2 570,000 14.8 8.4 824,000 12.9 10.6 7.0 oz/t 201,000 9.1 1.8 482,000 16.3 7.9 683,000 14.2 9.7 8.0 oz/t 118,000 10.2 1.2 412,000 17.8 7.4 530,000 16.2 8.6 9.0 oz/t 55,000 12.6 0.7 351,000 19.4 6.8 406,000 18.5 7.5

10.0 oz/t 41,000 13.6 0.6 318,000 20.5 6.5 359,000 19.7 7.1 11.0 oz/t 90,000 14.9 0.4 287,000 21.5 6.2 317,000 20.9 6.6 12.0 oz/t 23,000 15.9 0.4 267,000 22.3 5.9 290,000 21.8 6.3

Source: SRK

14.15 Relevant Factors SRK is not aware of any unusual environmental, permitting, legal, title, taxation, socio-economic, marketing, or political factors that may materially affect the Crescent mineral resources as of the date of this report. The Project is in a detailed mine planning phase of development. Costs from that development planning were incorporated into the CoG calculation for the resource statement. The primary consideration for the conversion of resources to reserves will be the selected mining method and the control of dilution when mining these narrow mineralized zones.

Crescent Mine,

Kellogg, ID

Figure 14.4.1

Plan View of Drilling and Underground Workings Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.4.2

Plan View of Mineral Domains at the 3940 Elevation Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.4.3

Cross-Section View of Mineralized Veins and Model Blocks Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.4.4

Modeled Oxidation on the South Vein Source: SRK, 2013

Oxide

Countess I-drifts

Crescent Mine,

Kellogg, ID

Figure 14.6.1

Box Plot of Silver by Vein (grades in oz/t) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.6.2

Box Plot of Silver in the South Vein by Sample Type (grades in oz/t) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.7.1

Cumulative Probability Plot for Silver in the Alhambra Vein Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.7.2

Cumulative Probability Plot for Silver in the South Vein Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.73

Cumulative Probability Plot for Silver in the Jackson Vein Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.1

Declustered Composites on the South Vein (Looking N35E, -12) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.2

Modeled Silver on the South Vein (Looking N35E, -12) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.3

Declustered Composites on the Alhambra Vein (Looking N35E, -12) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.4

Modeled Silver on the Alhambra Vein (Looking N35E, -12) Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.5

Distribution of Model Blocks and Composites Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.6

Swath Plot – Alhambra East-West Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.7

Swath Plot – Alhambra North-South Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.8

Swath Plot – South Vein East-West Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.11.9

Swath Plot – South Vein North-South Source: SRK, 2013

Crescent Mine,

Kellogg, ID

Figure 14.12.1

Classification for the South Vein Source: SRK, 2013

Measured Indicated Inferred

Crescent Mine,

Kellogg, ID

Figure 14.12.2

Classification for the Alhambra Vein Source: SRK, 2013

Measured Indicated Inferred



15 Mineral Reserve Estimate Mineral Reserves have not been estimated for the Project, and are not appropriate at this stage of Project development. No formal engineering or economic work that would enable identification of reserves has yet been carried out.



16 Mining Methods The Crescent Mine operated until early September 2012. When operations were stopped the Countess Decline and the Big Creek #4 Adit (BC #4) were only partially completed, and had not yet been connected to create a secondary egress. Both the Decline and the Adit were advanced with mechanized mobile equipment. Vein development work, locally known as I-drifting, was being done using jackleg drills and mucked with mechanized mobile load haul dump (LHD) units. Material mined from the I-drifts was lower grade than expected because of significant dilution. Production goals emphasized tonnage rather than grade, which caused dilution of the vein material, and for mill feed grades to be lower than expected.

16.1 Current or Proposed Mining Methods

16.1.1 Current Mining Method

A modified cut and fill method was used during recent test mining. Conventional drilling, blasting and ground support techniques were employed. Material mined was loaded into underground trucks using rubber-tired LHD’s and transported to the surface. On the surface, the mineralized material was loaded into highway trucks and transported to the New Jersey Mill (NJMC) processing facility. Waste was stockpiled in designated areas near each portal.

16.1.2 Proposed Mining Method

Sections of the vein were selected for stoping based on modeled vein width and grade continuity. The mining plans include planned dilution of approximately 1 ft of width of barren material in addition to the vein width. For the South Vein, dilution was estimated at 25%; and for the Alhambra, 50%. Access and infrastructure development were designed to support the cut and fill method and sized based on mining equipment and production rate requirements. The production schedule was prepared for a production rate starting at approximately 100 t/d during the first 34 weeks required to complete the secondary egress. Following completion of the secondary egress, production will be scaled up to 400 t/d over 18 months.

The Project has two known veins, the Alhambra and the South Vein. Both are steeply dipping, narrow and will require ground support to mine effectively. Geological interpretation suggests that the Jackson Vein could link the South Vein fault with the Alhambra fault. The areas where the Jackson and South Veins are coincident have mill feed-grade mineralization.

16.1.3 Selection of Mining Method

There is limited geotechnical data available for the South Vein area. Based on knowledge from historic mining throughout the district, a modified cut and fill mining method called resuing was selected. The South Vein will be accessed via attack ramps from the main ramp, and a mill feed pass system will be built where the attack ramps meet the vein. The Alhambra Vein will be developed as captive cut and fill utilizing slushers, chutes and rail haulage. A mill feed pass system will be developed that passes mill feed to the Big Creek Number 4 (BC#4) adit which will be the main haulage level. Mill feed will be hauled to surface using rail haulage equipment.



16.1.4 Mine Development

Development mining to complete the secondary egress will be done with one and two boom jumbos, 4 yd3 LHD’s (Load Haul Dump units) and 17 to 20 t trucks to haul development waste to the surface.

Approximately 3,700 ft of the Countess Decline has been completed at a grade of -13%. The ramp has a cross section that is 10 ft wide and 13 ft high with an arched back. Figure 16.1.4.1 illustrates a typical cross section of the Countess Decline to complete the secondary egress. The decline will be extended for another 3,329 ft at a grade of -17% to intersect the BC #4 cross-cut at the 3250 level. The volume/tonnage of the ramp includes an allowance for muck bays, sumps and refuge chambers.

Approximately 2,900 ft of the BC #4 cross-cut has been completed at a grade of +0.5%. The cross-cut has a cross section that is 10 ft wide and 13 ft high with an arched back. To complete the secondary egress the cross-cut will be driven an additional 4,048 ft at a grade of -0.5% from its current face to the point where it intersects the Countess Decline at the 3250 level. The design of the cross-cut includes an allowance for muck bays, sumps and refuge chambers.

Following completion of the secondary egress, the BC #4 cross-cut will provide the main haulage system for transporting men and materials. A mill feed/waste pass system will be installed adjacent to the Countess Decline. A mill feed loading chute, with an air gate, will be constructed on the 3250 level near the intersection of the Countess Decline, and the BC #4 cross-cut for loading material into rail cars to transport to surface.

A long section depicting the proposed development is shown on Figure 16.1.4.1. Figure 16.1.4.2 is a typical cross section of the Countess Decline and the BC#4 cross-cut development.

16.1.5 South Vein Development and Mining

Per recommendations by rock mechanics consultant Wilson Blake, the Countess Decline will be kept a least 50 ft from the South Vein (Blake, 2013). Access to the South Vein from the Decline will be via attack ramps 9 ft wide and 11 ft high, at maximum grade of -20%. Each attack ramp will enable 50 vertical feet of vein to be mined beginning with an I-drift and followed by successive overhand mining cuts using resuing to separately blast mill feed and waste. Figure 16.1.5.1 shows a cross section on the I-drift with the initial mining cut on the South Vein. Figure 16.1.5.2 illustrates the South Vein stope cut with cement backfill and waste backfill in the previous cut.

16.1.6 Alhambra Vein Development and Mining

Development cross-cuts about 1,000 ft long will be driven from the Decline to the Alhambra Vein on the 3450, 3650 and 3850 ft elevations. Level drifts will be driven along the strike of the vein to locate the mineralized shoots for mining. Track will be installed on each of the Alhambra cross-cuts and levels. The Alhambra stopes will be developed as captive cut and fill, and the mill feed will be hauled by rail to the mill feed pass system to be developed at the Countess Decline.

The Alhambra Vein will also be mined cut and fill but will be captive, utilizing slushers, chutes and man ways. By separating the mineralization into multiple areas with separate mill feed passes there will be multiple mining faces available for scheduling, which will facilitate achieving targeted mill feed tonnage and grade blending. As the mill feed is defined further, more working faces will be developed. Figure 16.1.6.1 shows a typical stope configuration in the Alhambra Vein.



Mill feed cars will be emptied onto an existing surface stockpile. From the stockpile, the mill feed will be loaded onto 20 ton highway trucks and hauled to the New Jersey Mill.

16.2 Parameters Relevant to Mine Designs and Plans

16.2.1 Geotechnical

The proposed dimensions of underground workings are summarized in Table 16.2.1.1. These dimensions are determined by equipment requirements for the existing and planned fleet. The recommended ground support for planned development is 6 ft and 4 ft split sets placed on 4 ft centers (Blake, 2013). The 6 ft split sets will generally be in the back and the 4 ft split sets in the ribs. Wire mesh will be placed above 5 ft from the sill and across the back. The Project is well above the areas of high natural rock stresses and a rock burst is highly unlikely (Blake, 2013).

Table 16.2.1.1: Dimensions of Working Areas

Type H – ft W- ft Countess Decline 10 13 Track Drift 8.5 8 I Drifts – w/Jackleg 8 7.5 I Drifts – w/Jumbo 12 7.5 Access/Attack Ramp 9 9

16.2.2 Hydrological

The planned mining area is well above the local ground water table. The only ground water inflows expected during mining are small volumes of perched ground water.

16.3 Mine Optimization Mine design was based on the wireframes of the mineralized zones. No mine optimization has been completed at the time of this report.

16.3.1 Development and Operations

The major development is partially completed and the surface infrastructure is in place. The remaining development is the 3,329 ft to complete the Countess Decline, and the 4,048 ft extension of the BC #4 cross cut. During development, test mining will be done on the South Vein to refine the resuing and backfill systems before going into full production. Development mucking will be done by two 4 yd3 LHD’s dumping into 17 and 20 t trucks which will haul the waste rock to a surface stockpile. When stoping commences, some of the waste material will be transferred directly to stopes for backfilling.

16.3.2 Ground Support

Planned ground support will consist of 6 ft and 4 ft split sets placed on 4 ft centers. The 6 ft split sets will generally be in the back and the 4 ft split sets in the ribs. Wire mesh will be placed up 5 ft from the sill, across the back and to within 5 ft of the opposing sill. Drilling for ground support will be undertaken by the development jumbos or jacklegs



Additional geotechnical studies will be done to refine ground support requirements as mining progresses.

16.3.3 Drilling

Drilling within the stopes will be carried out using jackleg drills, or in some cases a single boom jumbo. Air and water for the drills will be run through the access/attack ramp into the stope areas.

16.3.4 Blasting

Mining will use standard blasting techniques and vary depending on the ground conditions, drill spacing and presence of ground water. The primary blasting agent will be ANFO with stick explosives used in wet conditions. Additional technical studies will be done to refine the blasting techniques.

16.3.5 Backfill

The stopes will be developed using a reusing mining method. This requires the mill feed and waste be selectively blasted and removed. The waste will be left in the stope and will become the working floor when the next cut of mill feed is mined. A thin layer of cement will be placed on top of the waste to prevent mill feed dilution in the next cut of mining.

16.3.6 Ventilation

Total required airflow in the mine is estimated at 240,000 cfm. Air intake will be through the BC #4 and Hooper portals and exhausted through the Countess Decline. Dedicated ventilation fans will be located at each stope access area. Fresh air intake fans will be installed at the BC #4 and as necessary. The air flow maybe seasonally reversed.

The BC #4 portal will serve as the main entrance for men and materials. The Countess Decline will be the secondary escape-way. Surface mounted fans will be used to provide a negative pressure ventilation system. A total fan power of 200 kW will provide the approximately 240,000 cfm required to ventilate the mine.

16.3.7 Utilities: Compressed Air, Water and Electric Power

During development of the Countess Decline, compressed air and electric power will be delivered to the mine by diesel powered compressors and generators at the Countess Portal. At the BC #4, compressed air will be piped to the portal from the Big Creek mine office and electric power lines will be extended from the mine office to the portal. After the Countess Decline breaks through to the BC #4 Adit, all utilities will come from the BC #4 portal.

16.3.8 Health and Safety

Refuge stations will be located within a maximum fifteen minute walk from all working areas. These refuge stations will be constructed in the drifts, or be the portable self-contained type that can be relocated as required by the mining schedule.

