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NI 43-101

Technical Report

On the Mineral Resource and Mineral Reserve Estimation

for the Björkdal Gold Mine, Sweden

August 28, 2013

Effective Date: March 31, 2013

Prepared by:

James A. Currie, P.Eng. George Friesen, P.Eng. Gordon Clarke, P.Geol.

Björkdal Gold Mine, August 2013

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Table of Contents

1. EXECUTIVE AND TECHNICAL SUMMARY .......................................................................................... 1

2. INTRODUCTION ................................................................................................................................ 6

3. RELIANCE ON OTHER EXPERTS ......................................................................................................... 7

4. PROPERTY DESCRIPTION AND LOCATION ........................................................................................ 8

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY .............. 12

6. HISTORY ........................................................................................................................................ 13

7. GEOLOGICAL SETTING AND MINERALIZATION .............................................................................. 14

7.1 Regional Geology ........................................................................................................................ 14

7.2 Local and Property Geology ........................................................................................................ 15

7.3 Mineralization ............................................................................................................................. 17

8. DEPOSIT TYPE ................................................................................................................................. 20

9. EXPLORATION ................................................................................................................................ 21

10. DRILLING ........................................................................................................................................ 22

11. SAMPLING PREPARATION ANALYSES AND SECURITY .................................................................... 24

11.1 Core Logging Protocol ................................................................................................................. 24

11.2 Sample Collection ....................................................................................................................... 24

11.2.1 Diamond Drill Core .............................................................................................................. 24

11.2.2 Reverse Circulation Samples ............................................................................................... 25

11.3 Core Recovery ............................................................................................................................. 25

11.4 Analytical Procedures and Protocol ............................................................................................ 25

11.5 Quality Assurance/Quality Control Program (“QA/QC”) ............................................................ 25

11.5.1 Certified Standards ............................................................................................................. 26

11.5.2 Blind Blank Insertion ........................................................................................................... 26

11.5.3 Duplicates ............................................................................................................................ 28

11.5.4 Chip Sampling ...................................................................................................................... 28

11.6 Sample Security ........................................................................................................................... 29

12. DATA VERIFICATION ....................................................................................................................... 30

13. MINERAL PROCESSING AND METALLURGICAL TESTING ................................................................ 31

14. MINERAL RESOURCE ESTIMATES ................................................................................................... 32


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14.1 Topographical Survey .................................................................................................................. 33

14.2 Database Compilation ................................................................................................................. 33

14.3 Drilling ......................................................................................................................................... 34

14.4 Chip Sampling.............................................................................................................................. 38

14.5 Geological Interpretation ............................................................................................................ 39

14.6 Sample Data Processing .............................................................................................................. 40

14.6.1 Statistical Analysis ............................................................................................................... 40

14.6.2 Domain ................................................................................................................................ 41

14.6.3 Sampling Method ................................................................................................................ 42

14.6.4 Removal of Outlier Grades .................................................................................................. 46

14.6.5 Composting ......................................................................................................................... 50

14.7 Variography ................................................................................................................................. 51

14.7.1 Variogram Parameters ........................................................................................................ 52

14.7.2 Variography Interpretation ................................................................................................. 52

14.8 Block Modelling ........................................................................................................................... 52

14.8.1 Density ................................................................................................................................ 55

14.9 Grade Estimation ........................................................................................................................ 55

14.9.1 Kriging Plan.......................................................................................................................... 55

14.9.2 Validation ............................................................................................................................ 56

14.10 Selective Mining Units ................................................................................................................ 57

14.11 Depletion ..................................................................................................................................... 57

14.12 Reconciliation .............................................................................................................................. 57

14.12.1 Open Pit Production June 30, 1993 to February 1, 1998 ............................................... 58

14.12.2 Open Pit Production February 1, 1998 to July 31, 1999 ................................................ 59

14.12.3 Open Pit Production January 1, 2009 to January 1, 2013 .............................................. 61

14.12.4 Underground Production 2008 to 2013 ......................................................................... 62

14.13 Resource Classification ............................................................................................................... 64

14.14 Resource Evaluation .................................................................................................................... 65

14.15 Revised Open Pit Resources ........................................................................................................ 67

15 MINERAL RESERVES ESTIMATE ...................................................................................................... 70

15.1 Introduction ................................................................................................................................ 70


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15.2 Pit Reserves ................................................................................................................................. 70

15.2.1 Optimisation Parameters Pit Reserves ............................................................................... 70

15.2.2 Pit Reserves Results ............................................................................................................ 71

15.3 Underground Reserves ............................................................................................................... 72

15.4 Methodology: .............................................................................................................................. 73

15.4.1 Creating Solids .................................................................................................................... 73

15.4.2 Ore Reserves from the Block Model ................................................................................... 73

15.4.3 Mining Costs and Determination of Economic Cut off Grade for Underground ................ 77

16.0 MINING METHODS ......................................................................................................................... 79

17. RECOVERY METHODS ..................................................................................................................... 81

17.1 General ........................................................................................................................................ 81

17.2 Recovery Description .................................................................................................................. 81

17.2.1 Crushing .............................................................................................................................. 81

17.2.2 Grinding ............................................................................................................................... 81

17.2.3 Gravity Processing ............................................................................................................... 81

17.2.4 Flotation .............................................................................................................................. 82

17.2.5 Concentrate Dewatering ..................................................................................................... 82

17.2.6 Sampling .............................................................................................................................. 82

17.3 Production Data .......................................................................................................................... 82

17.4 Tailings Maintenance Facility ...................................................................................................... 83

18 PROJECT INFRASTRUCTURE ........................................................................................................... 84

19 MARKET STUDIES AND CONTRACTS .............................................................................................. 85

20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ....................... 86

21. CAPITAL AND OPERATING COSTS .................................................................................................. 87

21.1 Capital Costs ................................................................................................................................ 87

21.2 Operating Costs ........................................................................................................................... 87

22 ECONOMIC ANALYSIS ..................................................................................................................... 89

23 ADJACENT PROPERTIES .................................................................................................................. 90

24 OTHER RELEVANT DATA AND INFORMATION ............................................................................... 91

25 INTERPRETATION AND CONCLUSIONS .......................................................................................... 92

26 RECOMMENDATIONS .................................................................................................................... 93


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27. REFERENCES ................................................................................................................................... 94

28. APPENDIX 1 – VARIOGRAPHY Y (From: Ellis and Newall 2013) ..................................................... 95

29. CERTIFICATE OF AUTHOURS ........................................................................................................ 101

List of Figures

Figure 4.1 Location of Björkdal Gold Mine, Northern Sweden (modified from google maps) ............... 8

Figure 4.2 Björkdal License Boundaries (boundary shape files from Mining Directorate of Sweden) . 11

Figure 7.1 Regional Geology Skellefte District (Mercier-Langevin et al. 2013)..................................... 14

Figure7.2 Local and Property Geology (Billstrm et al. 2009) ............................................................... 16

Figure 11.1 Summary of Blank Assaying for Drillhole Samples by Year .................................................. 27

Figure 11.2 Log Probability Plot Comparing Duplicate Underground Chip Samples by Fire Assay at the Björkdal Laboratory (MILL_1, MILL2 and MILL_AVE) and Leachwell Assay at the ALS Chemex Laboratory (Piteå) (Ellis and Newall 2013). ............................................................ 29

Figure 14.1 Topographical Survey of Björkdal Mine as of January 1st 2013 also Showing Extent of Underground Development (orange) (Ellis and Newall 2013) ............................................. 33

Figure 14.2 Location of DDH Drill holes at Björkdal (Ellis and Newall 2013) .......................................... 37

Figure 14.3 Location of RC Drill holes at Björkdal (Ellis and Newall 2013) ............................................. 37

Figure 14.4 Location of Underground Chip Samples at Björkdal (Ellis and Newall 2013) ...................... 38

Figure 14.5 Location of Björkdal Domains (Ellis and Newall 2013) ......................................................... 40

Figure 14.6 Log Probability Plot of all Samples by Domain (Ellis and Newall 2013) ............................... 41

Figure 14.7a Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 11 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 43

Figure 14.7b Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 12 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 43

Figure 14.7c Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 13 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 44

Figure 14.7d Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 14 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 44

Figure 14.7e Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 21 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 45


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Figure 14.7f Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 22 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013) .................................................................................................................................... 45

Figure 14.8a Log Probability Plot Showing Top Cut Level For Au For Domain 11 (Ellis and Newall 2013)47

Figure 14.8b Log Probability Plot Showing Top Cut Level For Au For Domain 12 (Ellis and Newall 2013)47

Figure 14.8c Log Probability Plot Showing Top Cut Level For Au For Domain 13 (Ellis and Newall 2013)48

Figure 14.8d Log Probability Plot Showing Top Cut Level For Au For Domain 14 (Ellis and Newall 2013)48

Figure 14.8e Log Probability Plot Showing Top Cut Level For Au For Domain 21 (Ellis and Newall 2013)49

Figure 14.8f Log Probability Plot Showing Top Cut Level For Au For Domain 22 (Ellis and Newall 2013)49

Figure 14.9 Standard Histogram Showing Sample Length for All Mineralized Samples (Ellis and Newall 2013...................................................................................................................................... 50

Figure 14.10 Cross Section Showing Extent of Mineralization in Indicator Model (Ellis and Newall 2013) . ......................................................................................................................................... 54

Figure 14.11 Isometric View of Björkdal Underground Development Block Model (Ellis and Newall 2013) ......................................................................................................................................... 63

Figure 14.12 Plan View Illustrating Resource Classification (Ellis and Newall 2013) ................................ 65

Figure 15.3.1 Underground Zones – Björkdal Mine ................................................................................... 72

Figure 15.4.2 3D View of Mined Out Areas and Stopes in Ore Reserve. Mined out areas are shown in light green. ........................................................................................................................... 77

List of Photos

Photo 4.1 View of the Björkdal Pit, Looking East..…………….……………………………………………………………9

Photo 5.1 Typical Forest near the Björkdal Mine… .......................................................................... 12

Photo 7.1 Granodiorite-LimestoneContact.……………………….………………………………………………………..17

Photo 7.2 Typical Veining Terminating At The Dipping Structural Contact………….……………………….18

Photo 7.3 Typical Vein Structures in face of cross cut drift…………………………………………………………..19

List of Tables

Table 1.1 Open Pit & Underground Mineral Resource Estimate – (Ellis and Newall 2013) .................. 2

Table 1.2 Revised Open Pit Resources and Underground Resources .................................................... 3

Table 1.3: Open Pit Reserves – Elgin March 2013 .................................................................................. 4

Table 1.4 Underground Reserves – Elgin March 2013 ........................................................................... 4

Table 4.1 Björkdal Property ................................................................................................................... 9


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Table 6.1 Björkdal Annual Gold Production (2006 -2012) ................................................................... 13

Table 7.1 Legend for Figure 7.1 (Mercier-Langevin et al. 2013) .......................................................... 15

Table 10.1 Historical Drill-hole Summary .............................................................................................. 22

Table 10.2 Drilling Completed by Gold-Ore and Elgin ........................................................................... 23

Table 11.1 Summary of Certified Standards Used ................................................................................. 26

Table 14.1 Rejected Drill holes Due to Contamination and Reliability Issues ....................................... 34

Table14.2 Summary of Björkdal Drill hole Database ............................................................................ 36

Table 14.3 Summary of Björkdal Chip Sample Database ....................................................................... 39

Table 14.4 Summary of Björkdal Domains ............................................................................................. 40

Table 14.5 Statistical Analysis of Raw Samples >0.3g/t by Domain....................................................... 41

Table 14.6 Statistical Analysis of Raw Samples >0.3g/t by Domain and Sampling Method .................. 42

Table 14.7 Summary of Top-Cut Values ................................................................................................. 46

Table 14.8 Statistical Analysis of Top-Cut Samples >0.3g/t by Domain ................................................. 46

Table 14.9 Statistical Analysis of Composites >0.3g/t ........................................................................... 51

Table 14.10 Summary of Indicator Block Model Prototype Parameters ................................................. 53

Table 14.11 Summary of Block Model Fields ........................................................................................... 54

Table 14.12 Björkdal Kriging Plan (OK) .................................................................................................... 56

Table 14.13 Open Pit Production Data – June 30, 1993 to February 1, 1998.......................................... 58

Table 14.14 Block Model Open Pit Evaluation – June 30, 1993 to February 1, 1998. Measured and Indicated Resources ............................................................................................................. 59

Table 14.15 Open Pit Production Data – February 1, 1998 to July 31, 1999 ........................................... 60

Table 14.16 Block Model Open Pit Evaluation – February 1, 1998 to July 31, 1999. Measured and Indicated Resources ............................................................................................................. 60

Table 14.17 Open Pit Production Data – January 1, 2009 to January 1, 2013......................................... 61

Table 14.18 Block Model Open Pit Evaluation - – January 1, 2009 to January 1, 2013. Measured and Indicated Resources ............................................................................................................. 62

Table 14.19 Underground Development Production Data (On-vein drives and cross cuts - not including stopes) .................................................................................................................................. 63

Table 14.20 Block Model Evaluation Of Underground Development (On-vein drives and cross cuts only - not including stopeing) ...................................................................................................... 64

Table 14.21 Björkdal Underground Resource Estimate (WAI) (Ellis and Newall 2013) ........................... 66

Table 14.22 Björkdal Open Pit Resource Estimate (WAI) (Ellis and Newall 2013) ................................... 67

Table 14.23 Parameters for Elgin Revised Open Pit Resource Estimate ................................................. 68


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Table 14.24 Open Pit Resource Estimate Using Optimized Pit Shell ....................................................... 69

Table 15.1 Parameters for Elgin Revised Open Pit Reserves Estimate .................................................. 70

Table 15.2 Björkdal Pit Reserves ............................................................................................................ 71

Table 15.3.1 Björkdal Underground Reserves as at April, 2013 ............................................................... 72

Table 15.4.2 Ore Reserve Calculations ..................................................................................................... 74

Table 15.4.3 Underground Cut-Off Grade Cost Inputs ............................................................................. 78

Table 17.1 Production Data 2006-2012 ................................................................................................. 83

Table 21.2.1 Björkdal Production Data 2012 ............................................................................................ 87


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1. EXECUTIVE AND TECHNICAL SUMMARY

Elgin Mining Inc. (“Elgin” or the “Company”) (TSX: ELG) has prepared this National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (“NI 43-101”) compliant technical report on the mineral resource and mineral reserve estimates for the Björkdal gold mine, Sweden. The report was prepared by the management of Elgin under the direction and supervision of James A. Currie, P.Eng., Chief Operating Officer of Elgin, George Friesen, P.Eng. Technical Manager of Elgin and Gordon Clarke, P. Geol., former Vice President Exploration of Elgin acting together as the Qualified Persons for this report. Elgin is a producing issuer as defined by NI 43-101.

The Björkdal gold mine (the “Björkdal Mine”) is located in northern Sweden, 28 km northwest of the town of Skelleftea. This area of Sweden has had active mining for over 100 years. There is excellent infrastructure in place including grid power, paved roads, educated labour force and readily accessible technical support. Sweden has no overriding royalties on the mine production.

The Björkdal property consists of 6 mining concessions and 15 exploration permits as listed in Table 4.1 and shown on Figure 4.2. The 6 mining concessions and 16 exploration permits comprising the Björkdal Mine are owned 100% by Björkdalsgruvan AB and Björkdal Exploration AB, both of which are wholly owned subsidiaries of Gold-Ore Resources Ltd. (“Gold-Ore”).

On May 1, 2012 Elgin and Gold-Ore completed a business combination (the “Arrangement”) whereby Elgin acquired all of the issued and outstanding common shares of Gold-Ore. Delisting of the Gold-Ore common shares from the Toronto Stock Exchange (TSX) occurred on May 3, 2012 and Elgin graduated from a TSX Venture listed Company to a Toronto Stock Exchange listed Company on May 4, 2012.

The Björkdal Mine has operated more or less continuously for 25 years mainly processing ore mined from the open pit. In 1999 the mine was closed for two years due to the prevailing low gold price. Gold-Ore acquired an option to purchase the mine in 2006 and in December 2008 exercised the option and purchased the mine for a total of $8.8 million in a combination of shares and cash. In January of 2009, Gold-Ore declared commercial production. Since that date, a significant portion of production has been from underground. Over 1.0 million ounces of gold has been produced from the mine. For 2012, the production rate was 1.27 million tonnes and the mine produced 42,839 ounces of gold. The mine is currently operating as a combination open pit and underground mine with approximately equal amounts of ore coming from each source.

The Björkdal deposit is located in the prolific Skelleftea mining district, a Paleoproterozoic greenstone belt primarily known for volcanogenic massive sulphide deposits containing high gold values. The gold at the Björkdal Mine is associated with a sheeted quartz vein complex hosted in moderately north dipping granodiorite. The veins are near vertical and terminate at a hanging wall structure that separates the host rocks from overlying metasedimentary and metavolcanic rocks.

Wardell Armstrong International Ltd. (“WAI”), an independent contractor from the United Kingdom, prepared an updated block model and calculated an updated resource classification for the Björkdal Mine using survey information as of January 1st 2013 and incorporating drill data from 2,245 holes up to April 23, 2013 and underground chip samples up to December 12, 2012. The block model used Datamine software and Ordinary Kriging. The resource classification is in accordance with the Australian Code for


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Reporting of Exploration Results, Mineral Resources and Ore Reserves [JORC Code (2004)]. The WAI block model was reviewed and approved by Messrs. Friesen, Currie and Clarke and the resource estimates are summarized in Table 1.1. The WAI resource estimates were used in the estimation of a revised open pit resource and reserve estimates by Mr. Currie and underground reserve estimates by Mr. Friesen.

