Download pdf - Technical Report: Blue River Resource Update

Prepared for:

Commerce Resources Corp.

Prepared by:

Albert Chong, P.Geo.

Tomasz Postolski, P.Eng.

Ramon Reyes Mendoza, P.Eng.

Tony Lipiec, P.Eng.

Behrang Omidvar, P.Eng.

Effective Date: 22 June 2012

Project No. 168967

Commerce Resources Corp. Blue River Tantalum-Niobium Project British Columbia, Canada

NI 43-101 Technical Report on Mineral Resource Update

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical

Report for Commerce Resources Corporation (Commerce) by AMEC

Americas Limited (AMEC). The quality of information, conclusions,

and estimates contained herein is consistent with the level of effort

involved in AMEC’s services, based on: i) information available at the

time of preparation, ii) data supplied by outside sources, and iii) the

assumptions, conditions, and qualifications set forth in this report. This

report is intended for use by Commerce subject to the terms and

conditions of its contract with AMEC. Except for the purposes

legislated under Canadian provincial securities law, any other uses of

this report by any third party is at that party’s sole risk.

AMEC Americas Limited 111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3 Tel (604) 664-4315 Fax (604) 669-9516 www.amec.com

CERTIFICATE OF QUALIFIED PERSON

Albert Chong, P.Geo. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3

Phone: (604) 664-4116 E-mail: [email protected]

I, Albert Chong, P.Geo., am employed as a Principal Geologist with AMEC Americas Limited.

This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British

Columbia, Canada, NI 43-101 Technical Report on Mineral Resource Update” with an effective date

of 22 June 2012 (the “Technical Report”).

I am a Professional Geoscientist in the Province of British Columbia (P.Geo. #23773). I graduated

from McMaster University, Hamilton, Ontario with a B.Sc. degree in Geology, and from the University

of Tasmania with a M.Sc. degree in Exploration Geoscience.

I have practiced my profession for 27 years since graduation. I have been directly involved in green

fields and brown fields exploration, mining operations, consulting, and resource estimation of base

metal, precious metal and rare metal deposits.

As a result of my experience and qualifications, I am a Qualified Person as defined in National

Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101). I am the Qualified

Person responsible for Sections 1 to 12, 20, and 23 to 27 of the Technical Report.

I visited the Blue River property from 11 to 16 July 2010, 27 to 30 June 2011, and 6 to 14 September

2011.

I am independent of Commerce Resources Corporation as independence is described by Section

1.5 of NI 43–101.

I have been involved as an independent consultant on the Blue River Ta-Nb Project since 2010.

I have read NI 43–101 and this report has been prepared in compliance with that Instrument.

As of the date of this certificate, to the best of my knowledge, information and belief, the Technical

Report contains all scientific and technical information that is required to be disclosed to make the

Technical Report not misleading.

“signed and stamped”

Albert Chong, P.Geo.

Dated: 4 July 2012



Tomasz Postolski, P.Eng. AMEC Americas Limited



I, Tomasz Postolski, P.Eng., am employed as a Senior Geostatistician with AMEC Americas Limited.




I am a Professional Engineer in the Province of British Columbia (P.Eng. #34784). I have graduated

from The University of Mining and Metallurgy, Krakow, Poland with a Magister Inzynier degree in

Geological Engineering, and from the University of British Columbia with a Master of Applied

Science degree also in Geological Engineering. I have completed the Citation Program in Applied

Geostatistics at the Centre for Computational Geostatistics at the University of Alberta.

I have 18 years of consulting, mine operations, and academic experience specializing in

geostatistical mineral resource estimation and geological evaluation of gold, copper, rare earth

metals and other mineral deposits in Canada and abroad.



Person responsible for Section 14 and those portions of the Summary, Interpretation and

Conclusions, and Recommendations that pertain to this Section of the Technical Report.

I visited the Blue River property 27 to 30 June 2011.


1.5 of NI 43–101.

I have been involved with mineral resource estimation on the Blue River Ta-Nb Project since 2010.






Tomasz Postolski, P.Eng.

Dated: 4 July 2012



Ramon Mendoza Reyes (P.Eng.) AMEC Americas Limited



I, Ramon Mendoza Reyes (P.Eng.) am employed as a Principal Mining Engineer with AMEC Americas Limited. This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British



I am a Professional Engineer in the province of British Columbia. I graduated in 1989 from the

National Autonomous University of Mexico with a bachelor’s degree in Mining Engineering, and in

2003 completed a M.Sc. Degree in Mining & Earth Systems Engineering from the Colorado School

of Mines in Golden, Colorado, USA. I have practiced my profession for 22 years, and have

previously been involved with mine designs, mine planning and mine operations for base metal,

disseminated sulphide and industrial mineral projects in North America and South America.



Person responsible for Sections 15, 16, and 18 and those portions of the Summary, Cost Estimates

Interpretation and Conclusions, and Recommendations that pertain to the mining sections of the

Technical Report.

I visited the Blue River property in British Columbia from 12 to 14 July 2010.


1.5 of NI 43–101.

I have been involved with the mining aspects of the Blue River Tantalum–Niobium Project since

January 2010.






Ramon Mendoza Reyes, P.Eng.

Dated: 4 July 2012



Ignacy (Tony) Lipiec (P.Eng.) AMEC Americas Limited


Tel: 604-664-3130 E-mail: [email protected]

I, Ignacy (Tony) Lipiec (P.Eng.) am employed as a Principal Metallurgical Engineer with AMEC Americas Limited. This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British



I am a Professional Engineer in the province of British Columbia. I graduated from the University of

British Columbia with a B.A.Sc. degree in Mining & Mineral Process Engineering, in 1985. I have

practiced my profession for 27 years, and have previously been involved with metallurgical design

and process engineering for precious metal, base metal and specialty product projects in North

America and South America.



Person responsible for Sections 13, 17, 18, 21 and those portions of the Summary, Interpretation

and Conclusions and Recommendations that pertain to those sections of the Technical Report.

I did not visit the Blue River property.


1.5 of NI 43–101.

I have been involved as an independent consultant with the Blue River Ta-Nb Project since 2010.






Tony Lipiec, P.Eng.

Dated: 4 July 2012



Behrang Omidvar, P.Eng. AMEC Americas Limited



I, Behrang Omidvar, P.Eng., am employed as a Financial Analyst with AMEC Americas Limited.



22 June 2012 (the “Technical Report”).

I am a Professional Engineer in the Province of British Columbia (P.Eng. #35500). I have graduated

from The University of British Columbia with a Mechanical Engineering degree.

I have eight years of experience in engineering, project management and financial analysis for

mining and other projects. I have prepared cash-flow models and conducted financial and

throughput analyses of numerous mines and development properties in Canada and internationally.



Person responsible for Sections 19, 21, 22 and those portions of the Summary, Interpretation and

Conclusions and Recommendations that pertain to those Sections of the Technical Report.

I have not visited the Blue River property.


1.5 of NI 43–101.

I have been involved as an independent consultant with the Blue River Ta-Nb Project since 2010.






Behrang Omidvar, P.Eng.

Dated: 4 July 2012


Blue River Tantalum–Niobium Project

British Columbia, Canada


Project No.: 168967 TOC i 22 June 2012

C O N T E N T S

1.0 SUMMARY ................................................................................................................................... 1-1 1.1 Terms of Reference ......................................................................................................... 1-1 1.2 Key Outcomes ................................................................................................................. 1-1

1.2.1 Mineral Resource Update ............................................................................... 1-1 1.2.2 2011 PEA Outcomes ...................................................................................... 1-2

1.3 Project Setting ................................................................................................................. 1-2 1.4 Tenure, Surface Rights, Royalties, and Agreements ...................................................... 1-2 1.5 Environment, Permitting, and Socio-Economics ............................................................. 1-3 1.6 Geology and Mineralization ............................................................................................. 1-3 1.7 Exploration ....................................................................................................................... 1-4 1.8 Exploration Potential ........................................................................................................ 1-4 1.9 Drilling .............................................................................................................................. 1-4 1.10 Sample Preparation, Analysis, and Security ................................................................... 1-5 1.11 Data Verification .............................................................................................................. 1-6 1.12 Metallurgical Testwork ..................................................................................................... 1-7 1.13 Mineral Resource Estimation........................................................................................... 1-8 1.14 Mineral Resource Statement ........................................................................................... 1-9 1.15 Preliminary Economic Assessment ............................................................................... 1-10

1.15.1 2011 PEA ...................................................................................................... 1-10 1.15.2 Proposed Mining Method .............................................................................. 1-10 1.15.3 Geotechnical Considerations ........................................................................ 1-10 1.15.4 Dilution Considerations ................................................................................. 1-11 1.15.5 Drilling and Blasting ...................................................................................... 1-11 1.15.6 Mine Development ........................................................................................ 1-11 1.15.7 Mineralized Material and Waste Haulage ..................................................... 1-12 1.15.8 Mine Services ............................................................................................... 1-12 1.15.9 Mine Production Forecasts ........................................................................... 1-13 1.15.10 Process Design ............................................................................................ 1-13 1.15.11 Tailings and Waste Management ................................................................. 1-14 1.15.12 Planned Project Infrastructure ...................................................................... 1-14 1.15.13 Markets ......................................................................................................... 1-15 1.15.14 Capital Costs ................................................................................................ 1-16 1.15.15 Operating Costs ............................................................................................ 1-17 1.15.16 Financial Analysis ......................................................................................... 1-18 1.15.17 Sensitivity Analysis ....................................................................................... 1-20

1.16 Interpretation and Conclusions ...................................................................................... 1-21 1.16.1 2012 Mineral Resource Estimate Update ..................................................... 1-21 1.16.2 2011 PEA ...................................................................................................... 1-22 1.16.3 Project Opportunities .................................................................................... 1-23 1.16.4 Project Risks ................................................................................................. 1-23

1.17 Recommendations ......................................................................................................... 1-24

2.0 INTRODUCTION .......................................................................................................................... 2-1 2.1 Terms of Reference ......................................................................................................... 2-1 2.2 Qualified Persons ............................................................................................................ 2-1 2.3 Site Visits and Scope of Personal Inspection .................................................................. 2-1 2.4 Effective Dates ................................................................................................................ 2-2 2.5 Information Sources and References .............................................................................. 2-3





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2.6 Previous Technical Reports............................................................................................. 2-4

3.0 RELIANCE ON OTHER EXPERTS .............................................................................................. 3-1 3.1 Mineral Tenure ................................................................................................................ 3-1 3.2 Surface Rights ................................................................................................................. 3-1 3.3 Royalties and Agreements .............................................................................................. 3-1 3.4 Environmental, Permitting, and Liability Issues ............................................................... 3-2 3.5 Markets ............................................................................................................................ 3-2

4.0 PROPERTY DESCRIPTION AND LOCATION ............................................................................ 4-1 4.1 Project Ownership ........................................................................................................... 4-1 4.2 Mineral Tenure ................................................................................................................ 4-1 4.3 Surface Rights ................................................................................................................. 4-4 4.4 Royalties and Agreements .............................................................................................. 4-4 4.5 Permits ............................................................................................................................. 4-4 4.6 Environment .................................................................................................................... 4-4 4.7 Social and Community Impact ......................................................................................... 4-4 4.8 Comment on Section 4 .................................................................................................... 4-5

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY ......................................................................................................................... 5-1 5.1 Accessibility ..................................................................................................................... 5-1 5.2 Climate ............................................................................................................................. 5-1 5.3 Local Resources and Infrastructure ................................................................................ 5-2 5.4 Physiography ................................................................................................................... 5-2 5.5 Comment on Section 5 .................................................................................................... 5-3

6.0 HISTORY ...................................................................................................................................... 6-1 6.1 Pre-Commerce Exploration ............................................................................................. 6-1 6.2 Commerce Exploration .................................................................................................... 6-1 6.3 Commerce Mineral Resource Estimates ......................................................................... 6-2

7.0 GEOLOGICAL SETTING AND MINERALIZATION ..................................................................... 7-1 7.1 Regional Geology ............................................................................................................ 7-1 7.2 Project Geology ............................................................................................................... 7-3

7.2.1 Metasedimentary Rocks ................................................................................. 7-3 7.2.2 Gneisses and Schists ..................................................................................... 7-5 7.2.3 Amphibolites ................................................................................................... 7-6 7.2.4 Intrusive Rocks ............................................................................................... 7-6 7.2.5 Pegmatite Dykes .......................................................................................... 7-14

7.3 Structural Geology and Metamorphism ......................................................................... 7-14 7.4 Geochronology .............................................................................................................. 7-16 7.5 Carbonatites .................................................................................................................. 7-16

7.5.1 Fir Carbonatite .............................................................................................. 7-16 7.5.2 Verity Carbonatite ......................................................................................... 7-17 7.5.3 Exploration Targets ...................................................................................... 7-18

7.6 Mineralogy ..................................................................................................................... 7-20 7.6.1 Ferrocolumbite .............................................................................................. 7-20 7.6.2 Pyrochlore ..................................................................................................... 7-21 7.6.3 Fersmite ........................................................................................................ 7-21 7.6.4 Fenite Mineralization .................................................................................... 7-21 7.6.5 Mineral Zoning .............................................................................................. 7-21





Project No.: 168967 TOC iii 22 June 2012

7.7 Comment on Section 7 .................................................................................................. 7-22

8.0 DEPOSIT TYPES ......................................................................................................................... 8-1 8.1 Comment on Section 8 .................................................................................................... 8-3

9.0 EXPLORATION ............................................................................................................................ 9-1 9.1 Grids and Surveys ........................................................................................................... 9-1 9.2 Geological Mapping ......................................................................................................... 9-1 9.3 Geochemical Sampling .................................................................................................... 9-1

9.3.1 Stream Sediment Sampling ............................................................................ 9-1 9.3.2 Soil Sampling .................................................................................................. 9-2 9.3.3 Rock Chip, Grab, and Channel Sampling ...................................................... 9-3

9.4 Bulk Sampling .................................................................................................................. 9-4 9.5 Research Programs ......................................................................................................... 9-4 9.6 Comment on Section 9 .................................................................................................... 9-5

10.0 DRILLING ................................................................................................................................... 10-1 10.1 Core Drilling Strategy .................................................................................................... 10-4

10.1.1 Core Sizes .................................................................................................... 10-4 10.1.2 Collar Surveys .............................................................................................. 10-4 10.1.3 Down-Hole Surveys ...................................................................................... 10-5 10.1.4 Oriented Drill Core ........................................................................................ 10-5 10.1.5 Core Handling ............................................................................................... 10-5 10.1.6 Core Recovery .............................................................................................. 10-5

10.2 Drill Intercepts ................................................................................................................ 10-6 10.3 Comment on Section 10 ................................................................................................ 10-6

11.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY ........................................................ 11-1 11.1 Sampling Methods ......................................................................................................... 11-1 11.2 Metallurgical Sampling .................................................................................................. 11-2 11.3 Density Determinations ................................................................................................. 11-2

11.3.1 Density Check Program ................................................................................ 11-4 11.4 Analytical Laboratories .................................................................................................. 11-4 11.5 Sample Preparation and Analysis ................................................................................. 11-5 11.6 Quality Assurance and Quality Control ......................................................................... 11-5

11.6.1 Assessment of Precision .............................................................................. 11-6 11.6.2 Assessment of Accuracy ............................................................................ 11-12 11.6.3 Assessment of Laboratory Bias .................................................................. 11-16 11.6.4 Assessment of Contamination .................................................................... 11-18 11.6.5 Assay QA/QC Conclusions ......................................................................... 11-24

11.7 Databases ................................................................................................................... 11-25 11.8 Security ........................................................................................................................ 11-25 11.9 Comment on Section 11 .............................................................................................. 11-26

12.0 DATA VERIFICATION ................................................................................................................ 12-1 12.1 Database Verification .................................................................................................... 12-1 12.2 Site Visits ....................................................................................................................... 12-1

12.2.1 Drill Collar Location Check ........................................................................... 12-2 12.2.2 Inspection of Drill Core and Verification of Mineralization ............................ 12-2

12.3 Comment on Section 12 ................................................................................................ 12-4

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................................................. 13-1 13.1 Head Samples for Initial Testing ................................................................................... 13-2





Project No.: 168967 TOC iv 22 June 2012

13.2 Phase I Testing .............................................................................................................. 13-3 13.2.1 Grinding Size ................................................................................................ 13-3 13.2.2 Roughing and Cleaning Gravity Concentration ............................................ 13-3

13.3 Phase II Testing ............................................................................................................. 13-6 13.3.1 Flotation Tests .............................................................................................. 13-6

13.4 Phase III Testing ............................................................................................................ 13-8 13.4.1 2011 and 2012 Work .................................................................................... 13-8 13.4.2 Review of Concentrate Treatment Options .................................................. 13-8

13.5 Accuracy of Assaying .................................................................................................... 13-9 13.6 Comment on Section 13 ................................................................................................ 13-9

14.0 MINERAL RESOURCE ESTIMATES ......................................................................................... 14-1 14.1 Introduction .................................................................................................................... 14-1 14.2 Assay Data and Capping ............................................................................................... 14-1 14.3 Composites .................................................................................................................... 14-1 14.4 Exploratory Data Analysis ............................................................................................. 14-2 14.5 Contact Analysis ............................................................................................................ 14-4 14.6 Variography ................................................................................................................... 14-7 14.7 Carbonatite Solid Modeling ........................................................................................... 14-7 14.8 Block Model Dimensions ............................................................................................... 14-7 14.9 Assignment of Lithology and Specific Gravity to Blocks ............................................... 14-8 14.10 Block Model Grade Estimate ......................................................................................... 14-8 14.11 Block Model Validation .................................................................................................. 14-9

14.11.1 Visual Validation ........................................................................................... 14-9 14.11.2 Global Grade Bias Check ........................................................................... 14-12 14.11.3 Local Grade Bias Check (Swath Plots) ...................................................... 14-12 14.11.4 Selectivity Check ........................................................................................ 14-14

14.12 In Situ Block Model Carbonatite Reconciliation........................................................... 14-16 14.13 Mineral Resource Classification .................................................................................. 14-16 14.14 Reasonable Prospects for Economic Extraction ......................................................... 14-19

14.14.1 Market Study............................................................................................... 14-19 14.14.2 Commodity Price ........................................................................................ 14-19 14.14.3 Physical Assumptions ................................................................................. 14-19 14.14.4 Operational Considerations ........................................................................ 14-19 14.14.5 Economic Assumptions .............................................................................. 14-20 14.14.6 Economic Cut-Off ....................................................................................... 14-20

14.15 Mineral Resource Statement ....................................................................................... 14-21 14.16 Comparison of Mineral Resources .............................................................................. 14-23 14.17 Comment on Section 14 .............................................................................................. 14-24

15.0 MINERAL RESERVE ESTIMATE .............................................................................................. 15-1

16.0 MINING METHODS .................................................................................................................... 16-1 16.1 Introduction .................................................................................................................... 16-1 16.2 Optimization ................................................................................................................... 16-1

16.2.1 Assumptions ................................................................................................. 16-1 16.2.2 Mining Method .............................................................................................. 16-2 16.2.3 Mineral Resources considered for the 2011 PEA ........................................ 16-3 16.2.4 Production Rate ............................................................................................ 16-4

16.3 Geotechnical Conditions ................................................................................................ 16-4 16.4 Conceptual Mining Method ............................................................................................ 16-5





Project No.: 168967 TOC v 22 June 2012

16.4.1 Backfill Considerations ................................................................................. 16-5 16.5 Stoping Design .............................................................................................................. 16-6

16.5.1 Stability Analysis and Ground Support ......................................................... 16-6 16.5.2 Stope Geometry ........................................................................................... 16-7 16.5.3 Mining Sequence .......................................................................................... 16-7 16.5.4 Conceptual Mine Design .............................................................................. 16-7 16.5.5 Mining Dilution and Recovery ....................................................................... 16-7

16.6 Drilling and Blasting ....................................................................................................... 16-7 16.7 Mine Development ......................................................................................................... 16-8 16.8 Mineralized Material and Waste Rock Haulage ............................................................ 16-9 16.9 Mine Services .............................................................................................................. 16-12 16.10 Mine Development and Production Forecasts ............................................................ 16-12 16.11 Mine Equipment Requirements ................................................................................... 16-14 16.12 Mine Infrastructure ....................................................................................................... 16-14 16.13 Mining Personnel ......................................................................................................... 16-14 16.14 Comment on Section 16 .............................................................................................. 16-14

17.0 RECOVERY METHODS ............................................................................................................ 17-1 17.1 Introduction .................................................................................................................... 17-1 17.2 Plant Design .................................................................................................................. 17-1 17.3 Comminution (Crushing, Storage, and Grinding) .......................................................... 17-2 17.4 De-Sliming and Flotation ............................................................................................... 17-3 17.5 Filtration ......................................................................................................................... 17-3 17.6 Concentrate Pre-Treatment ........................................................................................... 17-3 17.7 Chlorination and Distillation ........................................................................................... 17-4 17.8 Product / Materials Handling ......................................................................................... 17-4 17.9 Energy, Water, and Process Materials Requirements .................................................. 17-4 17.10 Comment on Section 17 ................................................................................................ 17-4

18.0 PROJECT INFRASTRUCTURE ................................................................................................. 18-1 18.1 Introduction .................................................................................................................... 18-1 18.2 Site Layout ..................................................................................................................... 18-1 18.3 Buildings ........................................................................................................................ 18-1

18.3.1 Mine Service Building ................................................................................... 18-1 18.3.2 Truck Shop ................................................................................................... 18-3 18.3.3 Warehouse ................................................................................................... 18-3 18.3.4 Process Building ........................................................................................... 18-3 18.3.5 Crushing and Screening Circuit .................................................................... 18-3 18.3.6 Portal Infrastructure ...................................................................................... 18-4 18.3.7 Explosives Storage ....................................................................................... 18-4 18.3.8 Aggregate Crushing and Concrete Batch Plants .......................................... 18-4

18.4 Roads and Logistics ...................................................................................................... 18-4 18.4.1 Access Road................................................................................................. 18-4 18.4.2 Haul Road ..................................................................................................... 18-5

18.5 Co-Disposal Storage Facilities ...................................................................................... 18-5 18.5.1 Drystack Considerations ............................................................................... 18-5 18.5.2 Evaluation of Potential Sites ......................................................................... 18-6 18.5.3 Site Selection ................................................................................................ 18-7 18.5.4 Facility Design .............................................................................................. 18-8 18.5.5 Co-Disposal Facility Geohazards Considerations ........................................ 18-9 18.5.6 Co-Disposal Facility Stability Considerations ............................................... 18-9





Project No.: 168967 TOC vi 22 June 2012

18.5.7 Co-Disposal Facility Surface Water Run-Off Considerations ..................... 18-10 18.5.8 Co-Disposal Facility Closure Considerations ............................................. 18-11

18.6 Avalanche Hazard ....................................................................................................... 18-11 18.7 Water Supply, Distribution, and Treatment Systems .................................................. 18-11 18.8 Waste Considerations ................................................................................................. 18-12 18.9 Accommodation ........................................................................................................... 18-12 18.10 Power and Electrical .................................................................................................... 18-12 18.11 Fuel .............................................................................................................................. 18-13 18.12 Comment on Section 18 .............................................................................................. 18-13

19.0 MARKET STUDIES AND CONTRACTS .................................................................................... 19-1 19.1 Introduction .................................................................................................................... 19-1 19.2 2011 PEA Market Studies ............................................................................................. 19-1 19.3 2011 PEA Commodity Price .......................................................................................... 19-1

19.3.1 Tantalum ....................................................................................................... 19-1 19.3.2 Niobium ......................................................................................................... 19-2

19.4 Price Assumption Discussion ........................................................................................ 19-2 19.5 Comment on Section 19 ................................................................................................ 19-4

20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ....... 20-1 20.1 Environmental Assessment for Mining Projects ............................................................ 20-1 20.2 Project Studies .............................................................................................................. 20-2 20.3 Environmental Setting and Review of Environmental Baseline .................................... 20-3 20.4 Closure Considerations ................................................................................................. 20-9 20.5 Current Environmental Liabilities ................................................................................. 20-10 20.6 2011 PEA Closure Plan ............................................................................................... 20-10 20.7 Permitting..................................................................................................................... 20-10 20.8 Considerations of Social and Community Impacts ...................................................... 20-12

20.8.1 First Nations ................................................................................................ 20-13 20.8.2 Local Communities ..................................................................................... 20-14

20.9 Comment on Section 20 .............................................................................................. 20-14

21.0 2011 PEA CAPITAL AND OPERATING COSTS ....................................................................... 21-1 21.1 2011 PEA Basis of Estimate ......................................................................................... 21-1 21.2 2011 PEA Capital Costs ................................................................................................ 21-1

21.2.1 Infrastructure................................................................................................. 21-1 21.2.2 Material Handling .......................................................................................... 21-2 21.2.3 Process Plant................................................................................................ 21-2 21.2.4 Mining ........................................................................................................... 21-2 21.2.5 Contingency Costs ....................................................................................... 21-3 21.2.6 Indirect Costs ................................................................................................ 21-3 21.2.7 Sustaining Capital ......................................................................................... 21-4 21.2.8 Mine Closure................................................................................................. 21-4 21.2.9 Capital Cost Estimate Summary ................................................................... 21-4 21.2.10 2011 PEA Operating Costs .......................................................................... 21-5 21.2.11 Capital and Operating Cost Discussion ........................................................ 21-6


22.0 2011 PEA ECONOMIC ANALYSIS ............................................................................................ 22-1 22.1 2011 PEA Valuation Method ......................................................................................... 22-1 22.2 2011 PEA Financial Model Parameters ........................................................................ 22-2

22.2.1 Mineral Resources and Mine Life ................................................................. 22-2





Project No.: 168967 TOC vii 22 June 2012

22.2.2 Metallurgical Process ................................................................................... 22-2 22.2.3 Commodity Prices and Foreign Exchange ................................................... 22-2 22.2.4 Taxes ............................................................................................................ 22-2 22.2.5 PEA Financial Results .................................................................................. 22-4 22.2.6 2011 PEA Cash Costs .................................................................................. 22-4 22.2.7 2011 PEA Sensitivity Analysis ...................................................................... 22-6 22.2.8 Financial Analysis Discussion ...................................................................... 22-7


23.0 ADJACENT PROPERTIES ........................................................................................................ 23-1

24.0 OTHER RELEVANT DATA AND INFORMATION ..................................................................... 24-1

25.0 INTERPRETATION AND CONCLUSIONS ................................................................................ 25-1 25.1 Mineral Resource Update (Effective Date 22 June 2012) ............................................. 25-1 25.2 2011 PEA ...................................................................................................................... 25-2

25.2.1 Opportunities ................................................................................................ 25-3 25.2.2 Risks ............................................................................................................. 25-4

26.0 RECOMMENDATIONS .............................................................................................................. 26-1

27.0 REFERENCES ........................................................................................................................... 27-1

T A B L E S

Table 1-1: Blue River Project Estimated Mineral Resources; Effective Date 22 June 2012,

Tomasz Postolski, P.Eng, Qualified Person ...................................................................... 1-9 Table 1-2: Summary of Estimated Capital Costs .............................................................................. 1-17 Table 1-4: Summary Financial Analysis at Various Discount Rates (base case is highlighted) ....... 1-19 Table 2-1: Site Visit and Areas of Report Responsibilities .................................................................. 2-2 Table 6-1: Blue River Exploration History Summary ........................................................................... 6-1 Table 9-1: Soil Sample Campaigns ..................................................................................................... 9-2 Table 9-2: Rock Sample Campaigns ................................................................................................... 9-3 Table 10-1: Drill Campaign Summary ................................................................................................. 10-2 Table 10-2: Upper Fir Deposit Trench and Bulk Samples ................................................................... 10-2 Table 10-3: Example Drill Hole Intercept Summary Table .................................................................. 10-7 Table 11-1: 2005 – 2010 Specific Gravity Determinations by Campaign ........................................... 11-3 Table 11-2: 2005 – 2010 Specific Gravity Constants .......................................................................... 11-3 Table 11-3 Control Sample Insertion Rate Summary ........................................................................ 11-7 Table 11-4: Cumulative Frequency ARD Summary for Tantalum ....................................................... 11-8 Table 11-5: Cumulative Frequency ARD Summary for Niobium ........................................................ 11-9 Table 11-6: Cumulative Frequency ARD Summary for Tantalum (Mean > than 50 ppm Ta) ............. 11-9 Table 11-7: 2010 Nb XRF(F) Blue River SRM Control Chart Summary ........................................... 11-16 Table 11-8: 2010 Ta XRF(F) Blue River SRM Control Chart Summary ........................................... 11-16 Table 11-9: Pulp Check Between-Laboratory Bias ........................................................................... 11-17 Table 12-1: AMEC Site Visit Confirmation of Mineralization ............................................................... 12-3 Table 13-1: Head Assay Grades, Bulk Samples BS-2F and BS-2G ................................................... 13-3 Table 13-2: Results from F81 .............................................................................................................. 13-7 Table 13-3: Results of a Sequential Hydrochloric Acid Leach of Flotation “Middling” ........................ 13-8





Project No.: 168967 TOC viii 22 June 2012

Table 14-1: Capped Assays vs. 2.5 m Composites Statistics inside Carbonatites ............................. 14-1 Table 14-2: Composite Statistics in Carbonatite ................................................................................. 14-2 Table 14-3: Ta2O5 and Nb2O5 Correlogram Parameters in Carbonatite ............................................. 14-7 Table 14-4: Block Model Dimensions .................................................................................................. 14-8 Table 14-5: Estimation Parameters for Ta2O5 and Nb2O5 ................................................................... 14-9 Table 14-6: Mean Grades for NN and ID3 Models ............................................................................ 14-12 Table 14-7: Blue River Project Estimated Mineral Resources; Effective Date 22 June, 2012,

Tomasz Postolski, P.Eng, Qualified Person .................................................................. 14-22 Table 14-8: Blue River Project Sensitivity of Estimated Mineral Resources to Tantalum Price;

Effective Date 22 June 2012, Tomasz Postolski, P.Eng, Qualified Person ................... 14-23 Table 16-1: Minimum Stope Dimensions for Constraining the Subset of Mineral Resources within

Designed Stopes .............................................................................................................. 16-2 Table 16-2: Blue River Project Estimated Mineral Resources Supporting 2011 PEA; Effective Date

29 September 2011, Tomasz Postolski, P.Eng., Qualified Person .................................. 16-3 Table 16-3: Rock Mass Characteristics by Rock Group ..................................................................... 16-4 Table 16-4: Major Joint Sets ............................................................................................................... 16-5 Table 16-5: Stope Faces and Hydraulic Radius .................................................................................. 16-6 Table 16-6: Mine Development and Production Forecasts ............................................................... 16-13 Table 16-7: Mining and Tailings Facility Equipment Requirements .................................................. 16-15 Table 16-8: Mining Personnel Requirements .................................................................................... 16-16 Table 20-1: Provincial Permits, Approvals, Licences, and Authorizations ........................................ 20-11 Table 20-2: Federal Permits, Approval, Licences, and Authorizations ............................................. 20-11 Table 21-1: Summary of Estimated Capital Costs (CAD, 2011 constant dollars) ............................... 21-5 Table 21-2: Average Life-of-Mine Operating Cost Summary (CAD, 2011 constant dollars) .............. 21-5 Table 22-1: Summary Financial Analysis at Various Discount Rates ................................................. 22-4 Table 22-3: Sensitivity Summary in CAD, 8% Discount Rate ............................................................. 22-6 Table 26-1: Recommendations Summary ........................................................................................... 26-1

F I G U R E S

Figure 1-1: Sensitivity Summary, 8% Discount Rate .......................................................................... 1-21 Figure 4-1: Project Location Map ......................................................................................................... 4-2 Figure 4-2: Blue River Mineral Tenure Map ......................................................................................... 4-3 Figure 7-1: Tectonic Belts of British Columbia and Carbonatite Occurrences ..................................... 7-2 Figure 7-2: Blue River Project Local Geology Map .............................................................................. 7-4 Figure 7-3: Blue River Local Geology Legend (for Figure 7-2) ............................................................ 7-5 Figure 7-4: Deposit Area Surface Geology Map .................................................................................. 7-7 Figure 7-5: Drill Collar and Vertical Section Locations ......................................................................... 7-8 Figure 7-6: Longitudinal Section A – A’ (view SE) ................................................................................ 7-9 Figure 7-7: Geology Section 5796740 N ............................................................................................ 7-10 Figure 7-8: Geology Section 5796425 N ............................................................................................ 7-11 Figure 7-9: Fold Indicators (Hole F08-150: 121.8 m to 129.8 m) ....................................................... 7-13 Figure 7-10: Fold Indicators (Hole F08-150: 143.5 m and 147.0 m) .................................................... 7-13 Figure 7-11: Fold Indicators (Hole F08-151: 204.0 m to 204.5 m) ....................................................... 7-14 Figure 7-12: Exploration Target Location Surface Map ........................................................................ 7-19





Project No.: 168967 TOC ix 22 June 2012

Figure 10-1: Drill Collar Plan ................................................................................................................ 10-3 Figure 11-1: 2005 to 2008 Quarter-Core Duplicate Pair Cumulative Frequency ARD Chart ............... 11-8 Figure 11-2: 2010 Drill Core Assay Program Cumulative Frequency ARD Chart .............................. 11-10 Figure 11-3: 2009 Re-assay Program Ta XRF(F) Cumulative Frequency ARD Chart ...................... 11-11 Figure 11-4: 2009 Re-assay Program Nb XRF(F) Cumulative Frequency ARD Chart ...................... 11-11 Figure 11-5: 2005 to 2008 Ta ICP-MS BR-01 SRM Control Chart .................................................... 11-13 Figure 11-6: 2005 to 2008 Nb ICP-MS BR-01 SRM Control Chart .................................................... 11-13 Figure 11-7: 2009 Ta XRF(F) Blue River SRMs Control Chart\ ......................................................... 11-14 Figure 11-8: 2009 Nb XRF(F) Blue River SRMs Control Chart .......................................................... 11-15 Figure 11-9: 2010 Acme versus Stark Nb XRF(F) Check Pair RMA Chart ........................................ 11-18 Figure 11-10: 2005 – 2008 Blank Ta ICP-MS Performance Chart ...................................................... 11-19 Figure 11-11: 2005 - 2008 Blank Nb ICP-MS Performance Chart .................................................... 11-20 Figure 11-12: 2009 Ta XRF(F) Blank Performance Chart .................................................................... 11-21 Figure 11-13: 2009 Nb XRF(F) Blank Performance Chart ................................................................... 11-22 Figure 11-14: 2010 Ta XRF(F) Blank Performance Chart .................................................................... 11-23 Figure 11-15: 2010 Nb XRF(F) Blank Performance Chart ................................................................... 11-24 Figure 13-1: Sample BS-2F – Gravity Separation (Different Grinds) ................................................... 13-4 Figure 13-2: Sample BS-2G – Gravity Separation (Different Grinds) .................................................. 13-5 Figure 13-3: Overall Rougher and Cleaner Recovery vs Grade by Centrifugal Gravity Concentration 13-5 Figure 13-4: Upgrading by Wilfley and Mozley Units ........................................................................... 13-6 Figure 14-1: Ta2O5 Histograms and Probability Plot within Carbonatite .............................................. 14-3 Figure 14-2: Nb2O5 Histograms and Probability Plot within Carbonatite .............................................. 14-4 Figure 14-3: Ta2O5 Contact Plots between Carbonatite and Fenite ..................................................... 14-5 Figure 14-4: Nb2O5 Contact Plots between Carbonatite and Fenite .................................................... 14-6 Figure 14-5: Ta2O5 ID3 Model within Carbonatite – Plan 1,146.25 .................................................... 14-10 Figure 14-6: Ta2O5 ID3 Model within Carbonatite – Section N 5,796,932.5 ...................................... 14-10 Figure 14-7: Nb2O5 ID3 Model within Carbonatite – Plan 1,146.25 ................................................... 14-11 Figure 14-8: Nb2O5 ID3 Model within Carbonatite – Section N 5,796,932.5 ...................................... 14-11 Figure 14-9: Swath Plot for Ta2O5 ID3 Model ..................................................................................... 14-13 Figure 14-10: Swath Plot for Nb2O5 ID3 Model ..................................................................................... 14-13 Figure 14-11: Herco Grade – Tonnage Curves for Ta2O5 ID3 Model ................................................... 14-15 Figure 14-12: Herco Grade – Tonnage Curves for Nb2O5 ID3 Model .................................................. 14-15 Figure 14-13: Resource Classification – Plan 1,161.25 ....................................................................... 14-18 Figure 14-14: Resource Classification – Section N 5,796,882.5 .......................................................... 14-18 Figure 16-1: Conceptual Mine Layout Plan (plan view projection) ..................................................... 16-10 Figure 16-2: Aerial View of the Mining Area from Upper Portal ......................................................... 16-11 Figure 17-1: Concentration and Refining of Blue River Mineralization ................................................ 17-2 Figure 18-1: Proposed Site Layout Plan ............................................................................................... 18-2 Figure 19-1: Ta Price Trend ................................................................................................................. 19-3 Figure 19-2: Nb Price Trend ................................................................................................................. 19-4 Figure 22-1: Sensitivity Summary in CAD, 8% Discount Rate ............................................................. 22-7





Project No.: 168967 TOC x 22 June 2012

A P P E N D I C E S

A p p e n d i x A : List of Claims





Project No.: 168967 Page 1-1 22 June 2012

1.0 SUMMARY

1.1 Terms of Reference

AMEC Americas Limited (AMEC) was commissioned by Commerce Resources

Corporation (Commerce) to prepare a NI 43-101 compliant Mineral Resource update

and technical report (the Report) on the wholly-owned Blue River tantalum–niobium

Project (the Project), located within the North Thompson River valley of east–central

British Columbia (B.C.), Canada.

This technical report supports the findings of the Mineral Resource update and also

includes summaries from a Preliminary Economic Assessment study completed on the

Blue River Project with an effective date 29 September 2011 (2011 PEA). Results

from the 2011 PEA mining studies have not changed in terms of their outcomes as

their underlying assumptions remain reasonable.

1.2 Key Outcomes

1.2.1 Mineral Resource Update

The key findings of the Mineral Resource update (effective date 22 June 2012) are

summarized as follows:

Indicated Category: 51.8 million tonnes @ 192 ppm Ta2O5 and 1,490 ppm Nb2O5

Inferred Category: 8.8 million tonnes @ 186 ppm Ta2O5 and 1,660 ppm Nb2O5

The Mineral Resource update uses the same assumptions from the 2011 PEA for the

following items:

Ta and Nb metal prices

Mining method and mining extraction factor

Processing method and recovery factor

CAPEX and OPEX costs

Block Unit Value cut-off values of US$40/t for the bulk mining method and US$58/t

for the selective mining method.

The Mineral Resources have significantly increased in tonnage mostly due to a

reduction in the block unit value cut-off by eliminating back-fill costs and, to a lesser

extent, additional infill diamond drilling.






1.2.2 2011 PEA Outcomes

From the 2011 PEA, the following work and outcomes are considered to remain

reasonable as their underlying assumptions have not changed.

Estimated Internal rate of return: 9.1% (before tax)

Estimated Net present value: CAD$18.5 million at 8% discount rate (before tax)

Estimated Payback: 6.3 years

Average diluted grade in the conceptual mine plan to the mill:

185 ppm Ta2O5 and 1,591 ppm Nb2O5

Conceptual Operating cost: CAD$38.44/t milled (mining ~ 55% of operating cost)

Conceptual Capital cost: CAD$379 million (process ~ 31% of initial capital cost)

Proposed product: High purity Ta and Nb chloride product that is suitable for

several markets

Conceptual Mine Life: 10 years based upon the mineral resources (effective date

29 September 2011) defined using information to the end of 2009 drilling

NPV Sensitivity: The Upper Fir deposit is most sensitive to changes in exchange

rate, commodity prices, and mining costs

The above key outcomes contain forward looking information. The assumptions and

risks regarding those assumptions are summarized and explained in more detail in

Sections 1.15, and 1.16 of the Report.

1.3 Project Setting

The Blue River Project is situated 250 km north of the city of Kamloops, approximately

90 km south of the town of Valemount and 23 km north of the community of Blue

River, in the North Thompson River valley of east–central British Columbia. The

property is accessed from BC Highway 5 (Yellowhead Highway) via a 4 km

well-groomed gravel road. Within the Project area, access is by forestry service and

logging roads or by helicopter.

1.4 Tenure, Surface Rights, Royalties, and Agreements

The Project comprises 249 two-post claim, four-post claim, and mineral cell title

submission (MCX) claims in good standing that encompass just over 1,000 km2

(105,195 ha) within the Kamloops Mining Division. These claims are wholly-owned by

Commerce. Currently, all of the mineral claims are valid until 31 March 2021. All but

two of the mineral claims are on Crown land.






There are surveyed parcels along the western edge of the property with surface rights

held by other parties which overlap the property mineral tenure claims. Commerce is

not aware of any material issues that would prevent negotiation for access or surface

rights of these surveyed parcels should they be required in the future. The overlapped

claims do not host mineral resources, and currently no carbonatites are known within

these claims.

There are no known royalties, back-in rights, agreements, or encumbrances attributed

to the claims.

1.5 Environment, Permitting, and Socio-Economics

Commerce has been pro-active with regard to environmental and socioeconomic

issues. Environmental monitoring, baseline studies and site investigations have been

ongoing at the Blue River Project site since the summer season of 2006. Kinetic test

work for acid rock drainage and metals leaching was initiated in 2010. Additional

environmental baseline programs are expected to continue, as required through 2012.

First Nations engagement, with respect to exploration activities, began in 2007, and

will continue for the duration of the Project. The Blue River Project lies on lands which

comprise part of the traditional territory of the Simpcw First Nation. First Nations

engagement, with respect to exploration activities, began in 2007. Public engagement

to date has included meetings with local councils and informal discussions with local

land-owners.

1.6 Geology and Mineralization

The Blue River deposit is hosted within polyfolded carbonatite rocks. The carbonatites

intrude Late Proterozoic supracrustal rocks which lie on the north-eastern margin of

the Shuswap Metamorphic Complex within the Omineca terrane. The Blue River

carbonatites are hosted in the Mica Creek assemblage of the Horsethief Creek Group.

Two units of the Mica Creek assemblage underlie much of the Project area.

Carbonatites were emplaced as dikes or sills into the metasedimentary host rocks prior

to regional deformation and metamorphism. Regional deformation has folded the

carbonatite and its host rocks. Contacts between carbonatite and the host

metasediments are typically sharp and mantled by zones of metasomatized host rock,

known as fenite. The carbonatite has average thicknesses of 30 m, ranging between

5 m to about 90 m thick, and with strike lengths ranging between 50 m to 1,100 m.

Both dolomitic carbonatites and calcitic carbonatite occur at Blue River.






Mineralization comprises niobium- and tantalum-bearing minerals that have

crystallized in carbonatite by primary magmatic concentration and in fenite. Primary

economic minerals, with their generic end-member formulae, are ferrocolumbite

((Fe,Mn,Mg)(Nb,Ta)2O6,) and pyrochlore ((Ca,Na,U)2(Nb,Ti,Ta)2O6(OH,F)).

In the opinion of the QPs, knowledge of the deposit setting, lithologies, structural and

alteration controls on mineralization, and mineralization style are sufficient to support

mineral resource estimation.

1.7 Exploration

The Blue River area has been the subject of intermittent exploration since the

discovery of vermiculite-bearing carbonatite rock in 1949. Since Project acquisition in

2000, Commerce has completed surface mapping, trenching, soil, rock chip, grab and

channel sampling, core drilling, metallurgical testing, bulk sampling, environmental

baseline studies, mineral resource estimation, and a PEA on the Project.

In the opinion of the QPs, the exploration programs completed to date are appropriate

to the style of the deposits and prospects within the Project. The exploration and

research work supports the genetic and affinity interpretations.

1.8 Exploration Potential

The Upper Fir carbonatite has exploration potential directly northward of known

deposit extents based on soil sample results. Additional resource definition drilling is

warranted.

The Bone Creek and Fir carbonatites have additional exploration potential along and

across strike, based on soil sample anomalies. Additional in-fill soil sampling is

warranted prior to diamond drilling to assess for potential connections with the Upper

Fir carbonatite. In addition, Commerce has identified numerous tantalum-in-soil

anomalies from geochemical programs that require follow up.

In the opinion of the QPs, the Project retains significant exploration potential for

additional carbonatite-hosted tantalum–niobium mineralization.

1.9 Drilling

AMEC received a drilling database from Commerce that had a database closure date

of 29 September 2011. The database comprises a total of 269 core drill holes within






the Upper Fir, Bone Creek and Fir (Lower) carbonatites consisting of 54,065 m of HQ

(63.5 mm) diameter core.

Of the 269 core drill holes, 237 drill holes totalling 50,395 m of HQ diameter core, and

12,736 samples are used to support the Mineral Resource update.

Six geotechnical drill holes comprising 1,271 m of HQ diameter oriented core were

completed during 2010. In addition, optical and acoustic televiewer oriented core

surveys were completed for two 2010 holes, four pre-2010 holes, and eighteen 2011

campaign holes during 2010 and 2011.

Core recovery is very good within the waste and carbonatite rocks (typically >95%).

The only area that may have core recovery issues would be within the fenite rocks

located in the immediate hanging wall to the carbonatite.

Core sampling methods and approaches have been consistent through the 2005 to

2011 drill programs and the protocols are consistent with industry standard. In the

opinion of the QPs, the quantity and quality of the collar, down-hole survey, lithology,

and geotechnical data collected in the exploration and infill drill programs completed by

Commerce are sufficient to support mineral resource estimation.

1.10 Sample Preparation, Analysis, and Security

Drill hole samples were collected from an area approximately 1,600 m north–south by

1,000 m east-west. Average spacing between drill-hole intercepts in the Mineral

Resource area varies from 40 to 50 m.