Carbon monoxide and oxygen sensors are located at strategic positions underground. These sensors can be remotely readable at the security center. If the sensors indicate a high concentration of toxic gases outside of the blast time, alarms will indicate to control personnel that a fire has



occurred. A stench gas emergency warning system is in place. This system can be activated to warn those working underground a fire or other emergency. USC is a participant in Central Mine Rescue, a Silver Valley organization that provides 24-hour mine rescue services for all participating members. As the workforce increases, mine rescue teams will be trained and equipped to deal with underground emergency situations. Mine rescue assistance agreements are in place with the other mines in the area.

At this time the mine has no second egress. This problem will be eliminated when the Countess Decline connects with the BC #4. MSHA has agreed to let development and incidental mining occur while this connection is being created.

16.3.9 Organization and Staffing

At full production the Crescent Mine will have a total staff of 49, consisting of 17 salaried and 32 hourly employees. A summary of the planned staffing requirements is shown in Table 16.3.9.1. The organizational chart is shown in Figure 16.3.9.1.

Table 16.3.9.1: Staffing at Full Production

Role Number of Staff Management and Supervisory, Salaried 6 Technical Services, Salaried 11 Miners, Mechanics and other laborers, Hourly 32 Total 49

16.4 Mine Production Schedule The mine will begin operation at 100 t/d and gradually increase production to 400 t/d over an 18 month period. The mine life, based on the currently known Measured, Indicated and Inferred resource is approximately 5.5 years.



16.4.1 Mine Production

The production schedule is shown in Table 16.4.1.1.

Table 16.4.1.1: Production Schedule

Total Year Item Units or Average 0 1 2 3 4 5Mine Production

Alhambra Vein Development ft ft 7,735 1,500 1,235 2,000 2,000 1,000Raise Development ea 10 0 3 3 3 1Vein Development Tons Mill Feed kt 13 2 2 3 3 2 0Alhambra Stope Mill Feed kt 134 0 25 32 32 34 11Alhambra Mill Feed Production kt 147 2 27 35 35 35 11

South Vein I drift ft 4,800 1,600 1,600 800 800 0I Drift Mill Feed Tons kt 15 5 5 2 2 0South Vein Stope Tons kt 290 7 50 68 70 70 25South Vein Mill Feed Production kt 305 7 54 73 73 73 25Total Mill Feed - kt 451 9 81 108 108 108 36

Mill Feed Summary Average Undiluted Grade Alhambra Ag oz/t 12.6 12.6 12.6 12.6 12.6 12.6 12.6South Vein Ag oz/t 15.7 15.7 15.7 15.7 15.7 15.7 15.7Contained Metal Alhambra Ag koz 1,843 31 335 444 444 444 143South Vein Ag koz 4,796 107 856 1,146 1,146 1,146 395Total Contained Ounces koz 6,639 138 1,192 1,590 1,590 1,590 539

The summary of production and ounces produced is shown in Table 16.4.1.2.



Table 16.4.1.2: Production Summary

Total Year Item units or Avg. 0 1 2 3 4 5UNDERGROUND MINING Preproduction Development

Development Footage Countess Decline ft 3,329 3,329 Big Creek # 4 ft 4,048 4,048 Attack Ramps ft 188 188

7,565 7,565 - - - - -Post Production Development

Attack Ramps ft 1,002 407 238 238 119 - Mill Feed Pass ft 1,381 711 670 3250 Alhambra Drift ft 491 491 3450 Alhambra Drift ft 1,007 504 504 3650 Alhambra Drift ft 1,002 1,002 3850 Alhambra Drift ft 734 734 Hooper Decline ft 4,000 3,200 800

9,617 0 5,313 2,212 1,974 119 -Mine Production

Alhambra Vein Development ft ft 7,735 1,500 1,235 2,000 2,000 1,000 Raise Development ea 10 0 3 3 3 1 Vein Development Tons Mill Feed kt 13 2 2 3 3 2 0 Alhambra Stope Mill Feed kt 134 0 25 32 32 34 11 Alhambra Mill Feed Production kt 147 2 27 35 35 35 11

South Vein I drift ft 4,800 1,600 1,600 800 800 0 I Drift Mill Feed Tons kt 15 5 5 2 2 0 South Vein Stope Tons kt 290 7 50 68 70 70 25 South Vein Mill Feed Production kt 305 7 54 73 73 73 25 Total Mill Feed - kt 451 9 81 108 108 108 36

Mill Feed Summary Average Undiluted Grade Alhambra Ag oz/t 12.6 12.6 12.6 12.6 12.6 12.6 12.6 South Vein Ag oz/t 15.7 15.7 15.7 15.7 15.7 15.7 15.7 Contained Metal Alhambra Ag koz 1,843 31 335 444 444 444 143 South Vein Ag koz 4,796 107 856 1,146 1,146 1,146 395 Total Contained Ounces koz 6,639 138 1,192 1,590 1,590 1,590 539

PROCESSING Flotation Milling

Diluted Tons Alhambra Mill Feed kt 220 4 40 53 53 53 17 South Vein kt 381 9 68 91 91 91 31 Total Tons In kt 601 12 108 144 144 144 48 Contain In Contained Oz – Ag koz 6,639 138 1,192 1,590 1,590 1,590 539 Diluted Grade – Ag oz/t 11.1 11.3 11.0 11.0 11.0 11.0 11.1 Recover Ag 92% koz 6,108 127 1,096 1,463 1,463 1,463 495 Concentrate Mass 420 t 13,221 275 2,373 3,167 3,167 3,167 1,072

Grade Control

Following is a description of the grade control practices that will be implemented with the resumption of mining at Crescent. This proposed sampling system will quantify silver content as multiple points in the mining process. The following list itemizes the types of samples that will be collected:



1. Chip channel samples; 2. Muck samples; 3. Haul truck samples; 4. Rail car samples and 5. Mill feed samples.

Each sample type is described below.

Chip Channel Sampling

When driving exploration drifts on a vein or developing a stope with an I-drift development drift on a vein, chip channel samples will be taken on every face blasted. The resultant assays will be used to mark up the face for the next round and, as a rule, the next round will not be drilled and blasted until after assay results are received and the face can be marked up by a supervisor or geologist.

After washing down the face and back, the geologist will determine where samples will be taken across the face. Samples will be taken across horizontally across the face and the entire width of a face will be sampled. Individual sample widths will be based on geologic features such as color, sulfide, mineralization, structure and rock type, but no individual sample will be longer than 2 ft. Each individual sample should be at least 5 lb in order to obtain a representative result from each individual sample.

The face of each round blasted will be sampled with two sets of chip channel samples, one about 2 ft above the floor of the stope or I-drift and one about 5 ft above the floor. A third chip channel sample will be taken across the back of each round, in the middle of the round. (For a 6 ft round, the sample will be 3 ft from the face.)

Samples will be identified by mine level, stope and location within the stope with unique numbers or coding. The geologist will draw sketch maps of the back and face and indicate on the maps where each sample was collected. Photographs of each face will be taken to enable assay data to be correlated visually with geologic features and to assist in the preparation of geologic maps that will be provided to the miners to assist the miner is mining mineralized material with minimum dilution.

The geologist collecting the samples is responsible for transporting the samples to a secure area for transport to the assay laboratory.

Muck Samples

Muck samples will be the responsibility of the miner who is mucking mineralized material from and face where mineralized material has been blasted. For each round, the miner will collect three muck samples. Each sample will be approximately 5 lb and will be collected evenly across the width of the muck pile. One sample will be collected from the front of the muck pile (the part furthest away from the face of unbroken rock); the second sample will be collected from the middle of the muck pile after approximately half the muck pile has been mucked. The third sample will come from near the face of the muck pile.

Samples will be identified similar to the chip channel samplers for mine level, stope and location within the stope. Collected samples will be brought out of the mine at the end of shift or passed off to a shift boss or geologist to transport to a secure storage area to await transport to the assay laboratory.



Haul Truck Samples

Truck samples will be collected in two locations. The first is on the surface when the truck transporting the mineralized material to surface dumps it into a stockpile for each round. The truck driver hauling material out of the mine will collect approximately 2.5 lb of material from the muck pile after the driver dumps the truck on the surface. Material hauled to surface will be stockpiled by round and the samples from each round will be combined.

When material is hauled to the mill, truck drivers will collect a sample from the material after dumping it on the mill lay down pad. Samples from every five truck loads will be composited. Samples will be identified if possible, but it is expected that material hauled to the mill will be mixed as trucks must be fully loaded for hauling to the mill.

Rail Car Samples

Following completion of the secondary egress and installation of rail in the BC #4 cross-cut, mineralized material will be hauled from the mine with rail cars pulled by a battery driven locomotive. There will be 6 to 10 cars in every train hauled from the mine, and before dumping the cars on the surface about a pound of material will be collected and composited by shifts.

Mill Samples

Mill head samples are taken after the mineralized material is crushed and screened to size and fed into the ball mill. An automated sampling apparatus is part of the mill circuit, which is described in Section 15, Recovery Methods. Tailings samples are also collected by an automatic sampler.

Assay data collected from the several sampling methods and locations and will be analyzed and reconciled using several criteria to reconcile the grade of material being produced at the face with the material being recovered and sent to tails at the mill.

After commercial production is achieved, weekly and monthly reconciliations will be ongoing to check actual silver recovery with what the face and muck samples are indicating for silver produced. When reconciliations vary more than acceptable between mill and mine, sampling protocols and practices will be reviewed to insure that practices are following the protocols.

16.5 Waste and Stockpile Design Waste Rock Storage Facility

Development waste rock will be hauled from the mine to the surface in 18 ton trucks. It will be dumped near the portal of the Countess Decline and the Big Creek #4 cross cut, then loaded into highway trucks and hauled to the North American Dump. The North American Dump area has the capacity to for the planned development waste.

The North American Dump surface rights are currently owned by others. USC plans to purchase the surface rights. If negotiations are not successful, USC plans to negotiate with the timber company that holds the surface rights adjacent to the Countess Decline and the Big Creek #4.

Waste rock developed in the stopes is planned to be used as stope backfill.



Mill Feed Stockpiles

Mill feed hauled to the surface will be stored in temporary stockpiles at the portals. It will then be hauled in highway trucks to the New Jersey Mill. There is adequate room for these stockpiles.

16.6 Mining Fleet and Requirements The Crescent mine has an existing fleet of mining equipment, most of which is in good condition. The existing major mining equipment is shown in Table 16.6.1. The Project will require two additional 18 t haul trucks and, two 4 yd3 LHD’s. There is no requirement for additional drill jumbos. The additional mining equipment is shown in Table 16.6.2. The prices shown are for delivered and assembled equipment.

Table 16.6.1: Existing Mining Equipment

Quantity Item Quantity Item 5 Locomotives 1 LHD – 1.5 yd3 13 Granby Mill Feed Cars, 4 t 1 LHD – 1.25 yd3 6 Rocker Dump Mill Feed Cars, 2 t 1 Jumbo – Track 2 Track Mucker 1 Jumbo – Single Boom 1 LHD – 3.5 yd3 1 Jumbo – Double Boom 1 LHD – 2.0 yd3

USC also owns a fleet of support equipment including diesel powered generators and compressors, slushers, prill loader, skid steer loader, on highway dump truck, loaders to load the dump truck, and underground rated utility vehicles which are sufficient to operate the mine.

Table 16.6.2: Cost of Additional Mining Equipment

Quantity Item Per Unit Total Remarks/Basis 2 Haul Truck – 18 t $496,000 $992,000 Quote 2 LHD – 4.0 yd3 $555,000 $1,110,000 Quote Total $2,102,000

16.7 Mine Dewatering No dewatering is necessary to mine the resource area discussed above.

PROPOSED COUNTESS DECLINE

ALHAMBRA VEIN ACCESS

EXISTING COUNTESS DECLINE

BIG CREEK #4 - EXISTINGBIG CREEK #4 CROSS CUT - PROPOSED

HOOPER LEVEL - EXISTING

FILE NAME: Detail_Sketches_BCH.dwg

DRAWING TITLE:

IF THE ABOVE BARDOES NOT MEASURE 1 INCH,

THE DRAWING SCALE IS ALTERED

PROJECT:

PREPARED BY:

PROPOSED DEVELOPMENT LONG SECTIONDESIGN: KH

DRAWN: BCH

REVIEWED: KH

APPROVED: KH

CRESENT SILVER PROJECT AFIGURE 16.1.4.1

202500.030

DRAWING NO.: REV. NO.