Table 1.1 Open Pit & Underground Mineral Resource Estimate – (Ellis and Newall 2013)

TONNAGES AND GRADES

Resource Type

Cut-off

Measured Indicated Measured +

Indicated Inferred

Au (g/t)

Tonnage (000’s t)

Au (g/t)

Tonnage(000’s t)

Au (g/t)

Tonnage (000’s t)

Au (g/t)

Tonnage (000’s t)

Au (g/t)

Open Pit 0.30 113 1.68 9,205 1.07 9,318 1.08 7,320 1.06

Underground 0.60 162 2.42 10,976 2.17 11,139 2.17 9,003 2.21

Totals* 275 2.12 20,181 1.67 20,457 1.67 16,323 1.69

Ounces

Cut-off

Measured (ounces)

Indicated (ounces)

Measured + Indicated (ounces)

Inferred (ounces)

Open Pit 0.30 6,111 317,727 323,839 250,237

Underground 0.60 12,624 765,319 777,944 640,972

Totals* 18,735 1,083,046 1,101,783 891,209

Evaluation (4.0 m Selectivity) * Totals may differ due to rounding

Using the WAI block model for open pit resources James Currie of Elgin completed a Pit Optimization Study to calculate revised open pit resources. SRK Consulting has prepared a geotechnical review of the Björkdal open pit wall angles. The pit slope parameters used in the pit design for the open pit reserves are based on the specifications in Figure 7-3 from the report SE-384 “Inter-Ramp Slope Stability Analysis for Björkdal Mine” prepared by David Saiang (Saiang 2012). The inter-ramp slope height used in the final design was reduced to 20m from 40m for safety purposes, with 10m catch benches. A bench slope angle of 80 degrees was selected (a safe range of 70-90 degrees was recommended) and will be achieved through pre-split blasting. Table 1.2 lists underground resource estimates by WAI along with the revised open pit resource estimates by Elgin.


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Recommendations include an underground detailed bulk sample study, streamlining and updating stope surveys to produce shapes and models on an ongoing basis and continuation of efforts to reduce mining dilution by cable bolting along with improvements in sampling, drilling, and blasting practices to improve selectivity of higher grade ore.

As well, remnants and pillars above the active mine need to be re-evaluated as potential ore and viability of mining by future open pit methods. A plan for integrating drilling and sample information directly with the mine database through new software should be implemented. Exploration drilling should be continued to allow conversion of inferred resources to indicated, expanding the mine resource to help extend the mine life beyond 2020.

Table 1.2 Revised Open Pit Resources and Underground Resources

* Open pit resources calculated by WAI were adjusted based on an optimized pit shell by James A. Currie, P.Eng., of Elgin. ** Totals may differ due to rounding. Mineral reserves are included in the mineral resources. Mineral resources that are not mineral reserves

do not have demonstrated economic viability.

Using the WAI block model for the open pit and underground resources, proven and probable reserves were calculated by Elgin. Pit reserves were calculated by Mr. Currie and underground reserves were calculated by George Friesen. Reserves included drilling up to April 2013 and were adjusted for depletion up to the end of March 2013. Open pit reserves are summarized in Table 1.3 and underground reserves are summarized in Table 1.4.

TONNAGES AND GRADES

Resource Type Cut-off

Measured Indicated Measured +

Indicated Inferred

Au (g/t)

Tonnage(000’s t)

Au (g/t)

Tonnage(000’s t)

Au (g/t)

Tonnage(000’s t)

Au (g/t)

Tonnage(000’s t)

Au (g/t)

Open Pit* 0.32 51.8 1.65 6,469 1.10 6,521 1.11 2,216 1.57

Underground 0.60 162 2.42 10,976 2.17 11,139 2.17 9,003 2.21

Totals** 213.8 2.23 17,445 1.77 17,660 1.78 11,219 2.08

Ounces

Cut-off

Measured (ounces)

Indicated (ounces)


Inferred (ounces)

Open Pit* 0.32 2,746 229,611 232,357 111,907

Underground 0.60 12,624 765,319 777,944 640,972

Totals** 15,370 994,930 1,010,301 752,879


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Table 1.3: Open Pit Reserves – Elgin March 2013

Reserves* Tonnes (1000’s)

Grade

(g/t) Ounces

Proven 49.6 1.48 2,366

Probable 5,454.4 1.04 182,858

Total** 5,504.0 1.05 185,224

* Cut-off - 0.34 g/t; Gold Price $1500/Oz USD; Strip Ratio 4.5; 25% dilution ** Totals may vary due to rounding. Mineral reserves are included in the mineral resources. Mineral resources that are not mineral reserves

do not have demonstrated economic viability.

Table 1.4 Underground Reserves – Elgin March 2013


Grade

(g/t) Ounces

Proven 36.2 2.02 2,357

Probable 2,054.7 2.07 136,732

Total** 2,090.9 2.07 139,089

* Cutoff grade 1.15 g/t; 2.5 m selectivity; Gold Price $1500/Oz USD; 80% mining recovery; 30% mining dilution ** Totals may vary due to rounding

The Björkdal Mine produces approximately 42,000 - 46,000 ounces of gold annually by processing 1.3 million tonnes of ore. The mine is fully permitted and operates 7 days a week all year round. Four separate gold concentrates are produced and sold to smelters. The ore is sourced by open pit and underground mining techniques by a combination of owner operated and contractors’ fleet.

A new reserve estimate was completed in July 2013 using a base case gold price of US$1,500 per ounce which is the average market price over the last three years. In the latter half of Q1 and Q2 2013, the price of gold has been extremely volatile and generally trending downwards. Currently it is at a mid $1300 per ounce level. Should the price of gold remain at or below $1,300 for an extended period a new reserve estimate should be considered with a higher cut-off grade.

The proven and probable reserve estimate for the open pit unit as of the end of March, 2013 is 5.5 million tonnes of ore grading 1.05 grams gold per tonne. At the current mining rate of about 700,000 tonnes of ore per annum, the open pit has sufficient reserves for approximately 8 years of production.


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The proven and probable reserve estimate for the underground unit as of the end of March, 2013 is 2.09 million tonnes of ore grading 2.07 grams gold per tonne. At a current production rate of about 600,000 tonnes of ore per annum, the underground has sufficient reserves for approximately 3.5 years of production. There is a significant amount of tonnage in the resource category underground. As the underground access is continually developed and the infrastructure is established, the mine will continually drill the ore body to further define the resources and convert them into reserves. The mine has been successful at replacing the reserves mined underground for the last four years.


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2. INTRODUCTION

Elgin Mining Inc., is based in Vancouver Canada and is a listed issuer on the TSX. Elgin trades under the symbol ELG.

The information used in this report is largely based on company records and the operating history of the Björkdal Mine since commencement of production in 1988. Additional information has been provided by WAI with respect to mineral resources calculations.

The Qualified Persons responsible for this report have conducted several visits during the last six months (see Certificate of each Qualified Person for details of most recent site visit) and have a thorough understanding of all aspects of the operation.

This Report includes technical information which requires subsequent calculations or estimates to derive sub-totals, totals and weighted averages. Such calculations or estimations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, Elgin does not consider them to be material. Further, the Report summarizes the professional opinion of the authors and includes conclusions and estimates that have been based on professional judgment and reasonable care.

Said conclusions and estimates are consistent with the level of detail of this study and based on the information available at the time this Report was completed. All conclusions and estimates presented are based on the assumptions and conditions outlined in this Report.

This Report is to be issued and read in its entirety. Written excerpts from this Report may not be used without the express written consent of the authors or officers of Elgin.


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3. RELIANCE ON OTHER EXPERTS

The authors have assumed that previous companies’ reports, maps and other geological data that are listed in the References Section were complete and accurate. Parts of these reports have been used for this report. In particular the qualified Persons for this report are relying on mineral resource estimates completed by other experts.

WAI has prepared several resource estimates for the Björkdal property. For the resource estimates referenced in this technical report, the authors have relied on the WAI report titled “Björkdal Resource Estimate and Open Pit Reserve Estimation”. The WAI report number is: MM832, dated 08/07/2013 and was authoured by Phil Newall, Bsc (ARSM), CEng, FIMMM and Richard Ellis BSc., MSc., (MCSM) FGS. The WAI block model was reviewed and approved by Messrs. Currie, Friesen, and Clarke and used in the estimation of a revised open pit resource and for open pit reserve estimates by Mr. Currie and underground reserve estimates by Mr. Friesen.


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4. PROPERTY DESCRIPTION AND LOCATION

The Björkdal Mine is located in northern Sweden, 28km northwest of the town of Skelleftea, in Vasterbotten County, and some 750km north of Stockholm (see Figure 4.1).

Figure 4.1 Location of Björkdal Gold Mine, Northern Sweden

The terrain around Björkdal is relatively subdued with low hills and numerous shallow lakes as shown in Photo 4.1. Glacial till forms the main soil cover over the area.


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Photo 4.1 View of the Björkdal Pit, Looking East

The Björkdal property consists of 6 mining concessions and 15 exploration permits as listed in Table 4.1 and shown on Figure 4.2. The concessions and permits are owned 100% by Björkdalsgruvan AB and Björkdal Exploration AB, both of which are wholly owned subsidiaries of Gold-Ore. On May 1, 2012 Elgin and Gold-Ore completed an Arrangement whereby Elgin acquired all of the issued and outstanding common shares of Gold-Ore.

Table 4.1 Björkdal Property

PERMIT NAME PERMIT TYPE SIZE (ha) EXPIRY DATE

Björkdal Exploitation Concessions

Häbbersfors K nr 1 Exploitation 98.6894 Jan. 1, 2031

Häbbersfors K nr 2 Exploitation 34.8839 Feb. 2, 2025

Häbbersfors K nr 3 Exploitation 18.8864 Ap. 29, 2027

Häbbersfors K nr 4 Exploitation 5.0012 Nov. 21, 2025

Habbersfors K nr 5 Exploitation 21.8263 Mar. 6, 2034

Habbersfors K nr 6 Exploitation 23.4887 Apr. 24, 2038

TOTAL

202.7759


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PERMIT NAME PERMIT TYPE SIZE (ha) EXPIRY DATE

Björkdal Exploration Permits

Björkdal nr 26 Exploration 978.80 Renewal pending

Björkdal nr 10 Exploration 712.64 Dec. 18, 2015

Björkdal nr 25 Exploration 967.70 May 9, 2014

Björkdal nr 19 Exploration 225.00 Oct. 18, 2015



Björkdal nr 27 Exploration 578.77 May 5, 2014

Lillträsket nr 2 Exploration 246.97 Oct. 14, 2015

Norrberget nr 200* Exploration 50.00 Renewal pending

Norrberget nr 300* Exploration 37.50 May 23, 2014

Norrberget nr 400* Exploration 291.00 Oct. 01, 2013

Vidmyran nr 100 Exploration 1197.50 Renewal pending

Olofsberg nr 101* Exploration 42.70 Feb 15, 2016

Björkdal nr 29 Exploration 1073.89 Nov. 30, 2013

Björkdal nr 30 Exploration 64.03 Feb 23, 2014

TOTAL

6659.08

* Exploration permits that are subject to a 2% NSR in favour of North Atlantic Natural Resources AB


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Figure 4.2 Björkdal License Boundaries (boundary shape files from Mining Directorate of Sweden)


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5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The mine is accessed from Skelleftea by good paved roads with a driving time of around 30 minutes. Skelleftea is serviced by regular flights from Stockholm (flying time - 1 hour) and onwards to the rest of the world.

The area has a long history of mining and has developed an excellent infrastructure including good roads, relatively low-cost hydro power and a skilled workforce. Mining is accepted as a socially responsible and necessary contributor to the local economy. As a working mine, it is fully supported by power, water and communications.

This area of Sweden has a subarctic climate with mild summers and cold snowy winters. The climate is however moderated by the Gulf Stream in the Atlantic Ocean. While winters are cold they are much less so than similar latitudes in other parts of the world. The average low temperature for January is -14°C. The short summers are also reasonably warm for latitudes near the Arctic Circle. The average daily high temperature in July is 19°C. Yearly precipitation is low at under 600 mm, with August being the wettest month at over 70 mm. Due to its high latitude, July is typified by an average of 21 hours of daylight while the average for December is four. The Björkdal Mine operates year round.

The vegetation around Björkdal consists of dominantly managed forests of spruce and birch as shown in Photo 5.1 with some areas of cultivated land.

Photo 5.1 Typical Forest near the Björkdal Mine


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6. HISTORY

The Björkdal deposit was originally discovered in 1983 by Terra Mining AB from anomalous gold values in till samples, which were based on a 2 kilometer sample spacing, later tightened up to 75m. Anomalous bedrock values were obtained in 1985 and definition drilling began in early 1986.

Definition drilling was coincident with metallurgical testwork and positive feasibility studies that were completed in May 1987. Terra Mining commenced mining operations at the Björkdal Mine in July 1988. In 1996 Terra Mining was bought out by William Resource Ltd. (“William”). William continued to operate the mine until the end of June, 1999 when they were petitioned into bankruptcy. The assets were bought through public auction in June 2001 by International Gold Exploration, which operated the mine from September 2001 until 2003 when it was acquired by Minmet plc. Up to that date, some 38 million tonnes of waste and 15 million tonnes of ore had been mined out of the open pit yielding approximately 26.6 tonnes of gold.

In 2006, Gold-Ore acquired an option from Minmet plc to purchase the holding company for the mine. On December 31, 2007 Gold-Ore exercised its option and acquired all the shares of Björkdalsgruvan AB. During exploration and development of the Björkdal Mine, Gold-Ore generated cash flow from gold sales from the operation of the plant at the mine, fed by stockpiled material, some newly-mined open pit material, and material from underground development operations. In January 2009, Gold-Ore’s management concluded that there were sufficient mineral reserves and resources at the Björkdal Mine for at least a five year mine life and declared commercial production.

In May 2012, Elgin acquired all of the issued and outstanding common shares of Gold-Ore. Gold-Ore’s common shares were delisted from the TSX and Elgin graduated from a TSX Venture listed Company to a TSX listed Company. Table 6.1 shows annual gold production up to 2012 subsequent to Gold-Ore’s acquisition from Minmet and the acquisition by Elgin.

Table 6.1 Björkdal Annual Gold Production (2006 -2012)

Year Ounces of Gold Produced

2006 20,503

2007 19,198

2008 29,296

2009 37,568

2010 40,729

2011 40,762

2012 42,839


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7. GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Björkdal Mine is located in the Skellefte district in northern Sweden which has most recently been summarized by Mercier-Langevin et al. (2013) as forming a west-northwest trending belt of about 120×30 km that consists mainly of Early Proterozoic volcanic rocks with lesser amounts of intrusive and sedimentary rocks (Figure 7.1). The Skellefte belt is thought to overlie the dominantly sedimentary rocks of the Bothnian Basin that are exposed to the south. To the north, it is in contact with, and probably underlies, the well-preserved volcanic, sedimentary, and intrusive rocks of the Arvidsjaur group. The Skellefte belt comprises a >3-km thick volcanic-dominated succession that forms the Skellefte group and the overlying >4-km thick volcano-sedimentary succession that forms the Vargfors group.

The Skelleftea Group consists of two metavolcanic units overlain by a metasedimentary unit. The lower volcanic unit consists of felsic lavas and a volcaniclastic rock sequence which grades upwards into a series of bimodal rocks. The metasedimentary unit consists of mainly greywackes. In the transition zone between the bimodal metavolcanic unit and the metasedimentary unit, several massive sulphide deposits are present.

At least two phases of deformation have folded the supracrustal rocks in the Skellefte district and the regional metamorphism is greenschist facies, but locally reaches lower amphibolite facies. The supracrustal rocks have been intruded by at least three generations of granites ranging in age from approximately 1.95 and 1.78 Ga.

Figure 7.1 Regional Geology Skellefte District (Mercier-Langevin et al. 2013)


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Table 7.1 Legend for Figure 7.1 (Mercier-Langevin et al. 2013)

7.2 Local and Property Geology

The Björkdal gold deposit is located in the eastern portion of the Skelleftea District. The intrusion hosting the gold bearing veins and the adjacent metavolcanic and marble units form a circular structure that is almost completely enclosed by greywacke dominated metasedimentary rocks (Figure 7.2).


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Figure7.2 Local and Property Geology (Billstrm et al. 2009)


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The gold-bearing quartz veins occur at the contact between a tonolite or granodiorite intrusion with the overlying supracrustal rocks. The intrusive has been described as dome-shaped intrusive. This granotoid pluton or sill has intruded a sequence consisting of a lower felsic metavolcanic unit with interbedded marbles and an upper felsic to intermediate metavolcanic unit with associated metasediments.

The Björkdal granodiorite, contacts the overlying metasediments and metavolcanic rocks as shown in Photo 7.1. A strong biotite alteration halo has been mapped in the immediate area of the pit and affects the entire package of rocks over a 2 kilometer by 2 kilometer area. The intrusion/country rock contact is interpreted to be a significant thrust fault.

Photo 7.1 Granodiorite-Limestone Contact

7.3 Mineralization

Björkdal lies within the Skellefte District and lies some 14km northeast of the well known Boliden Mine. The district contains more than 85 pyritic Zn-Cu-Au-Ag massive sulphide deposits, of which 28 have been mined since 1924, and 5 are currently in operation. Fifty-two deposits have over 0.1Mt ore and the combined total of these is 161Mt at an average grade of 1.9g/t Au, 47g/t Ag, 0.7% Cu, 3.0% Zn and 0.4% Pb. They are characterized by high Au, As, Sb and Hg.