Commerce regularly collected specific gravity measurements at 3 m core intervals

using a water immersion method. Check sampling from field-collected core samples

was completed by Met Solve Laboratories of Burnaby, B.C. for the 2005 – 2009

campaigns with good correlation to the field measurements recorded in the exploration

database. Check specific gravity determinations for the 2010 campaign have yet to be

completed.

The entire carbonatite intersection and shoulder samples on each side of the

intersection are sampled; samples are typically 1 m in length and geological contacts

are generally respected. Half core is sent for analysis.

Acme Analytical Laboratories (Acme) in Vancouver was the primary laboratory for

sample preparation of the 2005 to 2008 drill core samples. Acme is an independent

mineral testing laboratory registered under ISO 9001.






PRA / Inspectorate Laboratories (Inspectorate) in Richmond, B.C., is the primary

sample preparation laboratory for the 2009 to 2011 drill core samples. Inspectorate is

also an independent mineral testing laboratory that reportedly works to internationally-

recognized standards such as ISO and ASTM. The Inspectorate-Vancouver

laboratory received ISO9001:2000 accreditation in 2006 and 2009.

Acme has been the primary analytical laboratory since 2005 up to and including 2011

drill core samples. In October 2011 the Acme-Vancouver laboratory received formal

approval of its ISO/IEC 17025:2005 accreditation.

Sample preparation for samples that support the Mineral Resource estimate has

followed a similar procedure for all of Commerce’s drill programs. The preparation

procedure is consistent with industry-standard methods for sampling within carbonatite

deposits.

Analyses were completed at Acme Analytical Laboratories. Between 2005 and 2008,

Ta and Nb were analysed by ICP-MS following a lithium metaborate / tetraborate

fusion and nitric acid digestion. Analysis in 2009 and 2010 was by X-Ray fluorescence

methods following a lithium metaborate fusion XRF(F) and ICP methods.

Overall, the drill programs included insertion of blank, duplicate and standard

reference material samples at a rate that meets industry-accepted standards of

insertion rates. AMEC concludes the Blue River sample results show imprecision but

no consistent bias and that the ICP-MS results from 2005 to 2008, and the XRF(F)

results supporting the 2009 and 2010 drilling are suitable for use in mineral resource

estimation. Caution should be applied in assigning a high level of confidence to the

pre-2010 tantalum and niobium analytical results until precision and accuracy issues

are resolved.

Independent data audits have been conducted, and indicate that the sample collection

and database entry procedures are acceptable. Sample security, storage facilities,

and chain of custody procedures are consistent with industry standards.

The QPs are of the opinion that the quality of the specific gravity, tantalum and

niobium analytical data are sufficiently reliable to support mineral resource estimation

and that sample preparation, analysis, and security are generally performed in

accordance with exploration best practices and industry standards.

1.11 Data Verification

Based on site visit inspections, data quality checks, and a minimum 5% database

verification completed by AMEC, the QPs are of the opinion that the collar coordinates,






down-hole surveys, lithologies, and assay data are considered sufficiently free of error

and that the data quality are suitable to support mineral resource estimation.

1.12 Metallurgical Testwork

Testwork began in 2009 and continued into 2010 to develop a process flowsheet for

the Blue River Project. The testwork was based on material produced from two bulk

samples, BS-2F and BS–2G. Mineralogical analysis was performed to obtain

knowledge regarding the occurrence of the tantalum and niobium within the material.

The testwork to date has primarily taken place in three phases:

Phase I – focused on the recovery of the tantalum–niobium minerals by gravity

although grinding and mineralogy investigations were also performed.

Phase II – focused on the recovery and upgrading of the tantalum–niobium

minerals by flotation.

Phase III – continued optimization of the process flowsheet for the production of a

tantalum-niobium mineral concentrate.

Phase I that showed gravity could concentrate the material to a low-grade product, but

that upgrading increasingly gave lower levels of metallurgical recovery as grade was

sought.

Work in Phase II saw the use of flotation concentration technology similar to that being

used for niobium-bearing carbonatites at Iamgold’s Niobec Mine in Quebec, Canada.

There was immediate success in the first phases of the work. Although there are

several stages to the concentration, the overall level of equipment, risk, and complexity

to produce a saleable or treatable concentrate is lower than the gravity route.

In the opinion of the QPs, the following conclusions are applicable:

Tantalum and niobium occur as ferrocolumbite and pyrochlore, which are

amenable to conventional flotation and proven refining processes with estimated

recoveries of 65% to 70%. For the purposes of the financial analysis in Section 22

of this Report, it was assumed that the process plant will have a 65% recovery for

Ta and 69% recovery for Nb in the flotation stage.

The metallurgical testwork has shown that it is possible to collect the tantalum and

niobium minerals into a concentrate suitable for extraction of the metals into

saleable products. The first step of the process uses typical grinding followed by

flotation. There is confidence that the secondary treatment or metal extraction of






the material is possible by an existing method such as aluminothermic reduction

followed by chlorine refining. These results are suitable to support estimation of

mineral resources for the deposits.

1.13 Mineral Resource Estimation

The resource model was constructed inside carbonatite using 237 diamond drill holes

totalling 50,395 m of HQ diameter core and 12,736 samples. Geological

interpretations were provided by Commerce to AMEC in the form of electronic 3D solid

wireframes.

Capped drill core assays were composited down the hole to a fixed length of 2.5 m

respecting lithological boundaries. Exploratory data analysis (EDA) was performed on

the composites. The coefficients of variation are low and support the use of linear

grade interpolation methods such as inverse distance methods.

Blocks within in the model were coded by lithology solids. Specific gravity values were

assigned by lithological unit. Ta2O5 and Nb2O5 grades were estimated in the

carbonatite using an inverse distance to the power of 3 (ID3) interpolation method. A

four-pass interpolation approach was used with each successive pass having greater

search distances.

The block model grades were validated by visual inspection comparing composites to

block grades on-screen, declustered global statistics checks, local biases checks using

swath plots, and finally model selectivity checks. No issues were identified that would

materially affect the Mineral Resource update.

The current mineral resource classification at Blue River is restricted to Indicated or

Inferred based on the following:

Confidence limits drill hole spacing studies

Concerns over analytical precision and provisional accuracy for the sample dataset

from 2005 to 2009

Required metallurgical testwork on the final stage of the proposed metallurgical

process is still ongoing to support proof-of–concept.

To assess reasonable prospects for economic extraction, the updated Mineral

Resources have been constrained using a “Stope Analyzer”. AMEC assumed that the

Blue River deposit would be mined utilizing self-supported, underground bulk mining

methods under a conceptual scenario that considers mining and processing at a rate

of 7,500 tonnes per day. Mining and economic parameters applied were based on the






2011 PEA assumptions which are still considered reasonable by the AMEC QPs.

Since the block unit value is estimated using commodity prices expressed in US

dollars, the costs and assumptions are also in US dollars.

1.14 Mineral Resource Statement

The Mineral Resource update is classified in accordance with the 2010 CIM Definition

Standards for Mineral Resources and Mineral Reserves, whose definitions are

incorporated by reference into NI 43-101. The Mineral Resource update with effective

date 22 June 2012 is summarized in Table 1-1.

Table 1-1: Blue River Project Estimated Mineral Resources; Effective Date 22 June 2012,

Tomasz Postolski, P.Eng, Qualified Person

Ta price

[US$/kg]

Confidence

Category Tonnes

Ta2O5

[ppm]

Nb2O5

[ppm]

Contained

Ta2O5

[1000s of kg]

Contained

Nb2O5

[1000s of kg]

317 Indicated 51,780,000 192 1,490 9,930 76,900

Inferred 8,800,000 186 1,660 1,600 14,600

Notes:

1. Assumptions include commodity prices of US$317/kg Ta, US$46/kg Nb, process recoveries of 65.4%

for Ta2O5 and 68.2% for Nb2O5, US$24/tonne mining cost, US$13/tonne process and refining cost,

US$3/tonne G&A cost.

2. Mineral resources are amenable to underground mining methods and have been constrained using a

“Stope Analyzer”

3. An economic cut-off was based on the estimated operating costs assuming either the bulk or selective

mining method from the PEA mine plan. The block unit value cut-off ranged from US$40/t (bulk) to

US$58/t (selective)

4. Mining losses = 0%, external dilution = 0%; planned internal dilution within the minimum stope size is

included

5. In situ contained oxide reported. Discrepancies in contained oxide values are due to rounding.

The Mineral Resources used for 2011 PEA were those with an effective date of

29 September 2011 as follows:

Indicated: 36.35 million tonnes at 195 ppm Ta2O5 and 1,700 ppm Nb2O5

Inferred: 6.4 million tonnes at 199 ppm Ta2O5 and 1,890 ppm Nb2O5.

There is a considerable increase in resource tonnes for the current Mineral Resource

update relative to the 29 September 2011 tonnage estimate where the Indicated

category has increased by 42% and the Inferred category by 37%. This increase in

tonnes is mostly due to (1) lowering the bulk mining method block unit value cut-off

from US$52/t to US$40/t by eliminating backfill costs and, to a lesser extent,

(2) additional infill diamond drilling.






1.15 Preliminary Economic Assessment

1.15.1 2011 PEA

This section incorporates assumptions, analysis and findings of the Preliminary

Economic Assessment that has an effective date of 29 September 2011.

The preliminary mine plan presented in this section is partly based on Inferred Mineral

Resources that are considered too speculative geologically to have the economic

considerations applied to them that would enable them to be categorized as Mineral

Reserves, and there is no certainty that the Preliminary Economic Assessment based

on these Mineral Resources will be realized.

The information relevant to the preliminary mine plan, prepared during the 2011 PEA,

is included in this section and has not been updated because AMEC considers that the

assumptions supporting the outcomes remain reasonable.

In a similar manner, AMEC’s opinion is that the assumptions made with respect to

infrastructure, recovery methods, estimation of capital and operating costs, marketing

studies and metal price assumptions and the resulting financial analysis remain

reasonable.

The effective date of the 2011 PEA results therefore remains 29 September 2011.

1.15.2 Proposed Mining Method

The 2011 PEA was developed assuming a sub-level open stoping mining method with

no backfill and no pillar recovery. In agreement with Commerce, a processing rate of

7,500 t/d was assumed for Mineral Resource estimation and for the conceptual design

of an underground mine for the Blue River Project. Price assumptions used in mine

planning were US$317/kg tantalum metal and US$46/kg niobium metal contained in

oxide product.

1.15.3 Geotechnical Considerations

Rock types of the Blue River Project have been grouped into two main geotechnical

domains: Intrusive and Layered Rocks. The Intrusive group encompass carbonatite

and fenite rocks, while the Layered Rocks group encompass gneiss and amphibolite

rocks. Generally, the rock mass ratings (RMR) indicated rock that can be considered

to be “good” in RMR terms.






1.15.4 Dilution Considerations

Material deemed to be mined by bulk mining methods represents 84% of the

resources. Within the mineable shapes there was internal dilution of 2% waste rock. It

was assumed that during mining 2% of waste material would be added as external

dilution and 2% of the broken material would not be recovered from the stopes due to

operational conditions.

The geotechnical investigation indicates that an extraction ratio of 67.5% is

reasonable. Applying this to the subset of the Mineral Resources considered in the

mine plan results in an overall mining extraction of 58% and provides 25.0 Mt of

material as run-of-mine (ROM) production to be processed. Applying internal and

external mining dilution, the overall subset Mineral Resource grades were diluted to

185 ppm of Ta2O5 and 1,591 ppm of Nb2O5 for mine planning purposes.

1.15.5 Drilling and Blasting

AMEC considered the implementation of conventional drilling methods. Due to the

high precipitation in the region and water continuity along fractures and rock layers,

wet conditions were assumed for development and stoping areas. The use of bulk-

blasting systems based on emulsion type explosives was assumed.

1.15.6 Mine Development

The deposit will be accessed through two main portals, the Upper and Lower Portals;

these portals are located in places where the deposit crops out on the hillside. The

Upper Portal will be located at Elevation 1,150 m. It will be used as the main entry and

will have most of the mine services; the Lower Portal will be located at Elevation

1,030 m and will be used for haulage trucks access. Access to the portals will be by a

road upgraded from existing exploration roads.

The mine will be accessed by adits driven in pairs from the portals. For this study, all

entries, ramps, drifts and crosscuts are considered to be 5 m wide by 5 m high with

semi-arched backs. Ramps will be driven at grades to a maximum of 15% to provide

access to the production areas. Two ramps or adits will be driven to each area to

provide single-way traffic of haulage trucks and to facilitate the implementation of

ventilation circuits.

Top access crosscuts are driven from the main ramps to each level on vertical

intervals between 20 to 30 m. Stope access crosscuts are driven at level from west to

east. Bottom access crosscuts are driven to function as mucking drifts. Underground






mine services such as ventilation and air heating, compressed air, water for drilling

and power supply will be provided to the mine via the adits. The total underground

development was estimated at 92,500 m for the life-of-mine.

1.15.7 Mineralized Material and Waste Haulage

Radio remote-controlled load-haul-dump units (LHDs) will be used to extract the

mineralized material out of the stope beyond the safety of the brow. The mineralized

material from stopes will be loaded directly to the haulage trucks that will be spotted at

the end of the crosscut. Underground trucks will haul the mined material through the

access drifts and ramps, unloading into the primary crusher surface stockpile near the

Lower Portal.

The PEA plan envisages that tailings material will be dry-stacked, and waste material

will be stored in the same general area. For convenience, the combined tails and

waste rock storage area is referred to as the “co-disposal facility”.

Waste from development will be initially utilized for construction of a structural shell for

the co-disposal site on surface, which will be located between Elevations 1,400 m and

1,600 m in an area east of the processing plant site.

A conveyor will be used to transport this material from the mine to a stockpile by the

plant site. A surface road developed at +10% grade will connect the plant with the

co-disposal site.

Trucks will haul waste from the plant to the co-disposal site when required.

1.15.8 Mine Services

Underground mine services such as ventilation and air heating, compressed air, water

for drilling and power supply will be provided to the mine via adits from the portals.

The sub-station for the main power distribution system and the air compressors will be

installed in facilities located adjacent to the Upper Portal.

Other mine services will include all the systems and supplies needed for the mining

operations, including: explosives storage, communications, monitoring and control

systems, road maintenance and mine equipment maintenance.

Portable self-contained refuge stations will be provided for the mine and will be located

at convenient locations.






1.15.9 Mine Production Forecasts

Production was estimated at 2.7 Mt/a of mill feed on average for 10 years. The first

year was considered as preproduction, leaving nine years of full-scale production.

At this preliminary level of study the stope mining sequence was not defined and

therefore average grades were used for each year in the mine plan.

There is opportunity to increase the net present value (NPV) of the project by mining

higher-grade zones early in the mine life providing that the sequence and overall

recovery of the stopes is not negatively affected.

Mine equipment and personnel requirements were defined and are appropriate to the

proposed production plan.

1.15.10 Process Design

The design for the process facilities considered a nominal processing capacity of

7,500 t/d. Where data were not available at the time of flowsheet development, AMEC

developed criteria for sizing and selection of equipment based on comparable industry

applications, benchmarking, and the use of modern modelling and simulation

techniques.

The mineral processing and the refining are based on conventional technology and

industry-proven equipment. A mineral processing method using a standard grind-

flotation procedure to make a concentrate of ferrocolumbite-pyrochlore is assumed for

Blue River material.

Metallurgical testing indicates a mineral concentrate assaying about 30% combined

Nb–Ta pentoxide within the recovery range of 65% to 70% is possible.

Run of mine mineralized material will be crushed to minus 5/8" and fed into a

comminution circuit comprised of a rod and ball mill using cycloning for classification.

After grinding, the flotation feed will be first de-slimed using high frequency fine

screens and cyclones. The coarse product resulting from de-sliming will be sent to

four concentration steps that will include pyrrhotite flotation, carbonate flotation, and

magnetite separation with all three concentrates from these processes reporting to

tailings. The fourth step, pyrochlore flotation, will recover a concentrate which is

reground and cleaned in five stages of cleaning. The mass of material will be reduced

substantially, to less than 1% of the feed into the plant.






This concentrate would be further processed to produce marketable separate Ta and

Nb products. The proposed processes are mature, are already used industrially, and

consist of reducing the concentrate to metals as ferroalloys in a standard

aluminothermic furnace followed by chlorinating the alloys and distilling the product to

separate high purity metal chlorides, TaCl5 and NbCl5. Recoveries from concentrate to

pure chlorides are expected to be 97%. Both Ta and Nb chloride products are then

readily converted and marketed as high purity oxides Ta2O5 and Nb2O5 respectively.

1.15.11 Tailings and Waste Management

The PEA design for tailings and waste management is to construct a co-disposal

drystack facility.

A series of tailings storage location screening assessments were carried out during

2008, 2009, and 2010 focusing mainly on conventional tailings storage. Some

evaluation of potential waste rock storage sites and a tailings drystack facility was

undertaken during studies completed in 2009.

Site WSF 3, about 8 km from the proposed plant site, and identified in 2009, was

selected as the preferred location for a tailings drystack. WSF 3 has flatter slopes than

other site alternatives and therefore has less stability concerns. It is also located in

closer proximity to the proposed plant site, although uphill haulage will be required,

and will not require crossing any major creeks. The facility was designed to hold a

total storage volume of approximately 20.9 Mm3.

Surface water management systems will likely be required for the tailings drystack to

divert non-contact (clean) water around the facility, and to collect run-off water which

had been in contact with the tailings drystack area.

1.15.12 Planned Project Infrastructure

The planned Upper and Lower Portals will be located about 4 km from the plant site.

At the front of the Upper Portal (service portal) sufficient space will be provided to

accommodate the required facilities for operation.

The plant service building will be a multi-purpose complex in a two-story building to be

located east of the process building. A 24 m by 36 m truck shop on the southwest end

of the site will be operated by a qualified contractor. The warehouse will be a 24 m by

50 m Coverall-type fabric building.






The potable water system and layout is designed to service buildings and a

process/administration workforce of 120 persons. Potable water for the mining area

will be constructed as part of the portal/underground works.

A modular sewage treatment system will be installed as part of the initial construction

infrastructure. Waste-water treatment sludge will be trucked away to a nearby

municipal facility or approved landfill. Waste lubrication and hydraulic oils from vehicle

maintenance will be stored in dedicated tanks and sent to a recycling facility offsite.

Contractors and employees will commute from nearby towns such as Blue River and

Valemount during construction. No on-site permanent accommodation will be

provided for personnel. It is assumed that the workforce, including management staff,

will reside in the nearby communities and will commute, via buses on a daily basis.

The road access design includes a short new road with a 7.2 m wide gravel surface

from the existing road to plant site about 80 m in length and a 1.5 km new service road

from the existing road to the Upper Portal and upgrades to the current access road.

There are two main haul roads, one from the mine portal to site of about 4 km length,

and the other from the site to the co-disposal facility, with a length of about 8 km.

An existing 80 m-long bridge crossing over the Thompson River has a limited load

capacity and might not qualify for crossing heavy loads during construction or long-

term use during the life of the mine. Therefore a new bridge has been included in

capital cost estimate. Using the existing railway for shipment should be investigated in

the next phase of study.

Ammonium nitrate, blended emulsion, and explosives will be delivered to site on

demand by contractors. Fuel will be delivered to the mine site using tanker trucks. A

crushing and stockpiling facility will be required during construction to provide crushed

product for roads and surfacing. AMEC considers that there is no need to establish a

concrete batch plant, and that concrete supply from nearby towns will be more

economic.

1.15.13 Markets

Commerce has prepared assessments of the tantalum and niobium markets which

outline the supply and demand for tantalum and niobium. The tantalum assessment

was prepared by a tantalum market expert, although he is not independent of

Commerce. His analysis reflects the general consensus of other analysts regarding

the tantalum market expressed in publicly available information.






The niobium assessment was prepared by an independent niobium expert and also

reflects the general consensus of analysts in publicly-available information for the

niobium market.

As the Project is still at an early evaluation stage, Commerce has not initiated requests

from potential buyers for expression of interests from potential buyers of the proposed

Blue River products and has not negotiated any purchase or off-take agreements.

The tantalum price assumption used in the 2011 PEA is based on 4th quarter 2010

information. The tantalum price moved significantly higher through 2011. AMEC has

checked publicly available tantalum and niobium metal prices as at May 2012 and

found the Ta and Nb prices used for both the current Mineral Resource estimate and

the 2011 PEA remain as reasonable assumptions, which are US$317/kg tantalum

metal and US$46/kg niobium metal.

1.15.14 Capital Costs

All costs are expressed in constant first quarter (Q1) 2011 Canadian (CAD) dollars.

No allowance has been included for escalation, interest or financing fees, taxes or

duties, or working capital during construction. The level of accuracy for the estimate is

+40/-20% of estimated final costs, as per the Association of Advanced Cost Estimators

(AACE) Class 5 (scoping level) definition.

Contractor-mining is not envisaged for pre-production development. The estimate

covers the direct field costs of executing the project, plus the Owner’s indirect costs

associated with design, construction, and commissioning. The preproduction costs are

capitalized and include all the expenditures before Year 1 of production.

The total estimated capital cost to design and build the Blue River tantalum project at

7,500 t/d capacity is CAD$379 million. The estimate is summarized in Table 1-2.






Table 1-2: Summary of Estimated Capital Costs

Item

Total

(CAD$000’s)

2011

(CAD$000’s)

2012

(CAD$000’s)

Project year

1 2

Production year

-2 -1

Capital expenditure

Initial Capital Infrastructure 29,491 10,322 19,169

Process Initial Capital 116,240 40,684 75,556

Mining Initial Capital 89,420

89,420

Material Handling 8,000

8,000

Contingency 43,613 12,751 30,862

Indirect/Owner Costs 92,268 29,627 62,641

Total 379,032 93,385 285,647

1.15.15 Operating Costs

All operating costs are expressed in Canadian (CAD) dollars. The operating costs for

the Blue River project are based on an Owner-operated mining fleet and process

facility and have been prepared in first quarter 2011 Canadian dollars. Operating

costs over the life-of-mine are estimated at CAD$38.44/t milled.

Operating costs include the three key areas of mining, process, and overall general

and administrative costs (G&A). The estimates are based on the staffing level,

consumables, and expenditures detailed as part of the underground mine plan and

process design. Average operating costs are listed in Table 1-3.

Table 1-3: Average Life-of-Mine Operating Cost Summary

Summary of Average Production Costs

LOM Total

(CAD$000’s)

Cost per

Tonne

Milled

(CAD$/t)

Mining 528,937 21.16

Process 338,500 13.54

Material Handling 18,516 0.74

G&A 74,998 3.00

Sub-total 960,951 38.44






1.15.16 Financial Analysis

The financial analysis is partly based on Inferred Mineral Resources that are

considered too speculative geologically to have the economic considerations applied

to them that would enable them to be categorized as Mineral Reserves, and there is

no certainty that the Preliminary Assessment based on these Mineral Resources will

be realized. Approximately 15% of the Mineral Resources that were included in the

financial model are classified as Inferred Mineral Resources.

The results of the economic analyses discussed in this section represent

forward-looking information as defined under Canadian securities law. The results

depend on inputs that are subject to a number of known and unknown risks,

uncertainties and other factors that may cause actual results to differ materially from

those presented here.

Information that is forward-looking includes:

Mineral Resource estimates

Assumed commodity prices and exchange rates

Estimated capital and operating costs

The proposed mine production plan

Projected recovery rates

Infrastructure construction costs and schedules

Assumptions that an EA will be approved by Provincial and Federal authorities.

The Project has been evaluated using a discounted cash flow (DCF) analysis. Cash

inflows consist of annual revenue projections for the mine and two years of

preproduction. Cash outflows such as capital and operating costs are subtracted from

the inflows to arrive at the annual cash flow projections.

The resulting net annual cash flows are discounted back to the date of valuation end-

of-year 2010 dollars and totalled to determine NPVs at the selected discount rates.

The IRR is calculated as the discount rate that yields a zero NPV. The payback period

is calculated as the time needed to recover the initial capital spent.

The Project Base Case (8% discount rate) returns an NPV of CAD$18.5 million before

tax. Table 1-4 summarizes the NPV for the Project at a range of discount rates, with

the Base Case highlighted.






Table 1-4: Summary Financial Analysis at Various Discount Rates

(base case is highlighted)

Summary of Cash Flow Pre-tax

Cumulative net cash flow

Undiscounted CAD$000 236,631

Net present value

Discounted at 5% CAD$000 80,349



Discounted at 8% (Project Base Case) CAD$000 18,487


Discounted at 10% CAD$000 (13,514)

Internal rate of return % 9.1

Payback period Years 6.3

Note: An exchange rate of US$0.95 to CAD$1.00 is used for all years of the financial model.

The cash cost value represents the cost incurred to produce 1 kg of primary product

after deducting the revenue from sales of secondary products. Since the price

analysis for the report was performed around Ta price variation, Ta is chosen as the

main product and Nb is treated as the secondary product for the assessment of cash

cost.

Using the Brook Hunt convention for reporting C1 cash costs1, the C1 cash cost of

tantalum is CAD$24.91/kg contained in oxide product (after credit for niobium

contribution) as shown in Table 1-5.

The cash cost for production of tantalum during the earlier years of the proposed

mining operation is CAD$57/kg and decreases over the life of the mine. The major

driver behind the changing costs is the decrease in the mining costs over the life-of-

mine. In the last three years of operation, the revenue generated from niobium

exceeds the total operating costs (mining, processing and G&A). The mining cost for

the entire Project (i.e. mining cost of both tantalum and niobium) drops from an

average of CAD$24/t in the first few years to CAD$18/t in the last year of full

production.

It is reasonable to expect that there has been cost escalation since the base of first

quarter 2011 but there has been no adjustment for this in the Report other than in the

sensitivity analysis.

1 Brook Hunt, established in 1975, is a global group that specializes in in-depth market analysis across the mining and metals

industries. Brook Hunt has established a method of comparison of costs between projects, countries and commodities that is considered an industry standard. C1 cash costs are defined by Brook Hunt as: the costs of mining, milling and concentrating, on-site administration and general expenses, property and production royalties not related to revenues or profits, metal concentrate treatment charges, and freight and marketing costs less the net value of by-product credits.






Table 1-5: Life of Mine Cash Cost Summary

Section

LOM Total

(CAD$000’s)

Cost per Tonne

Milled

(CAD$/t)

Cost per Kilogram Ta Payable

(CAD$/kg)

Cash costs

Mining 528,937 21.16 220.13

Process 338,500 13.54 140.87

G&A 74,998 3.00 31.21

Material Handling 18,516 0.74 7.71

Sub-total 960,951 38.44 399.22

Credits

Nb 901,094 36.04 375.01

Sub-total 901,094 36.04 375.01

Adjusted cash costs

Total 59,857 2.40 24.91

Note: The figures in this table do not include considerations of working capital or royalty payments

1.15.17 Sensitivity Analysis

The Upper Fir deposit is most sensitive to changes in exchange rate, mining costs,

and commodity prices. Since the sales currency is US dollars and operational costs

are in Canadian dollars, a rising US dollar value versus Canadian dollar value

improves the mine profitability. The Project is more sensitive to changes in operating

expenditures than capital expenditures.

The Project IRR increases to 14.4% at a Ta price of US$380/kg and the Project NPV

increases to CAD$125 million at an 8% discount rate. Sensitivities are illustrated in

Figure 1-1 for the 8% discount Base Case rate.






Figure 1-1:Sensitivity Summary, 8% Discount Rate

1.16 Interpretation and Conclusions

1.16.1 2012 Mineral Resource Estimate Update





The Mineral Resource update is based on information of reasonable quantity and

quality. The lithological, geotechnical, and collar location, down-hole survey, and drill

core sample data collected by Commerce in the exploration and delineation drill

programs meet and exceed industry standard practice.

The deposit is amenable to conventional underground mining methods with estimated

mining recovery that may vary from 65 to 85% depending on the mine and stope

layout and the success in which pillars can be mined on retreat.

Tantalum and niobium occur within the minerals pyrochlore and ferrocolumbite and are


recoveries of 65% to 70%.






High-quality technical grade tantalum and niobium products proposed for production

at-site are suitable for several markets. As the Project is still at an early evaluation

stage, Commerce has not initiated requests from potential buyers for expression of

interests from potential buyers of the proposed Blue River products and has not

negotiated any purchase or off-take agreements.

The Mineral Resources have significantly increased in tonnage mostly due to a

reduction in the block unit value cut-off by eliminating back-fill costs and, to a lesser

extent, additional infill diamond drilling.

The Mineral Resource update uses the same assumptions from the 2011 PEA for the

following items:

Ta and Nb metal prices (US$317/kg tantalum metal and US$46/kg niobium metal)






1.16.2 2011 PEA



Estimated internal rate of return: 9.1% (before tax)

Estimated net present value: CAD$18.5 million at 8% discount rate (before tax)

Estimated payback: 6.3 years



Conceptual operating cost: CAD$38.44/t milled (mining ~ 55% of cost)

Conceptual capital cost: CAD$379 million (process ~ 31% of initial cost)


several markets

Conceptual mine life: 10 years based upon the mineral resources (effective date

29 September 2011)

NPV sensitivity: The Upper Fir deposit is most sensitive to changes in exchange

rate, mining costs, and commodity prices.






The above key outcomes for the 2011 PEA contain forward looking information. The

assumptions and risks regarding those assumptions are explained in the body of the

Report.

AMEC has checked the publicly available tantalum and niobium metal prices as at May

2012 and found the Ta and Nb price assumptions used for both the current Mineral

Resource estimate and the 2011 PEA to remain reasonable. A higher tantalum price

would improve profitability and also increase the mine life. Additional exploration

potential could also provide additional mine life. A two or more times capital payback

is possible.

1.16.3 Project Opportunities

As a result of engineering work during the 2011 PEA, a lower block unit value cut-off

can be achieved by revising the mine design to eliminate back-fill costs. This

approach was used to support the current 22 June 2012 Mineral Resource update,

which in turn has increased the Mineral Resource tonnage at the Project. The

increase in Mineral Resources provides more flexibility for future mining studies and

hence opportunities to improve the Project NPV are as follows:

Optimization of the mine plan by mining higher-grade zones earlier in the mine life

providing that a practical mining sequence can be implemented and the overall

recovery of the Mineral Resources is not negatively affected

Optimization of the mine layout to minimize development costs

Advanced geotechnical studies to identify and understand ground conditions which

could allow an increase in the size of stopes and production drifts

Optimization of the supply and pricing of reagents for the refining.

1.16.4 Project Risks

In the opinion of the QPs, the top five risk factors identified for the Project are:

The current Mineral Resource estimate is supported by current tantalum and

niobium prices which are higher than historic average prices and may not reflect

long term prices.

Commerce has not initiated requests from potential buyers for expression of

interests from potential buyers of the proposed Blue River products and has not

negotiated any purchase or off-take agreements.

The proposed refining methods have been used in commercial applications but

have not been demonstrated in test work of Blue River material.






Testwork to date has not considered factors such as water recycling. A water

treatment plant may be required and may result in increased capital costs.

The 2011 PEA financial analysis is partly based on Inferred Mineral Resources

(effective date 29 September 2011) that are considered too speculative

geologically to have the economic considerations applied to them that would

enable them to be categorized as Mineral Reserves, and there is no certainty that

the Preliminary Economic Assessment based on these Mineral Resources will be

realized.

A more comprehensive risk factor list can be found in Section 25 of the Report.

1.17 Recommendations

AMEC recommends the following work programs:

Project management, field work, and desk top studies total about $2.0 million and

include the following: (1) project management and administration costs; (2) field costs

comprising a re-sampling program for campaigns with poor precision and accuracy to

improve confidence in their analyses, structural geology studies, and manpower and

field support costs; (3) core farm security improvements; (4) on-going marketing work;

(5) mining trade-off studies to optimize mine and stope design; (6) resource modeling

trade-off studies to optimize grade distribution; and (7) a mineral resource estimate

update.

The re-sampling program should focus on re-assaying samples within an area where

the first five years of mining is likely to occur.

An additional mineral resource update is recommended after all the 2011 drilling data

has been analysed, verified, updated into the drilling database, and interpreted. This

mineral resource update would include all drilling information up to and including the

2011 campaign plus outcomes from any mining, resource modeling, or metallurgical

optimization studies.

Additional diamond drilling is recommended totalling about $3.2 million for drilling,

sampling, assaying, and logging costs. The recommended drilling is to focus on the

volume within the first 5 years of the conceptual mine plan. The recommended

program has about 40 diamond drill holes comprising about 10,000 m of HQ diameter

coring for resource infill and step-out drilling and about 8 diamond drill holes

comprising around 2,000 m of PQ diameter coring for metallurgical testwork purposes.






2.0 INTRODUCTION

2.1 Terms of Reference

AMEC Americas Limited (AMEC) was commissioned by Commerce Resources

Corporation (Commerce) to prepare a NI 43-101 compliant Mineral Resource update

and technical report on the wholly-owned Blue River tantalum–niobium Project (the

Project), located within the North Thompson River valley of east–central British

Columbia (B.C.), Canada.

This technical report (the Report) supports the findings of the Mineral Resource update

and also includes a summary of the Preliminary Economic Assessment study

completed on the Blue River Project with an effective date 29 September 2011 (2011

PEA) . Results from the 2011 PEA mining studies have not changed in terms of their

outcomes as their underlying assumptions remain reasonable.

Commodity prices are quoted in US dollars. All other costs are in Canadian dollars

unless otherwise indicated. Volumes, weights, and distances are metric unless

otherwise indicated.

2.2 Qualified Persons

The Qualified Persons (QPs) for the Report are AMEC employees, based out of

AMEC’s Vancouver office, as follows:

Mr. Albert Chong, P.Geo., Principal Geologist

Mr. Tomasz Postolski, P.Eng., Senior Geostatistician

Mr. Ramon Mendoza Reyes, P.Eng., Principal Engineer

Mr. Tony Lipiec, P.Eng., Principal Metallurgical Engineer

Mr. Behrang Omidvar, P.Eng., Financial Analyst

2.3 Site Visits and Scope of Personal Inspection

Mr. Chong visited the property on a number of occasions, between 11 to 16 July 2010,

27 to 30 June 2011, and 6 to 14 September 2011. During the site visits Mr. Chong

inspected outcrops, drill hole collar locations, and reviewed the geologic interpretation.

Mr. Chong reviewed procedures for diamond drilling, logging, sampling, core storage,

and sample shipment. The drilling used for the current Mineral Resource update was

reviewed during the site visits by Mr. Chong. In 2011 Mr. Chong also completed an

aerial inspection of proposed future infrastructure locations.






Mr. Mendoza Reyes conducted a site visit from 12 to 16 July 2010. During the site

visit Mr. Mendoza inspected sites amenable for locating potential infrastructure.

Mr. Postolski conducted a site visit from 27 to 30 June 2011. During the site visit

Mr. Postolski was unable to access the property due to a road wash-out but was able

to complete an aerial inspection of proposed future infrastructure locations. During this

visit Mr. Postolski was able to visit the core logging facilities, observe core logging

procedures, and review progress on the mineral resource geology interpretation.

A summary of QP site visits is shown in Table 2-1.

Table 2-1: Site Visit and Areas of Report Responsibilities

Qualified Person Site Visit Sections of Report Responsibility

Albert Chong, P.Geo. 11 to 16 July 2010

27 to 30 June 2011

6 to 14 September 2011

Sections 1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11,

12, 20, 23, 24, 25, 26, and 27

Tomasz Postolski, P.Eng. 27 to 30 June 2011

Section 14 and those portions of the

Summary, Interpretation and Conclusions,

and Recommendations that pertain to that

Section

Ramon Mendoza Reyes, P.Eng 12 to 14 July 2010 Sections 15, 16, 18, and those portions of

the Summary, Capital and Operating

Costs, Interpretation and Conclusions,

and Recommendations that pertain to

those Sections

Tony Lipiec, P.Eng No site visit Sections 13 17, 18, 21, and those portions

of the Summary, Interpretation and

Conclusions, and Recommendations that

pertain to those Sections

Behrang Omidvar, P.Eng No site visit Sections 19, 21, 22, and those portions of

the Summary, Interpretation and

Conclusions, and Recommendations that

pertain to those Sections

2.4 Effective Dates

Information in this Report has a number of cut-off dates, as follows:

Effective date of the 2012 Mineral Resource update, the subject of this Report, is

22 June 2012

o The database closure date is 29 September 2011 and includes all drill

information up to the end of 2010.






o Results from 34 holes completed in 2011 were compared by AMEC to the

existing resource model and the results were found to be reasonably

consistent.

o The 2011 preliminary drill results were inspected by AMEC at the Project site

on vertical cross-sections by AMEC during September 2011, and the results

were found to be consistent with the geological interpretation.

Effective date of the 2011 PEA technical report work and the mineral resources upon

which the PEA is based is 29 September 2011. In the opinion of AMEC:

The 2011 PEA was completed within the last 12 months which is a reasonable

timeframe.

The 2011 PEA basis and underlying assumptions for the mining study remain

reasonable and hence the outcomes also remain reasonable.

The overall effective date of the Report is based on the completion date of the Mineral

Resource update and review of the latest available drilling information and is 22 June

2012. There has been no material change to the Project scientific and technical

information between the effective date of the Report, and the signature date.

2.5 Information Sources and References

Information for the Report was obtained from work completed by AMEC, and materials

provided by, and discussions with, Commerce personnel and third-party contractors

retained by Commerce.

Additional information was provided by Commerce and third party personnel in the

areas of environmental studies and permitting. This information was based on reports

prepared by third-party consultants retained by Commerce.

Marketing studies were provided to AMEC by Commerce. AMEC has reviewed the

information provided and considers that the studies support the commodity prices used

in the mineral resource estimates and financial evaluation.

Except where noted, Report figures were generated by AMEC.

Reports and documents listed in the Section 3, Reliance on Other Experts and Section

27, References sections of this Report were also used to support preparation of the

Report. Additional information was sought from Commerce personnel where required

to support preparation of this Report.






2.6 Previous Technical Reports

Commerce has previously filed the following technical reports on the Project:

Chong, A., Postolski, T., Mendoza, R., Lipiec, T., and Omidvar, B. 2011: Commerce

Resources Corporation, Blue River Tantalum–Niobium Project, Blue River, British

Columbia, Preliminary Economic Assessment: unpublished technical report prepared

by AMEC Americas Ltd. for Commerce Resources Corporation, effective date

September 29, 2011.

Chong, A., and Postolski, T., 2011: Commerce Resources Corporation, Blue River Ta-

Nb Project, Blue River, British Columbia, NI 43-101 Technical Report: unpublished

technical report prepared by AMEC Americas Ltd. for Commerce Resources

Corporation, effective date 31 January 2011.

Stone, M., and Selway, J., 2010: Independent Technical Report, Blue River Property,

Blue River, British Columbia, Canada: unpublished technical report prepared by

Caracle Creek International Consulting Inc. for Commerce Resources Corporation,

effective date 24 August, 2009.

Gorham, J., 2007: Technical Report on the Upper Fir Ta-Nb Bearing Carbonatite:

unpublished technical report prepared for Commerce Resources Corporation, effective

date 20 June 2007.






3.0 RELIANCE ON OTHER EXPERTS

The QPs state that they are qualified persons for those areas as identified in the

appropriate QP “Certificate of Qualified Person” attached to this Report. The authors

have relied upon and disclaim responsibility for information derived from the following

reports pertaining to mineral tenure, surface rights, royalties, environment and

permitting, and marketing.

3.1 Mineral Tenure

The AMEC QPs have not reviewed the mineral tenure, nor independently verified the

legal status, ownership of the Project area or underlying property agreements. AMEC

has fully relied upon, and disclaims responsibility for, information derived from legal

experts for this information through the following document:

Letter from Clark Wilson LLP titled Commerce Resources Corp. – Mineral Claim

Title Opinion to Mr. Albert Chong, dated 25 April 2012.

Information from this letter has been used in Section 4.2 of this Report.

3.2 Surface Rights

The AMEC QPs have not reviewed the status of the Project surface rights, nor

independently verified the surface rights status of the Project area. AMEC has fully

relied upon, and disclaims responsibility for information derived from experts for this

information through the following document:

Letter from David Hodge, President Commerce Resources Corp. to Mr Albert

Chong, entitled “Commerce Resources Corp – Blue River Property Encumbrances

and Surface Rights”, dated 08 May 2012.

This information has been used in Section 4.3 of this Report.

3.3 Royalties and Agreements

The AMEC QPs have not reviewed the royalties and agreements for the Project, nor

independently verified the royalties and agreements status of the Project area. AMEC

has fully relied upon, and disclaims responsibility for information derived from experts

for this information through the following document:






Letter from Clark Wilson LLP titled Commerce Resources Corp. – Mineral Claim

Title Opinion to Mr. Albert Chong, dated 25 April 2012, including Schedule A:

Claims List, and Schedule B: Officer’s Certificate from Mr. David Hodge, President

and Chief Executive Officer of Commerce Resources Corp.

This information has been used in Section 4.4 of this Report.

3.4 Environmental, Permitting, and Liability Issues

The AMEC QPs have not reviewed the permitting requirements, nor independently

verified the permitting status of the Project area. AMEC has fully relied upon, and

disclaims responsibility for information derived from experts for this information through

the following document:

Letter from Sage Resource Consultants Ltd. titled Commerce Resources Corp.

Upper Fir Deposit Mineral Resource Update – Independent Professional Opinion

on Environmental Permitting and Liability to Mr. Albert Chong and dated 30 April

2012.

This information has been used in Section 20 of this Report.

3.5 Markets

The AMEC QPs have relied on tantalum and niobium market analyses derived from

experts for this information through the following documents:

Confidential e-mail: from Dr. Axel Hoppe titled “Ta-Pricing 2011” dated

01 February 2012 to Jenna Hardy, Technical Services Manager, Commerce

Resource Corporation. Received 02 February 2012.

Memo from Dr. Axel Hoppe titled “Ap#6 Introduction to Tantalum

Markets_Finalpdf_2June09.pdf” received 18 October 2010

Memo from Michel Robert titled “Niobium_v3jh.doc” received 18 October 2010

Memo from Michel Robert titled “Niobium_v3jh.doc” received 19 October 2010

Dr. Hoppe is an internationally acknowledged leader in the tantalum field and is

Chairman of the Board of Directors for Commerce Resources.

Mr. Robert has extensive experience in niobium markets and is independent of the

company.






This information has been used in Section 19 of this Report, and to assess reasonable

prospects of economic extraction in Section 14.10.






4.0 PROPERTY DESCRIPTION AND LOCATION

The Project is located within the North Thompson River valley of east-central British

Columbia 25 km to 60 km north and northeast of the community of Blue River, British

Columbia (Figure 4-1).

The NTS sheets which cover the Project are: 83D.004-.006; 83D.014-.016;

83D.024-.027; 83D.034-.037; 83D.045-.047.

The Project is centered at approximately 52° 19' N latitude and 119° 10' W longitude.

4.1 Project Ownership

The Project is wholly-owned by Commerce and held in the name of Commerce

Resources Corp.

4.2 Mineral Tenure

The Project comprises 249 two-post claim, four-post claim, and mineral cell title

submission (MCX) claims in good standing that encompass just over 1,000 km2

(105,373 ha) within the Kamloops Mining Division. The claim boundaries are shown in

Figure 4-2.

A table listing claim details is included in Appendix A.

Currently, all of the mineral claims are valid until 31 March 2021. Commerce filed a

2010 work program by the end of day on 24 June 2011 thus grouping two blocks of

claims (Wasted and Joined) with the larger claims area. About $969,000 worth of

eligible 2010 work was filed for assessment credit at a filing cost of approximately

$48,700. This work has been approved by the British Columbia Ministry of Energy,

Mines and Petroleum Resources, bringing all Blue River claims to a common expiry

date of 31 March 2021.

Property boundaries are established in accordance with the Mineral Tenure Act of

British Columbia. Commerce has staked the claims by a combination of ground and

on-line staking.

Two-post and four-post claims were established through a legacy system of ground

staking which involved physically establishing claim posts on the ground.

MCX claims are established using the Government of British Columbia’s Mineral Titles

Online (MTO) staking system. MTO is an Internet-based mineral titles administration






system that allows the mineral exploration industry to acquire and maintain mineral

titles by selecting the area on a seamless digital GIS map of British Columbia. The

electronic Internet map allows selection of single or multiple adjoining grid cells. Cells

range in size from approximately 21 ha (457 m x 463 m) in the south to approximately

16 ha at the north of the province. All boundaries are oriented north–south and east–

west.

Figure 4-1: Project Location Map

Figure courtesy Commerce Resources Corp., 2011.






Figure 4-2: Blue River Mineral Tenure Map

Note: Grid is in metres for UTM NAD83 Zone 11. The British Columbia Yellowhead Highway 5 is

adjacent to the property claim boundary. Figure courtesy Dahrouge Geological Consulting Ltd., 2012.






4.3 Surface Rights

Commerce holds no surface rights on the property. Legal access to the property is

provided through the British Columbia Mineral Tenure Act. The Act provides for a

recorded claim holder to use, enter and occupy the surface of a claim or lease for the

exploration and development or production of minerals or placer minerals, including

the treatment of ore and concentrates, and all operations related to the exploration and

development or production of minerals or placer minerals and the business of mining.

Access to surface rights held by third parties typically requires compensation to be

paid.

There are surveyed parcels with surface rights held by other parties which overlap the

property mineral tenure claims. These parcels occur along the western edge of the

property and most are tree farm licences. Commerce is not aware of any material

issues that would prevent negotiation for access or surface rights of these surveyed

parcels should they be required in the future. The claims do not host mineral

resources, and currently no carbonatites are known within the claims.

4.4 Royalties and Agreements

There are no royalties, back-in rights, payments, or other agreements or

encumbrances to which the Blue River property is subject to other than the annual

claim maintenance fees due to the government as set out by the British Columbia

Mineral Tenure Act and Regulations.