09/06/2013DATE:

SRK JOB NO.:

H:\United Silver\900 Dwg GIS\Dwg_Working\Detail_Sketches_BCH.dwg

LOOKING NORTHDIP - 15°

NOT TO SCALE

10'

13'

AIR DUCTØ3.5'

5'

18 TON TRUCK

ARCHED BACK

LEAKY FEEDER (RBs)

AIR WATER OVERUP AND MESH

DITCH

ELECTRICALCABLE

7.5'

7.5'


DRAWING TITLE:

PROJECT:

PREPARED BY:

COUNTESS DECLINE AND BD#4DEVELOPMENT CROSS SECTION


202500.030


09/04/2013DATE:

SRK JOB NO.:


FEET

210



DESIGN: XXX

DRAWN: BCH

REVIEWED: KH

APPROVED: XXX

8'

5'

3.5'

7'

72°

DRILL HOLES

SOUTH VIEN


DRAWING TITLE:

PROJECT:

PREPARED BY:

SOUTH VEIN I-DRIFT CROSS SECTION


202500.030


09/04/2013DATE:

SRK JOB NO.:


FEET

52.50



DESIGN: XXX

DRAWN: BCH

REVIEWED: KH

APPROVED: XXX

9'

7'

6'

3.5'

72°

BLAST HOLE

WASTEBACKFILL

CEMENT FILL

CEMENT BACKFILL ONLY1/2 - 1 FT THICK

DRILL HOLES

WASTE MATERIAL TO BEBLASTED DOWN FOR BACKFILL


DRAWING TITLE:

PROJECT:

PREPARED BY:

SOUTH VEIN STOPE CUT


202500.030


09/04/2013DATE:

SRK JOB NO.:


FEET

52.50



DESIGN: XXX

DRAWN: BCH

REVIEWED: KH

APPROVED: XXX

MILL HOLEMANWAY &VENTILATION

BACKFILL

PILLAR

BACKFILL

PILLAR

LEVEL DRIFT 10 H

+210'

MANWAY &MATERIALS

MANWAY &MATERIALS

MINING LEVEL

8'


DRAWING TITLE:



PROJECT:

PREPARED BY:

TYPICAL ALHAMBRA STOPE DEVELOPMENTDESIGN: KH

DRAWN: BCH

REVIEWED: KH

APPROVED: KH


202500.030


09/06/2013DATE:

SRK JOB NO.:


FEET

30150


Kellogg, Idaho

Figure 16.3.9.1

Mine Organization Source: SRK, 2013

Mine Manager

1 – 0

Safety, Health &

Secretary

Environmental

1 – 0

1 – 0

Purchasing

(Full time in 2014)

1 – 0

Operations

Maintenance

Technical Services

0 – 27

1 – 4

7 – 2



17 Recovery Methods Mineralized material from the South Vein was mined from three sublevels on 50 ft vertical intervals, at approximately 3,940, 3,890, and 3,840 ft amsl. Mineralized material from each sublevel was stockpiled separately, and then hauled to the New Jersey Mill. Care was taken at the mine and the mill to keep the mineralized material from each sublevel segregated so the milling results are representative of the material mined by sublevel. Table 17.1 presents the calculated head grade of the milled South vein material.

The calculated head grade was determined by adding the grams of silver in concentrate, as determined by the concentrate assays used for preliminary payments, to the grams of silver in the tailings as indicated by the tailings assays. The tailings were sampled by an automatic slurry sampler at the mill where a 3 ml sample is taken every two minutes. The mill feed tonnage was obtained from a weightometer on the ball mill feed belt which was calibrated prior to the commencement of milling operations. A sample of the ball mill feed was collected manually every half-hour and combined into a single sample per shift. Mill feed moisture content was found by weighing this sample wet and weighing the sample again after drying in an oven.

Table 17.1: Head Data for South Vein Mining

Source of South Vein Mill Feed Tons Processed (dry) Ag Grade (oz/t)Total 12,607.0 7.57 3940 Sublevel 3,703.0 8.98 3890 Sublevel 5,893.3 6.56 3840 Sublevel 3,010.7 7.80 Source: NJMC, 2013

Only silver and copper were assayed for on a shift-by-shift basis. Copper assays were not obtained for the final 30% of the concentrate tonnage so the copper head grade was not calculated.

17.1 Operation Results Table 17.1.1 presents the New Jersey mill’s mass balance for the South vein mill feed that was processed at the mill in 2012 and early 2013.

Table 17.1.1: Mass Balance for the South Vein Milling Campaign 2012 & 2013

Product Dry Weight Tons Ag (oz/t) Ag Distribution by Mass (%) Feed – Calculated 12,607.0 7.57 100.0% Feed- Assay 12,607.0 7.27 Concentrate 250.6 340.67 89.5% Tails 12,356.4 0.80 10.5% Source: NJMC, 2013

The overall silver recovery to final concentrate was 89.5% with an average concentrate grade of 340.67 oz of silver per ton. However, silver recovery varied significantly by mining sublevel, and increased with decreased oxidation at depth. The 3940 sublevel had the highest degree of visible oxidation; the 3840 level had the least. Selected metallurgical data by mining sublevel is presented in Table 17.1.2.



Table 17.1.2: Assay and Recovery Data for the South Vein by Sublevel

Mining Sublevel

Dry Weight Tons

Calculated Head Grade – Ag

(oz/t)

Tailings Grade – Ag (oz/t)

Ag Recovery to Concentrate

(%) 3940 3,703.0 8.98 1.32 85.3% 3890 5,893.3 6.56 0.68 89.6% 3840 3,010.7 7.80 0.38 95.1% Source: NJMC, 2013

Silver recovery increased by nearly 10% from the 3940 to the 3840 sublevel. The head grade between these two levels is similar. Though these levels are only 100 ft apart vertically, there was a noticeable visual difference in mill feed. The 3940 mill feed had a reddish-brown color and copper oxide could also be found while the 3840 mill feed had a more gray or “fresh” appearance. Other process variables, such as grind and flotation reagents, remained relatively constant during the processing of the sublevels. Electron microprobe of the tailings samples, which are stored at the mill site, would help to confirm if oxidation was the primary factor in the variable recovery.

17.1.1 Concentrate Sales

A total of 250.6 t of concentrate were produced from processing 12,607 t of mill feed for a ratio of concentration equal to 50.3. The concentrate was sold to Essential Metals of Kellogg, Idaho and Ocean Partners of Wilton, Connecticut. Table 17.1.1.1 provides a summary of all the concentrate produced during the South vein milling campaign of 2012 and 2013. Copper content is not included because it is not a saleable product, and 53 t of concentrate that were sold to Ocean Partners were not assayed for copper. The copper assay varied from 13% to 15%. It may be helpful in the future to re-assay the splits from the shipments to Ocean Partners for copper to complete the metallurgical balance.

Table 17.1.1.1: Selected Elements in Concentrate

Dry Ton Ag (oz/t) As (%) Sb (%)250.6 340.7 4.95 3.79 Source: NJMC, 2013

17.1.2 Mill Reconciliation with Mine Production

Chip sample results from each I-drift face were used to estimate the grade of material mined and sent to the mill. Separate stockpiles were designated and managed for each of the 3940, 3890 and 3840 I-drift levels, both at the mine portal and at the mill. The tons and grades calculated from each I-drift was compared to the tons and grade of mill feed for each of the I-drifts. Results of this comparison are reported in Table 17.1.2.1.

Table 17.1.2.1: Mine to Mill Reconciliation

Mining Sub

Level

Mine Mill Variance

Mine to Mill Dry Tons oz/t Contained oz Dry Tons oz/t Contained oz Tons Oz Ag %

3950 3,749 14.5 54,276 3,703 9.0 33,241 -46 -21,035 -39% 3900 5,796 7.7 44,343 5,893 6.6 38,638 97 -5,705 -13% 3850 2,802 9.7 27,060 3,010 7.8 23,487 208 -3,573 -13%

Source: USC, 2013



The mill reported significantly lower grade and contained ounces than calculated from chip samples, especially for the 3940 I-drift, which had a different materials handling protocol. After the material handling process was improved, the discrepancy was reduced to 13% of contained ounces for the 3890 and 3840 I-drifts. This remaining discrepancy is believed to be due to sampling bias, loss of fines in handling, incorrect tonnage factor, or some combination of the three.

In September 2012, SRK Consulting reviewed USC’s sampling procedures and recommended that individual chip samples be limited to no more than 2.0 ft in width, and weigh at least 5 lb. SRK also recommended mapping the test drifts in greater detail each day.

17.2 Processing Methods In early 2011, USC and New Jersey Mining Company (NJMC) formed a joint venture (NJMC 65.2% and USC 34.8%) to expand and operate the New Jersey mill capacity to 440 t/d (18.4 t/h). Construction started in the spring of 2011 and was completed in June of 2012. The new mill was designed to produce a single flotation concentrate and included a new crushing plant, larger mill building, new fine mill feed bin, new ball mill, a refurbished large rougher flotation cell, new cleaner flotation cells, a new paste thickener, a concentrate filter, new pumps, and associated conveyors and piping. The new mill was commissioned in late June of 2012 with mill feed from the Crescent’s South vein. A total of 12,607 t of Crescent South vein mineralized material were processed from June 2012 through April 2013. The flow sheet for the expanded plant is shown in Figure 17.2.1.

17.3 Plant Design and Equipment Characteristics The new and expanded mill was designed to produce a single flotation concentrate. RoM material is fed by a front-end loader into a bin with a belt feeder that feeds into a 15 inch by 24 inch jaw crusher. Feed is crushed to minus 3 inches and then conveyed to a vibratory screen with a 0.5 inch by 2 inch slotted screen-cloth. Screen undersize drops into a 385 t fine mill feed bin while the screen oversize is conveyed to a Metso HP100 cone crusher where the closed-side setting is about 0.5 inch. The cone crusher product is then conveyed back to the screen in a closed loop for rescreening.

The fine mill feed bin discharges onto the bin discharge belt conveyor which, in turn, discharges onto the ball mill feed belt conveyor. A weightometer is located on the ball mill feed belt and the real-time feed rate in ton per hour is displayed and monitored by the mill operators. A variable-frequency drive (VFD) on the under-bin conveyor allows the mill operators to manually adjust the speed of the fine mill feed bin discharge conveyor to vary the ball mill feed as required. A screw-type lime feeder, also controlled with a VFD, is used to add lime to the ball mill feed. Lime is added at a rate of about 1.12 lb/t to depress pyrite in the flotation circuit.

The ball mill is a 250 kW Marcy-type ball mill that is 8 ft in diameter and 13 ft long. Mill feed is fed to the ball mill via a spout feeder. A trommel screen on the ball mill discharge removes pebbles, wood, and other debris. The ball mill discharge is pumped to a 10 inch Warman hydro-cyclone where cyclone underflow reports back to the mill for additional grinding, and cyclone overflow is conveyed via gravity in a 3 inch diameter pipeline to the rougher flotation cells.

The cyclone overflow slurry is sampled every two minutes by an automatic sampler and sample size is 3 ml. The slurry is fed to a vibratory screen for additional trash removal before being fed to the



rougher flotation circuit. Flotation reagents are fed into the rougher flotation circuit via metering pumps. The reagents added are the collector, Z-55, and a 10% NaCN solution which is used to depress pyrite. The rougher circuit consists of one Wemco 144 and five Wemco 66D flotation cells. Rougher concentrate is fed to two banks of three Wemco 36 cleaner flotation cells operated in series for two stages of cleaning. No reagents are added at the cleaner cells. Concentrate from the second stage of cleaner cells forms the final concentrate which is piped to a 5 ft diameter thickener where it is then transferred to agitated tanks for storage. Cleaner tailings are recycled back to the first rougher flotation cell. The final concentrate is either pumped into a concrete mixing truck for delivery to a local refinery or filtered with a plate and frame pressure filter and placed in 1 t supersacks for delivery to a foreign smelter.

Tailings from the rougher flotation circuit are pumped into two Deep Cone Thickeners (DCT) operated in parallel to produce paste tailings. Flocculant is added to the tailings stream to promote settling in the thickeners and produce a clear overflow. Thickener overflow is piped to a process water storage tank. Make-up water from an adjacent well is added to the process water tank. A pump then distributes the process water to various locations throughout the mill. Underflow from the paste thickeners is pumped to a paste tailings stack east about 500 ft east of the mill building. Peristaltic pumps are used to deliver the tailings to the stack where the tailings form a slope of about 6% with very little bleed water.

17.3.1 Crushing and Grinding Circuits

The South vein mill feed was crushed to pass 0.5 inch and the crushing plant throughput averaged about 66 t/h. Very little clay was observed in the South vein mill feed. A mill circuit survey determined the Operating Work Index of the ball mill to be 9.65 kWh/t (Rust, 2012). At the time of the survey, the ball mill produced a cyclone overflow with 80% passing 72 µm from feed that was 74% passing 0.25 inch. The particle size distribution from the grinding circuit survey is shown in Figure 17.3.1.1.

Note that the grind achieved during the South vein milling campaign was significantly finer than the grind achieved by previous testing and bulk sampling (80% passing 72 µm versus 80% passing 86 to 100 µm).

17.3.2 Flotation Circuit

Two collectors, Z-55 and Aerofloat 242, were tested during the South vein milling campaign. No significant difference between either silver recovery or concentrate grade obtained from the reagents was noted. Thus, Z-55 was selected as the primary collector because it does not require MIBC for frothing, resulting in cost savings. The flotation reagent scheme is presented in Table 17.3.2.1.

