technical report on the dome mountain gold-silver project ... mountain_2010.pdftechnical report on...

Technical Report on theDome Mountain Gold-Silver Project

Omineca Mining DivisionBritish Columbia, Canada

NTS 93L / 10ELat. 54°44'N Long. 126°37'W

Prepared For

Eagle Peak Resources Inc.&

Metal Mountain Resources Inc.Suite 413, Bentall 3595 Burrard Street

Vancouver, BC, CanadaV7X 1G4

Prepared By

Gary Giroux, P.Eng.Giroux Consultants Ltd.

April 2010

Dome Mountain Technical Report April 2010

Giroux Consultants Ltd. Page i

TABLE OF CONTENTS1 SUMMARY........................................................................................................................................... 12 INTRODUCTION AND TERMS OF REFERENCE ................................................................................ 33 RELIANCE ON OTHER EXPERTS ...................................................................................................... 54 PROPERTY DESCRIPTION AND LOCATION ..................................................................................... 6LAND TENURE ................................................................................................................................... 6

5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY... 116 HISTORY ........................................................................................................................................... 137 GEOLOGICAL SETTING ................................................................................................................... 16REGIONAL GEOLOGY...................................................................................................................... 16LOCAL GEOLOGY ............................................................................................................................ 16PROPERTY GEOLOGY .................................................................................................................... 20

8 DEPOSIT TYPES ............................................................................................................................... 259 MINERALIZATION............................................................................................................................. 2610 EXPLORATION................................................................................................................................ 2811 DRILLING ........................................................................................................................................ 29PRE-2009 DRILLING......................................................................................................................... 292009 DRILLING................................................................................................................................. 30

12 SAMPLING METHOD AND APPROACH ......................................................................................... 32PRE-2009 DRILLING......................................................................................................................... 322009 DRILLING................................................................................................................................. 322009 UNDERGROUND SAMPLING................................................................................................... 32

13 SAMPLE PREPARATION, ANALYSES AND SECURITY ................................................................ 33PRE-2009 DRILLING......................................................................................................................... 332009 DRILLING................................................................................................................................. 332009 UNDERGROUND SAMPLING................................................................................................... 34

14 DATA VERIFICATION...................................................................................................................... 35STANDARDS .................................................................................................................................... 35BLANKS ............................................................................................................................................ 36PULP DUPLICATES .......................................................................................................................... 37REJECT DUPLICATES...................................................................................................................... 38FIELD DUPLICATES ......................................................................................................................... 39METALLIC GOLD DUPLICATES........................................................................................................ 40

15 ADJACENT PROPERTIES .............................................................................................................. 4216 MINERAL PROCESSING AND METALLURGICAL TESTING ......................................................... 4417 MINERAL RESOURCE ESTIMATES ............................................................................................... 4918 OTHER RELEVANT DATA AND INFORMATION ............................................................................ 6519 INTERPRETATION AND CONCLUSIONS ....................................................................................... 6620 RECOMMENDATIONS .................................................................................................................... 6821 REFERENCES................................................................................................................................. 7022 SIGNATURE PAGE ......................................................................................................................... 7123 CERTIFICATE OF QUALIFIED PERSON......................................................................................... 72


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List of FiguresFigure 1 Location Map .......................................................................................................8Figure 2 Tenure Block Map................................................................................................9Figure 3 Regional Geology ..............................................................................................17Figure 4 Local Geology....................................................................................................18Figure 5 Vein Systems ....................................................................................................21Figure 6 Drill Hole Location Map .......................................................... insert after page 29Figure 7 Cross Section 653100 E ....................................................................................31Figure 8 Standards vs Time Plot......................................................................................35Figure 9 Blanks vs Time Plot ...........................................................................................36Figure 10 Pulp Duplicate Scatter Plot ................................................................................37Figure 11 Reject Duplicate Scatter Plot .............................................................................38Figure 12 Core Duplicate Scatter Plot................................................................................39Figure 13 Metallic Gold Results .........................................................................................40Figure 14 Metallic Duplicate Scatter Plot ...........................................................................41Figure 15 Three Dimensional Vein Solids ..........................................................................50Figure 16 Cumulative Frequency Plot for Au in Pre-2009 vs 2009 Drilling .........................55Figure 17 Plan View of Veins Showing Underground Development ...................................62Figure 18 Grade Tonnage Curves for Total Resource in the Total Resource

in the Undiluted, FW Diluted and HW Diluted Boulder Main Vein.......................64

List of TablesTable 1 Claim Details......................................................................................................10Table 2 History Summary ...............................................................................................15Table 3 Surface Drilling Summary ..................................................................................29Table 4 QA/QC Standard Failures ..................................................................................36Table 5 Cyanide Leaching Test Results..........................................................................46Table 6 Assay Statistics Sorted by Zone.........................................................................51Table 7 Assay Statistics Sorted by Lithology...................................................................52Table 8 Assay Statistics Sorted by Colour ......................................................................53Table 9 Capping Results Sorted by Zone .......................................................................53Table 10 Capped Assay Statistics Sorted by Zone ...........................................................54Table 11 1 Metre Composite Statistics Sorted by Zone ....................................................56Table 12 Summary of Semivariogram Parameters ...........................................................57Table 13 Indicated Resource within the Mineralized Veins ...............................................63Table 14 Inferred Resource within the Mineralized Veins .................................................63Table 15 Sensitivity of Indicated Resource to Gold Cut-off Grade.....................................66Table 16 Phase 1 Budget .................................................................................................68Table 17 Phase 2 Budget .................................................................................................69