4.5 Permits

Permits required to support Project development are discussed in Section 20.

4.6 Environment

Environmental studies are discussed in Section 20.

4.7 Social and Community Impact

The social and community impact assessments of the Project are discussed in

Section 20.






4.8 Comment on Section 4

The AMEC QPs conclude the following conclusions are appropriate:

Legal opinion supplied supports Commerce’s ownership of the mineral tenure on

the Project.

Work filed in 2011 with relevant regulatory authority updates all Blue River claims

to a common expiry date of 31 March 2021.

There are no known royalties, back-in rights, agreements, or encumbrances

attributed to the claims.

There are no known material issues at the Report effective date that would prevent

negotiation for surface rights to access the Project should they be required in the

future.

Exploration activities have been conducted within the regulatory framework

required by the B.C. Government.

Additional permits will be required for Project development.






5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1 Accessibility

The Project is located 23 km north of the community of Blue River, British Columbia,

approximately 250 km north of the city Kamloops and approximately 90 km south of

the town of Valemount. The property is accessed from B.C. Highway 5 (Yellowhead

Highway) via a 4 km well-groomed gravel road.

The Upper Fir and Bone Creek deposits can be reached from the Bone and Gum

Creek forestry service roads which branch east from Highway 5 approximately 23 km

north of Blue River. The east side of the property can be reached by forest service

roads along the west side of Kinbasket Lake and up Howard Creek. Logging roads on

Serpentine, Bone, Hellroar and Mud Creeks allow four-wheel drive and quad bike

access to most of the property. Access to remaining portions of the property is by

helicopter.

5.2 Climate

The local climate is typical of the interior of British Columbia. The area is part of a “wet

belt” that occupies part of eastern British Columbia. Heavy snow falls almost every

winter, in which temperatures stay close to the freezing point when maritime air

dominates. Rain is frequent in other seasons. Summer days are typically warm or

occasionally hot, with thunderstorms often spawning over the nearby mountains.

In July, the average daily temperature is 16.4°C and the average rainfall accumulation

is 97.5 mm for Blue River (Environment Canada Climate Normals 1971–2000 web site:

http://climate.weatheroffice.gc.ca/climate_normals/index_e.html. In January, the

average daily temperature is -9°C and the average snowfall accumulation is 109 cm

for Blue River. The average snow depth is 83 cm in February. Local rainfall and

snowfall accumulations on parts of the property may be much higher due to elevation

and orographic effects.

Drilling is feasible from mid-May through early to mid-October. Snowfall can exceed

10 m on the property making winter drilling very expensive and difficult, but not

impracticable. It is expected that any future mining activity could be conducted year-

round.






5.3 Local Resources and Infrastructure

There is no existing Project infrastructure. Exploration activities are currently supplied

from Kamloops, Clearwater and Valemount.

The city of Kamloops currently supports mining operations at the New Afton and

Highland Valley mines, and mineral exploration for the surrounding area. Services for

mining operations are reasonably available at Prince George, Vancouver, or

Edmonton.

Power transmission lines, rail, paved, and gravel roads are all adjacent to the Project

near the Yellowhead Highway. The Yellowhead Highway runs sub-parallel to the

North Thompson River. The community of Blue River has a municipal airport for light

aircraft and helicopter support.

The main line of the Canadian National Railway passes through the western part of the

property. Sidings currently exist at Lempriere and Blue River, located 16.5 km north

and 23.7 km south of the Upper Fir deposit respectively. The flat area immediately

north of Bone Creek may be suitable for a siding and is 4.9 km from the Upper Fir

deposit.

The BC Hydro 136,000 volt supply line for the North Thompson valley passes through

the west side of the property adjacent to the rail line. The 20 megawatt Bone Creek

run-of-river hydroelectricity project owned by Transalta Corp was commissioned in

June 2011, and is adjacent to the Project area.

Infrastructure requirements as detailed within the PEA for Project development are

discussed in Section 18 of this Report.

5.4 Physiography

The Project topography ranges from 700 m to 3,100 m elevation above sea level and

is located largely along the steep, west-facing slopes of the Monashee Mountains, to

the east of the North Thompson River.

The highest peak, Mt. Lempriere, is 3,183 m. Ice fields and glaciers dominate the

higher elevations on the property. Significant major tributaries feeding into the North

Thompson River in the area include Serpentine, Pyramid, Gum, Bone, Hellroar and

Mud Creeks.

Mountain slopes are typically covered by thick undergrowth consisting of grasses,

buck brush, devil’s club, and shrubs of willow, alder, rhododendron, huckleberry,






currants, gooseberry, thimbleberry, and raspberry. White spruce is common in

replanted logging areas. Former trails and flat wet areas are typically overgrown by

dense alder and willow. Areas not subjected to recent logging are covered by dense

stands of hemlock, cedar, fir and white pine. Within the area, the tree line is at

approximately 2,000 m elevation. Except for the Paradise Lake, Felix, Howard Creek

and Gum Creek localities, all other carbonatites are below the tree line, and outcrop

exposure is generally poor.


The existing and planned access, infrastructure, availability of staff, the existing power,

water, and communications facilities, the methods whereby goods could be

transported to any proposed mine, and any planned modifications or supporting

studies are reasonably well-established. There is sufficient area in the Project tenure

and in the vicinity of the Upper Fir deposit to support construction of plant, mining and

disposal infrastructure. The requirements to establish such infrastructure are

reasonably well understood by Commerce.

In the opinion of the QPs, the access, physiography, local services, plus existing and

planned infrastructure can support the declaration of mineral resources for the Upper

Fir deposit.

It is expected that any future mining operations will be able to be conducted year-

round.






6.0 HISTORY

6.1 Pre-Commerce Exploration


discovery of vermiculite-bearing carbonatite rock in 1949. A summary of exploration

activities on the Project is described below and summarized in Table 6-1.

Table 6-1: Blue River Exploration History Summary

Year Company Exploration

1949–1951 Oliver E. French Staking and prospecting; discovered vermiculite-bearing

carbonatite near Blue River, discovered uranpyrochlore in

dolomitic carbonatite

1952–1955 St. Eugene Optioned property, geological mapping, prospecting, stripping,

trenching, and sampling

1967-1968 Vestor Staking, reconnaissance surface mapping in the area south of

Paradise Lake

1976 J. Kruszewski Re-staked the area as the Verity and AR claims

1977–1978 J. Kruszewski /

E. Meyers

Magnetometer and scintillometer surveys, trenching and sampling

1980 AMC Optioned property, discovery of Fir and Bone Creek carbonatites

1980–1982 AMC 3,954.2 m of NQ diamond drilling at Verity, Mill, Fir and Bone

Creek

1989 Diegel et al. Government survey discovered two new carbonatite localities

near Serpentine Creek and Gum Creek

2000–Present Commerce Surface mapping, trenching, soil sampling, geophysics, diamond

drilling, metallurgical testing, bulk sampling

Abbreviations: St. Eugene = St. Eugene Mining Corporation Ltd.; Vestor = Vestor Exploration Ltd; AMC =

Anschutz (Canada) Mining Ltd.; Commerce = Commerce Resources Corp.

6.2 Commerce Exploration

In 2000, Commerce acquired the Property, confirmed known tantalum mineralization at

the Fir and Verity carbonatites, and explored for new carbonatite deposits. During the

summer of 2002, the Upper Fir Carbonatite showing was discovered and defined by

core drilling between 2005 and 2011.

Additional work undertaken by Commerce included surface mapping, trenching, soil,

rock chip, grab and channel sampling, geophysics, metallurgical testing, bulk

sampling, and mineral resource estimation.






6.3 Commerce Mineral Resource Estimates

During 2009–2010, Commerce commissioned Caracle Creek International Consulting

Inc. (CCIC) to prepare a mineral resource estimate for the Upper Fir. This estimate

was based on the interpretation of 168 Upper Fir drill holes completed during 2005 to

2008. An initial NI 43-101 compliant mineral resource estimate for the Upper Fir

tantalum and niobium bearing carbonatite was completed in early 2010 by CCIC

(Stone and Selway, 2010).

During 2011, Commerce commissioned AMEC to prepare a mineral resource estimate

for the Upper Fir deposit (Chong and Postolski, 2011) and a preliminary economic

assessment (PEA) (Chong et al., 2011). The current 2012 mineral resource estimate

discussed in Section 14 is an update of the estimate that supports the 2011 PEA.






7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The regional geology is taken largely from Currie (1976), Pell (1987 and 1994),

Gorham (2007), Stone and Selway (2010), and Chong and Postolski (2011a).

British Columbia is divided into three discrete areas hosting carbonatites and alkaline

rocks (Figure 7-1):

Eastern Area: the Foreland Belt, east of the Rocky Mountain Trench

Central Area: the eastern edge of the Omineca Belt

Western Area: the core of the Omineca Belt

The age of emplacement for carbonatites and alkaline rocks of the Eastern and

Central Areas typically range between Devonian–Mississippian (ca. 330–380 Ma).

Some occurrences from the core, or Western Area of the Omineca Belt might be older

(ca. 570 to 770 Ma).

The Eastern Area hosts northwest to southeast trending carbonatite occurrences.

They are often associated with syenite intrusions. Most of the Eastern Area

carbonatites have relatively high niobium and rare earth element (REE) levels, and

little or no tantalum. Some known Eastern Area carbonatite or carbonatite-associated

properties are: Aley, Prince, Ice River and Rock Canyon Creek.

The Central Area carbonatite intrusions occur along the eastern edge of the Omineca

Belt. The carbonatites of the Omineca Belt commonly have high concentrations of

niobium but low rare earth element (REE) values. Known carbonatite complexes

include the Blue River (Upper Fir, Fir, and Verity systems) and Mud Lake areas.

The Western Area includes both intrusive and extrusive carbonatites and syenitic

gneisses in the core of the Omineca Belt. Examples are Mount Copeland, Mount

Grace, and Three Valley Gap.

All of the alkaline and carbonatite complexes and their host rocks within the Omineca

Belt rocks were deformed and metamorphosed during the Jurassic-Cretaceous

Columbian Orogeny and have been subjected to upper amphibolite facies

metamorphism.






Figure 7-1: Tectonic Belts of British Columbia and Carbonatite Occurrences

Note: Adapted after Pell (1994). Figure courtesy of Dahrouge Geological Consulting Ltd., 2011.






7.2 Project Geology

The Project is located in the Central Area within the north-eastern margin of the

Shuswap Metamorphic Complex. The area comprises polyfolded, metamorphosed

Late Proterozoic (ca. 700–550 Ma) supracrustal rocks and is bounded on the east and

west by steep, Eocene age, west-side down normal faults in the southern Rocky

Mountain Trench to the east and the North Thompson valley to the west. The Malton

gneissic complex lies to the north.

The supracrustal rocks are part of a belt dominated by the Late Proterozoic Horsethief

Creek Group and the overlying Kaza Group. The belt is continuous from the northern

Selkirk Mountains in the southeast, through the Monashee Mountains, and into the

Caribou Mountains in the northwest.

The carbonatites within the Blue River Project area are hosted in the Mica Creek

assemblage of the Horsethief Creek Group (Figure 7-2 and Figure 7-3).

7.2.1 Metasedimentary Rocks

Two units of the Mica Creek assemblage underlie much of the study area. The units

are at least 1,000 m thick and comprise the lower pelite unit, and the stratigraphically

overlying semipelite–amphibolite unit (refer to Figure 7-2 and Figure 7-3).

The Mica Creek metasedimentary rock types include biotite gneiss, muscovite–biotite

schist and gneiss, garnet–muscovite–biotite schist and gneiss, calc-silicate–biotite

gneiss, amphibolite, garnet amphibolite, and calc-amphibolite.

The high intensity of deformation precludes determination of tops in metasedimentary

rocks, thus relative ages of individual units are not clear. Layering in the gneiss has

been previously interpreted as relict bedding, but likely is dominantly compositional

segregation due to the metamorphic grade.






Figure 7-2: Blue River Project Local Geology Map

Note: The Upper Fir deposit is located approximately 25 km north of the town Blue River near the western

claim boundary and Yellowhead Highway. See Figure 7-3 for the local geology legend. Figure is courtesy

Dahrouge Geological Consulting Ltd., 2012.






Figure 7-3: Blue River Local Geology Legend (for Figure 7-2)

Note: Figure courtesy Dahrouge Geological Consulting Ltd., 2012.

7.2.2 Gneisses and Schists

Metamorphosed quartzo-feldspathic biotite gneiss is the most abundant lithology that

crops out at surface. Biotite gneiss is ubiquitous and is inter-layered with all other

lithologies on the property.

Outcrops are moderately weathered with characteristic 0.2 to >1 m thick layers of

uniform, massive, medium grained quartz–feldspar–-biotite ± muscovite divided by

recessive schistose bands or fine partings. Fresh surfaces have a uniform,

equigranular, salt and pepper texture of quartz, feldspar and biotite. Muscovite occurs

as thin schistose partings ranging from trace to abundant amounts. Sub-one

millimeter to several centimeter diameter red garnet porphyroblasts occur in varying

amounts. These units are interpreted to represent deformed and moderately

re-crystallized turbidite. Calc-silicate-bearing biotite gneiss has pale green bands a

few centimetres thick that may be related to microscopic traces of actinolite ± diopside.






7.2.3 Amphibolites

The amphibolite units occur as lenses within all gneiss and schist units. They are

typically medium-grained, massive to moderately foliated amphibolites and can contain

red garnets (almandine) typically <1 cm in diameter. Plagioclase and hornblende

occur in varying proportions forming rocks ranging from tonalite to hornblendite

composition (<10% hornblende to >90% hornblende respectively). Weak mineral

lineations are present as observed by the alignment of hornblende. Rare banding is

observed at centimetre scale.

Locally, calc-amphibolite units are distinguished by an increase in mineral grain size, a

strong contrasting black and white colour, local presence of garnets, and effervescent

reaction with dilute hydrochloric acid. The amphibolite units are interpreted as

metamorphosed mafic sills, dikes, and possibly subaqueous flows.

7.2.4 Intrusive Rocks

Ultramafic Rocks

Ultramafic rocks associated with the carbonatites include fine- to medium-grained

pyroxenites and cumulate pyroxene–hornblendites. The ultramafic units likely

represent a metamorphosed ultramafic intrusion associated with mafic volcanism

(amphibolite) roughly the same age as the intruded metasediments (Figure 7-4).

Carbonatite

Both dolomitic carbonatite and calcitic carbonatite occur at Blue River. Dolomitic

carbonatite is often referred to as magnesio-carbonatite, rauhaugite, or beforsite.

Coarse-grained, calcitic carbonatite is also often referred to as calcio-carbonatite or

sövite.

The Blue River Fir carbonatite is approximately 330 million years old (and possibly

older) based on U–Pb geochronology data. The carbonatites were emplaced as dikes

or sills into the metasedimentary rocks prior to the regional deformation and

metamorphism that occurred c.a. 170 Ma.

The carbonatites forms sill-like bodies with average thicknesses of 30 m, ranging

between 5 m to about 90 m thick, and with strike lengths ranging between 50 m to

1,100 m (Figures 7-4 to 7-8). Bedding-parallel foliation in metasedimentary gneisses

and the contacts of carbonatite intrusions generally strike 335° and 155° with shallow

to moderate northeast and southeast dips.






Figure 7-4: Deposit Area Surface Geology Map

Note: The interpreted Upper Fir deposit surface expression is shown in blue. Inset shows the deposit

location at the Project. Figure courtesy Dahrouge Geological Consulting Ltd., 2012.






Figure 7-5: Drill Collar and Vertical Section Locations

Note: See Figure 7-6 for the longitudinal section A – A’; Figure 7-7 for section 5796740 N; Figure 7-8 for

section 5796425 N. Carbonatite surface expressions are coloured blue. Bulk sample locations are noted

at BS1 and BS2. Figure courtesy Dahrouge Geological Consulting Ltd., 2012.






Figure 7-6: Longitudinal Section A – A’ (view SE)

Note: The figure illustrates the carbonatite geometry and its NE – SW geological continuity approximately perpendicular to the local SE trending fold

hinge orientations. Upper Fir Carbonatite = blue coloured domain. View is to the south-east. Section influence is +/- 25 m. Figure by AMEC, 2012.

Overburden

Magnesiocarbonatite

Calciocarbonatite

Silicocarbonatite

Fenite

Skarn

Amphibolite

Garnet amphibolite

Pegmatite

Garnet gneiss

Diopside gneiss

Gneiss

Mylonite

Ultramafic

LITHOPLT: Lithology Colours

Viewport1

Bone Carbonatite

Upper Fir

Carbonatite

Legend

A (NE) A’ (SW)

50 m






Figure 7-7: Geology Section 5796740 N

Note: Illustrates the carbonatite geometry, its east-west geological continuity, and relationship between drilled thickness versus true thickness.

Additional 2010 drilling (F10 prefixed holes) has reasonably improved the local geological interpretation. Upper Fir Carbonatite = blue coloured

domain. View is to the north. Section influence is +/- 25 m. Figure by AMEC, 2012.

Overburden

Magnesiocarbonatite

Calciocarbonatite

Silicocarbonatite

Fenite

Skarn

Amphibolite

Garnet amphibolite

Pegmatite

Garnet gneiss

Diopside gneiss

Gneiss

Mylonite

Ultramafic


Viewport1

Bone Carbonatite

Upper Fir

Carbonatite

Fenite

Legend

West East

50 m






Figure 7-8: Geology Section 5796425 N

Note: Illustrates the carbonatite geometry, east-west geological continuity, folding, and relationship between drilled. Additional 2010 drilling (F10

prefixed holes) has improved the local geological interpretation. Upper Fir Carbonatite = blue coloured domain. View is to the north. Section influence

is +/- 25 m. Figure by AMEC, 2012.

Overburden

Magnesiocarbonatite

Calciocarbonatite

Silicocarbonatite

Fenite

Skarn

Amphibolite

Garnet amphibolite

Pegmatite

Garnet gneiss

Diopside gneiss

Gneiss

Mylonite

Ultramafic


Viewport1

Bone Carbonatite

Upper Fir

Carbonatite

Fenite

Legend

West East

(A)

(B)

(C)

50 m






Dolomitic and calcitic carbonatite usually form separate bodies but can occur together

within single intrusions. At Blue River, dolomitic carbonatite typically makes up the

cores of the carbonatite bodies. Crosscutting or gradational relationships can be

observed from one variety of carbonatite into another.

Dolomitic and calcitic carbonatites are medium to coarse-grained and have secondary

tectonically-imposed textures. A cataclastic (porphyroclastic) texture is common in all

the carbonatites. Most exposures display layering defined by varying quantities of

accessory minerals. Accessory minerals include amphibole, pyroxene, phlogopite,

olivine, magnetite, apatite, pyrite/pyrrhotite, ilmenite, zircon, and the tantalum and

niobium bearing minerals pyrochlore and ferrocolumbite.

Contacts between carbonatite and the host metasediments are typically sharp and

mantled by zones of metasomatized host rock, known as fenite.

At Blue River, the fenite rocks commonly, but not always, envelope the carbonatite

rocks and can extend up to 50 m from the carbonatite intrusions. The

metasedimentary host rocks are characterized by foliated calcite-richterite-biotite

(± apatite, ± vermiculite) rock. Of lesser importance are contact metasomatic veins

commonly less than 1 m thick that comprise amphibole-pyroxene (± vermiculite

± carbonate).

Fold indicators in core intersections observed in holes F08-150 and F08-151 (Figure

7-8: locations A, B, and C) are shown in Figure 7-9 to Figure 7-11 respectively.






Figure 7-9: Fold Indicators (Hole F08-150: 121.8 m to 129.8 m)

Notes: Left: Hole F08-150: 121.8m to 129.8m. HQ diameter diamond drill core. Hole was drilled vertical.

Compositional layering (L) of biotite-quartz gneiss is typically at a high angle to the core axis indicating a

sub-horizontal attitude when related to the sub-vertical dip of the drill hole. A ptygmatic fold is also

observed (T). Top of hole is towards the top left of photo. Right: Hole F08-150: 125 m. Indications of

folding include asymmetric parasitic folds with short and long limbs (P) bracketed by sub-horizontal

compositional layering. Centimetre scale ptygmatic folds (T) are noted. Photos from Chong (2010).

Figure 7-10: Fold Indicators (Hole F08-150: 143.5 m and 147.0 m)

Note: Left: F08-150: 143.5 m. Right: F08-150: 147.0 m. Indications of folding include high and low

angle layering indicating possible fold closures. Top of hole is towards the top left of photos. Photos from

Chong (2010).

P

T

L, T






Figure 7-11: Fold Indicators (Hole F08-151: 204.0 m to 204.5 m)

Note: F08-151: 204.0 m to 204.5 m (top to bottom). Fold indicators include repetition of carbonatite to

biotite-quartz gneiss and back into carbonatite. Upper carbonatite in figure has a contact at a high angle

to core axis indicating that it is a flat lying contact (upper right dashed line). Middle gneiss has a trend of

compositional layering with high - to low - to high angles relative to core axis (middle curved dashed line

and lower middle dashed line). Black box highlights a possible fold closure which gives a characteristic

bulls-eye appearance to the layering. Top of hole is towards the top left of photo. Photos from Chong

(2010).

7.2.5 Pegmatite Dykes

Pegmatite dykes and pods up to 500 m long and 15 m thick crosscut all lithologies

throughout the property. At least some of the pegmatites are folded. Among the

pegmatites, two mineralogically-distinct types exist:

Two-mica (± garnet, ± tourmaline) granitic pegmatites

Syenitic pegmatites with minor biotite (± amphibole, ± pyroxene).

Where pegmatites cross-cut carbonatite, a coarse-grained skarn assemblage of

calcite, amphibole and/or diopside is developed at the contact.

7.3 Structural Geology and Metamorphism

The structural geology is summarized largely from Kraft (2010), Ghent et al. (1977),

Simony et al. (1980), and Raeside and Simony (1983).






The style of structural deformation at the Project directly impacts the carbonatite

geometry. A deformation model was developed on behalf of Commerce by J. Kraft

during 2009 and early 2010. The structural deformation model was confirmed and

enhanced by field work completed by J. Kraft during July and September of 2010 and

by structural interpretation of sub-surface information by Gervais (2011). The following

descriptions include the 2010 supporting observations and interpretations. Reviews by

SRK (Couture and Nash, 2011a, 2011b) and Touchstone Geoscience Inc.

(Touchstone) (Lee, 2012) generally support the deformation model.

Three phases of compressional deformation have been mapped throughout most of

the region, from the northern Selkirk Mountains into the Cariboo Mountains. At the

Project, at least two additional deformation events are observed.

The first deformation event (D1) produced large recumbent folds (F1) with limbs

approximately 50 km long and an associated early foliation (S1). Features from F1 are

not observed within the immediate deposit area.

The second deformation event (D2) is associated with peak, mid-amphibolite facies

metamorphism with an associated foliation (S2). In general the S1 and S2 foliations

can rarely be distinguished from one another in the field and are mapped as S1+S2.

D2 has created boudinage, or pinch and swell features attributed to competency

contrasts between rock type layers.

The third deformation event (D3) is characterized by centimetre to decametre scale

recumbent folds (F3) that deform the S2 foliation. Axial planar schistosity that deforms

S1+S2, within micaceous lithologies such as fenite, is known as S3. The style and

attitude of F3 folds are variable, but axial planes are generally southeast-dipping.

Touchstone (Lee, 2012) suggests that transpositional folding associated with D2 may

play a larger part in carbonatite geometry than previously considered.

The fourth deformation event (D4) is characterized by inclined, southwest trending

folds that re-fold larger F3 folds in the deposit area. Open to tight upright folds with

east to southeast hinges occur sporadically. D4 is suggested to include thrust faulting

with top to the southwest vergence.

The fifth deformation event (D5) is described as a brittle extensional event

characterized by normal faults with slickensides, weak quartz-pyrite alteration of wall

rocks, and cross-cutting relationships with D4 structures.

A distinctive augen gneiss unit intersected below the carbonatite in many holes may

represent a high strain zone defining a lower boundary for the carbonatites.






Folding is observed in wall rocks adjacent to the carbonatite in outcrop and in drill

core, and is interpreted for carbonatite intercepts on large scale geological sections.

Within carbonatite, compressional deformation with weak southeast elongation is

suggested by zones with cataclastic to mylonitic foliation (Chudy and Ulry, 2012) and

weakly to moderately developed mineral lineations defined by amphiboles.

Carbonatite bodies are folded at metres to deposit scale, however their thickness and

massive (non-layered) nature makes observation of folding indicators within the

carbonatite comparatively difficult to observe in outcrop or drill core.

7.4 Geochronology

The geochronology is summarized largely from Pell (1994), Simonetti (2008), and

Gervais (2009).

An uranium–lead date of about 325 Ma was obtained from zircons from the Verity

carbonatite. A lead–lead date of 332.5 ± 5.7 Ma age was obtained from zircons for the

Upper Fir carbonatite. A preliminary uranium–lead date of 328 ± 30 Ma was obtained

from zircons from the Mud Lake area carbonatite. Zircons separated from syenite at

Paradise Lake yielded a uranium–lead age of about 340 Ma and lead-lead ages of

about 351 and 363 Ma.

Ongoing research on U/Th/Pb dating of zircons and monazites from the property

shows a complex thermal history, indicating that the age of emplacement of the Blue

River area carbonatites may be older than initial results have shown (pers. comm.

L. Millonig to J. Gorham).

7.5 Carbonatites

The Upper Fir and Bone Creek carbonatites are assessed in the mining plan

discussed in Section 16 of this report. However, for Report completeness purposes,

all of the significant carbonatite deposits are described in this subsection.

7.5.1 Fir Carbonatite

Information summarized in this subsection on the Fir deposit is from the British

Columbia Geological Survey website.

The Fir showing is located 1.25 km north of the Bone Creek carbonatite. Carbonatite

consisting of dolomitic and lesser calcitic carbonatite occurs as sills within the quartz-

hornblende-mica schist of the Semipelite Amphibolite division of the Horsethief Creek






Group. Other lithologies include amphibole-biotite schist, biotite-muscovite gneiss and

amphibole-biotite-garnet gneiss.

The Fir carbonatite has a likely strike extent of at least 400 m in a northerly direction

based on outcrop exposures. A 2 m exposure of dolomitic carbonatite was located

400 m north of the discovery outcrops. Additional exploration potential occurs south of

the Fir discovery outcrops based upon tantalum and niobium soil anomalies. Dolomitic

outcrops are coarsely crystalline and typically weather white. Accessory minerals in

the carbonatites include apatite, amphibole, olivine, magnetite, pyrite, pyrrhotite,

pyrochlore, and columbite. The dolomitic carbonatite is almost devoid of biotite and

magnetite. Three distinct textures were observed: breccias composed of tightly-

packed dolomite fragments within a finely crystalline dolomite groundmass; a

porphyritic texture with ghost dolomitic crystals in a fine-grained matrix; and a massive

texture with local banding of accessory minerals.

Tantalum and niobium mineralization in the Fir carbonatite occurs as the minerals

pyrochlore and columbite. The Fir carbonatite has the highest background niobium

and tantalum values of all carbonatites in the area. Tantalum averages greater than

0.015 per cent. Sampling of the discovery outcrops returned assays of 1.02% Nb2O5,

0.06% Ta2O5, and 6.31% P2O5. A sample from drill hole BC19 returned values of

0.18% tantalum and 8.51% phosphate from 119.0–119.6 m. The anomalous value

may be the result of a pyrochlore–apatite band in the sample.

7.5.2 Verity Carbonatite

The Verity carbonatite is located about 40 km north of the community of Blue River.

The Verity carbonatite has the most varied stratigraphy of all the carbonatites in the

area.

The Verity carbonatite consists of banded dolomitic and calcitic carbonatite that locally

intrude each other. It occurs as a 15 m to 30 m thick sill within quartz-hornblende-mica

schist of the Horsethief Creek Group. It can be traced up the hillside for 800 m to the

east–northeast.

A tectonic breccia showing hairline fractures is common in the dolomitic carbonatite. A

banded texture caused by layering of the accessory minerals apatite, amphibole,

olivine, magnetite, vermiculite, biotite, pyrite, pyrrhotite, pyrochlore, columbite, and

zircon is common in calcitic carbonatite and less developed in the dolomitic

carbonatite. Coarse olivine and apatite in calcitic dolomite form bands 1 cm to 5 cm

thick. Magnetite occurs as discontinuous lenses in calcitic carbonatite layers; the

lenses can be much as 20 cm in diameter.






Tantalum and niobium mineralization in the Verity carbonatite occurs in the minerals

pyrochlore and columbite. The pyrochlore and columbite crystals occur as

octahedrons that can reach 4 cm diameter. Calcitic carbonatite at the Verity

occurrence also contains greater than 10.8% phosphate and has abundant apatite

relative to other nearby carbonatites at the Project. Rare earth elements are

interpreted to be hosted in fluorine-rich carbonate.

7.5.3 Exploration Targets

Geochemical sampling has outlined a number of potential exploration targets,

including:

Upper Fir Extension target: strong tantalum anomalies on four adjacent soil lines

north–northwest of Bulk Sample Pit #2 indicate that the carbonatite subcrop likely

extends to the north, past the current drill coverage

Bone Creek Extension target: strong tantalum-in-soil anomalies on two widely-

spaced lines centered at UTM 5,797,000 N indicate a near surface carbonatite

body that is on strike with the Bone Creek carbonatite

Fir Exploration target: strong tantalum-in-soil anomalies on widely-spaced lines

located north, south and above the known Fir showing indicate possible extensions

of the Fir carbonatite

Mt. Cheadle Exploration target: a large diffuse tantalum-in-soil anomaly, with

several spikes, stretching over 2 km, is located north of Gum Creek and along

strike from the Upper Fir carbonatite

3050 Road target: Strong tantalum anomalies on soil lines to the north and east of

current drilling on the Upper Fir deposit near 3050 Road indicate another

carbonatite body may be located above the known extents of the deposit.

Target locations are indicated on Figure 7-12.






Figure 7-12: Exploration Target Location Surface Map

Note: Figure courtesy Dahrouge Geological Consulting Ltd., 2012.






Soil sample results indicate the Upper Fir carbonatite has exploration potential directly

northward of known deposit extents. Additional resource definition drilling is

warranted.

Soil sample results indicate the Bone Creek and Fir carbonatites have additional

exploration potential along, and across, strike. Additional in-fill soil sampling is

warranted prior to diamond drilling to assess for potential connections with the Upper

Fir carbonatite.

7.6 Mineralogy

The discussion in this section is summarized largely from observations during AMEC’s

site visits, Chudy (2008 and 2010), Chudy and Ulry (2012), Woolley and Kempe

(1989), Aaquist (1982a), and Mariano (2000).

There are two principal and one minor niobium- or tantalum-bearing minerals known at

the Project.

The minerals, using generic end-member compositions, are:

Ferrocolumbite ............ (Fe,Mn,Mg)(Nb,Ta)2O6

Pyrochlore ..... (Ca,Na,U)2(Nb,Ti,Ta)2O6(OH,F)

Fersmite ..... (Ca,Ce,Na)(Nb,Ta,Ti)2(O,OH,F)6.

7.6.1 Ferrocolumbite

Ferrocolumbite occurs predominantly in medium to coarse-grained, granoblastic

dolomitic carbonatites which typically form relatively thin intervals (<6 m) or occur at

margins of thicker intervals of carbonatite. These carbonatites contain the amphibole

minerals winchite and barroisite. Ferrocolumbite forms subhedral to anhedral,

sometimes strongly poikilitic, individual grains or agglomerates of grains.

Mineral liberation analyses show that the majority (~80%) of liberated grains are less

than 110 μm in diameter. Locally, individual grains and agglomerates of

ferrocolumbite may exceed 2 cm in diameter.

Ferrocolumbite grains from marginal zones may contain large amounts of tiny

inclusions such as thorite (Th-silicate), monazite (La, Ce-phosphate) and pyrochlore.

Ferrocolumbite may also occur sporadically as inclusions in apatite and amphibole. It

is often associated with layers and micro-veins of apatite that fill the interstices

between anhedral ferroan-dolomite grains.






7.6.2 Pyrochlore

Pyrochlore occurs predominantly in the fine-grained and porphyroblastic dolomitic

carbonatite containing the amphibole richterite, which is commonly developed in the

central portions of carbonatite intervals greater than 10 m thick. Such zones are less

abundant or absent in thinner carbonatite intersections. Pyrochlore is the only

tantalum mineral in the calcitic carbonatites that occurs in accessory amounts. Black

and brownish-yellow coloured varieties of pyrochlore are present.

The majority of the pyrochlore occurs as liberated grains in the dolomitic matrix. The

vast majority (~ 85%) of pyrochlore forms subhedral to anhedral, rounded grains less

than 200 μm in diameter. There are local larger grains and agglomerates, as well as

accumulations or veins less than a few tens of centimetres in width with high

pyrochlore abundance. This style of mineralization can result in locally high tantalum

values (> 450 ppm Ta).

Pyrochlore also occurs as inclusions in amphiboles (richterite), fluorapatite, and in

ferrocolumbite. In some rare cases the pyrochlore grains can be coated with a thin

film of pyrrhotite or pyrite.

7.6.3 Fersmite

Fersmite occurs as anhedral inclusions in apatite, primarily at the Verity carbonatite

and is considered a minor economic mineral at the Project.

7.6.4 Fenite Mineralization

Mineralization in the fenite is dominantly ferrocolumbite, concentrated in apatite-rich

layers. Ilmenite, with ferrocolumbite inclusions, appears to be a subdominant source

of both niobium and tantalum. Niobium and tantalum grades within fenite at Blue River

are considered to be sub-economic, but locally fenite may provide grade-bearing

mining-dilution material.

7.6.5 Mineral Zoning

Mineral zoning, or distribution, of ferrocolumbite and pyrochlore within the carbonatites

is not clear due to the variable thicknesses and polyfolded geometry of the carbonatite.

On-going research (Chudy and Ulry, 2012) supports the association of dominantly

pyrochlore mineralization with richterite-bearing, fine-grained, foliated carbonatite

defining zones of retrograde deformation. Ferrocolumbite mineralization is dominant in

winchite–barroisite-bearing, coarse-grained gneissic carbonatite. Carbonatite of






intermediate porphyroclastic textures may contain both minerals. Further work is

required to improve the understanding of the mineral zoning and to locate potential

material types required to support metallurgical testwork.


In the opinion of the QPs, knowledge of the deposit settings, lithologies, and structural

and alteration controls on mineralization are sufficient to support mineral resource

estimation.

The mineralization style of the Project deposit is sufficiently well understood to support


Prospects and targets are at an earlier stage of exploration, and their lithologies,

structural, and alteration controls on mineralization are at present not well understood

and hence more work is required to support estimation of mineral resources.






8.0 DEPOSIT TYPES

The mineralization identified to date at the Project is consistent with magmatic,

carbonatite-associated deposits. Carbonatite-associated deposits are classified as

either magmatic, replacement, or residual. Global examples of magmatic carbonatite

complexes or deposits include: Oka, Niobec (St. Honore) (Quebec); Kovdor (Russia);

Iron Hill (Colorado); and Gardiner (Greenland) (Mitchell, 2010). Examples of

replacement carbonatite deposits are Rock Canyon (B.C.), Bayan Obo (China), and

Palabora (South Africa). Araxa and Catalao (Brazil) are classified as residual

carbonatite deposits due to the degree of lateritic weathering. Carbonatites are the

main source of niobium +/- tantalum, and important sources of rare earth elements.

Magmatic carbonatite deposits have the following common features (Birkett and

Simandl, 1999).

Commodities: Niobium, tantalum, rare earth elements, phosphate, vermiculite,

copper, titanium, strontium, fluorine, thorium, uranium, magnetite.

Geological Setting: Carbonatites intrude all types of rocks and are emplaced at a

variety of depths. Carbonatites occur mainly in a continental environment, rarely in

oceanic environments (Canary Islands) and are generally related to large-scale,

intra-plate fractures, grabens or rifts that correlate with periods of extension and

may be associated with broad zones of uplift.

Age of Mineralization: Carbonatite intrusions are early Precambrian to Recent in

age; they appear to be increasingly abundant with decreasing age. In British

Columbia, carbonatites are mostly upper Devonian, Mississippian or Eocambrian

in age.

Host Rocks: Host rocks are varied, including calcite carbonatite (sövite), dolomite

carbonatite (beforsite), ferroan or ankeritic calcite-rich carbonatite

(ferrocarbonatite), magnetite-olivine-apatite ± phlogopite rock, nephelinite, syenite,

pyroxenite, peridotite and phonolite. Carbonatite lava flows and pyroclastic rocks

are not known to contain economic mineralization. Country rocks are of various

types and metamorphic grades.

Deposit Form: Carbonatites commonly occur as small, pipe-like bodies, dikes,

sills, small plugs or irregular masses. The typical pipe-like bodies have sub-

circular or elliptical cross sections and are up to 3-4 km in diameter. Magmatic

mineralization within pipe-like carbonatites is commonly found in crescent-shaped

and steeply-dipping zones. Metasomatic mineralization occurs as irregular forms

or veins. Residual and other weathering-related deposits are controlled by

topography, depth of weathering and drainage development.






Deposit Mineralogy:

o Magmatic: bastnaesite, pyrochlore, columbite, apatite, anatase, zircon,

baddeleyite, magnetite, monazite, parisite, fersmite.

o Replacement/Veins: fluorite, vermiculite, bornite, chalcopyrite and other

sulphides, hematite.

o Residual: anatase, pyrochlore and apatite, locally crandallite-group minerals

containing rare earth elements.

Gangue Mineralogy: Calcite, dolomite, siderite, ferroan calcite, ankerite, hematite,

biotite, titanite, olivine, quartz.

Alteration: A fenite halo (alkali metasomatized country rocks) commonly surrounds

carbonatite intrusions; alteration mineralogy depends largely on the composition of

the host rock. Most fenites are zones of desilicification with addition of Fe3+, Na

and K.

Mineralization Controls: Intrusive form and cooling history control primary igneous

deposits (fractional crystallization). Tectonic and local structural controls influence

the forms of metasomatic mineralization. The depth of weathering and drainage

patterns control residual pyrochlore and apatite deposits, and vermiculite deposits.

Many features of the mineralization identified within the Project to date are analogous

to magmatic carbonatite deposits, in particular the Oka (Husereau Hill) and Niobec (St.

Honoré) deposits in Quebec.

Key features of the Blue River deposits supporting a magmatic carbonatite model are:

Commodities: niobium and tantalum

Geological Setting: occurs along the eastern portion of the Omineca Crystalline

Belt and hence its tectonic setting is along a large scale zone with associated uplift

Age of Mineralization: data yields results of about 330 Ma which is consistent with

other British Columbia carbonatite deposits

Host Rocks: dolomite and calcite-rich carbonatite intrusion rocks

Deposit Form: the Blue River carbonatites occur as sills and dykes

Deposit Mineralogy: ferrocolumbite and pyrochlore

Gangue Mineralogy: dolomite, calcite, amphibole (richterite), quartz, pyroxene,

phlogopite, olivine, magnetite, apatite, pyrite/pyrrhotite, ilmenite, and zircon

Geochemistry: high strontium levels (>5,000 ppm)

Alteration: Fenite halos occur around most carbonatites at Blue River






Mineralization Controls: Carbonatites are the main host rocks to the Ta and Nb

rich minerals pyrochlore and ferrocolumbite. The Blue River carbonatites have

been deformed by multiple episodes of folding and faulting. The internal cooling

history of the deposit is not clear. The spatial distribution of the Ta and Nb rich

minerals varies throughout the carbonatite .


In the opinion of the QPs,

A polyfolded sill-like carbonatite model suitably describes the Blue River deposits

The deposit concepts being applied as the basis for exploration planning at the

Project are reasonable.






9.0 EXPLORATION


discovery of vermiculite-bearing carbonatite rock in 1949.

Commerce acquired parts of the current property in 2000 and initiated exploration for

new carbonatite deposits which culminated in the drilling of the Upper Fir carbonatite

and delineation of the Fir and Upper Fir–Bone Creek carbonatite system.

9.1 Grids and Surveys

All surveys to date are in UTM NAD83 Zone 11 coordinates.

In 2007, orthophotography and Lidar surveys were flown to create a 1:2,000 base map

of the Upper Fir area. A topography map with 2 m contour intervals was created by

Eagle Mapping Ltd.

9.2 Geological Mapping

Geological and structural mapping has been completed at 1:2,500 scale on a

continuous basis since 2006. The mapping area coverage is between Bone and Gum

Creeks; and from the North Thompson River to the ridge top located about 3 km east

on the slope that is known locally as either Fir or Cedar Mountain. The mapping area

includes the Fir, Upper Fir, Bone Creek (considered a single system) and Gum

carbonatites plus the nearby Hodgie Zone.

Mapping was used to determine the outcrop of carbonatite and provide geological and

structural data.

9.3 Geochemical Sampling

9.3.1 Stream Sediment Sampling

Reconnaissance stream sediment sampling was completed during 2001 to 2003, and

2006 to 2007. During 2008, 531 stream sediment samples were collected and

analysed for the streams throughout the entire property. The key exploration

pathfinder elements at Blue River are tantalum, niobium, and rare earth elements.

Detailed sample analysis using microscopic mineral characterization was utilized,

focusing on identifying pyrochlore, apatite, richterite, and monazite as pathfinder

minerals.






During the 2009 field season, a total of 20 stream pan concentrate samples were

taken in the Fir and Mud Creek areas to follow up on creek-mouth areas inaccessible

during the 2008 field season. Samples were analyzed at Acme Laboratories in

Vancouver B.C. (Acme) primarily for tantalum, niobium, rare earth elements,

phosphate, and carbonate using Acme’s 4B trace element package (lithium

metaborate fusion-ICP-MS technique).

Several samples anomalous in tantalum and niobium indicate that the Fir carbonatite

likely extends further south.

9.3.2 Soil Sampling

Soil sampling (B-horizon) has proven the best way to follow up on stream pan

concentrate sampling in the Blue River area as the niobium–tantalum-bearing

ferrocolumbite and pyrochlore are residual in soils. The key exploration pathfinder

elements from soil sampling are tantalum and niobium. During 2002, follow-up on an

anomalous stream sediment sample led to the discovery of the Upper Fir carbonatite.

Reconnaissance soil sampling was completed during 2001 to 2003 and 2006 to 2010

(Table 9-1). During the 2009 season, 1,694 soil samples were collected from several

area grids. Sample grids typically have 200 m spaced lines and samples are taken at

25 m intervals. Soil sampling in 2010 followed up on 2009 soil sampling anomalies on

the west slope of Mount Cheadle and extended the grid south towards Gum Creek.

A total of 477 samples were taken by Dahrouge and analyzed by Acme using Acme’s

4Bpackage (lithium metaborate fusion-ICP-MS technique). A broad area of tantalum

and niobium anomalies stretching north of Gum Creek indicates a possible extension

of the Fir/Upper Fir carbonatite system.

Table 9-1: Soil Sample Campaigns

Year

Number of

Samples Grids

2001 144 Verity, Fir



2006 308 Fir, Bone Cr, Switch Cr

2007 1,996 Fir, Bone Cr, South Fir, Switch Cr, Hellroar, Tailings area, Mt. Cheadle,

Pyramid Creek

2008 4,081 Bone Creek, Hellroar Creek, Mount Cheadle, Mud Lake, and Upper Fir

2009 1,694 Fir, Upper Fir, Hellroar, Switch Creek

2010 477 Mt. Cheadle






9.3.3 Rock Chip, Grab, and Channel Sampling

Rock samples have been taken as part of prospecting and mapping activities on the

Property since 2000 (Table 9-2). Rock samples from various new and known

localities, primarily on the Wasted claims located west of the North Thompson River,

were taken during 2010 prospecting to test for or verify the presence and abundance

of tantalum–niobium and rare earth mineralization. A total of 25 in situ bedrock grab

and chip samples were taken at Miledge Creek, the Hodgie Zone, the Felix, and the

Mona carbonatites.

The Mona carbonatite, south of the Fir carbonatite, is interpreted to be a late stage

remobilization of carbonatite from the Fir system. One continuous chip sample of

fenite at the Mona carbonatite averaged 1,474 ppm Nb and 0.66% total rare earth

elements (TREE) over 28 cm. Two grab samples of carbonatite from the Mona

averaged 1,743 ppm Nb and 1.36% TREE and 6 ppm Nb and 3.66% TREE

respectively.

Table 9-2: Rock Sample Campaigns

Year

Number of

Samples Type Area

2000 15 Chip, grab Verity, Roadside

2001 17 Chip, grab, float Fir, Bone Creek, Roadside

2002 20 Grab, float Fir, Bone Cr, Serpentine, Gum Cr

2003 3 Grab Upper Fir

2006 131 Chip, grab, float Upper Fir, Bone Cr, Mt Cheadle, Gum Cr,

Paradise L, Verity, Switch Cr, Serpentine

2007 110 Chip, grab, float Upper Fir, Bone Cr, Serpentine, Paradise L, Mt

Cheadle, Windfall Cr, Howard Cr, Gum Cr,

Pancake Cr

2008 117 Chip, grab, float, channel Upper Fir, Bone Cr ultramafic, Felix, Gum Cr,

Hodgie, Little Chicago, Mud Cr, Serpentine

2009 113 Chip, grab, float, channel Fir, Gum, Mud Cr, Howard Cr, Paradise,

Switch Cr-Roadside,

2010 25 Chip, grab, regolith Wasted Claims, Felix, Mona, Hodgie

Three samples of the Felix carbonatite confirmed that it is not a tantalum-bearing unit,

with Ta values below 1 ppm and TREE values below 0.01%. One sample of

clinopyroxene-amphibole schist from the Hodgie Zone returned only 0.02% TREE.