Table 17.3.2.1: Flotation Reagent Scheme

Reagent Description Addition Rate Location Lime Dry hydrated lime. 1.12 lb/t Ball Mill Feed Belt Z-55 Straight solution. 0.040 lb/t Rougher Cell No. 1 Sodium Cyanide 10% solution (mixed with water and lime). 0.015 lb/t Rougher Cell No. 1

Lime was the most important reagent to optimize concentrate grade. If the pH in the flotation circuit was less than 8.0, the silver grade of the concentrate declined substantially because more pyrite reported to the final cleaner concentrate. Pyrite was visible in the froth when the solution pH was less



than 8.0. If the solution pH was greater than about 8.5, the froth had a darker color and silver grade in the concentrate increased by about 63%.

The New Jersey mill flotation circuit has two cleaner flotation cell banks, which can be configured to run in either series or parallel. Both modes of cleaner cell operation were tested during the South vein milling campaign. Series operation produced a higher silver concentrate grade (414 oz/t) than parallel operation (358 oz/t). The silver head grade for both tests was approximately the same. Therefore, series operation was adopted.

17.3.3 Paste Tailings Disposal

Tailings from the rougher/scavenger circuit are pumped to two 12 ft diameter Deep Cone Thickeners (DCT). Diluted flocculant is added to the tailings as they are fed into the thickener feed well. Several flocculants were bench tested during the commissioning phase of the plant operation, and it was found that Z-Floc 565 produced the best settling characteristics with South vein mill feed. The dosage rate for the flocculant is 0.102 lb of dry flocculant per ton of mill feed. The flocculant is diluted to a 0.2% solution prior to injection into the tailings stream where it is again diluted with 4 gpm of process water. Overflow from the DCT thickeners is recycled back to the mill circuit, and no adverse impacts to the metallurgical performance of the mill were found. Underflow from the DCT thickeners is pumped to a paste tailings stack with the use of peristaltic hose pumps. Underflow density ranged from 60% to 70% solids and the tailings could be classified as paste in the high end of that range. The pressure in the paste tailings pipeline was continuously monitored at the hose pump discharge, and maximum pressures of about 100 psi were observed at high pulp densities.

Paste tailings were discharged from the pipeline at a slight elevation above the tailings impoundment floor. The paste flowed in anastomosing channels until it reached an approximately 6% angle of repose. No segregation of particle sizes was observed in the tailings stack. Mud cracks would form in the paste about two days after deposition. A small amount of bleed water was observed at the toe of the stack, and this clear water would either evaporate or infiltrate into the ground. This is the first application of paste tailings disposal in the Coeur d’Alene mining district.

17.4 Consumable Requirements Consumable requirements are shown in Table 17.4.1 below.

Table 17.4.1: Consumable Requirements

Reagent Description Addition Rate Location Makeup Water Water 102 gallons/t System Wide Lime Dry hydrated lime. 1.12 lb/t Ball Mill Feed Belt Z-55 Straight solution. 0.040 lb/t Rougher Cell No. 1 Sodium Cyanide 10% solution (mixed with water and lime). 0.015 lb/t Rougher Cell No. 1 Steel Grinding Balls 1.50 lb/t Ball Mill Feed


Kellogg, Idaho

Figure 17.2.1

Flotation Flowsheet Source: SRK, 2013


Kellogg, Idaho

Figure 17.3.1.1

Particle Size Distribution – Grinding Circuit Source: SRK, 2013



18 Project Infrastructure The Crescent is an old, well established mine that was operated between approximately 1910 and 1986. Following GFN’s purchase of the Project in late 2006, surface exploration and intermittent underground work resumed in 2007 and continued under USC’s direction from December 2009, when USC entered into an option agreement to acquire an 80% interest in the Project. There is currently a small staff in place which handles the care and maintenance of the mine. The existing infrastructure, mining equipment, and related support facilities are in good condition.

18.1 Infrastructure and Logistic Requirements

18.1.1 On-Site Infrastructure

The existing buildings are dedicated for offices, shops, and storage. They are in good condition and can be used for future mine operations. Power and water provided by public utilities are available on site.

18.1.2 Site Water Management

Industrial water supplies are obtained from the capture of naturally occurring ground water inside the mine. Disposal of excess mine water is discussed in Section 4.3 of this report.

18.1.3 Service Roads and Bridges

The Crescent mine site is accessible through a network of public roadways in the area, some of which currently support mining traffic. The main access is from I-90 using exit 54 for Big Creek Road. From the exit, it is approximately 2.1 miles south on paved road to the Crescent mine office.

The mine office and ancillary facilities are adjacent to the Hooper Tunnel, and located just above the valley bottom. Access from the office area to the Big Creek # 4 tunnel is along approximately 5,600 ft of two lane dirt road to reach a location about 500 ft above the Hooper Tunnel. This dirt road continues approximately 17,850 ft to the Countess decline located about 1,500 ft above the Hooper Tunnel. The company maintains the dirt road year around. The Project surface also contains a network of dirt roads that have been used for logging and are available to access diamond drilling platforms.

18.1.4 Mine Operations and Support Facilities

The Crescent Mine office was built in the 1950’s and was re-furbished in 2007 by GFN. The mine office and reception are found on the second story of the building above the shower and dry facilities. The existing office space is adequate for the projected staff.

There is a combination of three dry/shower rooms currently on the Crescent property. The Women’s dry is capable of accommodating 12 underground workers while the Men’s dry is split into two areas with a combined capacity for 70 underground workers.

The current warehouse is building is approximately 30 ft by 50 ft. It is equipped with electrical power, natural gas heating and outside water spigots. The surface mechanics shop is approximately 20 ft X 50 ft with an adjacent oil storage room, electrical parts warehouse and compressor room. The shop



is equipped with electrical power, natural gas heating, compressed air and firefighting equipment. Both buildings are adequate for needs of the mine.

Mine Fabrication and Machine (Mine Fab), a wholly owned subsidiary of USC, is a commercial fabrication, machining and repair facility that is equipped to provide repair services for all USC’s mobile and stationary equipment. The facility is located near Kellogg, Idaho.

The mine office and ancillary facilities have land line phone service provided by Frontier Communications. Cell phone service is not available at the mine office complex because it is located in a steep and narrow valley without access to a cell tower. The Countess Portal is serviced by a cell phone that connects to the land line at the mine office.

Numerous Femco mine phones are installed throughout the mine and in all refuge chambers. Underground personnel can contact other locations in the mine or on surface with the Femco phone system.

The current mine communications system will be replaced with a Leaky Feeder system when underground work resumes. The leaky feeder system is designed to provide portable communications through the mine. Mine supervisors and crews will be provided with hand held portable radios that will enable communications to any and all underground personnel, shops, refuge chambers and mine offices from any location within the mine. Complete implementation of the system will be achieved upon completion of the secondary egress.

There is currently no laboratory. The Company has developed plans to construct a new laboratory for its exclusive use. The laboratory will be sited at the New Jersey mill location and will analyze samples from both the mine and the mill.

The Company is also negotiating with American Analytical Services, Inc. of Osburn for assaying services with results delivered in 24 hours after sample delivery. That service would eliminate the need to construct a laboratory at the mill.

There is a 4,000 gallon (15,100 liter) off-road diesel storage tank located approximately 400 ft below the Crescent Mine parking lot and office buildings. This is a double-walled storage tank purchased new in 2011 and meets all State and Federal regulations for diesel storage. Diesel is transported via portable fuel tanks to equipment at the Big Creek #4 and Countess Portals as needed.

There is a 300 gallon (1,100 liter) on-road diesel storage tank and stand at the North end of the Crescent shop building. This storage tank area is equipped with properly sized spill containment in place.

The existing diesel storage is adequate for planned mine operations.

There are no bulk gasoline storage tanks on site. Gasoline is purchased in Kellogg as needed. Small quantities of gasoline are stored on site in MSHA approved metal fuel cans with pressure relief valves.

18.1.5 Process Support Facilities

All process support infrastructure is located off site, at the New Jersey Mill.



18.1.6 Additional Support Facilities

The disposal of the rock waste generated by the mining activities is discussed in Section 5.5.5. The solid waste disposal (garbage) and human waste are discussed below.

The Crescent Mine uses a local sanitation company, Water’s Garbage & Refuse (Waters) to provide roll-off dumpsters to the site. Waters collects the waste on a regular schedule and transports the waste to licensed solid waste disposal site.

There is a septic system in place adjacent to the parking lot of the mine office adequate for the more than 100 employees. The septic tank is pumped once per year.

The company owns Mine Fabrication and Machine, a complete fabrication, machining and shop servicing the local and regional mining industry in the western United States. Mine Fabrication is equipped to handle all repairs and fabrication that are beyond the capabilities of the mine.

18.1.7 Power Supply and Distribution

Electrical power and natural gas is purchased from Avista, the local utility company. Power is delivered to a 2,000 kW substation located near the mine office. The electric power distribution system includes all the surface building and feed for two 1,000 kVA power cables that carry electrical power into the mine via the Hooper Tunnel. Avista has assured the Company that there is sufficient electrical power available to meet Project demand.

Upon the resumption of underground work, surface electrical power distribution will be extended to the BC #4 portal. This will enable replacement of the portable diesel generator previously used to generate electricity in the BC#4 portal.

Upon completion of the secondary egress, electricity supplying the BC # 4 portal will be connected to the electrical distribution system installed from the Countess Portal. This will provide Avista supplied electrical power to the rest of the mine developed from the Countess Portal, enabling the diesel powered electrical generator at the Countess Portal to be removed from service.

18.1.8 Water Supply

Potable water is provided to the mine office and ancillary facilities by Central Shoshone Water District, a public utility.

Central Shoshone Water District provides potable water to the property at the South end of the office access road. There is sufficient pressure and flow to deliver water to main office pump room where pressure is increased and water is distributed to the shops, core shed, warehouse, geology building, hot water tanks, restrooms, sinks and spigots.

Emergency firefighting capability is provided by the Shoshone County fire protection district #2.

Industrial water supplies are obtained from the capture of naturally occurring ground water inside the mine.

18.1.9 Rail

The nearest rail siding, near Spokane, Washington, is about 65 miles from the Project.



18.1.10 Port

The nearest port, in Seattle, Washington, is about 350 miles from the Project. USC currently sells the silver concentrate to Ocean Freight. Ocean Freight is responsible for all freight from the mill to the smelter in Korea.

18.2 Tailings Management Area Tailings will be managed off-site, at the New Jersey Mill site.

18.3 Off-Site Infrastructure and Logistics Requirements No other off-site support facilities are required for the current project scope.



19 Market Studies and Contracts The process facility for this operation will produce a silver concentrate that will be sold to broker or to a commercial refiner. Sale price of concentrates is determined on world spot prices or London Metals Exchange market pricing. The Project will sell silver concentrate to Ocean Freight. Their charges include transportation to the port, sea freight, inland freight to a smelter in Korea and the smelting and refining charges.

19.1 Summary of Information A market study for the silver concentrate was not undertaken for this PEA.

19.1.1 Nature of Material Terms

United Silver currently sells its concentrate to Ocean Partners. The terms of the sales agreement are typical for the industry.

19.1.2 Commodity Price Projections

Figure 19.1.2.1 shows a graph of the LME silver price for January through late August, 2013. The price of silver has fallen from a high of US$32.23/oz on January 23, 2013 to a low of US$19.10/oz on July 9, 2013. Since then, the price has reached US$22.50. A silver price of US$20.00/oz is used as the basis of the Economic Analysis.

19.2 Contracts and Status The concentrate is currently sold to Ocean Partners Inc. Their charges include transportation to the port, sea freight, inland freight to a smelter in Korea and the smelting and refining charges.

United Silver plans to investigate alternative markets for the concentrates. The Sunshine refinery is modifying its process and maybe able to accept Crescent concentrates in the future. Should this transpire, Untied Silver would have the option of either taking position of the refined silver bullion and selling it on the open market or selling it directly to Sunshine. Marketing costs of bullion are assumed to be nil for this study.

United Silver has no forward sales agreements or hedge contracts.

19.2.1 Terms

The terms of the contract are summarized in Table 19.2.1.1.

Table 19.2.1.1: Smelting and Refining Costs

Item Amount Remarks Payable Silver 95% Treatment charge 750 US$ per dry metric tonne (dmt) Refining Cost 2.75 US$ per oz of Ag Received Penalty – As 3.50 US$ dmt per 0.10% over 0.10% Penalty – Sb 3.50 US$ dmt per 0.10% over 0.10%

Ocean Partners takes possession of the silver at the time United Silver is paid. This normally occurs when the concentrate leaves the mill. Ocean Partners sells the silver on the world market and any related costs are included in the smelting and refining charges.


