technical report and preliminary economic assessment … · little deer copper deposit...

TECHNICAL REPORT

AND

PRELIMINARY ECONOMIC ASSESSMENT (PEA)

OF THE

LITTLE DEER COPPER DEPOSIT

NEWFOUNDLAND, CANADA

Latitude 49o 32’08 North

Longitude 56o 06’07 West

For

THUNDERMIN RESOURCES INC.

AND

CORNERSTONE RESOURCES INC.

By

P&E Mining Consultants Inc.

Suite 202 - 2 County Court Blvd

Brampton, Ontario,

L6W 3W8

NI-43-101F1

TECHNICAL REPORT No. 227

Mr. Eugene Puritch, P.Eng.

Dr. Wayne Ewert, P.Geo.

Mr. Kirk Rodgers, P.Eng.

Mr. James L. Pearson, P.Eng.

Mr. David Orava, P.Eng.

Mr. Alfred Hayden, P.Eng.

Effective Date: November 1, 2011

Signing Date: December 15, 2011

The effective date of this report is

November 1, 2011

{SIGNED AND SEALED}

[Eugene J. Puritch]

Eugene J. Puritch, P.Eng.

Date of Signature: December 15, 2011

{SIGNED AND SEALED}

[Wayne Ewert]

Dr. Wayne Ewert, P.Geo.


{SIGNED AND SEALED}

[James L. Pearson]

James L. Pearson, P.Eng.


{SIGNED AND SEALED}

[David Orava]

David Orava, P.Eng.


{SIGNED AND SEALED}

[Kirk Rodgers]

Kirk Rodgers, P.Eng.


{SIGNED AND SEALED}

[Alfred Hayden]

Alfred Hayden, P.Eng.


TABLE OF CONTENTS

1.0 SUMMARY ............................................................................................................................... i

1.1 MINERAL RESOURCES AND POTENTIALLY MINEABLE MINERAL

RESOURCES ................................................................................................................ i 2.0 INTRODUCTION ..................................................................................................................... 1

2.1 TERMS OF REFERENCE ........................................................................................... 1 2.2 SOURCES OF INFORMATION ................................................................................. 2 2.3 UNITS AND CURRENCY .......................................................................................... 2 2.4 GLOSSARY AND ABBREVIATION OF TERMS .................................................... 2

3.0 RELIANCE ON OTHER EXPERTS ........................................................................................ 4 4.0 PROPERTY DESCRIPTION AND LOCATION ..................................................................... 5

4.1 LITTLE DEER PROPERTY LOCATION ................................................................... 5 4.2 PROPERTY DESCRIPTION AND TENURE ............................................................. 5 4.3 PERMITS AND OBLIGATIONS ................................................................................ 7

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY........................................................................................................... 8 5.1 ACCESS ....................................................................................................................... 8 5.2 CLIMATE ..................................................................................................................... 8 5.3 LOCAL RESOURCES ................................................................................................. 8 5.4 INFRASTRUCTURE ................................................................................................... 8 5.5 PHYSIOGRAPHY ........................................................................................................ 8

6.0 HISTORY ................................................................................................................................ 10 6.1 PREVIOUS RESOURCE ESTIMATES .................................................................... 11

7.0 GEOLOGICAL SETTING AND MINERALIZATION ......................................................... 14 7.1 REGIONAL ................................................................................................................ 14 7.2 GEOLOGY OF THE LITTLE DEER PROPERTY ................................................... 14 7.3 MINERALIZATION OF THE LITTLE DEER DEPOSIT ........................................ 16

8.0 DEPOSIT TYPES ................................................................................................................... 17 8.1 METALLOGENIC MODEL – VMS DEPOSITS ...................................................... 17 8.2 CYPRUS-TYPE VMS DEPOSITS ............................................................................ 17 8.3 LITTLE DEER DEPOSIT MODEL ........................................................................... 18

9.0 EXPLORATION ..................................................................................................................... 20 9.1 RECENT EXPLORATION (2010-2011) ................................................................... 20

10.0 DRILLING .............................................................................................................................. 21 11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ................................................ 24 12.0 DATA VERIFICATION ......................................................................................................... 25

12.1 SITE VISIT AND INDEPENDENT SAMPLING ..................................................... 25 12.2 QUALITY ASSURANCE/QUALITY CONTROL REVIEW ................................... 26

12.2.1 Performance of Certified Reference Materials ............................................... 26 12.3 PERFORMANCE OF BLANK MATERIAL............................................................. 26

12.3.1 Performance of Secondary Lab Checks ......................................................... 26 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ....................................... 27

13.1 INTRODUCTION ...................................................................................................... 27 13.2 MINERALOGY .......................................................................................................... 27 13.3 GRINDING ................................................................................................................. 27 13.4 FLOTATION .............................................................................................................. 27

14.0 MINERAL RESOURCE ESTIMATE .................................................................................... 29 14.1 INTRODUCTION ...................................................................................................... 29 14.2 DATA SUPPLIED ...................................................................................................... 29 14.3 DATABASE VALIDATION ..................................................................................... 29 14.4 BULK DENSITY ........................................................................................................ 29

14.5 DOMAIN MODELING .............................................................................................. 30 14.6 COMPOSITING AND COMPOSITE SUMMARY STATISTICS ........................... 31 14.7 TREATMENT OF EXTREME VALUES .................................................................. 32 14.8 CONTINUITY ANALYSIS ....................................................................................... 32 14.9 BLOCK MODEL ........................................................................................................ 33 14.10 RESOURCE ESTIMATION & CLASSIFICATION ................................................. 34 14.11 MINERAL RESOURCE ESTIMATE ........................................................................ 35 14.12 VALIDATION ............................................................................................................ 36 14.13 POTENTIALLY MINEABLE MINERAL RESOURCE ESTIMATE ...................... 36

15.0 MINERAL RESERVE ESTIMATES ..................................................................................... 39 16.0 MINING METHODS .............................................................................................................. 40

16.1 LONGHOLE LONGITUDINAL RETREAT MINING METHOD ........................... 40 16.2 MINE AND STOPE DEVELOPMENT ..................................................................... 40 16.3 STOPING .................................................................................................................... 41 16.4 SCHEDULE ................................................................................................................ 43

17.0 RECOVERY METHODS ....................................................................................................... 45 18.0 PROJECT INFRASTRUCTURE ............................................................................................ 46

18.1 SITE SURFACE INFRASTRUCTURE ..................................................................... 46 19.0 MARKET STUDIES AND CONTRACTS ............................................................................ 48 20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT ........................................................................................................ 49 21.0 CAPITAL AND OPERATING COSTS ................................................................................. 51

21.1 CAPITAL COST ESTIMATES ................................................................................. 51 21.1.1 Mine Capital Costs ......................................................................................... 51 21.1.2 Processing Plant Capital Costs ....................................................................... 53 21.1.3 Surface Infrastructure Capital Costs ............................................................... 54 21.1.4 Mine Closure Capital Costs ............................................................................ 55 21.1.5 Other Capital Costs......................................................................................... 56

21.2 OPERATING COST ESTIMATES ............................................................................ 56 21.2.1 Mining ............................................................................................................ 56 21.2.2 Mineral Processing ......................................................................................... 57 21.2.3 Other Related operating Costs ........................................................................ 58 21.2.4 General and Administration ........................................................................... 58

22.0 ECONOMIC ANALYSIS ....................................................................................................... 59 22.1 ECONOMIC CRITERIA ............................................................................................ 59

22.1.1 Physicals ......................................................................................................... 59 22.1.2 Revenue .......................................................................................................... 59 22.1.3 Costs ............................................................................................................... 60

22.2 CASH FLOW .............................................................................................................. 60 22.3 BASE CASE CASH FLOW ANALYSIS .................................................................. 62 22.4 SENSITIVITY ANALYSIS ....................................................................................... 62

23.0 ADJACENT PROPERTIES .................................................................................................... 64 24.0 OTHER RELEVANT DATA AND INFORMATION ........................................................... 65 25.0 INTERPRETATION AND CONCLUSIONS ........................................................................ 66 26.0 RECOMMENDATIONS ........................................................................................................ 67

26.1 RESOURCE UPGRADE ............................................................................................ 68 27.0 REFERENCES ........................................................................................................................ 69 28.0 CERTIFICATES ..................................................................................................................... 70

APPENDIX I. UNDERGROUND MINE PLAN DRAWINGS ............................................ 76

LIST OF TABLES

Table 1.1 Summary of Little Deer Mineral Resources ........................................................... ii

Table 1.2 Resource Summary ................................................................................................ iii Table 1.3 Capital Costs (Life of Mine) ................................................................................... v Table 1.4 Mine Operating Cost per Tonne Milled Summary ................................................. vi Table 1.5 Proposed Exploration Program and Budget .......................................................... vii Table 1.6 Preliminary Budget for Project Development to Pre-Feasibility Study Level ..... viii

Table 4.1 Mineral Licence and Claims Status, Little Deer Property ...................................... 6 Table 6.1 Summary of Historical and Thundermin-Cornerstone Exploration

On the Little Deer Property ................................................................................... 10 Table 6.2 Summary of Micon Little Deer Mineral Resources .............................................. 11 Table 6.3 Summary of RPA Little Deer Mineral Resources ................................................. 12

Table 10.1 Highlights of Drill Intercepts from the 2010/2011 Drill Program ........................ 21 Table 13.1 Summary of Locked Cycle Test Results ............................................................... 28

Table 14.1 Drillhole Database Summary ................................................................................ 29 Table 14.2 Bulk Density Values .............................................................................................. 30 Table 14.3 Domain Composite Summary Statistics ................................................................ 32 Table 14.4 Block Model Setup ................................................................................................ 34

Table 14.5 Summary of Little Deer Mineral Resources ......................................................... 35 Table 14.6 Domain Validation Statistics ................................................................................. 36

Table 14.7 Potentially Mineable Mineral Resources .............................................................. 37 Table 14.8 Resource Summary ............................................................................................... 38 Table 16.1 Summary of Estimated Mine and Stope Development ......................................... 41

Table 16.2 Stoping Drilling and Blasting Para metres ............................................................ 42 Table 16.3 Thundermin / Cornerstone Resources Inc. Stoping Productivities ....................... 43

Table 16.4 Mine Development Summary ............................................................................... 44

Table 16.5 Stope Development Summary ............................................................................... 44

Table 21.1 Capital Cost Schedule and Summary .................................................................... 51 Table 21.2 Mine Development Capital Cost Schedule and Summary .................................... 52

Table 21.3 Stope Development Capital Cost Schedule and Summary ................................... 52 Table 21.4 Underground Equipment Capital Cost Summary .................................................. 53 Table 21.5 Process Plant Capital Cost Summary .................................................................... 54

Table 21.6 Surface Infrastructure Capital Cost Summary ...................................................... 55 Table 21.7 Closure Bond ......................................................................................................... 55 Table 21.8 Mine Operating Cost per Tonne Milled Summary ................................................ 56 Table 21.9 Mine Operating Cost per Tonne of Stope Ore ...................................................... 57

Table 21.10 Process Plant Operating Cost Per Tonne Milled ................................................... 57 Table 21.11 Other Related Operating Costs .............................................................................. 58 Table 22.1 After-Tax Cash Flow Summary ............................................................................ 61

Table 22.2 Base Case Cash Flow Analysis ............................................................................. 62 Table 22.3 Sensitivity Item Values ......................................................................................... 63 Table 22.4 Summary of Sensitivity Analysis .......................................................................... 63 Table 26.1 Preliminary Budget for Project Development to Pre-Feasibility Study Level ...... 67

Table 26.2 Proposed Exploration Program and Budget .......................................................... 68

LIST OF FIGURES

Figure 1.1 Longitudinal Section .............................................................................................. iv

Figure 4.1 Location of the Little Deer Property ....................................................................... 5 Figure 4.2 Little Deer Property Claims Map ............................................................................ 7 Figure 5.1 View of Deer Pond, Looking South West ............................................................... 9 Figure 6.1 Schematic Cross Section, Little Deer Deposit ...................................................... 13 Figure 7.1 Simplified Geology and Location of Past-Producing Mines in Newfoundland .... 15

Figure 7.2 Local Geology of the Little Deer and Whalesback Mine Area ............................. 16 Figure 8.1 Schematic Diagram of a VMS Deposit ................................................................. 17 Figure 8.2 Schematic Model Illustrating a Possible Explanation for Two Copper

Stringer Zones–Paleovolcanic Listric Normal Faults ........................................... 19 Figure 10.1 Drillhole Location ................................................................................................. 23

Figure 12.1 Site Visit Sample Results for Copper .................................................................... 25 Figure 13.1 Locked Cycle Test Flowsheet ............................................................................... 28

Figure 14.1 Isometric Projection of Mineral Resource Domains ............................................. 30 Figure 14.2 Assay Sample Length Histogram .......................................................................... 31 Figure 14.3 Decile Analysis Results ......................................................................................... 32 Figure 14.4 Experimental Semi-Variograms ............................................................................ 33

Figure 14.5 Isometric Projection of Block Classification ........................................................ 35 Figure 14.6 Little Deer Domain Swath Plot ............................................................................. 36

Figure 18.1 Little Deer Copper Deposit Infrastructure ............................................................ 47 Figure 22.1 Sensitivity Graph ................................................................................................... 63

P&E Mining Consultants Inc. i

Thundermin Resources Inc. Little Deer Deposit PEA Report No. 227

1.0 SUMMARY

The following report was prepared to provide a National Instrument 43-101 (“NI 43-101”)

compliant Technical Report and Preliminary Economic Assessment (“PEA”) of the copper

mineralization contained in the Little Deer Copper Deposit (“Deposit”), located approximately

10 kilometres (“km”) north of the town of Springdale in north-central Newfoundland, Canada.

The Deposit is subject to a joint venture arrangement (the “LDJV”) between Thundermin and

Cornerstone who jointly own a 100% interest in the Deposit on a 50/50 basis with Thundermin

as the operator.

This report was prepared by P&E Mining Consultants Inc. (“P&E”) at the joint request of Mr.

John Heslop, President and CEO of Thundermin Resources Inc. (“Thundermin”), a Toronto-

based resource company and Mr. Brooke Macdonald, President of Cornerstone Resources Inc.

(“Cornerstone”), a Newfoundland-based resource company.

The Little Deer property comprises four mineral licenses containing a total of 276 staked claims

covering a total area of approximately 6,530 hectares (“ha”) (the “Property”). The LDJV has a

100% interest in the Property which comprises mineral licences 12196M, 10215M, 10214M and

16456M. All claims are in good standing as of the effective date of this report.

The Deposit is located in the northeastern sector of the Property at approximate UTM (NAD27,

Zone 21) grid co-ordinates 571,000E, 5,493,000N (approximately 49 32‟08” north latitude by 56

06‟07” west longitude).

The project site is easily accessible via a series of gravel roads which extend northwards from

paved highway Route 392 which connects Springdale to the small community of Little Bay

20 km to the northeast.

There are excellent local resources and infrastructure to support exploration and mining activities

and personnel are readily available from the town of Springdale, Newfoundland.

The area is characterized by a series of northeast-trending ridges and valleys which reflect the

underlying geological controls.

1.1 MINERAL RESOURCES AND POTENTIALLY MINEABLE MINERAL

RESOURCES

The Little Deer Copper Deposit (“Deposit”) was initially mined from 1970 to 1972 by British

Newfoundland Exploration Limited (“BRINEX”) via a 1,044 m drift on the 244 m (800) level of

the Whalesback Mine located approximately 800 m northeast of the Deposit. Operations at Little

Deer ceased in 1972 with the closure of the Whalesback Mine. In 1973, the Deposit was leased

by Green Bay Mining Company Limited (“Green Bay”) and they accessed the shallower portion

of the Deposit via a 329 m decline from surface. Development and mining was carried out

between 1973 and 1974 at which time operations ceased due to low copper prices.

The 2011 drill program (December 2010 – June 2011) comprised a total of 12,576 m in 25 holes.

The program was designed to update and expand the existing NI 43-101 mineral resource

(Puritch and Ewert, 2011). The updated mineral resource estimate for the Deposit is based on

assay results from 48,432 m of drilling 82 holes completed by Thundermin and Cornerstone

since June 2007 and assay data from a total of 102 surface and 122 underground historical holes

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that were drilled by BRINEX between 1961 and 1970 and Mutapa Gold Corporation between

1998 and 2000. The historical information was recovered from the archives of the Newfoundland

and Labrador Department of Natural Resources in St. John‟s, Newfoundland.

TABLE 1.1

SUMMARY OF LITTLE DEER MINERAL RESOURCES(1)(2)(3)(4)(5)(6)(7)

Resource Classification/Zone Tonnes Cu% Cu lbs. (M)

Indicated Mineral Resources

Little Deer Zone 1,911,000 2.37 99.8

Inferred Mineral Resources

Little Deer Zone 1,240,000 1.93 52.8

Little Deer Footwall Zone 1,711,000 2.04 77.0

Little Deer Footwall Zone Splay 797,000 2.64 46.2

Total Inferred Resources 3,748,000 2.13 175.9

(1) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The

estimate of mineral resources may be materially affected by environmental, permitting, legal, title, taxation,

socio-political, marketing, or other relevant issues.

(2) The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there

has been insufficient exploration to define these Inferred resources as an Indicated or Measured mineral

resource and it is uncertain if further exploration will result in upgrading them to an Indicated or

Measured mineral resource category.

(3) The mineral resources in this press release were estimated using the Canadian Institute of Mining,

Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and

Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.

(4) Ordinary Kriging was used for Cu grade interpolation.

(5) Grade capping of 15% Cu utilized on composites.

(6) A variable bulk density based on numerous field measurements was used for tonnage calculations.

(7) A copper price of US$3.42/lb. (May 31 2011 24 month trailing average) and an exchange rate of

US$0.95US=C$1.00 was utilized to derive the 1% Cu cut-off grade. Mining costs were C$40/t, process

costs were C$15/t and G&A was C$5/t. Concentrate freight and smelter treatment charges were C$10/t

mined. Concentrate mass pull was 7%, process recovery was 97%, smelter payable was 96% and Cu

refining was US$0.07/lb.