Sampling on the Wasted claims along Miledge Creek on the west side of the North

Thompson River was undertaken to aid in condemnation of potential waste rock or






tailings storage sites. The highest value found was a pyrite-muscovite schist yielding

49.5 ppm Nb, 3.3 ppm Ta, and 0.05% TREE.

9.4 Bulk Sampling

A bulk sample program was undertaken in the fall of 2008 by Dahrouge as part of on-

going evaluation of the Upper Fir carbonatite. Approximately 2,000 t from a 10,000 t

permitted volume were extracted from three bulk sample pits (BS-1, BS-2, and BS-3;

refer to locations shown in Figure 7-5) and placed into 75 t to 150 t stockpiles that

were comprised largely of -50 cm carbonatite material. The stockpiles were stored on

a lined, well-drained pad at the Project site for later metallurgical testing.

For each pit area, geological mapping was completed along with sampling of blast-

hole material and channel samples of the bench walls. Both gneissic and

porphyroclastic metamorphic textures and structures were exposed in the sample pits.

Microscopic examination of oxide phases in the bulk sample material indicated that

pyrochlore was the dominant mineral in pit BS-1 with the exception of benches at the

upper and lower contacts. Ferrocolumbite with subordinate amounts of pyrochlore

forms the mineralization in pits BS-2 and BS-3.

Pit BS-1 was excavated in dominantly fine- to medium-grained, granular, foliated,

apatite-bearing, dolomitic carbonatite. Pits BS-2 and BS-3 were excavated in

dominantly light-grey, coarse-grained, porphyroclastic, apatite-bearing, dolomitic

carbonatite. Crosscutting veins of dark green actinolite–calcite–diopside that are as

much as 20 cm wide are common. Contacts in each pit are marked by approximately

1 m of fenite, with contorted layers of dolomitic carbonatite up to 10 cm thick.

Material from the 2008 bulk sampling program appears to provide a sufficient range of

tantalum and niobium grades to represent the Upper Fir carbonatite mineralization for

initial metallurgical testing.

Both pits, BS-1 and BS-2, were stabilized and grass seeded, while pit BS-3 was

backfilled and reclaimed during 2009 and 2010. The bulk sampling permit expired on

31 December 2009 and has not been renewed, although bulk sample material remains

stockpiled on the site for future testwork.

9.5 Research Programs

Doctoral (Thomas Chudy 2009 to present) and post-doctoral (Leo Millonig 2010 to

present) studies on the geology, petrology and microscopy of the carbonatite






mineralization at Blue River are underway at the University of British Columbia.

Publication of these studies is expected in late 2012.


In the opinion of the QPs, the exploration programs completed to date are appropriate

to the style of the deposits and prospects within the Project. The exploration and

research work supports the genetic and affinity interpretations.

The project retains significant exploration potential for carbonatite-hosted tantalum–

niobium mineralization.






10.0 DRILLING

Core (diamond) drilling is the most extensively used exploration tool at Blue River.

AMEC received a drilling database from Commerce named CCE_Upper_Fir that had a

database closure date of 29 September 2011. The database comprises a total of 269

core drill holes within the Upper Fir, Bone Creek and Fir (Lower) carbonatites

consisting of 54,065 m of HQ (63.5 mm) and NQ (47.6 mm) diameter coring.

Table 10-1 lists the core drilling campaigns at the Upper Fir and Bone Creek deposits.

Core drilling by Commerce commenced in 2005 and continues to the present. Drill

locations are included in Figure 10-1.

Of the 269 core drill holes, 237 drill holes totalling 50,395 m of HQ diameter core

(12,736 samples) are used to support the Mineral Resource estimate. Of the 269 core

holes, Commerce drilled 11 holes in the Fir carbonatite area which are not part of this

Mineral Resource update. Of the 269 core holes, 21 legacy holes were drilled by

operators prior to Commerce’s involvement in the project. No sample intervals are

present in the database for the 21 legacy holes as the assay data could not be verified.

Therefore, these holes have not been used to estimate grades in this Mineral

Resource update, but they were used to interpret geology.

An additional 34 core holes, totalling 8,715 m of HQ drill core were drilled in 2011. The

2011 holes were not used in the preparation of this 2012 Mineral Resource update

presented in this report.

Preliminary results from 34 holes drilled in 2011 were reviewed by AMEC after

completion of the current Mineral Resource estimate update. Drill log lithology

information from these holes was reviewed on screen against the 3D carbonatite

model used in the resource estimation. The new 2011 drilling information generally

supports the geological interpretation. Some discrepancies were observed which

warrant local re-interpretation for future updates. Laboratory analyses of the 2011 drill

core samples are now complete and the quality control results are being reviewed by

Commerce. The locations of the 2011 drill collars are shown in Figure 10-1 together

with the 2005-2010 drill holes.

Table 10-2 lists the three bulk sample pits (BS-1, BS-2, and BS-3) and four trenches

(TR0, 0A, 1, and 1A) which have been sampled. These were only used for geology

interpretation and domain modeling, not for grade interpolation.






Table 10-1: Drill Campaign Summary

Category Deposit Operator Year # Holes Series Type # Metres # Samples % Samples

Resource Bone Creek Commerce 2005 4 CF05-01 to CF05-04 HQ 300 14 <1%

Resource Upper Fir Commerce 2005 4 CF0505 to CF0508 HQ 505 44 <1%

Resource Upper Fir Commerce 2006 17 CF0601 to CF0617 HQ 3,021 1,139 9%

Resource Upper Fir Commerce 2007 18 F0718 to F0735 HQ 4,310 1,053 8%

Resource Upper Fir Commerce 2008 118 F08-36 to F08-153 HQ 23,723 5,126 40%

Resource Upper Fir Commerce 2009 22 F09-154 to F09-176 HQ 5,587 842 7%

Resource Upper Fir Commerce 2010 54 F10-177 to F10-230 HQ 12,949 4518 36%

Resource Subtotal 237 50,395 12,736 100%

Historical Bone Creek AMC 1980-1981 17 BC-1 to BC-17 NQ 697 na-

Historical Fir AMC 1981 4 BC1-18 to BC-21 NQ 829 na-

Target Fir

(twins)

Commerce 2001-2002 11 F-01 to F-11 HQ 2,144 na-

Target Subtotal 32 3,670

Total Drilling 269 54,065

Notes:

Abbreviations: AMC = Anschutz Mining Corp.

The Commerce 2011 campaign comprises 34 HQ diameter drill holes totalling 8,715 m. These holes were not completed during the database audit or block

modeling for this 2012 Mineral Resource update. Preliminary analyses for the 2011 sampling are complete but pending QA/QC assessment.

Table 10-2: Upper Fir Deposit Trench and Bulk Samples

Category Deposit Operator Year

Number Series Type

#

(m)

Metallurgy Upper Fir Commerce 2008 3 BS01 to BS03 Bulk Sample 138

Exploration Upper Fir Commerce 2006 4 TR0, 0A, 1, 1A Trench-Chip 73






Figure 10-1: Drill Collar Plan

Note: Figure courtesy of Dahrouge Geological Ltd., 2012.






10.1 Core Drilling Strategy

The holes are collared on three primary drill roads (Upper, Middle, Lower roads and

several intermediate trails) that are oriented sub-parallel to the Upper Fir carbonatite

along the hillside. Drill pad set-ups are spaced about 50 m apart along drill roads.

The majority of the known portions of the Upper Fir deposit are defined on 50 m

centres. The at-depth Bone Creek carbonatite has only been intersected by a limited

number of drill holes.

The drill hole orientations appear to be approximately sub-perpendicular to the

carbonatites. The relationship between sample length and true thickness varies with

the dip of holes. True thicknesses are slightly less than drilled intercepts.

10.1.1 Core Sizes

Core holes are typically HQ diameter (96 mm) producing core with a diameter of

63.5 mm. Hole-diameter reductions due to poor ground conditions generally are not

an issue at the Project. Approximately 45% of the holes are vertical. The remaining

inclined holes typically have azimuths of 090° or 270° and dips that range from -60° to

-80°. Drill hole depths range from a minimum of 32 m and a maximum of 388 m,

averaging about 200 m.

10.1.2 Collar Surveys

The drill hole collars are spotted in the field with a hand-held global positioning system

(GPS) instrument and oriented with a Brunton compass. In 2011, an azimuth

positioning survey system (APS) was introduced to aid in lining up the drill and to

obtain a more accurate preliminary collar location and orientation.

During 2008 and 2009, collars were surveyed using a laser theodolite system by Steve

Mosdell of Align Surveys of Louis Creek, B.C.

In 2010, McElhanney Associates of Vancouver (McElhanney) undertook a differential

GPS survey of all 2010 drill collars, as well as all historic collars still visible, including

all Steve Mosdell’s work that could be verified, and all roads. McElhanney also

established six reference markers on the property to support future surveys.

Commerce advised AMEC that some earlier drill collar monuments have been

destroyed due to subsequent construction activities.

McElhanney conducted a follow-up differential GPS survey during September 2011.

The survey included 2011 campaign drill collars, the three bulk sample pits, new road

and trail construction, plus many culvert locations.






10.1.3 Down-Hole Surveys

In 2010, a Flexit Multishot down hole orientation tool was introduced for down-hole

surveys. A Flexit Single Shot tool was used for ten holes when the multishot tool was

not available or not working. The dip and azimuth of vertical and inclined drill holes

were typically tested at five (vertical) to ten (inclined) points in each hole using the

Flexit Single Shot tool. The Multishot tool recorded a reading every 3 m down-hole,

yielding a much more reliable dataset. The instrumentation records magnetic

inclination, azimuth, temperature and magnetic susceptibility at each survey depth.

The Flexit instruments were calibrated at the start of each field season from 2007 to

2010. In 2011, a Reflex gyroscopic down-hole orientation tool was introduced, with a

backup survey using a Flexit Multishot tool. Acid test was the main down-hole survey

method for the 2005-2006 drill campaigns.

10.1.4 Oriented Drill Core

Six geotechnical drill holes comprising 1,271 m of HQ diameter oriented core were

completed during 2010. In addition, an orientation survey was conducted by DGI

Geoscience Inc., Mississauga, Ontario, using optical and acoustic televiewers capable

of providing oriented drillhole information on two 2010 holes and four pre-2010 holes.

The optical and acoustic televiewer surveys were expanded to 18 drillholes during the

2011 campaign.

10.1.5 Core Handling

Core was obtained using wire-line methods and placed in wooden core trays. Core

trays were placed near the core barrel so that the core was placed in the tray in the

same orientation as it came out of the barrel. Rubble, which was rarely encountered,

was piled to about the length of the whole core that its volume would represent. Trays

were marked with drill hole name and box number. The end of every run is marked by

a wooden block depth marker.

The core trays are transported by pick-up truck down to the core logging facility at the

community of Blue River, B.C.

10.1.6 Core Recovery

Core recovery was determined prior to sampling. Typically, recovery measurements

were completed before detailed logging was initiated. Standard core recovery forms

were usually completed for each hole by the field assistant or geologist.






Core recovery is very good within the waste and carbonatite rocks (typically >95%).

The only area that may have core recovery issues would be within the fenite rocks

located in the immediate hanging wall to the carbonatite.

10.2 Drill Intercepts

Table 10-3 contains examples of the types of drill intercepts that have been returned

for the Blue River deposit areas. Typical drill hole orientations are indicated on the

cross-sections included in Section 7 of this Report. Due to the dip of the carbonatite,

drilled thicknesses reported in the table are slightly longer than true thicknesses.


In the opinion of the QPs, the quantity and quality of the lithological, geotechnical,

collar location and down-hole survey data collected in the exploration and infill drill

programs completed by Commerce are sufficient to support mineral resource

estimation.

Core logging meets industry standards for tantalum and niobium exploration within

a carbonatite setting

Collar surveys have been performed using industry-standard instrumentation

Down-hole surveys were performed using industry-standard instrumentation

Recovery data from core drill programs are acceptable

Drill hole orientations are generally appropriate for the mineralization style, and

have been drilled at orientations that are optimal to the orientation of mineralization

for the bulk of the deposit area

Drill hole spacing is sufficient to show geometry of the deposit and an

understanding of the variability of the grade and thicknesses

No material factors were identified with the data collection from the drill programs

that could affect mineral resource estimation.






Table 10-3: Example Drill Hole Intercept Summary Table

Drill Hole Easting Northing Elevation Azimuth Dip From (m) To (m)

Length**

(m)

Ta

(ppm)

Nb

(ppm)

F0720 352816.7 5796722.0 1207.2 272.8 -60 38.4 86.4 48.0 120 2,032

F0720

146.7 158.5 11.8 129 1,456

F0720

244.1 248.8 4.7 150 1,240

F0728 352811.4 5796442.0 1214.3 91.8 -60 27.3 30.0 2.7 13 745

F0728

93.0 150.4 57.4 188 1,417

F08-064 352782.3 5796404.0 1211.6 0 -90 95.7 130.4 34.7 172 1,754

F08-084 353094.7 5796738.0 1317.0 0 -90 110.0 162.0 52.0 180 1,494

F08-084

185.0 202.3 17.3 143 990

F08-084

214.3 223.4 9.1 144 1,504

F08-084

230.2 233.7 3.5 155 929

F08-084

249.6 253.8 4.2 157 878

F08-084

262.3 268.1 5.8 180 1,003

F08-113 352904.7 5796762.0 1240.3 271.8 -60 99.5 130.0 30.5 196 622

F08-113

130.0 132.1 2.1 111 258

F08-113

132.1 138.0 5.9 190 1,258

F08-113

175.0 202.8 27.8 169 616

F08-150 353006.3 5796416.5 1301.8 0 -90 129.0 136.8 7.8 143 2,258

F08-150

150.8 165.8 15.0 109 2,264

F08-150

190.0 202.1 12.1 169 969

F09-169 353033.3 5796716.0 1291.4 0 -90 78.4 126.9 48.5 185 1,658

F09-169

156.1 165.0 8.9 129 396

F09-169

165.0 167.0 2.0 224 395

F09-169

167.0 198.6 31.6 155 492

F10-200 352879.7 5796436.0 1238.9 1.8 -90 65.6 69.5 3.9 157 1,361

F10-212 352949.7 5796427.7 1273.2 91.8 -60 130.6 132.1 1.5 323 2,095

F10-212

132.1 134.9 2.8 109 1,110

F10-212

134.9 144.1 9.2 144 2,482

F10-212

144.1 145.8 1.7 58 459

F10-212

145.8 159.0 13.2 154 2,875

F10-212

183.4 189.8 6.4 161 809

Note: ** “Length” column represents drill intercept, or drilled thicknesses.






11.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY

11.1 Sampling Methods

Samples were collected from an area approximately 1,600 m north–south by 1,000 m

east-west. Sampling is from a combination of vertical and inclined holes drilled from

common collar locations. This results in a drill hole or sample spacing which increases

with depth. Average spacing between drill hole intercepts in the mineral resource area

varies from 40 to 50 m.

Core sampling method and approach has been consistent through the 2005 to 2011

drill programs. Core was boxed on site and delivered each day to a core facility in

Blue River where the core was logged and sawn. Core logging involved both

geotechnical and geological information. Geotechnical logging included measuring

core recovery per core run, rock quality designation (RQD), fracture roughness and

orientation. Core recovery and RQD were generally good for most drill core, with

typically greater than 95% recovery. The geological logging included observations of

colour, lithology, texture, structure, mineralization, and alteration. All drill core was

digitally photographed under natural outdoor, or fluorescent indoor lighting prior to

splitting, of reasonable quality. In 2011, additional ultraviolet light digital photography

of core was introduced to better characterize structure and mineral variation within the

carbonatite. All digital photos are stored in Commerce’s computerized archiving

system.

The sampling procedure used to collect core at Blue River is as follows:

The entire carbonatite intersection and shoulder samples on each side of the

intersection are sampled

Samples intervals, generally 1 m in length, are marked on the core by a geologist

Sample intervals are assigned a unique sample number

The geological contacts are generally respected

Specific gravity measurements of the carbonatite are performed at approximately

3 m spacing

Carbonatite samples are checked over the entire sample interval with a GR-130

miniSPEC gamma ray spectrometer for the presence of U and Th

Core is sawn in half by diamond saw

Half of the core is sent for analysis

Half of the core is stored in labelled core boxes for reference or further sampling.






11.2 Metallurgical Sampling

The bulk sampling program conducted in 2008 to provide metallurgical samples is

discussed in Section 9.3. Additional information on the metallurgical sampling is

included in Section 13.

Metallurgical samples were collected from bulk sample material originating from BS-2

(approximately 173 t) during January 2009 and from BS-1 (approximately 6 t) during

November 2009. BS-2 samples were ferrocolumbite dominant and selected to best

represent the average tantalum-niobium grades for the carbonatite. BS-1 samples

were selected to best reflect pyrochlore dominant mineralization.

The two bulk samples were crushed to a particle size of <1 inch diameter. After

crushing, each group of samples was homogenized separately by a standard coning

and quartering procedure. The blended samples were bagged into one tonne bags

and put into storage. One tonne of each sample was delivered to Met Solve

Laboratory (Metsolve) in Burnaby in B.C. to air dry and further reduce the size to -

10 mesh for bench testing.

11.3 Density Determinations

Commerce collected specific gravity (SG) measurements in 2010 and 2011 covering

the spatial and temporal aspect of all drill campaigns and considering the various

lithologies present. The methodology implemented was a water immersion method

and determines the specific gravity by the following formula:

SG = (weight in air) / (weight in air – weight in water)

A 10 to 20 cm piece of whole, dry, HQ core is weighed dry on an Ohaus triple beam

balance and the weight recorded. The weight in water is determined by attaching the

core by a long nylon fishing line to the Ohaus balance, lowering the core piece into a

large plastic tub located immediately below the scale and filled with purified water.

The weight of the core while immersed is then recorded, and applied to the formula for

determining the SG. Porous core samples of fenite are coated with a thin veneer of

lacquer or sealant to seal any voids or fractures present in the sample.

Calibration weights are occasionally used to verify the accuracy of the balance beam.

The table used to complete the measurements is made of wood construction and

tested for level by the technician.

The SG results are summarized in Table 11-1 and Table 11-2.






Table 11-1: 2005 – 2010 Specific Gravity Determinations by Campaign

Year Count (%) Comment

2005 10 1%

2006 45 2%

2007 96 5%

2008 216 11%

2009 51 3%

2010 1,495 78%

Total 1,913 100%

Other 12

Includes samples < 0.1 m length, outliers, and

rock types of mylonite or altered gneiss

Table 11-2: 2005 – 2010 Specific Gravity Constants

Rock Type Count Min Max Mean CV Comments

Mg-Carbonatite 845 2.71 3.49 2.97 0.01

Ca-Carbonatite 72 2.85 3.28 3.01 0.03

Silico-

Carbonatite 4 2.98 3.06 3.03 0.01 Insufficient number of samples for constants

Carbonatite-

Total 921 2.71 3.49 2.97 0.01

Amphibolite 235 2.64 3.29 3.02 0.03

Calc-silicate 42 2.89 3.24 3.08 0.03

Fenite 101 2.82 3.06 2.96 0.01

Gneiss 467 2.65 3.21 2.82 0.03

Pegmatite 58 2.56 2.70 2.62 0.01

Quartzite 10 2.6 2.71 2.64 0.01 Insufficient number of samples for constants

Schist 20 2.70 3.09 2.89 0.04

Skarn 48 2.81 3.22 3.03 0.03

Ultramafic 8 2.98 3.21 3.10 0.03 Insufficient number of samples for constants

Vein 3 2.77 2.87 2.81 0.02 Insufficient number of samples for constants

Total 1,913 2.56 3.49 2.93 0.04

Others 12

Includes samples < 0.1 m length, outliers, and

rock types of mylonite or altered gneiss

Note: By volume, the rock types with insufficient number of samples are not material to the Mineral

Resource update.






11.3.1 Density Check Program

Specific gravity check measurements were completed during the 2010 season by

Commerce. One hundred and nine samples of the 2005 – 2009 drill campaigns were

sent to Met Solve Laboratories of Burnaby, B.C. for SG determinations by wax coat

preparation followed by water immersion method. The check values compared well to

the water immersion field measurements recorded in the exploration database with a

correlation of determination (R2) of 0.98.

The check samples completed do not cover the 2010 drill campaign samples which

comprise approximately 78% of the SG data used to support the Mineral Resource

update. In the opinion of AMEC, the risk related to no checks for the 2010 SG data is

considered not materially significant for this estimate based upon the strong correlation

observed for the checks completed to date. For the upcoming 2012 field season,

AMEC recommends Commerce-Dahrouge continue using the water immersion

method and complete a 5% check of the 2010 and 2011 density determinations using

wax immersion preparation followed by the water immersion method.

11.4 Analytical Laboratories

Acme Analytical Laboratories (Acme) in Vancouver was the primary laboratory for

sample preparation of the 2005 to 2008 drill core samples. PRA/Inspectorate

Laboratories (Inspectorate) in Richmond, B.C., was the primary laboratory for sample

preparation of the 2009 to 2011 drill core samples. Acme was the primary laboratory

for sample analysis since 2005 up to and including 2011 drill core samples.

Acme is an independent mineral testing laboratory which, in 1996, became the first

commercial geochemical analysis and assaying laboratory in North America registered

under ISO 9001. The laboratory has maintained its registration in good standing since

then. ISO 9001 addresses data and organizational management to ensure

appropriate output of all product and client service. Acme regularly participates in

proficiency testing and in October 2011 the Vancouver laboratory received formal

approval of its ISO/IEC 17025:2005 accreditation from Standards Council of Canada

for Au by fire assay. Acme is working on adding additional accredited methods but

acknowledges there are no internationally recognized routine programs for Nb and Ta

proficiency testing.

Inspectorate is also an independent mineral testing laboratory that reportedly works to

internationally-recognized standards such as ISO and ASTM. The Vancouver

laboratory received ISO9001:2000 accreditation in 2006 and 2009 and participates in

proficiency testing programs.






11.5 Sample Preparation and Analysis

Between 2005 and 2008 sawn core samples were shipped to Acme where the entire

sample was crushed in a jaw crusher to 70% passing -10 mesh (2 mm) from which a

250 g riffle split sample was pulverized in a ring-and-puck mill to 85% passing

200 mesh (75 µm).

Split core samples from the 2009 to 2011 drill programs were shipped to Inspectorate

where the entire sample was crushed to 80% passing 10 mesh and a 300 g split of the

crushed material was pulverized to 100% passing 200 mesh. In 2011, pulverization

was to 95% passing 200 mesh.

Between 2005 and 2008 drill core samples were analyzed at Acme for rare earth

metals and refractory elements including Ta and Nb by inductively coupled plasma

mass spectrometry (ICP-MS) following a lithium metaborate/tetraborate fusion of a

0.2 g sample followed by dilute nitric acid digestion of the fused pellet. Major oxides

and several minor elements were analyzed by ICP-emission spectrometry (ICP-ES)

using the same procedure preceding the instrumental analysis. In addition, 36

elements were analyzed by ICP-MS of a 0.5 g sample digested in aqua regia.

Between 2009 and 2011 drill core samples were analyzed at Acme by X-ray

fluorescence analysis following a lithium metaborate fusion of a 2 g sample (XRF(F)).

11.6 Quality Assurance and Quality Control

Assessing the accuracy of Ta and Nb results presents challenges not encountered

with other commonly analysed metals. Base and precious metal assays have a wide

selection of certified reference materials (CRMs) with their associated round robins

and proficiency assessment programs. Without such feedback mechanisms, assay

laboratories can be expected to show poorer agreement for the less commonly

analysed metals. Some host minerals of these elements may be resistant to strong

acid dissolution and/or form unstable solutions in dilute acid take-up after dissolution,

which may impact ICP-MS and ICP-AES results. XRF(F) determinations may be

impacted by background corrections that impact instrument calibration, particularly at

low concentrations or short counting times.

Quality control procedures used by Commerce to monitor laboratory results have

evolved over the life of the project. Between 2005 and 2007 Commerce inserted very

few blank, duplicate, or standard reference material (SRM) control samples. During

this period, analysis of several pulp check samples was completed at six different

laboratories. In 2008 the control sample insertion frequency was increased to an

average of 3% for each of blanks, quarter core field duplicates, and SRM control






samples. In 2009 control sample insertion rates were increased to an average of 5%

per control sample type and pulp duplicates were added. Similar control sample

insertion rates were used for analysis of 2010 and 2011 drill core samples. Control

sample insertions are summarized in Table 11-3.

11.6.1 Assessment of Precision

Duplicates are used to assess laboratory precision. Duplicates in this case are

samples of the same material assayed at the same laboratory, using the same

procedure, and ideally analyzed in the same batch. Duplicate paired results are

assessed using Cumulative Frequency Absolute Relative Difference (ARD) charts.

For resource estimation purposes, a generally acceptable precision is achieved if 90%

of the duplicate pairs have an ARD less than 10% for pulp duplicates, less than 20%

for coarse reject duplicates, and less than 30% field duplicates.

Two hundred and twenty-nine quarter-core field duplicates were submitted between

2005 and 2008 as part of the regular QC program to support ICP-MS results. These

show a marginally acceptable precision for Ta ICP-MS (35% ARD at 90% cumulative

frequency, Figure 11-1). Nb paired results also have marginally acceptable precision.






Table 11-3 Control Sample Insertion Rate Summary

Year Primary Analysis

No. of Primary Samples

1/4 Core Field

Duplicates Insertion

Rate

Coarse Reject


Rate Pulp


Rate

2010 XRF(F) 4,518 105 2.3% 272 6.0% 258 5.7%

2009 XRF(F) 794 48 6.0% 0 0.0% 48 6.0%

2008 ICP-MS 5,882 191 3.2% 0 0.0% 268 4.6%

2007 ICP-MS 1,017 36 3.5% 0 0.0% 0 0.0%

2006 ICP-MS 1,140 2 0.2% 0 0.0% 0 0.0%

2005 ICP-MS 58 0 0.0% 0 0.0% 0 0.0%



Blue River SRMs

Insertion Rate

"BR-01" SRMs

Insertion Rate Blanks

Insertion Rate

2010 XRF(F) 4,518 291 6.5% 0 0.0% 248 5.5%

2009 XRF(F) 794 49 6.2% 0 0.0% 34 4.3%

2008 ICP-MS 5,882 0 0.0% 206 3.5% 222 3.8%

2007 ICP-MS 1,017 0 0.0% 49 4.8% 63 6.2%

2006 ICP-MS 1,140 0 0.0% 0 0.0% 48 4.2%

2005 ICP-MS 58 0 0.0% 3 5.2% 0 0.0%



1/4 Core Field

Checks Insertion

Rate

Coarse Reject

Checks Insertion

Rate Pulp

Checks Insertion

Rate

2010 XRF(F) 4,518 105 2.3% 272 6.0% 258 5.7%

2009 XRF(F) 794 0 0.0% 0 0.0% 49 6.2%

2008 ICP-MS 5,882 0 0.0% 0 0.0% 232 3.9%

2007 ICP-MS 1,017 0 0.0% 0 0.0% 373 36.7%

2006 ICP-MS 1,140 0 0.0% 0 0.0% 102 8.9%

2005 ICP-MS 58 0 0.0% 0 0.0% 58 100%

Table abbreviations: XRF(F) = X-ray fluorescence analysis following a lithium metaborate fusion; ICP-

MS=inductively coupled plasma mass spectrometry; Blue River SRMs = In 2008, fifteen SRMs were prepared for

Commerce by Process Research Associates (PRA, later Inspectorate) using core samples from the Blue River

carbonatite; "BR-01" SRMs = In 2005 SRM control sample BR-01 was prepared for Commerce by Acme, using

sample material from the Verity Carbonatite.






Figure 11-1: 2005 to 2008 Quarter-Core Duplicate Pair Cumulative Frequency ARD Chart

Forty-eight quarter-core and 48 pulp duplicates were submitted to support the 2009

drill program XRF(F) results. One hundred and five quarter core, 272 coarse reject,

and 258 pulp duplicate pairs were submitted to support the 2010 drill program results.

Cumulative frequency ARD chart results for duplicate pairs submitted as part of the

2005 to 2010 quality control program are summarized in Table 11-4 and Table 11-5.

Table 11-4: Cumulative Frequency ARD Summary for Tantalum

90% Cumulative Frequency ARD

Year Pulp

Coarse

Reject

Quarter

Core

2005 – 2008 ICP -

37%

2009 XRF(F) >100%

70%

2010 XRF(F) 15% 40% 55%

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

|Pair D

iffe

rence| /

(P

air M

ean)

Cumulative Frequency (Percentile Rank)

ARD Chart: Quarter Core Duplicate Pairs 2005 to 2008 Ta ICP-MS

ACME Ta ICP-MS ACME Ta ICP-MS>50






Table 11-5: Cumulative Frequency ARD Summary for Niobium


Year Pulp

Coarse

Reject

Quarter

Core

2005 – 2008 ICP -

40%

2009 XRF(F) 20%

40%

2010 XRF(F) 10% 18% 45%

The 2009 pulp and quarter core duplicate precision is poor for Ta and Nb XRF(F). The

2010 pulp and coarse reject duplicates achieve acceptable precision for Nb XRF(F)

but marginal to poor precision for Ta XRF(F). As sample grades approach the lower

detection limits of the given analytical procedure, the ARD typically increases. The

2010 Ta pulp and coarse reject duplicate precision improves to an acceptable level for

paired duplicates with a mean grade greater than 50 ppm (Table 11-6, Figure 11-2).

Table 11-6: Cumulative Frequency ARD Summary for Tantalum (Mean > than 50 ppm Ta)


Year Pulp

Coarse

Reject

Quarter

Core

2005 – 2008 ICP -

37%

2009 XRF(F) 89%

68%

2010 XRF(F) 11% 22% 50%






Figure 11-2: 2010 Drill Core Assay Program Cumulative Frequency ARD Chart

Note: Black line represents cumulative frequency with no exclusion of paired data. Green line represents

cumulative frequency after exclusion of pairs with mean grade < 50 ppm. Acme 8x refers to Acme’s

Group 8 single element analysis by XRF(F).

In 2008, 300 pulps were resubmitted after initial assaying as a separate batch to Acme

for ICP-MS analysis. This work was part of a check program investigating accuracy

and precision differences between ICP-MS and XRF(F). The ARD at 90% cumulative

frequency was 28% for Ta and 22% for Nb. An acceptable level of precision was not

achieved even at a 50 ppm Ta cut-off. However, no significant bias between the

duplicate pairs for either Ta or Nb was apparent.

An XRF(F) re-assay program of 2008 coarse rejects was initiated in 2009. Pulp

duplicate results for 118 pairs show acceptable precision was achieved for Nb but not

for Ta (Figure 11-3 and Figure 11-4).

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

|Pair D

iffe

rence| /

(P

air M

ean)


ARD Chart: Commerce Pulp Duplicate Pairs

ACME Acme Ta 8x (XRF-F) ACME Acme Ta 8x (XRF-F)>50






Figure 11-3: 2009 Re-assay Program Ta XRF(F) Cumulative Frequency ARD Chart

Note: Black line represents cumulative frequency with no exclusion of paired data. Green line represents

cumulative frequency after exclusion of pairs with mean grade < 50 ppm.

Figure 11-4: 2009 Re-assay Program Nb XRF(F) Cumulative Frequency ARD Chart

Note: Green line represents cumulative frequency with no exclusion of paired data.

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

|Pair D

iffe

rence| /

(P

air M

ean)


ARD Chart: Pulp Duplicate Pairs for 2009 Re-assay Ta XRF(F)

ACME Ta XRF(F) ACME Ta XRF(F)>50

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

0% 10% 20% 30% 40% 50% 60% 70% 80% 90% 100%

|Pair D

iffe

rence| /

(P

air M

ean)


ARD Chart: Pulp Duplicate Pairs for 2009 Re-assay Nb XRF(F)

ACME Nb XRF(F) ACME Nb XRF(F)>0






11.6.2 Assessment of Accuracy

Standard Reference Material (SRM) control samples are used to assess laboratory

accuracy. Accuracy is commonly measured as a bias between the Best Value (BV) of

the SRM and the SRM results from the drill sample batches. Biases less than ±5% are

commonplace and within the industry are widely accepted for the purposes of resource

estimation. A bias between five and ten percent is considered marginal and usually

warrants investigation. A bias exceeding 10% is considered significant and usually

warrants remedial action before samples are used to support resource estimations.

In 2005 SRM control sample BR-01 was prepared for Commerce by Acme, using

sample material from the Verity Carbonatite. The BR-01 SRM did not undergo round

robin testing: the best value (BV) is not certified and is used provisionally.

In 2008, fifteen SRMs (Blue River SRMs) were prepared for Commerce by Process

Research Associates (PRA, later Inspectorate) using core samples from the Blue River

carbonatite. Three samples of each standard were sent to six laboratories, for Ta

analysis, and seven laboratories for Nb analysis. Analytical procedures included

XRF(F) and XRF/pressed pellet, ICP-AES, ICP-MS, ICP-M, and atomic absorption.

The Blue River SRM round robin program used a small number of laboratories,

providing less reliable estimates of the BVs. The calculated 95% confidences of the Ta

BVs are unacceptably wide (greater than 5% and commonly greater than 10%). The

95% confidences for most of the Nb BVs are acceptably narrow (less than 5%), but the

small number of laboratories in the round robin increases the risk that the confidence

estimate is fortuitously small. The BVs could change significantly if more laboratories

were used in the round robin. The calculated BVs of these 15 SRMs are used on a

provisional basis and conclusions on accuracy are supported with results from check

samples.

Two hundred and six BR-01 SRMs were inserted by Acme into sample batches to

support the 2005 to 2008 ICP-MS drill sample results. Control sample results were

generally within ±5% of the best value. Less than 5% of the results exceeded

±2 standard deviations of the mean grade of all BR-01 results (Figure 11-5 and Figure

11-6).

Forty-seven Blue River SRMs were inserted to support the 2009 XRF(F) drill sample

results. There is an apparent negative bias for Ta results greater than 150 ppm and

no apparent bias for Nb results (Figure 11-7 and Figure 11-8). These charts show the

spread of Blue River SRM batch results around the expected value. Ta results are

quite variable; Nb results are quite tight.






Figure 11-5: 2005 to 2008 Ta ICP-MS BR-01 SRM Control Chart

Figure 11-6: 2005 to 2008 Nb ICP-MS BR-01 SRM Control Chart

80859095

100105110115120125130135140145150

2005

2007

2007

2007

2007

2007

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

Sta

nd

ard

BR

-01

, T

a 4

B p

pm

In order Assayed, by Date

SRM BR-01 Ta 2005 to 2008 ICP-MS

Data

Mean +/- 2

std.dev.s

Best Value

Moving Average

1.05 x Best

0.95 x Best

High Clusters

Low Clusters

600

650

700

750

800

850

900

950

1000

2005

2007

2007

2007

2007

2007

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

2008

Sta

nd

ard

BR

-01

, N

b 4

B p

pm

In order Assayed, by Date

SRM BR-01 Nb 2005 to 2008 ICP-MS

Data

Mean +/- 2

std.dev.s

Best Value

Moving Average

1.05 x Best

0.95 x Best

High Clusters

Low Clusters






Figure 11-7: 2009 Ta XRF(F) Blue River SRMs Control Chart\

y = 0.9214x + 4.3841R² = 0.9664

0

50

100

150

200

250

300

350

400

0 50 100 150 200 250 300 350 400

Ta p

pm

XR

F(F)

fro

m b

atch

ru

ns

Ta ppm XRF(F) Best value

2009 Drill Program Blue River SRM Ta XRF(F) Results

2009 Batch SRM Results Result <± 5% of expected value y=x Linear (2009 Batch SRM Results)






Figure 11-8: 2009 Nb XRF(F) Blue River SRMs Control Chart

Two hundred and ninety-one Blue River SRMs were inserted to support the 2010

XRF(F) drill sample results. There was no significant bias observed for Nb (Table

11-7). Ta results show a predominantly negative bias up to -9.7% relative to the

provisional BVs for most grade ranges (Table 11-8).

y = 1.0198x - 29.878R² = 0.9982

0

500

1000

1500

2000

2500

3000

3500

4000

4500

0 500 1000 1500 2000 2500 3000 3500 4000 4500

Nb

pp

m X

RF(

F) f

rom

bat

ch

run

s

Nb ppm XRF(F) Best value

2009 Drill Program Blue River SRM Nb XRF(F) Results

2009 Batch SRM Results Result < 5% of expected value y=x Linear (2009 Batch SRM Results)






Table 11-7: 2010 Nb XRF(F) Blue River SRM Control Chart Summary

Lab

Standard

Sample

Best

Value 95% CI 95%CI/BV

Mean

Value Bias

Number of

Samples

Acme Uf-STD-03 3048 ±83 3% 3112 2.1% 33

Acme Uf-STD-04 3907 ±117 3% 4012 2.7% 34

Acme Uf-STD-06 2297 ±86 4% 2390 4.0% 6

Acme Uf-STD-07 1753 ±72 4% 1818 3.7% 37

Acme Uf-STD-08 244 ±29 12% 238 -2.4% 31

Acme Uf-STD-09 313 ±29 9% 316 0.9% 29

Acme Uf-STD-10 956 ±92 10% 996 4.2% 26

Acme Uf-STD-12 1478 ±85 6% 1540 4.2% 28

Acme Uf-STD-13 1744 ±74 4% 1810 3.8% 25

Acme Uf-STD-14 282 ±35 12% 274 -2.7% 31

Acme Uf-STD-15 2524 ±95 4% 2601 3.1% 11

Table 11-8: 2010 Ta XRF(F) Blue River SRM Control Chart Summary

Lab

Standard

Sample

Best

Value

95%

CI

95%CI/

BV Mean Value Bias

Number

of

Samples

Acme Uf-STD-03 74 ±18 24% 73 -1.8% 27

Acme Uf-STD-04 122 ±17 14% 136 11.5% 28

Acme Uf-STD-06 172 ±20 12% 169 -1.7% 6

Acme Uf-STD-07 179 ±17 10% 178 -0.8% 31

Acme Uf-STD-08 175 ±11 6% 158 -9.7% 28

Acme Uf-STD-09 193 ±15 8% 174 -9.7% 24

Acme Uf-STD-10 221 ±14 7% 210 -5.1% 23

Acme Uf-STD-12 241 ±17 7% 233 -3.2% 24

Acme Uf-STD-13 280 ±28 10% 281 0.4% 23

Acme Uf-STD-14 256 ±20 8% 235 -8.0% 25

Acme Uf-STD-15 377 ±31 8% 377 -0.1% 8

11.6.3 Assessment of Laboratory Bias

Check samples, typically pulp samples, are sent to secondary laboratories to assess

for between-laboratory bias. Between laboratory bias is assessed here using

Reduction to Major Axis (RMA) charts which take into consideration that laboratory

results are independent and that one set of the paired results may be more “erratic”

than the other. Between-laboratory biases less than ±5% are generally considered

acceptable.






Pulp check sample analyses were not systematically performed between 2005 and

2008, and SRMs were not consistently included with the check sample batches.

Therefore the results for this period are considered an indicator of laboratory bias, but

in isolation are not considered to be definitive. Since 2009, most check sample

programs are supported with adequate control samples.

A summary of RMA calculated biases between primary and secondary lab results by

year, shown in Table 11-9, indicates a general reduction in between-laboratory bias for

Nb between 2005 and 2009 drill programs. Ta bias has an acceptable level of ±5%

between 2006 and 2010 drill programs.

Table 11-9: Pulp Check Between-Laboratory Bias

Year Ta Bias (%) Nb Bias (%)

Check

Laboratory

2005 -43.6 11.65 GDL

2006 -1.9 - Becquerel

2007 -6.6 36.8 ActLab

2007 -0.6 22 ALS

2008 -3.4 -9.4 GDL

2008 4.0 5.0 Acme

2009 -0.21 2.61 Stark

2010 -5% -1% Stark

Figure 11-9 shows the RMA chart used to assess between-laboratory bias for 2010 Nb

XRF(F) check pair results.






Figure 11-9: 2010 Acme versus Stark Nb XRF(F) Check Pair RMA Chart

All Excluding Excluding

Data Outliers (1) Outliers (2)

N 254 253 252

Percent Rejected 0.0% 0.4% 0.8%

R squared 0.99 0.99 0.99

R 0.99 1.00 1.00

slope m 1.02 1.01 1.01

intercept b -2.39 1.05 2.79

error in slope 1% 1% 1%

error in intercept 15.7 15.7 15.7

Bias -2% -1% -1%

11.6.4 Assessment of Contamination

Blanks are used to assess contamination during sample preparation and analysis.

Commerce uses coarse crushed, optical quartz material sourced from Jim Coleman

Crystal Mines located near Hotsprings, Arkansas for their blank material.

0

1000

2000

3000

4000

5000

6000

0 1000 2000 3000 4000 5000 6000

STA

RK

IC

P N

b p

pm

ACME XRF (F) Nb ppm

Commerce Pulp Check Pair Results for ACME XRF(F) vs STARK ICP-AES

Data

Outliers (1)

Fit, all data

Fit, exclude Outliers (1)

Outliers (2)

Fit, excludes Outliers(2)






Three hundred and five coarse blank control samples were inserted to support the

2005 to 2008 ICP-MS drill sample results; Thirty-two coarse blanks were inserted to

support the 2009 XRF(F) drill sample results. Two hundred and forty-eight coarse

blanks were inserted to support the 2010 XRF(F) drill sample results.

There is no indication of systematic contamination for 2005 to 2008 ICP-MS results

(Figure 11-10 and Figure 11-11). The blank performance charts show a few blank

sample results well in excess of background values. These results may indicate

periodic contamination but may also be a result of sample swaps or transcription

errors. The impact of these outliers in considered minor.

Figure 11-10: 2005 – 2008 Blank Ta ICP-MS Performance Chart

Note: This blank versus previous sample chart is prepared assuming samples were assayed in sample

number order.

0

20

40

60

80

100

120

140

160

180

0 50 100 150 200 250 300 350 400

Bla

nk

Ta p

pm

Previous Sample Ta ppm

QTZ BLK Blank Ta ICP-MS 2005 to 2008






Figure 11-11: 2005 - 2008 Blank Nb ICP-MS Performance Chart


number order.

There is no indication of systematic contamination of 2009 Ta XRF(F) results but Nb

XRF(F) results show an apparent systematic carry-over contamination of blanks for Nb

grades greater than 2,000 ppm (Figure 11-12 and Figure 11-13).

0

200

400

600

800

1000

1200

1400

1600

1800

0 500 1000 1500 2000 2500 3000 3500 4000

Bla

nk

Nb

pp

m

Previous Sample Nb ppm

QTZ BLK Blank Nb ICP-MS 2005 to 2008






Figure 11-12: 2009 Ta XRF(F) Blank Performance Chart


number order.

0

10

20

30

40

50

60

70

80

90

0 50 100 150 200 250 300 350 400 450 500

Ta_

pp

m X

RF(

F)

Previous Sample Ta_ppm XRF(F)

QTZ BLK Blank Ta_ppm XRF(F) 2009






Figure 11-13: 2009 Nb XRF(F) Blank Performance Chart


number order.

Blank performance charts for 2010 show no systematic contamination for Ta or Nb

XRF(F) results (Figure 11-14 and Figure 11-15).

0

10

20

30

40

50

60

0 500 1000 1500 2000 2500 3000 3500 4000

Nb

_p

pm

XR

F(F)

Previous Sample Nb_ppm XRF(F)

QTZ BLK Blank Nb_ppm XRF(F) 2009






Figure 11-14: 2010 Ta XRF(F) Blank Performance Chart

0

10

20

30

40

50

60

70

0

10

20

30

40

50

60

70

BR

17

70

18

BR

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90

75

BR

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24

BR

18

11

49

BR

18

01

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BR

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00

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49

BR

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97

BR

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99

BR

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BR

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71

BR

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45

BR

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48

BR

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BR

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53

BR

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95

BR

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90

04

BR

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10

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BR

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80

05

BR

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61

25

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60

05

BR

21

61

45

BR

20

40

05

BR

22

31

24

BR

22

90

06

BR

22

90

98

BR

22

01

15

BR

22

20

50

BR

22

50

50

BR

22

70

47

XR

F(F)

Ta

pp

m

Sample Number

Blanks Ta XRF(F) for 2010

Blanks

LDL

10* LDL






Figure 11-15: 2010 Nb XRF(F) Blank Performance Chart

11.6.5 Assay QA/QC Conclusions

Quality control sample insertion rates are sufficient to allow assessment of precision,

laboratory bias, and contamination. Accuracy assessment remains provisional due to

the difficulty of obtaining standards with suitable quality.

ICP-MS results for the 2005 to 2008 drill program have marginally acceptable

precision. XRF(F) precision is poor for 2009 results. XRF(F) precision is acceptable

for 2010 XRF(F) results for Ta and Nb grades above 50 ppm. Except for 2009 Ta

(XRF(F) results, no systematic accuracy or contamination problems are indicated for

these programs.

A re-assay program of 2008 coarse rejects resulted in poor precision for Ta XRF(F)

and acceptable precision for Nb XRF(F) results.

ICP-MS results supporting the 2005 to 2008 drill program and initial XRF(F) results

supporting the 2009 and 2010 drill programs are considered suitable for use in mineral

0

10

20

30

40

50

60

70

0

10

20

30

40

50

60

70

BR

17

70

18

BR

17

80

50

BR

17

90

75

BR

18

00

99

BR

18

10

99

BR

18

20

97

BR

18

30

45

BR

18

50

16

BR

18

60

92

BR

18

71

16

BR

19

00

26

BR

19

10

65

BR

19

20

50

BR

19

30

84

BR

19

50

02

BR

19

70

43

BR

19

81

09

BR

20

01

04

BR

20

20

29

BR

20

40

50

BR

20

50

69

BR

20

80

90

BR

21

00

16

BR

21

30

05

BR

21

40

05

BR

21

60

05

BR

21

61

95

BR

21

90

04

BR

22

00

33

BR

22

10

54

BR

22

21

43

BR

22

40

06

BR

22

51

25

BR

22

61

47

BR

22

80

48

BR

23

00

49

BR

23

30

41

BR

23

60

12

BR

24

50

45

BR

24

61

45

BR

24

80

36

XR

F(F)

Nb

pp

m

Sample Number

Blanks Nb XRF(F) Nb

Blanks

LDL

10* LDL






resource estimation. Caution should be applied in assigning a high level of confidence

to the 2005 to 2009 results until precision and accuracy issues are resolved.