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List of Appendices

APPENDIX I List of Drill-holes Used in Resource AnalysisAPPENDIX II Semivariograms Plots for Gold and silverAPPENDIX III Specific Gravity MeasurementsAPPENDIX IV 2009 Drill-hole Intersection CompositesAPPENDIX V Assayers Canada Fire Assay and ICP Analytical ProcedureAPPENDIX VI Ore Reference StandardsAPPENDIX VII Acme Fire Assay Analytical ProcedureAPPENDIX VIII Assayers Canada Metallic Gold Analytical ProcedureAPPENDIX IX Memo from Moose Mountain Technical Services Re: Dome Mountain Vein

ModelingAPPENDIX X Petrographic Evaluation of a Zinc, Copper, Lead, Silver and Gold-Bearing

Composite from Dome Mountain, British ColumbiaAPPENDIX XI Proposal on Metallurgical Testing - Dome Mountain ProjectAPPENDIX XII Cyanidation Test Reports


Giroux Consultants Ltd. Page 1

1 SUMMARY

Metal Mountain Resources Inc. (Metal Mountain) is an associate company of Eagle Peak

Resources Inc. (Eagle Peak) as both companies have the same management. In 2007 Eagle

Peak optioned the mineral tenures that cover the past-producing Dome Mountain Gold-Silver

Mine. Over the next two years Eagle Peak optioned and purchased four additional tenure

blocks in the Dome Mountain area. The tenure blocks were acquired for the purpose of

conducting further exploration work and re-opening the mine. On October 16, 2009 Metal

Mountain purchased all five tenure blocks from Eagle Peak.

The Project is situated approximately 38 kilometres due east of Highway 16 and the Town of

Smithers in northwest B.C. and is close to existing CN rail, power and highway infrastructure.

The deep-water ports of Prince Rupert, Kitimat and Stewart are located approximately 400 km

to the west.

After acquiring the property, Eagle Peak completed surface diamond drilling and underground

mapping and sampling programs to verify results by previous operators. The company drilled a

total of 42 HQ diamond drill-holes (5705 metres) to in-fill the existing drill pattern and collected

193 chip samples from the underground workings. This work culminated in the calculation of

the NI 43-101 Mineral Resource Estimate presented in this report.

Exploration drilling of coincident soil geochemistry and 3D induced polarization anomalies by

Eagle Peak in 2009 intersected a quartz vein that assayed 19.07 grams per tonne gold over the

0.6 metre interval from 49.4 to 50.0. This intersection is located approximately 400 metres

northeast of the 1290 portal and may represent the discovery of a new vein.

The project area is underlain by Early to Middle Jurassic calc-alkalic island arc rocks of the

Telkwa, Nilkitkwa, and Smithers Formations of the Hazelton Group. Structurally the area is a

northwest trending horst of folded and faulted Jurassic and Cretaceous volcanic and

sedimentary rocks bounded to the west and east by grabens containing Late Cretaceous and

younger rocks. This structure is hypothesized as the main control for the gold-silver

mineralization.

Several orogenic (mesothermal) quartz-carbonate-sulphide veins with economic potential have

been identified within the area of the Dome Mountain Project. During the period from 1991 to



1993, Timmins Nickel Inc. mined 43,900 tonnes at an average grade of 12.0 grams per tonne

gold from the Boulder Main Vein. Underground mining was conducted by shrinkage stope

methods accessed from trackless drift developments on the 1290 and 1370 levels.

A NI 43-101 compliant Mineral Resource Estimate for the Boulder Vein System has been

calculated using 285 surface diamond drill holes and 37 underground diamond drill holes. As

there has been no economic evaluation completed at this time to establish a cut-off grade, a

reasonable 5 g/t cutoff for an underground operation has been highlighted. This resource

occurs principally above the 1290 level and has been adjusted to remove the volumes

previously mined.

Indicated Resource within the Mineralized Veins

Au Cutoff(g/t)

Tonnes > Cutoff(tonnes)

Grade>Cutoff Contained MetalAu (g/t) Ag (g/t) Au (ozs) Ag (ozs)

4.00 143,000 14.76 72.83 67,900 334,8005.00 138,000 15.10 73.93 67,000 328,0006.00 133,000 15.46 75.36 66,100 322,200

Inferred Resource with the Mineralized Veins

Au Cutoff(g/t)



4.00 173,000 12.43 57.03 69,100 317,2005.00 154,000 13.42 60.63 66,500 300,2006.00 137,000 14.43 64.49 63,500 284,000

The best opportunities for increasing the mineral resources on the Dome Mountain Project are

down-plunge to the east on the Boulder Main and Boulder Footwall veins followed by the

Argillite Vein below the 1290 level.

Two phases of additional work are recommended on the Dome Mountain Project. Phase 1

would consist of a Preliminary Assessment to determine if the indicated resource is economic,

additional in-fill drilling to upgrade the inferred resource, exploration drilling down-plunge to the

east on the Boulder Vein System, and extension of the existing 3D induced polarization and soil

geochemistry surveys. Cost of the phase program is estimated at $1.464 M. Phase 2 would be

for following up Phase 1 positive exploration results by diamond drilling. Costs for Phase 2 are

estimated at $1.156 M.



2 INTRODUCTION AND TERMS OF REFERENCE

Metal Mountain Resources Inc. (MMR) is an associate company of Eagle Peak Resources Inc.