From detailed studies undertaken in the district, other types of deposits have also been identified:

High sulfidation epithermal Au model interpreted for Boliden;

Mesothermal gold deposits - Björkdal (Intrusion related gold) and Akerberg;

Porphyry Au-Cu-Mo at Tallberg;

Minor Ni in komatiities; and

Minor Li and Cs in pegmatites.


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The gold mineralization at Björkdal is associated with a steeply dipping sheeted quartz vein complex hosted within granodiorite. The quartz veins strike at 70 to 85 degrees azimuth (mine grid). Mine grid is 29.67 degrees west of true north. The veins are near vertical and are developed in zones of fracturing perpendicular to the hanging wall contact with the overlying supracrustal rocks. The hanging wall contact dips to the north at 30 to 35 degrees and is interpreted as a thrust fault. The veins occur preferentially in the granodiorite and metavolcanic rocks, and terminate at the hanging wall fault structure and gradually wane and die out or “keel” with depth away from the contact. Photo 7.2 shows typical veins in the north wall of the pit, terminating at the dipping structural contact.

Photo 7.2 Typical Veining Terminating At The Dipping Structural Contact

In general, the gold bearing quartz veins are 0.1 to 2.0 metre thick. Photo 7.3 shows a typical vein structure in the face of a cross cut drift. The quartz veins mineralogy is simple being dominated by quartz with minor amounts of sulphides, Bi-tellurides (e.g. tsumoite), tourmaline and scheelite. Tourmaline is a common constituent of all quartz veins and appears to be more abundant in veins with enhanced gold content. Biotite, calcite and actinolite are also present. The sulphides are dominated by pyrite with lesser pyrrhotite and trace chalcopyrite. The distribution of pyrite is somewhat erratic, large, euhedral crystals which occur in the interior of the quartz veins, whereas an intense fine-grained impregnation commonly characterizes the wall-rock contact. At the wall-rock contacts, a weak alteration zone is developed, up to 30 cm wide. The ore body has been drilled to a depth of approximately 400 m below surface and remains open along strike and in both up and down-dip directions.

Genetic theories suggest that the rheology contrast between the more brittle granodiorite and metavolcanic rocks and the overlying supracrustal rocks during structural deformation in particular, movement along the contact resulted in fracturing in the underlying rocks and the emplacement of the sheeted quartz vein complex.


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Photo 7.3 Typical Vein Structures in face of cross cut drift


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8. DEPOSIT TYPE

The Björkdal ore body has the characteristics of an intrusion-related gold system (IRGS). This deposit classification is relatively new, having been proposed by Thompson et al. (1999). The most distinctive style of gold mineralization is sheeted arrays of mineralized parallel quartz veins preferentially located at the top of an igneous intrusion.

Studies related to IRGS deposits have been concentrated on well documented occurrences in Alaska and the Yukon. Characteristics have been defined by Lang and Baker (2001) and consist of:

1. metaluminous, subalkalic intrusion of intermediate to felsic compositions that lie near the boundary between ilmenite and magnetite series;

2. carbonic hydrothermal fluids;

3. a metal assemblage that variably combines gold with elevated Bi, W, As, Mo, Te, and/or Sb and low concentrations of base metals;

4. a low sulphide mineral content, mostly <5 vol%, with a reduced ore mineral assemblage that typically comprises arsenopyrite,pyrrhotite and pyrite and lacks magnetite or hematite;

5. a really restricted, commonly weak hydrothermal alteration;

6. a tectonic setting well inboard of inferred or recognized convergent plate boundaries;

7. a location in magmatic provinces best or formerly known for tungsten and/or tin deposits

The Björkdal deposit exhibits almost all of the diagnostic characteristics of IRGS’s with the notable exception being a lack of arsenopyrite, which is environmentally favourable. The Skelleftea district is not known for tungsten or tin deposits but this may be due to a lack of exploration and discovery for these types of deposits.

Since there is very little outcrop in the Skellefte district and the Björkdal ore body is not known to have a distinct geophysical signature, it was discovered by systematic overburden till sampling. Once located the ore body has been explored by systematic drilling along the hanging wall contact and the top of the granodiorite intrusion. Elgin has not explored outside the immediate mine area but future exploration could be guided by assaying both till samples and drill core for pathfinder elements (in particular Bi, W, Te, and Sb) along with gold.


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9. EXPLORATION

Elgin has not undertaken any exploration work other than drilling directly related to the ore body.


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10. DRILLING

During the period between 1986 and 2004, a total of 1,148 holes were completed at the Björkdal Mine by previous operators. Table 8.1 lists the number of holes for each of; diamond drill, (“DDH”) direct circulation (“DC”) and reverse circulation (“RC”) drilling. During that same period, 6,110 DC grade control holes were drilled. Problems were identified with down-hole grade contamination, particularly in DC holes. Section 14.3 of this report details how these problems were dealt with and which holes were rejected from any resource estimates.

Table 10.1 Historical Drill-hole Summary

Drill-hole Types Number of Holes

DC 343

DDH 128

RC 677

Sub Total 1,148

Grade control 6,110

Total 7,258

In March, 2006, Gold-Ore commenced physical work activities at Björkdal by collaring a portal for the Eastern Tunnel. The tunnel was designed to provide access for diamond drill rigs to test for the extension of the ore body mined in the open pit. Drilling from the surface was a less attractive option due to the presence of a thick and barren supracrustal package of rocks that overlie the host intrusive body. Underground access also provided access for mapping, bulk samples and some feed for the processing plant.

Underground diamond drilling for exploration, development and grade control has been essentially continuous since 2006. A total of 527 diamond drill holes for 63,288 metres and 461 reverse circulation drill holes for 16,744 metres have been drilled on-lease since Gold-Ore became involved at Björkdal in 2006 up to the end of 2012. All diamond drill core is 55 mm in diameter. This drilling data is incorporated into the data base for the resource and reserve estimates. Table 10.2, lists the number of holes and total meterage drilled on the mining lease by year.


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Table 10.2 Drilling Completed by Gold-Ore and Elgin

Year Hole Type Underground Open Pit

# of holes Meterage # of holes Meterage

2006 Core 91 7,954 - -

2007 Core 109 10,454 19 3,303

2008 Core 40 2,577 - -

2009 Core 43 5,892 9 469

2010 Core 30 5,112 37 2,756

RC - - 76 2,978

2011 Core 52 10,271 15 1,325

RC - - 127 3,862

2012 Core 48 8,490.4 34 4,685.1

RC - - 258 9,904

Total 413 50,750 575 29,282.1


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11. SAMPLING PREPARATION ANALYSES AND SECURITY

11.1 Core Logging Protocol

During the logging core process, the geologist lays out the samples for the sampling crew. Samples were typically taken at a nominal one-metre length by Gold-Ore. In the fall of 2012, Elgin introduced sampling procedures with sample lengths adjusted to accommodate lithology changes and quartz veining contacts.

Prior to the fall of 2012, the minimum sample length was 0.6 metres and if a sample contained a vein, a maximum width of 1.1 metres was used. If a vein exceeded 1.1 metres wide, the vein was divided into two samples. A maximum sample length in unmineralized rock was 1.7 metres. In general though, the practice was to sample at continuous 1 metre intervals through areas of interest. In the fall of 2012, it was decided that sampling should be dictated more by lithology, in particular quartz veining and mineralization. Under the revised sampling procedures, sample lengths vary from a minimum of 0.15 metres to a maximum of 1.5 metres.

The division between samples is marked with a red lumber pencil by the logging geologist to guide designated core samplers. Two part sample tags are placed in the box and one part of the tag is placed in the sample bag for shipment and the other half remains in the core box.

Generally, when there is a change in lithology from a known ore-bearing unit to a non-ore-bearing unit, samples are taken one metres past the contact. If there are quartz veins and/or sulphides or visible gold in what is typically a non-ore-bearing unit, samples are taken to confirm that the unit does not have economic mineralization. Generally, one sample will be taken on each side of these samples.

11.2 Sample Collection

11.2.1 Diamond Drill Core

During 2006, a decision was taken to sample “whole core” as opposed to sawing and sampling “half core” for all holes proximal to the production areas. This decision was taken for a number of technical and financial reasons and is the current practice at the Björkdal Mine. Only core from exploration targets distal from the mine are handled with half core techniques. Prior to sampling the drill core, geologists collect geotechnical, lithology, alteration and mineralization data and layout samples.

Gold-Ore had all core data collected to the nearest 0.1 metre (10 cm). The historical Gold-Ore practice was to have the data from the core written on a log format designed by Gold-Ore for the Björkdal Gold Deposit. This data was later typed into spreadsheets and finally transferred manually into the Björkdal Access database. In early 2013, Elgin revised the logging format so that core data was collected to the nearest 0.01 metre (1 cm) and was directly entered into a new logging spread sheet on laptop computers. The new format allowed the digital capture of alteration, quartz vein intervals, mineralization and structural information. During the latter half of 2013, it is planned to have the features of the revised spread sheet incorporated into a commercial logging program that will allow easy direct importing of assay information and subsequent export of the logging and assay information into the access database.


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All core is photographed with a digital camera prior to sawing or sampling. The photos are archived in an electronic folder with the corresponding drill log.

11.2.2 Reverse Circulation Samples

During reverse circulation drilling, samples for logging and assay are collected every metre. Samples for shipment to the ALS Chemex lab are collected in water permeable cloth sample bags commonly referred to as Hubco bags. These samples are dried in the Hubco bag in a large drying oven at the Björkdal on-site core facility prior to shipment to ALS Chemex in Piteå, Sweden. All logging is completed by using an archive chip tray.

11.3 Core Recovery

The geologists determine core recovery by physically measuring the amount of core or length of core between each wooden depth marker in the core box. Core recoveries at the Björkdal Mine are generally considered to be excellent to very good, due to the competent nature of the host rocks.

11.4 Analytical Procedures and Protocol

All core and reverse circulation gold assays are performed by ALS Chemex at Piteå, Sweden. The samples are trucked to the lab and on the return trip, sample reject material is returned to the mine site and archived.

The Chemex lab at Piteå only performs gold assays by cyanide extraction using the “Leachwell” process. The lab does not have fire assay facilities, but can prepare a pulp for shipment elsewhere for fire assay or other analytical procedures, if required. The Leachwell process is basically a bottle roll or bulk leach extractable gold (“BLEG”) procedure that utilizes a catalyst to accelerate the gold leach cycle. This assay technique is often called “Accelerated Cyanide Leach” or ACL.

The ALS Chemex lab has an upper limit of detection of 300.00 g/t gold. All assays that exceed that level are reported as >300.00 gold, but Elgin uses 300.00 g/t as the assay for the purpose of the database. The lower detection limit is 0.01 grams per tonne.

The Leachwell process has been used at the Björkdal Mine for several years and prior to 2006 over 40,000 samples were analyzed using this process and it was chosen for all subsequent drill programs. The advantage of using this process is that a large (500 grams) sample is leached or assayed. The strong nugget effect at the Björkdal Mine is somewhat mitigated by using a 500 gram sample as opposed to a standard one-assay-ton fire assay. In addition, assay results are generally available within seven days of sample delivery to Piteå.

11.5 Quality Assurance/Quality Control Program (“QA/QC”)

The Björkdal Mine employs a rigorous QA/QC program for all of the drill campaigns. The QA/QC involves the insertion of certified standards and blind blanks into the sample stream.


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11.5.1 Certified Standards

All of the standards used during the drill programs (both Reverse Circulation and Diamond Drill ) at the Björkdal Mine were manufactured by Rocklabs Ltd/Scott Technology Limited of New Zealand. The Leachwell analytical process utilized by ALS Chemex has demonstrated good accuracy and precision with the Rocklabs standards. A summary of the rock standards used is given in Table 11.1.

Table 11.1 Summary of Certified Standards Used

Standard Name Grade Au (g/t) Manufacturer Year of Use

OxF65 0.805 Rocklabs Ltd. 2009-2010

OxJ64 2.366 Rocklabs Ltd. 2009-2010

OxL63 5.865 Rocklabs Ltd. 2009-2011

OxH37 1.286 Rocklabs Ltd. 2010-2012

OxK69 3.583 Rocklabs Ltd. 2010-2012

OxL93 5.841 Rocklabs Ltd. 2012-2013

OxG84 0.922 Rocklabs Ltd. 2012-2013

The standards are purchased in 2.5 kg plastic bottles and approximately 80 grams of standard are weighed out and placed in small plastic sealed bags for shipment with the core samples to ALS Chemex. Gold-Ore inserted standards in the sample stream at the rate of one standard per drill hole. In early 2013, Elgin instigated a practice of one blank and one standard for each RC hole and alternating the insertion of blanks and standards for every 20 samples for diamond drilling. The results for standards have been manually monitored for results greater than 3 standard deviations of the recommended gold concentration. For these values, the lab is contacted the data reviewed and if required reassaying carried out. Elgin is in the process of streamlining this process with logging and sampling software with built in QA/QC that will integrate logging and sampling with the mine database.

11.5.2 Blind Blank Insertion

Blanks are inserted into the drillhole sample stream (both RC and DDH ). In addition for diamond drilling, blanks are added after samples with noted visible gold. The blanks are analyzed using the same cyanide-based Leachwell assay as the core samples.

The blank material is obtained from a dimension stone outlet (Granitti Natursten AB of Piteå, Sweden) that sourced the rock from a granite quarry in Finland.


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A threshold of 0.10 ppm gold or greater has been applied as a level to identify any blanks with problematic gold values. There has been some minor cross-contamination during sample preparation, but in 93% of the cases this has occurred after high-grade samples were prepared. It was for this reason Elgin has instigated the placement of blanks after every noted occurrence of visible gold. A summary of blank samples entered into the drilling database is shown in Figure 11.1. The small number of samples affected by cross-contamination by high-grade samples does not cause a bias for the general sample population.

Figure 11.1 Summary of Blank Assaying for Drillhole Samples by Year


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11.5.3 Duplicates

Since whole core for diamond drilling and complete samples for reverse circulation drilling are submitted there are no duplicate samples.

11.5.4 Chip Sampling

From 2008 onwards following the commencement of underground mining operations at Björkdal underground chip sampling has been carried out for production grade control. Chip samples are taken using hammer and chisel as horizontal channels across a heading. The start and end points of each sample are surveyed by Björkdal surveyors using a Trimble Total Station instrument.

Chip samples are generally sampled based on lithology i.e chip samples can be comprised entirely of quartz vein material only. Prior to the change in core sampling procedures (Section 11.2.1) by Elgin, drill holes were generally sampled based on a minimum sample length of 1m and as such can comprise both quartz vein and wall rock material within the same sample. In these cases the chip samples and the drill hole samples are therefore not directly comparable samples. The underground chip samples are assayed at the Björkdal laboratory by fire assay with duplicate samples run independently and an average value returned.

Coarse duplicate samples have been taken of the underground chip samples to compare the results of assaying by fire assay at the Björkdal laboratory and by Leachwell assaying at ALS Chemex (Piteå). A total of 289 coarse duplicate samples were taken and submitted for analysis. The duplicate samples submitted to the Björkdal laboratory were subsequently halved and each sample submitted for separate analysis by fire assay to and returned two independent results (MILL_1 and MILL_2). The average of the two results was then taken and recorded (MILL_AVE). Because of the presence of coarse gold a direct comparison of the duplicate assay results on a sample by sample basis has not been carried out, however the distribution of the sample grades is compared by a log probability plot in Figure 11.2. Based on this, the ALS leachwell and Björkdal fire assay results exhibit very similar grade distributions, suggesting no obvious bias is evident between the assaying at the two laboratories.


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Figure 11.2 Log Probability Plot Comparing Duplicate Underground Chip Samples by Fire Assay at the Björkdal Laboratory (MILL_1, MILL2 and MILL_AVE) and Leachwell Assay at the ALS Chemex Laboratory (Piteå) (Ellis and Newall 2013).

11.6 Sample Security

All drill core and reverse circulation drill samples are processed in a dedicated coring logging facility at the mine site. The facility is very well equipped with roller logging tables, a drying furnace, core cutting saw, and separate, ventilated crushing room.

Every aspect of the core cutting, sampling, bagging and shipping are handled by Björkdal trained personnel and is a high level of professionalism. Samples are placed in heavy reusable nylon bags and sealed with zap straps. Samples are then placed in wooden crates and trucked to the commercial laboratory in Piteå.

Assaying is completed by ALS Chemex, a reputable commercial lab that operates to industry standards. Security of the samples is entirely within industry standards.

A completely separate laboratory facility handles all assaying for the gold production. This lab is located in the processing plant building on the mine site. The lab is operated by Björkdal personnel trained for assaying standards.

In the authors’ opinion, the sample preparation, sample security and analytical procedures are within industry standards and are completely adequate for the Björkdal operation.


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12. DATA VERIFICATION

Gordon Clarke, one of the authours of this report received and reviewed analytical data collected by Elgin. Historical data verification was carried out previously by WAI (Newall and Wheeler 2005). Prior to Elgin becoming the operator data verification was carried out by Robert Wasylyshyn, President of Gold –Ore.

Under the supervision of Mr. Clarke assay results were transferred directly from ALS certified assay reports into the Björkdal database. Results for inserted standards were checked. Results of blanks are monitored and a threshold of 0.1 ppm gold or greater is used to identify any problematic gold values. When identified it is generally a consequence of high gold values in a preceding sample. Elgin has instigated a policy of blank insertion after all noted occurrences of visible gold to avoid possible contamination of subsequent samples.