11.7 Databases

Collar, down-hole survey, geology, specific gravity and assays were stored in a

Gems™ database format. Prior to GEMS™, data capture occurs in a variety of

formats from hand logs and Excel files.

Assay data are received from the laboratories via comma-separated value (CSV) data

files. Collected data are subjected to validation prior to upload into the database using

built-in program triggers that automatically check the data. Verification checks include

collar co-ordinates, surveys, lithology, and assay data. After data are imported into

GEMS™, visual checks are completed to ensure that data placement was correct

within the various database fields.

Commerce has migrated to a Fusion data management system by Century Systems

Technologies Inc. during 2011 for all data captured during 2011 exploration activity.

Pre-2011 data are being migrated to the Fusion database on an incremental basis with

completion anticipated during 2012.

11.8 Security

The drill crew transports drill core to the Commerce field office in Blue River at the end

of every shift by pick-up truck. The boxes are laid out in order on saw horses and

inspected by the project manager. Dahrouge geologists and technicians log and

sample the drill. The core storage, logging and sampling facilities are not secured.

Samples are placed in pails and stored in the locked quonset hut for security prior to

shipping.

A commercial delivery service, Monashee Painting and Services of Blue River, B.C.,

transports the samples, to the preparation laboratory in Richmond, B.C. Sample sheet

manifests are submitted with the core samples. The manifests include information on

the operator, sample preparation laboratory, and a sample list. Sample rejects

returned from the laboratory are stored in the onsite quonset hut.

Archived 2010 and 2011 core is stored on pallets at the Blue River field office. The

pre-2010 core is stored on pallets in a field located at Valemount, B.C. During 2011,

Commerce initiated construction of a cold storage building and fenced compound at

the Valemount storage facility. Commerce plans to move the archived core into the

new storage facility at Valemount during 2012.







The majority of drill core samples used to support the 2012 Mineral Resource estimate

were prepared at an independent laboratory before submission to the primary

laboratory for analysis. All drill core samples were analysed by independent

laboratories using ICP and XRF methods.

Typically, drill programs included insertion of blank, duplicate and standard reference

material samples. QA/QC submission rates meet industry-accepted standards for

insertion rates.

Quality control work completed by Commerce between 2005 and 2009 identified

laboratory precision and accuracy concerns. Achieving precision for tantalum by

XRF(F) and ICP-MS methods was difficult and is likely related to the difficulty of

digesting tantalum prior to ICP-MS analysis and the relatively high detection limit for

tantalum by XRF methods. The XRF(F) detection limit initially provided was close to

the lower economic cut-off for tantalum. Submission of the 2010 drill program samples

was delayed while Commerce and their primary laboratory, Acme, worked together to

address this concern. Procedures for XRF(F) analysis were modified and quality

control sample results supporting the 2010 drill program sample analysis indicate

Acme has achieved acceptable precision and accuracy for tantalum and niobium using

XRF(F) methods.

During preparation of the 2012 Mineral Resource update reported in Section 14,

AMEC and Commerce examined the sample preparation and analysis of Blue River

core. The principal findings of this work were as follows:

SRM control samples indicate provisionally acceptable levels of accuracy are for

the most part, achieved for Nb by either XRF(F) or ICP-MS methods;

SRM control samples indicated a low bias may exist for Ta by either XRF(F) or

ICP-MS methods;

Poor laboratory precision is evident for Ta and Nb results collected in 2008 and

2009, but no consistent bias is evident;

Acceptable precision is evident for 2010 Ta and Nb results;

Significant inter-laboratory grade biases are evident for Ta and Nb in the 2005 and

2006 sampling; these samples represent a small portion of the database and

would likely not materially impact the resource estimate;

Acceptable inter-laboratory biases are achieved in the remaining 2007 to 2010

sampling programs;






AMEC concludes that the Blue River sample results show imprecision but no

consistent bias and that the ICP-MS results from 2005 to 2008, and the XRF(F) results

supporting the 2009 and 2010 drilling are suitable for use in mineral resource

estimation. Caution should be applied in assigning a high level of confidence to the

pre-2010 tantalum and niobium analytical results until precision and accuracy issues

are resolved.

AMEC concludes the SG field determinations are reasonable for the 2005 – 2009

campaigns. Additional SG checks at an independent laboratory are recommended for

the 2010 and 2011 drill campaigns using a wax immersion preparation, followed by a

water immersion determination.

Collected data are subjected to validation prior to upload into the database using built-

in program triggers that automatically check the data. Verification checks include

collar co-ordinates, surveys, lithology, and assay data. The checks are appropriate,

and consistent with industry standards. Independent data audits have been

conducted, and indicate that the sample collection and database entry procedures are

acceptable.

Sample security has relied upon the fact that the samples were always attended or

locked in appropriate storage facilities. Chain-of-custody procedures consist of filling

out sample submittal forms that are sent to the laboratory with sample shipments to

make certain that all samples are received by the laboratory;

Current sample storage procedures and indoor storage areas are consistent with

industry standards. Security for core on pallets stored outdoors in unfenced fields is

under improvement by Commerce.

The QPs are of the opinion that the quality of the specific gravity, tantalum and

niobium analytical data are sufficiently reliable to support mineral resource estimation

and that sample preparation, analysis, and security are generally performed in

accordance with exploration best practices and industry standards.






12.0 DATA VERIFICATION

Commerce implemented an industry-acceptable quality control program to manage

logging, sampling, and analysis. This section summarizes the database and site visit

verification work by AMEC.

12.1 Database Verification

AMEC completed a minimum 5% database verification check and concluded the collar

coordinates, down-hole surveys, lithologies, and assay databases are sufficiently free

of error (Chong and Postolski, 2011; Thompson, 2011). Other verification reviews

include calculation checks on the density determinations, rock quality designation, and

total core recovery using source record data. AMEC concludes that the density

determinations, total core recovery, and the rock quality designation database values

have been calculated appropriately.

Down-hole survey data quality was checked by AMEC for proper magnetic deviation

adjustments and for potential, erroneous, down-hole surveys that result in drill hole

deviations that exceed typical deviations expected for drill holes. AMEC concludes the

down-hole deviation surveys are reasonable and suitable for mineral resource

estimation (Chong and Postolski, 2011; Thompson, 2011).

12.2 Site Visits

The AMEC QPs were assisted by Dahrouge managers, geologists and technicians

during their site visits.

The AMEC geology QP, Albert Chong, P.Geo. visited the Project during 11 to 16 July

2010, and 6 to 14 September 2011 to verify the 2005–2010 drilling, sampling and

surface mapping campaigns. Outcrops, drill core and sampling method protocols were

reviewed to verify the data, the exploration protocols, and the resulting geological

interpretation.

During the September 2011 site visit, the AMEC QP was assisted by SRK structural

geologists J.F. Couture, P.Geo., and I. Nash to verify the enhanced structural geology

interpretation of the Upper Fir deposit.

The AMEC Mining QP, Ramon Mendoza Reyes, P.Eng., visited the Project during

2010 to assess the site for infrastructure planning purposes.






12.2.1 Drill Collar Location Check

The AMEC audit included 32 drill collars of 237 drill holes from the 2005–2010 drill

campaigns, or approximately 14% of the mineral resource drill holes.

AMEC checked the location of the drill collars using a hand-held Garmin GPS Map 60

CSX unit. The 2006 to 2009 drilling campaigns were verified based upon holes from

10 setups. All holes checked were within ±8 m and most were within ±3 m. The drill

hole locations with discrepancies greater than 3 m were found to be related to

disturbance of markers from drill road construction.

Upon completion of a drill hole, 4" x 4" wooden posts are placed into the hole. The drill

hole collar casings are in some cases still in place. Steel plates with the drill hole

names have been cemented into the ground adjacent to the hole collars making hole

identification relatively easy.

At the recommendation of AMEC, Commerce-Dahrouge commissioned McElhanney

Land Surveying of Vancouver (McElhanney) to re-survey all the available 2005–2009

drill collars, and complete primary surveys for the 2010 drilling. During November

2011, McElhanney completed a second primary survey of the 2011 drilling and the

bulk sample pits (BS1 and BS2). The drill hole collar database has S. Mosdell data for

the 2005–2008 campaigns, both S. Mosdell and McElhanney data for the 2009

campaign, and McElhanney data for the 2010 campaign.

12.2.2 Inspection of Drill Core and Verification of Mineralization

During the site visits, the AMEC QP checked drill log entries by re-logging drill core

from 14 (6%) Upper Fir deposit drill holes with a focus on as-logged lithology,

structural deformation observations, mineralization, sample intervals, and visual

checks on the total core recovery (Chong, 2010, 2012). Based upon the holes

reviewed, the AMEC QP concludes the database records reasonably reflects the as-

logged drill core observations by Commerce observed in the audited drill core.

AMEC collected and submitted 31 quarter-core samples to Acme in Vancouver for

preparation and analysis by Package 4B ICP-MS methods. A comparison of AMEC

results with matched interval results reported in the resource database reasonably

support the grades reported in the resource database (Table 12-1). The mean grade

of the original and check results show noteworthy agreement for a comparison of this

size and type.






Table 12-1: AMEC Site Visit Confirmation of Mineralization

Original Half Core Samples AMEC Quarter Core Samples

DHID From (m) To (m) Ta (ppm) Nb (ppm) Ta (ppm) Nb (ppm)

F07-028 135 136 421 4,405 244 2,404

F08-150 129.7 131 210 4,739 143 1,727

F08-151 194.3 195 219 2,887 187 2,084

F08-150 134 135 181 2,742 189 2,667

CF0612 105.3 106.3 365 1,494 287 1,305

F08-151 195 196 91 2,040 78* 1,786*

F08-150 158 159 86 1,559 134* 2,034*

F08-150 133 134 101 1,365 118 1,566

F07-028 134 135 128 1,261 64** 553**

F07-028 133 134 75 910 141** 1,734**

F08-151 192 193 143 744 140 798

CF0612 104.3 105.3 194 506 146 367

CF0612 106.3 107.3 122 466 146 565

CF0612 103.3 104.3 121 360 131 328

CF0612 108.2 109.2 43 128 112 392

F10-222 238 239 969 6,904 900 6,653

F10-184 249.5 250.76 382 4,462 483 5,141

F10-184 158.41 159.6 241 1,565 235 1,426

F10-184 260.46 261.57 252 1,526 327 2,006

F10-222 278 279 243 1,404 421 3,796

F10-222 248 249 213 1,455 218 1,430

F10-222 263 264 72 1,564 99 1,876

F10-184 196.6 197.84 176 1,057 218 1,338

F10-184 168.57 170 100 1,110 63 693

F10-184 228.22 229.3 188 926 145 749

F10-184 187.45 188.68 174 683 190 733

F10-220 301 301.5 173 380 110 281

F10-222 310 311 148 342 166 363

F10-220 317 318 120 226 155 228

F10-222 289.44 290.45 70 189 50 156

F10-220 308 309 26 69 37 83

Count 31 31 31 31

Mean 195 1,596 196 1,525

Relative Difference 0% -4%

Notes: The table is generally sorted by pre-2010 sampling versus 2010 sampling and then by Ta and Nb grade. Grey shaded records indicate samples with potential sample swaps.

* possible sample swap on the AMEC check sample analyses.

** possible sample swap on the AMEC check sample analyses.







Based upon the database and site visit verification audits completed by AMEC, the

opinion of the QPs is that the collar coordinates, down-hole surveys, lithologies, and

assays are considered sufficiently free of error and that the data are suitable to support







13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

Testwork began in 2009 and continued into 2010 to develop a process flowsheet for

the Blue River Project. The testwork was based on material produced from two bulk

samples, BS-2F and BS–2G. Mineralogical analysis was performed to obtain

knowledge regarding the occurrence of the tantalum and niobium within the material.

Given the complexities with assaying for the tantalum, a fair amount of effort also went

into developing the appropriate routine for the assaying of samples.

The 2009-2010 testwork primarily took place in two phases:

Phase I – focused on the recovery of the tantalum–niobium minerals by gravity

although grinding and mineralogy investigations were also performed.

Phase II – focused on the recovery and upgrading of the tantalum–niobium

minerals by flotation.

Work has continued in 2011 and 2012:

Phase III – continued optimization of the process flowsheet at the laboratory scale

for the production of a tantalum-niobium mineral concentrate.

A large amount of work was performed in Phase I that showed gravity could

concentrate the material to a low-grade product, but that upgrading increasingly gave

lower levels of metallurgical recovery as grade was sought. Mineralogical work

completed before and during this phase of work showed that the tantalum was not

present as tantalite but rather as the minerals ferrocolumbite and pyrochlore, which

limits recovery by the gravity route due to the low differential specific gravity between

pyrochlore and gangue minerals.

Work in Phase II saw the use of flotation concentration technology similar to that being

used for niobium-bearing carbonatites at Iamgold’s Niobec Mine in Quebec, Canada.

There was immediate success in the first phases of the work. Although there are

several stages to the concentration, the overall level of equipment, risk, and complexity

to produce a saleable or treatable concentrate is lower than the gravity route.

Process development work is continuing in this area, but for the purposes of the PEA,

the process suggested by Test F81 was selected as the basis of initial concentration

design because recoveries were good (approx. 70% for Ta) for this type of

mineralization and because a combined grade of 10% Ta–Nb was achieved. It is

expected that with further work, a combined grade of 30% Ta–Nb should be

achievable.






In both work phases, the emphasis of concentration techniques was to create a

material which would be easily upgraded by hydrometallurgical methods,

pyrometallurgical methods, or a combination of both. These processes would permit

the separation of Ta from Nb, allowing payment for both products. To this end, an in-

depth review was completed of those technologies for the production of high-value

intermediate products and final products. There is confidence that the concentrate

could be reduced to metal by the aluminothermic process. Subsequently there would

be chlorination of the granulated metal alloy product and distillation of the anhydrous

metal chloride products to produce high purity Nb and Ta chlorides. Tantalum chloride

is the precursor to capacitor grade Ta powder and can be marketed as such.

However, both Ta and Nb chlorides can also be hydrolyzed and calcined to generate

high purity oxide products for other applications.

In 2011 and 2012, work has continued into Phase III which is the optimization of work

conducted in Phase II. This optimization work has concentrated on de-sliming of

flotation feed, rejection flotation of carbonates and pyrrhotite, and subsequent flotation

of a tantalum-niobium concentrate for processing through extractive metallurgy.

Although progress has been made in the testwork, there is no material change from

the results indicated by the Phase I and II testwork. As a result there is no impact on

the performance assumptions, or the capital and operating costs in the PEA.

Confidence in the process design has been enhanced.

13.1 Head Samples for Initial Testing

In 2009, two bulk samples, BS-2F and BS–2G, sourced from a small pit in the Upper

Fir zone, and weighing approximately 200 t in total, were contract-crushed to a particle

size of <1 inch diameter. After crushing, each group of samples was homogenized

separately by a standard coning and quartering procedure. The well-mixed samples

were bagged into one tonne bags and put into storage. One tonne of each sample

was delivered to Met-Solve Laboratory (Met-Solve) in Burnaby B.C. to air dry and

further reduce the size to -10 mesh for bench testing.

Met-Solve is a commercial mineral and metallurgical testing facility that is independent

of Commerce, and specializes in mineral beneficiation and hydrometallurgical

testwork.

The mineralogical examinations of all the bulk samples taken during the 2008

exploration program are described by Chudy (2009). Additional mineralogical

examinations were performed on some of the test products during the mineral

processing investigations.






The head assays, established using X-ray fluorescence (XRF) analysis, for the two

bulk samples are tabulated in Table 13-1.

Table 13-1: Head Assay Grades, Bulk Samples BS-2F and BS-2G

Sample Ta (ppm) Nb (ppm)

BS-2F 194 1,300

BS -2G 114 764

13.2 Phase I Testing

13.2.1 Grinding Size

Each sample was subjected to gravity separation tests at five different grind sizes of

80% passing 500 µm, 230 µm, 100 µm, 74 µm and 45 µm to determine the liberation

size using a centrifugal concentrator. A standard seven-pass procedure was used to

simulate continuous gravity concentrator action.

This work indicates that the liberation size for both samples is coarser than P80 of

76 µm. The relative position of the curves (Figure 13-1 and Figure 13-2) indicates that

effective liberation for gravity is likely achievable at a grind size slightly coarser than

120 µm. The results for niobium are similar.

Assaying of the individual size fractions of the tailings from the BS-2G tests indicate

that there are still a few locked particles between 74 µm and 106 µm when ground to

P80 112 µm but that material coarser than 150 µm does not contain any tantalum.

Given the natural size distribution obtained in grinding, this implies that effective

liberation for processing, is about P80 of 125 µm for gravity treatment and slightly

coarser for flotation (P80 up to 160 µm). These numbers are in line with independent

findings from the mineralogical examination of all bulk samples of the 2008

exploration program.

13.2.2 Roughing and Cleaning Gravity Concentration

With the establishment of the grind size and initial gravity results, it was decided to

progress with the gravity concentration work. The two samples were treated with a

centrifugal concentrator, using 10 consecutive stages for rougher concentration

followed by three cleaning stages of the combined rougher concentrates. Four

different grind sizes were tested for each sample. All results were similar, with

recoveries falling off quickly in cleaning and inability to raise the grades any higher






than Ta 3,500 ppm (0.35% Ta). Results from sample BS-2G are shown on Figure

13-3.

Large batch samples of 60 kg were tested using a Falcon Centrifugal Gravity

Concentrator in 10 consecutive stages to produce a rougher and a scavenger

concentrate at a grind size of P80 100 µm. The rougher concentrate only was

screened to produce three size fractions as follows:

+74 µm

37 to 74 µm

-37 µm.

Figure 13-1: Sample BS-2F – Gravity Separation (Different Grinds)

0

10

20

30

40

50

60

70

80

90

100

0 200 400 600 800 1000 1200 1400 1600 1800 2000

Reco

very

(%

)

Grade Ta (ppm)

BS-2F - Gravity Separation Tantalum Grade/Recovery Curves

P80 500 um

P80 233 um

P80 112 um

P80 76 um

P80 45 um






Figure 13-2: Sample BS-2G – Gravity Separation (Different Grinds)

Figure 13-3: Overall Rougher and Cleaner Recovery vs Grade by Centrifugal Gravity

Concentration

Each fraction was then cleaned by gravity using a Wilfley shaking table, with a

medium-size deck. Results were similar to the gravity separation using centrifugal

separator only, with no improvement in recoveries or grades. These fractions were

also tested using a Mozley table concentrator to determine the upgrading

characteristics of the products. Results showed that while it would be possible to

increase the grades by up to six times at the laboratory level, the recoveries would

drop accordingly. The results are shown in Figure 13-4.

0

10

20

30

40

50

60

70

80

90

100

0 200 400 600 800 1000

Reco

very

(%

)

Grade Ta (ppm)

BS-2G - Gravity Separation Tantalum Grade/Recovery Curves

P80 650 um

P80 250 um

P80 120 um

P80 76 um

P80 45 um

0

10

20

30

40

50

60

70

80

90

100

0 500 1000 1500 2000 2500 3000 3500 4000

Ta R

eco

very

, %

Ta Grade, ppm

Ta Grade Recovery Curve P80 116um

P80 83 um

P46 um

P80 238 um






Tests were also performed to determine the benefits of additional steps, such as de-

sliming and de-sulphidization. These procedures were incorporated into the testwork

but essentially AMEC was of the opinion that concentration by gravity as the primary

method was not the optimum choice for Project development.

Figure 13-4: Upgrading by Wilfley and Mozley Units

13.3 Phase II Testing

13.3.1 Flotation Tests

Testwork in flotation has centered around using the same procedures in de-sliming

and flotation as used at the Niobec Mine.

Flotation testwork achieved higher recoveries and rougher grades than the gravity

method. While the rougher stage gave good results, the initial cleaning stage tests

were problematic due primarily to non-optimized conditions at this preliminary stage of

testing. These tests indicated that a total oxide grade of more than 30% combined

Nb2O5 and Ta2O5 is achievable although not at high recoveries. Later flotation

testwork used improved de-sliming equipment. Testwork with this equipment and

further flotation work indicated optimum ranges are similar to those obtained at Niobec.

Approximately 11% of the tantalum and 11% of the niobium was lost in this de-sliming

stage. A further 8% of the tantalum and 6% of the niobium were lost in the carbonate

and pyrrhotite rejection steps.

Tests were also performed to optimize the kinetics of the rougher tantalum–niobium

flotation and to test reagent conditions. It has been shown that control of the pH

through the stages is critical. The use of a tallow diamine acetate collector has also

proven to be important. Although this reagent is no longer available as a commercial

product, current practitioners such as the Niobec Mine now purchase the two main

reagents (the amine and acetic acid) and prepare the collector at site.

40

50

60

70

80

90

100

100 1,000 10,000 100,000

Ta R

eco

very

(%

)

Grade Ta (ppm)

Gravity Concentration

Mozley Characterization

Wilfley Table






Process development work is continuing in this area, but for this report, the process

suggested by Test F81 (see Table 13-2) has been chosen as the basis of initial

concentration design as recoveries were good (approx. 70% Ta) for this type of

mineralization and due to a combined grade of 10% Ta–Nb (equivalent to

approximately 14% combined oxides) being achieved. It is expected that with further

work, a combined grade of 30% combined oxides should be achievable.

Table 13-2: Results from F81

Mass Assay Recovery

Products Ta Nb S Ta Nb S

% ppm ppm % % % %

Cyclone Overflow #1 16.5 49 338 0.35 6.9 7.4 9.0

Cyclone Overflow #2 7.1 73 410 0.41 4.4 3.8 4.5

Carbonate Concentrate 28.0 25 151 0.17 5.9 5.6 7.3

Pyrrhotite Concentrate 1.7 152 275 27.38 2.2 0.6 73.5

Magnetic product 0.1 56 395 21.59 0.0 0.0 2.3

Stage 5 Pyrochlore Cleaner Con 0.6 12,839 86,732 1.14 69.8 72.7 1.1

Stage 5 Pyrochlore Cleaner Tail 0.4 228 1816 0.15 0.7 0.9 0.1







Stage 1 Pyrochlore Cleaner Con 11.2 810 5,570 0.11 77.6 82.1 1.9


Total Pyrochlore Rougher Concentrate 25.1 367 2,492 0.07 78.7 82.3 2.8

Flotation Tails 21.5 10 10 0.02 1.8 0.3 0.7

Calculated Feed 117 760 0.64 100 100 100

Assayed Feed 113 764

In addition to flotation tests, preliminary dilute hydrochloric acid leaching tests were

performed. These indicated that low- to intermediate-grade gravity and flotation

products can be upgraded significantly with negligible loss of Ta + Nb. The final

upgrading flowsheet will be based on an economic comparison between pay metal

losses from physical beneficiation and the cost of acid plus stabilization/disposal of the

leach products.

Table 13-3 presents results of a four-stage hydrochloric acid (pH 2, pH 1.2,

6N/1h/100°C, 6N/5h/100°C) on a flotation middling product.






Table 13-3: Results of a Sequential Hydrochloric Acid Leach of Flotation “Middling”

Products Weight Assay (ppm) Distribution (%)

(g) Ta Nb Ta Nb

Stage 1 Filtrate 180.0 0.002 0.00 0.0001 0.000004

Stage 2 Filtrate 220.0 0.031 0.34 0.001 0.002

Stage 3 Filtrate 255.0 0.024 0.63 0.001 0.003

Stage 4 Filtrate 210.0 1.491 12.27 0.056 0.053

Filter Cake 10.7 51,813 453,168 99.9 99.9

Calculated Head 26,824 234,609 100.0 100.0

Assayed Head 20.0 27,663 245,813

There was an indication of the technical feasibility of upgrading by acid leaching with

negligible solution loss of Ta + Nb. The final leach residue assay is > 50% Ta + Nb.

The Stage 3 and 4 strong acid leaches were designed to investigate the possibility of

dissolving Ta + Nb, but the minerals appear to be entirely resistant to this relatively

aggressive leach.

13.4 Phase III Testing

13.4.1 2011 and 2012 Work

A series of tests were performed by AcmeMet, a metallurgical testing facility located in

Vancouver, Canada, in the period up to the first quarter of in 2012. These tests were

performed to optimize the concentration of the tantalum and niobium into a mineral

concentrate. These tests focused on optimizing the flowsheet developed in Phase II.

This optimization work has concentrated on the reduction of reagents involved in the

de-sliming of flotation feed, rejection flotation of carbonates and pyrrhotite and the

subsequent flotation of a tantalum-niobium concentrate for processing through

extractive metallurgy. Although progress has been made in the testwork and the

levels of reagents, there is no major material change from the results indicated by the

Phase I and II testwork. Thus, the performance assumptions, capital and operating

costs are unchanged from those presented in the 2011 PEA. The Phase III testwork is

ongoing.

13.4.2 Review of Concentrate Treatment Options

In Phases I and II, the emphasis of concentration techniques was to create a material

which would be easily upgraded by hydrometallurgical methods, pyrometallurgical

methods, or a combination of both. This has led to a review of those technologies for

the production of high value intermediate products and final products. The process






selected is a process which has been used commercially for the extraction of the

tantalum. The first stage of the process involves the reduction of the concentrate into

metal through the use of the aluminothermic process. Subsequently there would be

chlorination of the granulated metal alloy product and distillation of the anhydrous

metal chloride products to produce high purity Nb and Ta chlorides. Tantalum chloride

is the precursor to capacitor-grade Ta powder, so would be marketed in this form.

Niobium chloride can be sold as a chemical precursor. Both Ta and Nb chloride

products can be readily converted and marketed as high purity Ta2O5 and Nb2O5

oxides respectively.

13.5 Accuracy of Assaying

A review of all calculated and measured feed assay results for tests using bulk sample

BS-2G was performed to check on the accuracy of the chemical analysis and the tests

results. It was decided to continue the assaying of low values, such as tailings, in

duplicate on separate aliquots; this procedure will continue as these assays could

introduce variations to results.


In the opinion of the QPs, the following conclusions are applicable:





Ta and 69% recovery for Nb in the flotation stage.

The metallurgical testwork has shown that it is possible to collect the tantalum and

niobium minerals into a concentrate suitable for extraction of the metals into

saleable products. The first step of the process uses typical grinding followed by

flotation. The secondary treatment or metal extraction of the material is possible

by an existing method such as aluminothermic reduction followed by chlorine

refining. These results are suitable to support estimation of mineral resources for

the deposits.






14.0 MINERAL RESOURCE ESTIMATES

14.1 Introduction

The current resource block model was constructed inside carbonatite only. All

surrounding lithologies including fenite carry fairly low Ta2O5 and Nb2O5 grades and are

considered sub-economic. Generally assay data exist only for carbonatite. There are

some assay values for fenite and other wall rocks but not in sufficient numbers to

support creation of a block model for these lithologies.

14.2 Assay Data and Capping

The resource model was constructed inside carbonatite using 237 diamond drill holes.

Collar, survey, lithology and assay files were exported from the database as csv files,

imported into MineSight® commercial mine modeling software, and combined into a

drill hole assay file.

AMEC conducted grade capping on original samples that are mostly 1 m long.

Capping was required to limit the influence of outliers. The choice of capping was

based on visual inspection of histograms and probability plots. The amount of capping

was small; top-cuts of 1,000 ppm Ta2O5 and 10,000 ppm Nb2O5 were used in

carbonatite. Only six Ta2O5 samples and 13 Nb2O5 samples were capped resulting in

an expected metal removal of 0.24% Ta2O5 and 0.63% Nb2O5.

14.3 Composites

Capped drill core assays were composited down the hole to a fixed length of 2.5 m.

Compositing of Ta2O5 and Nb2O5 was performed in MineSight® software honouring

geologic boundaries. Composites with length less than 1.25 m were merged with the

previous composite. AMEC confirmed that the pre- and post-compositing Ta2O5 and

Nb2O5 means were identical and that compositing resulted in a reduced variability as

indicated by lower CV (CV=coefficient of variation; CV=standard deviation / mean).

This exercise demonstrated that no bias was introduced during compositing. Table

14-1 shows a summary of this check for carbonatite.

Table 14-1: Capped Assays vs. 2.5 m Composites Statistics inside Carbonatites

Variable

Assay

Capped Mean

Assay

Capped CV

2.5 m

Composites

Mean

2.5 m

Composites

CV

Mean diff

(from assays

capped to comps)

Ta2O5 187.4 0.47 187.4 0.35 0.0%

Nb2O5 1465.4 0.91 1465.6 0.78 0.0%






14.4 Exploratory Data Analysis

Exploratory data analysis (EDA) was performed on the composites to better

understand the data used in the resource estimation. This type of investigation reveals

the underlying characteristics of the data. Table 14-2 contains a summary of

univariate statistics for Ta2O5 and Nb2O5 in carbonatite.

Table 14-2: Composite Statistics in Carbonatite

Area/Variable No. Mean Min Max Std. Dev. CV

Carbonatite

Ta2O5 4,171 187.4 3.1 598.9 64.9 0.35

Nb2O5 4,171 1,465.6 7.2 8,293.6 1,147.9 0.78

Note: CV is the Coefficient of Variation and is equal to the standard deviation divided by mean.

Figure 14-1 and Figure 14-2 show arithmetic and log histograms and probability plots

of Ta2O5 and Nb2O5 composites in carbonatite. Both distributions are positively

skewed, and the Nb2O5 distribution is approximately lognormal. The coefficients of

variation are low and support the use of linear grade interpolation methods such as

inverse distance methods.






Figure 14-1: Ta2O5 Histograms and Probability Plot within Carbonatite

Note: Figure prepared by AMEC, June 2012.






Figure 14-2: Nb2O5 Histograms and Probability Plot within Carbonatite


14.5 Contact Analysis

AMEC calculated contact profiles on composite data using in-house software to

analyze grade behaviour at lithology boundaries. There were sharp differences in

grade for each of the variables at the carbonatite/fenite boundary, meaning that values

from outside the carbonatite should be disregarded in the interpolation process of

Ta2O5 and Nb2O5 grade inside the carbonatite. Figure 14-3 and Figure 14-4 show

contact profiles for respectively Ta2O5 and Nb2O5 grade at carbonatite and fenite

boundary.






Figure 14-3: Ta2O5 Contact Plots between Carbonatite and Fenite







Figure 14-4: Nb2O5 Contact Plots between Carbonatite and Fenite







14.6 Variography

Variograms and correlograms are tools used to quantify the spatial variability of a

variable in a geological domain. AMEC used both in-house software and commercially

available Sage2001 software to produce variogram maps, and to construct down-the-

hole and directional correlograms for carbonatite composites. Ta2O5 and Nb2O5

correlograms were created within the entire carbonatite zone. Two spherical models

were used to fit the experimental correlograms; a summary of their parameters is

shown in Table 14-3.

Table 14-3: Ta2O5 and Nb2O5 Correlogram Parameters in Carbonatite

Metal C0*

1st

Structure 2nd

Structure

Rotation (°) Range (m) Rotation (°) Range (m)

C1* Z X Y X Y Z C2* Z X Y X Y Z

Ta2O5 0.336 0.554 -44 0 80 14.2 29.8 21.8 0.110 -44 0 80 32.8 158.4 92.2

Nb2O5 0.118 0.321 -157 -11 90 8.6 11.1 14.6 0.561 -157 -11 90 61.4 214.3 361.3

*C0 – nugget effect; C1-contribution of the 1st structure to the sill; C2-contribution of the 2nd

structure to the sill; sill

has been standardized to value of 1.

The first rotation uses a left hand rule around positive Z axis, the second rotation is a

right hand rule around positive X axis and finally the third rotation is a right hand rule

around positive Y axis. The nugget effects (C0) were modelled from the down-hole

correlograms.

14.7 Carbonatite Solid Modeling

Geological interpretations were provided by Commerce to AMEC in the form of 3D

solids in DXF format. The solids were created by Dahrouge geologists using GEMS™

geological modeling software for the major lithologies with the exception of gneiss,

which was left as a default. The carbonatite solids were provided as 48 structural

(different strike, dip and / or pitch) domains. AMEC reviewed the geological

interpretations and 3D solids and considers them to be appropriate for resource

estimation work.

14.8 Block Model Dimensions

The block model consists of unrotated regular blocks. The block model framework

parameters are listed in Table 14-4.






Table 14-4: Block Model Dimensions

Axis Origin* Block Size (m) No. of Blocks Model Extension (m)

X 352,350 5 250 1,250

Y 5,795,850 5 390 1,950

Z 925 2.5 244 610

Note: *Origin is defined as the bottom southwest corner of the model, located at the lowest combined

northing and easting coordinates and the lowest elevation.

14.9 Assignment of Lithology and Specific Gravity to Blocks

Blocks in the block model were coded by lithology solids. A block was tagged by a

particular solid code if at least 50% of the block volume belonged to this solid. The

volume of each lithology solid was then compared with the volume of the blocks inside

a particular solid. The block model and corresponding lithology solid volumes

compared within ±1%.

Resource block model specific gravity was not estimated; instead a specific gravity

value was assigned by lithology to all blocks in the block model (including blocks

outside of carbonatite) as follows:

2.97 value was assigned to all blocks in carbonatite

2.96 value was assigned to all blocks in fenite

2.82 value was assigned to all blocks in gneiss

3.02 value was assigned to all blocks in amphibolite

2.62 value was assigned to all blocks in pegmatite

3.03 value was assigned to all blocks in skarn

All the above specific gravity values were derived as described in Section 11.3.

14.10 Block Model Grade Estimate

Ta2O5 and Nb2O5 grades were estimated in the carbonatite using an inverse distance

to the power of 3 (ID3) interpolation method. A four-pass interpolation approach was

used with each successive pass having greater search distances. A hard boundary

was used, meaning that composites from outside the carbonatite were not used in the

interpolation process. Estimation was done separately within each domain of the

carbonatite folds. Forty-eight different structural domains were identified and used in

the estimation process.

Table 14-5 shows the estimation search parameters for Ta2O5 and Nb2O5.






Table 14-5: Estimation Parameters for Ta2O5 and Nb2O5

Domain Pass

Search Ellipse Min.

No.

Comp

Max.

No.

Comp

Max.

Comp.

/Hole

Rotation (°) Ranges(m)

Z X Z X Y Z

Common to

all domains

1 differ

by

domain

differ

by

domain

differ

by

domain

50 50 5 5 8 2

2 100 100 5 3 8 2

3 150 150 5 2 8 2

4 300 300 50 2 8 2

The rotation angles of the search ellipse are the same for each pass, but they are

different for each of the 48 structural domains. They reflect average strike, dip, and

pitch of each fold limb / domain.

14.11 Block Model Validation

The block model grades were validated by visual inspection comparing composites to

block grades on-screen, declustered global statistics checks, local biases checks using

swath plots, and finally model selectivity checks.

14.11.1 Visual Validation

AMEC completed a visual inspection of composites and blocks in vertical sections and

plan views. Figures 14-5 to 14-8 show colour-coded Ta2O5 or Nb2O5 composites and

corresponding ID3 block models on plan and in section. The model generally honours

both Ta2O5 and Nb2O5 data well, and grade extrapolation is well-controlled where

sufficient data exist.






Figure 14-5: Ta2O5 ID3 Model within Carbonatite – Plan 1,146.25


Figure 14-6: Ta2O5 ID3 Model within Carbonatite – Section N 5,796,932.5







Figure 14-7: Nb2O5 ID3 Model within Carbonatite – Plan 1,146.25


Figure 14-8: Nb2O5 ID3 Model within Carbonatite – Section N 5,796,932.5







14.11.2 Global Grade Bias Check

The ID3 block models were checked for global bias by comparing the average grade

(with no cut-off) from these models with that obtained from nearest-neighbour (NN)

model estimates (Table 14-6). The NN estimator produces a globally unbiased

estimate of the average value when no cut-off grade is imposed and is a good basis

for checking the performance of different estimation methods. Table 14-6 shows that

global biases are well below the AMEC-recommended guideline of ±5% (relative

difference).

Table 14-6: Mean Grades for NN and ID3 Models

Model Ta2O5 Nb2O5

Nearest Neighbour 188.3 1,449.6

Inverse Distance (ID3) 188.5 1,439.2

% Diff (ID3 – NN)/NN 0.1% -0.7%

AMEC also estimated the impact of outlier capping on the estimated global mean of

the model. A comparison of global means of capped and uncapped ID3 models

showed the amount of metal removed by capping is minor (0.2% for Ta2O5 and 0.6%

for Nb2O5); it is almost identical to the expected metal removal based on composite

analysis (see Section 14.2).

14.11.3 Local Grade Bias Check (Swath Plots)

Checks for local biases were performed for Ta2O5 and Nb2O5 by creating and

analyzing local trends in the grade estimates using swath plots. This was done by

plotting the mean values from the NN estimate versus the ID3 estimates in east-west,

north-south and vertical swaths or increments. Swath intervals are 50 m in both the

northerly and easterly directions, and 10 m vertically. Swath plot checks using only

Indicated blocks for the ID3 Ta2O5 model are shown on Figure 14-9. Figure 14-10

shows corresponding swath plots using only Indicated blocks for Nb2O5.






Figure 14-9: Swath Plot for Ta2O5 ID3 Model


Figure 14-10: Swath Plot for Nb2O5 ID3 Model







In the upper row of the swath plots, the black line represents the ID3 model grades,

the red line represents the NN model grades, and the blue line represents the

composite grades. In the lower row of swath plots, the lines represent the number of

blocks contained in each swath, and the number of composites. Because the NN

model is declustered, it is a better reference to validate the resource block model.

Composites are not declustered and only provide an indicative check. Swath plot

checks conducted by AMEC show that there are no local biases between ID3 and NN

models for estimated Ta2O5 and Nb2O5 in the carbonatite.

14.11.4 Selectivity Check

Selectivity analysis for Ta2O5 and Nb2O5 was completed using the Discrete Gaussian

Model for change of support from composite size to a selective mining unit (SMU) size.

This was done using AMEC in-house software (Herco). The aim of this analysis was

to assess whether the estimated resource reasonably represents the recoverable

resources (represented by Herco curves) relative to the proposed mining method. The

selectivity analysis assumed a 10 m by 10 m by 5 m block as the smallest SMU size

for Blue River.

The results of the Herco analysis are generally discussed in terms of smoothness. An

over-smoothed model may over-estimate the tonnes and under-estimate the grade.

The model with an appropriate amount of smoothing will follow the Herco grade and

tonnage curves for values corresponding to different economic, or grade cut-offs.

The Herco analyses were undertaken using only Indicated blocks in order to obtain a

good understanding of the model selectivity, or smoothness. Inferred blocks are often

extrapolated and over-smoothed for lack of data are not recommended for use in this

analysis. Herco grade–tonnage curves checks using only Indicated blocks for ID3

Ta2O5 model are shown in Figure 14-11. Figure 14-12 shows the corresponding Herco

checks for Nb2O5. On both graphs, the upward-trending blue line represents the ID3

model grades, while the paired red line represents the Herco model grades. The

downward trending blue line represents the ID3 model tonnage, while the paired red

line represents the Herco model tonnes.

The Herco selectivity analyses show that the Ta2O5 and Nb2O5 ID3 models are

properly smoothed for the cut-offs of interest. These models are used for tabulating

the Blue River Mineral Resources.






Figure 14-11: Herco Grade – Tonnage Curves for Ta2O5 ID3 Model


Figure 14-12: Herco Grade – Tonnage Curves for Nb2O5 ID3 Model







14.12 In Situ Block Model Carbonatite Reconciliation

There is a 6% change in block model carbonatite volume going from the block model

used for the 2011 PEA to the current block model. The reason for this difference is a

change in the structural geology interpretation in carbonatite, resulting in a slight

increase of the overall carbonatite 3D solid volume.

There are 23% of blocks from the 2011 PEA block model that are not in the current

model, and there are 27.4% of blocks from the current model that are not in the block

model used for the 2011 PEA. The reason for these changes is again the result of the

structural geology interpretation of the carbonatite which modified somewhat the

position of the carbonatite in 3D space.

The differences in average grade from the block model used for the 2011 PEA to the

current block model within only those blocks estimated in both models are 1.4% for

Nb2O5 and 1.5% for Ta2O5.

14.13 Mineral Resource Classification

The Mineral Resources were classified in accordance with the 2010 Canadian Institute

of Mining, Metallurgy, and Petroleum (CIM) Definition Standards for Mineral Resources

and Mineral Reserves, whose definitions are incorporated by reference into NI 43-101.

Mineral resources are required to be classified as Inferred, Indicated, and Measured

according to increasing confidence in geological, grade continuity, and other aspects

impacting the resources. However, there are no regulatory specifications as to the

procedure to use to achieve that classification.

In addition to criteria such as sufficient geological continuity, grade continuity, and data

integrity, one AMEC guideline for resource classification is to have drill hole spacing

sufficient to predict potential production with reasonable precision. As such AMEC

conducted drill hole spacing studies taking into account both grade continuity and ore

tonnage / volume uncertainty.

Based on these drill hole spacing studies AMEC established the following criteria for

classification of mineral resources at Blue River:

Inferred Mineral Resources:

Minimum one drill hole

Distance to the closest composite less than 100 m






Indicated Mineral Resources:

Minimum two drill holes


Distance to the second closest composite less than 60 m

Measured Mineral Resources:

Minimum three drill holes


Distance to the second closest composite less than 30 m

The above criteria are somewhat more restrictive than the criteria given in the. 2011

PEA, because the tonnage / volume uncertainties are now better understood and were

incorporated in the most recent drill hole spacing studies.

The current mineral resource classification at Blue River is restricted to Indicated or

Inferred, based on the following:

Confidence limits drill hole spacing studies

Concerns over analytical precision and provisional accuracy for the sample dataset

from 2005 to 2009

Required metallurgical testwork on the final stage of the proposed metallurgical

process is still ongoing to support proof-of–concept.

Eighty-two per cent of the carbonatite blocks are classified as Indicated. Seventeen

per cent of the carbonatite blocks are classified as Inferred. One per cent of the block

model in carbonatite is unclassified. Blocks that fall into the unclassified category are

within carbonatite solids that were intersected typically by one isolated drill hole. The

geological continuity and volume of those solids cannot be reasonably assumed.

Figure 14-13 and Figure 14-14 show examples of the resource classification.






Figure 14-13: Resource Classification – Plan 1,161.25

Note: The following block colour scheme is used in the figure: Green – Indicated; Yellow – Inferred; Red

– Unclassified; drill hole projection ± 2.5 m. Figure prepared by AMEC, June 2012.

Figure 14-14: Resource Classification – Section N 5,796,882.5

Note: The following block colour scheme is used in the figure: Green – Indicated; Yellow – Inferred; Red

– Unclassified; drill hole projection ± 20 m; view north. Figure prepared by AMEC, June 2012.






14.14 Reasonable Prospects for Economic Extraction

To assess reasonable prospects for economic extraction, AMEC assumed that the

Blue River deposit would be mined utilizing self-supported mining methods under a

conceptual scenario that considers mining and processing at a rate of 7,500 tonnes

per day. Mining and economic parameters applied for the 2012 estimate were

adjusted based on the results from the 2011 PEA.

14.14.1 Market Study

The marketing assumptions are discussed in Section 19.

14.14.2 Commodity Price

Commodity price assumptions are discussed in Section 19.

14.14.3 Physical Assumptions

Tantalum-niobium mineralization is hosted in carbonatite

Continuous mineralization is found in moderately flat and wide carbonatite bodies

with modest dips

Mineralized areas 20 m to 70 m in height are expected in several zones

Steep topography provides access to the mineralized areas in the form of adits on

the hillsides

Fair to good rock conditions are expected in the majority of the deposit

Identified faulted zones may require wider pillars to avoid unstable mining

conditions.

14.14.4 Operational Considerations

The underground mining methods envisaged are sublevel open stoping and room

and pillar without backfill

Mining recovery is assumed to vary from 65% to 85% depending on the mine and

stope layout and the success in which pillars can be mined on retreat. The rest of

the resource is expected to remain in place as pillars for stability considerations.

A bulk mining method with minimum stope size of 10 m x 10 m rooms with 15 m

height is assumed in order to attain a relatively high production rate of 7,500

tonnes per day






A more selective method with stopes size of 10 m x 10 m rooms with 5 m height is

assumed to capture higher-grade material located on the thinner edges of the

mineralized zones

An external dilution factor was not considered during this estimation

Planned internal dilution within the minimum stope size is included

The concentration method considered is flotation followed by a refining process on

site; global recoveries to obtain metal grade products were assumed to be 65.4%

for tantalum and 68.2% for niobium.

14.14.5 Economic Assumptions

Since the block unit value (see Section 14.14.6) is estimated using commodity prices

expressed in US dollars, the costs and assumptions are also expressed in US dollars.

The following operating cost and price assumptions were adopted from the 2011 PEA,

but rounded. The exchange rate used to calculate the block unit value is US$0.92 =

CAD$1.00.