(EPR). Between June 2007 and March 2009 EPR acquired, through four option agreements

and one purchase, the five tenure blocks that make up the Dome Mountain Project. On October

16, 2009 MMR purchased all five tenure blocks from EPR and on January 29, 2010 proceeded

to exercise its option on the block containing the historic underground workings on the Boulder

Vein System as described in a NI 43-101 Technical Report by Scott Wilson Roscoe Postle

Associates Inc. (Scott Wilson RPA) dated April 28, 2008 (Rennie, 2008).

Gold mineralization was first located on the property in the late 1800s and considerable surface

and underground work was done in 1923-24. Resumption of exploration in the 1980s led to the

discovery of the Boulder Vein System in 1985. Underground mining was initiated in August

1991 by Timmins Nickel Inc. (TNI) and its joint venture partner, Habsburg Resources Inc. (HRI)

and ceased in May 1993 due to legal and financial problems. During the period of operation

43,900 tonnes at an average grade of 12.0 grams per tonne gold were reportedly mined. The

ore was transported and milled on a toll basis at the Equity Silver mill near Houston, British

Columbia and at the Westmin mill near Stewart, British Columbia. In 2008 EPR conducted soil

geochemistry and 3D induced polarization surveys over a 120 hectare block centered on the

mine workings. In 2009 EPR completed the work recommended by Scott Wilson RPA, including

diamond drilling, to verify and report the historic resource as a current resource consistent with

NI 43-101. EPR also undertook a small diamond drilling program in 2009 to search for

additional resources.

MMR acquired the properties with the intention of re-opening the past-producing Dome

Mountain Mine and exploring for additional resources. Since acquiring the properties, MMR has

completed a NI 43-101 resource estimate and is continuing with the metallurgical,

environmental and engineering studies initiated by EPR. .

This report reviews the current status of the technical aspects of the Dome Mountain Project,

discloses the results of the 2010 resource estimate by Giroux Consultants Ltd. and makes

recommendations for further work. The report was prepared in compliance with the

requirements of National Instrument 43-101 Standards of Disclosure for Mineral Deposits and is



intended to be used as a supporting document to be filed with the British Columbia Securities

Commission and the TSX Venture Exchange.

Gary Giroux, P. Eng. of Giroux Consultants Ltd. (GCL) completed a site visit to the property on

March 7, 2010. He inspected the 1290 level and observed the vein exposures in two stopes.

SOURCES OF INFORMATIONAnthony L'Orsa, P. Geo. provided the summary of the property history.

Mineral tenure information was obtained by IGS from the BC Mineral Titles On-Line website

(https://www.mtonline.gov.bc.ca/mtov/home.do).

Information on tenure option and purchase agreements came from the corporate office of MMR.

The sources of written information reviewed for this report are listed in Section 21 - References.



3 RELIANCE ON OTHER EXPERTS

The author of this report has relied on the historical exploration data, reports and certain legal

documents provided to him by MMR and on reports available in hardcopy or digital form from

the British Columbia Ministry of Energy, Mines & Petroleum Resources.

Daryl J. Hanson, P. Eng. of In-Depth Geological Services (IGS) spent a total of twelve months

as a consultant working on the Dome Mountain Project between December 2009 and February

2010. He supervised the 2009 diamond drilling program, the creation of the drill hole database,

the interpretation of drill sections, the development of the vein solids model, and the

underground mapping and sampling. Additionally he wrote all sections of this report except for

section 17 - Mineral Resource Estimate.

The results of the metallurgical testing were provided by H.M. Bolu, P.Eng.

Except for the purposes legislated under provincial securities laws, any use of this report by any

third party is at that Party's sole risk.



4 PROPERTY DESCRIPTION AND LOCATION

The subject mineral property is located approximately 38 kilometres due east of the town of

Smithers in northwest British Columbia at 126°37' W longitude and 54°44' N latitude. The

property is within the Omineca Mining Division on NTS Map Sheet 93L 10E (Figure 1).

The Dome Mountain property consists of forty one (41) contiguous mineral claims and one

mining lease comprising a total area of 10,970.9 hectares. The configuration of the claim blocks

is shown in Figure 2 and the details are listed in Table 1.

There are no existing surface rights other than the mining lease over the property known to the

author. The tenures are part of the Lake Babine Band (Nedut'en) First Nation traditional use

territory.

LAND TENURETenure Block A, which includes the mining lease, was acquired by Eagle Peak Resources Inc.

in June 2007 via an option agreement with Angel Jade Mines Ltd., Kevin Coswan, Judith

L'Orsa, and Anthony L'Orsa (the Dome Royalty Group). On October 16, 2009 Eagle Peak sold

its interest in the block to Metal Mountain Resources Inc. and on January 29, 2010 Metal

Mountain exercised its option with the vendors. As a result of these transactions, Metal

Mountain has acquired a 100% interest in the tenures subject to a 2% Net Smelter Return

(NSR) royalty payable to the Dome Royalty Group. Metal Mountain has the right to buy back

50% of this royalty for $1,000,000.

Eagle Peak acquired tenure Block B by option from Guardsmen Resources Inc. (Guardsmen)

on March 12, 2009 and subsequently sold this interest to Metal Mountain on October 16, 2009.

Metal Mountain has the right to acquire a 100% interest in the tenures by making staged

payments totaling $500,000 by March 31, 2011. On exercise of the option, Guardsmen would

retain a 3% NSR royalty. Up to 24 months after exercising the option, Metal Mountain has the

right to purchase up to 2% of the royalty by paying $50,000 for each one-tenth of one percent of

the NSR to be purchased.