Kellogg, Idaho

Figure 19.1.2.1

Silver Price, January – August 2013 Source: Kitco, 2013



20 Environmental Studies, Permitting and Social or Community Impact The Crescent Mine and New Jersey Mill are located in a designated Superfund site, registered on EPA’s National Priorities List (NPL) (IDD048340921). The first cleanup actions occurred in 1987 and then resumed in 1989, continuing to the present. Contaminants, including lead, arsenic, zinc, and cadmium, from historical mining in the Silver Valley area were found in Coeur d’Alene Lake and the Spokane River system to about 60 miles downstream. An estimated 62 Mt of mine wastes were dumped directly into the South Fork of the Coeur d’Alene River and its tributaries between 1884 and 1968. Cleanup actions are determined by the Records of Decision (ROD) issued by the EPA in a public involvement process. Currently, the RODs are divided into three cleanup areas called operable units, which include:

Operable Unit 1 = populated areas of the Bunker Hill Box is a-square-mile area was addressed first because of the high levels of contaminants in soil and in blood lead levels in humans;

Operable Unit 2 = non-populated, non-residential areas; Operable Unit 3 = areas of mining-related contamination in the South Fork and main Coeur

d’Alene River watersheds including Coeur d’Alene Lake and the Spokane River to Upriver Dam in Washington State.

The Crescent Mine and New Jersey Mill are located in Operable Unit 3 (OU3).

In 2010, the EPA issued a draft Focused Feasibility Study (FFS) for the Upper Basin of OU3. This Upper Basin FFS focused on developing remedies for identified areas of concern in the drainage of the South Fork Coeur d’Alene River stretching from Mullan to Kingston within the Silver Valley. The draft FS identified, by name and location, both active and historic mining sites or features, and allocated specific remedies to these sites/features. When the draft FFS was issued for public comment, it received widespread criticism for assigning expensive remedies to active mining sites that are currently operating under approved state and federal permits, as well as to many historical sites with little or no factual information on site conditions.

As a result of this overwhelming public response, the EPA has modified its approach to assessing mining sites for inclusion in the scope of remedial actions. Active mining sites operating under state and federal permits are now considered as contingency sites, and are not to be targets for investigation or clean up, so long as they remain in compliance with their permits (Hydrometrics, Inc., 2011).

The State of Idaho’s position is that unless the owner was an active participant in the release, they will be held harmless for remediation (personal communication with B. Schuld, 2010).

20.1 Required Permits and Status According to statements from USC, the mine and mill are fully permitted and all permits are in compliance. SRK has not independently verified the status of the permits. Because the mine is on patented mining claims (privately-owned land), only a limited number of environmental permits and authorizations are required for mining and milling operations. For the most part, these permits are all issued by the State of Idaho.



20.1.1 Summary of Operating and Environmental Permits

Table 20.1.1: Summary of Existing Permits

Permit CommentsMine Water Rights License – UG No Water Rights License is required because the mine is located on

Patented Mining Claims (Private Property) Water Rights License – Surface No Current Permit – These permits are obtained on an annual and as

needed basis. Water Discharge Permit Water is discharged under Sunshine Mine NPDES Permit ID#-000006-0.

Permitted under a 1973 Agreement with Sunshine Silver Mines Corp. Stormwater Discharge Permit #-IDR050000 (Permit Tracking # -IDR05CA75). Explosives Permit ATF # 9-ID-079-33-4K-00329 Reclamation Bond None Required Mill Cyanidation Permit Idaho Permit # CN-000027 This permit includes the reclamation plan Reclamation Bond No bond required Permit to Appropriate Water Ground water permit # 94-07509 Stormwater Discharge Permit # IDR05A383

20.2 Environmental Study Results A Modified Phase I Environmental Assessment Report for the site (LFR, 2007), noted five “Recognized Environmental Conditions,” including:

a. Location within a Superfund-Designated Area. b. Listing within the EPA “remedial investigation/feasibility study” for the presence of adit

drainage, upland waste rock, and surface disturbance with potential for erosion. c. Surface Water Discharge Contaminant Contributions and NPDES Permitting. d. Upland Soil and Waste Rock Contamination. e. Underground Contaminant Sources and Ground Water Impacts.

As suggested by Hydrometrics, Inc. (2011) and supported by SRK, the area-wide encumbrances of the EPA Superfund cleanup will not affect the ability of USC to mine and process mill feed from the Crescent Silver Project under the current permit authorizations, but may remain a longer-term risk of future liability should the EPA reverse course and again pursue existing operations to shoulder the burden of the cleanup efforts.

20.3 Environmental Issues The regulations applicable to the socio-environmental, archaeological and cultural aspects are composed of a set of rules and regulations that govern the environmental issues relevant to mining in Idaho. This set of regulations includes all national and sector rules which focus on the protection, preservation, and sustainable management of natural resources, the application of air, water, noise soil, flora, and fauna quality standards.

The Area of Direct Influence (ADI) is defined as the area where the Project can potentially directly affect the environment during construction and operation. These areas include:

ADIs, such as the mine, process plant, tailings facilities, and other ancillary facilities; Hydrological areas of influence of the Big Creek ravine; Meteorological areas of influence such as downwind areas that can be affected by dust; and



Socioeconomic areas of influence such as the communities of Kellogg, Osborn and Wallace.

As all of the areas are in compliance with regulatory requirements, direct influence is expected to be minimal.

20.4 Operating and Post Closure Requirements and Plans USC plans to operate the Crescent Mine as an underground operation, bringing mineralized material and waste rock to the surface, as required by operational needs. Waste rock will be stockpiled on the surface, in areas already designated to hold waste stockpiles. Mineralized material will be transported to a toll mill. The mill will be responsible for tailings disposal (including permitting).

20.5 Post-Performance or Reclamations Bonds The State of Idaho does not require Reclamation Bonds for underground mines.

20.6 Social and Community The Project is located in an extensive historic mining district, with nearby silver and base metal mines currently in operation. The community would benefit from up to 49 full-time employees working at the Project. Most of these positions could be filled from current residents of the Kellogg area, which has the resources to support the proposed workforce.

20.7 Mine Closure At mine closure, the mine will be closed in accordance with then current closure regulations.

20.8 Reclamation Measures During Operations and Project Closure Activities at the mine will have a very limited footprint during operations. Reclamation activities are not planned until closure. At closure, the waste rock dumps and portal areas will be recontoured and seeded. The mine opening will have bulkheads installed in them to seal the openings. A semi-passive water treatment system will be installed to handle drainage from the mine. Exact requirements of this system have not yet been fully defined.

20.9 Reclamation and Closure Cost Estimate Reclamation costs are shown in Table 20.9.1

Table 20.9.1: Closure Cost Estimates

Item Cost US$(000s)Dump Reclamation $200 Water Treatment System $110 Consultants $25 Bulkheads $70 Total $405Source: Hydrometrics, Inc. 2011b



21 Capital and Operating Costs

21.1 Capital Cost Estimates The Crescent Project is a partially developed underground property with extensive development and equipment available, currently on care and maintenance status. Mill feed will be processed on a contract basis at an operating mill. Mine offices, dry and ancillary facilities already exist, and can be used without additional development.

A summary of total estimated capital expenditures for the Project is presented in Table 21.1.1.

Table 21.1.1: Capital Cost Summary

Initial Capital Cost Item Cost US$ (000s)Mine Development $4,120 Mine Equipment $2,102 Underground Communication $111 Utilities/Electrical $132 Exploration $650 Contingency (25%) $1,779 Initial Capital Total $8,894Mine Development $2,668 Other $481 Laboratory $209 Rail Installation $157 Sustaining Capital $160 Mine Closure $405 Contingency (25%) $1,020 LoM Total Capital $13,993Source: SRK, 2013

The support for this cost estimate is provided in the sections below.

21.1.1 Preproduction Capital

Estimated costs for underground development before production begins are shown in Table 21.1.1.1. This capital cost is to extend both the Countess Decline and the Big Creek #4 to establish the second means of egress and allow production to begin. A number of attack ramps will be developed during this time. Additional excavation to create a refuge station, as required by law and broken rock rehandling stations, to improve efficiency, will be required

Table 21.1.1.1: Preproduction Mine Development Cost

Description Productivities Units Quantity Unit Rate Cost US$(000s)Countess Decline 17 ft/day ft 3,329 $552 $1,838Big Creek # 4 16 ft/day ft 4,048 $550 $2,226Attack Ramps 5.5 ft/day ft 188 $298 $56Total Preproduction Development $4,120

In order to meet development requirements additional mobile mining equipment will be required. In addition, a “leaky feeder” underground communications system will be installed. Utilities for electric power and compressed air will be run from the mine maintenance shop to the BC #4 to eliminate diesel generators and compressor. Exploration drilling is planned to confirm the mill feed zones.



Table 21.1.1.2: Other Preproduction Capital Requirements

Description Cost US$(000s)Mine Equipment $2,102Underground Communications $111Utilities/Electrical $132Exploration Cost $650Other Preproduction Capital $2,995

Table 21.1.1.3 is a breakdown of the mine equipment capital requirement.

Table 21.1.1.3: Cost of Additional Mining Equipment

Quantity Item Unit Total 2 Haul Truck – 18 t $496,000 $992,000 2 LHD – 4.0 yd3 $555,000 $1,110,000

Total $2,102,000

21.1.2 Post Development Capital

Once production begins, capital expenditures to develop the mill feed will continue. Additional attack ramps and access drifts for the Alhambra drift will be developed during the life of the mine. Immediately after the Countess Decline and the BC#4 meet, work will begin on the mill feed pass to allow mill feed to be cheaply moved by rail out of the BC#4 drift. A decline will also be driven from the BC#4 to the Hooper level to explore the South Vein. Table 21.1.2.1 details the capital expenditures for this development. Table 21.1.2.2 details other capital expenditures that will be required during the project life.

Table 21.1.2.1: Post Production Development

Description Productivities Unit Quantity Unit Rate Cost US$(000s) Attack Ramps 5.5 ft/day ft 1,002 $298 $299 Mill Feed Pass 6 ft/day ft 1,381 $267 $369 3250 Alhambra Drift 16.5 ft/day ft 491 $250 $123 3450 Alhambra Drift 16.5 ft/day ft 1,007 $250 $252 3650 Alhambra Drift 16.5 ft/day ft 1,002 $250 $251 3850 Alhambra Drift 16.5 ft/day ft 734 $250 $184 Hooper Decline 14 ft/day ft 4,000 $298 $1,192 Post Production Development Cost $2,668

Table 21.1.2.2: Other Post Development Capital

Description Cost US$ (000s) Laboratory $209 Rail for BC#4 $157 Sustaining Capital $160 Mine Closure $405 Other $481 Total Other Post Development Capital $1,412



Capital cost estimates were developed using both detailed cost estimates and vendor quotations. A 25% contingency for omissions has been added to all capital estimates in the economic analysis

21.2 Operating Cost Estimates

21.2.1 Basis for Mining Operating Cost Estimates

Mining equipment cost on a US$/day basis were developed from Infomine CostMine Underground Mining Equipment cost guide (Infomine, 2012). Operating Cost includes fuel and lubricants, tires, maintenance parts and labor, and wear items. Diesel price used was US$3.50 per gallon. A breakdown of equipment costs are shown in Table 21.2.1.1.

Table 21.2.1.1: Equipment Operating Rates

Equipment Operating Cost

US$/Day JACKLEGS $8 LHD – 2.0 CU YD MTI $155 U/G TRUCK 4.3 YD $78 ANFO LOADER $5 SCISSOR LIFT $32 PERS CARRIER –KUBOTA $30 DR JUMBO – 2 BOOM $54 50 HP FANS $15 400 kw GENERATOR $1693 COMPR DSL 750 CFM $429 CAT FEL 6 CU YD $233 FORK LIFT $32 PICK-UP $82

Labor rates for mining hourly labor are shown in Table 21.2.1.2.

Table 21.2.1.2: Hourly Labor Rates (US$)

PERSONNEL US$/Hr Fringes Base O.T. Burden Cost

Bonus/Hr /Day 0% 39.58% /M. Day SHIFT BOSS $ 44.00 $ - $ 440.00 $ - $ 174.15 $ 614.15 LEAD MINER (A) $ 30.00 $ 15.00 $ 450.00 $ - $ 178.11 $ 628.11 MINER (B) $ 26.00 $ 13.00 $ 390.00 $ - $ 154.36 $ 544.36 OPERATOR (Starter) $ 24.00 $ 12.00 $ 360.00 $ - $ 142.49 $ 502.49 TRUCK $ 24.00 $ 12.00 $ 360.00 $ - $ 142.49 $ 502.49 MECHANIC/ELECT $ 27.00 $ - $ 270.00 $ - $ 106.87 $ 376.87 SURFACE $ 20.00 $ - $ 200.00 $ - $ 79.16 $ 279.16

Mining unit costs were developed based on daily productivity estimates and assumed working two-10 hour shifts per day, seven days per week. Cost included labor and equipment, materials (primarily roof support and utilities installation) and supplies, including blasting drill steel and bits, vent tubing and hoses and fittings. Mining G&A includes an allowance for miscellaneous equipment and supplies that cannot be attributed to direct mining cost. Underground mill feed haulage includes rail haulage of mill feed from the mill feed pass to the surface. Table 21.2.1.3 outlines the productivity estimates and the calculated unit rates for the various cost components of mill feed production. Alhambra and South Vein production include the costs for drilling and blasting waste in the stopes and the cemented fill that is used for backfill.