Mining cost – bulk mining method ................................. $US24/tonne

Mining cost – selective mining method .......................... $US42/tonne

Processing and refining cost ......................................... $US13/tonne

General and Administration ............................................. $US3/tonne

Base case scenario price of tantalum$US317/kg Ta metal in an oxide product

Base case scenario price of niobium $US46/kg Nb metal in an oxide product

14.14.6 Economic Cut-Off

Block Unit Value

The block model was adapted to represent the two payable metal contents in terms of

Block Unit Value (BUV) in US$/t using the following formula:

BUV = (Ta2O5 grade in ppm * Ta recovery factor * Ta price in US$/g *

proportion of 2Ta:Ta2O5) + (Nb2O5 grade in ppm * Nb recovery factor * Nb

price in US$/g * proportion of 2Nb:Nb2O5)

For the base case scenario:

BUV = (Ta2O5 * 0.654 * 0.317 * 0.819) + (Nb2O5 * 0.682 * 0.046 * 0.699)






The tool “Stope Analyzer” from Vulcan® was utilized to identify the blocks that exceed

the cut-off value while complying with the aggregation constraint of minimum stope

size. This tool “floats” a stope with the specified dimensions and flags each block

when the average block unit value of the contained blocks within a stope exceeds the

designated cut-off value.

For constraining resources deemed to be mined by underground methods, the use of

this tool as an alternative to a conventional economic grade-shell provides an

advantage based on the ability to aggregate blocks into the minimum stope

dimensions and the automatic elimination of outliers that do not comply with this

condition.

For purposes of estimating the current model Mineral Resources, an underground

mining cut-off value of US$40/t BUV was established for the material susceptible to be

mined by bulk methods; the direct operating cost estimate for the bulk mining method

is the result of the 2011 PEA mine plan and its associated costs. For material to be

mined by a selective mining method a cut-off value of US$58/t BUV was adopted; the

direct operating cost for the selective mining method is an estimate of the associated

costs when a room and pillar mining method is assumed and was based on the work

performed during the 2011 PEA.

14.15 Mineral Resource Statement

The Mineral Resources were classified in accordance with the 2010 CIM Definition

Standards for Mineral Resources and Mineral Reserves, whose definitions are

incorporated by reference into NI 43-101.

Table 14-7 shows the estimated mineral resources. The Indicated Mineral Resources

are 51.78 million tonnes at 192 ppm Ta2O5 and 1,490 ppm Nb2O5. Inferred Mineral

Resources are 8.8 million tonnes at 186 ppm Ta2O5 and 1,660 ppm Nb2O5.






Table 14-7: Blue River Project Estimated Mineral Resources; Effective Date 22 June,

2012, Tomasz Postolski, P.Eng, Qualified Person

Ta price

[US$/kg]

Confidence

Category Tonnes

Ta2O5

[ppm]

Nb2O5

[ppm]

Contained

Ta2O5

[1000s of kg]

Contained

Nb2O5

[1000s of kg]

317 Indicated 51,780,000 192 1,490 9,930 76,900

Inferred 8,800,000 186 1,660 1,600 14,600

Notes:

1. Assumptions include commodity prices of US$317/kg Ta, US$46/kg Nb, process recoveries of 65.4%

for Ta2O5 and 68.2% for Nb2O5, US$24/tonne mining cost, US$13/tonne process and refining cost,

US$3/tonne G&A cost


“Stope Analyzer”



US$58/t (selective)


included


This Mineral Resource estimate is supported by a base case price assumption of

US$317/kg Ta, which is higher than historic average prices. The Ta and Nb metal

prices assumptions are the same as those used in the 2011 PEA and are considered

still reasonable based on publicly available information on the current market prices.

Table 14-8 shows the sensitivity of the Blue River Mineral Resources to tantalum metal

price. Sensitivities are based on a fluctuating metal price but could also represent

fluctuating mining or processing costs or metallurgical recoveries or a combination of

all of these factors.






Table 14-8: Blue River Project Sensitivity of Estimated Mineral Resources to Tantalum

Price; Effective Date 22 June 2012, Tomasz Postolski, P.Eng, Qualified

Person

Ta price

[US$/kg]

Confidence

Category

Mass

[tonnes]

Ta2O5

[ppm]

Nb2O5

[ppm]

Contained

Ta2O5

[1000s of kg]

Contained

Nb2O5

[1000s of kg]

470 Indicated 55,050,000 189 1,430 10,430 78,750

Inferred 9,800,000 182 1,610 1,800 15,700

381 Indicated 54,230,000 190 1,440 10,310 78,270

Inferred 9,300,000 184 1,630 1,700 15,300

317 Indicated 51,780,000 192 1,490 9,930 76,900

Inferred 8,800,000 186 1,660 1,600 14,600

272 Indicated 47,700,000 194 1,560 9,250 74,400

Inferred 8,100,000 187 1,700 1,500 13,900

238 Indicated 43,170,000 196 1,650 8,440 71,270

Inferred 7,500,000 188 1,760 1,400 13,200

Notes:

1. Ta price was varied and all other assumptions remained the same as base case. Base case is in bold.


“Stope Analyzer”.



US$58/t (selective)


included.


The Mineral Resources have been assessed for reasonable prospects for economic

extraction using assumptions based on similar deposits. Economic viability of the

Mineral Resource can only be demonstrated by Pre-Feasibility and Feasibility Studies,

and there is no assurance that the stated resources can be upgraded in confidence

and converted to mineral reserves. Since underground mining methods are

envisioned, the mining recovery may vary from 65% to 85% depending on the success

in which pillars can be mined on retreat and/or fill is utilized.

14.16 Comparison of Mineral Resources

The Indicated Mineral Resources in the current model total 51.78 million tonnes at

192 ppm Ta2O5 and 1,490 ppm Nb2O5. Inferred Mineral Resources total 8.8 million

tonnes at 186 ppm Ta2O5 and 1,660 ppm Nb2O5.






The Indicated Mineral Resources in the model used for 2011 PEA were 36.35 million

tonnes at 195 ppm Ta2O5 and 1,700 ppm Nb2O5. Inferred Mineral Resources were 6.4

million tonnes at 199 ppm Ta2O5 and 1,890 ppm Nb2O5.

There is a considerable increase in resource tonnes for the current Mineral Resource

update relative to the 29 September 2011 tonnage estimate where the Indicated

category has increased by 42% and the Inferred category by 37%. This increase in

tonnes is mostly due to (1) lowering the bulk mining method block unit value cut-off

from US$52/t to US$40/t by eliminating backfill costs, and to a lesser extent (2)

additional infill diamond drilling.


The QPs are of the opinion that the Mineral Resources for the Project, which have

been estimated using core drill data, have been completed using industry best

practices, and conform to the requirements of CIM (2010). The QPs are not aware of

any known environmental, permitting, legal, title, taxation, socio-economic, marketing,

political or other relevant factors that could materially affect the Mineral Resource

estimate. From the 2011 PEA, the Project is most sensitive to the American to

Canadian exchange rate, the operating costs, and the Ta and Nb metal commodity

prices.






15.0 MINERAL RESERVE ESTIMATE

No Mineral Reserves have been estimated for the Project.






16.0 MINING METHODS

16.1 Introduction

This mining section incorporates assumptions, analysis and findings of the Preliminary

Economic Assessment that had an effective date of 29 September 2011 (the 2011

PEA) by AMEC (Chong et al., 2011). The mineral resources used in the 2011 PEA

mine plan were those with an effective date of 29 September 2011.

The preliminary mine plan presented in this section is partly based on Inferred Mineral

Resources that are considered too speculative geologically to have the economic


Reserves, and there is no certainty that the Preliminary Economic Assessment based

on these Mineral Resources will be realized.

The information relevant to the preliminary mine plan prepared during the 2011 PEA

based on a bulk mining method is included in this section and has not been updated

because AMEC considers that the assumptions supporting the outcomes remain

reasonable. The effective date of the 2011 PEA results remains 29 September 2011.

16.2 Optimization

16.2.1 Assumptions

Mining assumptions used to define the Mineral Resource estimate in Section 14 were

adapted from the 2011 PEA mine design.

The block model consists of regular blocks with dimensions of 5 m x 5 m in the

horizontal plane and 2.5 m vertically; no rotation was adopted.

The block model was adapted to represent the unit value per tonne of material

considering two payable metal contents. This value was named the block unit value

(BUV) in US$/t and was estimated using the following formula:

BUV = (Ta2O5 grade in ppm * Ta recovery factor * Ta price in US$/g *

proportion of 2Ta: Ta2O5) + (Nb2O5 grade in ppm * Nb recovery factor *

Nb price in US$/g * proportion of 2Nb:Nb2O5)

Price assumptions were US$317/kg tantalum metal and US$46/kg niobium metal,

contained in oxide product.






Tantalum and niobium mineral occurrences are amenable to conventional flotation and

refining processes with estimated overall recoveries of 65.4% and 68.2%.

The block model was valued using the following formula, incorporating the specific

grades in each block:

BUV = (Ta2O5 * 0.654 * 0.317 * 0.819) + (Nb2O5 * 0.682 * 0.046 * 0.699)

The 29 September 2011 Mineral Resource estimates used an underground mining

cut-off value of US$52/t for the material susceptible to be mined by bulk methods and

a cut-off value of US$59/t for material to be mined by the selective methods. Mineral

Resources considered in the preliminary mine plan are those tabulated in Table 16-2.

The “Stope Analyzer” tool from commercially-available Vulcan® software was utilized

to identify blocks within the resource model for which the block unit value (BUV)

exceeded the cut-off value while complying with an aggregation constraint of a

specified minimum stope size. This tool “floats” a stope with previously specified

dimensions and flags each block when the average block unit value of the contained

blocks within a stope exceeds the designated cut-off value.

The use of this tool is an alternative to a conventional economic grade-shell for

constraining Mineral Resources deemed to be mined by underground methods, and

provides an advantage based on the ability to aggregate blocks into the minimum

stope dimensions and the automatic elimination of outliers that do not comply with this

condition. Stope dimensions are detailed in Table 16-1.

Table 16-1: Minimum Stope Dimensions for Constraining the Subset of Mineral

Resources within Designed Stopes

Mining Method Width (m) Length (m) Height (m)

Sub-level Open Stoping 10 10 15

Room and Pillar 10 10 5

16.2.2 Mining Method

Blue River Project is located largely along the steep, west-facing slopes of the

Monashee Mountains and it is closely positioned to, and just east of, the North

Thompson River. As mineralization is close to surface, extraction of the mineralized

material could potentially be by either open pit or underground methods, or a

combination of both.






Initial assessment identified technical challenges and increased costs for tailings and

waste rock disposition related to local topography, stream courses and climate. For

this reason a decision was made to consider an underground mining scenario for PEA

purposes.

16.2.3 Mineral Resources considered for the 2011 PEA

The 2011 PEA was based on a mineral resource estimate announced in February

2011 (“Blue River Ta-Nb Project NI 43-101 Technical Report, Blue River, British

Columbia” by AMEC with effective date 31 January 2011). For this estimate, AMEC

used the drill results up to the end of 2009, which includes 183 drill holes comprising

37,446 metres of HQ drill core and 8,218 sawn core samples to develop the mineral

resource estimate. By inspection, the volume of rock (tonnes and grade) above cut-off

considered in the 2011 PEA are not materially different from the volume of rock above

the same cut-off in the updated 22 June 2012 mineral resource estimate.

Table 16-2 shows the estimated mineral resources used in the 2011 PEA. AMEC

cautions that Mineral Resources are not Mineral Reserves as they do not have

demonstrated economic viability.

Table 16-2: Blue River Project Estimated Mineral Resources Supporting 2011 PEA;

Effective Date 29 September 2011, Tomasz Postolski, P.Eng., Qualified

Person

Ta price

[US$/kg]

Confidence

Category Mass [tonnes]

Ta2O5

[ppm]

Nb2O5

[ppm]

Contained

Ta2O5

[1,000s of kg]

Contained

Nb2O5

[1,000s of kg]

317 Indicated 36,350,000 195 1,700 7,090 61,650

Inferred 6,400,000 199 1,890 1,300 12,100

Notes:

1. Assumptions include US$317/kg Ta, US$46/kg Nb, 65.4% Ta2O5 recovery, 68.2% Nb2O5 recovery,

US$32/tonne mining cost, US$17/tonne process and refining cost. Mining losses = 0% and dilution =

0%.


“Stope Analyzer”.


mining method. The block unit value cut-off ranged from US$52/t (bulk) to US$59/t (selective)


To support the 2010 Mineral Resource estimate underground mining methods were

envisioned (room and pillar or variants), with mining recovery assumed to vary from

65% to 85% depending on the success in which pillars could be mined on retreat

and/or fill is utilized.






16.2.4 Production Rate

In 2009, AMEC carried out an “order of magnitude” financial analysis to evaluate a

range of processing throughput rates. This analysis indicated that the Project required

processing rates of greater than 5,000 t/d to provide economies of scale that allowed a

reasonable economic return on investment.

A processing rate of 7,500 t/d was assumed for Mineral Resource estimation and for

the conceptual design of an underground mine for the Project.

It is AMEC’s opinion that this mining rate is reasonable based on the geometry of the

deposit and the mining method selected.

16.3 Geotechnical Conditions

In the second half of 2010 AMEC carried out a geotechnical program to provide design

guidelines for the Blue River Project. This program included:

A geotechnical site investigation program;

Training for site geologists and geological technicians in oriented core logging;

QA/QC site visits during the geotechnical drilling program; and

Engineering analysis and recommendations applicable to the underground mining

method.

Rock types have been grouped into two main geotechnical domains: Intrusive and

Layered Rocks. The Intrusive group encompasses carbonatite and fenite rocks while

the Layered Rocks group encompasses gneiss and amphibolite.

Based on rock strength analysis and assessment of rock quality designation (RQD),

joint spacing, joint condition and groundwater condition assumptions, the rock mass

characteristics detailed in Table 16-3 were obtained for each rock group.

Table 16-3: Rock Mass Characteristics by Rock Group






A structural set analysis showed similarities between the Intrusive and Layered Rock

groups with a strong concentration of flat joints dipping between 0° to 30° to the

east/northeast and sub-vertical joints striking northwest/southeast. The major joint

sets were identified as shown in Table 16-4.

Table 16-4: Major Joint Sets

16.4 Conceptual Mining Method

The major items governing the selection of underground mining methods include the

geology and geometry of the deposit and geotechnical properties of the mineralized

material and country rock.

AMEC followed the technique proposed by Nicholas (1992) to rank the underground

mining methods suitable for the Blue River deposit. AMEC classified the deposit as a

thick, tabular, and flat deposit with relatively uniform low grades and moderate

geotechnical conditions in the deposit as well as in the host rock. The three methods

with higher ranking are:

Sub-level stoping

Cut-and-fill

Sub-level caving

The first two methods were considered for this study. The first was evaluated for

areas where the carbonatite is greater than 15 m in thickness and the second for

thinner areas towards the edges of the carbonatite folds.

16.4.1 Backfill Considerations

During economic analysis it was found that the cut-and-fill method was likely to prove

uneconomic and it was excluded from the mine plan.

The 2011 PEA was developed assuming a sub-level open stoping mining method with

no backfill and no pillar recovery.






16.5 Stoping Design

The Upper Fir deposit geological model shows thicknesses between 20 m to 80 m and

strike lengths between 50 m to 200 m in the east–west direction. Transverse

dimensions in the north–south direction range between 100 m to 500 m. Vertical

stopes oriented east–west with maximum dimensions of 30 m high, 15 m wide and

60 m long were selected for preliminary analysis.

Based on these stope dimensions, the following stope faces and hydraulic radius are

defined (Table 16-5).

Table 16-5: Stope Faces and Hydraulic Radius

Face A Face B Face C

Longwall Endwall Back

Height (m) 30 30

Width (m) 15 15

Length (m) 60 60

Hydraulic radius (m) 10 5 6

16.5.1 Stability Analysis and Ground Support

As part of geotechnical analysis, AMEC carried out stability analysis based on the

empirical Mathews/Potvin Stability Graph method.

Results of the analysis indicated:

Face A – Longwall: This wall is expected to be composed of competent Intact to

Blocky rock mass. The main jointing is sub-horizontal crosscut by two sub-vertical

joint sets. The stope wall is vertical and should be inherently stable unless

disturbed by over-blasting. Sloughing up to 0.5 m thick is expected. Cable bolting

may be required in certain places.

Face B – Endwall: Numerical modeling suggests that the stress is relatively high

compared with the rock strength and the joint set configuration favours sliding

failure. Cable bolts are required to provide stability. A square cable pattern of 2 m

by 2 m is recommended and will be installed from the sill drift fanning upwards to

cover the unsupported span between level drifts.

Face C – Back: This face is horizontal and as such is expected to be in a relaxed

stress state. Gravity driven rock fall is likely the dominant failure mode. Face

reinforcement is required with the development of a self supporting rock arch over






the back face. A 15 m wide face requires cable bolting in square pattern of 2 m by

2 m and cable bolts of 9 m long.

16.5.2 Stope Geometry

AMEC made adjustments to the preliminary stope design to reduce the demand for

cable bolting; specifically the Back was reduced to a 5 m-wide entry. This requires the

stope drilling pattern to be changed from vertical to fanned holes.

16.5.3 Mining Sequence

The stope face will advance from the east end of each block to the west in a retreat

manner. Once the maximum unsupported stable length of the stope has been

reached, a pillar will be established, and mining will resume from a new stope.

16.5.4 Conceptual Mine Design

AMEC developed a conceptual design for one of the potential areas to be mined,

geological domain A110. This domain was selected because it was one of the thicker

areas within the deposit and represented a reasonable test area for mine design.

16.5.5 Mining Dilution and Recovery

Material deemed to be mined by bulk mining methods represents 84% of the Mineral

Resources. Within the mineable shapes there was internal dilution of 2% waste rock.

It was assumed that during mining 2% of waste material would be added as external

dilution and 2% of the broken material would not be recovered from the stopes due to

operational conditions.

The geotechnical investigation indicates that an extraction ratio of 67.5% is

reasonable. Applying this factor to subset the Mineral Resources considered in the

mine plan results in an overall mining extraction of 58% and provides 25.0 Mt of

material as run-of-mine (ROM) production to be processed. Applying internal and

external mining dilution, the overall subset Mineral Resource grades were diluted to

185 ppm of Ta2O5 and 1,591 ppm of Nb2O5 for mine planning purposes.

16.6 Drilling and Blasting

AMEC assumed conventional drilling methods: electro-hydraulic jumbos for face drift

rounds and electro-hydraulic long-hole drills for stope drilling.






Due to the high precipitation in the region and water continuity along fractures and rock

layers, wet conditions were assumed for development and stoping areas requiring the

use of emulsion type explosives.

AMEC assumed the utilization of specifically-designed tanks to store emulsion

explosives in trucks delivering and loading explosives at development faces and

stopes. A centralized blasting system where all blasts are initiated sequentially from a

single location at the end of each shift will be used to reduce potential safety and

ventilation risks.

16.7 Mine Development

The deposit will be accessed through two main portals, the Upper and Lower Portals,

located in positions where the deposit outcrops on the hillside. The Upper Portal will

be located at Elevation 1,150 m. It will be used as the main entry and will have most

of the mine services. The Lower Portal will be located at Elevation 1,030 m and will be

used for haulage trucks access. Access to the portals will be by a road upgraded from

existing exploration roads. Location plans for the portals are included in Section 18.

The mine will be accessed by adits driven in pairs from the portals. All entries, ramps,

drifts and crosscuts will be 5 m wide by 5 m high with a semi-arched back to

accommodate the haulage trucks plus mine services, including pipelines for

compressed air, drill and drainage water, for ventilation ducts and electric and

communication cables.

Ramps will be driven at grades to a maximum of 15% to provide access to the

production areas. Two ramps or adits will be driven to each area to provide single-way

traffic of haulage trucks and to facilitate ventilation.

Top access crosscuts are driven from the main ramps to each level on vertical

intervals between 20 m to 30 m. Stope access crosscuts are driven along the levels

from west to east. Bottom-access crosscuts are driven to function as mucking drifts.

Ground support will be by grouted bolts in all man-entry drifts with steel mats, wire

mesh and plates assumed to be installed in 20% of the areas.

The length of development necessary was calculated by designing centrelines of

ramps and drifts to provide access to all potentially mineable areas of the mine.

Additional development for re-mucking cut-outs, sumps, substation rooms, storage and

any other excavation needed for infrastructure was included as an allowance of 20% of

the semi-permanent development entries. The total development was estimated at

92,500 m during the life-of-mine.






A conceptual preliminary overall mine plan is included as Figure 16-1. Figure 16-2

displays an aerial view of the mining area from the Upper Portal.

16.8 Mineralized Material and Waste Rock Haulage

The PEA envisages that tailings material will be dry-stacked and waste rock will be

stored in the same general area. For convenience, the combined storage area is

referred to as the “co-disposal facility”.

Radio remote-controlled load-haul-dumps (LHDs) will be used to extract the

mineralized material from beyond the safety of the stope brow. This material will then

be loaded directly into the haulage trucks that will be spotted at the end of each stope

crosscut.

Underground trucks will haul the mined material through the access drifts and ramps,

unloading into the primary crusher surface stockpile located close to the Lower Portal.

Crushed material will be transferred to the process plant by a belt conveyor.

Waste from development will initially be utilized for construction of a structural shell for

the tailings co-disposal site located between Elevations 1,400 m and 1,600 m in an

area east of the process plant. The conveyor will be used to transport this material in

batches from the mine to a stockpile by the plant site.

A road developed at +10% grade will connect the plant with the co-disposal site.

Surface trucks will haul waste from the plant to the co-disposal site.






Figure 16-1: Conceptual Mine Layout Plan (plan view projection)

Note: image figure colours may appear darker than reference key colours due to over-plotting of design layers. Figure prepared by AMEC, 2011.






Figure 16-2: Aerial View of the Mining Area from Upper Portal

Note: image figure colours may appear darker than reference key colours due to over-plotting of design layers. Light brown blocks in background of

figure are the potentially-mineable blocks. Figure prepared by AMEC, 2011.





Project No.: 168967 Page 16-12 22 Jun e2012

16.9 Mine Services

Underground mine services such as ventilation and air heating, compressed air, water

for drilling and electric power supply will be provided to the mine via adits from the

portals. The main power substation and the air compressors will be installed in

facilities located adjacent to the Upper Portal.

Other mine services will include all the systems and supplies needed for the mining

operations, including explosives storage, communications, monitoring and control

systems, road maintenance and an underground equipment maintenance facility.

Portable self-contained refuge stations will be provided for the mine and will be located

at convenient locations. Refuge stations provide a common assembly area in the

event of a mine fire or other emergency and are portable so they can be easily

relocated to the next active area.

16.10 Mine Development and Production Forecasts

The forecast preproduction development is 12,000 m and the annual development

over the ten-year mine life is forecast to decrease from 15,000 m in the first full

production year to about 2,150 m in the final year.

Production was estimated at 2.7 Mt/a of mineralized material. Following the

preproduction development year full production is maintained for nine years followed

by decreased production in Year 10 as the subset of the Mineral Resources

considered in the mine plan are exhausted.

At this preliminary level of study the stope mining sequence was not defined and

therefore average grades were used for each year in the mine plan.

There is opportunity to increase the net present value (NPV) of the project by mining

higher-grade zones early in the mine life providing that the sequence and overall

recovery of the stopes is not negatively affected.

The development and production forecasts are shown in Table 16-6.






Table 16-6: Mine Development and Production Forecasts

Production Schedule Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5

Nominal Production Rate tpd

7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360 360

Production of Mineralized Material t/year

2,700,000 2,700,000 2,700,000 2,700,000 2,700,000

Ta2O5 Grade ppm

185 185 185 185 185

Nb2O5 Grade ppm

1,591 1,591 1,591 1,591 1,591

Development (total) m/year 12,000 15,000 12,834 10,980 9,394 8,038

Capital Development m/year 12,000 5,000 1,216 1,040 890 761

Operational Development m/year 0 10,000 11,618 9,940 8,505 7,276

Production Schedule Units Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Total

Nominal Production Rate tpd 7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360

Production of Mineralized Material t/year 2,700,000 2,700,000 2,700,000 2,700,000 700,000 25,000,000

Ta2O5 Grade ppm 185 185 185 185 185 185

Nb2O5 Grade ppm 1,591 1,591 1,591 1,591 1,591 1,591

Development (total) m/year 6,877 5,884 5,034 4,307 2,153 92,500

Capital Development m/year 651 557 477 408 0 23,000

Operational Development m/year 6,225 5,326 4,557 3,899 2,153 69,500






16.11 Mine Equipment Requirements

Table 16-7 shows the major mining equipment and the support equipment required

each year for the life of mine and includes allowances for equipment utilization and

availability due to maintenance downtime. The table includes equipment required for

the tailings and waste co-disposal facility.

The “Equipment additions” reflect additional requirements based on increased activity

levels and replacements assuming typical useful operating lives for the units.

16.12 Mine Infrastructure

The mine infrastructure planned for the Project is listed in the capital costs section of

this report and includes establishment of mine portals, access roads to portals

underground maintenance bays, ventilation and heating systems, air compressors, fuel

tanks, explosives magazines, pumps, electrical transformers and ancillary equipment.

16.13 Mining Personnel

The mine is scheduled for three 8-hour shifts per day, 360 days per year. This will

require an equivalent of 4.5 mine crews working a rotating schedule for activities

scheduled seven days a week and three crews for activities scheduled five days a

week. The annual personnel requirements are shown in Table 16-8. This includes

management and supervisory personnel and personnel to operate and maintain

equipment used to service the surface roads and the tailings disposal site.


In the opinion of the QPs, the following conclusions are appropriate:

The deposits are amenable to underground mining, and the PEA developed

assuming a Base Case sub-level open stoping mining method with no backfill is

still valid

Material deemed to be mined by bulk mining methods is a subset of the Mineral

Resources supporting the 2011 PEA, and represents 84% of the Mineral

Resources estimated at that time; within the stopeable shape, an additional 2% of

waste was identified as internal dilution.






Table 16-7: Mining and Tailings Facility Equipment Requirements

Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Major Equipment Units Required

Jumbo - 2 boom 4 4 4 4 3 3 3 3 2 2 2

Longhole dril l - 5 5 5 5 5 5 6 6 6 4

Bolter 3 3 3 3 3 2 2 2 2 1 2

Emulsion Truck 3 4 4 3 3 3 3 3 2 2 3

Scissor Lift 4 4 4 4 3 3 3 3 2 2 2

LHD - Ejector - 7 m3 3 5 5 5 5 5 5 5 4 4 4

Trucks 3 9 9 9 8 8 8 8 8 8 7

Major Equipment Purchases

Jumbo - 2 boom 4 - - - - 2 1 - - - -

Longhole dril l - 5 - - - 3 2 - - - -

Bolter 3 - - - - 1 1 - - - -

Emulsion Truck 3 1 - - - 2 1 - - - -

Scissor Lift 4 - - - - 2 1 - - - -

LHD - Ejector - 7 m3 3 2 - - - 3 2 - - - -

Trucks 3 6 - - - 4 4 - - - -

Support Equipment Purchases

Low profile U/G Motor Grader 1 - - - - 1 - - - - -

Crane truck 1 - - - - 1 - - - - -

Fuel/lube vehicle 2 - - - - 2 - - - - -

Service truck w/ scissor l ift 2 - - - - 2 - - - - -

Forklift/Cable reeler 1 - - - - 1 - - - - -

Mancarrier - 16 person 2 - - - 2 - - - 2 - -

Shotcrete Machine 2 - - - 2 - - - 2 - -

Mechanics truck w/ flat deck 2 - - - 2 - - - 2 - -

Supply truck w/ flat deck 2 - - - 2 - - - 2 - -

Crew cab 4 - - - 4 - - - 4 - -

Mine rescue van 1 - - - 1 - - - 1 - -

Surface Equipment Required

Grader 2 2 2 2 2 2 2 2 2 2 2

Water Truck 1 1 1 1 1 1 1 1 1 1 1

Compactor 1 1 1 1 1 1 1 1 1 1 1

Dozer 3 4 4 4 3 3 3 3 3 3 3

Excavator 1 2 2 2 2 2 2 2 2 2 1

Haul Truck 5 13 13 13 13 13 13 13 13 13 10

Surface Equipment Purchases

Grader 2 - - - - - 1 - - - -

Water Truck 1 - - - - - - - - - -

Compactor 1 - - - - - - - - - -

Dozer 3 1 - - - - 2 1 - - -

Excavator 1 1 - - - - 1 - - - -

Haul Truck 5 8 - - - - - - - - -






Table 16-8: Mining Personnel Requirements

Mine Operations Personnel Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Operators for Major Equipment

Jumbo - 2 boom 14 17 15 13 11 9 8 7 6 5 8

Longhole drill 0 14 14 14 15 15 15 15 16 16 13

Bolter 21 26 22 19 16 14 12 11 9 8 10

Emulsion Truck 9 13 11 10 9 8 8 7 6 6 7

Scissor Lift 14 17 15 13 11 9 8 7 6 5 8

LHD - Ejector - 7 m3 9 20 19 18 17 16 16 15 15 14 14

Trucks 9 34 32 31 31 30 30 30 29 29 26

Operators for Support Equipment

Low profile U/G Motor Grader 5 5 5 5 5 5 5 5 5 5 5

Service truck w/ scissor lift 9 9 9 9 9 9 9 9 9 9 9

Forklift/Cable reeler 3 3 3 3 3 3 3 3 3 3 3

Operators for Surface Equipment

Grader 9 10 9 9 9 9 9 9 9 9 9

Water Truck 5 5 5 5 5 5 5 5 5 5 5

Compactor 3 6 3 3 3 3 3 3 3 3 3

Dozer 14 23 23 18 14 14 14 14 14 14 14

Excavator 3 6 6 6 6 6 6 6 6 6 3

Haul Truck 23 59 59 59 59 59 59 59 59 59 50

Maintenance

Crane truck 3 3 3 3 3 3 3 3 3 3 3

Fuel/lube vehicle 5 9 9 9 9 9 9 9 9 9 7

Mechanics truck w/ flat deck 10 18 18 18 18 15 15 15 15 15 10

Supply truck w/ flat deck 10 18 18 18 18 15 15 15 15 15 10

Maintenance Shop 13 20 14 9 5 8 7 5 3 2 6

Mine Services

Services installations 5 5 5 5 5 5 5 5 5 5 3

Grouting/shotcrete 5 5 5 5 5 5 5 5 5 5 3

Construction 5 5 5 5 5 5 5 5 5 5 3

Level maintenance 2 2 2 2 2 2 2 2 2 2 2

Nipper 2 2 2 2 2 2 2 2 2 2 2

General Mine Administration, Technical and Services

Mine Administration 19 19 19 19 19 19 19 19 19 19 15

Maintenance 11 11 11 11 11 11 11 11 11 11 8

Technical Services 23 23 23 23 23 23 23 23 23 23 20

Subtotal Mine Operations 263 407 384 364 348 336 331 324 317 312 279






The geotechnical investigation indicates that an extraction of 67.5% can be

achieved, resulting in an overall mining recovery of 58%. It was assumed that an

additional 2% of waste material would be added as external dilution and 2% of

mineralization losses were incurred due to operating conditions. An overall mining

recovery factor of 58% of the estimated Mineral Resources was considered in this

study; this accounts for 25.0 Mt of mineralized material that is a subset of the

Mineral Resources as run-of-mine (ROM) production to be processed inclusive of

diluting material.

For the preliminary mine plan, considering waste inside stopes and external

dilution, the overall subset Mineral Resource estimate grades were diluted to

185 ppm of Ta2O5 and 1,591 ppm of Nb2O5 for the ROM estimates.

Production was estimated at 2.7 Mt/a of mineralized material to be extracted over

10 years; the first year was considered as preproduction, leaving nine years of

full-scale production.

Developments were modeled following a decreasing activity level from the

beginning to the end of the life of mine. A total of 92,500 m of development was

estimated.

The deposit will be accessed through two main portals, Upper and Lower, and a

series of adits from the portals. Top access crosscuts will be driven from the main

ramps to each level on vertical intervals between 20 to 30 m. Stope access

crosscuts will be driven at level from west to east. Bottom access crosscuts will be

driven to function as mucking drifts. Underground mine services such as

ventilation and air heating, compressed air, water for drilling and power supply will

be provided to the mine via the adits.

Radio remote-controlled load-haul-dump units (LHDs) will be used to extract the

mineralized material from the stope beyond the safety of the brow. The

mineralized material from stopes will be loaded directly to the haulage trucks that

will be spotted at the end of the crosscut. The trucks will drive down the ramps

and will exit the mine at the Lower Portal. The trucks will deliver the mined

material to a surface stockpile at the primary crusher close to the portal. Crushed

material will be transferred to the process plant by a belt conveyor.

Waste from development will be initially utilized for construction of a structural shell

of the tailings co-disposal site on surface, which is located in an area east of the

processing plant site. This material will be transferred to the plant site in batches

on the conveyor system.

Several areas of investigation are required to support a detailed mine plan. These

include underground geotechnical and geo-hydrological conditions for mine design,

the possible use of mining methods utilizing un-cemented backfill, and continuous






handling systems for mineralized material and waste rock. Opportunities include

possible better ground conditions than assumed. With improved ground conditions

the size of stopes and production drifts could be increased.

There is opportunity to increase the net present value (NPV) of the Project by

mining higher-grade zones early in the mine life providing that a practical mining

sequence can be implemented and the overall recovery of the Mineral Resources

is not negatively affected.

There is opportunity to increase the net present value (NPV) of the Project by

optimising the mine layout to minimize development costs.






17.0 RECOVERY METHODS

17.1 Introduction

This section on recovery methods incorporates assumptions, analysis and findings of

the Preliminary Economic Assessment that has an effective date of 29 September

2011.

The information relevant to the plant design supporting the financial analysis prepared

during the 2011 PEA is included in this section and has not been updated because

AMEC considers that the assumptions supporting the outcomes remain reasonable.

The effective date of the 2011 PEA results remains 29 September 2011.

17.2 Plant Design

The design for the process facilities considered a nominal processing capacity of

7,500 t/d. Where data were not available at the time of flowsheet development, AMEC

developed criteria for sizing and selection of equipment based on comparable industry

applications, benchmarking, and the use of modern modelling and simulation

techniques.

The mineral processing and the refining are based on conventional technology and

industry-proven equipment.

Run-of-mine (ROM) mineralized material from the underground will be crushed and

conveyed to the concentrator where the mineralization will be ground to liberate the

mineral values from the host rock and then separated by flotation. The bulk

tantalum-niobium concentrate produced will be filtered, dried, and introduced into the

refining plant. There the concentrate will undergo a thermal reduction which will

remove most of the gangue material and create a smaller, higher purity material for

chlorine processing. The distillation of the anhydrous metal chloride products will

produce separated high-purity Nb and Ta chlorides.

Tantalum chloride is the precursor to capacitor grade Ta powder, so would be

marketed in this form. Niobium chloride can be sold as a chemical precursor. Both Ta

and Nb chloride products can be readily converted and marketed as high-purity

technical grade Ta2O5 and Nb2O5 oxides respectively. The simplified flowsheet is

shown in Figure 17-1.






17.3 Comminution (Crushing, Storage, and Grinding)

The primary crushing station will be a fixed jaw crusher. Mine haul trucks will dump

ROM mineralized material into the ROM surface stockpile located close to the Lower

Portal. Mineralized material will be fed to the crusher using an apron feeder. Crushed

mineralized material will fall onto a conveyor and be fed to a fine crushing circuit at the

plant site which will further reduce the material to -8 mm. The material will be stored in

a fine ore silo. It will be withdrawn by feeder into a rod mill. The discharge from the

rod mill will flow into the cyclone feed pumpbox. The cyclone feed pump will transfer

the material to the cyclone circuit which will produce finished product in the overflow (a

P80 size of 100 µm). The cyclone underflow will report to a ball mill for additional

grinding. The discharge from the ball mill will join the rod mill discharge as feed to the

cyclone pumpbox.

Figure 17-1: Concentration and Refining of Blue River Mineralization

Note: Figure generated by AMEC, 2011






17.4 De-Sliming and Flotation

In both cases, the pyrrhotite and carbonate concentrates will join the de-sliming fines

in the tailings filtration system.

After the rejection flotation work, the carbonate tails will be processed through a

magnetic separator to recover any magnetite to tailings. The water will be exchanged

at this point to allow a higher level of control in the subsequent pyrochlore flotation to

recover tantalum and niobium.

Flotation of Nb-Ta-bearing minerals to a mineral concentrate will occur at a pH of 7.0

employing a pH modifier/promoter (fluosilicic acid), a collector (a Duomac-T

equivalent) and a frother (MIBC) as required. The pyrochlore tails will pass to the

tailings filtration system. The pyrochlore rougher concentrate will be reground and

cleaned in five stages with the same reagents. The mass of material will be reduced

substantially, to less than 1% of the feed into the plant. All cleaner tails will be sent

directly to the tailings filtration system.

17.5 Filtration

After de-sliming, magnetic separation, and flotation, the combined tailings will be

pumped to two separate tailings thickeners for water recovery. After thickening, the

material will be pumped into one of four tailings pressure filters. These filters will

reduce the moisture to a level for disposal to dry stacked tails.

Should a paste backfill option be considered during future studies, material after

filtration could be sent to a paste silo feed thickener. In this latter case, the material

could be withdrawn for use as required and transported to the underground mine portal

where it would be mixed with cement prior to use underground as backfill.

The pyrochlore concentrate product will be a much smaller mass and will first be sent

to a small concentrate thickener and filtered.

17.6 Concentrate Pre-Treatment

There are two options to pre-treat the flotation concentrate. If the concentrate grade is

between 10% and 30% Ta and Nb, then it is possible to perform a pre-leach with

strong acid to dissolve some of the gangue after which the material would be sent for

filtration. If the concentrate is more than 30% the non-acidified material would be sent

directly to filtration. After filtration, the concentrate would be dried and sent to the

concentrate receiving bin.






Material would be recovered from the concentrate bin to a blender where flux (calcium

fluoride, quicklime), iron oxide, scrap aluminum and fuse mix would be added. After

blending the material is charged to burn pits. The aluminothermic reduction occurs in

these pits. After smelting and solidification, the alloy ingot (with adhering slag) is

removed from the burn pit by crane. After further cooling, the alloy ingot and adhering

slag would be broken, crushed and separated through the use of magnetic separation.

The slag material would be disposed of to tailings while the alloy material containing

the tantalum and niobium would be charged to the chlorination system.

17.7 Chlorination and Distillation

The material, which is charged into the chlorination system, will be set into fixed

charge pots where chlorine will be added and the mixture will be heated to 350°C. The

mixed chloride product from chlorination of the ferroalloy smelting product will then be

distilled to achieve high purity Ta and Nb chlorides. These distillation products will be

captured by separate condensers.

17.8 Product / Materials Handling

A conveyor is planned to transport materials from the portal to the plant.

17.9 Energy, Water, and Process Materials Requirements

Power for the proposed operation will be sourced from B.C. Hydro. The most

appropriate source will be investigated during more detailed Project studies.


In the opinion of the QP, the metallurgical programs completed on the Blue River

Project have met their objective of identifying a processing method allowing for the

extraction of tantalum and niobium mineralization that has reasonable prospects of

being economic. Additional work is required to confirm the extractive metallurgy of the

concentrate, produce the target flotation concentrate grade and examine the response

of the process to variability of mineralization within the deposit and process conditions.

The following interpretations apply to the plant design and metallurgical testwork

results:










Ta and 69% recovery for Nb in the flotation stage. The refining process will have a

97% recovery for both Ta and Nb.

Optimization of the supply and pricing of reagents for the refining process may

support lower operating cost assumptions.

Metallurgical testing has not yet attempted to demonstrate that a 30% combined

oxidized concentrate grade as a feed for the refining stage is achievable.








18.0 PROJECT INFRASTRUCTURE

18.1 Introduction

This project infrastructure section incorporates assumptions, analysis and findings of

the Preliminary Economic Assessment that has an effective date of 29 September

2011.

The information relevant to the project infrastructure supporting the preliminary mine

plan prepared during the 2011 PEA is included in this section and has not been

updated because AMEC considers that the assumptions supporting the outcomes

remain reasonable. The effective date of the 2011 PEA results remains 29 September

2011.

18.2 Site Layout

The overall Project site layout plan is included as Figure 18-1. The planned Upper and

Lower Portals will be located about 4 km from the plant site. At the front of the Upper

Portal (service portal) sufficient space will be provided to accommodate the required

facilities for operation.

18.3 Buildings

18.3.1 Mine Service Building

The plant service building will be a multi-purpose complex in a two story building east

of the process building. The first floor (18 m by 40 m) will be a maintenance bay for

minor repairs and maintenance of the mobile equipment and will have an office for the

maintenance foreman, a small parts area, a tool crib, and a storage area for safety

equipment.

Part of the first floor will contain the men’s and women’s dry. The second floor will be

the administration offices. The complex will be connected to the process building by a

covered walkway.






Figure 18-1: Proposed Site Layout Plan

Note: Figure prepared by AMEC, 2011






18.3.2 Truck Shop

A 24 m by 36 m truck shop on the southwest end of the site will be operated by a

qualified contractor. The required equipment and tools for regular maintenance and

possible repair of the haul trucks are assumed to be supplied by the contractor.

18.3.3 Warehouse

To save cost and construction time, the warehouse will be a 24 m by 50 m

Coverall-type fabric building. One third or more of the building will be dedicated to

warehouse cold storage, and the remainder allocated to fire water and potable water

tanks and a fire pump skid. The building will be equipped with an interior liner as an

insulation layer to minimize the potential of freezing inside the warehouse. This will

save the cost of insulation and heat tracing of the tanks, pipes and all equipment and

is easier for operation and maintenance.

Forklifts, pallet racking, bins, and carousels will be provided for handling materials.

Flammable products such as solvents and paints will be stored separately.

The Coverall-type building life expectancy is 15 years and it will withstand design snow

and wind loads.

18.3.4 Process Building

The process building will be a 22 m by 52 m steel structure building sitting on a

concrete foundation. The mill foundation will be on bed rock; a geotechnical

investigation is required prior to determination of the final location of the mill.

A 15 m diameter tailing thickener will be located to the south of the process building

with a walkway/pipe rack connection to the building.

18.3.5 Crushing and Screening Circuit

The preliminary site layout is designed to take advantage of the topography to

minimize earthworks. The secondary crusher will be located in the northeast corner of

the site at a higher elevation.

The conveyor span from the secondary crusher discharge will rise at 12° (an access

safety constraint) to reach the height of the screen above the fine ore bin.

Screen oversize will pass via a return conveyor to the tertiary crusher.






18.3.6 Portal Infrastructure

A yard will be constructed in front of the Upper Portal, which will accommodate an

electrical substation and generator set, and office building, provision for storage,

ventilation infrastructure, and heater, air compressor, diesel storage, a first aid rescue

vehicle bay and a water tank.

The buildings and related facilities will be pre-engineered as much as possible.

Drinking water will be provided from the plant site water treatment plant by containers.

A portable wash room will be provided for workers. Sewage will be trucked to the site

wastewater treatment plant.

18.3.7 Explosives Storage

Ammonium nitrate, blended emulsion, and explosives will be delivered to site on

demand by contractors. A small storage magazine will be constructed at a distance of

about 200 m from the Upper Portal. Room for explosive storage will be provided by

excavating into the rock. The walls and roof will be reinforced and a lockable door will

be provided, as per the requirements of the Quantity-Distance Principles User’s

Manual published by the Explosives Regulatory Division of NRCan. The magazine will

hold boosters, delays, detonating cords, detonating caps, and other explosive

accessories.

18.3.8 Aggregate Crushing and Concrete Batch Plants

A crushing and stockpiling facility will be required during construction to provide

crushed product for roads and surfacing. The mobile plant assembly will include a jaw

crusher, screening plant, closed-circuit secondary crushing unit, and washing plant.

Concrete supply from nearby towns is assumed adequate for construction purposes.

An on-site concrete batch plant is not proposed for the project. Availability of existing

concrete supply should be examined in the next phase of study.

18.4 Roads and Logistics

18.4.1 Access Road

The road access design includes a short new road with a 7.2 m wide gravel surface

from the existing road to plant site about 80 m in length and a 1.5 km new service road

from the existing road to the Upper Portal and upgrades to the current access road.






An existing 80 m-long bridge crossing over the Thompson River has a limited load

capacity and might not qualify for crossing heavy loads during construction or

long-term use during the life of the mine. Therefore a new bridge has been included in

capital cost estimate. Using the existing railway for shipment should be investigated in

next phase of study.

18.4.2 Haul Road

Dual-lane traffic requires a travel width (16.2 m) of not less than three times the width

of the widest haulage vehicle used on the road (assumed to be trucks of the size of

CAT 775F with 5.4 m of overall width).

Single-lane traffic requires a travel width (11 m) of not less than two times the width of

the widest haulage vehicle used on the road.

Shoulder barriers should be at least three-quarters of the height of the largest tire on

any vehicle hauling on the road wherever a drop-off greater than 3 m exists. The

shoulder barriers are designed at 1.5:1 (H:V). The width of the barrier is excluded

from the travel width.

There is one main haul road, from the site to the waste rock stock pile, that will have a

length of about 8 km (refer to Figure 18-1). The road from the upper portal to the plant

will be a service road for personnel and supplies.

18.5 Co-Disposal Storage Facilities

The PEA design for tailings and waste management is to construct a co-disposal

drystack facility.

18.5.1 Drystack Considerations

Filtered tailings stacks are often referred to as “drystacks” and that nomenclature is

used in this chapter. However, these facilities are not “dry” per se as the tailings, while

placed in an unsaturated state, do have moisture contents that are typically 70% to

85% of saturation. Generally the tailings are filtered to within a few percent of optimum

standard Proctor moisture content, which is typically on the order of 15% (wt water :

wt solids). The main distinguishing feature from other tailings deposits is that filtered

tailings stacks are a solid rather than the more typical slurry and need to be

transported using mechanical means versus hydraulic methods.






Tailings drystacks are particularly suited to locations with flat or gently sloping storage

sites and arid environments where it is relatively easy to place and compact filtered

tailings into a stable geometry. However, unlike conventional slurried tailings, tailings

drystacks can also be placed on relatively steep terrain in a similar manner to

conventional waste rock facilities. Placement practices and drystack design can be

adapted to different environments and drystacks have been successfully constructed

in cold climates.