Lorne Brian Warren (Warren) granted an option to Eagle Peak to acquire tenure Block C on

February 29, 2008. The option was subsequently sold to Metal Mountain on October 16, 2009.



Metal Mountain currently has the right to acquire a 100% interest in the claims by making

staged payments totaling $120,000 and issuing 600,000 shares by March 19, 2011.

On exercise of the option, Warren retains a 2% NSR royalty with Metal Mountain having the

right to purchase 1% for $1,000,000.

A 100% interest in tenure Block D was purchased for $30,000 and 666,000 shares by Eagle

Peak from Quantum Speed Internet Products Inc. (Quantum) on September 25, 2008. On

October 16, 2009, the block was purchased by Metal Mountain. There is no NSR royalty

retained by Quantum.

Tenure Block E was optioned by Eagle Peak on July 1, 2008 from four partners: Anthony L'Orsa

(60%), Mary Sikkes (10%), Catherine L'Orsa (10%), Andrew L'Orsa (10%), and Suzanne L'Orsa

(10%) - the McKendrick Royalty Group. The option was sold to Metal Mountain on October 16,

2009, Metal Mountain has the right to acquire a 100% interest in the tenures by making

payments totaling $300,000 by July 1, 2015 and by issuing common shares totaling 150,000 by

December 31, 2010. On exercise of the option. the McKendrick Royalty Group is entitled to a

minimum annual royalty of $25,000 or a 2.5% NSR whichever is greater. Until July 1, 2018

Metal Mountain retains the right to purchase up to 1.25% of the royalty for $1,250,000.



Metal Mountain Resources Inc.

Dome Mountain ProjectBritish Columbia, Canada

Figure 1Location Map



Table 1Claim Details

Tenure Number Block Type Claim Name Good Until Area (ha)

238086 A claimREFER TO LOT

TABLE 2018.OCT.02 25.00238538 A claim COPE 1 2018.OCT.02 25.00308801 A lease 2010.SEP.14 54.78507597 A claim 2018.OCT.02 93.37507598 A claim 2018.OCT.02 74.70522324 A claim 2018.OCT.02 74.65548965 A claim 2018.OCT.02 373.14374166 B claim DOME 400 2012.JUL.02 500.00374168 B claim DOME 100 2012.JUL.02 500.00381072 B claim HOO 2012.JUL.02 25.00503165 B claim 2012.JUL.02 802.65503167 B claim 2012.JUL.02 485.32524847 B claim 2012.JUL.02 429.52525968 B claim HOO FRACTION 2012.JUL.02 18.67382560 C claim FREE GOLD - 1 2018.OCT.02 25.00382561 C claim FREE GOLD - 2 2018.OCT.02 25.00382562 C claim FREE GOLD - 3 2018.OCT.02 25.00382563 C claim FREE GOLD - 4 2018.OCT.02 25.00591933 D claim LITTLE MCKINNY 2012.JUL.02 37.32592283 D claim HILO 2012.JUL.02 447.59592285 D claim HILO 2012.JUL.02 466.25592286 D claim HILO 2012.JUL.02 466.24592288 D claim HILO 2012.JUL.02 466.41592289 D claim HILO 2012.JUL.02 466.21592290 D claim HILO 2012.JUL.02 466.09592291 D claim HILO 2012.JUL.02 465.99592292 D claim HILO 2012.JUL.02 447.21592293 D claim HILO 2012.JUL.02 447.21592294 D claim HILO 2012.JUL.02 447.21592295 D claim HILO 2012.JUL.02 465.89592296 D claim HILO 2012.JUL.02 223.59597303 D claim 2012.JUL.02 372.98329906 E claim DREA 2018.OCT.02 25.00557203 E claim 2018.OCT.02 111.81557458 E claim 2018.OCT.02 74.54557548 E claim 2018.OCT.02 74.56557615 E claim 2018.OCT.02 74.56572765 E claim 2018.OCT.02 372.98572766 E claim 2018.OCT.02 186.43582849 E claim 2018.OCT.02 111.79582851 E claim 2018.OCT.02 447.57582853 E claim 2018.OCT.02 223.70

total 10970.92



5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,INFRASTRUCTURE AND PHYSIOGRAPHY

AccessibilityThe claims are road accessible from Smithers by 64 km of mostly gravel all-weather roads.

From a point on Highway 16, 4 km south of Smithers, the route follows the Babine Lake

(Eckman) Road to km 39, then turns southeast on the Chapman Forest Service Road for 16 km

to km 69, and then winds generally uphill in a southwesterly direction for 4 km on the Dome

Mountain Mine access road to the 1290 Portal (see Figure 1).

ClimateThe area has a moderate climate with an average annual precipitation of approximately 510 mm

and annual snowfall of approximately two metres. The area is usually free of snow from June to

mid-October with temperatures ranging from a low of -40°C in December and January to a high

of 28°C in July and August.

Local ResourcesThe Town of Smithers, with a regional population of approximately 15,000, is a major centre for

resource industries operating in northwest B.C. It is located approximately 400 kilometres from

deep water ocean ports in Prince Rupert, Kitimat and Stewart, has an airport with daily service

to Vancouver, and has access to the to the CN rail-line. Several exploration companies and

diamond drill contractors have offices in Smithers. Smithers has readily available skilled mine

and construction labour as well as connections to electric power and natural gas.