Table 21.2.1.3: Mine Operating Productivities and Unit Cost

Description Productivity Unit Rate A Vein Development 7.5 ft/day $252 US$/ft Raise Development 39 t/d $45,000 US $/ea Alhambra Mill Feed Production 52 t/d $54 US $/t I drift 43 t/d $217 US $/ft Attack Ramp Wedge Allowance $3.50 US $/t South Vein Mill Feed Production 50 t/d $ 61 US $/t Underground Mill Feed Transport Allowance $2.00 US $/t Mine G&A Allowance $500,000 US $/yr

21.2.2 Mill Feed Transportation and Processing

Once mill feed is hauled and stockpiled on the surface, it will be hauled to the New Jersey Mill. Initial mill feed will be hauled from the Countess Decline in highway trucks. The historic cost of hauling from this location is US$7.00/t. Once the mill feed raise has been established, all mill feed will be hauled from the Big Creek #4. This haul will have a much shorter cycle time and the cost of mill feed haulage was factored down to US$3.00/t.

The mill feed processing cost of US$23.50/t was estimated by the New Jersey Mill. Table 21.2.2.1 summarizes the mill feed transportation and processing cost.

Table 21.2.2.1: Surface Mill Feed Transportation and Process Cost

Description Unit RateMill Feed Transportation – Initial $7.00 US $/t Mill Feed Transportation $3.00 US $/t Milling Cost $23.50 US $/t

21.2.3 General and Administrative Cost

General and Administrative costs are estimated US$2,200,000 per year, after the mine is in full production. Table 21.2.3.1 shows the G&A labor rates.

Table 21.2.3.1: G&A Labor Rates

Cost – US$ Total Qty Position Ea – Month Total – Month Burden Total Year

1 Mine Manager 10,000 10,000 3,889 13,889 166,668 1 Secretary 2,640 2,640 1,027 3,667 44,000 1 Safety 7,500 7,500 2,917 10,417 125,001 1 Purchasing 7,500 7,500 2,917 10,417 125,001 1 Engineer 9,000 9,000 3,500 12,500 150,001 1 Geologist – Chief 9,000 9,000 3,500 12,500 150,001 4 Geologist – Jr 5,000 20,000 7,778 27,778 333,336 1 CAD Technician 5,000 5,000 1,945 6,945 83,334 1 Survey 4,600 4,600 1,789 6,389 76,667 1 Janitor (Half Time) 1,760 1,760 684 2,444 29,334

13 Total 29,945 106,945 1,283,344



Total G&A costs at full production are shown in Table 21.2.3.2.G&A costs were ramped up during the first two years of the operation.

Table 21.2.3.2: G&A operating Cost Summary

Item US$ Labor and Fringes 1,283,344 Laboratory 407,502 Power 310,000 Pickups 57,600 Insurance 50,000 Communication 25,000 Small Vehicles 20,000 Road and Dump Maintenance 13,000 Natural Gas 12,000 Safety Equipment 12,000 Office Supplies 7,200 Total – Year $2,197,646



22 Economic Analysis The financial results of this report are based upon work performed by SRK and have been prepared on an annual basis. All costs are in 2013 US constant dollars.

22.1 Principal Assumptions and Input Parameters A financial model was prepared on an after-tax basis; the results are presented in this section. Key criteria used in the analysis are discussed in detail throughout this report. Financial assumptions used are summarized in Table 22.1.1 and 22.1.2.

Table 22.1.1: Market Parameters

Description Value Unit Market Price: Silver $20.00 US$/oz Smelter Payment 95.0% Refining $2.75 US$/oz Smelting Charge $750.00 US$/t-concTransportation $0.00 Included Insurance 0.05% Assumed Royalty 2.00% Of NSR Source: SRK, 2013

Table 22.1.2: Production Parameters

Description Value Unit Mine Life 6 Years RoM Mined (undiluted) 451 kt RoM Processed (diluted) 601 kt Recovered Ag 6,108 koz Payable Ag 5,803 koz

22.2 Cashflow Forecasts and Annual Production Forecasts The following contain the production and cost information developed for the Project. Table 22.2.1 is a summary of the annual mine production and cashflow over a 6-year mine life.

Table 22.2.1: Annual Mine Production and Cashflow Summary

Year Undiluted

RoM kt Diluted

RoM to Plant Recovered

Ag koz Free

Cashflow Discounted

Cashflow

2013 9 12 127 (10,264) (10,264)

2014 81 108 1,096 987 914 2015 108 144 1,463 6,049 5,186 2016 108 144 1,463 6,409 5,087 2017 108 144 1,463 7,164 5,266 2018 36 48 495 3,313 2,255 2019 0 0 0 206 130 Total 451 601 6,108 13,865 8,575

The economic analysis results shown in Table 22.2.2 indicate a NPV@8% of US$8.6 million (after-tax basis) and an IRR of 32%. The following provides the basis of the SRK LoM plan and economics:



A mine life of 6 years; An overall average process recovery rate of 92%; A tax loss carry forward of $2.67 million was included: A site operation cash cost of US$146.43/t processed; and Post Development Capital of US$5.1million will be required.

Project economic results and estimated cash costs are summarized in Table 22.2.2.

Table 22.2.2: Economic Results

Description Value units Production Summary

RoM Mined (undiluted) 451 kt RoM Processed (diluted) 601 kt

Estimate of Cash Flow

Gross Income $122,165

US$/t-process Refining ($26,775) $44.58

Gross Revenue $95,390 Royalty ($1,157) 1,157

Net Revenue $93,722 Operating Costs

Mining $33,084 $55.08 Processing $16,074 $26.76

G&A $10,864 $18.09 Total Operating $60,022 $99.93

Site Operation Cash Cost $87,954 $146.43 Operating Margin $34,211

Initial Capital $8,894 Post Development Capital $5,099

Income Tax $0 Cash Flow Available for Debt Service $13,865

NPV 8% $8,575 IRR 32%

22.3 Taxes, Royalties and Other Interests Taxes and other interests have been calculated for the Project.

Royalty applied at 2.0% Net Smelter Return; Mining Tax applied at a 1% rate of net cash flow; State Income Tax applied at a 7.6% rate; and Federal Income Tax applied at a 35% rate.

Depletion applied at a 15% rate of adjusted Gross Income



22.4 Sensitivity Analysis Sensitivity analysis for key economic parameters is shown in Table 22.4.1 and Figure 22.4.1 below. The Project is most sensitive to market prices (revenues). The Project’s sensitivities to capital and operating costs are similar but more susceptible to operating costs.

Table 22.4.1: Sensitivity Analysis of NPV @ 8% (US$000)

Description -20% -10% Base 10% 20%Silver Price 16.00 18.00 20.00 22.00 24.00 Revenues (9,859) (633) 8,575 17,188 23,324 Capital Costs 11,222 9,898 8,575 7,251 5,927 Operating Costs 17,092 13,484 8,575 3,665 1,245


Kellogg, Idaho

Figure 22.4.1

Sensitivities Source: SRK, 2013



23 Adjacent Properties The Project property is situated at the west end of the Silver Belt located between the Sunshine Mine approximately 5,000 ft to the east and the Bunker Hill Mine approximately 10,000 ft to the west. The Sunshine Mine has a historic production of about 365Moz of silver. The Bunker Hill Mine has produced about 192Moz of silver in addition to 3.6Mt of lead. Neither mine is currently producing.

The ground adjacent to the property on the west is controlled by The New Bunker Hill Mining Company. On the north, there are BLM staked claims and State of Idaho Leases that are controlled by the Lowell Miller Estate. On the east and south are claims controlled by Sunshine Silver Mines Corporation.



24 Other Relevant Data and Information The lower Project underground workings are partially flooded at present. This water level is controlled by the water maintained in the Bunker Hill Mine by the interconnected underground workings. At present, The New Bunker Hill pumps to keep the water level at about the 2,000 ft elevation, which is about 700 ft below the Hooper Tunnel in the Project. If The New Bunker Hill discontinues pumping, the water level will rise to no more than the daylight elevation of the Kellogg Tunnel at 2,350 ft elevation, which corresponds to 350 ft below the Hooper Tunnel in the Project. All the resources reported in this report are located above the Hooper Tunnel elevation, and therefore would not be flooded if pumping stopped.

Zones of low-grade uranium mineralization are present within the veins of the Project. Drill samples contain uranium as high as 245 ppm over 2 ft. Nine samples out of about 1,200 analyses returned 80 to 245 ppm U; the remainder returned no greater than 40 ppm U. Elevated radon readings were noted by previous operators in a crosscut near the Ellis Shaft on the Hooper Level. Proper mine ventilation should alleviate the buildup of any radon gasses.



25 Interpretation and Conclusions SRK finds no material impediments to the development of the Project. The State of Idaho has permitting and closure regulations that are amenable to mining, and the Silver Valley, where the Project is located, has over 100 years of profitable mining production history.

At PEA level, inferred resources have been used to evaluate potential mineral economics. Reserves have not yet been stated for the Project. SRK has supported USC in preparing the necessary drilling and sampling programs designed to upgrade Inferred resources. There is a strong potential to expand existing resources with additional drilling, especially at structural intersections where the newly identified Jackson Vein intersects the Alhambra and the South Vein.

From an economic perspective, the Project benefits from previous test mining in 2011/12, which provided essential information about the nature of mineralization, mining methods, processing and costs. These data underpin the economic evaluation of this PEA. Table 25.1 provides a summary of the risks and opportunities associated with Project at the current level of understanding.

Table 25.1: Relevant Risks and Opportunities

Project Element Economic Risk Level CommentRESOURCES Database Exploration Data Sufficiency/Adequacy Low Silver Valley mineralization is historically very

continuous. Additional drilling is recommended to confirm untested areas of the South Vein.

Assaying Low Recent drilling programs have had modern QA/QC and support historic results.

Surveying Moderate Collar surveys are potentially inaccurate due to survey methods. Down hole surveys need confirmation using alternative methods such as gyroscope for validation.

Geology Low Geology is sufficiently understood to direct drilling and future resource expansion.

Geology and Resource Modeling Geological modeling Moderate Absolute location of veins could be affected

by potentially inaccurate down hole surveys of deep core holes. This was largely mitigated by the location of the veins in the 2011/12 test mining campaign.

Resource modeling approach Low Geostatistical analysis Low Variography was not applied to the estimate. Resource estimate Low Resource risk is considered low but requires

validation and upgrade of some areas from additional drilling.

GEOTECHNICAL Slope Stability Geotechnical data adequacy Low Underground ground support requirements

and stability are well understood as a result of test mining.

Design Low No complications are envisioned based on existing drift performance.

Waste Rock Dump Geotechnical data adequacy Low Permits for placement High While waste rock volumes will be low, the

surface rights for waste placement have not yet been secured.

Design Low There is ample space for waste-rock placement, once surface rights are acquired.



Project Element Economic Risk Level Comment Geochemistry Low Idaho environmental requirements are not

stringent. MINEABLE RESERVES AND MINING Conversion of resources to reserves Not yet applicable Resources being evaluated in this PEA have

strong potential to be converted to reserves with additional drilling.

Proposed production schedule Moderate Key to the success of the Project will be the ability of the Company to extract tons of mill feed and waste on the proposed schedule. Mining methods planned , though innovative, are untested.

Dilution High Previous test mining demonstrated that equipment selection and mining methods were crucial to control dilution, which ultimately determines profitability.

Equipment fleet Low (Opportunity) USC’s proposed list of equipment is considered reasonable for this level of study. There may be opportunities to acquire used equipment at reduced cost or lease/rent equipment to reduce Capex.

Mining unit cost assumptions Low Mining costs for this study are very robust and informed from previous test mining.

Labor Low Labor rates for this study are very robust and informed from previous test mining.

METALLURGICAL TEST WORK/PROCESSING FACILITIES Metallurgical Test Work Low (opportunity) Existing test work and bulk testing indicates

good recovery. Metallurgical test work has not been completed on the oxide (upper) portion of the mill feed. Mineable material would increase if oxide mill feed recovery was adequate.

Mill feed type definition Moderate Recovery projections Moderate 92% recovery projected should be attainable

in the sulfide portion of the mill feed. Throughput Low Mill design is adequate for planned

production Process unit assumptions and reasonableness of rates

Moderate Mill has not been run to maximum capacity for a sustained period.