The Blue River site is expected to have about 1.5 m of precipitation annually (KCB

2009a). The site experiences a cold winter and steep topography which can present

challenges for construction of a tailings drystack. AMEC notes, however, there are

operating drystacks in even wetter environments.

The general drystack concept for the Blue River site is to have an outer shell zone with

a general tailings placement area located upstream of the shell. The shell zone would

consist of well-compacted tailings placed only when it can be assured that such

compaction can be achieved. The waste rock could be placed inter-layered or mixed

with the tailings in the shell or could be used as armour on the face of the shell to

prevent erosion of the tailings surface. The outer shell would support a general tailings

placement area where tailings could be placed in poor weather and with less ability to

achieve compaction during winter months. The same operating attempts at

compaction would be made in winter/wet weather conditions but it is more difficult to

get assured densities. As a consequence, having a general placement area for such

materials where structural integrity of the overall stack is not at jeopardy if lower

densities are achieved is sound tailings management. It would also be possible to

store waste rock in the general placement area where it could be encapsulated within

the tailings. In order to effectively co-dispose of waste rock in the general fill area,

waste rock and tailings disposal would have to be appropriately timed and managed.

Typically, slopes of tailings drystacks are designed to be 3H:1V or less to minimize

erosion of the tailings on the downstream slope. This is particularly a consideration for

closure.

Steeper slopes will be required at the Blue River site due to topographical constraints.

The steeper slopes are achievable but require a potentially wider shell zone of good

compaction, confirmed foundation conditions for the shell and the slopes will require

erosion protection.

18.5.2 Evaluation of Potential Sites

A series of tailings storage location screening assessments were carried out during

2008, 2009, and 2010 (KCB 2009a, 2009b; AMEC 2010a, 2010b, 2010c), focusing






mainly on conventional tailings storage. Some evaluation of potential waste rock

storage sites and a tailings drystack facility was undertaken in the 2009 studies.

The initial 2009 study (KCB 2009a) focused on the area in the immediate vicinity of the

deposits and was based on a desk study using available topography and other

information. Five conventional storage sites, using containment dams, were assessed.

The potential tailings drystack locations identified by KCB in 2009 are located on valley

sidehills and in flatter areas adjacent to the North Thompson River The tailings storage

facilities (TSFs) were assumed to be constructed of local borrow with slopes of

2.5H:1V and a settled density of 1.3 t/m3 was assumed for the tailings.

The second study carried out in 2009 (KCB 2009b) reviewed a number of alternative

sites for TSF storage away from the proposed mine site. The main focus of the study

was an industrial land parcel near Valemount, B.C., and consideration was given to

construction of a 40 Mt TSF, a plant, 30 kt of rock storage, and a rail-siding facility on

the property. Four other areas around Valemount were also considered as TSF

alternatives.

None of the options were considered ideal and additional studies were carried out over

an increasingly large area and with varying constraints to identify potential tailings

storage areas (AMEC 2010a, b, and c). Generally topographic and climatic conditions

are challenging for surface storage of tailings and no ideal location has been identified.

Given the current stage of project development and the constraints imposed, the input

provided should be considered as conceptual. Tailings storage site locations

determined by Klohn Crippen Berger (KCB) in 2009 were visually identified by AMEC

staff during a reconnaissance flight in July 2010 and were visited on foot by AMEC

personnel.

However, the proposed tailings storage sites have not been visited by geotechnical

personnel, there have been no site investigations and there have been no specific

technical analyses in support of facility layout. The suggestions presented in the PEA

for tailings storage were based on engineering judgment, AMEC’s experience with

filtered tailings, and standard industry practice for similar facilities.

18.5.3 Site Selection

Additional review for the PEA indicated that a site identified by KCB in 2009, termed

WSF3 and shown in Figure 18-1 as the co-disposal site, could be utilized, since the

volume of tailings storage had been revised downward from about 30 Mt as

conceptualized in 2009 to 22 Mt in the PEA. The facility was moved uphill slightly






relative to the 2009 study to take advantage of locally flatter areas and reduce the

footprint of the facility in steeper areas.

Additional optimization of the facility layout could be carried out in consideration of

local topography and stability; however the volumes and areas would not be expected

to be significantly altered.

18.5.4 Facility Design

The facility was laid out with 2H:1V slopes. Attempts were made to use flatter slopes;

however, due to the topography, they were not feasible. If the project is advanced to a

later stage, the slopes could potentially be locally flattened in specific areas of flatter

topography. The facility will have the following dimensions:

Total storage volume: 20.9 Mm3

Total footprint area: 534,000 m2

Total surface area of drystack: 559,000 m2

Volume of sloped (shell) portion of facility (shell portion of facility): 13.2 Mm3

Footprint surface area under slope (shell): 385,000 m2

Surface area of sloped (shell) portion of tailings: 416,000 m2.

Conceptually, the sloped portion of the drystack would form a structural shell,

supporting the general tailings placement area behind it. The shell zone would consist

of compacted filtered tailings possibly interlayered or mixed with waste rock. Stringent

placement control and compaction would be required in the shell. It is possible that

the limits of the shell zone could be optimized at a later stage. The general tailings

placement zone would allow for placement in wet and cold weather when there would

be less assured compaction. Operating practices for the general tailings placement

zone will be identical to those required in the shell area.

Depending on the timing of any mine start-up and the tailings production schedule, it

may be necessary to construct a starter berm from non-tailings material to provide

sufficient storage for the first winter of tailings placement. The starter berm could be

constructed of non acid generating waste rock, general rockfill or granular overburden

obtained from a local borrow source.

Based on the volumetric proportions of the shell and general placement areas and

assuming a six-month period available for shell construction, there may be a deficit

with respect to the amount of tailings and waste rock available to construct the shell






zone at certain points during operation. It is likely that the configuration of the facility

could be optimized during design to overcome this deficit. If supplemental fill material

were required in the shell, it could consist of general rock fill or granular overburden.

The option of storing some of the tailings produced during the winter for placement in

the drystack during the summer construction season could also be considered.

18.5.5 Co-Disposal Facility Geohazards Considerations

The area in the vicinity of the deposits has steep-sided valleys with glaciers located in

the upper portions of many of the catchments and visible avalanche tracks. The

WSF3 site is located on a side-hill near the mouth of the valley, away from the glaciers

located farther up the valley. It is also located on a spur minimizing the risk of

avalanches. A site-specific geohazards evaluation for the site has not been

completed, however, and should be carried out as the Project is advanced.

18.5.6 Co-Disposal Facility Stability Considerations

The 2H:1V slope required because the topography at the Blue River site is steep

compared to typical tailings drystacks. However, in such situations, flatter slopes are

often developed to minimize operating and closure erosion concerns. From a purely

structural stability perspective, provided the shell zone of appropriate size can be

developed while meeting compaction criteria, the use of 2H:1V slopes is acceptable.

Additionally, the general fill zone behind the shell is intended to allow placement of

tailings during poor weather and will potentially have lower strength and likely areas of

wetter tailings. The shell zone is therefore necessary to support the facility and will

require significant compaction and placement control to provide sufficient resistance.

The foundation is expected to consist of colluvium and/or silty to sandy tills which

should not be a concern with regard to overall stability, although this will require

confirmation as the project is advanced. A toe berm or shear key may be required for

stability of the drystack depending on specific foundation conditions. This should be

assessed in future design stages.

The Project is located in a moderately seismic area (KCB 2009a) and stability under

seismic loading will have to be considered during design of the facility. Because of

their unsaturated nature, the general placement tailings are considered unlikely to

liquefy and the shell tailings will not present a concern because the compacted nature

of the downstream shell will result in dilative behaviour under shear further improving

liquefaction resistance. Rain, ice, and snow may potentially create isolated areas in

the general placement area that are saturated and poorly compacted, and therefore

locally susceptible to liquefaction. It is highly unlikely that the entire zone would liquefy






and over time the pore water in any saturated zones would seep out of the stack,

leaving it unsaturated and resistant to liquefaction in the long term. Although

liquefaction is not considered a concern, deformations due to increased loading during

seismic events will occur and should be quantified during detailed design phases.

For the Project base case it is assumed that the tailings facility does not require lining.

A requirement to line the facility would involve clearing and stripping the entire footprint

at start up, placing and compacting a bedding layer over the entire footprint, installing

the liner system over the entire footprint, and placing a protective cover layer over the

entire liner.

18.5.7 Co-Disposal Facility Surface Water Run-Off Considerations

Two surface water management systems will likely be required for the tailings

drystack.

The first system would divert non-contact (clean) water around the facility.

The second would collect run-off water which had been in contact with the tailings

drystack area.

Contact run-off water would have to be collected and potentially treated prior to

release, whereas the clean water would be diverted around the facility and into the

creeks downstream of the tailings facility.

The non-contact water diversion system would consist of a ditch located beyond the

final footprint of the drystack and would be a long-term structure. The contact water

collection system would consist of a ditch around the perimeter of the drystack

footprint and would be reconstructed annually, so as to be located slightly ahead of the

advancing drystack footprint. The perimeter ditch would require sediment-control

structures within it to help contain any tailings mobilized from the drystack. The

perimeter ditch would direct the contact run-off water to a collection and sedimentation

pond where eroded tailings could settle out and water treatment could be carried out if

required.

Located as it is on a side-slope and near the nose of the ridge, there is little catchment

area uphill of the drystack facility and it may be feasible to combine the two water

management systems, particularly upon closure. Combining the two systems would

increase the volume of water collected and potentially requiring treatment, however,

the increased water volume may also contribute to dilution of the contact water.






18.5.8 Co-Disposal Facility Closure Considerations

At mine closure, the surface of the drystack would be sloped/contoured for drainage,

covered with an appropriate material and vegetated to enhance erosion protection.

The shell face will require armouring to reduce erosion. The perimeter ditches will be

increased in size and lined with rip-rap following mine closure to minimize the amount

of run-off water that comes into contact with the stack, further reducing the potential for

erosion. Depending on the chemistry and flows, it may be possible to combine the

water diversion ditch with the closure perimeter ditch.

The Project base case is that the uranium and thorium levels are sufficiently low to not

be a concern. If the levels were found to be an issue, however, it would affect closure

requirements. In addition to the potential difficulty of treating any seepage, if radon

gas emission was above regulatory levels, the cover system would have to be

designed to contain radon gas. Specific details of the cover would be based on

applicable regulations but it may well require that the cover last for a very long period

of time necessitating rock armouring among other considerations.

18.6 Avalanche Hazard

No allowance for avalanche protection for any infrastructure has been considered at

this stage, but considering the steep slope and possible heavy snow at the area an

avalanche study is recommended during more detailed studies.

18.7 Water Supply, Distribution, and Treatment Systems

The potable water system and layout for the process and administrative area is

designed to service buildings and a workforce of 120 persons. Potable water for the

mining section will be constructed as part of the portal and underground infrastructure.

Raw water will be provided from a well and a prefabricated water treatment module will

treat water to standard drinking water requirements. Treated water will be stored in a

potable water tank.

Fire protection water for the construction camp and later for all buildings will be

provided by a prefabricated diesel-driven fire and electric jockey pump on a skid that

will be located in the coverall building.

Both the potable and the dedicated fire water tanks will be located under the coverall

building adjacent to the fire pump skid and the water treatment module.






Potable and fire water will be distributed to the plant buildings through separate pipes.

All water mains will be buried to a depth in excess of 3 m, or will be insulated providing

an equivalent degree of cover to prevent freezing.

18.8 Waste Considerations

Waste-water treatment sludge will be trucked away to a nearby municipal facility or

approved landfill.

Waste lubrication and hydraulic oils from vehicle maintenance will be stored in

dedicated tanks and sent to a recycling facility offsite. Their disposal will be contracted

to an approved contractor.

A modular sewage treatment system will be installed as part of the initial construction

infrastructure. A small package treatment plant will provide treatment to the domestic

sewage at the site. Effluent from this plant will meet specified water discharge

guidelines prior to discharge into the environment.

Buried gravity sewer lines and manholes will collect and direct sewage from the

service building and truck shop to an equalization tank adjacent to the wastewater

treatment module for treatment.

18.9 Accommodation

Contractors and employees will commute from the nearby towns, such as Blue River

and Valemount, during construction.

No on-site permanent accommodation will be provided for personnel. It is assumed

that the workforce, including management staff, will reside in the nearby communities

and will commute, via buses, on a daily basis. For safety reasons, no private vehicles

will be permitted on the site access road or at the site.

18.10 Power and Electrical

Power supplies in the region have been assumed to be sufficient for Project

requirements, and no allocation for additional power line construction has been

included.

The BC Hydro 136,000 volt supply line for the North Thompson valley passes through

the west side of the property adjacent to the rail line. The 20 megawatt Bone Creek

run-of-river hydroelectricity project, owned by Transalta Corp., was commissioned in






June 2011, and is adjacent to the Project area its powerhouse is located approximately

4.4 km south of the Lower Portal.

The mill will be the greatest consumer of power. The high voltage line from the grid

would go to a main substation close to the mill. From this sub-station, lower voltage

power will be distributed to the mill, offices, maintenance shop, and other

infrastructure, and to the proposed mine substation. The PEA assumes power will be

supplied to the mine via the portals. The sub-station for the main power distribution

system and the air compressors will be installed in facilities located adjacent to the

Upper Portal.

18.11 Fuel

Fuel will be delivered to the mine site using tanker trucks. The fuel storage tanks will

be single-walled within a lined containment berm. Tank design will comply with the

appropriate regulatory requirements.



The project infrastructure supporting the preliminary mine plan prepared for the

2011 PEA remains appropriate and current for the conceptual mining method,

mineral processing method, treatment plant and planned throughput rate.

Infrastructure envisaged includes a plant, plant service building, truckshop, potable

and process water systems, a sewerage system, co-disposal site, underground

mining operation, conveyor system, and various haul and access roads. The

planned Upper and Lower Portals will be located about 4 km from the plant site.

The co-disposal facility will be about 8 km from the plant site.

Facilities to support mine operations will require construction.

No on-site permanent accommodation will be provided for personnel. It is

assumed that the workforce, including management staff, will reside in the nearby

communities and will commute, via buses, on a daily basis.

Geohazards are present in the area, and will require careful consideration in future

studies.

Water management studies, in particular for the co-disposal site, will be required.






19.0 MARKET STUDIES AND CONTRACTS

19.1 Introduction

This section includes a summary from the 29 September 2011 Preliminary Economic

Assessment (the PEA) by AMEC (Chong et al., 2011). AMEC has reviewed recent

publicly available information for Ta metal prices and Nb metal prices as at May 2012

and found that the Ta and Nb prices used for both the 22 June 2012 Mineral Resource

update and the 2011 PEA to remain as reasonable assumptions, which are US$317/kg

tantalum metal and US$46/kg niobium metal.

19.2 2011 PEA Market Studies

For the 2011 PEA, Commerce prepared assessments of the tantalum and niobium

markets which outlined their supply and demand. The tantalum assessment was

prepared by a tantalum market expert employed by Commerce who is not

independent. His analysis reflected the general consensus of other analysts regarding

the tantalum market expressed in publicly available information.

The 2011 PEA niobium assessment was prepared by an independent niobium expert

and also reflected the general consensus of analysts in publicly-available information

for the niobium market.


for expression of interests from potential buyers of the proposed Blue River products

and has not negotiated any purchases or off-take agreements.

19.3 2011 PEA Commodity Price

19.3.1 Tantalum

Tantalum is commonly quoted in two separate forms:

Ta2O5 in tantalite concentrate: a non-refined, tantalum-bearing concentrate of

variable composition and trace element content; and

Tantalum metal scrap (99.9% pure Ta): this form of tantalum product receives a

premium price in the market relative to tantalite concentrate.

Over the five years from 2005 to 2010 tantalite concentrate prices ranged from

US$75/kg contained Ta2O5 to US$100/kg contained Ta2O5 (US$35/lb to US$45/lb). In






the same period tantalum metal scrap prices ranged from US$110/kg Ta to US$180/kg

Ta metal (US$50/lb to US$80/lb).

In 2010, prices rose dramatically in response to changing market conditions including

reduced production, increased concerns about conflict-tantalum production in Africa,

depletion of known strategic stockpiles, and curtailed exports from China. In mid-

October 2010 the price for Ta2O5 in tantalite concentrate was US$195/kg and for

tantalum metal scrap was US$280/kg.

The higher price for tantalum metal scrap compared to the price for Ta2O5 in

concentrate is considered a proxy to the added value Commerce should recognize by

refining the Blue River concentrate to high purity Ta2O5.

In AMEC’s opinion, the base case price for tantalum (US$317/kg) is reasonable for

constraining mineral resources based on recent market conditions, but notes it is

significantly higher than historical prices. There is a risk that using current price

assumptions may not reflect the long term price of Ta and Nb, particularly in the

present volatile market conditions.

19.3.2 Niobium

Niobium generally trades as Nb metal, or ferroalloy, and the price has remained

relatively constant at US$44.08/kg Nb metal (US$20/lb Nb) over the last several years.

A base case price of US$46/kg Nb metal was assumed.

19.4 Price Assumption Discussion

Review of recent publicly available information on Ta and Nb prices by AMEC notes

that the market prices have not changed significantly enough to warrant altering the

2011 PEA price assumptions (Ta US$317/kg and Nb US$46/kg) for the current mineral

resource update (Figure 19-1 and Figure 19-2).

AMEC believes the market studies prepared for the 2011 PEA provide a reasonable

basis for the long-term Ta and Nb prices used in the 2011 PEA and that the

assumption that there will be a market for future mine production is also reasonable.






Figure 19-1: Ta Price Trend

Note: Table and Ta price trend from www.metalprices.com on 22 May 2012. US$92/lb Tantalite is

equivalent to US$317/kg Tantalum.

2011 PEA Price $US 92 /lb Tantalite

http://www.metalprices.com/






Figure 19-2: Nb Price Trend

Note: Table and Nb price trend from www.metalprices.com on 22 May 2012. US$21/lb Nb is equivalent to

$US 46/kg Nb.


In the opinion of the QPs, the following conclusions can be drawn from the marketing

strategy used to support the PEA:

Commerce has prepared assessments of the tantalum and niobium markets which

outline the supply and demand for tantalum and niobium.

As the Project is still at an early evaluation stage, Commerce has not initiated

requests for expression of interests from potential buyers of the proposed Blue

River products and has not negotiated any purchase or off-take agreements.

The price assumptions from the 2011 PEA are still reasonable and are suitable to

use in the current (22 June 2012) mineral resource estimate.

PEA 2011 $US 21/lb Nb

http://www.metalprices.com/






20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR

COMMUNITY IMPACT

20.1 Environmental Assessment for Mining Projects

The Blue River Project will require approval under the Federal and Provincial

environmental assessment (EA) processes prior to applying for the necessary permits

and authorizations for construction and mine operation. This section discusses the

environmental assessment and permitting process as it stands today, and describes

the principal licences and permits which would be required for the Blue River Project.

The British Columbia Environmental Assessment Office (BCEAO) and the Canadian

Environmental Assessment Agency (CEAA) would both conduct an environmental

review of the Upper Fir Project, as defined respectively under the B.C. Environmental

Management Act and the Canadian Environmental Assessment Act.

Overall the environmental review of a project is a process that could take up to

24 months to complete. The process would include the development of several

important documents by Commerce, including the Project Description, Assessment

Information Requirements and an Environmental Impact Assessment application,

followed by the review of these documents by the public, interested stakeholders, First

Nations and regulators.

Both the Provincial and Federal processes have defined timelines for project review

though these timelines are not yet currently harmonized and past attempts to do so

have not been overly successful. Though this is under review, enabling regulations

have not yet been passed, and under present legislation, the Federal timeline is likely

to be the longer of the two at up to two years to review the Environmental Assessment

application and make a decision, while the provincial application review stage is 180

days plus 45 days for decision. Additionally, the Federal clock is stopped each time

CEAA submits comments to the proponent for review, to respond to, or revise; this can

possibly extend the time line.

There is also a need for additional time on the front end of the process to develop the

Project Description. This document must first be accepted by the regulators, and then

will be used to develop Assessment Information Requirements or Terms of Reference

which must be put out for public review prior to acceptance. These Assessment

Information Requirements will define the content required for the Environmental

Assessment application. The Federal government allows 90 days for this, while the

Provincial government has no timeline on this process.






The environmental review and assessment process results in a decision with respect

to whether or not the Project should be issued an Environmental Assessment

Certificate by the Provincial government, as well as receive Federal Ministerial

approval based on the recommendations put forth to the Minister of Environment in a

Comprehensive Study Report prepared by the Major Projects Management Office.

Both are required for a project to proceed to permitting and development.

20.2 Project Studies

Environmental monitoring, baseline studies and site investigations have been ongoing

at the Blue River Project site since the summer season of 2006 with the selection of

local and regional studies areas for each biophysical discipline.

Field studies completed by specialist consultants independent of Commerce

Resources include:

Site hydrology (2006–present);

Snow course depths (2007, 2008);

Fisheries and aquatics (2006–2008);

Soils, flora and fauna assessments (2006, 2007), including studies of rare,

threatened and endangered plants (2007), breeding birds (2007) and terrestrial

ecosystem mapping (2006, 2007), wildlife studies and habitat suitability mapping

(2006–2008);

Geochemistry, mineralized material and waste rock characterization with baseline

ABA and metals analyses (2007, 2008);

Surface water and sediment quality (2006–present);

Groundwater (2007–2009); and

Terrain stability assessment for roads (2007–present).

Kinetic test work for ARD/ML was initiated in June 2010 and is ongoing; results of this

work will give an indication of the type of management strategies required for handling

PAG waste rock.

Additional environmental baseline programs are expected to continue, as required

through 2012.

Monitoring of meteorology, air quality, hydrology, and water quality will continue

throughout the construction, operation, closure and post-closure phases.






It is anticipated that with this work and the results of the PEA in hand, Commerce will

have the necessary information to start development of a draft Project Description

which is a prerequisite to entering the Environmental Assessment process. These

data will provide a strong base to initiate meetings with the BCEAO and CEAA, as well

as with key provincial regulators such as Ministries of Energy and Mines, Forests

Lands and Natural Resource Operations, and Environment, to discuss specific project

requirements under the Provincial and Federal environmental assessment processes.

Summaries will be prepared of baseline data collected and work plans as appropriate

for submission, review and input by Federal and Provincial regulators.

20.3 Environmental Setting and Review of Environmental Baseline

Characterization of existing environmental conditions, which began in 2006, is an

important component of the risk management and permitting process for the Blue

River Project.

The Blue River Project area is located within the B.C. Ministry of Environment

Thompson-Nicola Region (Region 3), what was the Ministry of Forests and Range

Headwaters Forest District and Fisheries and Oceans Canada Sub-district 29J

(Clearwater). It falls towards the northern end of the Kamloops Land Resource

Management Plan (LRMP) which was approved by the province in 1995. This LRMP

is the first plan of its kind in British Columbia in that it is a locally-developed plan that is

designed to guide land and resource management decisions in a way to balance

community needs, environmental concerns and economic values. This LRMP is

termed a sub-regional integrated land use plan in that it establishes the framework for

land use and resource management objectives and strategies. The Plan requires that

more detailed operational plans which are subsequently developed be consistent with

the management strategies and objectives defined in the LRMP.

Following implementation of the B.C. Mountain Caribou Recovery Strategy, site-

specific objectives and strategies for caribou management were developed (first in

2006, then updated in 2009) which included objectives pertaining to mineral

exploration. The Blue River Project area falls within the Wells Gray-Thompson caribou

planning unit (unit 4A). In this case, the relevant portion of the LRMP was repealed

and replaced under a Government Actions Regulation (GAR) by a

caribou-management strategy, specifically the identification of Ungulate Winter Range

(UWR) zones. The GAR order states that exploration and mine development activities

within the UWR are considered by the government to be an acceptable risk to caribou,

and are allowed to proceed without requiring an exemption from the Ministry of

Forests, Lands and Natural Resource Operations.






The bulk of the infrastructure proposed for the Blue River Project falls within an area

identified as a Modified Harvest Zone. In such zones, operational activity is expected

to be considerate of caribou habitat and disturbance in caribou areas. Commerce is

committed to seeking specialized professional advice to minimize or eliminate

disturbances to caribou using best management or other practices. Wildlife studies to

date suggest that caribou are recorded only rarely in the area of the proposed Project.

The larger Blue River Project area lies on the eastern side of the south-flowing North

Thompson River where claims held encompass portions of the Bone, Gum,

Moonbeam, Paradise Lake, Pyramid, and Serpentine Creek watersheds. The overall

relief is moderate to steep, with an average elevation of 1,625 m, a maximum elevation

of 3,225 m, and a minimum elevation of 580 m. Some small glaciers exist in the

easternmost part of the Project area, and moderate to steep forested slopes rise

above the North Thompson River valley. The North Thompson River drainage

continues south to join the South Thompson River at Kamloops, B.C.

Proposed Project infrastructure is located along a western-facing slope immediately

above a gravelly part of the North Thompson River valley, and includes portions of the

Bone and Gum Creek watersheds as well as residual areas draining directly into the

North Thompson River. The tree line is located at approximately 2,000 m elevation,

and the Upper Fir deposit centre is located at 1,180 m elevation along a network of

previously-constructed logging and skid roads. Much of the area of the proposed

Project infrastructure, including the Upper Fir deposit had been logged prior to the

commencement of mineral exploration. Naturally-occurring outcrop is generally poor,

with limited exposure of underlying country rocks along road cuts and locally in

streams.

The area has a continental climate which is subject to frequent modification by

maritime air masses from the Pacific Ocean. The area is part of a "wet belt" which

occupies part of eastern British Columbia. Heavy snow falls occur almost every

winter, in which temperatures stay close to the freezing point when maritime air

dominates. The most severe cold spells may send thermometer readings below -

40°C/F. Rain is frequent in other seasons. Summer days are warm or occasionally

hot, with thunderstorms often spawning over the nearby mountains. The optimal field

exploration season is mid-June through mid-September.

Streams within the Project Area are generally characterized by a snowmelt-dominated

peak rising in April or May and peaking sometime between June and July. Rain-on-

snow events occasionally occur in this region and these can enhance both winter flows

and spring peaks. In addition, late fall rainstorms are common, recharging soil

moisture heading into winter and producing short-duration peak flows. Low flows

occur generally from the end of November to March, and in hot summer months, with






the lowest flows commonly occurring in January or February. No wetlands have been

identified within areas of proposed Project infrastructure.

Surficial materials were typed for texture, drainage, moisture and nutrient regime, and

parent material. This not only supports ecosystem classification, but also provides

baseline soils data for eventual environmental impact assessment and reclamation

planning. Materials range from colluvial veneers to fluvial plains, glaciofluvial terraces,

rock, and morainal blankets. On the Upper Fir slope, morainal materials are most

common, while fluvial and glaciofluvial deposits are limited to lower elevations near the

valley bottom. Silts and sands are the most prevalent soil textures, although mixed

fragments, rubble, and gravel also occur. Overall the deposits are relatively shallow,

although blankets, which have more than 1 m of surficial materials (e.g., glacial till),

are much more common than veneers, which have less than 1 m of surficial materials.

Soil moisture and nutrient regime are generally average over the majority of the site

and have formed in place within morainal and glaciofluvial landforms.

Brunisols are the most dominant soil order found, while podzols are secondary,

becoming more prominent in areas with increasing rainfall and elevations above

1,500 m. Soil quality and quantity appear to be adequate to support soil salvage and

reclamation activities in the area of the proposed development. Soil fertility suggests

normal levels of soil nutrients as compared to other mine sites in B.C. Soils are

moderately acidic (i.e., pH 4-5.5) and considered normal for mesic, conifer-dominated,

forested vegetation in similar areas. Soils are very rapidly to imperfectly-drained with

soil moisture regimes ranging from sub-xeric to hygric and soil nutrient regimes limited

to moderate and rich.

Terrestrial Ecosystem Mapping (TEM) was completed to describe terrestrial

ecosystems according to the bioterrain base and standards established by British

Columbia’s Resource Information Standards Committee. The Blue River Project falls

within the Cariboo Mountain Ecosection of the Northern Columbia Mountains. Two

biogeoclimatic zones are found in the project study area. These are the Interior Cedar

Hemlock (ICH) zone, which occurs at lower elevations, and the Engelmann Spruce

Subalpine Fir (ESSF) zone which occurs at higher elevations above the ICH zone.

Biogeoclimatic zones, subzones and variants within the Study Area were classified

using the Ministry of Forests Biogeoclimatic Ecosystem system. The following

subzones/variants are present within the larger project study area:

Wells Gray Wet Cool Interior Cedar – Hemlock Variant (ICHwk1)

Mica Very Wet Cool Interior Cedar – Hemlock Variant (ICHvk1)

Northern Monashee Wet Cold Engelmann Spruce – Subalpine Fir Variant

(ESSFwc2)






Wet Cold Engelmann Spruce –Subalpine Fir Woodland Subzone (ESSFwcw)

Wet Cold Engelmann Spruce – Subalpine Fir Parkland Subzone.

Much of the area is forested, although avalanche chutes punctuate the landscape at

seemingly regular intervals and forest harvesting has been extensive at lower to mid

elevations.

In terms of vegetation, the area directly covered by the proposed Blue River Project

infrastructure is relatively small and generally characterized by common plant

communities associated with five biogeoclimatic subzones (ICHwk1, ICHvk1,

ESSFwc2, ESSFwcw and ESSFwcp). Based on an assessment of biogeoclimatic

units in the Study Area and the B.C. Conservation Data Centre species at risk list for

the Headwaters Forest District (B.C. Conservation Data Centre 2006), at least 37

ranked plants may occur within the larger study area. However, a field study of rare

vascular plants identified only four populations of two Provincially-listed rare plants

(Galium trifidum ssp. trifidum and Carex paysonis) occurring outside the current project

envelope. No Federally-listed plant species or plant communities were identified.

The region encompassing the proposed Project infrastructure is likely home to many

terrestrial wildlife species including black and grizzly bears, deer, moose and mountain

goats; birds are likely to include osprey, eagle, woodpecker and raven, migratory

songbirds, raptors; and numerous small mammals.

Wildlife species of concern, whose confirmed distribution intersects that of the Blue

River Project area, include the blue-listed grizzly bear (Ursus arctos) and red-listed

mountain caribou (Rangifer tarandus caribou). Mountain caribou presence on the

larger area of the mining claims making up the property and in the general area has

been confirmed through ongoing government radio-telemetry studies. Mountain

caribou are a Federally- and Provincially-listed species and are of considerable

concern to the public. They have the greatest potential to interact with the Project

property in early winter.

Due to deep snow pack in the region, and considerable management and local

interest, other species of concern include mountain goat (Oreamnos americanus) and

moose (Alces alces). Mountain goat winter range exists throughout the area, with

occupied ranges within 2 km of the area of potential Project development. Moose are

the most heavily-hunted ungulate in the area. Moose winter range occurs throughout

the North Thompson valley, with valley wetlands and early seral stage habitats of

prime importance.

Habitat suitability was assessed for wildlife focal species selected based on their

at-risk status under the British Columbia Conservation Data Centre (CDC) and the






Committee on the Status of Endangered Wildlife in Canada (COSEWIC) systems, and

their level of local concern. Species include mountain caribou (Rangifer tarandus

tarandus; southern populations), moose (Alces alces), grizzly bear (Ursus arctos),

mountain goat (Oreamnos americanus), and marten (Martes americana). Specific

habitat notes were also recorded for other species such as deer (Odocoileus sp.), elk

(Cervus canadensis), wolves (Canis lupis), and black bear (Ursus americanus).

Habitat ratings suggest moderate and high suitability for mountain caribou during the

early winter season in the ICH. Moderate ratings were also assigned to marten in

many sites in the ICH. No sites were rated as high or moderate suitability for mountain

goat in the immediate areas of potential Project infrastructure.

Regional and site specific fisheries studies show that bull trout and mountain whitefish

are utilizing lower Gum Creek for rearing. Bull trout consisted of both juveniles and

young-of-the-year suggesting that it is being used for spawning by this species. The

lower reach of Bone Creek is being utilized by coho salmon, parr, and torrent sculpin.

Benthic invertebrate data were also collected. Habitat available for fish within Gum

and Bone creeks is limited to their lowermost reaches, near their mouths. Gum Creek

fish habitat use is limited to the lowermost portion of the creek, from the mouth to

600 m upstream before a falls/gradient barrier (>20%) and fish distribution in Bone

Creek is limited to the section from its confluence with the North Thompson River to

approximately 2,100 m upstream before an impassable water fall.

Water quality studies were conducted at various sites within the Project area from

2006 to the present, with the objective of providing a long-term record of the relative

chemical stability of the project area. Samples were analyzed for physical variables,

anions, nutrients, total organic carbon, and total and dissolved metals. Data for each

site were compared to the Canadian Council of Ministers of the Environment (CCME)

and B.C. water quality guidelines (BCWQG).

Total suspended solids (TSS) and turbidity values tend to be the highest at sample

sites in the North Thompson River and Bone Creek, the latter the result of small scale

debris flows upstream caused by larger rain events. Metals that exceeded the

applicable aquatic life protection guidelines included total and dissolved aluminum,

total and dissolved cadmium, total chromium, total cobalt, total and dissolved copper,

total and dissolved iron, and total lead, total manganese, total selenium, total silver,

total thallium and total zinc. Of these, dissolved cadmium, total cobalt, dissolved

copper, total and dissolved iron, total manganese, total selenium, total silver, total

thallium and total zinc did not exceed the applicable guidelines in most years. In most

years, concentrations of total chromium and total copper tend to be naturally elevated

in the North Thompson River and Bone Creek compared to other sites. As the Project

is at the exploration stage, these values reflect natural background values. While

elevated metal concentrations were noted in the sediment samples from selected






streams, chromium was the only metal to exceed CCME Sediment Guidelines in one

sample collected from Bone Creek.

Samples from surface rocks and drill core were collected and tested as part of the

Upper Fir acid rock drainage/metal leaching (ARD/ML) characterization program.

Laboratory tests included static acid-base accounting, total inorganic carbon and a

standard multi-element ICP suite on material solids.

A subset of samples were submitted for additional testing including solids trace

element/rare earth element (REE) chemistry, short-term leach extraction tests, acid

buffering characteristic curves (ABCC) and mineralogical analysis including

petrography and Rietveld XRD. All carbonatite samples tested were classified as non-

potentially acid generating (non-PAG). Paste pHs for nearly all host rock samples

were near-neutral to alkaline indicating currently available buffering capacity in the

samples at the time of testing.

Most country rock in the Upper Fir deposit was characterized by generally low to

moderate sulphide content (<1% sulphur) and low to moderate neutralization potential

predominantly provided by slower reacting silicate minerals. However, a minor

proportion of country rock (~10% of samples) was associated with elevated sulphide

content (>1% sulphur). These rock units showed a range in acid generation potential

classifications, and in particular, a significant proportion of gneiss (~52% of gneiss

samples) was considered potentially acid generating (PAG). The majority of

amphibolite (~85%) and pegmatite (~65%) samples were classified as non-acid

generating (non-PAG), with a minor proportion classifying as PAG, typically associated

with higher sulphide samples. Fenite material was considered to be non-PAG.

Based on this initial characterization program, though some proportion of waste rock

appears to be PAG, it would appear that the Blue River Mineral Resource has an

overall low potential for acid rock drainage/metal leaching (ARD/ML) generation,

especially if waste segregation strategies can be incorporated into proposed mining

methods. Kinetic test work on two composite samples remains ongoing and results

will be incorporated into planning for additional sampling as well as modeling of PAG.

No work has yet been completed on the ARD/ML potential of tailings.

Hydrogeologic investigations show that groundwater elevation in bedrock roughly

mimics topography in the Project area, so flow in the vicinity of the Upper Fir Deposit is

generally from east to west. Groundwater depth in boreholes was observed to range

from 15 m to 130 m in the area of the deposit, and is near the surface at lower

elevations. Most groundwater flow through bedrock occurs through fractures, and the

bulk hydraulic conductivity of bedrock was estimated to range from 10-8 m/s to






10-6 m/s. Due to fracture-control, flow through bedrock is likely complex, and may not

be well connected to the surface.

Of the six groundwater samples collected in 2009, the British Columbia Ministry of the

Environment Water Quality Criteria for Freshwater Aquatic Life (BCMOE) and the

Canadian Water Quality Guidelines for the Protection of Aquatic Life (CCREM) criteria

for aquatic life were exceeded in five samples for fluoride, one sample for aluminum,

three samples for chromium, one sample for copper, and three samples for zinc. No

other parameters exceeded these criteria. However, as the boreholes used in this and

past hydrogeology studies of the Project site are deep exploration holes, and were not

developed or purged prior to sampling, sample chemistry may not be accurately

characterizing existing groundwater conditions.

Current air quality at the site is considered excellent with limited influence from road

traffic and forestry activities. No site specific data has been collected to date.

An initial review of environmental conditions and planned project features indicates

that proactive design and mitigation can be successful in addressing environmental

impacts associated with developing, constructing, operating and closing the proposed

Blue River Project. As with other projects in the many B.C. mines located in

mountainous terrains, water management will be a key issue.

20.4 Closure Considerations

Commerce has engaged in progressive reclamation activities during exploration since

geological work and drilling began to focus on the area of, and around, the Upper Fir

resource.

Conceptual closure planning for the Blue River Project involves staged reclamation

and closure over the life of the exploration, development, construction and operation of

the mine. This will include appropriate contouring and revegetation of any waste

dumps, the Upper and Lower Portals, the drystack tailings storage facility, closure of

exploration roads, trails and platforms, closure of mine roads and removal of all mine

facilities, as well as post-closure management and monitoring plans for a defined

period of time.

It is expected that the initial design of the drystack tailings storage facility, water

storage, diversion and water management structures, waste rock dumps, and other

mine and plant facilities will be integrated into closure designs for each component as

well as for the mine as a whole.






20.5 Current Environmental Liabilities

Current environmental liabilities are believed to be restricted to exploration drilling

programs. Existing disturbances due to exploration include drill pads, trails and

access ways, which are remediated in an ongoing program of progressive closure

once Commerce establishes exploration has been completed in a particular area.

Under the existing exploration permit, a reclamation bond is in place which will cover

the cost of any outstanding reclamation from these activities.

20.6 2011 PEA Closure Plan

For the purposes of the PEA, a closure estimate of CAD$10 million was incorporated

in the financial analysis. The figure was obtained by benchmarking to similar-size

mines with the same level of complexity.

20.7 Permitting

Following environmental assessment approval, permits needed for construction and

mine operations can be issued. In B.C. there is an option to apply for concurrent

permitting. This allows a review for permit applications to be processed at the same

time as the environmental assessment is being conducted, resulting in the permits

required for construction being issued shortly after a positive environmental review

decision.

Table 20-1 and Table 20-2 provide a listing of possible federal and provincial permits

that will be required for construction, mine operations, closure and post-closure.

This listing cannot be considered comprehensive due to the complexity of government

regulatory processes, which evolve over time, and the large number of minor permits,

licences, approvals, consents, and authorizations, and potential amendments that will

be required throughout the life of the mine. The permit requirements will be reviewed

and updated as the Project advances.






Table 20-1: Provincial Permits, Approvals, Licences, and Authorizations

Provincial Permits Description ACT

Notice of Work Approval for exploration and site programs

to be conducted to gather geological and

other site information

Mines Act

Mines Act permit Approval to construct, operate and reclaim

mine and its infrastructure

Mines Act

Mining Lease Land occupancy for mine (sub-surface

rights)

Mineral Tenure Act

Surface Lease Surface land occupancy for mine and site

infrastructure

Land Act

Licence of Occupation Land occupancy for other features (e.g.

borrow pits)

Land Act

Statutory Right of Way Land occupancy for linear features Land Act

Waste Discharge Permit – Water Approval to discharge mine effluent and

sewage into the environment

Environmental

Management Act

Waste Discharge Permit – Air Approval to discharge air emissions into the

environment

Environmental

Management Act

Occupant Licence to Cut Approval to remove timber (mine,

infrastructure, borrow areas)

Forest Act

Road Use permits Approval to use existing forestry roads Forest and Range

Practices Act

Special Use permit Approval to construct new roads Forest Practices Code

of B.C.

Water Licence Approval to construct, maintain and

decommission water works

Water Act

Section 9 Approval Approval for changes in and about a stream Water Act

Section 8 Approval Approval for short term use of surface water Water Act

Authorization for Public Highway Use Approval to use public highways Transportation Act

Exemption Permit Approval to haul concentrate (if required) Transportation Act

Construction Permit To construct a potable water system Drinking Water

Protection Act

Table 20-2: Federal Permits, Approval, Licences, and Authorizations

Federal Permits Description ACT

Navigable Waters Approval Approval to build bridges across streams Navigable Waters Protection Act

Section 35(2) Authorization Allows harmful alteration or disruption of

fish habitat (HADD) (e.g. bridge upgrade)

Fisheries Act

Explosives Magazine Licence Approval to store explosives Explosives Act

Radio Licences Approval to operate radios Radio Communications Act






20.8 Considerations of Social and Community Impacts

Socioeconomic and cultural heritage studies have not yet been initiated for the Blue

River Project. Basic community profiling has been completed of the individual

communities nearby as well as relevant regional government and planning

organizations. This work shows that as with other areas of rural B.C., there is a large

dependence on primary resource industries, and overall the population of the area is in

decline.

The Blue River Project is located in the North Thompson River valley within the

Thompson-Nicola Regional District (TNRD). TNRD functions as a partnership of

11 member municipalities (Ashcroft, Barriere, Cache Creek, Chase, Clearwater,

Clinton, Kamloops, Logan Lake, Lytton, Merritt and Sun Peaks) as well as 10 electoral

areas whose voices at the Board table are representative of many small

unincorporated communities, member municipalities and electoral areas.

The area has a population of over 122,286 (2006 census) and a total area of

45,279 km2. The Regional District is active in providing over 115 services including

planning and building inspection, emergency preparedness and 911 services,

recreation, utilities, TV rebroadcasting, river buoys, transit, tourism, economic

development as well as environmental health services, which include waste reduction,

mosquito and weed control.

The closest town to the Project is the small community of Blue River located about

20 km south of the project. Blue River is an unincorporated village, located at the

confluence of the Blue and North Thompson Rivers along the Yellowhead Highway

about halfway between Kamloops, B.C. and Jasper, Alberta. It currently has a

declining population of about 260 residents, with a local economy supported by

logging, tourism, and transportation industries. Accommodation is available for

exploration crews by way of hotels and rental housing. Commerce maintains an active

presence in the town with a field office open during the exploration season.

The Project is about 90 km south of the village of Valemount, B.C., a rural community

of about 1,150 situated between the Rocky, Monashee, and Cariboo Mountains. It is

the nearest community to the west of Jasper National Park, and is also the nearest

community to Mount Robson Provincial Park.

Outdoor recreation is popular in summer and winter; hiking, skiing, snowmobiling, and

horseback riding are common activities. Economic activities include logging, railway,

transport and tourism. Valemount is considered a fully-serviced village, boasting high-

speed wireless internet, train, bus and highway service. The town serves as a supply

centre for another 700 people who live in the Regional District of Fraser-Fort George,






from Albreda to Small River. Today Valemount’s economy is based on logging and a

rapidly-growing tourism industry.

20.8.1 First Nations

The Blue River Project lies on lands which comprise part of the traditional territory of

the Simpcw First Nation. Simpcw First Nation is a member of Secwepemc (Shuswap)

Nation Tribal Council (SNTC), a political organization, which works on matters of

common concern, including the development of self-government and the settlement of

the aboriginal land title questions. The SNTC is involved with natural resource

management within the Secwepemc Nations territory and the creation of economic

development opportunities for Secwepemc communities. Commerce is very aware of

its responsibility to appropriately engage local and regional First Nations early in the

planning and development stages of the project.

On behalf of Commerce, members of the Simpcw First Nation completed

Archaeological Overview Assessments (AOA) over all areas of proposed disturbance

related to Commerce exploration activities, as well as over key areas of potential

project infrastructure. No concerns were noted by the archaeology field technicians

and exploration activities were approved to proceed by the Simpcw archaeologist with

no further recommendations for work necessary in the areas surveyed.

Traditional Knowledge/Traditional Use (TK/TU) studies, as well as a detailed

archaeological impact assessment will need to be undertaken and will also involve

Simpcw First Nation participation. Such studies may identify areas and seasons

where Simpcw have engaged in traditional activities such as hunting, fishing, gathering

and spiritual ceremonies, and the outcomes will be used to inform the overall design

and operation of the Project.

First Nations engagement, with respect to exploration activities, began in May 2007,

and will be continuing for the duration of the project. Engagement activities have

included presentations and discussions with Chief and Council and Sustainable

Resources Department staff, one-on-one meetings and a site visit by elders.

On 25 October 2010, Simpcw First Nation and Commerce signed a confidential

Exploration Agreement with respect to exploration activities on the Blue River project,

which formalized a process for ongoing discussion regarding all exploration activities,

recognizes the traditional cultural, heritage, and environmental interest of the Simpcw,

and ensures that benefits from the project are realized by Simpcw First Nation.

Commerce has also committed to involve the Simpcw in environmental plans to gain

from their knowledge of the region, as well as to keep them informed of project goals.






20.8.2 Local Communities

The Blue River Project is only at the exploration and early economic evaluation stage;

however, to introduce Commerce and its Project to its communities, Commerce has

hosted one community meeting in each of Blue River and Valemount, and has made

presentations to the Valemount Council. The Valemount Mayor and Council have also

toured the property and continue to receive regular updates on the project. Periodic

community newsletters provide updates on the Blue River Project; these are

distributed in the town of Blue River and are readily available on Commerce’s website.

In the summer of 2010, Commerce hosted a community barbeque in Blue River to

thank its neighbours and the local people for their assistance and support over the

past exploration seasons. As the project moves forward, open houses/information

sessions and meetings will take place in other local communities such as Barriere,

Clearwater and Chu Chua.