The increase in tonnage in the updated mineral resource estimate for the Deposit compared to the

previous estimate (Pressacco, 2010), is due to a reinterpretation of the sectional data for the

Deposit, the inclusion of all of the historical assay data recovered from the archives, the assay

data from the 25 new holes drilled in the 2011 drill program and the use of length weighted bulk

density data for individual assay samples that was not used previously.

A potentially mineable portion of these mineral resources was determined as a basis for a

Preliminary Economic Assessment of the Deposit. The envisaged mining methods are estimated

to experience mining dilution in the order of 20% at zero grade. Mine recovery (extraction) is

estimated to be 90%. A summary of Potentially Mineable Mineral Resources, including dilution

and recovery, is presented in Table 1.2.

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TABLE 1.2

RESOURCE SUMMARY

Potentially Mineable Mineral Resources

Description Resource Au Ag Co Cu

Tonnes g/t g/t % %

Mineral Resources Included 5,652,500 0.055 2.279 0.022 2.212

Diluted Mineral Resources 6,783,000 0.046 1.899 0.019 1.843

Total Potentially Mineable Mineral Resources 6,104,699 0.046 1.899 0.019 1.843

*Note: Some values have been rounded. The totals are accurate summations of the columns of data.

The Potentially Mineable Mineral Resources contain Inferred Mineral Resources which have not

been sufficiently drilled to confidently demonstrate economic viability. In addition, the work

undertaken on the Little Deer Project to date is considered to be at conceptual levels of study

only. As such, and according to the NI 43-101 Regulations, it is not possible to declare a mineral

reserve of any kind.

A conceptual mining and processing plan has been developed to assess the potential of

economically extracting metals from the Deposit.

The envisaged mining plan includes a preliminary -15% ramp access to the Deposit followed by

a combined ramp and shaft access. The ramp would ultimately extend to the -1,000 m elevation,

1,126 m below surface. The shaft would ultimately extend to the -730 m elevation, 875 m below

surface. Ventilation raises would be developed at the deposit extremities.

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Figure 1.1 Longitudinal Section

Footwall drifts will be developed from the shaft or ramp at 90 m intervals, to provide cross-cuts

for access to the deposit.

The selected mining method is Longhole Longitudinal Retreat. At 30 m intervals above the

footwall drift elevation, stope drifts will be developed to the full width of the deposit. These

drifts will provide access for the successive operations of slot raise development, blasthole

drilling and blasting and backfill placement. Removal of the mineralization from the stope will

be accomplished at the footwall drift elevation. Cemented hydraulic tailings will provide the

majority of the backfill placed. This will be supplemented with waste rock. Stope mining would

commence from the top of the mine and progress downwards in successive three sublevel

increments or “stoping blocks”.

It is estimated that 208 stopes would be mined over the mine life. This would generate an

average of 1,800 tonnes per day (“tpd”) composed of 1,477 stoping tonnes (“t”) and 323 t from

the drift and raise development in the Deposit.

SGS Mineral Services of Lakefield, Ontario (“SGS”) carried out a characterization and flotation

test program on a composite sample from the Deposit (Imeson, 2010). A Bond ball mill index of

13.2 kWh/T (14.6 kWh/t) was measured, indicating a material of average hardness. Rougher

flotation tests at a grind of 90 microns with a moderately elevated pH of 9–9.5 using lime and

isopropyl xanthate as collector yielded 99% recovery at a concentrate grade of 12% Cu,

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indicating excellent performance. A regrind size of about 30 microns was indicated. Locked

cycle testing yielded approximately 97% copper recovery and concentrate grades of 28% Cu. It

was noted that there are some minor issues indicated with the pyrrhotite which may impact

recovery or concentrate grade.

Based on these data, a conventional process flowsheet was selected, including crushing and

grinding to a 90 micron grind at a rate of 1,800 t/d, followed by flotation recovery of copper to a

rougher concentrate. The rougher concentrate would be reground to minus 30 microns and

cleaned in a three stage flotation circuit to yield a final concentrate containing copper at a

marketable grade. The concentrate would be filtered to an assumed 8% moisture content for

shipment. Power requirements for the milling process are estimated to be approximately

28 kWh/t.

The Little Deer Project has minimal infrastructure requirements due to its location close to the

Trans-Canada Highway and due to the existence of infrastructure established during its previous

operating history.

Electric power for the Little Deer Project will come from the provincial electrical substation

located just outside Springdale on Highway 392. A tailings storage strategy would be developed

based on an assessment of the existing tailings impoundment area and other potential storage

sites.

The proposed Little Deer Project would be developed, operated and closed in accordance with

environmental and health and safety regulatory requirements.

The estimated total capital costs for the Project total approximately $303.4 million (see Table

1.3). This is composed of approximately $98.5 million in preproduction capital costs and

$204.9 million in sustaining capital costs. This includes the costs for the ongoing development of

the mine.

TABLE 1.3

CAPITAL COSTS (LIFE OF MINE)

Description Total

Mine Development 120.4

Stope Development 75.8

Shaft Related Equip 13.0

Mine Equipment 19.9

Misc. U/G 1.0

Processing Plant 45.3

Surface Infrastructure 18.0

Closure Bond 6.0

Powerline Construction 2.0

Purchase Royalties 2.0

Total 303.4


The estimated total average operating cost of the mine is $47.32 per tonne of ore milled. This is

composed of the components listed in Table 1.4.

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TABLE 1.4

MINE OPERATING COST PER TONNE MILLED SUMMARY

Description $CAN/t

Mining

Stoping Costs 21.41

Underground Haulage 3.16

Underground Hoisting Services Costs 1.39

Mineral Processing

Process Plant Operating 13.32

Cemented Hydraulic Tailings Backfill 2.75

Tailings to Tailings Impoundment Area 0.19

Tailings Pond Water Treatment 0.08

G&A Costs 5.00

Total Operating 47.32

The project was evaluated on an after-tax cash flow basis and generates a net cash flow of

$165.9 million. This results in an after-tax Internal Rate of Return (IRR) of 21.5% and an after-

tax Net Present Value (NPV) of $86.7 million when using a 6% discount rate. In the base case

scenario, the project has a payback period of 3.8 years. The copper price used in this PEA is

US$3.75/lb and the US$/CAN$ exchange rate used in the PEA is 0.95. At forecast metal prices

and exchange rates the break-even copper price is estimated to be US$1.14/lb Cu payable at an

average operating cost of CAN$47.32 per ore tonne ore processed.

This after-tax base case NPV is most sensitive to the $CAN/$US exchange rate followed by the

Cu metal price, Cu head grade and metallurgical recoveries, followed by the capital and

operating costs.

P&E concludes that the Deposit has economic potential as an underground mining and milling

operation producing copper concentrates.

P&E recommends that Thundermin and Cornerstone advance the project with extended and

advanced technical studies particularly in metallurgical, geotechnical and environmental matters

with the intention to advance the project to a pre-feasibility stage. This would include:

Perform a comprehensive program of metallurgical testing on representative

samples of the mineralized zone, to assess and confirm expected recoveries,

reagent usages, process flow sheets and other associated operating issues.

Carry out hydrogeological and hydrological analyses of the project site and

surrounding area. Carry out a more detailed geotechnical assessment of ground

conditions to be able to estimate the ground support required and expected waste

rock dilution of the mill feed with more confidence in subsequent studies.

P&E is also of the opinion that Thundermin and Cornerstone should undertake further

exploration work and the following program is recommended for the Deposit for the period

October 1, 2011 to June 30, 2012, with a $2.0 M budget:

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A small program involving the re-assaying of standards and other check samples

using aqua regia and four acid digestions to try and determine if the estimated

resource grade may, in fact, be higher than estimated as discussed in Section

12.2.1 of this report.

Additional diamond drilling to test for extensions of the copper mineralization at

depth and along strike. Infill diamond drilling on approximately 50 m centres to

upgrade the Inferred Resources to the Indicated Resource category. The infill

drilling, which will be required in order to undertake a pre-feasibility or feasibility

study on the Deposit, should commence at shallower levels of the Deposit and

proceed to depth.

Borehole Pulse EM surveys on selected deep drill holes.

Differential GPS surveys on all new drill hole collars.

Revised NI 43-101 mineral resource estimate following completion of the

recommended diamond drill program.

It is anticipated that this work will be undertaken in two phases: Phase 1 (approx. October 1 to

December 31, 2011 and Phase 2 (January 1 to June 30, 2012), as shown in Table 1.5.

TABLE 1.5

PROPOSED EXPLORATION PROGRAM AND BUDGET

Phase 1 CAN$

4,500 m of diamond drilling at $120.00 per m 540,000

Differential GPS surveying of all new drill holes 2,500

Re-assaying 500

Total Phase 1 543,000

Phase 2

12,000 m of diamond drilling at $120.00 per m 1,440,000


Borehole Pulse EM surveys on 5-6 holes 40,000

Revised NI 43-101 Mineral Resource estimate 25,000

Total Phase 2 1,508,000

The estimated drilling costs are “all-in” costs and include direct drilling costs, salaries and

wages, assaying, room and board, truck rentals, management fees etc.

The combined cost of the recommended program of site investigation and exploratory drilling is

estimated to be $5.25 million. A breakdown of this cost is provided in Table 1.6.

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TABLE 1.6

PRELIMINARY BUDGET FOR PROJECT DEVELOPMENT TO PRE-FEASIBILITY STUDY LEVEL

Proposed Budget Elements CAN&

Hydrogeological and hydrological analyses $200,000

Geotechnical test work. $750,000

Infrastructure $200,000

Engineering design of a TMF and waste rock pile $500,000

Pre-Feasibility level metallurgical testwork including both bench scale and

limited pilot scale. $400,000

Resource Upgrade drilling from inferred to indicated and measured category

(see Table 1.5) $2,000,000

Preliminary environmental and socio-economic impact assessment work in the

project area and data collection for a EIA $100,000

Test work on tailings characterization and treatment options. $150,000

Examination of land acquisition options and acquisition cost. $50,000

Preparation of a Pre-Feasibility Study $900,000

Total Budget $5,250,000

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2.0 INTRODUCTION

2.1 TERMS OF REFERENCE

The following report was prepared to provide a National Instrument 43-101 (“NI 43-101”)

compliant Technical Report and Preliminary Economic Assessment (“PEA”) of the copper

mineralization contained in the Little Deer Copper Deposit (“Deposit”), located approximately

10 km north of the town of Springdale, Newfoundland, Canada. Thundermin Resources Inc.

(“Thundermin”), the project operator, and its joint venture partner Cornerstone Resources Inc.

(“Cornerstone”) own, on a 50/50 basis, a 100% interest in the Deposit and adjacent property (the

“Property”).

This report was prepared by P&E Mining Consultants Inc. (“P&E”) at the joint request of

Mr. John Heslop, President and CEO of Thundermin, a Toronto-based resource company and

Mr. Brooke Macdonald, President of Cornerstone, a Newfoundland-based resource company.

The corporate offices for Thundermin and Cornerstone are as follows:

Thundermin Resources Inc. Cornerstone Resources Inc.

Suite 201, 133 Richmond Street West 26 Kyle Avenue

Toronto, ON Mount Pearl, NL

M5H 2L3 A1N 4R5

Tel: 647-344-1167 Tel: 709-745-8377

Fax: 416-364-5098 Fax: 709-747-1183

This report has an effective date of November 1, 2011.

Mr. Eugene Puritch, a Qualified Person (“QP”) under the regulations of NI 43-101, conducted a

site visit and independent verification sampling program at the Property on May 16, 2011

(Puritch and Ewert, 2011).

In addition to the site visit, P&E has held discussions with technical personnel from Thundermin

and Cornerstone regarding all pertinent aspects of the project and carried out a review of all

available literature and documented results concerning the Property. The reader is referred to

those data sources, which are outlined in the References, Section 26.0 of this report, for further

detail.

The present Technical Report is prepared in accordance with the requirements of NI 43-101F1 of

the Ontario Securities Commission (“OSC”) and the Canadian Securities Administrators

(“CSA”).

The Mineral Resources in the estimate are considered compliant with the Canadian Institute of

Mining, Metallurgy and Petroleum (“CIM”) Standards on Mineral Resources and Reserves,

Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions.

The purpose of the current report is to provide an independent, NI 43-101 compliant, Technical

Report and PEA of the Deposit. P&E understands that this report will be used for internal

decision making purposes and may be filed as required under TMX regulations. The TMX

Group owns and operates Toronto Stock Exchange and TSX Venture Exchange The report may

also be used to support public equity financings.



2.2 SOURCES OF INFORMATION

This report is based, in part, on internal company technical reports, maps and technical

correspondence, published government reports, press releases and public information as listed in

the References (Section 27) at the end of this report. Several sections from reports authored by

other consultants have been directly quoted or summarized in this report, and are so indicated

where appropriate.

With regard to certain sections of the current report, the authors have drawn heavily upon

selected portions or excerpts from material contained in a NI 43-101 technical report prepared by

P&E as noted below:

Puritch, E.J., and Ewert, W.D., 2011: Technical Report and Resource Estimate Update on the

Little Deer Copper Deposit Newfoundland, Canada, dated August 5, 2011.

2.3 UNITS AND CURRENCY

Unless otherwise stated all units used in this report are metric. Copper values are reported in

pounds per tonne (“lbs Cu/t”) unless some other unit is specifically stated. The CAN$ is used

throughout this report unless otherwise specifically stated.

2.4 GLOSSARY AND ABBREVIATION OF TERMS

In this document, the following terms have the meanings set forth below unless the context

otherwise requires.

“$” and “CAN$” means the currency of Canada

“AAS” means Atomic Absorption Spectroscopy

“AA” is an acronym for Atomic Absorption, a technique used to measure metal

content subsequent to fire assay

“asl” means above sea level

“Au” means gold

“C” means degrees Celsius

“CIM” means the Canadian Institute of Mining, Metallurgy and Petroleum

“cm” means centimetres

“Co” means Cobalt

“Cornerstone” means Cornerstone Resources Inc.

“Cu” means Copper

“CSA” means the Canadian Securities Administrators

“E” means east

“el” means elevation level

“Ga” means gigayear, a unit of a billion years

“ha” means Hectare

“km” means kilometre

“kwh/t” kilowatt hour per tonne

“lbs Cu/t” means pounds of copper per tonne

“m” means metre

“M” means million

“Ma” means millions of years

“mm” means millimetres



“MMER” means metal mining effluent regulations

“Mt” means millions of tonnes

“N” means north

“NE” means northeast

“NI 43-101” means National Instrument 43-101

“NPI” means Net Profit Interests

“NTS” means National Topographic System

“NW” means northwest

“NSR” means an acronym for net smelter return, which means the amount

actually paid to the mine or mill owner from the sale of ore, minerals and

other materials or concentrates mined and removed from mineral

properties, after deducting certain expenditures as defined in the

underlying smelting agreements

“oz./T” means ounces per short ton

“P&E” means P&E Mining Consultants Inc.

“PEA” means a Preliminary Economic Assessment

“Property” means the Little Deer Property

“ppb” means parts per billion

“ppm” means parts per million

“S” means south

“SE” means southeast

“SEDAR” means the System for Electronic Document Analysis and Retrieval

“SW” means southwest

“t” means tonnes (metric measurement)

“t/a” means tonnes per year

“Thundermin” means Thundermin Resources Inc.

“TN” means True North

“tpd” means tonnes per day

“TSX-V” means the TSX Venture Exchange

“US$” means the currency of the United States

“UTM” means Universal Transverse Mercator

“SWRPA” means Scott Wilson Roscoe Postle Associates

“W” means west



3.0 RELIANCE ON OTHER EXPERTS

P&E has assumed, and relied on the fact, that all the information and existing technical

documents listed in the References (Section 27) of this report are accurate and complete in all

material aspects. While we carefully reviewed all the available information presented to us, we

cannot guarantee its accuracy and completeness. We reserve the right, but will not be obligated

to revise our report and conclusions if additional information becomes known to us subsequent to

the date of this report.

Copies of the tenure documents were reviewed by P&E and an independent but cursory

verification of claim title was performed using the Mineral Rights Inquiry form found on the

Newfoundland and Labrador Department of Natural Resources‟ website

(http://gis.gov.nl.ca/mrinquiry/mrinquiry.asp). Operating permits and licenses, and work

contracts were not reviewed. P&E has not verified the legality of any underlying agreement(s)

that may exist concerning the licenses or other agreement(s) between third parties but has relied

on, and believes it has a reasonable basis to rely upon, Mr. Andrew Hussey, P.Geo., Lands

Manager for Cornerstone and senior geologist for the Little Deer Joint Venture (“LDJV”) to have

conducted the proper legal due diligence in this regard.

Select technical data, as noted in the report, were provided by Thundermin and Cornerstone, and

P&E has relied on the integrity of such data.

A draft copy of the report has been reviewed for factual errors by the clients and P&E has relied

on Thundermin‟s and Cornerstone‟s knowledge of the Property in this regard. All statements and

opinions expressed in this document are given in good faith and in the belief that such statements

and opinions are not false and misleading at the date of this report.

http://gis.gov.nl.ca/mrinquiry/mrinquiry.asp



4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 LITTLE DEER PROPERTY LOCATION

The Little Deer Property is located approximately 10 km north-northeast of the town of

Springdale in north-central Newfoundland (see Figure 4.1) at approximate UTM (NAD 27,

Zone 21) grid coordinates 571,000E and 5,493,000N (approximately 49o32‟,08” north latitude

and 56o06‟07” west longitude).

Figure 4.1 Location of the Little Deer Property

(Source: Pressacco, 2009)

4.2 PROPERTY DESCRIPTION AND TENURE

The Property comprises 4 mineral licenses containing a total of 276 staked claims covering a

total area of approximately 6,530 hectares (Figure 4.2). Surface rights are not part of the land

holdings and the claim boundaries of all the map-staked claims are currently established by

geographic (UTM grid) reference. The claim boundaries of all ground-staked claims are

established by placement of claim posts along the claim lines and at the corners of the claims.

Such claims have not been land surveyed.

A schedule of claims has been provided by Thundermin and is presented in Table 4.1 and shown

in Figure 4.2.



TABLE 4.1

MINERAL LICENCE AND CLAIMS STATUS, LITTLE DEER PROPERTY (AS OF JUNE 16, 2011)

Licence

Number

Number

of

Claims

Issuance

Date

Assessment

Year

Renewal

Date

Expiration*

Date

Expenditur

es Required

Expenditures

Due Date

Licence

Holder

Surface

Rights Status

10214M 4 15-May-95 17 Not

Applicable 15-May-15 $0.00

Not

Applicable

Weyburn

Investments

Ltd.