While the previous and current practices of manual transfer from assay reports into the database has been considered adequate, Elgin is instigating a new automated procedure for importing drilling and assay information into the database. This will be carried by the implementation of GeoSparkTM software that has been modified specifically for the Björkdal mine.

The database provided to WAI for resource estimates was checked for validation errors and was used for Elgin’s revised open pit resource estimate and pit and underground reserve estimates.


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13. MINERAL PROCESSING AND METALLURGICAL TESTING

Mineral processing is described in section 17 of this report. There have been no comprehensive metallurgical testing studies carried out since a pilot plant test work report prepared by Davy McKee in 1987 for Terra Mining AB. The results of this study were used to design the existing plant and the report is located in the Björkdal mine archives.


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14. MINERAL RESOURCE ESTIMATES

Following on from previous modeling studies undertaken by Adam Wheeler in 2004/2005, the drill hole sample database was updated to incorporate drilling information up to April 23, 2013. As with the previous study, the same drill holes identified as having data validation issues were rejected and various flag values were set which were subsequently used in the resource classification. Previous studies used a multiple-indicator kriging (“MIK”) approach to produce a block model that contains ore proportions and associated grades. This methodology was selected owing to the complexity and frequency of quartz veins while quickly allowing for a range of cut off grades to be considered. The current resource estimation adopted a dynamic anisotropy approach to modeling of the vein system and used ordinary kriging (“OK”) for grade estimation.

Initial resource estimation by WAI was classified in accordance with the Australian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves [JORC Code (2004)]. Criteria for defining resource categories were derived from the geostatistical studies and were based upon the search volumes used during grade estimation. There are no material differences between JORC Code 2004 and CIM definition standards on mineral resources and reserves.

• Measured resources were classified corresponding to those blocks encountered during the first search of the grade estimation. Measured resources were based on the drill hole data only (DDH and RC) and required 3 drill holes. A 17.5m x 17.5m drill hole spacing was used for all domains with the exception of the Lakezone domain where 12.5m x 12.5m drilling spacing was required for the allocation of measured resources.

• Indicated resources were classified by use of wireframes based upon those blocks generally encountered during the second search of the grade estimation. Indicated resources were based upon the drill hole data (DDH and RC) and the chip sample data. A 35m x 35m sample spacing was used for all domains with the exception of the Lakezone domain where a 25m x 25m sample spacing was required for the allocation of indicated resources. The majority of indicated resources within the open pit area are therefore located within the areas of drilling based on a 30m x 30m drilling grid. Indicated resources located within the underground area are defined as the areas directly surrounding the underground development and including areas of the Lakezone domain.

• Inferred resources were classified as all remaining blocks based on a maximum sample spacing of 60m. Inferred resources were based on the drill hole data (DDH and RC) and chip sample data. Inferred resources are generally located in the most peripheral areas of the deposit, however, a small area located with the central part of the open pit has also been classified as inferred resources and was historically drilled using DC drilling. DC drilling has been rejected from this study due to downhole contamination issues associated with this drilling method. Drilling by DDH or RC drilling will be required to upgrade this area into indicated resources.

Elgin revised the WAI open pit resource estimate in accordance with an internal policy of reporting resources within an optimized shell


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14.1 Topographical Survey

Topography utilized for both open pit and underground was produced by Björkdal staff (Figure 14.1). It was made available as wireframes in AutoCad format and was up to date as of January 1st 2013.

Figure 14.1 Topographical Survey of Björkdal Mine as of January 1st 2013 also Showing Extent of Underground Development (orange) (Ellis and Newall 2013)

14.2 Database Compilation

Databases used comprised a drill hole database and underground chip sample database in Microsoft Access format and titled:

• Db_may_17_2013_only_UG.mdb, and;

• ProdBjörkdal_20121214.mdb.

The drill hole database included all drill hole assays up to the April 23, 2013 and the underground chip sample database included all chip assays up to 12th December 2012.


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14.3 Drilling

Elgin provided a drill hole database to WAI in Access format. Three main drilling methods have been carried out at Björkdal from 1986 onwards and include diamond drilling (DDH), direct circulation (DC) and reverse circulation (RC). There are known issues associated with the Björkdal drill hole database and are as summarized:

• Previous studies have identified a clear problem with downhole grade contamination (smearing of grades) from all DC drill holes. All DC drill holes have been removed from the drill hole database for the purposes of resource modeling.

• All historical (pre-2006) grade control drill holes (GC) have been removed from the drill hole database as a combination of RC and DC drilling was used to drill these holes and the drilling method is not recorded in the drill hole database. No historical grade control holes have therefore been used for the purposes of resource modeling;

• Previous studies highlighted that a number of historical DDH drill holes have known contamination and reliability issues. These drill holes have been removed from the drill hole database for the purposes of resource modeling. A list of the rejected drill holes is shown in Table 14.1.

It has been assumed that drill hole sample intervals, in which assay values are absent, were not sampled as they were not considered to contain mineralization. Assay values which are absent have therefore been replaced with zero values. Similarly, any assay values recorded in the drill hole database as <0 have been replaced with a zero assay value. Assay values greater than 0.3g/t Au which have the same grade as the preceding assay have been assigned a zero assay value as these values are the result of transcription errors in the database. Historical drill holes 87D003, 87D005, 87D007 and 93D186D were also rejected due to errors in the recorded z-elevation of the drill hole collar. The location of the final imported drill holes is shown in Figure 14.2 and Figure 14.3.

Table 14.1 Rejected Drill holes Due to Contamination and Reliability Issues

86042 86301 86343 87217 87257 88007 88744 93016 89ST513 94D006

86067 86302 86344 87218 87258 88008 88745 93026 89ST514 96D801

86077 86303 86345 87219 87259 88009 88746 93027 89ST515 96D802

86201 86304 87060 87220 87260 88201 88747 95046 89ST516 96D803

86202 86305 87061 87221 87261 88202 88748 86209A 89ST517 96D804

86203 86306 87062 87222 87262 88203 88749 86230X 89ST518 98D009

86204 86307 87063 87223 87263 88204 88750 86230Y 92K0011 DH03003


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86205 86308 87088 87224 87264 88205 88751 87243A 92K0012 DH03007

86206 86309 87100 87225 87265 88206 88752 87ST520 92K0013 DH03014

86207 86310 87101 87226 87266 88207 89023 87ST521 92K0014 DH03021

86208 86311 87102 87227 87267 88208 89024 87ST522 92K0015 DH03023

86209 86312 87103 87228 87268 88209 89025 87ST523 92K0016 DH03027

86210 86313 87104 87229 87269 88210 90001 87ST524 85D003 DH03029

86211 86314 87111 87230 87270 88211 90002 87ST525 87D003 DH03031

86212 86315 87114 87231 87271 88212 90003 87ST526 87D005 DH03032

86213 86316 87115 87232 87272 88213 90004 88ST501 87D007 DH03033

86215 86317 87116 87233 87273 88214 90005 88ST502 87D040 87007

86216 86318 87117 87234 87274 88215 90006 88ST503 87D041 87034

86217 86319 87118 87235 87275 88216 90007 88ST504 87D042

86218 86320 87119 87236 87277 88217 90008 88ST505 87D043

86219 86321 87120 87237 87301 88701 90009 88ST506 87D214

86220 86322 87121 87238 87302 88702 90010 88ST507 90D001

86221 86323 87122 87239 87303 88704 90011 88ST508 90D002

86222 86324 87123 87240 87305 88721 90012 88ST509 90D005

86223 86325 87124 87241 87306 88722 90013 88ST726 90D019

86224 86326 87201 87242 87307 88723 90014 88ST727 90D020

86225 86329 87202 87243 87308 88724 90015 88ST728 90D021

86226 86330 87203 87244 87309 88725 90016 88ST729 90D022

86227 86331 87204 87245 87310 88732 90018 88ST730 90D106

86228 86332 87205 87246 87311 88733 90019 88ST731 92D001


36

86229 86333 87206 87247 87312 88734 90020 89ST501 92D002

86230 86334 87207 87248 87313 88735 91008 89ST502 92D003

86231 86335 87208 87249 87314 88736 91056 89ST504 92D004

86232 86336 87209 87250 87315 88737 91058 89ST505 92D005

86233 86337 87210 87251 88001 88738 91059 89ST507 92D011

86234 86338 87211 87252 88002 88739 91060 89ST508 93D001

86235 86339 87212 87253 88003 88740 91068 89ST509 93D1240

86236 86340 87213 87254 88004 88741 91120 89ST510 94D001

86237 86341 87214 87255 88005 88742 92035 89ST511 94D004

86238 86342 87216 87256 88006 88743 92075 89ST512 94D005

A summary of the drill hole database used for the resource calculations following data validation is shown in Table 14.2.

Table14.2 Summary of Björkdal Drill hole Database

Drilling Type Number of Drill Holes Number of Samples

DDH 669 82,072

RC 1,576 114,364

TOTAL 2,245 196,436

The location of the drill holes is shown in Figures 14.2 and 14.3.


37

Figure 14.2 Location of DDH Drill holes at Björkdal (Ellis and Newall 2013)

Figure 14.3 Location of RC Drill holes at Björkdal (Ellis and Newall 2013)


38

14.4 Chip Sampling

Underground chip sampling has been carried out at Björkdal in the area of underground development from 2008 onwards. Horizontal chip samples of approximately 1m are taken across the width of the heading and comprise the full width of the vein. The location of the chip samples is shown in Figure 14.4.

Figure 14.4 Location of Underground Chip Samples at Björkdal (Ellis and Newall 2013)

Additional data processing associated with the Björkdal chip sample database has been carried out as summarised:

• Chip samples are taken for production planning purposes and one sample may only be taken for the quartz vein only portion of a heading. It was therefore decided to create dummy samples either side of each existing mineralized chip sample to replicate the ‘off vein’ footwall and hanging wall. Each dummy sample was assigned a zero grade and assumes that no mineralization is carried in the footwall or hanging wall.

• The chip samples are generally of around 1m in length; however, there are increasingly becoming chip samples of smaller more variable lengths being taken. To reduce the influence of these smaller samples, it was decided to standardize the minimum sample length to 1m. Therefore, all samples less than 1m in length were diluted up to 1m in length. Again, the dilution adjustment assumed a zero grade for the footwall and hanging wall.


39

• The chip sample orientation is highly variable and can comprise both across vein and along vein samples. It was decided to restrict the chip sample database to across vein samples only to limit any bias that may be introduced by sampling along the strike of the vein. All chip samples with azimuths >60⁰ and <120⁰ and all chip samples with azimuths >240⁰ and <300⁰ were removed from the chip sample database for the purposes of resource modeling. All indicated azimuths are relative to mine grid.

A summary of the chip sample database following validation is shown in Table 14.3.

Table 14.3 Summary of Björkdal Chip Sample Database

Number of Chip Samples with Assay

TOTAL 5,839

14.5 Geological Interpretation

Based on current geological and operational understanding of the Björkdal mineralization, a number of important conclusions with respect to the geological interpretation at Björkdal have been taken into account including:

The key geological controls applied to block modeling should be the lithological contacts, particularly the granodiorite/marble contact. The marble and cherty unit can be considered as barren. The Björkdal deposit is dominated by relatively thin sheeted quartz veins within a non-mineralized matrix.

The veins are generally sub-vertical, and on average strike at roughly 072° (mine grid). However, there are variations in this orientation within different areas of the deposit.

A wireframe depicting the extent of the mineralization at Björkdal was constructed from drill hole lithology data and grade data. The wireframe depicted the extent of the mineralized grano-diorite unit. The overlying marble unit is considered as barren while all material contained within the granodiorite mineralized envelope is considered to potentially contain mineralization. An additional low angle thrust limestone unit was also included in the northern part of the deposit and all material contained within this unit was also considered as barren. Additional domaining of the deposit has been carried out by mine area based on structural and grade domains defined by the operations staff at the Björkdal Mine. The location of the domains is shown in Figure 14.5 and summarized in Table 14.4


40

Figure 14.5 Location of Björkdal Domains (Ellis and Newall 2013)

Table 14.4 Summary of Björkdal Domains

Mine Area Domain Domain Code

Open Pit Central Pit 11

Open Pit South Wall and Pit 3 12

Open Pit South East Extension 13

Open Pit Nylund 14

Underground Main Zone 21

Underground Lakezone 22

14.6 Sample Data Processing

14.6.1 Statistical Analysis

Statistical analysis has been carried out by domain and by sampling method on the raw sample data (i.e without top cutting) contained within the mineralized envelope to identify any potential bias within the dataset.


41

14.6.2 Domain

Log probability plots showing Au grade for all samples by domain are shown in Figure 14.6 and statistical analysis is shown in Table 14.5. The open pit domains of Central Pit (domain 11), Southwall and Pit 3 (domain 12) and Southeast Extension (domain 13) exhibit very similar grade distributions and mean grades while the remaining open pit domain of Nylund (domain 14) exhibits a significantly lower grade population. The underground domains of Main Zone (domain 21) and Lakezone (domain 22) both exhibit higher grade populations to that of the open pit domains. The Lakezone domain exhibits a slightly different population distribution whereby fewer extreme grades are seen compared to other domains but a higher percentage of the assaying appears to be within the 1g/t Au to 10g/t Au range.

Figure 14.6 Log Probability Plot of all Samples by Domain (Ellis and Newall 2013)

Table 14.5 Statistical Analysis of Raw Samples >0.3g/t by Domain

DOMAIN

11 12 13 14 21 22

FIELD AU AU AU AU AU AU

NSAMPLES 3,631 4,583 3,106 649 11,213 667

MINIMUM 0.3 0.3 0.3 0.3 0.3 0.3

MAXIMUM 396.3 630.0 764.0 80.8 3290.0 277.0

RANGE 396.0 629.7 763.7 80.5 3289.7 276.7

MEAN 3.78 3.73 3.74 1.87 7.37 5.14


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DOMAIN

11 12 13 14 21 22

VARIANCE 248.00 367.27 453.41 32.38 2,105.80 272.63

STANDDEV 15.75 19.16 21.29 5.69 45.89 16.51

SKEWNESS 14.84 18.74 22.30 8.64 41.26 9.86

KURTOSIS 281.50 462.06 665.25 91.42 2,541.72 129.12

14.6.3 Sampling Method

Log probability plots showing Au grade for all samples by sampling method for each domain are shown in Figures 14.7 to 14.7f and statistical analysis of these is shown in Table 14.6. Where similar numbers of samples are present, the DDH and RC data exhibit very similar grade distributions. In the Main Zone (domain 21), the RC drilling appears to understate the DDH slightly; however, this is a consequence of the RC drilling being limited only to the upper levels of the Main Zone only by surface drilling through the open pit wall where-as the majority of the DDH drilling comes from underground drilling. The chip sampling in the Main Zone (domain 21) exhibits a higher grade distribution to both the DDH and RC and is a result of the difference in the sampling methodology between the chip samples and the drill hole samples. Chip samples are currently sampled based on lithology (i.e chip samples can be comprised entirely of quartz vein material only where-as drill holes prior to the fall of 2012 were generally sampled based on a minimum sample length of 1m and as such could comprise both quartz vein and wall rock material within the same sample). As a consequence of the difference in sampling methodology, the chip samples exhibit a higher grade sample distribution compared to both the DDH and RC drill holes. For the purposes of resource estimation it was decided to combine the drill hole database and the chip sample database and to test the validity of this by reconciliation.

Table 14.6 Statistical Analysis of Raw Samples >0.3g/t by Domain and Sampling Method

DOMAIN 11 11 12 12 13 13 14 14 21 21 21 22

DHTPE DDH RC DDH RC DDH RC DDH RC DDH RC CHIP DDH

FIELD AU AU AU AU AU AU AU AU AU AU AU AU

NSAMPLES 564 3,067 152 4,431 226 2,880 67 582 5,472 2,449 3,292 667

MINIMUM 0.3 0.3 0.3 0.3 0.3 0.3 0.3 0.3 0.3 0.3 0.3 0.3

MAXIMUM 396.3 321.9 630.0 483.6 764.0 217.4 65.2 80.8 300.0 1,602.4 3,290.0 277.0

RANGE 396.0 321.6 629.7 483.3 763.7 217.1 64.9 80.5 299.7 1,602.1 3,289.7 276.7

MEAN 5.20 3.52 14.56 3.35 9.41 3.29 2.93 1.75 5.51 4.95 12.28 5.14

VARIANCE 602.66 182.33 3,539.41 254.29 4,120.17 162.95 76.74 27.13 429.57 1,789.70 5,093.01 272.63

STANDDEV 24.55 13.50 59.49 15.95 64.19 12.77 8.76 5.21 20.73 42.30 71.37 16.51

SKEWNESS 11.68 14.82 8.12 18.27 9.85 10.84 5.93 9.19 9.53 27.98 32.47 9.86

KURTOSIS 158.03 292.25 75.43 440.21 101.84 141.89 36.98 108.30 110.36 926.27 1,383.77 129.12


43

Figure 14.7a Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 11 and by Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013)

Figure 14.7b Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 12 and by

Sampling Method (DHTYPEN1 = DDH, DHTYPEN2 = RC, DHTYPEN3 = CHIP) (Ellis and Newall 2013)


44

Figure 14.7c Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 13 and by


Figure 14.7d Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 14 and by



45

Figure 14.7e Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 21 and by


Figure 14.7f Log Probability Plot for Au by Sampling Method for >0.3g/t Au for Domain 22 and by



46

14.6.4 Removal of Outlier Grades

Analysis to identify outlier values within the dataset was carried out using log probability plots, decile analysis and contiguous analysis. Based on this, a summary of the top-cut levels used is set out in Table 14.7 and the log probability plots are shown in Figures 14.8a to 14.8f. A summary of the database following top-cutting is shown in Table 14.8.