Public engagement to date has included meetings with local councils (e.g., Valemount,

Barriere) and informal discussions with local land-owners.



The Blue River Project will require approval under the Federal and Provincial

environmental assessment (EA) processes prior to receiving the necessary permits

and authorizations for construction and mine operation.

Overall the environmental review of a project is a process that will take at least

18 months to complete. The process would include the development of several

important documents by Commerce, including the Project Description, Assessment

Information Requirements and an Environmental Impact Assessment application,

followed by the review of these documents by the public, interested stakeholders,

First Nations and regulators.

The environmental review and assessment process results in a decision with

respect to whether or not the Project should be issued an Environmental

Assessment Certificate by the provincial government, as well as receive federal

Ministerial approval based on the recommendations put forth to the Minister of

Environment in a Comprehensive Study Report prepared by the Major Projects

Management Office. Both are required for a project to proceed to permitting and

development.

Environmental monitoring, baseline studies and site investigations have been

ongoing at the Blue River Project site since the summer season of 2006 with the






selection of local and regional studies areas for each biophysical discipline. Field

studies completed by specialist consultants independent of Commerce include: site

hydrology (2006–present); snow course depths (2007, 2008); fisheries and

aquatics (2006–2008); soils, flora and fauna assessments (2006, 2007), including

studies of rare, threatened and endangered plants (2007), breeding birds (2007)

and terrestrial ecosystem mapping (2006, 2007); wildlife studies and habitat

suitability mapping (2006–2008); geochemistry, mineralized material and waste

rock characterization with baseline ABA and metals analyses (2007, 2008), surface

water and sediment quality (2006–present), groundwater (2007–2009) and terrain

stability assessment for roads (2007–present).

Kinetic test work for ARD/ML was initiated in June 2010 and remains ongoing;

results of this work will give an indication of the type of management strategies

required for handling PAG waste rock.

Additional environmental baseline programs are expected to continue, as required

through 2012.

A preliminary list of the Federal and Provincial permits required for operation of a

mine has been developed. This listing cannot be considered comprehensive due

to the complexity of government regulatory processes, which evolve over time, and

the large number of minor permits, licences, approvals, consents, and

authorizations, and potential amendments that will be required throughout the life

of the mine. The permit requirements will be reviewed and updated as the Project

advances.

Socioeconomic and cultural heritage studies have not yet been initiated for the

Blue River Project. Basic community profiling has been completed of the individual

communities nearby, as well as relevant regional government and planning

organizations. This work shows that as with other areas of rural B.C., there is a

large dependence on primary resource industries, and overall the population of the

area is in decline.

The Blue River Project lies on lands which comprise part of the traditional territory

of the Simpcw First Nation.

First Nations engagement, with respect to exploration activities, began in May

2007, and will be continuing for the duration of the project. Engagement activities

have included presentations and discussions with Chief and Council and

Sustainable Resources Department staff, one-on-one meetings and a site visit by

elders. On behalf of Commerce, members of the Simpcw First Nation completed

Archaeological Overview Assessments (AOA) over all areas of proposed

disturbance related to Commerce exploration activities, as well as over key areas

of potential project infrastructure. Traditional Knowledge/Traditional Use (TK/TU)






studies, as well as a detailed archaeological impact assessment will need to be

undertaken and will also involve Simpcw First Nation participation.

On 25 October 2010, Simpcw First Nation and Commerce signed an Exploration

Agreement with respect to exploration activities on the Blue River project, which,

amongst other aspects, formalized a process for ongoing discussion regarding all

exploration activities, recognizes the traditional cultural, heritage, and

environmental interest of the Simpcw, and ensures that benefits from the project

are realized by Simpcw First Nation. Commerce has also committed to involve the

Simpcw in environmental plans to gain from their knowledge of the region, as well

as to keep them informed of project goals.

Public engagement to date has included meetings with local councils (e.g.,

Valemount, Barriere) and informal discussions with local land-owners.






21.0 2011 PEA CAPITAL AND OPERATING COSTS


Assessment (the 2011 PEA) by AMEC (Chong et al., 2011).

21.1 2011 PEA Basis of Estimate

All costs in the 2011 PEA were expressed in constant first quarter (Q1) 2011 Canadian

dollars. No allowance had been included for escalation, interest or financing fees,

taxes or duties, or working capital during construction. The level of accuracy for the

estimate was +40 /-20% of estimated final costs, as per the Association of Advanced

Cost Estimators (AACE) Class 5 (scoping level) definition.

The estimate scope is limited to the battery limits of the plant and mine sites with no

allowance for off-site facilities.

The estimate covered the direct field costs of executing the project, plus the Owner’s

indirect costs associated with design, construction, and commissioning. The

preproduction costs were capitalized and included all the expenditures before Year 1

of production.

21.2 2011 PEA Capital Costs

The costs are divided into five areas: (1) infrastructure, (2) material handling,

(3) process, (4) mining, (5) contingency, and (6) indirect.

21.2.1 Infrastructure

Blue River initial direct civil infrastructure capital costs amount to CAD$30 million. This

area covers the infrastructure and facilities required to support the mine/mill operations

including site preparation, civil work, services, roads, explosive facilities and electrical

substation.

The planned surface conveyor system is included as a material handling item, and

does not appear in the civil infrastructure total.

Power supplies in the region have been assumed to be sufficient for Project

requirements, and no allocation for additional power line construction has been

included.






21.2.2 Material Handling

Blue River initial direct material handling capital costs amount to CAD$8 million. This

area covers belt conveyors and transfer stations and is based on in-house AMEC data

and benchmarking against comparable projects. The cost estimate for the primary

crusher, bin and structure is included in the process plant capital estimates.

21.2.3 Process Plant

Blue River initial direct process plant capital costs amount to CAD$116 million. This

area covers all the process equipment and structures from mills to tailing filters, as well

as pre-treatment and refinery facilities.

21.2.4 Mining

Blue River initial direct mining capital costs amount to CAD$89 million. Mining direct

capital costs include pre-production mining, capital development costs, mine mobile

equipment, and mine infrastructure.

Development to be completed prior to the commencement of production at full rate

was classified as pre-production development. This development was assumed to be

undertaken by Owner’s mining crews with unit costs rates as shown in the operating

cost section.

Development associated with semi-permanent excavations, when used for more than

two years, was treated as capital development. Based on preliminary designs, an

estimate of 20% of all development was treated as capital development.

The equipment hours required for each unit of activity and daily service equipment

requirements were estimated. The required equipment operating hours for each

equipment type were aggregated. Assuming typical yearly operating hours for each

type of equipment, AMEC has estimated minimum equipment fleets to forecast capital

expenditure.

AMEC used a database of budget costs for mine equipment. Where necessary, the

budget costs were factored to reflect Q1 2011 costs.

AMEC used a database of costs estimates for mining infrastructure and fixed service

equipment. Where necessary, the budget costs were factored to reflect Q1 2011

costs.






Initial direct mining capital requirements include major equipment for underground

operations (drilling, loading and hauling), support equipment for underground

operations, equipment for surface road maintenance and hauling and handling of

tailings from the process plant to the drystack (co-disposal) area.

Mining capital requirements were allocated to Year -1, but during more detailed

studies, consideration should be given to allocating the capital requirements over more

than one year, as it is likely that payments for equipment will be required prior to the

equipment being delivered to site.

In a similar manner, the development metreage achieved during the pre-production

year should be re-evaluated during more detailed studies, and a formal pre-production

development schedule with achievable monthly development metreage targets should

be developed.

21.2.5 Contingency Costs

Blue River contingency costs amount to CAD$44 million. Contingency accounts for

unforeseen costs within the project scope. Contingency costs were calculated using a

factor of 25% of civil infrastructure, material handling, process plant, and mine

infrastructure direct capital costs. A contingency factor of 5% was applied to the mine

mobile equipment direct capital costs. No contingency was calculated for

pre-production mining and capital development costs. The contingency factors are

considered appropriate for the level of engineering work performed in the preparation

of this Report. Input variables used in calculating the contingency are a result of

information gathered from previous projects and industry standards.

21.2.6 Indirect Costs

The indirect costs of CAD$92 million cover temporary construction facilities and

services, construction equipment, freight, vendor’s representatives, start-up and

commissioning, engineering, procurement and contract management (EPCM), working

capital, warehousing spares, and first fill. Indirect costs were calculated using a factor

of 30% of civil infrastructure, material handling, process plant, and mine infrastructure

direct capital and contingency costs. An indirect costs factor of 5% was applied to the

mine mobile equipment direct capital and contingency costs. No indirect cost was

calculated for pre-production mining and capital development costs.






21.2.7 Sustaining Capital

The sustaining capital costs of Blue River total CAD$116 million. The primary

sustaining capital components of the proposed mine are:

Underground mine development, CAD$34 million

Fleet replacement, CAD$73 million

For underground development, the cost of development of the entire mine life was

estimated, and then factors were applied to distribute this cost over the life-of-mine,

with costs decreasing as time increased. A unit cost in CAD$/m was then applied

against the annual metres.

The second portion of the initial truck purchases in the first production year was

categorized as sustaining capital. Equipment fleet replacement costs were based on

actual requirements as the useful life of each unit was reached. Major mobile

equipment replacements were considered in Years 5–6 of operation, smaller mobile

equipment was considered to be replaced in four-year intervals.

21.2.8 Mine Closure

A total of CAD$10 million was estimated for mine closure and was benchmarked to

similar-size mines with the same level of complexity.

21.2.9 Capital Cost Estimate Summary

The total estimated capital cost to design and build the Blue River tantalum project at

7,500 t/d capacity is CAD$379 million. The estimate is summarized in Table 21-1.






Table 21-1: Summary of Estimated Capital Costs (CAD, 2011 constant dollars)

Item

Total

(CAD$000’s)

(CAD$000’s)

(CAD$000’s)

Project year

1 2

Production year

-2 -1

Capital expenditure

Initial Capital Infrastructure 29,500 10,300 19,200

Process Initial Capital 116,200 40,700 75,600

Mining Initial Capital 89,400

89,400

Material Handling 8,000

8,000

Contingency 43,600 12,800 30,900

Indirect/Owner Costs 92,300 29,600 62,600

Total 379,000 93,400 285,600

Note: Summation discrepancy due to rounding.

21.2.10 2011 PEA Operating Costs

The operating costs for the Blue River project are based on an Owner-operated mining

fleet and process facility and are stated in first quarter 2011 Canadian dollars.

Operating costs over the life-of-mine are estimated at CAD$38.44/t milled.

Operating costs include the three key areas of mining, process, and overall general

and administrative costs. The estimates are based upon the staffing level,

consumables, and expenditures detailed as part of the underground mine plan and

process design.

Average operating costs are listed in Table 21-2.

Table 21-2: Average Life-of-Mine Operating Cost Summary

(CAD, 2011 constant dollars)

Summary of Average Production Costs

LOM Total

(CAD$000’s)

Cost per Tonne

Milled

(CAD$/t)

Mining 528,900 21.16

Process 338,500 13.54

Material Handling 18,500 0.74

G&A 75,000 3.00

Sub-total 960,900 38.44






21.2.11 Capital and Operating Cost Discussion

AMEC assumes the price assumptions, mining, and process recovery methods from

the 2011 PEA remain as reasonable assumptions. AMECs opinion is that the 2011

PEA capital and operating costs for the Project also remain reasonable for this

technical report.


It is the opinion of the QPs that:

The 2011 PEA price assumptions, conceptual mining, and conceptual recovery

methods are considered reasonable for the purposes of the technical report.

The assumptions used for capital and operating costs in the 2011 PEA are

considered reasonable for the purposes of this Technical Report.

Regarding the 2011 PEA results, the following key outcomes are concluded:

The total estimated capital cost to design and build the Blue River Project at an

assumed 7,500 t/d capacity is CAD$379 million

A total of CAD$10 million was included in the capital cost estimate for mine closure

and was benchmarked to similar-size mines with the same level of complexity







22.0 2011 PEA ECONOMIC ANALYSIS


Assessment (the 2011 PEA) by AMEC (Chong et al., 2011). In addition, AMEC

believes the PEA economic analysis assumptions and outcomes remain reasonable.

The results of the PEA economic analyses discussed in this section represent forward-

looking information as defined under Canadian securities law. The results depend on

inputs that are subject to a number of known and unknown risks, uncertainties and

other factors that may cause actual results to differ materially from those presented

here.

Information that is forward-looking includes:

Mineral Resource estimates

Assumed metallurgical recoveries

Assumed commodity prices, exchange rates, and markets for mine production

The proposed mine production plan

Projected recovery rates

Capital costs, operating costs, and schedules

Assumptions that an EA will be approved by Provincial and Federal authorities.

The financial analysis of the 2011 PEA was partly based on Inferred Mineral

Resources that were considered too speculative geologically to have the economic


Reserves, and there is no certainty that the Preliminary Assessment based on these

Mineral Resources will be realized. Approximately 15% of the Mineral Resources that

support the financial model had been classified as Inferred Mineral Resources.

22.1 2011 PEA Valuation Method

The Project is valued using a discounted cash flow (DCF) analysis, assuming all equity

financing (no debt). Cash flows are assumed to occur at the end of each period. Cash

inflows consisted of annual revenue projections for the mine. Cash outflows such as

capital and operating costs were subtracted from the inflows to arrive at the annual

cash flow projections.

The resulting net annual cash flows were discounted back to the date of valuation the

beginning of 2011 and totalled to determine the net present value (NPV) at the

selected discount rates. Constant 2011 dollars were used for the entire DCF model.






All NPVs in this report are on a pre-tax basis. The project’s internal rate of return

(IRR) was calculated as the discount rate that yields a zero NPV. The simple payback

period was calculated as the time needed to recover the initial capital spent from the

start of production

22.2 2011 PEA Financial Model Parameters

22.2.1 Mineral Resources and Mine Life

The model includes 36,349 kt of Indicated Mineral Resources as well as 6,385 kt of

Inferred Mineral Resources. For this study AMEC utilized average grades of

mineralized materials throughout the mine life at 195 ppm for Ta and 1,700 ppm for

Nb. The diluted grades as the result of the proposed mining method were assumed at

185 ppm Ta and 1,591 ppm Nb. After applying mine recovery and dilution factors, the

financial model assumes that the mine life is 10 years, assuming the plant will process

25 Mt at a 7.5 kt/d plant throughput rate (2.7 Mt/a).

22.2.2 Metallurgical Process

Recovery assumptions from the process plant include 65% recovery for Ta and 69%

recovery for Nb in the flotation stage. The refining process has an estimated 97%

recovery for both Ta and Nb.

22.2.3 Commodity Prices and Foreign Exchange

Publicly-available tantalum and niobium pricing information is very limited as the

markets tend to be based on long-term relationships between few buyers and sellers.

Slightly more information is available for niobium than for tantalum.

A tantalum price of US$317/kg of contained metal in the oxide product is supported by

the prices for tantalum reported on subscription news services.

The niobium price was set at US$46/kg of contained metal in the oxide product over

the life-of-mine.

An exchange rate of US$0.95 to CAD$1.00 is used for all years of the financial model.

22.2.4 Taxes

The discounted cash flow model is pre-tax. Publicly-available taxation information

suggests that the following taxes could be levied.






Taxation considerations comprise Provincial and Federal corporate income taxes and

BC Mineral taxes. The following discussion outlines the main Federal and Provincial

taxation and considerations for mining ventures in B.C.:

Federal taxes: Includes income tax, customs duties, fuel taxes, payroll taxes and

transaction taxes. The general rate of Federal income tax on active business

income earned by a corporation for 2011 is 16.5% and is legislated to decrease to

15% starting in 2012.

Provincial income tax: The general rate of BC Provincial income tax on active

business income earned by a corporation in the Province is 10%.

Provincial mineral taxes: The BC Mineral Tax provides for the Crown's financial

share of mineral production in two ways. The primary way is to receive 13% of a

producer’s profit that is in excess of a normal return on investment over the life of a

mine. This is referred to as Net Revenue Tax. To minimize any disincentive to

investment, the Province does not receive this share until the producer’s

investment and a reasonable return on it have been recovered. The second way is

to receive 2% of operating cash flow from production in each year. This is referred

to as Net Current Proceeds Tax. It is intended to provide compensation for

depletion of the resource when production yields less than a reasonable profit for

the producer. So that only one or the other share is paid, Net Current Proceeds

Tax is fully creditable against Net Revenue Tax.

Depreciation/Salvage Value

No depreciation is incorporated in the model.

Financing

The project is assumed to be 100% Owner-financed.

Capital Costs

The total estimated capital cost to design and build the Blue River Project at 7,500 t/d

capacity is CAD$379 million.

Operating Costs







Working Capital

No working capital is incorporated in the model. Working capital allowances required

to operate this Project are not expected to have a significant impact on the cash flow of

the mine.

Inflation

No inflation adjustments are incorporated in the model. Capital and operating costs

are based on first quarter 2011 Canadian dollars.

Royalty

The Project is not subject to any royalties.

22.2.5 PEA Financial Results

The Project Base Case (8% discount rate) returns an NPV of CAD$18.5 million and an

IRR of 9.1% before tax, and a 6.3 year payback period. Table 22-1 summarizes the

NPV for the Project at a range of discount rates, with the base case highlighted.

Table 22-1: Summary Financial Analysis at Various Discount Rates

Summary of Cash Flow

Pre-tax

Cumulative net cash flow

Undiscounted CAD$000 236,631

Net present value




Discounted at 8% (Project Base Case) CAD$000 18,487


Discounted at 10% CAD$000 (13,514)

Internal rate of return % 9.1

Payback period Years 6.3

Note: base case is highlighted. Exchange rate is US$0.95 to CAD$1.00.

22.2.6 2011 PEA Cash Costs

The cash cost value represents the cost incurred to produce 1 kg of primary product

after deducting the revenue from sales of secondary products. Since the price

analysis for the report was performed around Ta price variation, Ta was chosen as the






main product and Nb was treated as the secondary product for the assessment of

cash cost.

Cash costs are derived through the following formulae:

Production costs = (Mining + Process + G&A + Material Handling)

Cash cost = (Production costs - Revenue of Nb sales) ÷ tantalum production in kg

Using the Brook Hunt convention for reporting C1 cash costs2 , after credit for Niobium

contribution, the tantalum cash cost is calculated to be approximately $24.91/kg

contained in oxide product (equivalent to US$23.66/kg contained in oxide product with

the PEA study exchange rate of US$0.95 to C$1).

The tantalum cash cost was calculated to be approximately CAD$24.91/kg contained

in oxide product (after credit for niobium contribution) as shown in Table 22-2.

Table 22-2: Life of Mine Cash Cost Summary

Section

LOM Total

(CAD$000’s)

Cost per Tonne Milled

(CAD$/t)

Cost per Kg Ta Payable

(CAD$/kg)

Cash costs

Mining 528,900 21.16 220.13

Process 338,500 13.54 140.87

G&A 75,000 3.00 31.21

Material Handling 18,500 0.74 7.71

Sub-total 960,900 38.44 399.92

Credits

Nb (901,100) (36.04) (375.01)

Sub-total (901,100) (36.04) (375.01)

Adjusted cash costs

Total 59,800 2.40 24.91

Note: The figures in this table do not include considerations of working capital or royalty payments

The cash cost for production of tantalum during the earlier years of the proposed

mining operation is $57/kg and decreases over the life of the mine. The major driver

behind the changing costs is the decrease in the mining costs over the life-of-mine. In

the last three years of operation, the revenue generated from niobium exceeds the

2 Brook Hunt, established in 1975, is a global group that specializes in in-depth market analysis across the mining and metals

industries. Brook Hunt has established a method of comparison of costs between projects, countries and commodities that is considered an industry standard. C1 cash costs are defined by Brook Hunt as: the costs of mining, milling and concentrating, on-site administration and general expenses, property and production royalties not related to revenues or profits, metal concentrate treatment charges, and freight and marketing costs less the net value of by-product credits.






total operating costs (mining, processing and G&A). The mining cost for the entire

Project (i.e. mining cost of both tantalum and niobium) drops from an average of $24/t

in the first few years to $18/t in the last year of full production.

22.2.7 2011 PEA Sensitivity Analysis

The sensitivity analysis showed that the project was more sensitive to changes in

operating expenditures than capital expenditures. The project was most sensitive, in

order, to changes in exchange rate, operating expenditure, niobium price, tantalum

price and capital expenditure. Since the sales currency was US dollars and

operational costs were in Canadian dollars, a rising US dollar value versus Canadian

dollar value improved the mine profitability.

The project IRR increased to 14.4% and the NPV increased to CAD$125 million at an

8% discount rate if a Ta price of US$380/kg (20% increase) was assumed.

Sensitivities are summarized in Table 22-3 and Figure 22-1 for the 8% discount base

case rate.

Table 22-3: Sensitivity Summary in CAD, 8% Discount Rate

SENSITIVITY OF NPV @ 8%

Change in Factor

-30% -20% -10% 0% 10% 20% 30%

Facto

r

Exchange rate 448.7 269.5 130.0 18.5 (72.8) (148.8) (213.2)

Capital expenditure 117.9 84.8 51.6 18.5 (14.6) (47.8) (80.9)

Operating expenditure 190.5 133.1 75.8 18.5 (38.8) (96.2) (153.5)

Nb price (140.9) (87.8) (34.6) 18.5 71.6 124.7 177.9

Ta price (123.3) (76.1) (28.8) 18.5 65.8 113.0 160.3






Figure 22-1: Sensitivity Summary in CAD, 8% Discount Rate

22.2.8 Financial Analysis Discussion

It is reasonable to expect that there has been cost escalation since the base of first

quarter 2011 but this has not been quantified for this Technical Report. However, the

sensitivity analysis in this report shows the Project’s sensitivity to the capital and

operating costs.


In the opinion of the QPs:

The PEA financial analysis remains reasonable and current for the price

assumptions, conceptual mining method, mineral processing and recovery factors,

and costs.

Based on the assumptions in this Report, the financial analysis for the Blue River

project, using a discount rate of 8%, returns an NPV of about CAD$18.5 million

and an IRR of 9.1% before tax.

The Project is most sensitive, in order, to changes in exchange rate, operating

costs, niobium price, and less sensitive to changes in tantalum price , and least

sensitive to changes in capital costs.






23.0 ADJACENT PROPERTIES

There are no adjacent properties that are relevant to the Report.






24.0 OTHER RELEVANT DATA AND INFORMATION

AMEC is not aware of any other relevant data or required information for inclusion to

make the report not misleading.






25.0 INTERPRETATION AND CONCLUSIONS

25.1 Mineral Resource Update (Effective Date 22 June 2012)

Commerce has delineated a significant tantalum- and niobium-rich carbonatite deposit

near the town of Blue River in central eastern British Columbia. The company holds a

100% interest in the project.

Commerce has professionally executed an exploration program. The quantity and

quality of the lithological, geotechnical, and collar location, down-hole survey, and

drill-core sample data collected by Commerce in the exploration and delineation drill

programs meet and exceed industry standard practice.

The Blue River Project has very good access and supporting infrastructure.

The Blue River Mineral Resources have the following characteristics:

The mineralization is hosted by a polyfolded carbonatite sill swarm averaging 30 m

thick and 1,100 m long

Close-spaced drilling has confirmed local continuity of the carbonatite

The deposit is amenable to conventional underground mining methods with

estimated mining recoveries that may vary from 65 – 85% depending on the mine

and stope layout and the success in which pillars can be mined on retreat

Tantalum and niobium occur in the minerals pyrochlore and ferrocolumbite and are


recoveries of 65% to 70%

The Mineral Resource estimate is based on information of reasonable quality

There are reasonable prospects for economic extraction

The deposit and Mineral Resource geometry allows for large-scale and selective

mining methods

The Mineral Resources have significantly increased in tonnage mostly due to

reducing the block unit value (BUV) cut-off by eliminating back-fill costs and, to a

lesser extent, additional infill diamond drilling.










The updated Mineral Resources use the same assumptions from the 2011 PEA for the

following items:

Ta and Nb metal prices






It is expected that any future mining operations will be able to be conducted year-

round. High-quality technical grade tantalum and niobium products proposed for

production at-site are suitable for several markets.

Commerce has been pro-active with regard to environmental and socioeconomic

issues. Environmental monitoring, baseline studies and site investigations have been

ongoing at the Blue River Project site since the summer season of 2006. Kinetic test

work for acid rock drainage and metals leaching was initiated in 2010. Additional

environmental baseline programs are expected to continue, as required through 2012.

First Nations engagement, with respect to exploration activities, began in 2007, and

will continue for the duration of the Project. The Blue River Project lies on lands which

comprise part of the traditional territory of the Simpcw First Nation. First Nations

engagement, with respect to exploration activities, began in 2007. Public engagement

to date has included meetings with local councils and informal discussions with local

land-owners.


for expressions of interest from potential buyers of the proposed Blue River products

and has not negotiated any purchase or off-take agreements.

25.2 2011 PEA








Estimated internal rate of return: 9.1% (before tax)

Estimated net present value: CAD$18.5 million at 8% discount rate (before tax)

Estimated payback: 6.3 years



Conceptual project operating cost: CAD$38.44/t milled

Conceptual capital cost: CAD$379 million


several markets

Conceptual mine life: 10 years based upon the Mineral Resources (effective date

20 September 2011)

Most significant conceptual OPEX costs: mining (55%)

Most significant conceptual CAPEX cost: process initial capital (31%)

NPV sensitivity: The Upper Fir deposit is most sensitive to changes in exchange

rate, mining costs, and commodity prices.

The tantalum price assumption used in the 2011 PEA is based on 4th quarter 2010

information. The tantalum price moved significantly higher through 2011. AMEC has

checked the publicly available tantalum and niobium metal prices as at May 2012 and

found the Ta and Nb price assumptions used for both the current Mineral Resource

estimate and the 2011 PEA to remain reasonable.

A higher tantalum price would improve profitability and also increase the mine life.

Additional exploration potential could also provide additional mine life. A two or more

times capital payback is possible.

25.2.1 Opportunities

As a result of engineering work during the 2011 PEA, a lower block unit value cut-off

can be achieved by revising the mine design to eliminate back-fill costs. This

approach was used support the current 22 June 2012 Mineral Resource update, which

in turn has increased the Mineral Resource tonnage at the Project. The increase in

Mineral Resources provides more flexibility for future mining studies and hence

opportunities to improve the Project NPV are as follows:

Optimization of the mine plan by mining higher-grade zones earlier in the mine life,

providing that a practical mining sequence can be implemented and the overall

recovery of the Mineral Resources is not negatively affected

Optimization of the mine layout to minimize development costs






Advanced geotechnical studied to identify and understand ground conditions which

could allow an increase in the size of stopes and production drifts

Optimization of the supply and pricing of reagents for the refining.

25.2.2 Risks

The risk factors are:

The current Mineral Resource estimate is supported by current tantalum and

niobium prices which are higher than historic average prices and may not reflect

long term prices.

Commerce has not initiated requests for expressions of interest from potential

buyers of the proposed Blue River products and has not negotiated any purchase

or off-take agreements.



Testwork to date has not considered factors such as water recycling. A water

treatment plant may be required and may result in increased capital costs.

The 2011 PEA financial analysis is partly based on Inferred Mineral Resources

(effective date 29 September 2011) that are considered too speculative

geologically to have the economic considerations applied to them that would

enable them to be categorized as Mineral Reserves, and there is no certainty that

the Preliminary Economic Assessment based on these Mineral Resources will be

realized.

The Blue River Project will require approval under the federal and provincial

environmental assessment (EA) processes prior to receiving the necessary permits

and authorizations for construction and mine operation. Overall the environmental

review of a project is a process that can take up to 24 months to complete.

Traditional Knowledge/Traditional Use (TK/TU) studies, as well as a detailed

archaeological impact assessment will need to be undertaken.

The Project warrants additional work to examine the opportunities and mitigate the

risks. On completion of this work, Commerce may consider proceeding with a pre-

feasibility study.

Uranium and thorium are present in the resource and waste rocks. Any radon

produced in the mine and process plant is likely manageable with ventilation, dust

control, and monitoring. Expected CAPEX and OPEX costs will not be significantly

increased as a result of these safety measures.






An extensional faulting event has potential for displacements of greater than 10 m.

Such offsets would certainly impact deposit geometry and future mine designs.

Mining recovery may vary from 65% to 85% depending on the success in which

pillars can be mined on retreat and/or fill is utilized, however mining recovery could

be lower and dilution increased in areas with moderate dips greater than 10°.






26.0 RECOMMENDATIONS

AMEC recommends a work program for an estimated total cost of CAD$5.2 million.

The recommendations are based on the current Mineral Resource estimate with

effective date 22 June 2012. Table 26-1 summarizes the recommended work program

with estimated costs.

Table 26-1: Recommendations Summary

Task

Estimated

Budget Comment

Project Management +

Claims + Socio-Economic

+ Administration

$630,000

2012 Field work $625,000 Geology mapping; re-sampling; core review; structural

geology reviews; on-going research; assay QA/QC SRM

replenishment program.

Security (Valemount core

facility)

$75,000 On-going upgrades to Valemount facility.

Marketing $50,000

Mining trade-off studies $200,000 Trade-off studies on mine and stope design

Resource modeling trade-

off studies

$40,000 Studies to optimize grade distribution and domains

Mineral Resource Update

2013

$360,000 Incorporate 2011 drilling, 2012 exploration data, and new

information from mining or metallurgical optimization

trade-off studies.

Diamond drilling:

for mineral resource

definition

$2,660,000 About 40 diamond drill holes comprising an estimate

10,000 m of HQ diameter coring for infill and step-out

drilling to support confidence category upgrades

Diamond drilling:

to supply metallurgical

testwork

$560,000

About 8 diamond drill holes comprising approximately

2,000 m of PQ diameter coring

Metallurgical testwork $ - $1M budgeted for metallurgical testwork in the 2011

PEA. This work is on-going and results will lead to a

more definitive budget going forward.

Total $5,200,000

Project management, field work, and desk top studies total about $2.0 million and

includes the following: (1) project management and administration costs; (2) field

costs comprising a re-sampling program for campaigns with poor precision and

accuracy to improve confidence in their analyses, structural geology studies, and

manpower and field support costs; (3) core farm security improvements; (4) on-going

marketing work; (5) mining trade-off studies to optimize mine and stope design; (6)






resource modeling trade-off studies to optimize grade distribution; and (7) a mineral

resource estimate update.

The re-sampling program should focus on re-assaying samples within an area where

the first five years of mining is likely to occur.

An additional mineral resource update is recommended after all the 2011 drilling data

has been analysed, verified, updated into the drilling database, and interpreted. This

mineral resource update would include all drilling information up to and including the

2011 campaign plus outcomes from any mining, resource modeling, or metallurgical

optimization studies.

Additional diamond drilling is recommended totalling about $3.2 million for drilling,

sampling, assaying, and logging costs. The recommended drilling is to focus on the

volume within the first 5 years of the conceptual mine plan. The recommended

program has about 40 diamond drill holes comprising about 10,000 m of HQ diameter

coring for resource infill and step-out drilling and about 8 diamond drill holes

comprising about 2,000 m of PQ diameter coring for metallurgical testwork purposes.






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Summary Of Current Knowledge, 4 p.






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Upper Fir Carbonatite System: an Update of Current Knowledge with Implications

for Exploration. Confidential report for Commerce Resource Corporation. 54p.

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Currie, K.L. 1976: The Alkaline Rocks of Canada: Geol. Surv. Can., Bull. 239, 228 p.

Dahrouge, J., 2001a: 2000 Geologic Mapping and Sampling on the Verity Property: B.C.

Min. Energy, Mines Petr. Res. Ass. Rept 26550, 7 p.

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Dahrouge, J. and Reeder J., 2001: 2001 Geologic Mapping, Sampling and Geophysical

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Mount Cheadle area, Monashee Mountains: in Current Research, Part E, Geol.

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Property 20 June 2008: 48 p. plus appendices and maps.

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Project No.: 168967 Appendix 22 June 2012

A P P E N D I X A

LIST OF C L A I M S

SCHEDULE A

Blue River Property, located in the Kamloops Mining Division, north and northeast of Blue River, British Columbia, Canada.

Tenure Tenure Tenure Sub Map Number Claim Name Owner Type Type Number

374665 FIR 3 142572 (100%) Mineral Claim 083D025

374670 FIR 8 142572 (100%) Mineral Claim 083D035

380034 MARA 5 142572(100%) Mineral Claim 083D045

382164 FIR II 142572 (100%) Mineral Claim 083D035

506262 142572 (100%) Mineral Claim 083D

506263 142572(100%) Mineral Claim 083D

506264 142572 (100%) Mineral Claim 083D

506265 142572 (100%) Mineral Claim 083D

506267 142572 (100%) Mineral Claim 083D

506270 142572(100%) Mineral Claim 083D

506273 142572 (100%) Mineral Claim 083D

506274 142572(100%) Mineral Claim 083D

506387 142572 (100%) Mineral Claim 083D

506391 142572(100%) Mineral Claim 083D

506392 142572(100%) Mineral Claim 083D

506393 142572 (100%) Mineral Claim 083D

506395 142572 (100%) Mineral Claim 083D

506397 142572 (100%) Mineral Claim 083D

506399 142572 (100%) Mineral Claim 083D

506401 142572 (100%) Mineral Claim 083D

506402 142572 (100%) Mineral Claim 083D

506403 142572 (100%) Mineral Claim 083D

506405 142572 (100%) Mineral Claim 083D

506407 142572(100%) Mineral Claim 083D

506408 142572(100%) Mineral Claim 083D

506423 142572 (100%) Mineral Claim 083D

506425 142572(100%) Mineral Claim 083D

506426 142572 (100%) Mineral Claim 083D

506427 142572 (100%) Mineral Claim 083D

506428 142572 (100%) Mineral Claim 083D

506429 142572 (100%) Mineral Claim 083D

506430 142572 (100%) Mineral Claim 083D

506431 142572 (100%) Mineral Claim 083D

506433 142572 (100%) Mineral Claim 083D

506445 142572 (100%) Mineral Claim 083D

506450 142572 (100%) Mineral Claim 083D

506459 142572 (100%) Mineral Claim 083D

506461 142572 (100%) Mineral Claim 083D

506464 142572 (100%) Mineral Claim 083D

506466 142572 (100%) Mineral Claim 083D

506468 142572 (100%) Mineral Claim 083D

506473 142572 (100%) Mineral Claim 083D

Issue Date Good To Date Status Area (ha)

2000/feb/I 6 2021 /mar/31 GOOD 25.0

2000/feb/I 6 2021/mar/31 GOOD 25.0

2000/aug/18 2021/mar/31 GOOD 25.0

2000/oct/28 2021/mar/31 GOOD 500.0

2005/feb/08 2021/mar/3 I GOOD 98.623

2005/feb/08 2021/mar/3 1 GOOD 295.727

2005/fŁb/08 202 1/mar/31 GOOD 236.8

2005/feb/08 2021/mar/31 GOOD 79.069

2005/feb/08 2021/mar/31 GOOD 98.817

2005/feb/08 2021/mar/31 GOOD 1225.766

2005/feb/08 2021 /mar/31 GOOD 1619.061

2005/feb/08 202 1/mar/31 GOOD 1244.47

2005/feb/09 2021/mar/31 GOOD 98.638

2005/feb/09 2021/mar/31 GOOD 39.459

2005/feb/09 2021/mar/31 GOOD 39.46

2005/feb/09 2021/mar/31 GOOD 39.447

2005/feb/09 2021/mar/31 GOOD 39.452

2005/feb/09 2021/mar/31 GOOD 19.728

2005/feb/09 2021/mar/31 GOOD 79.084

2005/feb/09 2021/mar/3I GOOD 39.542

2005/feb/09 2021/mar/31 GOOD 19.768

2005/feb/09 2021/mar/31 GOOD 19.766

2005/feb/09 202 1/mar/31 GOOD 19.765

2005/feb/09 2021/mar/31 GOOD 591.699

2005/feb/09 2021/mar/31 GOOD 118.38

2005/feb/09 202 1/mar/31 GOOD 591.653

2005/feb/09 2021/mar/31 GOOD 157.847

2005/feb/09 202 1/mar/31 GOOD 39.439

2005/feb/09 2021/mar/31 GOOD 19.717

2005/feb/09 2021/mar/31 GOOD 551.916

2005/feb/09 2021/mar/31 GOOD 78.924

2005/feb/09 202 1/mar/31 GOOD 414.436

2005/fcb/09 202 1/mar/31 GOOD 315.765

2005/feb/09 2021 /mar/31 GOOD 533.482

2005/feb/09 2021/mar/31 GOOD 355.921

2005/feb/09 2021/mar/31 GOOD 236.589

2005/feb/09 202 1/mar/31 GOOD 473.37

2005/feb/09 2021/mar/31 GOOD 315.725

2005/feb/09 2021/mar/31 GOOD 78.95

2005/feb/09 2021/mar/31 GOOD 217.118

2005/feb/09 2021/mar/31 GOOD 355.271

2005/feb/09 2021/mar/31 GOOD 474.81

CW523273 I I

A-2

Tenure Tenure Tenure Sub Map Number Claim Name Owner Type Type Number Issue Date Good To Date Status Area (ha)

506475 142572(100%) Mineral Claim 083D 2005/feb/09 202l/mar/3l GOOD 395.675

507391 142572 (100%) Mineral Claim 083D 2005/feb/17 2021/mar/31 GOOD 553.698

530510 LIGHTNING 142572 (100%) Mineral Claim 083D 2006/mar/24 2021/mar/31 GOOD 494.525

530511 LIGHTNING 2 142572 (100%) Mineral Claim 083D 2006/mar/24 2021/mar/31 GOOD 395.741

530513 LIGHTNING 3 142572 (100%) Mineral Claim 083D 2006/mar/24 2021/mar/31 GOOD 217.556

537452 PYRAMID 1 142572(100%) Mineral Claim 083D 2006/jul/20 202I/mar/31 GOOD 493.795

537454 PYRAMID 2 142572 (100%) Mineral Claim 083D 2006/jul/20 2021/mar/31 GOOD 494.024

537456 PYRAMID 3 142572 (100%) Mineral Claim 083D 2006/jul/20 2021/mar/31 GOOD 197.674

550560 MUD 10 142572(100%) Mineral Claim 083D 2007/jan/29 2021/mar/31 GOOD 495.976



550565 MUD 14 142572 (100%) Mineral Claim 083D 2007/jan/29 2021/mar/31 GOOD 376.8803

550568 MUDI5 142572 (100%) Mineral Claim 083D 2007/jan/29 2021/mar/31 GOOD 178.5237

550603 ARIANEI 142572 (100%) Mineral Claim 0831) 2007/jan/30 2021/mar/31 GOOD 493.6076

550605 ARIANE2 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 493.8371

550607 ARIANF3 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 493.6181

550608 ARIANE4 142572 (100%) Mineral Claim 083D 2007/jan/30 202 1/mar/31 GOOD 493.8457



550612 ARIANE7 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/3I GOOD 473.8467


550614 ARIANE9 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 493.7679

550615 ARIANEIO 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/3l GOOD 474.1587

550616 ARIANEI 1 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 493.4837

550620 ARIANEI2 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 494.1158

550621 4512124519227384 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 474.5925

550622 ARIANEI3 142572 (100%) Mineral Claim 083D 2007/jan/30 202I/mar/31 GOOD 474.8547

550623 ARIANE 14 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 494.3431



550628 ARIANEI7 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 492.9972



550633 ARIANE 20 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 414.104

550636 ARIANE 20 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/3 1 GOOD 493.7078










550648 ARIANE 30 142572 (100%) Mineral Claim 083D 2007/Jan/30 2021/mar/31 GOOD 494.3939


550651 ARIANE 32 142572 (100%) Mineral Claim 083D 2007/jan/30 202l/mar/3l GOOD 493.2738

550652 ARIANE 33 142572 (100%) Mineral Claim 083D 2007/jan/30 202l/mar/31 GOOD 493.0544

CW523273 I . I

A-3


550655 ARIANE 34 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 1971623

550658 AR1ANE 35 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 492.9679






550666 ARIANE 41 142572 (100%) Mineral Claim 0831) 2007/jan/30 2021/mar/31 GOOD 494.1665











550681 ARIANE 53 142572 (100%) Mineral Claim 083D 2007/jan/30 202l/mar/31 GOOD 493.2689



550687 ARIANE 56 142572(100%) Mineral Claim 083D 2007/Jan/30 2021/mar/31 GOOD 473.5216










550704 ARIANE 66 142572 (100%) Mineral Claim 0831) 2007/jan/30 202l/mar/3l GOOD 494.2991


550707 ARIANE 68 142572 (100%) Mineral Claim 0831) 2007/an/30 2021/mar/31 GOOD 435.0935

550709 AR1ANE 69 142572 (100%) Mineral Claim 083D 2007/jan/30 2021/mar/31 GOOD 474.7668

550711 ARIANE 70 142572(100%) Mineral Claim 083D 2007/jan/30 2021/mar/3l GOOD 474.7626









550886 HELLROAR 142572(100%) Mineral Claim 0831) 2007/feb/01 2021/mar/31 GOOD 435.4711

550887 IIELLROARS 142572 (100%) Mineral Claim 083D 2007/feb/01 2021/mar/31 GOOD 475.2464

550888 BAT OUT OF HELL 142572 (100%) Mineral Claim 083D 2007/feb/01 2021/mar/31 GOOD 475.3964 TI IF MONSTER IS

550889 LOOSE 142572(100%) Mineral Claim 083D 2007/feb/Ol 2021/mar/31 GOOD 475.2026

CW5232731 I

BE


565127 PROSPER 1 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 475.099


565129 PROPSER 3 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 494.7982


565131 PROSPERS 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 494.798






565139 PROSPER II 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 494.7982

565140 PROSIER 12 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 475.1588






565147 PROSPER 17 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/3l GOOD 178.195

565148 PROSPER 18 142572 (100%) Mineral Claim 083D 2007/aug/28 202l/mar/31 GOOD 336.7507



565152 PROSPER 21 142572(100%) Mineral Claim 083D 2007/augI28 2021/mar/31 GOOD 495.164















565168 PROSPER 35 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/3 I GOOD 495.3947


565170 IROSPER 37 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 495.5681

565171 SHADOWI 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 495.6504

565172 SHADOW 2 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 436.1667

565173 SI IADOW 3 142572 (100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 495.6211


565175 SHADOWS 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 495.6219





CW5232731 I


565180 SHADOW 9 142572(100%) Mineral Claim 083D 2007/augI28 2021/mar/31 GOOD 475.9524


565182 SHADOW II 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 495.8402


565184 SHADOW 13 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/3I GOOD 495.9423


565186 SHADOW 15 142572 (100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 456.3659

565187 FALKOR 1 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 456.0848


565189 FALKOR 3 142572(100%) Mineral Claim 083D 2007/aug/28 202l/mar/31 GOOD 396.7453


565191 FALKOR 5 142572 (100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 374.6441


565193 FALKOR 7 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 4717231


565195 FALKOR 9 142572(100%) Mineral Claim 083D 2007/aug/28 202l/mar/3l GOOD 496.3699


565197 FALKOR II 142572 (100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 496.401


565199 MINI 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 39.7012




565203 FALKOR 15 142572 (100%) Mineral Claim 083D 2007/augI28 2021/mar/31 GOOD 496.169



565206 MINI 2 142572(100%) Mineral Claim 083D 2007/aug/28 2021/mar/31 GOOD 39.6939










565216 FALKOR 27 142572(100%) Mineral Claim 083D 2007/aug/28 202I/mar/31 GOOD 496.6269






565222 FALKOR 33 142572 (100%) Mineral Claim 083D 2007/aug/28 2021/mar/3 1 GOOD 397.1157



588427 WASTED 1 142572(100%) Mineral Claim 083D 2008/jul/l8 2021/mar/31 GOOD 494.2937

588428 WASTED 2 142572(100%) Mineral Claim 083D 2008/jul/l8 2021/mar/31 GOOD 474.3304

588429 WASTED 3 142572(100%) Mineral Claim 083D 2008/jul/18 2021/mar/31 GOOD 474.1507

CW5232731,1

Tenure Tenure Tenure Sub Map Number Claim Name Owner Type Type Number

588430 WASTED 4 142572 (100%) Mineral Claim 083D

589537 FELIX! 142572 (100%) Mineral Claim 083D

589538 FELIX2 142572 (100%) Mineral Claim 083D



589541 FELIX5 142572(100%) Mineral Claim 083D





589556 FELIX1O 142572 (100%) Mineral Claim 083D

589557 FELIXII 142572(100%) Mineral Claim 083D

589559 FELIXI2 142572 (100%) Mineral Claim 083D

589563 FELIXI3 142572 (100%) Mineral Claim 083D

798362 JOIN 142572 (100%) Mineral Claim 083D

Issue Date Good To Date Status Area (ha)

2008/jul/I 8 202 1/mar/31 GOOD 473.977

2008/aug/05 2021/mar/31 GOOD 496.3964

2008/aug/05 2021/mar/31 GOOD 496.3976

2008/augI05 2021 /mar/31 GOOD 377.1036

2008/aug/05 2021 /mar/31 GOOD 496.1697

2008/aug/05 2021/mar/31 GOOD 495.9655

2008/aug/05 2O21/mar/31 GOOD 436.3101

2008/aug/05 2021/mar/31 GOOD 376.6839

2008/aug/05 2021/mar/31 GOOD 475.8199

2008/aug/05 2021/mar/31 GOOD 396.317

2008/aug/05 202 1/mar/31 GOOD 495.4715

2008/aug/05 2021/mar/31 GOOD 415.9727

2008/aug/05 2021/mar/31 GOOD 495.1674

2008/aug/05 202 1/mar/31 GOOD 356.6787

201 0/jun/25 2021/mar/31 GOOD 177.9794

CW5232731 I