100%

Crown

Land

Good

Standing

10215M 20 9-Jan-95 17 Not

Applicable 9-Jan-15 $0.00

Not

Applicable

Weyburn

Investments

Ltd.

100%

Crown

Land

Good

Standing

16456M 20 23-Jun-05 6 23-Jun-15 23-Jun-25 $10,140.30 23-Jun-12

Weyburn

Investments

Ltd.

100%

Crown

Land

Good

Standing

12196M 232 24-May-02 10 24-May-12 24-May-22 $37,333.37 24-May-13

Cornerstone

Resources

Inc.

100%

Crown

Land

Good

Standing

Claims 276

Area

(km2)

65.3

*Note: Mineral licences in Newfoundland and Labrador may be held for a maximum of 20 years, after which time they must be converted to a Mining Lease.



The Little Deer project is subject to a Joint Venture between Thundermin and Cornerstone

(50/50 basis, with Thundermin acting as the operator) that was formed in June 2007 when an

option to acquire a 100% interest in the past-producing Deposit and adjoining claims from

Weyburn Investments Ltd (“Weyburn”) was signed. Details regarding terms of the agreement

were presented in a joint Thundermin and Cornerstone news release that was issued on

May 1, 2007. A summary of the JV terms can also be found on the websites of Thundermin

(http://www.thundermin.com/) and Cornerstone (http://www.cornerstoneresources.com/). On

July 12, 2011, Thundermin and Cornerstone exercised their option to acquire a 100% interest in

the Deposit and adjoining lands from Weyburn.

As of the effective date of this report, all the Little Deer claims are in good standing.

Figure 4.2 Little Deer Property Claims Map

4.3 PERMITS AND OBLIGATIONS

On-going exploration work, including the creation of drill access roads and drill platforms

requires the approval of the Newfoundland and Labrador Department of Natural Resources. All

required permits for such exploration work are currently in place.

http://www.thundermin.com/

http://www.cornerstoneresources.com/



5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY

5.1 ACCESS

The Property is located in the western Notre Dame Bay area of north-central Newfoundland,

approximately 10 km north northeast of the town of Springdale (see Figure 5.1). The project site

is easily accessed via a network of gravel roads which extend north from paved highway Route

392, which connects Springdale to the small community of Little Bay 20 km to the northeast.

5.2 CLIMATE

The climate of north-central Newfoundland is northern temperate generally with cold winters

and short, moderately hot summers. Temperatures range from approximately +22oC during the

summer to -15oC during the winter. Yearly precipitation averages approximately 1000 mm, with

Environment Canada reporting an average of 747 mm of rain and 253 cm of snow for Springdale

during the period 1970-2000.

It is expected that mining activity on the Property could be conducted year-round.

5.3 LOCAL RESOURCES

The Notre Dame Bay area has a long history of copper mining. Between 1860 and the end of

World War I, more than two dozen copper mines had been in production, including the Tilt

Cove, Betts Cove and Little Bay mines. Copper production peaked in the 1880‟s when

Newfoundland was the world‟s sixth largest copper mining area. The area still retains a strong

mining culture and local residents are supportive of the mining industry.

The nearby town of Springdale has a population of approximately 2,800 and is a service centre

for the Green Bay area, with general amenities and community services available. Springdale

also has several local diamond drilling contracting companies and an analytical laboratory. The

area also has a skilled work force, many of whom have experience working in the mineral

exploration and mining industry.

5.4 INFRASTRUCTURE

The project site is located immediately north of paved highway Route 392 and 20 km northeast

of the Trans-Canada Highway. An electrical power transmission line parallels Route 392 and a

high voltage electrical substation is located 10 km south southwest just outside Springdale.

The project area has several lakes and ponds which provide an ample supply of fresh water.

There are several deep water marine ports suitable for shipping future copper concentrates

located nearby (e.g. Little Bay, 10 km away; Goodyear‟s Cove, 32 km away).

5.5 PHYSIOGRAPHY

The regional physiography of the western Notre Dame Bay area is characterized by a series of

northeast-trending ridges and valleys which reflect the underlying geological controls (lithology

and fault structures). Elongated coastal bays, as well as inland drainage patterns and the

orientation of lakes, also generally parallel this structural trend.



The Property area exhibits gently to steeply rolling topography which is forested with spruce, fir

and birch. Hilltops are occasionally barren and low-lying areas and valleys are covered by bogs,

swamps, lakes and ponds. The Deposit area is located underneath and to the west of Deer Pond

at an elevation ranging from approximately 105 to 150 metres asl (Figure 5.1).

Figure 5.1 View of Deer Pond, Looking South West




6.0 HISTORY

A brief history of exploration and development work on the Property is presented below in Table

6.1.

TABLE 6.1

SUMMARY OF HISTORICAL AND THUNDERMIN-CORNERSTONE EXPLORATION ON THE LITTLE

DEER PROPERTY

Year Company Exploration

1952

Falconbridge

Nickel

Mines Ltd

Initial discovery of the Little Deer property.

1955 BRINEX General prospecting and soil geochemical surveys.

1960-1962 BRINEX

Detailed geological mapping, magnetic, electromagnetic and self-

potential geophysical surveys. Additional geochemical surveys

detected a series of copper anomalies extending from the north

shore of Little Deer pond to the east bay of the lake.

25 boreholes were advanced beneath the lake which revealed the

continuation of the mineralized zone over a strike length of 244 m

with an average width of 8 m.

1963 BRINEX 12 more boreholes advanced which indicated an easterly extension

of the mineralization at depth and a parallel (East) lens.

1965-1972 BRINEX

Extensive drilling on Property. Mining activities treated as a co-

development to the underground operations at the nearby

Whalesback Mine. Achieved by driving a 1,044 m tunnel at a

depth of 244 m (800 ft. level) which served as the main haulage

level. Limited development, no accurate production records from

this time.

Production was thought to be limited due to the secondary nature

of its development to Whalesback, the inadequate nature of the

exploration work (i.e. – there were no established mineable

reserves) and the premature closure of the Whalesback Mine due

to low copper prices.

1973-1974 Green Bay

Mining Co.

Little Deer Mine reopened. Financial difficulties and poor copper

prices caused operations to cease. Development limited to shallow,

low grade copper resources that were accessible from a 329 m

decline ramp driven from surface at the Little Deer Mine site.

1998-2000 Mutapa

Gold Corp.

Geological mapping, surface and borehole geophysical surveys. 12

diamond bore holes advanced for a total of 6,815 m of drilling.

Drilling focused on the possible west-south western strike

extension to the Duck Pond area.

2000 Mutapa

Gold Corp.

Mutapa Gold Corp. returned Property to owners due to low copper

prices and a change in business focus to the tech sector.

2007

Thundermin

&

Cornerstone

Option to earn a 100% interest in the Property acquired from

Weyburn. Initial program of diamond drilling (4,941.55 m in 8

DDH), line cutting, GPS surveying, and data compilation followed

by a program of diamond drilling (8,887.85 m in 17 DDH), GPS

surveying, data compilation, gyro surveying, whole rock sampling,



TABLE 6.1

SUMMARY OF HISTORICAL AND THUNDERMIN-CORNERSTONE EXPLORATION ON THE LITTLE

DEER PROPERTY

Year Company Exploration

borehole Pulse-EM surveying, along with 227 line km of airborne

versatile time domain electromagnetic (VTEM) surveying.

2008

Thundermin

&

Cornerstone

14 Boreholes advanced totalling 9,004 metres. 150 samples taken

for analysis. Down-hole geophysics using Pulse EM completed on

14 boreholes. 227 line kilometres of Versatile Time Domain

Electromagnetic and magnetic airborne survey was flown over a

portion of the Little Deer deposit and the adjoining Weyburn

licenses to the east.

2009

Thundermin

&

Cornerstone

Diamond drilling (11,377.0 m in 17 DDH), GPS surveying,

compilation, borehole Pulse-EM geophysical surveys, initial 43-

101 mineral resource estimate, prospecting.

2010 Thundermin

&

Cornerstone

Diamond drilling (11,501.6 m in 18 DDH, including 3 holes

drilled in December as part of 2011 drill program), line cutting,

GPS surveying, data compilation, borehole Pulse-EM geophysical

surveys, Induced Polarization (IP) geophysical survey, updated 43-

101 mineral resource estimate, initial metallurgical test work, and

prospecting.

6.1 PREVIOUS RESOURCE ESTIMATES

There are no technically supported historical resource evaluations of the sulphide mineralization

at Little Deer. Former staff at the Whalesback and Little Deer mines stated that no mineral

resources were attempted during the BRINEX period because the deposit shape, geometry and

grade characteristics were poorly understood. Mining at Little Deer was via a development drift

at the 244 m level (the 800 foot level) which was established from the Whalesback Mine located

approximately 1,800 m to the northeast.

At the cessation of the Green Bay Mining Company‟s operations in 1974, an unsupported

statement was released suggesting that a reserve of 210,200 t, grading 1.53% Cu remained above

the 245 m elevation. It should be noted that this estimate is historic in nature, has not been

reviewed by a QP and should not be relied upon.

In 2009, Micon prepared a NI 43-101 compliant Mineral Resource estimate for the Deposit

(Pressacco, 2009) using the Gemcom software package. Micon estimated that the deposit

contained Indicated Mineral Resources of 1,087,000 t grading 2.90% Cu and Inferred Resources

of 1,950,000 t grading 2.29%.

TABLE 6.2

SUMMARY OF MICON LITTLE DEER MINERAL RESOURCES AS OF AUGUST 14, 20091

Resource Classification Tonnes Cu% Cu lbs. (M)

Indicated Mineral Resources 1,087,000 2.90 69.5

Inferred Mineral Resources 1,950,000 2.29 98.5

(1) Pressacco, R. (2009). Technical report on the initial mineral resource estimate for the Little Deer Deposit,

Newfoundland, Canada; Unpublished document available at www.SEDAR.com. Dated August 14, 2009;

86 pp.



In 2010 Scott Wilson Roscoe Postle Associates Inc. (“RPA”) updated the Mineral Resource

estimate for the Deposit (Pressacco, 2010). RPA estimated that the deposit contained Indicated

Mineral Resources of 1,150,500 t grading 2.79% Cu and Inferred Mineral Resources, comprised

of the Little Deer and Footwall zones (Table 6.3), of 2,335,500 t grading 2.06% Cu.

TABLE 6.3

SUMMARY OF RPA LITTLE DEER MINERAL RESOURCES AS OF SEPTEMBER 30, 20101



Little Deer Zone 1,150,500 2.79 70.8


Little Deer Zone 1,227,300 2.21 59.8



(1) Pressacco, R. (2010). Mineral resource update for the Little Deer Project. Unpublished memorandum

available at www.SEDAR.com. Dated September 30, 2010; 26 pp.

P&E has not independently verified the mineral resource estimates presented in Table 6.2 and

Table 6.3 and makes no assurances as to their validity or economic viability, in whole or in part.

It should be further noted that these mineral resource estimates have been superseded by the NI

43-101 compliant mineral resource estimate prepared by P&E in 2011 (Puritch and Ewert, 2011)

as presented in section 14.0 of this report and utilized in preparation of this PEA analysis.



Figure 6.1 Schematic Cross Section, Little Deer Deposit




7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 REGIONAL

The island of Newfoundland is underlain by a wide variety of Precambrian and Palaeozoic rocks

ranging from the Proterozoic to Carboniferous period. The area in the vicinity of Little Deer

property is underlain by the Lushs Bight Group which underlies the Springdale Peninsula,

Sunday Cove Island and part of Pilley‟s Island and the Southwest Arm Area, Notre Dame Bay,

northern Newfoundland. The Lushs Bight Group is comprised mainly of sheeted diabase dykes

and basaltic flows and minor pyroclastic and ultramafic rocks. Pillow basalts are further

subdivided based upon the presence of diabase dykes, pillow breccia, intercalated tuff,

amygdules and hematization.

The Lushs Bight Group is a part of the Paleozoic Central Mobile Belt of the Newfoundland

Appalachians‟. It lies within the Notre Dame Subzone of the Dunnage tectono-stratigraphic zone

(Figure 7.1) This zone is characterized by remnants of a series of Cambrian and Ordovician

island-arcs and back-arc basins that were successively accreted to the North American

(Laurentian) and Gondwanan continental margins during the Ordovician and Silurian.

According to Kean et al. (1995), lithogeochemistry work shows that portions of the Lushs Bight

Group (with major rock units comprised of sheeted dykes and pillow lavas forming part of an

ophiolite sequence) formed in a „suprasubduction zone‟ environment as an incipient island arc.

7.2 GEOLOGY OF THE LITTLE DEER PROPERTY

The Deposit is hosted in a typical ophiolitic sequence which underlies most of the Springdale

Peninsula. Similar ophiolite sequences are known to host volcanogenic massive sulphide

(“VMS”) and related deposits elsewhere in Newfoundland, including the former producing

mines at Little Bay, Whalesback, Betts Cove, Tilt Cove, Gullbridge and Rambler (Figure 7.1).



Figure 7.1 Simplified Geology and Location of Past-Producing Mines in Newfoundland

(Source: Mercator Geological Services, 2010)

The major host lithology consists of steeply dipping mafic metavolcanic rocks with few

continuous stratigraphic marker units relative to copper mineralization as is commonly found in

VMS deposits. Occurrences of agglomerates, tuffs and chert-rich units are observed in the drill

core, but sometimes are not found in adjacent drill holes suggesting such units have been

deposited in small isolated depressions.



7.3 MINERALIZATION OF THE LITTLE DEER DEPOSIT

The Deposit contains mainly stringer and disseminated sulphide mineralization, with lesser

amounts of massive sulphide mineralization, associated mainly with Upper Cambrian age mafic

volcanic rocks of the Lushs Bight Group. The predominant sulphides present are pyrrhotite,

chalcopyrite, pyrite and sphalerite. The copper mineralization outlined seems to be stratiform in

overall form and generally follows the orientation of the host mafic volcanic units.

The copper mineralization is manifested as narrow intervals of massive sulphide, wider intervals

of semi-massive sulphide (i.e.-sulphide-matrix breccia), stringers, veinlets and disseminations.

The mineralogy in copper rich areas resembles that found at the Little Deer mine and is a

mixture of chalcopyrite and pyrrhotite with occasional occurrences of sphalerite. As evidenced

by drill hole data, the copper rich mineralization is present in a series of discrete lenses and zones

that are oriented in an en echelon pattern.

The host rocks consist of chloritized and epidotized pillow basalts and an intermediate chlorite

schist zone. The schist zone ranges from chlorite schist through chlorite-sericite schist and

quartz-sericite schist to sericite schist. The dominant alteration mineral is chlorite, mainly an

iron-rich variety known as repidolite.

The host volcanic sequence is bounded by two faults – the Davis Pond Fault and the Middle Arm

– Clam Pond Fault. There are several small faults in the schist zone. It was previously thought

that the mineralization was controlled by faulting but it is now thought that the schists have been

preferentially sheared due to mechanical weaknesses in the volcanic pile (Figure 7.2).

Figure 7.2 Local Geology of the Little Deer and Whalesback Mine Area

(Source: Claims shown are after Pressacco 2009)



8.0 DEPOSIT TYPES

8.1 METALLOGENIC MODEL – VMS DEPOSITS

VMS deposits typically occur as lenses of polymetallic massive sulphide which form at or near

the seafloor in submarine volcanic environments. They are formed by the focused discharge of

metalliferous hydrothermal fluids associated with seafloor hydrothermal convection. The host

rocks can be either volcanic or sedimentary. VMS deposits are major sources of zinc, copper,

lead, silver and gold and significant sources of cobalt, tin, selenium, manganese, cadmium,

indium, bismuth, tellurium, gallium and germanium.

VMS deposits typically feature a tabular to mound-shaped stratabound body comprised,

principally, of massive (>40%) sulphide, quartz and subordinate phyllosilicates and iron oxide

minerals and altered silicate wall-rock. These stratabound bodies are typically underlain by

discordant to semi-discordant stockwork veins and disseminated sulphides. The stockwork vein

systems, sometimes referred to as „pipes‟ are surrounded by a distinctive alteration halo which

may extend into the hanging wall strata above the VMS deposit. (Figure 8.1)

Figure 8.1 Schematic Diagram of a VMS Deposit

(Source: Galley et al., 2007)

8.2 CYPRUS-TYPE VMS DEPOSITS

Cyprus type VMS deposits form on the sea floor and are related to the formation of oceanic crust

and spreading zones where new crust is being created. These types of deposits are related to



extrusive, often pillowed basalts, usually underlain by ultramafic intrusive and cumulate rocks

(Taylor et al., 1995).

Oceanic crust is continually being formed at spreading centres on the sea floor. These

environments are hot and dynamic where lavas and basalts are erupted onto the sea floor.

Hydrothermal solutions and vents, also known as „black smokers‟ are vents and fissures where

hot solutions erupt on to the sea floor forming crusts, mounds or plumes of solids which „rain‟

down on the sea floor. Black smokers are thought to be the modern equivalent of volcanogenic

massive sulphides in the geological record.

Ophiolitic rocks are sections of oceanic crust that have been uplifted and exposed above sea level

and often emplaced onto continental crustal rocks. Ophiolite VMS deposits are generally copper-

rich and comprised of two distinct components – the vertical to sub-vertical stringer zone of vein

and disseminated sulphide beneath – and the more massive sulphides as tabular, blankets lying

parallel to the sea floor.

8.3 LITTLE DEER DEPOSIT MODEL

Little Deer is a Cyprus-type VMS deposit that occurs within the Cambro-Ordovician Lushs

Bight Group sequence of ophiolitic intermediate to mafic volcanic rocks. The main sulphide

mineralization consists of disseminated, stringer, and semi-massive to massive pyrite, pyrrhotite

and chalcopyrite with minor sphalerite. The main copper-bearing horizon strikes at

approximately 075 and dips approximately 75° to the south. Eight similar Cyprus-type VMS

copper deposits occur in the region and are also hosted by the Lushs Bight Group. They have

reported resources, of which the past-producing Whalesback mine is amongst the largest, at

approximately 3.8 million tonnes grading approximately 1% Cu (Van Staal, 2007). Seventeen

such deposits are known in Cyprus with similar characteristics to those occurring in the Lushs

Bight Group with the largest being Mavrovouni at approximately 25 Mt.