Concentrate quality Moderate Concentrate quality from mill runs was lower than predicted from test work

Offtake agreements Low (opportunity) Offtake contract in place. If concentrate quality can be improved, better terms could be negotiated or sales to others arranged to improve revenue.

ENVIRONMENTAL AND PERMITTING Status of statutory permits for current and future operations

Low to Moderate Mine is on patented claims, reducing permit requirements. All permits to operate are in place. Mine is located within a Superfund Site

Compliance of current operations with existing permits

Low The operation is currently in compliance.

Risks for future compliance of operations with permits

Unknown

Identification of environmental and social risks

Low Mine is in a historic mining district with local support. The mine is not a target for inclusion in Superfund cleanup cost.

Mine reclamation and closure plans and costs

Moderate Closure costs are minimal. Building removal may be required. Post closure discharge requirements need to be defined.



Project Element Economic Risk Level CommentINFRASTRUCTURE Power Low In place and adequate for present plan. Water Supply Low In place and adequate for present plan Access Low In place and adequate for present plan Transportation Low In place and adequate for present plan Surface facilities Low In place and adequate for present plan CAPITAL COSTS Capital cost programs Moderate Underground development costs are

estimated. Equipment capital based on vender quotes.

Sustaining capital Low For replacement vehicles only. OPERATING COSTS Forecast costs used in mineable resource determination

Moderate Costs are estimated, Ultimate mine profitability will be based on limiting dilution.

Currency split of domestic to foreign currency

Low All costs and revenue are in US dollars.

FINANCIAL MODEL Model verification Moderate Production schedule includes inferred

resources. Revenue calculations Moderate Revenue based on inclusion of inferred

resources. Tax Low Opportunity USC economics benefit from a tax loss carry

forward. MANAGEMENT AND STAFFING Low Mine is located in a mining district that has a

good source of trained miners. Management is experienced.



26 Recommendations

26.1 Recommended Work Programs and Costs Additional work recommended to advance the Crescent Project consists of the following key elements:

Confirmation drilling on the South Vein; Metallurgical testing of mineralization in oxide; Underground grade control laboratory; and Test mining to confirm mining methods and economics.

The total cost estimate for additional work is US$1,635,000 and is expected to be completed by the end of the third quarter of 2014 assuming test mining can be financed and initiated by the end of 2013. At that point the mine would go into full production. The work program is summarized in Table 26.1.

Table 26.1: Recommended Pre-Production Work Program Costs

Work Element Estimated Cost US$ Assumptions/Comments Underground Confirmation Core Drilling 650,000 20 holes to 540 ft @ US$60/ft Metallurgical Testwork of Oxide 100,000 Underground Grade Control Lab 60,000 Test mining (Year 1) 825,000 Does not include cost recovery from productionHydrogeochemical Studies 100,000 Total 1,735,000

Confirmation Drilling

The Mineral Resource is well defined by drilling at and above the Countess Decline elevation but has fewer intercepts at depth. SRK recommends a drilling program in the area of the South Vein between the Countess Decline and Hooper levels with the objective to confirm the location and thickness of the mineralization and to improve the classification of that material from Inferred to Indicated. The proposed underground drilling program consists of 20 core holes, 16 of which to be drilled downward from the Countess Decline and four of which to be drilled upward from the Hooper level. The total footage anticipated for the program is 10,830 ft at an all-in cost of US$60/ft for a total of US$650,000. At typical district drilling rates of 150 ft/day, the program will be completed in approximately 10 weeks. Positive results from the drilling program will validate the mine plan to connect the Countess Decline to Big Creek #4 and confirm future mining opportunities down to the Hooper level.

All of the exploration drill holes should get gyroscopic directional surveys to maximize accuracy of their down-hole trajectories, and to get the best available location data for the mineralized intercepts.

Metallurgical Testing

Metallurgical testing is needed to investigate improving recovery in the oxide portion of the South Vein, most of which has been left out of the mine plan. Testing will also be re required to determine how to make a cleaner concentrate. A higher grade concentrate will directly affect the overall economics of the Project because more refiners will be capable of handling the concentrate and the overall cost of refining the concentrate will be reduced. The estimated cost of metallurgical testing of



the oxide is US$100,000. Testing materials will be collected during test mining. The test work is expected to be completed by the end of the second quarter of 2014.

Underground Grade Control Laboratory

SRK recommends that USC continue to investigate using one of the planned underground facilities, possibly one of the emergency fire escape chambers as a laboratory for underground grade control. Mining will be dependent on rapid turnaround of analytical results from the advancing headings. There is an opportunity, using available portable X-ray scanning technology, to make an order-of-magnitude grade assessment within minutes to hours of sampling. The underground grade control laboratory would facilitate reduction of face samples to a pressed powder suitable for scanning. This first-pass grade control would be supported by conventional assays using certified local analytical laboratories. The cost of the equipment for the underground grade control laboratory is estimated at US$60,000. This includes the scanning device, crushing and pulverizing equipment and ventilation.

Test Mining

A limited-scale test mining program to prove the stope mining plan is recommended during initial development. MSHA will allow limited production while establishing a second means of egress. The main emphasis during development must be to establish egress and must be management’s primary goal. The test mining can be accomplished from existing I-drifts in the South Vein and from the 4100 elevation in the Alhambra Vein. The test mining must be carefully monitored to establish basic metrics for production scheduling and costing, and to determine if the current cost estimates can be met. The estimated costs for this test-phase of mining total US$825,000, which represents capital and operating costs not offset by production revenue. The test mining phase of development has a 34 week completion schedule.

Hydrogeochemical Studies

Additional geochemical and hydrological studies need to be performed to accurately define the post closure water quality and volume expected to drain from the mine that may require treatment prior to NPDES discharge.

Metallurgical Testing

Metallurgical testing is needed to investigate improving recovery in the oxide portion of the South Vein, most of which has been left out of the mine plan. Testing will also be re required to determine how to make a cleaner concentrate. A higher grade concentrate will directly affect the overall economics of the Project. More refiners will be capable of handling the concentrate and the overall cost of refining the concentrate will be reduced.



27 References ALS Chemex 2009. Geochemical technical information webpages and downloads.

http://www.alsglobal.com/Mineral/ALSContent.aspx?key=29#ICPAES Accessed ca. April 2009;

http://www.alsglobal.com/en/Our-Services/Minerals/Geochemistry/Capabilities/~/link.aspx?_id=9E011F041FF84D6FAC0FE84D203E5F42&_z=z, ME-ICP41 method .pdf, Accessed 10 September 2013.

Anderson, I. 1922. Letter to F.S. Norcross. Homestead, Oregon. 15 March 1922.

Baldry, J. 1981. Report on the Bunker Hill Main Mine: Crescent Mine and Concentrator. Bunker Hill company report. 18 December 1981.

Blake, W. 2013. Personal communications with M. Gross, T. Byberg, R. M. Robb, C. Durick. March, April, and May 2013, various dates.

Bunker Hill Company. 1982. Revised report by Radford, N.A. 1973.; “Geology of the Crescent Mine and its Relationships to the Coeur d’Alene Silver Belt. Osburn , Idaho. April 1973.” Revised by: Meyer, R.L., 1981 and Duff, J.K., 1982. The Bunker Hill Company. Kellogg, Idaho.

Bunker Hill Company. 1980. Crescent Mine Operating Highlights. December 1980.

Childs, H. M. 1928 and 1929. 27 Memoranda to Stanley Easton, Crescent Mine: Mining progress from Jan. 6, 1928 to Feb. 22, 1929.

Colson, J.B. 1958. The Geology of the Sunshine Mine and Adjacent Areas, paper read at A.I.M.E. Northwest Regional Conference, 18 April 1958. 10p.

Dalton, D.J. 1986. Memorandum to Jack Swanson, Bunker Hill Mining Company. November 15, 1986, 1p.

Derkey, P.D., Johnson, B.R., Carver, M. 1996. Digital Geologic Map of the Coeur d’Alene District, Idaho and Montana. USGS Open File Report 96-299.

Farmin, J.C. 1975. Ore Control Research: 1973 through April, 1975. Bunker Hill Mining Company research report. April 1975.

Fleck, R.J., Criss, R.E., Eaton, G.F., Cleland, R.W., Wavra, C.S., Bond, W.D. 2002. Age and Origin of Base and Precious Metal Veins of the Coeur d’Alene Mining District, Idaho. Economic Geology and the Bulletin of the Society of Economic Geologists. Vol. 97, no.1 pp.23-42.

Fryklund, V.C. and Weis, P.L. 1964. Ore Deposits of the Coeur d’Alene District, Shoshone County, Idaho: U.S. Geological Survey Professional Paper 445, 103 p.

G&T Metallurgical Services, Ltd. 2011. Metallurgical Testing of the Cresceent South Vein. July 8, 2011.

Hershey, O.H. 1916. Origin and distribution of ore in the Coeur d’Alene. Mining and Scientific Press, San Francisco, 32 p.

Hershey, O.H. 1922. Report on “Anderson” Mine: Report for F. W. Bradley, Bunker Hill & Sullivan M. & C. Co. Internal Report. 23 March 1922.



Hobbs, S.W., Griggs, A.B., Wallace, R.E., and Campbell, A.B. 1965. Geology of the Coeur d’Alene District, Shoshone County, Idaho: U.S. Geological Survey Professional Paper 478, 139 p.

Hydrometrics, Inc. 2011. Mining in the Silver Valley of North Idaho in Today’s Environmental Climate. July 12, 2011.

Hydrometrics, Inc 2011b Budgetary Cost Estimates for Specific EPA Remedies, July 12, 2011

Infomine. 2013. Web page: Mine & Mill Equipment Costs Estimator's Guide, Capital & Operating Costs, http://costs.infomine.com/costdatacenter/miningequipmentcosts.aspx. Accessed August 2013.

Julihn, C.E. and Horton, F.W. 1936. The Silver Belt and the Sunshine Mine of the Coeur d’Alene District: U.S. Bureau of Mines IC-6876, 16p.

Juras, D.W. 1977. Structural Evaluation of the Crescent Mine. Bunker Hill Company report. 17 March 1977.

Kitco. 2013. Web page: London Fix Historical Silver Price for 2013, http://www.kitco.com/scripts/hist_charts/yearly_graphs.plx, Accessed 19 August 2013.

Lewis, R.S., Burmester, R.F., Breckenridge, R.M., McFadden, M.D., Kauffman, J.D. 2002. Geologic map of the Coeur d’Alene 30 x 60 minute quadrangle, Idaho: Idaho Geological Survey Geological Map GM-33, scale 1:100,000.

LFR Inc. 2007. Modified Phase I Environmental Site Assessment Report, Crescent Mine Property Located in the Big Creek Drainage, Shoshone County, Idaho. 9 February 2007.

Radford, N.A. 2009. Personal communication with USC.

Radford, N.A. 1985. Geology of the Upper Crescent Mine, Kellogg, Idaho, Shoshone County, Idaho. Osburn, Idaho. August 1985.

Rust, W.C. 2011. Crescent South Vein Mill Test, June 11, 2011.

Rust, W.C. 2012. New Jersey Mill grinding circuit survey with South Vein material. August 21, 2012.

Schuld, Bruce. December 16, 2010. Personal communications with USC and Val Sawyer, SRK.

Schuld, Bruce. June 2013. Personal communications with Lisa Hardy, USC.

SRK Consulting. 2010. Updated NI 43-101 Technical Report on Resources, United Mine Services, Inc., Crescent Mine, Kellogg, Idaho. 21 May 2010. 81p.

Springer. 1983. Independent Reserve Estimation for Bunker Hill Mine.

Stoel Rives LLP. 2007. Mineral Rights Title Opinion, Stoel Rives LLP 900 SW Fifth Ave, Suite 2600, Portland Oregon. Letter dated 14 May 2007. 18p.

Strand, A.D. 2002. Hydrothermal Alteration in the Coeur d’Alene mining district, Idaho, USA. Master’s thesis, University of Auckland; Auckland, New Zealand. 86 pp.

USC. 2011. Press Release entitled “United Mining Group (TSX: UMG) and Gold Finder (GFN.V) Announce Accelerated Earn-In on Crescent Silver Mine Project”. 01 June, 2011.

USC. 2012. Press Release entitled “United Silver Corp. Acquires Gold Finder Interest in Crescent Mine Project”. 29 May 2012.



USC. 2013. Press Release entitled “United Silver Corp. Releases Positive Preliminary Economic Assessment and Reports Updated Estimated Resources”. 29 July 2013.

Wavra, C. S., Bond, W.D. and Reid, R.R. 1994. Evidence from the Sunshine Mine for Dip Slip Movement during Coeur d’Alene district Mineralization: Economic Geology v.89, p 515-527.