Table 14.7 Summary of Top-Cut Values

Mine Area Domain Domain Code Top-Cut (Au g/t)

Open Pit Central Pit 11 55

Open Pit South Wall and Pit 3 12 55

Open Pit South East Extension 13 50

Open Pit Nylund 14 15

Underground Main Zone 21 70

Underground Lakezone 22 20

Table 14.8 Statistical Analysis of Top-Cut Samples >0.3g/t by Domain

DOMAIN

11 12 13 14 21 22


NSAMPLES 3,631 4,583 3,106 649 11,213 667

MINIMUM 0.3 0.3 0.3 0.3 0.3 0.3

MAXIMUM 55 55 50 15 70 20

RANGE 54.7 54.7 49.7 14.7 69.7 19.7

MEAN 3.13 2.88 2.72 1.47 5.29 3.48

VARIANCE 54.88 52.57 45.10 6.48 145.49 26.67

STANDDEV 7.41 7.25 6.72 2.54 12.06 5.16

SKEWNESS 4.88 5.47 5.25 4.06 3.99 2.20

KURTOSIS 26.88 32.78 30.39 17.09 16.61 3.80


47

Figure 14.8a Log Probability Plot Showing Top Cut Level For Au For Domain 11 (Ellis and Newall 2013)

Figure 14.8b Log Probability Plot Showing Top Cut Level For Au For Domain 12 (Ellis and Newall

2013)


48

Figure 14.8c Log Probability Plot Showing Top Cut Level For Au For Domain 13 (Ellis and Newall

2013)

Figure 14.8d Log Probability Plot Showing Top Cut Level For Au For Domain 14 (Ellis and Newall

2013)


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Figure 14.8e Log Probability Plot Showing Top Cut Level For Au For Domain 21 (Ellis and Newall

2013)

Figure 14.8f Log Probability Plot Showing Top Cut Level For Au For Domain 22 (Ellis and Newall

2013)


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14.6.5 Composting

Standard histograms displaying the variation in sample length for all mineralised (>0.30g/t Au) samples from both drill hole and chip samples are shown in Figure 14.9. A 1m composite sample length has been selected to give a consistent level of support for geostatistical analysis and this has been applied to all of the drill holes. Each chip sample has been treated as a separate drill hole; therefore, no compositing of the chip samples has been carried out. However, given that the majority of chip samples are generally around 1m in length, and any smaller samples have been diluted up to 1m in length, this is considered to be acceptable. An indicator was then applied to the full length composite drill holes and chip samples to define the mineralization based on a cut-off grade of 0.30g/t Au. The composites were assigned an indicator field (IND) code of 1 for all composite samples greater than 0.30g/t Au and an IND code of 0 for all composite samples less than 0.30g/t Au.

Figure 14.9 Standard Histogram Showing Sample Length for All Mineralized Samples (Ellis and Newall 2013)

A summary statistical analysis of all composites >0.30g/t Au is shown in Table 14.9.


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Table 14.9 Statistical Analysis of Composites >0.3g/t

DOMAIN

11 12 13 14 21 22


NSAMPLES 3,576 4,592 3,127 666 11,890 834

MINIMUM 0.3 0.3 0.3 0.3 0.3 0.3

MAXIMUM 55 55 50 15 70 20

RANGE 54.7 54.7 49.7 14.7 69.7 19.7

MEAN 3.07 2.87 2.70 1.42 4.76 2.73

VARIANCE 7.24 7.22 6.62 2.43 10.86 4.15

STANDDEV 4.95 5.48 5.28 4.12 4.23 2.64

SKEWNESS 27.90 32.89 30.81 17.95 19.37 6.62

KURTOSIS AU AU AU AU AU AU

14.7 Variography

Variography was undertaken:

to estimate the presence of anisotropy in the deposit;

to derive the spatial continuity of mineralisation along the principal main anisotropic orientations;

to produce suitable variogram model parameters for use in geostatistical grade interpolation; and

to assist in selection of suitable search parameters upon which to base the resource estimation.

Variography was performed using Datamine Studio v3 software. Pairwise relative variograms were generated, with the spherical scheme model being used for modelling purposes. Variography was carried out by individual domain using the 1m drill hole composite data contained within that domain.


52

14.7.1 Variogram Parameters

Directional and down hole semi-variograms were carried out for indicator values above 0.3g/t Au values. In keeping with the quartz vein orientation strings, the semi-variograms were carried out in the general orientation of the quartz veins for each domain. Along strike semi-variograms were variably orientated at approximately 085/00; down-dip variograms were variably orientated at approximately 355/85; and cross strike variograms were variably orientated at approximately 175/05. The nugget variances were modeled from average downhole variograms based on a 1m lag reflecting the downhole drill hole composite spacing. The final modeled semi-variograms for each domain are contained in Appendix 1.

14.7.2 Variography Interpretation

The principal direction of continuity was selected from the generated experimental semi-variograms and modelled with three-structure spherical models. The three orthogonal orientations representing the predominant along-strike, down-dip and cross-strike directions were defined. Semi-variograms for domains 11, 12, 13, 14 and 21 indicated along strike and down dip ranges of approximately 30m to 35m, although, by around 15m over 2/3rds of the total variability (sill) has been reached. Semi-variograms for domain 22 indicated along strike and down dip ranges of approximately 25m and reflects the dominant drill hole spacing in this area. All semi-variograms indicate high nugget values of between 45% and 65%.

14.8 Block Modelling

The complexity and frequency of quartz veins makes it difficult to interpret and model the quartz veins individually. In addition to this, the large numbers of very thin veins (that can still contain appreciable ore grades) mean that model blocks have to be generated so that they include these veins. In previous studies, multiple-indicator modeling approaches (MIK) have been applied and are considered to be a logical approach given the complications owing to the complexity and frequency of the quartz veins. The MIK method produces ore proportions and associated grades within an individual block. The major limitation of the MIK modeling methodology is that it cannot be used for underground mine design as this method does not define the exact location of a vein which is required for mine design. A dynamic anisotropy method of directly modeling the veins from the grade data and using variable strike and dip orientations was therefore adopted.

Directional control strings were used to define the local variation in the strike and dip of the deposit. These orientation strings were used in the dynamic anisotropy procedure where a search of 60m x 60m (along strike x down dip) was used based on the indicator field (IND). The true dip and dip direction of each block was coded into the model to be used for controlling the search orientation used in the grade estimation. A parent cell size of 10m x 5m x 10m (along strike x across strike x vertical) was used with sub celling down to 5m x 1m x 5m within the mineralized zones and was used to reflect the general across strike sample thickness, the 5m bench height used and to accommodate the higher density of drill holes in the cross-strike direction. A minimum sub cell dimension of 2.5m in the along strike and vertical direction was also used. The base of the model was extended down to a depth of 500m and is considered the current drill base extent. The model has been rotated to 252⁰ to align with the predominant strike of the veins. A summary of the model prototype parameters used is shown in Table 14.10.


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Table 14.10 Summary of Indicator Block Model Prototype Parameters

Property Direction Metres (m)

Model Origin

X 780

Y -1130

Z -500

Parent Cell Size

X 10

Y 5

Z 10

Mineralized Model Cell Size

X 5

Y 1

Z 5

No. of Cells

X 235

Y 528

Z 51

The block model was coded according to open pit, pillar or underground areas. The pillar has been modelled between the open pit and underground areas. An average pillar thickness of 20m has been used. A sectional view of the block model is shown in Figure 14.10.


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Figure 14.10 Cross Section Showing Extent of Mineralization in Indicator Model (Ellis and Newall 2013)

An explanation of any additional fields added to the block model is shown below in Table 14.11.

Table 14.11 Summary of Block Model Fields

Code Description

ROCK Rock Type (MARB, GRAN or TILL)

ROCKN Rock Type Numeric

MINED 0=Unmined (in situ); 1=Mined

DENSITY Density

TRDIPDIR True Dip Direction

TRDIP True Dip

IND 1=Mineralised, 0=Non-mineralised

AU Au (Ordinary Kriging Interpolation)

Underground

Open Pit

Pillar


55

Code Description

AU_ID Au (Inverse Distance Squared Interpolation)

AU_NN Au (Nearest Neighbour Interpolation)

SVOL Search Volume Number

CLASS Resource classification (1=Measured; 2=Indicated; 3=Inferred)

AREA 1= Open Pit Area; 2=Underground Area; 3= Pillar Area

14.8.1 Density

A density of 2.71t/m3 has been used for the granodiorite and the overlying limestone unit. A density of 1.80 t/m3 has been used for the overlying till. The rock density values applied are consistent with the historical densities used at the Björkdal Mine.

14.9 Grade Estimation

Grade estimation was carried out using Ordinary Kriging (OK) as the principle interpolation method. Inverse distance weighting squared (IDW) and nearest neighbour (NN) interpolations were also carried out for comparative purposes. The OK method used estimation parameters defined by the variography. The estimation was performed only on mineralised vein material defined within the block model.

14.9.1 Kriging Plan

The OK estimation was run in a three pass kriging plan, the second and third passes using progressively larger search radii to enable the estimation of blocks unestimated on the previous pass. The search parameters were derived from the variography, with the first search distances corresponding to the distance at 2/3rds of the semi-variogram sill value and the second search distance approximating the semi-variogram range. Sample weighting during grade estimation was determined by semi-variogram model parameters for the OK method. Block discretization was set to 5 x 3 x 5 to estimate block grades. A summary of the kriging plan is shown in Table 14.12.


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Table 14.12 Björkdal Kriging Plan (OK)

DOMAINS 11, 12, 13, 14 and 21

Search Distance (m) Samples Drill Holes

Search Along Strike

Down-

Dip

Cross-Strike

Minimum Maximum Minimum Type

1st 17.5 17.5 1.5 7 15 3 Only DDH + RC

2nd 35 35 3 4 15 2 DDH+RC +CHIP

3rd 60 60 4.5 1 15 1

DOMAIN 22 (Lakezone)

Search Distance (m) Samples Drill Holes

Search Along Strike

Down-

Dip

Cross-Strike

Minimum Maximum Minimum Type

1st 12.5 12.5 1.5 7 15 3 Only DDH + RC

2nd 25 25 3 4 15 2 DDH+RC +CHIP

3rd 60 60 4.5 1 15 1

14.9.2 Validation

Following grade estimation, a statistical and visual assessment of the block model was undertaken to 1) assess successful application of the estimation passes; 2) ensure that as far as the data allowed, all blocks within mineralization domains were estimated, and; 3) the model estimates performed as expected. The model validation methods carried out included:

A visual assessment of grade;

Global statistical grade validation; and

SWATH plot (model grade profile) analysis.

Globally, no indications of significant over or under estimation were apparent in the model nor were any obvious interpolation issues identified. From the perspective of conformance of the average model grade to the input data, WAI considers the model to be a satisfactory representation of the sample data used and an indication that the grade interpolation performed as expected. In terms of conformance to the composite data, WAI consider the OK interpolation method to most closely represent the composite data. The resource estimate is based upon the OK grade estimation.


57

As a general comment, the validations only determine whether the grade interpolation has performed as expected. Acceptable validation results do not necessarily mean the model is correct or derived from the right estimation approach. It only means the model is a reasonable representation of the data used and the estimation method applied.

14.10 Selective Mining Units

The block model was subsequently regularized to two selective mining units (SMU’s). The first SMU comprised 5m x 4m x 5m blocks (reflecting a minimum mining width of 4m) and the second SMU comprised 5m x 2.5m x 5m blocks (minimum mining width of 2.5m). Dilution was applied assuming zero grade for all diluting material.

14.11 Depletion

The block models have been depleted using an open pit and underground survey understood to be up to date as of January 1, 2013. The underground survey comprises both development drives and stope surveys. The stope surveys are conducted using a radial laser scanning system, whereby the stopes are surveyed from a single point at the start of the heading. However, because of this there are issues associated with ’shadowing’ during the scanning whereby any overhangs located within the stope prevent a complete survey as the single point scanning system cannot survey behind these obstructions. To ensure correct and full depletion of these incomplete stope surveys, wireframes were manually constructed to encapsulate the furthest points of the stope surveys and these have subsequently been used to deplete the block model. It is recommended that the stope scanning system in use at the Björkdal Mine be updated to allow for complete stope surveying irrespective of any overhangs located within the stope.

14.12 Reconciliation

Reconciliation has been undertaken based on historical and recent production and survey data. Full block models were constructed and involved adjusting for areas of the open pit which have been historically mined out and then subsequently backfilled. For these areas, the tip material which forms the current surface, has been replaced with the in-situ granodiorite which comprised the historical unmined surface. The test reconciliation was completed on the historical recorded open pit production from the Björkdal Mine from 1993 until the end of mining operations in 2000. From 2000 until the start of underground operations in 2008, the majority of production at the Björkdal Mine came from the re-processing of stockpiles rather than open pit mining. Mining operations for the underground and open pit areas commenced in 2008 and 2009, respectively, and annual production records and survey data up until 2013 are also available from this period.


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14.12.1 Open Pit Production June 30, 1993 to February 1, 1998

A mine survey of the open pit at the Björkdal Mine was undertaken at the end of June 1993 and annual production figures for this year have been adjusted to reflect this date. A final mine survey dated February 1, 1998 was also undertaken. The open pit production figures between these dates are shown in Table 14.13 and show that a total of 4,235t of ore at 2.46g/t Au was mined from the open pit during this production period. It is understood that a minimum mining width of 2.5m was used in the open pit during this time. A cut-off grade of around 0.8g/t Au was used and relatively robust grade control procedures were implemented.

Table 14.13 Open Pit Production Data – June 30, 1993 to February 1, 1998

Production Period

Mill Feed

(kt)

Grade Au

(g/t)

Contained Au Metal

(kg)

Contained Au Metal

(oz)

1993 (from 30th June) 382 3.33 1,273 40,925

1994 872 2.63 2,294 73,767

1995 840 2.11 1,772 56,970

1996 877 2.31 2,026 65,151

1997 1,157 2.49 2,880 92,604

1998 (to 1st February) 106 1.70 181 5,810

Total 4,235 2.46 10,427 335,227

Notes: 1. Mill feed direct from run of mine i.e not stockpiles

Blocks contained between the two surveys were extracted from the 2.5m SMU block model. Additional mining factors of 5% mining dilution and 95% mining recovery were subsequently applied to the block model to reflect potential mining factors in the open pit at this time. The results of the block model evaluated above a range of cut-off grades and with these additional mining factors applied are shown in Table 14.14.


59

Table 14.14 Block Model Open Pit Evaluation – June 30, 1993 to February 1, 1998. Measured and Indicated Resources

Measured and Indicated Resources

Cut-Off (g/t)

Tonnage (kt)

Grade Au (g/t)

Contained Au Metal (kg)

Contained Au Metal (oz)

0.0 6,369 1.66 10,592 340,543

0.1 6,342 1.67 10,590 340,470

0.2 6,095 1.73 10,553 339,272

0.3 5,696 1.84 10,458 336,227

0.4 5,247 1.96 10,308 331,421

0.5 4,803 2.11 10,118 325,314

0.6 4,420 2.24 9,919 318,888

0.7 4,078 2.38 9,706 312,071

0.8 3,761 2.52 9,480 304,794

0.9 3,491 2.65 9,262 297,785

1.0 3,246 2.78 9,040 290,628

Notes: 1. 2.5m selectivity model 2. Additional mining factors of 5% mining dilution and 95% mining recovery applied

Based on a 0.8g/t Au cut-off, the block model adjusted for 5% mining dilution and 95% mining recovery, reports 3,761kt of measured and indicated mineral resources at a grade of 2.52g/t Au. This figure can be compared with 4,128kt at a grade of 2.48g/t Au reporting from production. Overall, the reconciliation results compare relatively well with the block model reporting slightly under with 9% less total contained metal than production for the same period.

14.12.2 Open Pit Production February 1, 1998 to July 31, 1999

A mine survey of the open pit at the Björkdal Mine was undertaken on February 1, 1998 and annual production figures for this year have been adjusted to reflect this date. A final mine survey dated July 31, 1999 was also undertaken and again annual production figures for this year have been adjusted to reflect this date. During this period, grade control procedures were reduced in the pit for purposes of cost reduction and consequently a reduction in head grade resulted. The open pit production figures between these dates are shown in Table 14.15 and show that a total of 1,540kt of ore at 1.65g/t Au was mined from the open pit during this production period. It is understood that a minimum mining width of 2.5m and a cut-off grade of around 0.8g/t Au was maintained for the open pit operations during this time.


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Table 14.15 Open Pit Production Data – February 1, 1998 to July 31, 1999

Production Period Mill Feed

(kt) Grade Au

(g/t)


Contained Au Metal

(oz)

1998 (from 1st February) 1,169 1.7 1,988 63,909

1999 (to 31st July) 371 1.5 556 17,870

Total 1,540 1.65 2,544 81,779

Notes: 1. Mill feed direct from run of mine i.e not stockpiles

Blocks contained between the two surveys were extracted from the 2.5m SMU block model. Additional mining factors of 50% mining dilution and 95% mining recovery were subsequently applied to the block model to reflect the increased mining dilution resulting from the reduced grade control. The results of the block model evaluated above a range of cut-off grades and with these additional mining factors applied are shown in Table 14.16.