The mineralogy of the deposit is predominantly copper with subsidiary cobalt and silver with

minor gold. Low to moderate zinc values are present, but the zinc is normally zoned away from

the copper. In this regard, the Deposit is closer to a Cyprus-type VMS deposit characterized by a

metal content that is usually restricted to copper, gold and, less commonly, zinc. Figure 8.2

presents one possible explanation for the presence of two copper stringer zones, whereby the

mineralization is deposited along paleo-volcanic listric normal faults.



Figure 8.2 Schematic Model Illustrating a Possible Explanation for Two Copper

Stringer Zones–Paleovolcanic Listric Normal Faults

(Source: Cornerstone Presentation, 2010)



9.0 EXPLORATION

9.1 RECENT EXPLORATION (2010-2011)

All exploration carried out on the Property prior to the 2010/2011 drilling program is

summarized in Section 6.0 of this report.

In 2011, Thundermin and Cornerstone completed a geological compilation of historical surface

and underground diamond drilling information dating back to the 1960‟s. This information was

obtained from the archives of the Newfoundland and Labrador Department of Natural Resources.

The conclusion of this compilation work was that there was potential to add significant resources

of high grade copper mineralization at shallower levels in the eastern portion of the deposit

above the -400 m elevation, particularly the -250 m elevation. This led to the drill program that

began in December 2010.



10.0 DRILLING

All drilling prior to the 2010/2011 drill program is summarized in section 6.0 of this report.

The aim of the 2010/2011 diamond drilling campaign was to increase the estimated mineral

resources outlined by RPA (Pressacco, 2010). The drilling focused on three main areas:

Above the -400 m elevation where historical drilling indicated a good potential

for outlining high grade resources in the eastern portion of the deposit, especially

above the -250 m elevation;

Along strike both east and west of the limits of the 2010 RPA resource outline

between the -650 m and -400 m elevations; and

At depth below the -650 elevation.

Two drills were utilized for the 2010/2011 drilling, with one drill testing the shallow portion of

the deposit and a second drill testing deeper targets.

Three holes (LD-10-39, LD-10-40 and LD-10-41) totalling 966 m were drilled in December

2010. These holes confirmed the high grade copper mineralization known to exist in the upper

portion of the deposit based on a review of historical data.

Twenty-two holes totalling 11,610 m were drilled between January and June 2011. Each

borehole intersected copper mineralization over varying widths. Hole LD-11-60 was abandoned

due to drilling difficulties. In total, twenty-five boreholes were advanced for a total of 12,576 m

of drilling.

A list of drillholes and significant intersections is provided in Table 10.1. Drillhole locations are

presented in Figure 10.1.

The results of the drill program are extracted from relevant news releases (Thundermin,

2010, 2011) and summarized in Table 10.1 below.

TABLE 10.1

HIGHLIGHTS OF DRILL INTERCEPTS FROM THE 2010/2011 DRILL PROGRAM

Hole No. East

(m)

North

(m)

Dip

(°)

Az

(°) From (m)

To

(m)

Interval

(m)*

Cu

(%)

LD-10-39 14,057 4,459 -37.1 321.6 208.6 209.1 0.5 13.4

and

213.9 218.1 4.2 4.6

and

233.9 250.4 16.5 5.0

incl.

233.9 239.0 5.1 6.1

incl.

244.9 250.4 5.5 9.2

LD-10-40 14,057 4,459 -35.8 315.0 294.5 295.2 0.7 2.4

LD-10-41 14,057 4,459 -36.1 335.1 202.6 203.0 0.4 5.1

and

219.2 222.2 3.0 2.1

and

229.7 235.6 5.9 4.5

LD-11-42 14,057 4,459 -63.0 305.5 306.8 308.0 1.2 1.0

LD-11-43 13,536 4,545 -56.5 331.3 No Significant

Values

LD-11-44 13,943 4,337 -48.1 318.8 413.8 415.4 1.6 9.3



TABLE 10.1

HIGHLIGHTS OF DRILL INTERCEPTS FROM THE 2010/2011 DRILL PROGRAM

Hole No. East

(m)

North

(m)

Dip

(°)

Az

(°) From (m)

To

(m)

Interval

(m)*

Cu

(%)

and

469.3 479.9 10.6 4.1

incl.

469.3 475.1 5.8 6.7

LD-11-45 13,536 4,545 -66.2 337.7 472.9 473.9 1.0 4.0

and

488.8 494.2 5.4 1.4


Values


Values

LD-11-48 13,536 4,545 -54.5 351.5 366.2 367.2 1.0 1.4

LD-11-49 13,943 4,337 -63.0 314.5 620.9 623.6 2.7 5.7

LD-11-50 13,749 4,530 -59.6 326.8 365.3 368.7 3.5 3.4

LD-11-51 13,749 4,530 -60.7 351.0 372.7 374.7 2.0 2.5

incl.

373.2 373.7 0.5 8.8

LD-11-52 13,943 4,337 -50.8 330.2 443.4 447.1 3.7 2.0

LD-09-

18A 13,518 4,133 -48.0 329.4

No Significant

Values

LD-11-53 13,817 4,277 -54.5 326.6 596.5 597.0 0.5 3.3

and

603.65 605.15 1.5 1.7

and

628.9 629.8 0.9 3.4

LD-11-54 13,754 4,228 -55.6 324.2 782.2 786.9 4.7 1.0

and

817.7 823.2 5.5 0.9

LD-11-55 13,517 4,131 -55.6 337.3 973.8 977.9 4.1 1.1

LD-11-56 13,754 4,228 -55.8 332.2 728.1 729.6 1.5 1.3


Values

LD-11-58 13,765 4,920 -42.0 154.8 149.85 150.45 0.6 2.5

and

173.0 175.9 2.9 3.5

LD-11-59 13,812 4,900 -44.7 134.2 178.0 178.95 0.95 3.0

and

185.1 191.1 6.0 2.1

incl.

189.0 190.1 1.1 8.6

LD-11-60 13,881 4,820 -42.4 100.4 Abandoned

LD-11-61 13,865 4,832 -40.2 99.4 86.7 87.3 0.6 1.2

LD-11-62 13,865 4,832 -40.5 116.8 73.4 74.8 1.4 2.2

and

87.0 89.7 2.7 1.6

*All indicated widths are core lengths



Figure 10.1 Drillhole Location

(Source: Thundermin Press Release, 2011)



11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

After logging of the core but prior to sampling, each interval to be sampled was subjected to a

number of procedures including accurate measurement of core angles, measurement of rock

quality designations (RQD), and photographing of both the wet and dry core. The geologist then

marked the sampling intervals to be submitted for analysis, assigned each interval a unique

sample tag in triplicate, noting the date, project, drill hole number, depth from, depth to and

sample width. Care was taken to ensure that the samples corresponded to either geological or

alteration intervals present in the core. Aside from some narrow intervals of fault gouge and

blocky core, no drilling, sampling or recovery factors were encountered that would materially

impact the accuracy and reliability of the analytical results. The drill core provided samples of

high quality, which were representative of any alteration, veining or sulphide accumulations that

were intersected by the drill hole. No factors were identified which may have resulted in a

sample bias.

After the intervals to be sampled were marked, the core was cut lengthways in half. One half of

the core sample was then placed in a plastic bag containing a sample tag for easy identification,

sealed and placed and further sealed in a container (fibre bag) for shipping to the assay lab. The

remaining half core was left in the core box for future reference.

Specific gravity measurements were made on all samples considered to represent a zone of

significant copper mineralization. In these cases, the specific gravities of all individually marked

samples were determined on the whole core sample by the core technician or geologist using the

Archimedes principle.

Once all the samples had been collected for a drill hole, they were transported under the direct

supervision of the geologist or core technician to the sample receiving facilities of Eastern

Analytical Ltd. Once all the samples for one drill hole had been cut, the remaining half core was

stored in a secure indoor location. A total of 541 samples of half cut drill core were taken during

the 2010/2011 drilling program at the Little Deer deposit.



12.0 DATA VERIFICATION

12.1 SITE VISIT AND INDEPENDENT SAMPLING

Mr. Eugene Puritch, P. Eng., visited the Property on May 16, 2011 for the purpose of doing the

site visit and completing an independent verification sampling program. The Little Deer core was

examined and 13 samples were taken in 11 holes by cutting ¼ splits of the remaining half core in

the box. An effort was made to sample a range of grades.

At no time were any employees of Thundermin or Cornerstone advised as to the identification of

the samples to be chosen during the visit.

The samples were selected by Mr. Puritch, and placed into sample bags which were sealed with

tape and placed in a rice bag. The samples were brought by Mr. Puritch to AGAT Laboratory,

(“AGAT”) in Mississauga, Ontario for analysis.

AGAT has developed and implemented at each of its locations a Quality Management System

(QMS) designed to ensure the production of consistently reliable data. The system covers all

laboratory activities and takes into consideration the requirements of ISO standards. AGAT

maintains ISO registrations and accreditations (ISO 9001 and ISO/IEC 17025).

Copper samples were digested using four acid and analyzed using atomic absorption

spectrometry (“AAS”) finish. Overlimits were run using peroxide fusion and AAS analysis.

A comparison of the results is presented in Figure 12.1.

Figure 12.1 Site Visit Sample Results for Copper

0

1

2

3

4

5

6

7

8

Cu (%)

Drill Hole

Little Deer ProjectSite Visit Sample Results for Copper

Original Cu %

AGAT result



12.2 QUALITY ASSURANCE/QUALITY CONTROL REVIEW

Thundermin and Cornerstone implemented a quality assurance/quality control (QAQC) program

for the drilling programs, with the addition of two different certified reference materials and a

pulverized blank material at a rate of approximately 1:20. In addition, 36 pulp samples were sent

to a secondary lab as verification on the principal lab. P&E reviewed all data as briefly discussed

in the following sub-sections.

12.2.1 Performance of Certified Reference Materials

Two certified reference materials were used for the drill programs, which were both purchased at

CAN Resource Laboratories Ltd. in Langley, BC. The one with the higher grade mean was

certified at 1.58% Cu, and the slightly lower grade reference material had a certified mean of

1.18% Cu.

There were 22 data points for the material grading 1.58% copper. The data were graphed, using

+/- 2 standard deviations from the mean for the warning limits and +/- 3 standard deviations

from the mean for the tolerance limits.

Two data points failed below the tolerance limit of -3 standard deviations. Six data points were

above the mean and the remaining 14 data points were all within -2 standard deviations.

The material grading 1.18% Cu had 81 data points. This standard performed very poorly with the

majority of the data points failing below -3 standard deviations. P&E examined the analysis

methods for the round robin characterization of the standards, as well as the method used at the

principal lab, in order to ascertain the possible source of error. The standards were characterized

using a four acid digest, while the principal lab used three acid. It is possible this is partly

responsible for the inaccuracy issues. The fact that the standards failed low is a cause for

concern, in that the resource grade may in fact be higher than estimated. This fact has been

discussed with Thundermin and Cornerstone, who are investigating this issue with the principal

laboratory.

12.3 PERFORMANCE OF BLANK MATERIAL

The blank material used was pre-pulverized and therefore did not go through the sample

reduction process – it monitored possible analytical contamination only. There were 82 data

points for the blank material and all were well below the upper threshold of three times the

detection limit.

12.3.1 Performance of Secondary Lab Checks

Thirty-six pulp samples were sent from Eastern Analytical to ALS Minerals of Vancouver for

verification purposes. The data correlation was excellent with all points falling on a 1:1 line, or

very close to it.

P&E declares the data acquired and analyzed by Thunderstone and Cornerstone to be satisfactory

for use in a resource estimate.



13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 INTRODUCTION

SGS Mineral Services of Lakefield, Ontario (“SGS”) were retained by Thundermin to conduct a

characterization and flotation test program on a composite sample from the Deposit (Imeson,

2010). The objectives of the initial metallurgical study were to examine the basic characteristics

of the material in terms of grindability and mineralogy. The study was also to conduct a scoping-

level flotation study to assign grade-recovery values to the test sample and to assess Co and

impurity levels in the concentrate.

13.2 MINERALOGY

The material graded 2.4% Cu which occurred almost exclusively as chalcopyrite. QEMSCAN

mineralogical characterization revealed that approximately 10.5% of the mass was iron

sulphides: pyrrhotite (85%) and pyrite (15%). The non-sulphides were mainly chlorites (51%),

quartz (15%) and plagioclase (7%). Liberation characteristics of the chalcopyrite indicated a

primary grind to be in the range of 150 microns. Regrinding to about 30-40 microns may be

necessary.

13.3 GRINDING

A Bond ball mill index of 13.2 kWh/T (14.6 kWh/t) was measured, indicating a material of

average hardness.

13.4 FLOTATION

Rougher flotation tests at a grind of 90 microns with a moderately elevated pH of 9–9.5 using

lime and isopropyl xanthate as collector yielded 99% recovery at a concentrate grade of 12% Cu,

indicating excellent performance. Somewhat coarser grinds (to at least 125 microns) may be

possible without significant copper loss.

Several regrind sizes and potential flowsheets were examined in batch cleaner tests. Although all

yielded satisfactory results, a regrind size of about 30 microns was indicated.

Locked cycle testing applying a standard rougher-cleaner circuit at primary grinds of 160 and

96 microns both yielded approximately 97% copper recovery and concentrate grades of 28% Cu.

The locked cycle testing did reveal some minor issues with pyrrhotite which was not stable

during the tests. This may impact recovery or concentrate grade.

A schematic of the locked cycle test flowsheet is presented in Figure 13.1.



Figure 13.1 Locked Cycle Test Flowsheet

A summary of the locked cycle test results are presented in Table 13.1.

TABLE 13.1

SUMMARY OF LOCKED CYCLE TEST RESULTS



14.0 MINERAL RESOURCE ESTIMATE

14.1 INTRODUCTION

The Mineral Resource Estimate prepared by P&E 2011 (Puritch and Ewert, 2011) and used as

the basis for the PEA presented in this report was prepared in accordance with the Canadian

Securities Administrators‟ National Instrument 43-101 and was estimated in conformity with

generally accepted CIM “Estimation of Mineral Resource and Mineral Reserves Best Practices”

guidelines using the commercially available Gemcom GEMS TM and Snowden Supervisor TM

software programs. All mineral resource estimation work reported by P&E for the Deposit was

carried out under the supervision of Mr. Eugene J. Puritch, P. Eng., an independent Qualified

Person according to NI43-101 regulations.

Mineral resources are not mineral reserves and do not have demonstrated economic viability. It

is not guaranteed that any part of the mineral resource will be converted into a mineral reserve.

14.2 DATA SUPPLIED

The database as received by P&E contained assay results from 48,432 m of drilling in

82 drillholes completed by Thundermin and Cornerstone since June 2007, and assay data from a

total of 102 surface and 122 underground historical holes that were drilled by BRINEX between

1961 and 1970 and Mutapa Gold Corporation between 1998 and 2000. The historical

information was recovered from the archives of the Newfoundland and Labrador Department of

Natural Resources in St. John‟s, Newfoundland and Labrador (Table 14.1).

TABLE 14.1

DRILLHOLE DATABASE SUMMARY

Type Number of Drillholes Total Metres

Historical Surface Drilling 102 23,546.42

Historical Underground Drilling 122 12,077.09

Current Surface Drilling 82 48,432.00

Total 306 84,055.51

14.3 DATABASE VALIDATION

Industry standard validation checks were completed on the supplied databases with no assay

entry errors detected. No significant validation errors were noted and P&E believes that the

supplied database is suitable for mineral resource estimation.

14.4 BULK DENSITY

The supplied database contained a total of 1,701 bulk density measurements. The supplied bulk

density measurements were used to estimate block density values (Table 14.2).



TABLE 14.2

BULK DENSITY VALUES

All Waste Little Deer Footwall Splay

Mean (t/m3) 3.00 2.98 3.05 2.99 3.06

Median (t/m3) 2.96 2.95 3.00 2.95 3.02

Mode (t/m3) 2.94 2.95 2.91 2.94 2.94

Standard Deviation 0.16 0.14 0.20 0.16 0.15

Sample Variance 0.03 0.02 0.04 0.02 0.02

Range 1.18 1.11 1.18 1.13 0.97

Minimum 2.72 2.73 2.72 2.72 2.81

Maximum 3.90 3.84 3.90 3.85 3.79

Count 1701 959 315 284 143

14.5 DOMAIN MODELING

Three mineralization domain models have been identified for the Deposit, named the Little Deer,

Footwall and the Little Deer Footwall Splay. Domain models were generated by P&E from cross

sectional polylines spaced every ten metres and oriented perpendicular to the trend of the

mineralization. The outlines of the polylines were determined by selecting Cu assay grades equal

to or greater than 1.0 % Cu with demonstrated continuity along strike and down dip, and include

low-grade material where necessary to maintain continuity between sections. The domains were

used for rock coding, statistical analysis and compositing limits (Figure 14.1).

Figure 14.1 Isometric Projection of Mineral Resource Domains



14.6 COMPOSITING AND COMPOSITE SUMMARY STATISTICS

Assay sample lengths ranged from 0.25 m to 5.30 m, with an average sample length of 1.37 m

(Figure 14.2). Two distinct sample length populations are evident, however, averaging 1.00 m

and 1.60 m. In order to ensure equal sample support a compositing length of 2.00 m was

therefore selected for use for mineral resource estimation.

Figure 14.2 Assay Sample Length Histogram

Length-weighted composites were calculated within the defined domains. The compositing

process started at the first point of intersection between the drillhole and the domain intersected,

and halted upon exit from the domain wireframe. Assays and composites were assigned a

domain rock code value based on the domain wireframe that the interval midpoint fell within.

The composite data were then exported to extraction files for grade estimation.

P&E generated summary statistics for 890 composite samples from the Little Deer domain,

164 composite samples from the Footwall domain, and 77 composite samples from the Splay

domain. P&E also computed multiple de-clustered means over a range of cell sizes in order to

provide accurate grades for model comparison and validation (Table 14.3).