InfrastructureSite infrastructure consists of two levels of drift development at the 1370 and 1290 metre

elevations. There is no surface infrastructure on site.

PhysiographyDome Mountain is a glacially rounded summit that reaches an elevation of 1 753 metres above

sea level and marks the most southerly occurrence of alpine terrain in the Babine Range.

Slopes on the mountain range from gentle to steep but cliffs are rare. Overburden cover

consists of alluvial clays, sands, and gravels overlying gravely boulder till. In the vicinity of the

Boulder Vein at approximately 1 300 metre elevation the overburden ranges from one to two

metres thick.



Vegetation cover consists of thick stands of mature balsam fir, lodge pole pine and spruce. At

elevations above 1 500 metres alpine meadows are common. Outcrop exposure on the

wooded slopes is poor and averages less than 1%.

The area is drained by several, small creeks, such as Fedral Creek and Boulder Creek, that flow

year round.



6 HISTORY

The Dome Mountain area has a long history of successful exploration that resulted in the

discovery of numerous gold bearing quartz-sulphide veins. The Boulder Vein has a complicated

history of development and production with various operators, option agreements and name

changes occurring over a short period of time between discovery by Noranda in 1985 and

cessation of operations in 1993. A synopsis of the exploration, development and production

history is listed in Table 2.

In 1989, D.R. Melling calculated a "reserve estimate" for the Boulder and Argillite veins of

318,312 tons at a grade of 0.345 oz/ton gold and 2.22 oz/ton silver. The Melling (1989)

estimate was subsequently reviewed by Derry Michener Booth & Wahl who reclassified the

Argillite Vein "reserves" of 70,316 tons grading 0.412 oz/ton gold and 3.45 oz/ton silver as

"possible reserves". These calculations and classifications pre-date NI 43-101 and should not

be relied upon for economic assessment.

Timmins Nickel Inc. commenced underground mining in August 1991 and stopped in 1993.

During this period 43,900 tonnes at an average grade of 12.0 grams per tonne gold were

reportedly mined from shrinkage stopes accessed from trackless drift developments on the

1290 and 1370 levels. The ore was shipped off-site to either the Equity Silver mill near

Houston, BC or to the Westmin Premier mill near Stewart, BC for toll milling.

In 1993 Roscoe Postle Associates Inc. (RPA, a predecessor company to Scott Wilson RPA)

prepared a historical resource estimate by a longitudinal, polygonal method. The "in situ proven

and probable reserves" reported for the Dome Mountain property were 181,780 tonnes grading

14.8 g/t gold, with "possible reserves " of 39,650 tonnes at an average grade of 12.6 g/t gold.

This is a historical estimate as defined by NI 43-101 and as such should not be relied upon for

investment decisions. These historical resources were contained within the following five zones

of the Boulder Vein System: the Boulder Main Vein, Boulder Vein HW, Boulder Vein FW,

Argillite Vein, and Argillite Vein HW. All high gold values were cut to 51.4 g/t gold.

EPR optioned the Block A tenures in 2007 and conducted orientation "Ultra-Trace" soil

geochemistry and 3D induced polarization (3DIP) surveys over the Boulder Vein System and its

projected eastern extension in 2008.



In 2009 EPR drilled 4817.2 metres in 42 HQ holes to fill in gaps in the Boulder Main Vein drill

pattern and to confirm the results of the historic drilling. Also in 2009 four exploration holes

totaling 888.2 metres were drilled to test coincident 3DIP and zinc soil geochemical anomalies.



TABLE 2HISTORY SUMMARY

Year Event

1898 Mineral occurrences on Dome Mountain first staked by W.B. Forrest

1923-24Surface and underground work was done by the Dome Mountain Mining Company Ltd. Workincluded 32 m of shaft sinking, 102 m of drifting and cross-cutting, and driving of adits on theForks Vein.

1924-80 No work recorded. Property was acquired by Silver Standard Mines Ltd., McIntyre Mines Ltd.,T. L'Orsa, K. Coswan, L. Warren and B. McGowen

1980 Panther Mines Ltd. and Reako Exploration Ltd. optioned L. Warren claims1981 Reako Exploration Ltd. optioned McIntyre Mines Ltd. claims1982 Panther Mines Ltd. and Reako Exploration Ltd. optioned Silver Standard Mines Ltd. claims

1984-85

Noranda Exploration Company Ltd. (Noranda) optioned claims from various parties andconducted extensive exploration work consisting of geological mapping, geophysical surveys,geochemical surveys, trenching and diamond drilling. The Boulder Vein was discovered bytrenching a zinc soil anomaly on the eastern strike extension of the Cabin Vein.

1985Canadian United Minerals Inc. (Canadian United) optioned the Noranda interest subject to aback-in right to re-acquire 50%. Canadian United then optioned a 75% interest to TeeshinResources Inc. (Teeshin).

1986 Canadian United drilled the Boulder Vein. Total Erickson Resources Ltd. (Total) acquiredNoranda's back-in rights.

1987

Canadian United formed a joint venture with Total and Teeshin. Surface and undergrounddiamond drilled, air-borne geophysical surveys ( DIGHEM III EM, magnetometer, and VLF-EM), and underground development (1370 adit) were carried out. The Argillite Vein wasdiscovered.

1988 Conceptual mine design and cost estimates were prepared by Dynatec Mining Limited.

1989 Teeshin became the operator and drilled 14 holes on the west and east extensions of theBoulder Zone. A feasibility study was completed by M.P.D. Consultants Inc.