Table 14.16 Block Model Open Pit Evaluation – February 1, 1998 to July 31, 1999. Measured and Indicated Resources


Cut-Off (g/t)

Tonnage (kt)

Grade Au (g/t)



0.0 2,371 1.06 2,510 80,694

0.1 2,362 1.06 2,509 80,678

0.2 2,258 1.11 2,498 80,315

0.3 2,098 1.18 2,471 79,442

0.4 1,913 1.27 2,428 78,061

0.5 1,756 1.36 2,381 76,546

0.6 1,597 1.45 2,323 74,675

0.7 1,468 1.54 2,267 72,878

0.8 1,355 1.63 2,211 71,070

0.9 1,245 1.73 2,148 69,061

1.0 1,156 1.81 2,092 67,257



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Based on a 0.8g/t Au cut-off, the block model adjusted for 50% mining dilution and 95% mining recovery reports 1,355kt of measured and indicated mineral resources at a grade of 1.63g/t Au. This figure can be compared with 1,540kt at a grade of 1.65g/t Au reporting from production. Overall, the reconciliation results compare moderately well with the block model again reporting under with 13% less total contained metal than production for the same period.

14.12.3 Open Pit Production January 1, 2009 to January 1, 2013

Mine surveys of the open pit at the Björkdal Mine are understood to be up to date as of January 1, 2009 and January 1, 2013. The open pit production figures between these dates are shown in Table 14.17 and show that a total of 2,279kt of ore at a grade of 0.90g/t Au was mined from the open pit during this production period. It is understood that a minimum mining width of 4.0m was used in the open pit during this time with a cut-off grade of approximately 0.4g/t Au.

Table 14.17 Open Pit Production Data – January 1, 2009 to January 1, 2013

Year Mill Feed

(kt) Grade Au

(g/t)



2009 557 0.88 489 15,724

2010 586 0.81 473 15,200

2011 569 0.91 519 16,674

2012 568 1.01 573 18,420

Total 2,279 0.90 2,053 66,018

Notes: 1. Mill feed direct from run of mine i.e not including stockpiles

Blocks contained between the two surveys were extracted from the 4.0m SMU block model. Additional mining factors of 30% mining dilution and 95% mining recovery were subsequently applied to the block model to reflect the potential mining factors in the open pit. The results of the block model evaluated above a range of cut-off grades and with these additional mining factors applied are shown in Table 14.18.


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Table 14.18 Block Model Open Pit Evaluation - – January 1, 2009 to January 1, 2013. Measured and Indicated Resources


Cut-Off (g/t)

Tonnage (kt)

Grade Au (g/t)



0.0 3,165 0.72 2,286 73,494

0.1 3,158 0.72 2,285 73,476

0.2 2,865 0.79 2,250 72,323

0.3 2,392 0.90 2,159 69,409

0.4 2,011 1.02 2,058 66,151

0.5 1,706 1.14 1,952 62,763

0.6 1,487 1.25 1,860 59,811

0.7 1,294 1.36 1,764 56,700

0.8 1,134 1.47 1,671 53,734

0.9 1,000 1.58 1,584 50,917

1.0 877 1.70 1,494 48,031


Based on a 0.4g/t Au cut-off grade, the block model adjusted for 30% mining dilution and 95% mining recovery reports, 2,011kt of measured and indicated mineral resources at a grade of 1.02g/t Au. This figure can be compared with 2,279kt at a grade of 0.90g/t Au reporting from production. Overall, the reconciliation results compare very well with the block model reporting an almost identical figure compared to the production results.

14.12.4 Underground Production 2008 to 2013

An underground mine survey of all underground development and understood to be up to date as of January 1, 2013 was used for the purposes of reconciliation. Stope surveys were also available, however because of issues identified with the quality of the stope surveying this data was not included in the reconciliation exercise. The reconciliation is therefore based on the underground development data only (i.e on vein drives and cross cuts only). A summary of the production figures derived from underground development and not including stopes since 2008 are shown in Table 14.9 and show that a total of 2,079kt of ore at a grade of 1.62g/t Au was mined during this production period.


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Table 14.19 Underground Development Production Data (On-vein drives and cross cuts - not including stopes)

Year Mill Feed

(kt) Grade Au

(g/t) Contained Au Metal

(kg) Contained Au Metal

(oz)

2008 90 2.24 202 6,505

2009 450 1.60 722 23,205

2010 457 1.79 818 26,297

2011 496 1.50 745 23,954

2012 443 1.47 653 21,004

Total 1,937 1.62 3,140 100,964

Notes: 1. Mill feed direct from run of mine i.e not including stockpiles

Blocks contained within the underground development were extracted from the 1m selectivity model as shown in Figure 14.11.

Figure 14.11 Isometric View of Björkdal Underground Development Block Model (Ellis and Newall 2013)


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All blocks without a gold grade were assigned zero gold grades. Total volumes and tonnages for all mined material contained within the block model are very similar (within 2%) to the total tonnages reported in the production figures. For the purposes of reconciliation, it has therefore been assumed that nearly all mined material from the underground development has subsequently been processed by the mill. Based on this, no cut-off grade or additional mining factors have been applied when evaluating the block model. The results of the block model evaluation for the underground development from 2008 to 2013 are shown in Table 14.20. The results are quoted for measured and indicated resources.

Table 14.20 Block Model Evaluation Of Underground Development (On-vein drives and cross cuts only - not including stopeing)

Year Tonnage

(kt) Grade Au

(g/t) Contained Au Metal

(kg) Contained Au Metal

(oz)

Total 2008 to 2013 1,909 1.46 2,794 89,839

Notes:

1. 1m selectivity model diluted to include all underground development

2. No cut-off grade or additional mining factors applied

The block model reports a total of 1,909kt of measured and indicated mineral resources at a grade of 1.46g/t Au. These figures can be compared with 1,937kt at a grade of 1.62g/t Au reporting from the production. Overall, given the assumptions involved, the reconciliation compares relatively well, albeit with the resource model reporting at a lower grade than production.

14.13 Resource Classification

Criteria for defining resource categories were derived from the geostatistical studies and were based upon the search volumes used during grade estimation. Measured resources were classified corresponding to those blocks encountered during the first search of the grade estimation. Measured resources were based on the drill hole data only (DDH and RC) and required 3 drill holes. A 17.5m x 17.5m drill hole spacing was used for all domains with the exception of the Lakezone domain where 12.5m x 12.5m drilling spacing was required for the allocation of measured resources.

Indicated resources were classified by use of wireframes based upon those blocks generally encountered during the second search of the grade estimation. Indicated resources were based upon the drill hole data (DDH and RC) and the chip sample data. A 35m x 35m sample spacing was used for all domains with the exception of the Lakezone domain where a 25m x 25m sample spacing was required for the allocation of indicated resources. The majority of indicated resources within the open pit area are therefore located within the areas of drilling based on a 30m by 30m drilling grid. Indicated resources located within the underground area are defined as the areas directly surrounding the underground development and including areas of the Lakezone domain.


65

Inferred resources were classified as all remaining blocks based on a maximum sample spacing of 60m. Inferred resources were based on the drill hole data (DDH and RC) and chip sample data. Inferred resources are generally located in the most peripheral areas of the deposit, however, a small area located with the central part of the open pit has also been classified as inferred resources and was historically drilled using DC drilling. DC drilling has been rejected from this study due to downhole contamination issues associated with this drilling method. Drilling by DDH or RC drilling will be required to upgrade this area into indicated resources. An example plan view of the block model showing the resource classification is shown in Figure 14.12.

Figure 14.12 Plan View Illustrating Resource Classification (Ellis and Newall 2013)

14.14 Resource Evaluation

A breakdown of the separate resource estimates for the underground and open pit areas are shown in Tables 14.21 and 14.22. Any mined-out resources have been depleted from the resource model (i.e. only remaining in-situ resources are quoted). The resources are also reported exclusive of any resources contained within the current pillar configuration between the open pit and the underground areas. The model has been evaluated at various cut-off grades and selective mining units. No additional mining factors such as (unplanned) mining dilution or mining recovery have been applied to the resource estimates.

The stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues, to the best knowledge of the authors. There are no known mining, metallurgical, infrastructure, or other factors that materially affect this mineral resource estimate, at this time.


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Table 14.21 Björkdal Underground Resource Estimate (Ellis and Newall 2013)

Minimum Mining Width 1.0m 1.0m 2.5m 2.5m 4m 4m

Cut Off Grade (g/t) 0.3 0.6 0.3 0.6 0.3 0.6

Measured

Tonnage (kt) 139 138 203 171 204 162

Au (g/t) 3.94 3.97 2.39 2.76 2.01 2.42

Metal kg 549 549 486 472 411 393

oz 17,658 17,637 15,620 15,164 13,212 12,624

Indicated

Tonnage (kt) 6,211 6,015 12,324 9,805 14,516 10,976

Au (g/t) 4.16 4.28 2.07 2.49 1.75 2.17

Metal kg 25,835 25,739 25,539 24,416 25,349 23,804

oz 830,626 827,530 821,085 784,998 815,004 765,319

Measured + Indicated

Tonnage (kt) 6,350 6,153 12,528 9,976 14,720 11,139

Au (g/t) 4.16 4.27 2.08 2.49 1.75 2.17

Metal kg 26,385 26,288 26,024 24,888 25,760 24,197

oz 848,284 845,167 836,705 800,163 828,215 777,944

Inferred

Tonnage (kt) 7,849 6,151 12,076 8,250 13,743 9,003

Au (g/t) 2.98 3.68 1.85 2.51 1.60 2.21

Metal kg 23,413 22,662 22,381 20,717 21,981 19,936

oz 752,760 728,601 719,564 666,069 706,692 640,972

Notes: 1. Mineral Resources are not reserves until they have demonstrated economic viability

based on a feasibility study or pre-feasibility study. 2. Mineral Resources are reported inclusive of any reserves. 3. Grade represents estimated contained metal in the ground and has not been adjusted for

metallurgical recovery.


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Table 14.22 Björkdal Open Pit Resource Estimate (Ellis and Newall 2013)

Minimum Mining Width 1.0m 1.0m 2.5m 2.5m 4m 4m

Cut Off Grade (g/t) 0.3 0.6 0.3 0.6 0.3 0.6

Measured

Tonnage (kt) 74 73 111 87 113 84

Au (g/t) 3.27 3.31 1.90 2.30 1.68 2.11

Metal kg 242 242 211 201 190 178

oz 7,790 7,770 6,792 6,462 6,111 5,725

Indicated

Tonnage (kt) 5,286 4,573 8,529 5,172 9,205 5,254

Au (g/t) 2.06 2.31 1.20 1.71 1.07 1.56

Metal kg 10,895 10,552 10,274 8,820 9,882 8,191

oz 350,298 339,255 330,320 283,579 317,727 263,361

Measured + Indicated

Tonnage (kt) 5,360 4,646 8,640 5,260 9,318 5,338

Au (g/t) 2.08 2.32 1.21 1.72 1.08 1.57

Metal kg 11,138 10,794 10,485 9,021 10,073 8,370

oz 358,088 347,025 337,112 290,041 323,839 269,086

Inferred

Tonnage (kt) 5,889 3,752 6,760 3,770 7,320 3,852

Au (g/t) 1.59 2.26 1.20 1.82 1.06 1.64

Metal kg 9,391 8,470 8,120 6,844 7,783 6,318

oz 301,939 272,307 261,068 220,034 250,237 203,140

NB – 1. Mineral Resources are not reserves until they have demonstrated economic viability

based on a feasibility study or pre-feasibility study. 2. Mineral Resources are reported inclusive of any reserves. 3. Grade represents estimated contained metal in the ground and has not been adjusted for

metallurgical recovery.

14.15 Revised Open Pit Resources

Elgin revised the WAI open pit resource estimate in accordance with an internal policy of reporting resources within an optimized shell. The resource shell was generated using Geovia Whittle™ based on the parameters in Table 14.23


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Table 14.23 Parameters for Elgin Revised Open Pit Resource Estimate

Parameter Unit Resources Notes

Mining USD/t 2.35 Provided by Björkdal personel based on previous year costs

Plant USD/t 7.93 Provided by Björkdal personnel based on previous year costs

G/A USD/t 1.71 Provided by Björkdal personnel based on previous year costs

Other USD/t 2.15 Provided by Björkdal personnel based on previous year costs

Total Mining Cost USD/t 2.35 Calculated values from the above cost parameters

Total Mill Cost USD/t 11.79 Calculated values from the above cost parameters

Mining Dilution % 25.0 Based on 2012 reconciliation from base model values to mill

Ore Losses % 0.0

Process Plant

Au Mill Recovery <1g/t % 87.0 optimization Data for WAI

Au Mill Recovery >1g/t % 90.0 optimization Data for WAI

Gold Price USD/oz 1,600

Geotechnical Specifications

Overall Slope Angle degrees 55.9 Based on SRK Geotechnical report SE-384

Ramp wall flattening degrees 3.9

Bench Face Angle degrees 80 Based on SRK Geotechnical report SE-384

Berm Width m 10 Based on SRK Geotechnical report SE-384

Bench Height m 5 Based on equipment size

Benches Per Berm # 4 Based on safe mining bench heights between berms


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Parameter Unit Resources Notes

Ramp Width m 22 Based on equipment size

Ramp gradient % 10 Based on safe mining geometries

Smelter Terms

`

Selling Cost % 6 The selling cost is 6% of the gold price

The optimization included measured, indicated and inferred class material at a 0.32g/t cut-off, which was calculated from the economic parameters. The resulting optimized pit shell was imported into Geovia Surpac™ to generate the pit resources which are summarized in Table 14.24.

Table 14.24 Open Pit Resource Estimate Using Optimized Pit Shell

TONNAGES AND GRADES

Resource Type

Cut-off Measured Indicated Measured +

Indicated Inferred

Au (g/t)

Tonnage (000’s t) Au (g/t) Tonnage(000’s t) Au (g/t)

Open Pit 0.32 51.8 1.65 6,469 1.10 6,521 1.11 2,216 1.57

Ounces

Cut-off Measured (ounces)

Indicated (ounces)


Inferred

(ounces)

Open Pit 0.32 2,746 229,611 232,357 111,907


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15 MINERAL RESERVES ESTIMATE

15.1 Introduction

The WAI block model was used in the estimation of the open pit reserve estimates by Mr. Currie and the underground reserve estimates by Mr. Friesen.

15.2 Pit Reserves

Elgin undertook a pit optimisation using the Mineral Resource Block Model prepared by WAI. The model was depleted to contain only those Mineral Resources, which have not been extracted as of January 17, 2013. Elgin used Geovia Whittle™ software for the optimisation, applying conceptual financial and technical parameters outlined in Table 15.1. The optimised pit was then used as the basis for the reserves pit, designed in Geovia Surpac™. Reserves totals are based on the January 17, 2013 surface less depletion to the end of March 2013.

The objective of this study was to obtain an economical, practically mineable design to be used for annual and life of mine scheduling.

15.2.1 Optimisation Parameters Pit Reserves

The Datamine block model was imported into Geovia Whittle™. The pit optimisation was based on the three year rolling average gold price of US$1500/oz and excluded the underground and pillar zones of the deposit. The full list of parameters used is detailed in Table 15.1.

Table 15.1 Parameters for Elgin Revised Open Pit Reserves Estimate

Parameter Unit Reserves Notes

Mining USD/t 2.35 Provided by Björkdal personnelbased on

previous year costs

Plant USD/t 7.93 Provided by Björkdal personnel based on

previous year costs

G/A USD/t 1.71 Provided by Björkdal personnel based on

previous year costs

Other USD/t 2.15 Provided by Björkdal personnel based on

previous year costs

Total Mining Cost USD/t 2.35 Calculated values from the above cost

parameters

Total Mill Cost USD/t 11.79 Calculated values from the above cost

parameters

Mining Dilution % 25.0 Based on 2012 reconciliation from base

model values to mill


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Parameter Unit Reserves Notes

Ore Losses % 0.0

Process Plant

Au Mill Recovery <1g/t % 87.0 Optimization Data for WAI

Au Mill Recovery >1g/t % 90.0 Optimization Data for WAI

Gold Price USD/oz 1,500 3 year average

Geotechnical Specifications

Overall Slope Angle degrees 55.9 Based on SRK Geotechnical report SE-384

Ramp wall flattening degrees 3.9

Bench Face Angle degrees 80 Based on SRK Geotechnical report SE-384

Berm Width m 10 Based on SRK Geotechnical report SE-384

Bench Height m 5 Based on equipment size

Benches Per Berm # 4 Based on safe mining bench heights between

berms

Ramp Width m 22 Based on equipment size

Ramp gradient % 10 Based on safe mining geometries

Smelter Terms

Selling Cost % 6 The selling cost is 6% of the gold price

15.2.2 Pit Reserves Results

The resulting optimised pit shell was used as the basis for the reserve pit designs. The final design includes access ramps and catch benches. The reserves are summarized in Table 15.2.

Table 15.2 Björkdal Pit Reserves


Grade

(g/t) Ounces

Proven 49.6 1.48 2,366

Probable 5,454.4 1.04 182,858

Total 5,504.0 1.05 185,224

* Reserves totals are based on the January 17 2013 surface less depletion to the end of March 2013


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15.3 Underground Reserves

Underground Mine Reserves were developed by sectioning the unbounded block model provided by WAI 2013) drawing profiles to represent the mineable portion of each vein system in the section, and connecting the profiles for each vein to create a mineable solid for that vein system.

Measured and Indicated Resources were evaluated for each minable shape, mining recovery and dilution were applied, and those shapes that qualified in terms of size and grade were added to the mineral reserve. Reserves of the underground mine are summarized in Table 15.3.1.