TABLE 14.3

DOMAIN COMPOSITE SUMMARY STATISTICS

Total Little Deer Footwall Splay

Mean 1.71 1.73 1.37 2.16

Declustered Mean 1.65 1.57 1.85 2.24

CV 1.22 1.25 1.12 0.89

Median 1.04 0.98 1.04 1.81

Standard Deviation 2.08 2.17 1.54 1.92

Sample Variance 4.33 4.72 2.36 3.70

Kurtosis 6.86 6.66 11.06 1.20

Skewness 2.25 2.25 2.68 1.22

Range 17.28 17.28 11.15 8.51

Minimum 0.001 0.001 0.001 0.001

Maximum 17.28 17.28 11.15 8.51

Count 1131 890 164 77

14.7 TREATMENT OF EXTREME VALUES

The presence of high-grade outliers for the composite data was evaluated by a combination of

decile analysis and review of probability plots. Decile analysis results indicate that minimal

capping is required, with 20% of the mineral content contained in the upper decile and 6% in the

upper percentile for Cu. (Figure 14.3). One composite grade was capped to the selected 15% Cu

threshold value prior to estimation.

Figure 14.3 Decile Analysis Results

14.8 CONTINUITY ANALYSIS

Domain-coded, composited sample data were used for continuity analysis. Strike orientations for

the domains were modeled using the known geometry of the mineralization. Dip and dip plane

orientations were modeled using orientations developed from variogram fans, which were



assessed for geological reasonableness. Normal-scores experimental semi-variograms aligned

with the best-fit orientation of the mineralization were then generated (Figure 14.4). Variogram

model ranges were checked and iteratively refined for each model. The nugget effect for each

vein was derived from the down hole experimental semi-variogram. Rotation is defined by the

Gemcom ADA convention in the defined block model space, and the variance contributions were

back-transformed and checked relative to the mineralization.

Based on the analysis of the resulting experimental semi-variograms, a strike range of 40.0 m, a

dip range of 40.0 m, and a cross-strike range of 10.0 m was selected as appropriate for mineral

resource estimation. Continuity ellipses based on the observed ranges were then generated and

used as the basis for estimation search ranges, distance calculations and mineral resource

classification criteria. Anisotropy was modeled with an average strike azimuth of 260°, -80S°

down dip on an azimuth of 170o and +10S° across strike on an azimuth of 170°.

Figure 14.4 Experimental Semi-Variograms

14.9 BLOCK MODEL

A rotated block model was established across the Deposit with the block model limits selected so

as to cover the extent of the mineralized domains, and the block size reflecting the generally



narrow widths of the mineralized zones and the drill hole spacing (Table 14.4). The block model

consists of separate models for estimated grade, rock code, percent, density and classification

attributes. A percent block model was used to accurately represent the volume and tonnage that

was contained within the constraining grade domains. As a result, the mineral resource

boundaries were properly represented by the percent model‟s capacity to measure infinitely

variable inclusion percentages. The volume represented by the historical underground workings

was subsequently depleted from the model.

TABLE 14.4

BLOCK MODEL SETUP

Dimension Origin Number of Blocks Block Size (m)

X 569,900 360 5

Y 5,492,200 120 5

Z -1000 240 5

Rotation 30° CCW

14.10 RESOURCE ESTIMATION & CLASSIFICATION

Block bulk density values were calculated using a single pass. Anisotropic inverse distance

squared (“ID2”) linear weighting of between three and six bulk density values was used for the

estimation of individual block bulk density values.

Ordinary Kriging (“OK”) of capped composite values was used for the estimation of block

grades, with the anisotropy defined by the axes of the search ellipse. A two-pass series of

expanding search volumes with varying minimum sample requirements was used for sample

selection, grade estimation and classification. Composite data used during grade estimation were

restricted to samples located within their respective domains.

During the first pass, three to six composites from two or more drillholes within a search

ellipsoid of 40.0 m x 40.0 m x 10.0 m were required for grade block estimation.

During the second pass, three to six composites from one or more drillholes were required for

grade block estimation. The search ellipse was expanded to ensure that all blocks within the

defined mineralization domains were estimated.

Mineral resources were classified in accordance with guidelines established by the CIM:

Mineral resource classification was implemented by generating three-dimensional envelopes

around those parts of the block model for which the drillhole data and grade estimates met the

required continuity criteria. The resulting classifications were iteratively refined until they were

geologically reasonable in order to prevent the generation of small, discontinuous areas of a

higher confidence category being separated by a larger area of a lower confidence areas.

Indicated resources were defined based on the results of the first pass, and then consolidated

into an envelope digitized around the central area of blocks estimated during this pass. This

process downgraded scattered and isolated higher confidence blocks and combined Indicated

mineral resources into a continuous unit, and upgrade scattered and isolated Inferred mineral

resources surrounded by higher confidence blocks. All remaining blocks estimated were

classified as Inferred, including all blocks in the Footwall and Splay domains (Figure 14.5).



Figure 14.5 Isometric Projection of Block Classification

14.11 MINERAL RESOURCE ESTIMATE

The mineral resource estimate for the Little Deer deposit is reported at a cut-off grade of 1.0 %

Cu (Table 14.5), with an effective date of June 18, 2011.

TABLE 14.5

SUMMARY OF LITTLE DEER MINERAL RESOURCES(1)(2)(3)(4)(5)(6)(7)



Little Deer Zone 1,911,000 2.37 99.8


Little Deer Zone 1,240,000 1.93 52.8


Little Deer Footwall Zone Splay 797,000 2.64 46.2


(1) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The

estimate of mineral resources may be materially affected by environmental, permitting, legal, title, taxation,

socio-political, marketing, or other relevant issues.

(2) The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there

has been insufficient exploration to define these Inferred resources as an Indicated or Measured mineral

resource and it is uncertain if further exploration will result in upgrading them to an Indicated or

Measured mineral resource category.

(3) The mineral resources in this press release were estimated using the Canadian Institute of Mining,

Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and

Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.

(4) Ordinary Kriging was used for Cu grade interpolation.

(5) Grade capping of 15% Cu utilized on composites.

(6) A variable bulk density based on numerous field measurements was used for tonnage calculations.

(7) A copper price of US$3.42/lb. (May 31 2011 24 month trailing average) and an exchange rate of

US$0.95US=C$1.00 was utilized to derive the 1% Cu cut-off grade. Mining costs were C$40/t, process

costs were C$15/t and G&A was C$5/t. Concentrate freight and smelter treatment charges were C$10/t

mined. Concentrate mass pull was 7%, process recovery was 97%, smelter payable was 96% and Cu

refining was US$0.07/lb.

View looking east.

Blue: Inferred

Green: Indicated



14.12 VALIDATION

The block model was validated visually by the inspection of successive section lines in order to

confirm that the block model correctly reflects the distribution of high-grade and low-grade

samples. Local trends were evaluated by comparing the OK block estimates to a nearest

neighbour estimate (“NN”) at zero cut-off along the strike of the Deposit (Figure 14.6). In

general the OK block estimates are in good agreement with the NN estimates, and demonstrate

no evidence of systematic bias in the model.

Figure 14.6 Little Deer Domain Swath Plot

As a further check on the model the average model block grade was compared to the NN block

average as well as the de-clustered mean and the average of the composite data. No significant

bias between the block model and the input data was noted (Table 14.6).

TABLE 14.6

DOMAIN VALIDATION STATISTICS

Domain Model Average

Cu %

NN Average

Cu %

Declustered

Mean Cu %

Composite

Average Cu %

Little Deer 1.65 1.65 1.57 1.86

Footwall 1.77 2.26 1.85 1.54

Splay 2.44 2.35 2.24 2.16

Total 1.90 1.90 1.65 1.84

14.13 POTENTIALLY MINEABLE MINERAL RESOURCE ESTIMATE

A potentially mineable portion of these mineral resources was determined as a basis for a PEA of

the Deposit. Dilution and mining losses were considered in mineable mineral resource

calculations. The results of this determination are provided below. P&E cautions that these

mineable mineral resources are not mineral reserves and do not have demonstrated economic

viability.



A longhole longitudinal retreat mining method was selected as the most likely mining method

that would be applied on the Deposit. Starting at the 80 m (below surface) elevation, mining

levels were envisaged every 30 m vertically. Based on this mining method, P&E determined a

minimum mining width of 2.0 m. This excluded an estimated 6,560 t of mineral resources that

are in areas less than 2.0 m wide. A summary of mineral resources that are considered to be

potentially mineable is presented in Table 14.7.

TABLE 14.7

POTENTIALLY MINEABLE MINERAL RESOURCES (BEFORE RECOVERY AND DILUTION)

Level Ore Tonnes Au (g/t) Ag (g/t) Co (%) Cu (%)

80 3,896 0.068 1.857 0.016 1.473

50 57,775 0.070 2.143 0.017 1.727

20 119,558 0.068 2.032 0.016 1.640

-10 172,402 0.084 2.695 0.020 2.209

-40 170,647 0.093 2.957 0.021 2.437

-70 131,635 0.083 2.779 0.020 2.281

-100 66,299 0.067 2.039 0.016 1.640

-130 116,471 0.077 2.832 0.020 2.315

-160 125,390 0.082 3.218 0.022 2.637

-190 147,752 0.116 3.258 0.022 2.446

-220 158,295 0.147 2.629 0.020 2.194

-250 164,122 0.108 2.764 0.022 2.477

-280 147,600 0.093 3.231 0.024 2.794

-310 104,899 0.062 3.112 0.025 2.600

-340 140,665 0.049 2.676 0.021 2.160

-370 181,343 0.047 2.327 0.020 2.032

-400 192,916 0.050 2.352 0.022 2.192

-430 228,385 0.048 2.452 0.023 2.221

-460 291,815 0.044 2.418 0.023 2.412

-490 341,753 0.041 2.317 0.024 2.464

-520 352,652 0.038 2.329 0.024 2.489

-550 328,385 0.035 2.111 0.024 2.408

-580 328,656 0.039 1.987 0.024 2.346

-610 289,207 0.044 1.794 0.024 2.126

-640 215,447 0.053 2.046 0.027 2.140

-670 159,733 0.059 2.305 0.027 2.153

-700 135,278 0.028 1.295 0.020 1.907

-730 68,428 0.016 0.903 0.020 1.555

-760 65,650 0.009 0.781 0.019 1.469

-790 117,803 0.007 0.630 0.019 1.353

-820 135,117 0.014 1.286 0.021 1.585

-850 124,512 0.024 2.233 0.025 1.884

-880 94,771 0.027 1.410 0.025 2.006

-910 70,019 0.043 2.408 0.024 2.018

-940 54,141 0.056 2.316 0.023 1.922

-970 35,999 0.073 2.018 0.021 1.701

-1000 13,083 0.078 2.395 0.020 1.693



TABLE 14.7

POTENTIALLY MINEABLE MINERAL RESOURCES (BEFORE RECOVERY AND DILUTION)

Level Ore Tonnes Au (g/t) Ag (g/t) Co (%) Cu (%)

Total 5,652,499 0.055 2.279 0.022 2.212

The longhole longitudinal retreat mining method is estimated to experience mining dilution in

the order of 20% at zero grade. Mine recovery (extraction) is estimated to be 90%. A summary

of Potentially Mineable Mineral Resources, including dilution and recovery, is presented in

Table 14.8.

TABLE 14.8

RESOURCE SUMMARY

Potentially Mineable Mineral Resources

Description Resource Au Ag Co Cu

Tonnes g/t g/t % %

Mineral Resources Included 5,652,500 0.055 2.279 0.022 2.212

Diluted Mineral Resources 6,783,000 0.046 1.899 0.019 1.843

Total Potentially Mineable Mineral Resources 6,104,699 0.046 1.899 0.019 1.843



15.0 MINERAL RESERVE ESTIMATES

The inferred mineral resources presented herein have not been sufficiently drilled to confidently

demonstrate economic viability. In addition, the work undertaken on the Little Deer Project to

date is considered to be at conceptual levels of study only. As such, and according to the NI 43-

101 Regulations, it is not possible to declare a mineral reserve of any kind as of the effective date

of this report.



16.0 MINING METHODS

The Little Deer Potentially Mineable Mineral Resource extends from the +110 m elevation to the

-1,000 m elevation; a vertical distance of 1,110 m. A conceptualized mining plan has been

developed to extract the deposit using mechanized trackless mining equipment.

The deposit would be accessed initially by a ramp from surface with a grade of -15%. The ramp

portal elevation would be at 126 m. The ramp would ultimately extend to a vertical depth of

1,126 m (to the -1000 m elevation). Due to the depth of the deposit, a vertical shaft will also be

constructed. The proposed shaft collar elevation would be at +145 m. The shaft would be sunk

conventionally to a vertical depth of approximately 875 m (to the -730 m elevation). The upper

part of the mine would be serviced by both the shaft and the ramp. The bottom section of the

mine, from the -1000 m to the -670 m level, would be serviced only by the ramp. Shaft sinking

would be carried out independently of the ramp and mine development.

The irregularity of the deposit along strike requires that a non-captive cut and fill mining method

would be the preferred mining method. To minimize the capital costs required to develop the

mine, a longhole longitudinal retreat mining method has been selected.

16.1 LONGHOLE LONGITUDINAL RETREAT MINING METHOD

The mining method selected is Longhole Longitudinal Retreat. Drilling, blasting and mucking

sublevels would be driven every 30 vertical metres. Cross-cuts averaging 27 metres in length

would be developed to the deposit from footwall drifts. Individual “stopes” would average 92 m

long by 30 m high by 4m wide. A slot/ventilation raise would be driven at the extremity of each

stope and on average every 25 m along the stope, between the sublevels. Successive rows of

drillholes will be blasted into the slot and open stope. Cemented hydraulic tailings backfill and

development waste would be placed in the stopes as they retreat. The Life-of-Mine (“LOM”)

schedule includes 208 stopes which would be mined at a rate of 1,800 tpd ore. Typically, this

corresponds to mining three sublevels concurrently (i.e. 600 tpd / level).

16.2 MINE AND STOPE DEVELOPMENT

All excavations in waste rock are classified as mine development. All development that directly

produces feed to the mill, is classified as stope development. The LOM schedule includes a total

of 22,826 m of mine development (see Table 16.1) plus an additional 6,976 cubic metres of shaft

station and loading pocket development.



TABLE 16.1

SUMMARY OF ESTIMATED MINE AND STOPE DEVELOPMENT

Level

FW

Drift

(m)

X-cut

to Vein

(m)

FW

Dr. to

Ramp

(m)

Ramp

(m)

Shaft to

FW Dr.

(m)

X-Cut to

Shaft

Orepass

(m)

Shaft

Ore

Pass

(m)

X-Cut to

Orepass

(m)

Internal

Ore Pass

(m)

X-Cut to

Vent Rse

(m)

80 191 55 15 312 10

50 193 56 20 202 10

20 198 54 28 202 10

-10 191 62 28 202 10

-40 192 72 35 202 10

-70 145 41 35 202 166 10 10

-100 154 44 38 202 10

-130 164 65 33 202 5 31 10

-160 156 58 33 202 105 10 87 5 31 10

-190 142 62 37 202 10

-220 141 59 36 202 5 31 10

-250 138 61 37 202 123 10 87 5 31 10

-280 99 45 33 202 10

-310 117 56 33 202 5 31 10

-340 164 72 29 202 117 10 87 5 31 10

-370 181 71 31 202 10

-400 270 85 34 202 10 62 10

-430 262 120 35 202 136 10 87 10 62 10

-460 286 186 35 202 10

-490 261 205 30 202 10 62 10

-520 262 178 32 202 88 10 87 10 62 10

-550 284 215 28 202 10

-580 243 176 28 202 10 62 10

-610 251 189 18 202 103 10 87 10 62 10

-640 199 118 16 202 10

-670 200 59 19 202 107 10 62 5 31 10

-700 134 60 24 202 32 10

-730 143 27 28 202 10

-760 132 47 30 202 10

-790 111 42 26 202 10

-820 110 59 28 202 10

-850 101 51 27 202 10

-880 74 36 31 202 10

-910 79 37 29 202 10

-940 62 31 30 202 10

-970 36 17 28 202 10

-1000 9 5 30 202 10

Total 6,077 2,879 1,087 7,595 945 80 615 95 591 370

Notes: In addition there will be 1,217m of exhaust raises, 875 m of shaft and 400m of ramp connecting to the spill

handling pocket in the shaft.

Some values have been rounded. The totals are accurate summations of the columns of data.

There is a total of 24,688 m of stope development required over the LOM.

16.3 STOPING

Using the Longhole Longitudinal Retreat stoping method, drifts would be developed to the full

width of the deposit every 30 vertical metres (“undercuts” and “overcuts”) from the access cross-

cuts. A 2.4 m by 2.4 m slot / ventilation / backfill raise would be excavated between sublevels at



25 m intervals along the strike of the deposit. This would include the initial slot / ventilation raise

at the end of the deposit and several raises for re-slotting, as well as access and ventilation.

Blastholes measuring 5.1 cm (2 inches) in diameter would be drilled from the sublevel either up

or down to adjacent sublevels. These blastholes would typically be drilled on a 1m by 1m

pattern, in order to break the rock into the open slot and stope. The blasting powder factor using

emulsion explosives is estimated to be 0.6 kg/t. An estimated 1,477 tonnes of mill feed would

need to be excavated on a daily basis from a combination of stopes. Stope development activities

would add another 323 tonnes mill feed to the total, to provide a combined 1,800 tpd of mill

feed. A summary of stope drilling and blasting para metres is presented in Table 16.2.

TABLE 16.2

STOPING DRILLING AND BLASTING PARA METRES

Tonnes of Mill Feed per Day 1,800

Stope Tonnes per Day 1,477

Mineralization Specific Gravity (t/m3) 3.00

Stope Tonnes per 25 m of Strike Length 8,661

Slot Raise Tonnes per Raise 441

Undercut Tonnes per 25 m of Strike Length 1,299

Longhole Retreat Stoping 6,921

Longhole Drilling Para metres @ 2' Dia Holes

Total Drilling Per Stope (m) 2,307

Drillholes Per Stope 90

Drilling Time Per Shift 10

Metres Drilled per Shift 76

Total Metres Drilled Per Day 152

Required Metres per Day 492

Blasting Para metres

Stemming Length Per Blasted Hole Length (m) 6.6

Load Length per Hole, (m) 18.9

Length of Holes Loaded Per Ring (m) 73

Stope mining would commence from the top of the mine and progress downwards in successive

three sublevel increments or “stoping blocks”. The initial stoping blocks would include the

blocks above the -10 m, -100 m and -190 m levels, where undercut drifts would typically be

established. Each stoping block would also have overcut drifts established 30 metres and

60 metres above the undercut drift. For example, in the case of the stoping block between the -10

m and -100 m elevations, overcut drifts would be developed at the -70 m and -40 m elevations.