1990 Teeshin acquired Canadian United's interest and drilled 18 diamond drill holes

1991Teeshin formed a joint venture with Timmins Nickel Inc. (Timmins). Teeshin changed its nameto Habsburg Resources Inc. (Habsburg). Mining commenced on the Boulder Vein and ore wasshipped direct to the Equity Silver Mill. The 1290 cross-cut was started.

1992 Mining Lease was approved. Mine operated with 28 employees.

1993 Mining was suspended due to Timmins' financial and legal problems. Total production was48,400 tons at an average grade of 0.35 oz/ton gold.

1994 Habsburg changed its name to Dome Mountain Resources Ltd.1996 Dome Mountain Resources Ltd. changed its name to DMR Resources Ltd. (DMR).2001 DMR is delisted

2005 DMR transferred ownership of the Mining Lease and their remaining claims to Angel JadeMines Ltd., K. Coswan, A. L'Orsa and J. L'Orsa (L'Orsa-Coswan-Angel Jade).

2007 Eagle Peak Resources Inc. (Eagle Peak) optioned the property from L'Orsa-Coswan-AngelJade.

2008 Eagle Peak conducted soil geochemistry and 3D induced polarization surveys over theBoulder Vein System and its projected extension to the east.

2009 Eagle Peak drilled 46 HQ diameter holes (4817.2 metres in 42 in-fill holes and 888.2 metres in4 exploration holes)



7 GEOLOGICAL SETTING

The following sections on the regional, local and property geological setting of the Dome

Mountain Project have been adapted from the Scott Wilson Roscoe Postle Associates Inc. NI

43-101 report dated April 28, 2008.

REGIONAL GEOLOGY

The Dome Mountain Project is situated in the Babine Range of west central British Columbia.

The Babine Range is a northwest trending horst of folded and faulted Jurassic and Cretaceous

volcanic and sedimentary rocks bounded to the west and east by grabens of Late Cretaceous

and younger rocks (Figure 3). The regional stratigraphy has been described by Tipper and

Richards (1976) and refined by MacIntyre et al. (1987).

Babine Range is underlain by Early to Middle Jurassic calc-alkalic island arc rocks of the

Telkwa, Nilkitkwa, and Smithers Formations. The Nilkitkwa Formation disconformably overlies

the Telkwa Formation which in turn disconformably overlies the Smithers Formation.

The structural setting is analogous to the Basin and Range province of the US Southwest and

structural development is probably related to Late Cretaceous to Early Tertiary extensional

tectonics. This tectonic event is characterized by northeast-trending shearing, which offsets the

horst and graben boundaries on major north-trending transcurrent faults. The structure of the

area is characterized by asymmetric to overturned, southeast-plunging folds that are truncated

by northeast-trending shear zones and northwest-striking high-angle reverse and normal faults.

LOCAL GEOLOGY

Lithology

The Dome Mountain area is predominantly underlain by the Lower to Middle Jurrasic Hazelton

Group island arc assemblage. The Telkwa Formation, at the base of the Hazelton Group, is the

thickest and most extensive formation. The Nilkitkwa Formation conformably to disconformably

overlies the Telkwa Formation and is an important host for mineral occurrences (Figure 4).



The Lower Jurassic Telkwa Formation has been subdivided into four mappable units which are

from oldest to youngest: (1) polymictic conglomerate (lJT1); (2) porphyritic andesite (lJT2); (3)

fragmental volcanic rock (lJT3); and (4) phyllitic maroon tuff (lJT4). Units 2 and 3 are

considered to be proximal vent facies rocks.

The Nilkitkwa Formation is composed of transgressive marine sediments that overlie rhyolite,

basalt and red epiclastic rocks. The formation has been subdivided into four mappable units. In

ascending stratigraphic order these units are (1) interbedded red epiclastics and amygdaloidal

flows (lJN1); (2) rhyolitic volcanic rocks (lJN2); (3) tuffaceous conglomerate, cherty tuff and

siltstone (lJN3); and (4) thin-bedded argillite, chert and limestone (lJN4).

The Smithers Formation (mJS) comprises fossiliferous sandstone and siltstone with intercalated

felsic tuff that was deposited during a marine transgression. It overlies the Nilkitkwa and Telkwa

Formations in a disconformable fashion. It is typically comprised of medium to thick-bedded,

dark grey limy siltstone and mudstone and weathers orange to brown. At Dome Mountain, the

thick-bedded siltstone grades laterally to a relatively thin unit of well-bedded dark grey

argillaceous limestone, limy siltstone, and wacke, with a few thin beds of pebble conglomerate

and chert.

Isolated fault bounded blocks of the Bowser Lake Group (Middle-Upper Jurassic Ashman

Formation) occur locally. These rocks conformably overlie the Smithers Formation. Late

Cretaceous to Tertiary lapilli tuffs and porphyritic andesite flows (uKEv) also outcrop locally in

fault bounded blocks

Outcropping intrusive rocks are rare on Dome Mountain. A few outcrops of dioritic intrusive

rocks with foliations parallel to the host rocks have been mapped and are considered to be

coeval with the Lower Jurassic volcanism. The 1987 airborne magnetic survey revealed several

positive magnetic features which suggest the presence of buried intrusives.

Structure

The predominant structural feature on the property is a southeast-trending, southeast plunging

and southwest-verging anticline. The lack of an axial planar fabric within this structure indicates



an origin due to vertical tectonic events. Doming over an inferred buried intrusive of Late

Cretaceous or Early Tertiary age is probable as suggested by a positive magnetic feature which

coincides with Dome Mountain. Alternatively, the vertical movements associated with the last

tectonic event could be considered as the probable cause of the anticlinal structure.