Table 15.3.1 Björkdal Underground Reserves as at April, 2013

Reserve Category Tonnes Grade Au gpt

Grams Au

Ounces Au

Proven 36,226 2.02 73,326 2,357

Probable 2,054,656 2.07 4,252,845 136,732

Proven and Probable 2,090,882 2.07 4,326,172 139,090

The active mine includes all veins in the South East Extension Zone on the South to Lake Zone on the North and between approximately -180 and -415 elevations. See Figure 15.3.1 below.

Figure 15.3.1 Underground Zones – Björkdal Mine


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The veins are generally parallel and reasonably close together, typically within 30 m of each other, requiring minimal waste development to access them. In practice, when developing the cross-cuts between veins, it is the mine’s experience to quite often encounter mineralized material, sometimes substantially above cut-off grade, in these crosscuts. Depending on the grade, this material either reports directly to the mill for processing or is placed on the low grade stockpile for processing in the future. This additional material from crosscuts is not modeled and is therefore not accounted for in the reserve calculations. It generally grades lower than the stopes and the on vein development.

Many of the veins in reserves, particularly in South, Main, 610, and 620 zones, already have good access to them and in some case are either developed or partially developed.

15.4 Methodology:

15.4.1 Creating Solids

Using the unbounded block model, with a 1x5x5m sub block size, the blocks for the measured and indicated resource were displayed and profiles were drawn on section for each of the vein systems at 5 to 10m intervals through the entire ore body. At the 1.15 gpt cut off the larger vein systems showed reasonable continuity and it was usually possible to draw profiles around the vein on section leaving few or no empty cells within the profile. Profiles of adjacent vein systems were displayed together, and checked to ensure that profiles from one system were not mixed in with those of the adjacent one. Profiles of each vein system were then adjusted as needed and then connected to create minable solids.

15.4.2 Ore Reserves from the Block Model

Ore reserves were calculated by estimating the tonnes and grade within the minable shapes using the block model and applying appropriate mining dilution and recovery factors. The mining dilution applied is 30% and recovery is 80%. The stopes that remained above cut-off grade are tabulated in Table 15.4.2. Please note the adjustment made at the bottom of the Table 15.4.2 for mine production period January to March of 2013. This is to account for surveys that were not included in WAI’s resource calculations and block model construction.

Although the applied dilution factor of 30% was considered appropriate for a number of years, it may be somewhat underestimated to what is actually being achieved. The mine is planning to conduct a bulk sample study in the future in a form of a typical test stope to help understand the deposit better and where, amongst other factors, the actual dilution will be accurately measured. Dilution tends to be significantly higher in some stopes and development headings.

The mine has started to reduce dilution in stopes and the on vein development headings by cable bolting the walls in stopes, keeping the development headings as narrow as possible, analyzing and improving sampling, drilling, and blasting practices, and being more selective on the material reporting as ore. Appropriate development and cable bolting equipment has been purchased and is on site.

The stopes classified as reserves are well drilled, reasonably well understood geologically, are for the most part un-mined , have accesses in place or planned, and mining methods, mining costs, ground conditions, and metallurgy are all known factors.


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Table 15.4.2 Underground Ore Reserve Calculations

Stope In Situ Grade in g/t

In Situ Tonnes

In Situ Metal in

kg

30% Diluted Tonnes

80% Recovered

Tonnes

Diluted Grade in

g/t

Recovered Metal in

kg

Recovered Metal in

Oz

610_v1 3.23 166,871 538.510 216,933 173,546 2.48 430.808 13,851

610_v2 2.91 42,428 123.435 55,156 44,125 2.24 98.748 3,175

610_v3 2.67 83,599 223.619 108,679 86,943 2.06 178.895 5,752

610_v4 5.04 20,032 100.890 26,041 20,833 3.87 80.712 2,595

610_v5 3.12 38,474 120.104 50,016 40,013 2.40 96.083 3,089

610_v6 2.55 5,887 15.022 7,653 6,122 1.96 12.017 386

610_v7 2.34 14,880 34.770 19,344 15,475 1.80 27.816 894

610_v8 4.24 51,443 218.374 66,876 53,501 3.27 174.699 5,617

610_v9 1.92 20,480 39.269 26,624 21,300 1.47 31.415 1,010

610_v10 2.87 8,607 24.701 11,189 8,951 2.21 19.761 635

610_v11 1.20 14,049 16.802 18,264 14,611 0.92 13.441 432

610_v12 3.35 12,968 43.439 16,859 13,487 2.58 34.751 1,117

610_v13 3.56 5,105 18.193 6,636 5,309 2.74 14.555 468

610_v14 3.22 42,155 135.621 54,802 43,841 2.47 108.497 3,488

610_v15 1.69 9,404 15.905 12,225 9,780 1.30 12.724 409

610_v16 2.12 8,222 17.415 10,688 8,551 1.63 13.932 448

610_v17 3.16 40,083 126.545 52,108 41,687 2.43 101.236 3,255

610_v18 3.35 47,710 159.973 62,023 49,618 2.58 127.978 4,115

610_v19 2.61 28,170 73.647 36,621 29,297 2.01 58.918 1,894

Total 610 3.10 660,566 3,553.953 858,736 686,989 2.38 1,636.985 52,630

620_v1 3.32 8,881 29.441 11,545 9,236 2.55 23.553 757

620_v2 3.84 7,815 30.005 10,159 8,127 2.95 24.004 772

620_v3 3.65 8,356 30.500 10,863 8,690 2.81 24.400 784

620_v4 4.67 17,194 80.321 22,353 17,882 3.59 64.257 2,066

620_V5 1.72 7,820 13.424 10,165 8,132 1.32 10.739 345

620_v6 3.18 12,401 39.463 16,121 12,897 2.45 31.570 1,015

620_v7 1.62 30,147 48.944 39,191 31,353 1.25 39.155 1,259


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In Situ Tonnes

In Situ Metal in

kg

30% Diluted Tonnes

80% Recovered

Tonnes

Diluted Grade in

g/t

Recovered Metal in

kg

Recovered Metal in

Oz

620_v8 2.29 14,643 33.601 19,036 15,229 1.77 26.881 864

620_v9 1.93 23,297 45.005 30,286 24,229 1.49 36.004 1,158

620_v10 2.73 14,707 40.204 19,120 15,296 2.10 32.163 1,034

620_v11 1.58 31,928 50.347 41,506 33,205 1.21 40.278 1,295

620_v12 2.90 6,934 20.127 9,014 7,211 2.23 16.102 518

620_v13 2.02 3,409 6.875 4,431 3,545 1.55 5.500 177

620_v14 1.88 20,026 37.732 26,033 20,827 1.45 30.185 970

620_v16 2.81 17,626 49.612 22,913 18,331 2.17 39.690 1,276

620_v17 3.80 19,425 73.785 25,253 20,202 2.92 59.028 1,898

620_v18 4.54 21,259 96.456 27,636 22,109 3.49 77.165 2,481

620_v19 5.13 6,395 32.778 8,314 6,651 3.94 26.222 843

Total 620 2.79 272,262 1,517.239 353,940 283,152 2.14 606.896 19,512

LZ_v1 2.85 11,533 32.857 14,993 11,994 2.19 26.286 845

LZ_v2 3.04 51,639 156.756 67,131 53,705 2.34 125.405 4,032

LZ_v3 2.47 18,096 44.683 23,524 18,820 1.90 35.747 1,149

LZ_v4 2.16 17,889 38.564 23,256 18,605 1.66 30.851 992

LZ_v5 2.48 25,396 62.972 33,015 26,412 1.91 50.378 1,620

LZ_v6 2.52 16,460 41.450 21,398 17,118 1.94 33.160 1,066

LZ_v7 1.68 67,637 113.712 87,928 70,343 1.29 90.969 2,925

LZ_v8 1.84 11,503 21.160 14,954 11,963 1.42 16.928 544

LZ_v9 1.50 47,593 71.289 61,871 49,497 1.15 57.031 1,834

LZ_v10 1.78 55,821 99.117 72,567 58,054 1.37 79.294 2,549

LZ_v11 1.98 13,182 26.068 17,137 13,709 1.52 20.854 670

LZ_v12 2.26 18,750 42.309 24,375 19,500 1.74 33.847 1,088

LZ_v13 2.35 24,968 58.587 32,459 25,967 1.80 46.870 1,507

LZ_v14 3.41 18,357 62.543 23,864 19,091 2.62 50.035 1,609

LZ_V15 2.04 62,525 127.653 81,282 65,026 1.57 102.122 3,283

LZ_v16 1.98 109,658 217.037 142,556 114,045 1.52 173.629 5,582


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In Situ Tonnes

In Situ Metal in

kg

30% Diluted Tonnes

80% Recovered

Tonnes

Diluted Grade in

g/t

Recovered Metal in

kg

Recovered Metal in

Oz

LZ_v17 1.68 37,901 63.560 49,272 39,417 1.29 50.848 1,635

LZ_v18 1.57 24,324 38.083 31,622 25,297 1.20 30.466 980

LZ_v19 3.40 2,709 9.211 3,521 2,817 2.62 7.369 237

LZ_v20 4.94 12,099 59.755 15,728 12,583 3.80 47.804 1,537

LZ_v21 2.02 14,215 28.734 18,480 14,784 1.55 22.987 739

LZ_v23 1.67 3,641 6.099 4,734 3,787 1.29 4.879 157

LZ_v24 2.11 12,584 26.537 16,360 13,088 1.62 21.230 683

LZ_v25 2.15 1,295 2.783 1,683 1,346 1.65 2.227 72

LZ_v26 1.61 3,737 6.032 4,858 3,886 1.24 4.826 155

LZ_v28 1.89 10,107 19.136 13,139 10,511 1.46 15.309 492

LZ_v30 2.00 32,957 65.876 42,844 34,275 1.54 52.701 1,694

Total LZ 2.12 726,577 3,085.131 944,550 755,640 1.63 1,234.053 39,676

MZ_v1 2.81 3,019 8.482 3,924 3,139 2.16 6.785 218

MZ_v2 1.57 11,817 18.599 15,362 12,289 1.21 14.879 478

MZ_v4 2.84 19,769 56.049 25,699 20,560 2.18 44.840 1,442

MZ_v5 2.04 10,600 21.617 13,780 11,024 1.57 17.294 556

MZ_v7 1.97 30,479 59.941 39,623 31,698 1.51 47.953 1,542

MZ_v8 2.90 25,862 74.988 33,621 26,897 2.23 59.990 1,929

MZ_v9 1.50 8,903 13.370 11,575 9,260 1.16 10.696 344

MZ_v10 1.99 12,729 25.347 16,547 13,238 1.53 20.278 652

MZ_v11 1.63 4,145 6.745 5,388 4,310 1.25 5.396 173

MZ_v12 3.84 11,393 43.794 14,811 11,849 2.96 35.035 1,126

MZ_v13 1.56 27,893 43.492 36,261 29,009 1.20 34.794 1,119

MZ_v14 4.59 28,816 132.332 37,460 29,968 3.53 105.865 3,404

Total MZ 2.58 195,424 1,009.511 254,051 203,241 1.99 403.804 12,983

SEE_v1 10.64 12,579 133.882 16,353 13,082 8.19 107.106 3,444

SEE_v2 4.38 16,371 71.663 21,283 17,026 3.37 57.331 1,843

SEE_v3 2.21 19,586 43.316 25,462 20,369 1.70 34.653 1,114


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In Situ Tonnes

In Situ Metal in

kg

30% Diluted Tonnes

80% Recovered

Tonnes

Diluted Grade in

g/t

Recovered Metal in

kg

Recovered Metal in

Oz

Total SEE 5.13 48,536 497.724 63,097 50,478 3.94 199.090 6,401

SZ_v1 1.87 30,723 57.576 39,940 31,952 1.44 46.061 1,481

SZ_v2 3.46 16,025 55.445 20,833 16,666 2.66 44.356 1,426

SZ_v3 1.96 95,231 186.397 123,800 99,040 1.51 149.117 4,794

SZ_v5 1.81 16,618 30.046 21,603 17,283 1.39 24.037 773

SZ_v6 2.40 77,768 186.559 101,098 80,878 1.85 149.247 4,798

SZ_v8 4.45 5,810 25.840 7,554 6,043 3.42 20.672 665

Total SZ 2.24 242,175 1,083.724 314,827 251,862 1.72 433.489 13,937

Total Stopes

2.63 2,145,540 5,642.896 2,789,202 2,231,361 2.02 4,514.317 145,139

Jan to Mar 2013 Production 140,479 1.34 188.145 6,049

Total Reserves 2,090,882 2.07 4,326.172 139,089

Figure 15.4.2 3D View of Mined Out Areas and Stopes in Ore Reserve. Mined out areas are shown in light green.

15.4.3 Mining Costs and Determination of Economic Cut off Grade for Underground

Costs and factors used to calculate the economic cut-off grade are tabulated below:


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Table 15.4.3 Underground Cut-Off Grade Cost Inputs

Factor/Cost US $

Au Price $us/oz. 1500.00

Cost of Sales (6%) 90.00

Net Revenue $US/oz. 1410.00

Mill Recovery 88%

Recovered Revenue/oz 1240.80

Annual Production Kt 600

Mining Cost/t 33.16

Mining Capital/t 5.58

Process Cost/t 7.93

G & A Cost/t 1.71

Other Costs/t 2.15

Total Cost/t 50.53

Total Cost/t less Capital 44.95

Cut-off Grade/t with Capital 1.29

Cut-off Grade/t no Capital 1.15

Based on actual 2012 costs

The economic cut-off is calculated to be 1.15 grams/tonne.


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16.0 MINING METHODS

The open pit mining is done by contract with Björkdal staff providing geological and technical oversight of the mining. The open pit produces approximately 700,000 tonnes of ore per year. Five meter benches are drilled by the contractor, NCC Roads AB, using an Atlas Copco Coprod production drill. Charging and blasting are also done by contractor (EPC Sverige AB) using a combination of emulsion and Anfo explosives. Mucking is done in 2.5m lifts to ore and waste profiles defined and marked on the blasted muck piles by Björkdal staff geologists. Contract mucking is done using 80 – 90 tonne excavators using 2.7m wide buckets to 45-65 tonne rock trucks, which haul either directly to the crushing circuit or a laydown adjacent to the jaw crusher. Contract mucking and hauling of both ore and waste is supplied by Benny’s Grav AB.

Reverse circulation drilling is done in production areas of the open pit to improve mining selectivity, contract operation performed by Styrud Jarvsoborr AB. Exploration drilling of the pit area is done using a combination of reverse circulation drilling by Styrud Jarvsoborr AB and diamond core drilling by Protek Norr AB.

The underground mining is conducted using a combination of contractor and owner-operated equipment and personnel. Engineering, geology, and technical oversight are provided by Björkdal staff. The Company has hired personnel to provide 50% of mine services with the other 50% provided by the contract miner Strabag Sverige AB. Maintenance of the company-owned mobile underground fleet and pumps is conducted by Björkdal staff in the newly-constructed underground service shop.

The Underground mine produces approximately 600,000 tonnes of ore per year. The underground mining method being utilized at the Björkdal Mine is longhole open stoping or blast hole open stoping with sub-level spacing of 20 metres. The mining contractor, Strabag Sverige AB, is responsible for all drift mining while the stope mining is done by Björkdal personnel. The underground ore haulage is done through a contract with Renfors Akeri AB.

Using this mining method, cross-cut drifts are excavated perpendicular to the gold bearing veins vertically every 20 metres. The veins are developed with sublevels driven along the vein. The veins are drilled with either an Atlas Copco Simba S7D or M7C drill from the lower level upwards. When production drilling is completed, an initial opening or slot raise is blasted. The broken rock is removed after each blast by the mucking sublevel using a Sandvik LH410 scooptram. Once the muck from the slot raise is removed, the production drill lines are blasted into the void produced by the slot raise. These lines are blasted in groups of 3 to 5 lines with enough muck removed between blasts to allow sufficient void for the following blast. When the entire mining block is blasted, the scooptram mucks out the remaining broken ore. The scooptram is fitted with a radio remote control to allow mucking of material beyond the blasted brow. By leaving small pillars, and changing the mining sequence, this mining method can be adapted to changing ground conditions, and can be modified to drill production holes either down towards the mucking sublevel or up towards a previously developed sublevel or known mining boundary, as described.

Exploration diamond drilling for the underground is performed through contract with Protek Norr AB. All underground electrical installations are the responsibility of an electrical contractor, Bergteamet (Electro) AB.


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By the end of 2013, the Company is planning to phase out the underground contractor, Strabag Sverige AB, and begin doing all underground development in house. Necessary equipment has been purchased and is on site.


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17. RECOVERY METHODS

17.1 General

The Björkdal plant was constructed by Davy McKee (UK) and uses conventional crushing and grinding followed by a combination of gravity and flotation processing techniques to recover gold to four separate concentrates. Production commenced in 1988 and increased to reach a maximum throughput of 1.32 Mtpa in 1998. No ore was treated from July 1999 through September 2001 due to a period of bankruptcy of the previous operator. Production re-commenced in September 2001 and has been continuous since that time, with the plant producing its 1 millionth ounce in 2010. In 2012 the plant processed 1.274 million tonnes of ore.

17.2 Recovery Description

17.2.1 Crushing

Ore is trucked to the plant and either dumped directly into a hopper or stockpiled on the ground beside the hopper. Ore is fed into a 1.2 x 1.0 metre Svedala jaw crusher and the crushed product is conveyed to a triple-deck screen. The -11mm product is conveyed directly to the fine ore bin. This material has a high moisture content and can cause blockages in the fine ore bin if frozen during winter months. A separate conveyor therefore exists to transfer the material, via a separate feeder, directly into the plant.