To maintain access for backfilling the upper 30 m of the stoping block (for example at the -10 m

elevation in the -10 m to -100 m stoping block) a 2.4m by 2.4m drift will be driven at the backfill

/ ore contact before the upper 30 m of the stoping block is mined.

The stope mining cycle would include longhole drilling, blasting, mucking and backfilling. The

overall stope mining productivity is estimated to be 150 tpd per stope. At any given time three

levels should be available for stope mining, each with a maximum of four stopes. On average

this would provide for an average production rate of 600 tpd per level and 1,800 tpd overall.



A summary of stoping productivities is presented in Table 16.3.

TABLE 16.3

THUNDERMIN / CORNERSTONE RESOURCES INC. STOPING PRODUCTIVITIES (TPD)

Operation Productivity

Drilling 457

Blasting 914

Mucking 457

Backfill 914

Average Stope Productivity 150

Average Productivity per Level 600

Maximum Number of Stopes / Level 4

16.4 SCHEDULE

Underground access would be provided by a ramp, driven at - 15% to the -1,000 m elevation and

an 875 m deep production shaft. It is envisioned that production from the upper 400 m of the

deposit would be provided through the ramp, while production from the lower part of the deposit

would be serviced through the production shaft. The ramp would be developed in the six stages:

Stage 1: From surface to the -190 level;

Stage 2: From the -190 to the -460 level;



Stage 5: From the -820 to the -910 level; and

Stage 6: from the -910 to the -1000 level.

The shaft and main ramp are scheduled to be driven independently. All excavations in waste rock

are classified as mine development. All development that directly produces feed to the mill, is

classified as stope development. A summary of yearly LOM development is presented in Table

16.4.



TABLE 16.4

MINE DEVELOPMENT SUMMARY

Description Units Total Yr 1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Ramp Portal m 50 50

Ramp m 7,946 832 1,251 1,944 277 1,393 428 977 844

Shaft m 875 27 500 348

Ramp

Crosscuts m 1,087 28 226 112 101 148 80 56 47 58 173 58

Footwall

Drifts m 6,077 155 1,216 634 588 977 506 465 563 371 505 97

Crosscuts m 4,369 533 1,306 360 447 297 445 341 294 270 76

Orepasses m 1,206 31 646 63 125 62 124 93 62

Ventilation

Raises m 1,217 299 153 128 121 260 26 77 79 74

Total m 22,827 1092 4056 5143 1517 3211 1633 1090 1070 1839 1871 305

Shaft

Stations m

3 6,676 3,338 3,338

Loading

Pocket m

3 300 300

Total m3 6,976 3,338 3,638

A summary of yearly LOM stope development is presented in Table 16.5.

TABLE 16.5

STOPE DEVELOPMENT SUMMARY

Description/Yr Units Total Yr 1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Stope Drift m 19,384 0 3,407 2,320 2,083 1,724 2,277 2,194 1,927 1,626 1,184 640

Slot Raise m 5,304 0 612 838 619 603 621 582 408 306 408 306

Total m 24,688 0 4,019 3,158 2,702 2,327 2,899 2,777 2,335 1,932 1,592 946



17.0 RECOVERY METHODS

A summary of available metallurgical testwork is presented in Section 13. Based on these data, a

conventional process flowsheet is selected, including crushing and grinding to a 90 micron grind

at a rate of 1,800 tpd, followed by flotation recovery of copper to a rougher concentrate. The

rougher concentrate would be reground to minus 30 microns and cleaned in a three stage

flotation circuit to yield a final concentrate containing copper at a marketable grade. The

concentrate would be filtered to an assumed 8% moisture content for shipment. Power

requirements for the milling process are estimated to be approximately 28 kWh/t.



18.0 PROJECT INFRASTRUCTURE

The Little Deer Project has minimal infrastructure requirements due to its location close to the

Trans-Canada Highway and due to the infrastructure established during its previous operating

history. Figure 18.1 shows the infrastructure in Little Deer area.

18.1 SITE SURFACE INFRASTRUCTURE

Site surface infrastructure includes site facilities, buildings, buildings furnishings and surface

mobile equipment. The site facilities include; the hydraulic tailings backfill plant and distribution

system; the tailings / waste rock co-disposal basin and dam; site roads; surface parking areas; the

fuel farm; lubrication and oil storage facilities; surface explosive magazines; yard piping; the fire

prevention and fighting system; the potable water treatment plant and storage tanks; the tailings

water treatment plant and pond and the water management pond building and site run-off. The

site buildings include; the main gate building; the surface mine shop; the warehouse and

warehouse equipment; the office trailers and the dry. Furnishings include; the surface mine shop

equipment and tools; the office furniture, computers, etc.; environmental equipment; dry

equipment; site communications and medical center equipment. Surface mobile equipment

includes; a road / ramp grader; an integrated tool carrier; a fuel/lube truck; a service truck; a

garbage truck; a personnel bus; an ambulance; a fire/ rescue truck and pickup trucks.

Power Supply

Electric power for the Little Deer Project will come from the provincial electrical substation

located just outside Springdale on Highway 392. The transmission line will be installed 8 km

along the paved Little Bay Road and then 2 km on a gravel road to site.

Tailings Management

A tailings storage strategy would be developed based on an assessment of the existing tailings

dam and other potential storage sites.

Waste Management

The waste rock dump(s) will be designed, built and closed out so as to minimize long-term

impact on the environment. Other waste materials will be recycled (e.g. spent lubricants) or

disposed of in accordance with provincial and federal regulations.

Hazardous Material Storage

Storage facilities for materials such as fuel, explosives and process chemicals have not been

detailed at this scoping study level. As the project proceeds, such facilities will be designed to

meet all relevant codes and regulations in order to protect employees, the public and the

environment.



Regional Resources

With the probable exception of an underground mining contractor, Thundermin and Cornerstone

should not have to go beyond the Province for any supplies or services. The regional labour force

includes experienced equipment operators and mine workers.

Figure 18.1 Little Deer Copper Deposit Infrastructure



19.0 MARKET STUDIES AND CONTRACTS

There were no market studies completed or contracts in place in support of this Technical

Report. The only commercial product produced by the project is copper concentrate.

Thundermin / Cornerstone will be paid once the copper concentrate has been delivered to the

smelter.



20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY

IMPACT

The Little Deer Project (“Project”) is located approximately 10 km north of Springdale in a

historic mining district characterized by moderately rugged topography, exposed bedrock,

isolated vegetated areas, ponds and historically disturbed land areas including a historic tailings

basin. The proposed Project includes:

An underground mine with a shaft and ramp along with ventilation and fill raises

to surface. Stopes would be developed below the 100 m level and backfilled using

a combination of cemented tailings and rock fill. The backfilling would be

primarily done for stope stability purposes and would reduce the amount of waste

rock and tailings requiring disposal on surface.

A 1,800 tpd throughput capacity mill with covered concentrate storage and load-

out areas. The concentrate would be trucked to a marine port.

A tailings management area that would be constructed in a historic tailings basin

that is largely confined by the natural topography. The new tailings would be

submerged underwater to inhibit oxidation and acid rock drainage. Water would

be recycled from the tailings pond to the mill. A water treatment plant and settling

and polishing ponds would be developed downstream of the tailings management

area. Water would be treated before it is released to the environment.

Surface infrastructure including shop, warehouse, offices, dry and ancillary

facilities. A power transmission line would be developed from the Springdale

substation to the site.

The environmental assessment (“EA”) and permitting process for mining and mineral processing

projects in Newfoundland and Labrador is well-established and is harmonized with the Federal

EA process requirements. The Project would require an environmental assessment with public

consultation under the Environmental Protection Act (SNL 2002 c.E-14.2). Once approved, the

Project would require a Certificate of Approval under s.78 of the Environmental Protection Act,

water rights under the Water Resources Act (SNL2002 c.W-4.01), a mill license, and other

operating permits.

Thundermin plans to further assess its environmental and social base line study, carry out public

consultation and determine permitting requirements with regulatory authorities.

The proposed Project would be developed, operated and closed in accordance with

environmental and health and safety regulatory requirements. It is expected that engineered

controls such as, but not limited to, double walled fuel storage tanks and spill response

procedures to eliminate or mitigate environmental risks, would be incorporated into the detailed

design of the project. Waste material management procedures would be in place and waste

would be disposed of in accordance with regulatory requirements.

Progressive mine rehabilitation and closure is required by Provincial legislation. The Mining

Regulations (sections 4 to 7) under the Mining Act (SNL 1999 c.M-15.1) require the mine

operator to develop and submit a Development Plan, Operational Plan, Rehabilitation and

Closure Plan and Annual Reports. Financial Assurance for relevant costs including ongoing

monitoring and site maintenance is required under Section 8 of the Mining Regulations. The

envisaged closure works include the decommissioning of the proposed underground mine. The

underground mine equipment and stored fuel and lubricants would be returned to surface and



disposed of in accordance with regulatory requirements. Mine openings to surface (i.e. shaft

collar, ramp portal and raise collars) would be sealed. The hoists, conveyances and headframe

would be sold and removed from site. The mill would be demolished and major equipment

salvaged. The projected closure costs for the Project are summarized in Section 21.1.4 “Closure

Bond”.



21.0 CAPITAL AND OPERATING COSTS

21.1 CAPITAL COST ESTIMATES

Capital costs include the cost of mine and stope development; shaft equipment and related

facilities; underground mining equipment; processing plant and related facilities; surface mobile

equipment; electrical power supply infrastructure; the purchase of the existing royalties; and the

project closure bonds. A summary of the Capital Cost Estimates for the Project is provided in

Table 21.1.

TABLE 21.1

CAPITAL COST SCHEDULE AND SUMMARY (M CAN$’S)

Preproduction Preproduction

Description Total Yr -2 Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Mine

Development 120.4 0.0 6.5 32.7 31.5 5.7 13.2 6.4 3.6 3.6 7.9 8.0 1.2

Stope

Development 75.8 0.0 0.0 12.8 9.4 8.2 6.9 8.9 8.6 7.4 6.2 4.8 2.7

Shaft Related

Equip 13.0 0.0 6.5 6.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Mine

Equipment 19.9 0.0 11.9 8.1 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Misc. U/G 1.0 0.0 0.3 0.5 0.3 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Processing

Plant 45.3 15.1 30.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Surface

Infrastructure 18.0 0.0 18.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Closure

Bond 6.0 1.5 4.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Powerline

Construction 2.0 0.0 2.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Purchase

Royalties 2.0 0.0 2.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0

Total 303.4 16.6 81.9 60.6 41.1 13.9 20.2 15.3 12.2 11.0 14.0 12.7 3.9


Details of these estimates are provided in the following subsections.

21.1.1 Mine Capital Costs

Mine capital costs include the cost of underground waste rock excavations, excavations in the

stoping area to open up the stopes for mining and fixed and mobile mining equipment costs.

Mine development includes all underground development excavations in waste rock. This

includes: the main access ramp; the shaft and all related shaft facilities and excavations; x-cuts to

the stoping areas; footwall drifts; ore passes and ventilation raises. A summary of mine

development capital costs and schedule is presented in Table 21.2.



TABLE 21.2

MINE DEVELOPMENT CAPITAL COST SCHEDULE AND SUMMARY (M CAN$’S)

Description Total Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Ramp Portal 0.8 0.8

Ramp 39.7 4.2 6.3 9.7 1.4 7.0 2.1 4.9 4.2

Shaft 30.6 1.0 17.5 12.2

Shaft

Stations 1.3 0.7 0.7

Loading

Pocket 0.1 0.1

Ramp

Crosscuts 4.3 0.1 0.9 0.4 0.4 0.6 0.3 0.2 0.2 0.2 0.7 0.2

Footwall

Drifts 21.3 0.5 4.3 2.2 2.1 3.4 1.8 1.6 2.0 1.3 1.8 0.3

Crosscuts 15.6 2 5 1 2 1 2 1 1 1 0

Orepass 1.8 0.0 1.0 0.1 0.2 0.1 0.2 0.1 0.1

Ventilation

Raises 4.9 1.2 0.6 0.5 0.5 1.0 0.1 0.3 0.3 0.3

Total 120.4 6.5 32.7 31.5 5.7 13.2 6.4 3.6 3.6 7.9 8.0 1.2


All development that directly produces feed to the mill is classified as stope development. This

includes: developing undercuts and overcuts in the stopes and slot / ventilation raises between

sublevels. A summary of stope development capital costs and schedule is presented in Table

21.3.

TABLE 21.3

STOPE DEVELOPMENT CAPITAL COST SCHEDULE AND SUMMARY (M CAN$’S)

Description Total Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Stope Drift 67.8 0.0 11.9 8.1 7.3 6.0 8.0 7.7 6.7 5.7 4.1 2.2

Slot Raise 8.0 0.0 0.9 1.3 0.9 0.9 0.9 0.9 0.6 0.5 0.6 0.5

Total 75.8 0.0 12.8 9.4 8.2 6.9 8.9 8.6 7.4 6.2 4.8 2.7

* Note: Some values have been rounded. The totals are accurate summations of the columns of data.

The mine equipment capital costs include: the shaft headframe, hoist room, hoist, loading pocket

and grizzly/rockbreaker infrastructure; all underground mobile and stationary equipment; all

related mine surface equipment; and the required underground infrastructure such as lunch

rooms, explosive magazines, etc. A summary the underground mine equipment capital costs, and

schedule of purchases, is presented in Table 21.4.



TABLE 21.4

UNDERGROUND EQUIPMENT CAPITAL COST SUMMARY (M CAN$’S)

Description Units Total Yr -1 Yr 1 Yr 2

Shaft Related Equipment

Headframe, Hoist Room, Hoists(2) Lot 12.0 6.5 5.5

Loading Pocket 1 0.5 0.0 0.5

Grizzly / Rockbreaker 2 0.5 0.0 0.5

Shaft Related Equipment Total 13.0 6.5 6.5

Mine Equipment

Development Jumbo - 2 Boom 2 2.0 1.0 1.0

Longhole ITH Drill 2 1.5 0.8 0.8

LHD - 6.1 cubic metres 2 2.6 1.3 1.3

Haul Trucks - 50t 3 4.0 1.3 2.7

Blasting Tractor 1 0.6 0.0 0.6

ANFO Loader 1 0.4 0.4 0.0

Cable Bolter 1 0.8 0.0 0.8

Lube Service Vehicle 1 0.3 0.3 0.0

Fuel truck 1 0.4 0.4 0.0

Personnel Vehicle – Mechanical 1 0.1 0.1 0.0

Personnel Vehicle – Electrical 1 0.1 0.1 0.0

Boom Truck 1 0.3 0.3 0.0

Grader 1 0.4 0.4 0.0

Tractors 3 0.2 0.2 0.0

Alimak 1 0.3 0.3 0.0

Shotcrete Machine 1 0.1 0.1 0.0

Personnel Carrier 1 0.3 0.3 0.0

Misc. Underground equipment 2.1 1.3 0.8

Misc. Surface Equipment 3.5 3.5 0.2

Mine Equipment Total 19.9 11.9 8.1

Miscellaneous U/G Infrastructure Lot 1.0 0.3 0.5 0.3


21.1.2 Processing Plant Capital Costs

The capital costs of the process plant include direct costs such as site preparation, all concrete

work, all structural work, process plant equipment and installation, piping, and all electrical

equipment and instrumentation. Indirect process plant capital costs include field supervision and

expenses, construction equipment, engineering design and layouts, spare parts and commission

costs. A summary of the process plant direct and indirect capital costs is presented in Table 21.5.



TABLE 21.5

PROCESS PLANT CAPITAL COST SUMMARY (M CAN$’S)

Description Estimated Cost

Direct Costs

Site Preparation 2.1

Concrete 4.6

Structural 4.3

Equipment 11.1

Equipment Erection 1.5

Piping 2.1

Electrical 3.4

Instrumentation 0.9

Miscellaneous 0.4

Total Direct Costs 30.4

Indirect Costs

Field Supervision 2.1

Field Expenses 1.3

Temporary Facilities 0.9

Construction Equipment 1.5

Craft Benefits 1.9

Engineering 3.9

Freight 1.2

Spare Parts 0.9

Start-up 0.2

Engineering Fee 0.9

Total Indirect Costs 14.9

Total Cost 45.3


The Process Plant construction expenditures are expected to occur in Yr -2 (1/3 of cost) and Yr -

1 (2/3 of the cost).

21.1.3 Surface Infrastructure Capital Costs

Surface infrastructure capital costs include site facilities, buildings, buildings furnishings and

surface mobile equipment.

The capital cost of site facilities includes; the cost of the hydraulic tailings backfill plant and

distribution system; the tailings / waste rock co-disposal basin and dam; site roads; surface

parking areas; the fuel farm; lubrication and oil storage facilities; surface explosive magazines;

yard piping; the fire prevention and fighting system; the potable water treatment plant and

storage tanks; the tailings water treatment plant and pond and the water management pond

building and site run-off.



Buildings capital costs include; the main gate building; the surface mine shop; the warehouse and

warehouse equipment; the office trailers and the dry. The buildings furnishings include; the

surface mine shop equipment and tools; the office furniture, computers, etc.; environmental

equipment; dry equipment; site communications and medical centre equipment.

Surface mobile equipment capital costs include; a road / ramp grader; an integrated tool carrier; a

fuel/lube truck; a service truck; a garbage truck; a personnel bus; an ambulance; a fire/ rescue

truck and pickup trucks. The surface infrastructure capital cost summary is presented in Table

21.6.

TABLE 21.6

SURFACE INFRASTRUCTURE CAPITAL COST SUMMARY (M CAN$’S)

Description Estimated Cost

Site Facilities 12.0

Buildings 2.1

Buildings Furnishings 1.8

Surface Mobile Equipment 2.1

Total 18.0

21.1.4 Mine Closure Capital Costs

The capital cost of removing the shaft headframe, collar house, hoists and hoist room and

securing the surface underground mine openings is estimated to equal the salvage value of these

facilities. A closure bond will be required to remove the process plant, for final tailings

construction and seeding; the tailings spillway, final water treatment and remove surface

infrastructure and final clean up. This closure bond will be required during the pre-production

period. Details of the capital cost and payment schedule for the related closure bond is presented

in Table 21.7.