Western Regional Climate Center (WRCC). 2009. Period of Record Monthly Climate Summary. Kellogg, Idaho. Period of Record: 02/01/1905 to 12/31/2005. http://www.wrcc.dri.edu/cgi-bin/cliMAIN.pl?idkell. Accessed ca. April 2009 and 10 September 2013.

White, B.G. 1998a. New Tricks for an Old Elephant: Revising Concepts of Coeur d’Alene Geology. Mining Engineering. Vol. 50, no. 8, pp. 27-35. August.

White, B.G. 1998b. Diverse Tectonism in the Coeur d’Alene Mining District, in Berg, R.B. ed. Belt Symposium III, Montana Bureau of Mines and Geology Special Publications 112, p 254-265.

White, B.G. and Winston, D. 1977. The Revett-St. Regis “Transition Zone” Near the Bunker Hill Mine, Coeur d’Alene District, Idaho. Society of Economic Geologists, Coeur d’Alene Field Conference, Idaho – 1977, in Reid, R.R. and Williams, G. A. Idaho Bureau of Mines and Geology. Idaho Department of Lands. Bulletin 24. May 1982.

White, B.G. 1977. Stratigraphic Ore Control in the Crescent Mine. Bunker Hill Company report. March 15.



28 Glossary

28.1 Mineral Resources The mineral resources and mineral reserves have been classified according to the “CIM Standards on Mineral Resources and Reserves: Definitions and Guidelines” (November 27, 2010). Accordingly, the Resources have been classified as Measured, Indicated or Inferred, the Reserves have been classified as Proven, and Probable based on the Measured and Indicated Resources as defined below.

A Mineral Resource is a concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.

An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.

An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough to confirm both geological and grade continuity.

28.2 Mineral Reserves A Mineral Reserve is the economically mineable part of a Measured or Indicated Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate information on mining, processing, metallurgical, economic and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified. A Mineral Reserve includes diluting materials and allowances for losses that may occur when the material is mined.

A ‘Probable Mineral Reserve’ is the economically mineable part of an Indicated, and in some circumstances a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility



Study. This Study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified.

A ‘Proven Mineral Reserve’ is the economically mineable part of a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction is justified.

28.3 Definition of Terms The following general mining terms may be used in this report.

Table 28.3.1: Definition of Terms

Term DefinitionAssay The chemical analysis of mineral samples to determine the metal content. Capital Expenditure All other expenditures not classified as operating costs. 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 mill feed.

Crushing Initial process of reducing mill feed particle size to render it more amenable for further processing.

Cut-off Grade 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 mill feed. 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 mill feed area or stope. Gangue Non-valuable components of the mill feed. Grade The measure of concentration of gold within mineralized rock. Hangingwall The overlying side of a mill feed area or slope. Haulage A horizontal underground excavation which is used to transport mined mill feed. Hydrocyclone A process whereby material is graded according to size by exploiting centrifugal

forces of particulate materials. Igneous Primary crystalline rock formed by the solidification of magma. Kriging An interpolation method of assigning values from samples to blocks that minimizes

the estimation error. Level Horizontal tunnel the primary purpose is the transportation of personnel and

materials. Lithological Geological description pertaining to different rock types. LoM Plans Life-of-Mine plans. LRP Long Range Plan. Material Properties Mine properties. Milling A general term used to describe the process in which the mill feed is crushed and

ground and subjected to physical or chemical treatment to extract the valuable metals to 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. Ongoing Capital Capital estimates of a routine nature, which is necessary for sustaining operations. Pillar Rock left behind to help support the excavations in an underground mine. Sedimentary Pertaining to rocks formed by the accumulation of sediments, formed by the erosion

of other rocks. Shaft An opening cut downwards from the surface for transporting personnel, equipment,

supplies, mill feed and waste. Sill A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by the

injection of magma into planar zones of weakness.



Term DefinitionSmelting A high temperature pyrometallurgical operation conducted in a furnace, in which the

valuable metal is collected to a molten matte or doré phase and separated from the gangue components that accumulate in a less dense molten slag phase.

Stope Underground void created by mining. 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 statistical representation of the characteristics (usually grade).

28.4 Abbreviations The following abbreviations may be used in this report.

Table 28.4.1: Abbreviations

Abbreviation Unit or Term A ampere AA atomic absorption A/m2 amperes per square meter ANFO ammonium nitrate fuel oil Ag silver Au gold AuEq gold equivalent grade °C degrees Centigrade CCD counter-current decantation CIL carbon-in-leach CoG cut-off grade cm centimeter cm2 square centimeter cm3 cubic centimeter cfm cubic feet per minute ConfC confidence code CRec core recovery CSS closed-side setting CTW calculated true width ° degree (degrees) dia. diameter dmt dry metric tonnes EIS Environmental Impact Statement EMP Environmental Management Plan FA fire assay ft foot (feet) ft2 square foot (feet) ft3 cubic foot (feet) g gram gal gallon g/L gram per liter g-mol gram-mole gpm gallons per minute g/t grams per ton ha hectares HDPE Height Density Polyethylene hp horsepower HTW horizontal true width



Abbreviation Unit or Term ICP induced couple plasma ID2 inverse-distance squared ID3 inverse-distance cubed IFC International Finance Corporation ILS Intermediate Leach Solution kA kiloamperes kg kilograms km kilometer km2 square kilometer koz thousand troy ounce kt thousand tons kt/d thousand tons per day kt/y thousand tons per year kV kilovolt kW kilowatt kWh kilowatt-hour kWh/t kilowatt-hour per metric ton L liter L/sec liters per second L/sec/m liters per second per meter lb pound LHD Long-Haul Dump truck LLDDP Linear Low Density Polyethylene Plastic LOI Loss On Ignition LoM Life-of-Mine m meter m2 square meter m3 cubic meter masl meters above sea level mg/L milligrams/liter mm millimeter mm2 square millimeter mm3 cubic millimeter MME Mine & Mill Engineering Moz million troy ounces Mt million tons MTW measured true width MW million watts m.y. million years NGO non-governmental organization NI 43-101 Canadian National Instrument 43-101 OSC Ontario Securities Commission oz troy ounce % percent PLC Programmable Logic Controller PLS Pregnant Leach Solution PMF probable maximum flood ppb parts per billion ppm parts per million QA/QC Quality Assurance/Quality Control RC rotary circulation drilling RoM Run-of-Mine RQD Rock Quality Description SEC U.S. Securities & Exchange Commission sec second SG specific gravity SPT standard penetration testing t U.S. ton (2,000 pounds) t/h tons per hour t/d tons per day t/y tons per year



Abbreviation Unit or Term TSF tailings storage facility TSP total suspended particulates µm micron or microns, micrometer or micrometers V volts VFD variable frequency drive W watt XRD x-ray diffraction y year

SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Preliminary Economic Assessment – Crescent Silver Project Appendices


Appendices

SRK Consulting (U.S.), Inc. NI 43-101 Technical Report – Preliminary Economic Assessment – Crescent Silver Project Appendices


Appendix A: Certificates of Qualified Persons

SRK Consulting (US) Inc. 5250 Neil Road, Suite 300 Reno, Nevada 89502 T: (775) 828-6800 F: (775) 828-6820 [email protected] www.srk.com

CERTIFICATE OF AUTHOR

I, J. B. Pennington, CPG, do hereby certify that:

1. I am Principal Mining Geologist of SRK Consulting (U.S.), Inc., 5250 Neil Road, Suite 300, Reno, Nevada 89502.

2. This certificate applies to the technical report titled “NI 43-101 Technical Report, Preliminary Economic Assessment, Crescent Silver Project, Kellogg, Idaho, USA” with an Effective Date of July 22, 2013 (the “Technical Report”).

3. I graduated with a Bachelor of Science Degree in Geology from Tulane University, New Orleans, La., USA; May 1985; and a Master of Science Degree in Geology from Tulane University, New Orleans, La., USA; May 1987. I am a Certified Professional Geologist through membership in the American Institute of Professional Geologists, C.P.G. #11245. I have been employed as a geologist in the mining and mineral exploration business, continuously, for the past 26 years, since my undergraduate graduation from university. My relevant experience for the purpose of the Technical Report is: • Project Geologist, Archaen gold exploration with Freeport-McMoRan Australia Ltd. Perth Australia,

1987-1989; • Exploration Geologist, polymetallic regional exploration, Freeport-McMoRan Inc; Papua, Indonesia,

1990-1994; • Chief Mine Geologist, mine geology and resource estimation, Grasberg Cu-Au Deposit, Freeport-

McMoRan Inc, Papua, Indonesia 1995-1998; • Corporate Strategic Planning: Geology and Resources, Freeport-McMoRan Inc., New Orleans, LA.,

1999; • Independent Consultant: Geology, Steamboat Springs, CO., 2000; • Senior Geologist, environmental geology and mine closure, MWH Consulting, Inc., Steamboat

Springs, CO., 2000-2003; • Principal Mining Geologist, precious and base metal exploration, resource modeling, and mine

development, SRK Consulting (U.S.), Inc., 2004 to present; • Experience in the above positions working with, reviewing and conducting resource estimation and

feasibility studies in concert with mining and process engineers; and • As a consultant, I have participated in the preparation of NI 43-101 Technical reports from 2006-

2009.

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 of NI 43-101.

5. I visited the Crescent Silver Mine property on April 22, 2013 for 2 days. 6. I am responsible for the preparation of Sections 2-12, 14, and portions of Sections 1, 25 and 26

summarized therefrom. 7. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101. 8. I have not had prior involvement with the property that is the subject of the Technical Report. 9. I have read NI 43-101 and Form 43-101-F1 and the sections of the Technical Report I am responsible for

have been prepared in compliance with that instrument and form.

U.S. Offices: Anchorage 907.677.3520 Denver 303.985.1333 Elko 775.753.4151 Fort Collins 970.407.8302 Reno 775.828.6800 Tucson 520.544.3688

Mexico Offices: Zacatecas 52.492.927.8982 Querétaro 52.442.218.1030

Canadian Offices: Saskatoon 306.955.4778 Sudbury 705.682.3270 Toronto 416.601.1445 Vancouver 604.681.4196 Yellowknife 867.873.8670

Group Offices: Africa Asia Australia Europe North America South America

SRK Consulting (U.S.), Inc. Page 2 10. As of the aforementioned Effective Date, to the best of my knowledge, information and belief, the

sections of the Technical Report I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 24th Day of September, 2013. “Signed” “Sealed” ________________________________ CPG # 11245

J. B. Pennington, CPG

QP_Cert_Pennington_Jay_2013

SRK Consulting (U.S.) Inc. 5250 Neil Road, Suite 300 Reno, Nevada 89502 T: (775) 828-6800

F: (775) 828-6820 [email protected] www.srk.com

CERTIFICATE OF AUTHOR

I, Kent W. Hartley, BSc, PE Mining, do hereby certify that:

1. I am Principal Consultant of SRK Consulting (U.S.), Inc., 5250 Neil Road, Suite 300, Reno, Nevada, 89502.

2. This certificate applies to the technical report titled “NI 43-101 Technical Report, Preliminary Economic Assessment, Crescent Silver Project, Kellogg, Idaho, USA” with an Effective Date of July 22, 2013 (the “Technical Report”).

3. I graduated with a degree in Mining Engineering from Michigan Technological University in 1979. I have worked as an Engineer for a total of 30+ years since my graduation from university. My relevant experience includes mine planning and project engineering at a number of open pit and underground mines as well as construction management and cost estimating experience. I am a registered Professional Engineer in Nevada, license number 021612.

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 of NI 43-101.

5. I visited the Crescent Silver Mine property on August 8, 2012 for 2 days and on April 22, 2013 for 2 days. 6. I am responsible for the preparation of Sections 13, and 15-24, and portions of Sections 1, 25 and 26 of

the Technical Report. 7. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101. 8. I have not had prior involvement with the property that is the subject of the Technical Report. 9. I have read NI 43-101 and Form 43-101-F1 and the sections of the Technical Report I am responsible for

have been prepared in compliance with that instrument and form. 10. As of the aforementioned Effective Date, to the best of my knowledge, information and belief, the

sections of the Technical Report I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 24th Day of September, 2013. “Signed” “Sealed” ________________________________ PE Nevada #021612

Kent W. Hartley, BSc, PE Mining

U.S. Offices: Anchorage 907.677.3520 Denver 303.985.1333 Elko 775.753.4151 Fort Collins 970.407.8302 Reno 775.828.6800 Tucson 520.544.3688

Mexico Offices: Zacatecas 52.492.927.8982 Querétaro 52.442.218.1030

Canadian Offices: Saskatoon 306.955.4778 Sudbury 705.682.3270 Toronto 416.601.1445 Vancouver 604.681.4196 Yellowknife 867.873.8670

Group Offices: Africa Asia Australia Europe North America South America

ni 43-101 technical report preliminary economic assessment ... · srk consulting (u.s.), inc. ni...

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