The +11 mm material is transferred via a stockpile to a Svedala 4000 cone crusher. The crusher product is conveyed to a screen and the screen undersize is conveyed to the fine ore bins. The screen oversize is conveyed to a third stage of crushing which consists of a Sandvik H6800 and a Sandvik 865 cone crusher in parallel. The crushed products are conveyed to -8mm screens with the screen undersize reporting to the fine ore bin and the oversize being returned to the crushers.

The crushing plant operates on a three-shift basis during weekdays and two shifts on weekends. The average plant throughput during drift is approximately 250t/h.

17.2.2 Grinding

The crushed ore is conveyed on two parallel conveyors to a 2.7 x 3.7m rod mill and a 3.5 x 4.2m ball mill. Both mills were supplied by Morgardshammar; the rod mill is fitted with a 350kWmotor and the ball mill with a 1000kW motor. Both mills are rubber lined.

Mill discharges are pumped to a 1.5 mm horizontal vibrating screen. The screen oversize is returned to the ball mill and the undersize is pumped to Krebs D15B cyclones. Cyclone overflow passes to flotation and the undersize gravitates to the gravity section.

17.2.3 Gravity Processing

The cyclone underflow product gravitates to two primary Mineral Deposits HC1 spiral separator batteries. The spiral concentrate is cleaned using Outotec MC7000 spirals. The secondary spirals concentrate passes to Deister shaking tables. The table concentrate is cleaned on a third table and the


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cleaned concentrate is pumped into a storage tank. The cleaner table tailings reports into the "middlings concentrate".

The primary table tailings are reground in a small ball mill and the ground product are pumped into a Knelson KC12 concentrator. The Knelson concentrate reports into the storage tank. The concentrate in the storage tank is cleaned on a daily basis over a shaking table. The concentrate reports as the final "high grade gravity concentrate" and the tailings from the clean-up report to the "middlings concentrate". The tailings from primary spiral separators and Knelson concentrator KC12 is pumped to a 3 x 4.0m regrind mill (500kW). The reground product is returned to the 1.5 mm screen.

17.2.4 Flotation

The grinding cyclone overflow is re-cycloned and the underflow passes to a Knelsonconcentrator XD30. The discharge from the Knelson concentrator is pumped into two Sala 3.3 cubic metre cells. The tailings from these cells pass to a bank of five Outokumpu 8 cubic metre cells, combined with the overflow from the re-cycloned product.

The float tailings are the final plant tailings and are pumped directly into the tailings maintenance facility.

17.2.5 Concentrate Dewatering

The flotation concentrate is pumped to a Delkor Enviroclear thickener where lime and flocculant are added to assist settling. The thickened pulp is filtered in a vacuum drum filter. The final filter cake contains 10% moisture.

17.2.6 Sampling

Automatic samplers are installed on the middlings concentrate, the flotation concentrate, Knelson concentrate and the final plant tailings streams. The plant feed is not sampled due to the problems associated with sampling of coarse gold particles. The plant recovery is calculated from the weights of the concentrates produced, the plant feed tonnage and the tailings assays

17.3 Production Data

The production records for the Björkdal concentrator for the calendar years since Gold-Ore, and subsequently Elgin, have operated the mine site are given in Table 17.1.


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Table 17.1 Production Data 2006-2012

Year Throughput in tonnes

(*1000) Recovery

(%) Feed (g/t)

2006 1,209,600 86.8 0.61

2007 1,109,482 85.2 0.63

2008 1,169,768 87.1 0.89

2009 1,063,741 88.4 1.24

2010 1,154,650 88.9 1.23

2011 1,232,925 88.6 1.16

2012 1,274,125 87.7 1.19

17.4 Tailings Maintenance Facility

The tailings maintenance facility (“TMF”) is located in an area of gently undulating relief. Surface elevations range between 195m above datum across the higher ground of the north of the impoundment to about 140m above datum to the eastern end of the impoundment. Tailings from the concentrator are pumped a distance of about 1.5 kilometers in two stages from the processing plant to the TMF. The tailings are considered to be inert by the Swedish environmental authorities.


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18 PROJECT INFRASTRUCTURE

The Björkdal mine consists of administrative buildings, processing plant, maintenance facilities, assay laboratory, core logging facilities, tailings ponds, waste dumps and stockpiles. The administrative building, along with office space, contains kitchen facilities and a mine dry. The mine is not isolated and is readily accessible by paved roads and has access to low cost hydro power and water. On site communications include regular land line telephone and internet service.


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19 MARKET STUDIES AND CONTRACTS

There are worldwide gold markets into which the Company can sell and, as a result, the Company is not dependent on a particular purchaser with regard to the sale of the gold which it produces.

The Company currently has several contracts as more particularly described in Section 16 of this report.


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20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

All operations are fully permitted in accordance with Swedish environmental and health & safety legislation. The latest mining permit M25-10 was issued on the 24th of June 2010 and is in good standing Under Swedish law there is no limit on the permit but the government may make adjustments as required to meet any regulation changes. A new permit related to the mine tailings area is required before 2018 and the application process will begin in 2014.

Björkdal is permitted to use the Kage river as a water source for the processing plant. The allowed amount is 180m3 per hour. The plant uses approximately 150m3 per hour and of this, half is recycled from the tailings facility. Water used at the mine site for purposes other than the processing plant is sourced from dug wells.

Studies related to the mine operations include, ecological surveys, culture surveys, fish surveys in local streams, aquatic habitat surveys in local streams and waste characterization. A full environmental study is carried out every three years by a independent consultant.

Monitoring, control and management, policy and procedures are well documented and entirely appropriate to the type of operation. The mine has low sulphide content resulting in no acid rock drainage potential. Gold is recovered by mechanical and gravity processes with no use of cyanide. There are no harmful elements associated with the mine tailings and they have been declared non-toxic by the Swedish government. There are no issues with community impact. The mine is located in a part of Sweden that has a long history of mining activity and mining is accepted as a socially responsible and necessary contributor to the local economy. The mine is not located in reindeer habitat and therefore there are no issues with the indigenous Sami population.

The Company has a reclamation bond with the Swedish authorities in the amount of $2,426,539.


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21. CAPITAL AND OPERATING COSTS

21.1 Capital Costs

Since reopening the mine in 2001 Björkdal Mine has generated enough cash flow cover the sustaining capital expenditures required to maintain the planned production rates over the life-of-mine. These capital expenditures include, but are not limited to, costs for tailings expansion, replacement costs for equipment and capital development.

21.2 Operating Costs

Björkdal Mine has been operating consistently since 2001. The operating costs and production levels are well understood.

Table 21.2.1 shows the 2012 actual costs and the budgeted 2013 costs.

Table 21.2.1 Björkdal Production Data 2012

Item Unit 2012**

Open Pit Mining Costs $/t mined 2.35

Underground Mining Costs $/t mined 33.16

Stockpile Handling Costs $/t mined 1.49

Process Costs* $/t milled 11.79

OP Tonnes Mined t x 1000 3,447

UG Tonnes Mined t x 1000 762

OP Tonnes Processed t x 1000 568

UG Tonnes Processed t x 1000 580

Stockpile Tonnes Processed t x 1000 127

Total Tonnes Processed t x 1000 1,275

Gold Ounces Produced oz 42,839

Cash Operating Costs $/oz 978

* These costs include G&A and support costs ** 2012 average exchange rate of 6.78 SEK/USD

Elgin has provided 2013 guidance of 45,000 to 49,000 ounces of gold produced at a cash cost per ounce

of $1,040 to $1.145.


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The operating costs per unit of production have been relatively consistent from year to year. The cost for processing and mine support are expected to remain constant for the remainder of the mine life, except when they are changed due to inputs that affect the entire gold mining industry, including, but not limited to, changes in, fuel costs, reagent costs, exchange rates, labour costs and inflation. The unit costs for mining are expected to increase as the pit and underground deepen and they are also affected by the previously listed inputs.


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22 ECONOMIC ANALYSIS

Not applicable as the Björkdal Mine is currently in production.


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23 ADJACENT PROPERTIES

There is nothing to report in this section.


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24 OTHER RELEVANT DATA AND INFORMATION

At this time, all relevant data and information regarding the Björkdal Mine is included in other sections of this report.


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25 INTERPRETATION AND CONCLUSIONS

The Björkdal Mine in produces approximately 42,000 – 46,000 ounces of gold annually by processing 1.3 million tonnes of ore. The mine is fully permitted and operates 7 days a week all year round. Four separate gold concentrates are produced and sold to smelters. The ore is sourced by open pit and underground mining techniques by a combination of owner operated and contractors’ fleet.

A new reserve estimate was completed using a base case gold price of US$1,500 per ounce which is the average market price over the last three years. In the last few months the price of gold has been extremely volatile and generally trending downwards. Recently gold price has fluctuated between $1,200-$1,400/oz. Should the price of gold fall much lower a new reserve estimate should be considered with a higher cut-off grade.

The proven and probable reserve estimate for the open pit unit as of the end of March, 2013 is 5.5 million tonnes of ore grading 1.05 grams gold per tonne. At the current mining rate of about 700,000 tonnes of ore per annum, the open pit has sufficient reserves for approximately 8 years of production.

The proven and probable reserve estimate for the underground unit as of the end of March, 2013 is 2.09 million tonnes of ore grading 2.07 grams gold per tonne. At a current production rate of about 600,000 tonnes of ore per annum, the underground has sufficient reserves for approximately 3.5 years of production. There is a significant amount of tonnage in the resource category underground. As the underground access is continually developed and the infrastructure is established, the mine can continually drill the ore body to further define the resources and convert them into reserves. The mine has been successful at replacing the reserves mined underground for the last four years.


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26 RECOMMENDATIONS

Conduct the underground bulk sample study as planned to help understand the deposit better and gain a higher level of confidence in local tonnes, grade, dilution, and recoveries estimation.

Improve in the way stope reconciliations are performed by surveying the stopes more consistently and building the required shapes and models in Surpac as needed on a timely basis.

Continue with efforts to reduce dilution in stopes and the on vein development headings by cable bolting the walls in stopes, keeping the development headings as narrow as possible, analyzing and improving sampling, drilling, and blasting practices, and being more selective with the material reporting as ore. The potential gains from higher grade ore reporting to the mill are substantial.

The remnants and pillars above the active mine need to be re-evaluated to see what, if any, opportunities exist for further mining. The pillar between the open pit and underground is 20 m thick and may be potentially minable via the open pit in the future.

While the previous and current practices of manual transfer from assay reports into the database has been considered adequate, Elgin is instigating a new automated procedure for importing drilling, underground sampling and assay information into the database. This will be carried by the implementation of GeoSparkTM software that has been modified specifically for the Björkdal mine.

Exploration drilling should be continued to convert inferred resources to indicated and to expand the near mine resource base to help extend the mine life beyond 2020. The life of the underground mine can potentially extend well beyond the 3.5 years that are currently in reserve as the mineralization is still open to the North, South and East at depth.


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27. REFERENCES

Billstrom, K., Broman, C., Jonsson, E., Recio, C., Boyce, A.J., Torssander, P., (2009). Geochronological, stable isotopes and fluid inclusion constraints for a premetamorphic development of the intrusive-hosted Björkdal Au deposit, northern Sweden. International Journal Earth Sciences 98:1027-1052

Ellis, R., Newall, P.S., (2013). Technical Report Björkdal Resource Estimation. Wardell Armstrong International Ref 61-1239 Report Number MM832.

Internal Corporate Records.

Lang, J.R., Baker, T., Hart, C.J.R. & Mortensen, J.K., (2000). An exploration model for intrusion-related gold systems. Society of Economic Geologists Newsletter 40, p.1-15.

Mercier-Langevin, P., McNicoll, V., Allen, R.L., Blight, J.H.S., Dube, B., (2013). The Boliden gold-rich volcanicogenic massive sulphide deposit, Skellefte district, Sweden: new U-Pb age constraints and implications at deposit and district scale. Miner Deposita 48:485-504

Newall, P.S., Wheeler, A. (2005). Resource Evaluation and Pit Optimisation of the Björkdal Gold Mine, Northern Sweden.. Wardell Armstrong International Ref 61-0302 Report No: MM119.

Saiang, D., (2012). Inter-Ramp Slope Stability Analysis for Björkdal Mine. SRK Consulting Project Number SE-384.

Thompson, J.F.H, Sillitoe, R.H., Baker, T., Lang, J.R. & Mortensen, J.K.,(1999). Intrusion-related gold deposits associated with tungsten-tinprovinces. Mineralium Deposita, v. 34, p. 197-217.

Wasylyshyn, R.S., Dickson, G.D., (2012). N-43-101 Technical Report on the Mineral Resource and Mineral Reserve Estimation for the Björkdal Gold Mine, Sweden. Gold-Ore Resources Ltd.


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28. APPENDIX 1 – VARIOGRAPHY Y (FROM: ELLIS AND NEWALL 2013)

Final Modelled Semi-Variogram – Domain 11


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97



98



99



100



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29. CERTIFICATE OF AUTHOURS

CERTIFICATE OF QUALIFIED PERSON

As an author of this Technical Report on the Björkdal Mine for Elgin Mining Inc. (“Elgin”), I, Gordon Clarke, P.Geol., of 12187 Cherrywood Drive, Maple Ridge, British Columbia, Canada, V2X 0X3, do hereby certify that:

1) This certificate pertains to the Technical Report dated August 28, 2013 and entitled “NI 43-101 Technical Report on the Mineral Resource and Mineral Reserve Estimation for the Björkdal Gold Mine, Sweden.” The effective date of the report is March 31, 2013;

2) I am the former Vice President, Exploration of Elgin and have carried out this assignment as a private consultant.

3) I am a member in good standing of the Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists (NAPEG member #1584). I graduated from Acadia University with a Bachelor of Science degree (Honours) in Geology in 1990 and from the University of New Brunswick with a Master’s of Science degree in Geology in 1994;

4) I have practiced my profession continuously for the last 14 years as a consultant and employee for a number of companies in a variety of countries and geological environments;

5) I am familiar with Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-101”) and, by reason of education, experience and professional registration, I fulfill the requirements of a Qualified Person as defined in NI 43-101;

6) My most recent personal inspection of the Björkdal Mine was from June 19 to June 28, 2013;

7) Prior to the completion of this Technical Report, I held the position Exploration Manager with Elgin from September 13, 2012 to December 31, 2012 and held the position of Vice President, Exploration with Elgin from January 1, 2013 to July 31, 2013. I directly supervised the work pertaining to the mineral resource and reserve estimates for the Björkdal Mine that was carried out by Elgin employees. In addition, I currently hold securities in Elgin. Accordingly, I am not independent of Elgin as defined by NI 43-101;

8) I am responsible for the preparation of sections 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14-14.14, 18, 20, 22, 23, 24 and 27 of this Technical Report, as well as those portions of sections 1, 2, 25, and 26 that pertain to the foregoing sections;

9) As of the date of this certificate, to the best of my knowledge, information and belief, the portions of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make this Technical Report not misleading; and

10) I have read the Canadian National Instrument 43-101 and the portions of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

Dated this 28th day of August, 2013

“Gordon Clarke”

Gordon Clarke, P.Geol. Private Consultant


As an author of this Technical Report on the Bjorkdal Mine for Elgin Mining Inc. (“Elgin”), I, George Friesen, P.Eng., do hereby certify that:


2) I am employed as the Manager of Technical Services of Elgin, located at Suite 201 – 750 West Pender Street, Vancouver, British Columbia, Canada V6C 2T7;

3) I am a member of the Northwest Territories Association of Professional Engineers and Geoscientists (NAPEG, member #1537). I graduated from University of British Columbia with a B.A.Sc. in Mineral Engineering in 1988;

4) I have practiced my profession continuously since 1988 and have relevant experience in the mining industry including the design, construction and operation of mines;


6) My most recent personal inspection of the Björkdal Mine was from June 10 to June 20, 2013;

7) I have been employed by Elgin as Manager of Technical Services since November 1, 2012. In addition, I currently hold securities in Elgin. Accordingly, I am not independent of Elgin as defined by NI 43-101;

8) I am responsible for the preparation and review of sections 3, 15, 16 and 24 of this Technical Report, as well as those portions of sections 1, 2, 25, and 26 that pertain to the foregoing sections;




“George Friesen”

George Friesen, P.Eng. Manager of Technical Services of Elgin Mining Inc.


As an author of this Technical Report on the Bjorkdal Mine for Elgin Mining Inc. (“Elgin”), I, James A. Currie, P.Eng., do hereby certify that:


2) I am employed as the Chief Operating Officer of Elgin, located at Suite 201 – 750 West Pender Street, Vancouver, British Columbia, Canada V6C 2T7, tel. 604.682.3366, email [email protected];

3) I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC, member #13368). I graduated from Queen’s University with a B.Sc. (Hons.) in Mining Engineering in 1979;

4) I have practiced my profession continuously since graduation and have 34 years of relevant experience in the mining industry including the design, construction and operation of mines;


6) My most recent personal inspection of the Björkdal Mine was from July 30 to August 1, 2013;

7) I have been employed by Elgin as Chief Operating Officer since August 1, 2012. In addition, I currently hold securities in Elgin. Accordingly, I am not independent of Elgin as defined by NI 43-101;

8) I am responsible for the preparation and review of 3, 14.15, 15, 16, 17, 19, 21 and 24 of this Technical Report, as well as those portions of sections 1, 2, 25, and 26 that pertain to the foregoing sections;




“James A. Currie”

James A. Currie, P.Eng. Chief Operating Officer of Elgin Mining Inc.

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