TABLE 21.7

CLOSURE BOND (M CAN$’S)

Description Total Yr -2 Yr -1

Remove headframe, collar house, hoists(2) and hoist room; Secure Surface

Openings Nil

Remove process plant 4.0 1.0 3.00

Final tailings dam work - 10ha @ $80k/ha plus $50k for design work 0.9 0.2 0.64

Spillway 0.1 0.1 0.1

Final water treatment (batch) 0.1 0.1 0.1

Remove surface infrastructure / clean-up 1.0 0.3 0.7

Total 6.0 1.5 4.5




21.1.5 Other Capital Costs

A capital cost of $2.0 million has been included in year 2 for the powerline construction to the

site. In addition, the cost of $2 million has been included in year 2 for purchasing outstanding

royalties.

21.2 OPERATING COST ESTIMATES

Operating costs include the cost of operating labour, maintenance labour, electrical power,

operating materials and supplies, reagents and fuel. The yearly operating cost varies from $40.12

to $50.41 per tonne milled. A summary of the average operating cost estimates for the Little

Deer Project is provided in Table 21.8.

TABLE 21.8

MINE OPERATING COST PER TONNE MILLED SUMMARY

Description $CAN/t

Mining

Stoping Costs 21.41

Underground Haulage 3.16

Underground Hoisting Services Costs 1.39

Mineral Processing

Process Plant Operating 13.32

Cemented Hydraulic Tailings Backfill 2.75

Tailings to Tailings Impoundment Area 0.19

Tailings Pond Water Treatment 0.08

G&A Costs 5.00

Total Operating 47.32


Details of these estimates are provided in the following subsections.

21.2.1 Mining

On average 1,477 tpd of mill feed will be mined by stoping. The balance of 323 tpd will be

extracted by stope development for a total of 1,800 tpd.

Stope operating costs includes the cost of material, consumables and labour for stope drilling,

blasting, mucking, pipe and accessories, and stope ventilation. The estimated operating cost, per

tonne of stope ore mined, is summarized in Table 21.9. The stope development costs have been

included in the capital costs for the mine.



TABLE 21.9

MINE OPERATING COST PER TONNE OF STOPE ORE

Description CAN$ per

Stoping T

CAN$ per

T of Mill

Feed

Stoping cost per Stoping Tonne

Drilling & Blasting $5.64

Ground Support $0.82

Mucking $1.72

Pipe & Accessories $0.10

Stope Ventilation $0.18

Cemented Hydraulic Tailings Backfill Elsewhere

Total Stoping $8.46

Services and Power Cost per Stoping Tonne $8.00

Staff Labour Cost per Stoping Tonne $3.38

Hourly Labour Cost per Stoping Tonne $6.27

Average Cost per Tonne of Stope Material $26.11 $21.41

Underground Haulage / t of Mill Feed $3.16

Hoisting Services $1.39

21.2.2 Mineral Processing

On average 1,800 tpd ore will be processed. The mineral processing operating cost includes the

cost of all material, consumables and labour required to process 1,800 tpd ore. This includes all

electrical power requirements, reagents, operating and maintenance supplies and labour, and a

5% contingency allowance. A summary of process plant operating costs, per tonne milled and

total cost per year, is presented in Table 21.10.

TABLE 21.10

PROCESS PLANT OPERATING COST PER TONNE MILLED

Item $/t $/annum

Operating Labour 2.61 1,714,800

Power 2.07 1,361,900

Reagents 3.69 2,426,600

Operating Supplies 1.08 711,700

Maintenance Labour 2.22 1,456,900

Maintenance Supplies 1.01 665,700

Total 12.69 8,337,600

Contingency, at 5% 0.63 416,880

Total Cost 13.32 8,754,480



21.2.3 Other Related operating Costs

Other related costs not included in the mining or processing tables are provided in Table 21.11.

TABLE 21.11

OTHER RELATED OPERATING COSTS

Description $/t

Hydraulic Tailings Backfill $2.75

Tailings Pumping to Tailings Pond $0.19

Tailings Pond Water Treatment $0.08

21.2.4 General and Administration

General and Administration (“G&A) costs include costs for staff, general maintenance, office

administration, safety equipment and personal protective equipment (“PPE”), and engineering

tools and professional services cost.

The estimated cost for G&A is approximately $5.00 per tonne milled.



22.0 ECONOMIC ANALYSIS

This Report is considered by P&E Mining Consultants Inc. to meet the requirements of a

Technical Report as defined in Canadian NI 43-101 regulations. This PEA is preliminary in

nature and includes Inferred 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 PEA will be realized. There is no guarantee that

Thundermin / Cornerstone will be successful in obtaining any or all of the requisite consents,

permits or approvals, regulatory or otherwise for the Deposit to be placed into production.

22.1 ECONOMIC CRITERIA

22.1.1 Physicals

Mine life:

Pre-production 18 months

Production Mining/Milling Year 1 to 10 for a total of 9.5 years

Decommissioning 6 months in Year 10.

Production rate 1,800 t per day

Total production:

Total ore production 6,104,700 t ore at 1.84 % Cu

Total concentrate production 388,800 t

Metallurgical para metres:

Process recovery 97%

Concentration ratio 15.7

Concentrate grade 28.1% Cu

Concentrate moisture content 8%

Total payable metal:

Copper 109,200 t of Cu

22.1.2 Revenue

The only commercial product anticipated is copper concentrate which is processed at an off-site

smelter. The copper price used in this PEA is US$3.75/lb. Revenues were calculated as Net

Smelter Returns (NSR‟s). The NSR payables were based on the following parametres.

Smelter treatment charge CAN$/DMT:$75.00/t

Concentrate shipping charge CAN$/WMT:67.00/t

Smelter payable 96%

Cu refining charge CAN$/DMT:$0.07/lb Cu

NSR royalty 1.0% NSR

The US$/CAN$ exchange rate used in the PEA is 0.95.

Net revenue:

Copper: CAN$829.2 million



22.1.3 Costs

Operating costs:

Total average cost: CAN$47.32 per t ore milled

US$ 1.14 per recovered pound Cu

CAN$ 1.16 per contained pound Cu

Capital costs:

Preproduction CAN$98.5 million

Sustaining CAN$204.9 million

Total capital costs CAN$303.4 million

These capital costs include the cost of; mine and stope development; the shaft headframe, hoists,

hoist room, shaft stations and loading pocket; the surface power line; mine equipment; surface

infrastructure; underground infrastructure; the process plant; the purchase of royalties and the

closure bond.

22.2 CASH FLOW

An after-tax cash flow (CF) model has been developed for the Little Deer Project. The model

does not take into account the following components:

Financing cost, other than interest included in capital lease rates

Insurance

Overhead cost for a corporate office

Taxes are estimated to be 30% of pre-tax cash flow. A cash flow summary is presented in Table

22.1. All costs are in 3rd quarter 2011 Canadian dollars with no allowance for inflation.



TABLE 22.1

AFTER-TAX CASH FLOW SUMMARY

Description Units / Yr -2 -1 1 2 3 4 5 6 7 8 9 10 Total

Waste t(000''s) 78 263 323 78 189 83 50 49 112 115 14 1,355

Development Ore t(000''s) 233 138 111 99 118 112 100 80 74 31 1,096

Cu % 1.73 1.95 1.97 1.93 1.89 1.86 1.84 1.85 1.37 1.54 1.82

Stope Ore t(000''s) 306 520 547 558 539 546 557 577 584 275 5,009

Cu % 1.83 1.92 1.97 1.94 1.89 1.85 1.81 1.89 1.67 1.57 1.85

Total Ore t(000''s) 539 657 657 657 657 657 657 657 657 306 6,105

Cu % 1.79 1.92 1.97 1.94 1.89 1.85 1.82 1.89 1.64 1.57 1.84

NSR Can$/tonne Ore 131.39 142.20 145.80 143.44 139.49 136.63 133.87 139.23 119.61 114.40 135.83

Revenue Can$(M's) 70.9 93.5 95.9 94.3 91.7 89.8 88.0 91.5 78.6 35.0 829.2

Operating Cost

Mining Can$(M's) 8.0 13.6 14.3 14.6 14.1 14.2 14.6 15.1 15.2 7.2 130.8

Cemented Hydraulic Tailings Backfill Can$(M's) 1.2 1.5 1.5 1.5 2.0 2.1 2.1 1.9 1.9 1.0 16.8

Tailings to Tailings Dam Can$(M's) 0.2 0.2 0.2 0.2 0.1 0.0 0.0 0.1 0.1 0.0 1.2

Tailings Pond Water Treatment Can$(M's) 0.1 0.1 0.1 0.1 0.1 0.1 0.1 0.1 0.1 0.0 0.5

Process Plant Can$(M's) 7.2 8.8 8.8 8.8 8.8 8.8 8.8 8.8 8.8 4.1 81.3

U/G Haulage Can$(M's) 0.2 2.3 2.5 1.6 2.1 1.7 1.6 1.6 2.1 2.4 1.1 19.3

U/ G Hoisting Services Costs Can$(M's) 0.4 1.1 1.2 1.1 1.0 1.0 1.1 1.1 0.5 8.5

G&A COSTS Can$(M's) 2.7 3.3 3.3 3.3 3.3 3.3 3.3 3.3 3.3 1.5 30.5

Total Operating Can$(M's) 0.0 0.2 21.6 30.3 30.8 31.7 31.1 31.1 31.5 32.4 32.9 15.4 288.9

Capital Cost

Development Cost Can$(M's) 6.5 45.6 40.9 13.9 20.2 15.3 12.2 11.0 14.0 12.7 3.9 196.2

Shaft Headframe, Hoist & Hoist Room, LP Can$(M's) 6.5 6.5 13.0

Power Line - 10km & Hookup Can$(M's) 2.0 2.0

Mine Equipment Can$(M's) 11.9 8.1 19.9

U/G Infrastructure Can$(M's) 0.3 0.5 0.3 1.0

Surface Infrastructure Can$(M's) 18.0 18.0

Process Plant Can$(M's) 15.1 30.2 45.4

Purchase Royalty Can$(M's) 2.0 2.0

Closure Bond Can$(M's) 1.5 4.5 6.0

Total Capital Can$(M's) 16.6 81.9 60.6 41.1 13.9 20.2 15.3 12.2 11.0 14.0 12.7 3.9 303.4

Pre-tax Cash Flow Can$(M's) -16.6 -82.1 -11.4 22.1 51.1 42.5 45.3 46.5 45.6 45.2 33.1 15.7 237.0

Cumulative Pre-tax Cash Flow Can$(M's) -16.6 -98.7 -110.1 -88.0 -36.9 5.6 50.9 97.5 143.0 188.2 221.3 237.0

Taxes @ 30% Can$(M's) 1.7 13.6 14.0 13.7 13.6 9.9 4.7 71.1

After Tax Cash Flow Can$(M's) -16.6 -82.1 -11.4 22.1 51.1 40.8 31.7 32.6 31.9 31.6 23.1 11.0 165.9

Cumulative After Tax Cash Flow Can$(M's) -16.6 -98.7 -110.1 -88.0 -36.9 3.9 35.6 68.2 100.1 131.7 154.9 165.9

After Tax IRR % 21.5%

After Tax NPV @ 6% Can$(M's) 86.7



22.3 BASE CASE CASH FLOW ANALYSIS

The following after tax cash flow analysis was completed:

Net Present Value NPV (at 0%, 5% 7% and 10% discount rate)

Internal Rate of Return IRR

Payback period

The summary of the results of the cash flow analysis is presented in Table 22.2.

TABLE 22.2

BASE CASE CASH FLOW ANALYSIS

Description Discount Rate Units Value

Non Discounted After Tax CF Can$(M) 165.9

Internal Rate of Return % 21.5%

NPV at

0% Can$(M) 165.9

5% Can$(M) 97.1

7% Can$(M) 77.1

10% Can$(M) 52.9

Project Payback Period in Years Years 3.82

The project was evaluated on an after-tax cash flow basis and generates a net cash flow of

$165.9 million. This results in an after tax Internal Rate of Return (IRR) of 21.5% and an after-

tax Net Present Value (NPV) of $86.7 million when using a 6% discount rate. In the base case

scenario, the project has a payback period of 3.8 years. At forecast metal prices and exchange

rates the break-even copper price is estimated to be US$1.14/lb Cu payable at an average

operating cost of Cdn$47.32 per ore tonne ore processed.

22.4 SENSITIVITY ANALYSIS

Project risks can be identified in both economic and non-economic terms. Key economic risks

were examined by running cash flow sensitivities to:

CAN$/US$ exchange rate

Copper metal price

Copper head grade

Copper metallurgical recovery

Operating costs, and

Capital costs

To determine what this project is most sensitive to, each of the sensitivity items were adjusted up

and down by 10% and 20% to see what effect it would have on the NPV at a 6% discount rate.

The value of each sensitivity item, at 80%, 90%, base, 110% and 120%, is presented in Table

22.3.



TABLE 22.3

SENSITIVITY ITEM VALUES

Item 80% 90% 100% 110% 120%

Cu Head Grade - % 1.47 1.66 1.84 2.03 2.21

Cu Metallurgical Recovery - % 77.6% 87.3% 97.0% 100.0%* 100.0%*

Cu Metal Price - US$/lb. $3.00 $3.38 $3.75 $4.13 $4.50

$Can/$US Exchange Rate 0.76 0.86 0.95 1.05 1.14

Opex - Can$/tonne $37.85 $42.59 $47.32 $52.05 $56.78

Capex - Can$(M) $242.7 $273.1 $303.4 $333.8 $364.1

*Note: 100% recovery is achieved with a 3% improvement in recovery over the base case.

The resultant after-tax NPV @ 6% value of each of the sensitivity items at 80% to 120% is presented

in Table 22.4 and Figure 22.1. This after-tax base case NPV is most sensitive to the $CAN/$US

exchange rate followed by the Cu metal price, Cu head grade and metallurgical recoveries, followed

by the capital and operating costs.

TABLE 22.4

SUMMARY OF SENSITIVITY ANALYSIS

Item

After Tax NPV @ 6% at the % Sensitivity Item Values –

CAN$(M)

80% 90% 100% 110% 120%

Cu Head Grade -1.6 42.9 86.7 130.0 173.2

Cu Metallurgical Recovery -1.6 42.9 86.7 100.5* 100.5*

Cu Metal Price -3.3 42.1 86.7 130.8 174.8

$Can/$US Exchange Rate 194.8 134.8 86.7 46.9 13.4

Opex 114.6 100.7 86.7 72.7 58.6

Capex 122.5 104.7 86.7 68.5 50.2

*Note: 100% recovery is achieved with a 3% improvement in recovery over the base case.

Figure 22.1 Sensitivity Graph

Sensitivity Graph

-$5

$20

$45

$70

$95

$120

$145

$170

$195

80% 90% 100% 110% 120%

Percent of Value

Aft

er

Ta

x N

PV

@ 6

% (

M)

Opex Capex Cu Head Grade

Cu Metallurgical Recovery Cu Metal Price $Cdn/$US Exchange Rate



23.0 ADJACENT PROPERTIES

There are no adjacent properties which materially affect the Property. The LDJV controls a

100% interest in adjoining mineral licences that cover a significant portion of the along-strike

extension of the host lithologies and/or structures found at Little Deer.



24.0 OTHER RELEVANT DATA AND INFORMATION

P&E is not aware of any other relevant data or information as of the effective date of this report.



25.0 INTERPRETATION AND CONCLUSIONS

The Little Deer VMS Copper Deposit occurs within the Cambro-Ordovician Lushs Bight Group

sequence of ophiolitic intermediate to mafic volcanic rocks. The main sulphide mineralization

consists of disseminated, stringer, and semi-massive to massive pyrite, pyrrhotite and

chalcopyrite with minor sphalerite.

The Little Deer Copper Deposit was modeled in compliance with the CIM Definitions and

Standards on Mineral Resources and Mineral Reserves, December 11, 2005. National Instrument

43-101 reporting standards and formats were followed in this document in order to report the

mineral resource in a fully compliant manner.

Diamond drill data from 48,432 m of drilling in 82 drillholes completed by Thundermin and

Cornerstone since June 2007, and assay data from a total of 102 surface and 122 underground

historical holes that were drilled by BRINEX between 1961 and 1970 and Mutapa Gold

Corporation between 1998 and 2000 were used for the Resource Estimate.

Exploration drilling can extend the known copper mineralized zones at depth and infill drilling

can convert Inferred Resources to Indicated Resources

P&E Mining Consultants Inc. offers the following interpretation and conclusions:

This Report is considered by P&E Mining Consultants Inc. to meet the

requirements of a Technical Report as defined in Canadian NI 43-101 regulations.

The economic analysis contained in this Report is based on Indicated and Inferred

Resources. The mineral resources in this PEA were estimated using the CIM

Standards on Mineral Resources and Reserves, Definitions and Guidelines

prepared by the CIM Standing Committee on Reserve Definitions and adopted by

CIM Council, December 11, 2005.

There is no guarantee that Thundermin / Cornerstone will be successful in

obtaining any or all of the requisite consents, permits or approvals, regulatory or

otherwise for Little Deer or that Little Deer will be placed into production.

The project was evaluated on an after-tax cash flow basis and generates a net cash

flow of $165.9 million after-tax. This results in an after tax Internal Rate of

Return (IRR) of 21.5% and an after-tax Net Present Value (NPV) of $86.7 million

when using a 6% discount rate. In the base case scenario, the project has a

payback period of 3.8 years. At forecast metal prices and exchange rates the

break-even copper price is estimated to be US$1.14/lb. Cu payable at an average

operating cost of CAN$47.32 per ore tonne ore processed.

The after-tax base case NPV is most sensitive to the $CAN/$US exchange rate

followed by the Cu metal price, Cu head grade and metallurgical recoveries,

capital and operating costs, respectively.

The Longhole Longitudinal Retreat mining method is determined to be the

preferred mining method to be applied at Little Deer because it is effective in

mining the resources at a relatively low cost while still maintaining dilution

around 20%, at zero grade and recovery at around 90%.