On a local scale, the sulphide bearing quartz veins are situated along east-trending shear zones

which are interpreted as structures reactivated during Late Cretaceous volcanism. The veins

trend both northwest and east-west, and are disrupted by northwest-trending post-ore faults.

The most prominent joint orientation is northeast, roughly perpendicular to major fold axes.

These steep, northwest-dipping C-joints also parallel prominent airphoto lineaments and several

major high-angle faults which offset stratigraphy.

PROPERTY GEOLOGY

The Dome Mountain Project consists of two principal zones of high grade gold-silver

mineralization known as the Boulder and Argillite Veins (Figure 5). This subdivision was

established by earlier mine workers for the purposes of "reserve" estimation and is a function of

vein orientation and host rock lithology. Both veins occur within folded fragmental volcanic

rocks of the Telkwa Formation and within amygdaloidal basalts and altered volcanic rocks of the

Nilkitkwa Formation. The Boulder Vein has hanging wall and footwall veins and the Argillite

Vein has a hanging wall vein. These additional veins are generally splays and shoots off the

main vein structures.

In addition to the Boulder Vein System, the project is host to the Cabin, Elk, Forks, 9800, Free

Gold, Ptarmigan, Eagle, Gem, Raven, Hawk, Chance, Hoopes, Jane, Elk and Pioneer veins.

The Cabin Vein is interpreted as the westward extension of the Boulder Vein. The other veins

mentioned are separate from the Boulder Vein system. A modest amount of drilling has been

carried out on these veins, but to date, no mineral resources have been defined.

The quartz veins are mineralized with a sulphide assemblage consisting of pyrite, sphalerite,

galena, and chalcopyrite. Wall rocks are typically altered and moderately deformed for several

metres on either side of the veins.



Vein Geometry and StructureIn Detail, the veins are not simple planar structures. They display variations in thickness, strike

and dip. They are gently curved or flexed and are concave towards the south. The veins occur

within a deformation zone averaging less than 10 m in thickness. The host rocks are

penetratively deformed (sheared) with foliation development most pronounced adjacent to the

veins. The veins and associated foliation cross-cut the bedding in the host rocks. The veins

display a diverse range of deformation structures. They may be massive, boudinaged,

brecciated, banded or tightly folded. Locally minor offsets occur along narrow shears which are

parallel to and at high angles to the veins.

The Boulder Main Vein has an average orientation of 100°/50°S and a strike length of

approximately 440 m. Dips tend to be steeper, 50° to 85°S, in the central and eastern portion of

the vein and flatter, 30° to 40°S, towards the western extremity. The vein varies in true width

from 0.7 m to 4.5 m but averages 1.45 m. Thickness and grade contours demonstrate that the

deposit pitches about 45° east within the plane of the vein. Small off-shoots or splays,

branching from the main vein structure, occur in the hanging wall and footwall of the main vein.

The mineralized zone is particularly thick in the areas of intersection. The thickness and grade

of the mineralized vein is most consistent when the hanging wall is amygdaloidal basalt.

The Argillite Vein has an average orientation of about 120°/41°S and a strike length of

approximately 240 m. It is a major splay or bifurcation of the Boulder Vein. The mineralization

varies in true width from 0.7 m to 4.75 m but averages 1.24 m. Correlations of the Argillite Vein

between sections are more difficult than for the Boulder Vein but may still be done with

reasonable confidence. The best Argillite Vein mineralization reportedly occurs where the shear

zone hosting the vein intersects less competent volcanic sediments. Small splays and offshoots

from the main structure are more common in the Argillite Vein.

AlterationEnveloping the Dome Mountain veins are alteration zones which extend several metres into the

wall rocks. These "bleached" zones are characterized by abundant carbonate, and sericite. In

close proximity to the vein contacts, the sericite is a distinctive lime green color. Locally,

euhedral pyrite is present in the altered zones. The alteration zones rarely contain significant

gold/silver mineralization except where they contain quartz-carbonate-sulphide stringers.



The Boulder Vein is characterized by a more pronounced alteration envelope than the Argillite

Vein - probably a function of host rock lithology. The correlation of alteration in section is an

important consideration for geological interpretation.

Alteration varies both in thickness and intensity and in general, gold mineralization and intensity

of alteration as positively correlated. Intensely altered rocks are schistose with an almost white

color and disseminated pyrite. Weakly altered rocks are marked by chlorite alteration of mafic

minerals.

MineralizationThe veins are characterized by quartz with lesser carbonate and sulphide mineralization.

Massive quartz-carbonate veins lacking sulphides are typically barren with respect to gold and

silver.

Quartz occurs as both a white massive variety and as a clear variety which is associated with

higher gold grades. Carbonate minerals (ankerite and calcite) occur as cream to beige crystals.

Small scale folds in the veins attest to continued movement after their formation.

Sulphide minerals in the Boulder Vein constitute approximately 10% of the vein mineralogy. In

decreasing order of abundance the sulphide minerals are: pyrite (6%), sphalerite (2.5%),

chalcopyrite (1%), and galena-tetrahedrite-arsenopyrite (<1%). Pyrite occurs are fine euhedral

cubic crystals disseminated throughout the wall rock alteration and quartz veins. Coarse

masses of pyrite also occur as well as some individual pyrite crystals up to one centimetre wide.