26.0 RECOMMENDATIONS

P&E concludes that the Little Deer Project has economic potential as an underground mining and

milling operation producing copper in concentrate

Note: This PEA is preliminary in nature and its mineable tonnage includes 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 will be realized. Mineral resources that are not

mineral reserves do not have demonstrated economic viability.

P&E recommends that Thundermin and Cornerstone advance the project with extended and

advanced technical studies particularly in metallurgical, geotechnical and environmental matters

with the intention to advance the project to a Pre-feasibility stage.

Specifically, it is recommended that Thundermin and Cornerstone take the following actions to

develop the project to a Pre-Feasibility Study level

Perform a comprehensive program of metallurgical testing on representative

samples of the mineralized zone(s), to assess and confirm expected recoveries,

reagent usages, process flow sheets and other associated operating issues;

Carry out hydrogeological and hydrological analyses of the project site and

surrounding area;

Carry out a more detailed geotechnical assessment of ground conditions to be able

to estimate the ground support required and expected waste rock dilution of the

mill feed with more confidence in subsequent studies;

Table 26.1 lists recommended actions and associated preliminary cost for the

recommendations. The total preliminary budget for the recommended activities

targeted at the development of the project to a Pre-Feasibility Study stage is

$5.25 million.

TABLE 26.1

PRELIMINARY BUDGET FOR PROJECT DEVELOPMENT TO PRE-FEASIBILITY STUDY LEVEL

Proposed Budget Elements Cost

Estimate

Hydrogeological and hydrological analyses $200,000

Geotechnical test work. $750,000

Infrastructure $200,000

Engineering design of a TMF and waste rock pile $500,000

Pre-Feasibility level metallurgical test work including both bench scale and limited

pilot scale. $400,000

Resource Upgrade drilling from inferred to indicated and measured category (see

Table 26.2) $2,000,000

Preliminary environmental and socio-economic impact assessment work in the

project area and data collection for a EIA $100,000

Test work on tailings characterization and treatment options. $150,000

Examination of land acquisition options and acquisition cost. $50,000

Preparation of a Pre-Feasibility Study $900,000

Total Budget $5,250,000



26.1 RESOURCE UPGRADE

Thundermin and Cornerstone recommend the following approximately $2.0 M on-going

exploration program for the Little Deer Copper Deposit for the period October 1, 2011 to

June 30, 2012.

A small program involving the re-assaying of standards and other check samples

using aqua regia and four acid digestion to try and determine if the estimated

resource grade may, in fact, be higher than estimated as discussed in Section

12.2.1 of this report.

Additional diamond drilling to test for extensions of the copper mineralization at

depth and along strike. Infill diamond drilling on approximately 50 m centres to

upgrade the Inferred Resources to the Indicated Resource category. The infill

drilling, which will be required in order to undertake a pre-feasibility or feasibility

study on the Deposit, should commence at shallower levels of the Deposit and

proceed to depth.

Borehole Pulse EM surveys on selected deep drill holes.

Differential GPS surveys on all new drill holes.

Revised NI 43-101 mineral resource estimate following completion of the

recommended diamond drill program.

It is anticipated that this work will be undertaken in two phases: Phase 1 (approx. October 1 to

December 31, 2011 and Phase 2 (January 1 to June 30, 2012), as shown in Table 26.2.

TABLE 26.2

PROPOSED EXPLORATION PROGRAM AND BUDGET

Phase 1 CAD$

4,500 m of diamond drilling at $120.00 per m 540,000


Re-assaying 500

Total Phase 1 543,000

Phase 2

12,000 m of diamond drilling at $120.00 per m 1,440,000


Borehole Pulse EM surveys on 5-6 holes 40,000

Revised NI 43-101 Mineral Resource estimate 25,000

Total Phase 2 1,508,000

The estimated drilling costs are “all-in” costs and include direct drilling costs, salaries and

wages, assaying, room and board, truck rentals, management fees etc.



27.0 REFERENCES

Bowman, B.A., Caldwell, R.J., (2010), An Investigation into Scoping Level Environmental

Characterisation of Little Deer Flotation Tailings prepared for Thundermin Resources Inc.

Project 12426-002 Final Report, SGS Minerals Services.

Galley, A.G., Hannington, M.D. and Jonasson, I.R., (2007), Vocanogenic Massive Sulphide

Deposits, in Goodfellow, W.D., ed., Mineral Deposits of Canada: A Synthesis of Major Deposit-

Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods:

Geological Association of Canada, Mineral Deposits Division, Special Publication No.5, p. 141-

161.

Imeson, D., (2010), An Investigation into the Recovery of Copper from the Little Deer Deposit

prepared for Thundermin Resources Inc.Project 12426-001 Final Report, SGS Minerals Services.

Kean, B.F., Evans, D.T.W., Jenner, G.A. (1995), Geology and Mineralization of the LushsBight

Group, Report 95-2. Geological Survey of Newfoundland and Labrador.

Pressacco, R., (2009), Technical Report on the Initial Mineral Resource Estimate for the Little

Deer Copper Deposit, Newfoundland, Canada. Micon International Limited.

Pressacco, R., (2010), Mineral Resource Update for the Little Deer Project. Scott Wilson Roscoe

Postle Associates Inc.

Puritch, E.J., and Ewert, W.D., 2011: Technical Report and Resource Estimate Update on the

Little Deer Copper Deposit Newfoundland, Canada, dated August 5, 2011.

Taylor, C., Zierenberg, A., Goldfarb, R., Kilburn, J., Seal II, R., Klienkopf, D. (1995) Volcanic-

Associated Massive Sulfide Deposits. USGS ofr-95-0831.

Van Staal, C.R., (2007), Pre-Carboniferous tectonic evolution and metallogeny of the Canadian

Appalachians in Goodfellow, W.D., ed., Mineral Deposits of Canada: A Synthesis of Major

Deposit –Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration

Methods: Geological Association of Canada, Mineral Deposits Division, Special Publication No.

5, p793-818.



28.0 CERTIFICATES

CERTIFICATE OF QUALIFIED PERSON

EUGENE J. PURITCH, P. ENG.

I, Eugene J. Puritch, P.Eng., residing at 44 Turtlecreek Blvd., Brampton, Ontario, L6W 3X7, do hereby certify that:

1. I am President of P & E Mining Consultants Inc. and am contracted independently by Thundermin Resources Inc.

and Cornerstone Resources Inc. 2. This certificate applies to the technical report titled “Technical Report and Preliminary Economic Assessment

(PEA) of the Little Deer Copper Deposit, Newfoundland, Canada”, (the “Technical Report”) with an effective

date of November 1, 2011.

3. I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as

obtaining an additional year of undergraduate education in Mine Engineering at Queen‟s University. In addition

I have also met the Professional Engineers of Ontario Academic Requirement Committee‟s Examination

requirement for Bachelor‟s Degree in Engineering Equivalency. I am a mining consultant currently licensed by

the Professional Engineers of Ontario (License No. 100014010). I am also registered in the Province of

Saskatchewan (APEGS No. 16216) and the Province of Newfoundland and Labrador (PEG No. 05998) and

registered with the Ontario Association of Certified Engineering Technicians and Technologists as a Senior

Engineering Technologist. I am also a member of the National and Toronto Canadian Institute of Mining and

Metallurgy.

I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify

that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past

relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

I have practiced my profession continuously since 1978. My summarized career experience is as follows:

Mining Technologist - H.B.M.& S. and Inco Ltd., ......................................................... 1978-1980

Open Pit Mine Engineer – Cassiar Asbestos/Brinco Ltd., ............................................... 1981-1983

Pit Engineer/Drill & Blast Supervisor – Detour Lake Mine, ........................................... 1984-1986

Self-Employed Mining Consultant – Timmins Area, ...................................................... 1987-1988

Mine Designer/Resource Estimator – Dynatec/CMD/Bharti, ......................................... 1989-1995

Self-Employed Mining Consultant/Resource-Reserve Estimator, .................................. 1995-2004

President – P & E Mining Consultants Inc, ................................................................. 2004-Present

4. I am responsible for co-authoring Sections 1, 11, 12, 14 as well as Section 25 of the Technical Report.

5. I have visited the Property on May 16, 2011.

6. I have had no prior involvement with the Property that is the subject of this Technical Report.

7. 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.

8. I am independent of the issuer applying the test in Section 1.5 of NI 43-101.

9. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance

therewith.


Signed Date: December 15, 2011

{SIGNED AND SEALED}

[Eugene J. Puritch]

Eugene J. Puritch, P.Eng.




WAYNE D. EWERT, P.GEO.

I, Wayne D. Ewert, P.Geo., residing at 10 Langford Court, Brampton, Ontario, L6W 4K4, do hereby certify that:

1. I am a principal of P & E Mining Consultants Inc. who has been contracted by Thundermin Resources Inc. and

Cornerstone Resources Inc.

2. This certificate applies to the technical report titled “Technical Report and Preliminary Economic Assessment

(PEA) of the Little Deer Copper Deposit, Newfoundland, Canada”, (the “Technical Report”) with an effective


3. I graduated with an Honours Bachelor of Science degree in Geology from the University of Waterloo in 1970

and with a PhD degree in Geology from Carleton University in 1977. I have worked as a geologist for a total of

42 years since obtaining my B.Sc. degree. I am a P. Geo., registered in the Province of Saskatchewan (APEGS

No. 16217), British Columbia (APEGBC No. 18965), the Province of Ontario (APGO No. 0866) and the

Province of Newfoundland and Labrador (PEG No. 06005).


that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past


My relevant experience for the purpose of the Technical Report is:

Principal, P&E Mining Consultants Inc. ..............................................................................2004 – Present

Vice-President, A.C.A. Howe International Limited ............................................................... 1992 – 2004

Canadian Manager, New Projects, Gold Fields Canadian Mining Limited ............................. 1987 – 1992

Regional Manager, Gold Fields Canadian Mining Limited ..................................................... 1986 – 1987

Supervising Project Geologist, Getty Mines Ltd. .................................................................... 1982 – 1986

Supervising Project Geologist III, Cominco Ltd. .................................................................... 1976 – 1982

4. I have not visited the Property that is the subject of this Technical Report.

5. I am responsible for authoring Sections 2 through 10, 23, 24, and as well as co-authoring Sections 1, 25 and 26

of this Technical Report.

6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101.

7. I have not had prior involvement with the project that is the subject of this Technical Report.


therewith.

9. 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 SEALED}

[Wayne Ewert]

________________________________

Dr. Wayne D. Ewert P.Geo.



KIRK RODGERS, P.ENG.

CERTIFICATE OF AUTHOR

I, Kirk H. Rodgers, P. Eng., residing at 378 Bexhill Rd., Newmarket, Ontario, do hereby certify that:

1. I am an independent mining consultant, contracted as Vice President, Engineering by P&E Mining Consultants Inc.

2. This certificate applies to the technical report titled “Technical Report and Preliminary Economic Assessment (PEA)

of the Little Deer Copper Deposit Newfoundland, Canada (the “Technical Report”) with an effective date of

November 1, 2011.

3. I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining. I subsequently attended

the mining engineering programs at Laurentian University and Queen‟s University for a total of two years. I have

met the Professional Engineers of Ontario Academic Requirement Committee‟s Examination requirement for

Bachelor‟s Degree in Engineering Equivalency.

I have been licensed by the Professional Engineers of Ontario (License No. 39427505), from 1986 to the present. I

am also a member of the National and Toronto Canadian Institute of Mining and Metallurgy.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that,

by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work

experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

My relevant experience for the purpose of the Technical Report is:

Underground Hard Rock Miner, Denison Mines, Elliot Lake Ontario ............................................... 1977-1979

Mine Planner, Cost Estimator, J.S Redpath Ltd., North Bay Ontario ................................................. 1981-1987

Chief Engineer, Placer Dome Dona Lake Mine, Pickle Lake Ontario ................................................ 1987-1988

Project Coordinator, Mine Captain, Falconbridge Kidd Creek Mine, Timmins, Ontario ................... 1988-1990

Manager of Contract Development, Dynatec Mining, Richmond Hill, Ontario ................................. 1990-1992

General Manager, Moran Mining and Tunnelling, Sudbury, Ontario................................................. 1992-1993

Independent Mining Engineer .....................................................................................................................1993

Project Manager - Mining, Micon International, Toronto, Ontario ................................................. 1994 - 2004

Principal, Senior Consultant, Golder Associates, Toronto, Ontario ............................................... 2004 – 2010

Independent Consultant, VP Engineering to P&E Mining Consultants Inc, Brampton Ontario .. 2011 – present

4. I am responsible for co-authoring the Sections 1, 22, 25, and 26.

5. I have not visited the Property that is the subject of this report.

6. 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.

7. I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.


9. I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.



{SIGNED AND SEALED}

{Kirk Rodgers}

Kirk Rodgers, P. Eng.



JAMES L. PEARSON, P.ENG.

CERTIFICATE OF AUTHOR

I, James L. Pearson, P.Eng., residing at 5 Clubhouse Court, Bolton, Ontario, Canada, L7E 0B3, do hereby certify

that::

1. I am an independent Mining Engineering Consultant, contracted by P& E Mining Consultants Inc.

2. This certificate applies to the technical report entitled “Technical Report and Preliminary Economic

Assessment (PEA) of the Little Deer Copper Deposit Newfoundland, Canada (the “Technical Report”) with an

effective date of November 1, 2011” (the “Technical Report”) dated November 1, 2011.

3. I am a graduate of Queen‟s University, Kingston, Ontario, Canada, in 1973 with a Bachelor of Science

degree in Mining Engineering. I am registered as a Professional Engineer in the Province of Ontario (Reg.

No. 36043016). I have worked as a mining engineer for a total of 37 years since my graduation.

I have read the definition of "qualified person" set out in National Instrument 43-101 (“NI 43-101”) and

certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101)

and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of

NI 43-101. My relevant experience for the purpose of the Technical Report is:

Review and report as a consultant on numerous exploration and mining projects around the world

for due diligence and regulatory requirements;

Project Manager and Superintendent of Engineering and Projects at several underground

operations in South America;

Senior Mining Engineer with a large Canadian mining company responsible for development of

engineering concepts, mine design and maintenance;

Mining analyst at several Canadian brokerage firms


5. I am responsible for authoring Sections 15, 16, 18 and 19 as well as co-authoring Sections 1, 14, 21, 22, 25

and 26 of the Technical Report;

6. I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

7. I have had no prior involvement with the property that is the subject of the Technical Report.

8. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with

that Instrument and Form.


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

Report not misleading.

Effective date: November 1, 2011


{SIGNED AND SEALED}

[James L. Pearson]

James L. Pearson, P. Eng.




DAVID A. ORAVA, P. ENG.

I, David A. Orava, M. Eng., P. Eng., residing at 19 Boulding Drive, Aurora, Ontario, L4G 2V9, do hereby certify

that:

1. I am an Associate Mining Engineer at P&E Mining Consultants Inc. and President of Orava Mine Projects Ltd.


(PEA) of the Little Deer Copper Deposit Newfoundland, Canada (the “Technical Report”) with an effective

date of November 1, 2011” (the “Technical Report”) dated November 1, 2011.

3. I am a graduate of McGill University located in Montreal, Quebec, Canada at which I earned my Bachelor

Degree in Mining Engineering (B.Eng. 1979) and Masters in Engineering (Mining - Mineral Economics Option

B) in 1981. I have practiced my profession continuously since graduation. I am licensed by the Professional

Engineers of Ontario (License No. 34834119).




My summarized career experience is as follows:

Mining Engineer – Iron Ore Company of Canada. ..................................................................... 1979-1980

Mining Engineer – J.S Redpath Limited / J.S. Redpath Engineering. ........................................ 1981-1986

Mining Engineer & Manager Contract Development – Dynatec Mining Ltd. ........................... 1986-1990

Vice President – Eagle Mine Contractors............................................................................................ 1990

Senior Mining Engineer – UMA Engineering Ltd. ............................................................................. 1991

General Manager - Dennis Netherton Engineering .................................................................... 1992-1993

Senior Mining Engineer – SENES Consultants Ltd. .................................................................. 1993-2003

President – Orava Mine Projects Ltd. .................................................................................. 2003 to present

Associate Mining Engineer – P&E Mining Consultants Inc. .............................................. 2006 to present


5. I am responsible for authoring Section 20 of this Technical Report.

6. I am an independent of the issuer applying all of the tests in Section 1.5 of NI 43-101.

7. I have had no prior involvement with the project that is the subject of this Technical Report.

8. I have read NI 43-101 and Form 43-101F1 and the Report has been prepared in compliance therewith.


contains all scientific and technical information that is required to be disclosed to make the Technical Report not

misleading.



{SIGNED AND SEALED}

[David Orava]

____________________________________

David Orava, M. Eng., P. Eng.




ALFRED S. HAYDEN, P. ENG

I, Alfred S. Hayden, P. Eng., residing at 284 Rushbrook Drive, Ontario, L3X 2C9, do hereby certify that:

1. I am currently President of:

EHA Engineering Ltd.,

Consulting Metallurgical Engineers

Box 2711, Postal Stn. B.

Richmond Hill, Ontario, L4E 1A7


(PEA) of the Little Deer Copper Deposit Newfoundland, Canada” (the “Technical Report”), with an effective


3. I graduated from the University of British Columbia, Vancouver, B.C. in 1967 with a Bachelor of Applied

Science in Metallurgical Engineering. I am a member of the Canadian Institute of Mining, Metallurgy and

Petroleum and a Professional Engineer and Designated Consulting Engineer registered with Professional

Engineers Ontario. I have worked as a metallurgical engineer for a total of 42 years since my graduation from

university.




4. I have not visited the Property that is the subject of this report.

5. I am responsible for authoring of Section 13 and 17 as well as co-authoring Sections 1 and 21 of the Technical

Report

6. I am independent of the issuer applying the test in Section 1.5 of NI 43-101.



therewith.


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

not misleading.



{SIGNED AND SEALED}

[Alfred Hayden]

__________________________

Alfred S. Hayden, P.Eng.



APPENDIX I. UNDERGROUND MINE PLAN DRAWINGS



A longitudinal section of the proposed mine layout.



Typical plans of proposed mine development are presented on the following pages.

technical report and preliminary economic assessment … · little deer copper deposit...

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