Often the pyrite crystals show evidence of crushing with the interstices filled with other

sulphides. Aggregates of fine-grained reddish brown sphalerite occur as irregular masses

associated with pyrite, galena, chalcopyrite and arsenopyrite. Chalcopyrite is commonly

intergrown with pyrite. Fine-grained tetrahedrite, galena and arsenopyrite occur as

disseminations, as thin fracture coatings, or as fine irregular masses with the other sulphides.

Even though gold grades as high as several grams per metric tonne are present, visible gold is

rare. Microscopic examination indicates that the gold usually occurs as minute grains along the

pyrite crystal margins and in microfractures within the pyrite crystals. Metallurgical test work



indicates an average grain size of 25 microns. Gold may be present as electrum since gold

analyses indicate 18% to 23% silver.

Silver values up to 514 grams per tonne have been reported from core assays although no

silver minerals have been identified. It appears that the silver values reflect the abundance of

galena and tetrahedrite as indicated by an analysis of tetrahedrite that contained 2% to 4%

silver.



8 DEPOSIT TYPES

The mineral deposits at Dome Mountain are structure-controlled orogenic (mesothermal) quartz-

carbonate-sulphide veins with associated gold and silver mineralization. Controlling structures

are east-west and northwest-southeast trending brittle fault zones that dip moderately to steeply

south and southwest. The host rocks are Lower to Middle Jurassic subaerial volcanic flows,

pyroclastic, and related volcanoclastic rocks.

According to Goldfarb et al (2005) orogenic gold deposits are generally located along deep-

crustal fault zones or in related second and third order shears and faults, particularly at jogs or

changes in strike along crustal fault zones. There is generally a spatial association with

granitoid bodies which are related to the orogenic event. Mineralization styles vary from

stockworks and breccias in shallow brittle regimes, through laminated crack-seal veins and

sigmoidal vein arrays in brittle-ductile crustal regions, to replacement and disseminated-type

mineralized bodies in deeper ductile environments. Most orogenic gold deposits contain 2 to 5

percent sulphide minerals (mainly pyrite and arsenopyrite) and have gold/ratios of 5 to 10 and

gold fineness >900. Host rock lithology is critical for concentration of gold in some provinces

where iron or carbon rich rocks along a flow path are important sinks for the release of gold from

hydrothermal solutions. Competency of the host lithology may influence the width of the

mineralized vein. Alteration intensity, width and assemblage may also vary with the host rock

lithology but carbonates, sulphides, muscovite, chlortie, K-feldspar, biotite, tourmaline and albite

are generally present.

The best known deposit of this type in the Canadian Cordillera is the Bralorne-Pioneer Mine

near Gold Bridge, BC.



9 MINERALIZATION

The most significant gold mineralization on the Dome Mountain Project is contained within five

zones (Figure 5) of the past-producing Boulder Vein System (listed in order of importance):

! Boulder Main Vein: This zone has a strike ranging from 090° to 100° and dips

ranging from 45° to 70° south. The vein has been defined over a strike length of 440

metres from section 652,900 UTME to 653,340 UTME and a down-dip extent of more

than 180 metres starting at the surface. The mineralized zone within the vein has a

shallow easterly plunge with an average horizontal width of 2.3 metres, ranging from

1.6 to 10.5 metres. The majority of the 43,900 tonnes of past production came from

this zone.

! Argillite Vein: The Argillite Vein is a splay off the Boulder Main Vein. It strikes at 120°

and dips to the southwest at angles ranging from 35° to 50° and it appears to flatten

along strike to the southeast. The zone is somewhat discontinuous over the defined

strike length of 240 metres and a maximum down-dip extent of 90 metres. The vein

has an average vertical thickness of 3.2 metres, ranging from 2.1 to 3.8 metres.

! Boulder Footwall Vein: This zone is sub-parallel to the Boulder Main Vein and is

located at the east end of the Boulder Vein System. It has been defined by drilling

over a strike length of 140 metres from 653,300 UTME to 653,440 UTME and over a

down-dip extent of approximately 200 metres. The zone has an average horizontal

width of 1.9 metres, ranging from 1.6 to 3.8 metres. The zone is open down-plunge

to the east and down-dip. Minor past-production came from this zone.

! Boulder West Hangingwall Vein: This vein is also sub-parallel to the Boulder Main

Vein. It is a somewhat discontinuous vein with an overall strike length of 100 metres

located between 652,960 UTME and 653,060 UTME and a defined down-dip extent of

approximately 100 metres. It has an average horizontal width of about 2.4 metres,

ranging from 1.8 to 3.8 metres.



! Boulder East Hangingwall Vein: This is a thin and somewhat discontinuous vein that

occurs over a strike length of about 50 metres at the east end of and sub-parallel to

the Boulder Main Vein.

In additional to the past producing Boulder Vein System , gold mineralization at the Dome

Mountain Project is present at twelve separate locations: Free Gold, Forks, 9800, Ptarmigan,

Elk, Eagle, Gem, Raven, Hawk, Chance, Hoopes, Jane Cabin, and Pioneer. The Free Gold and

Forks zones are classified as past-producers by the BC Government Minfile database while theothers are listed as either as prospects or showings. The Cabin Vein is considered to be the

westward extension of the Boulder Vein.

The veins occur in a roughly northwest-southeast 12 kilometre trend from southeast of Dome

Mountain to Mt. McKendrick. This trend may reflect the presence of a deep seated structure. A

modest amount of diamond drilling has been conducted on the various veins but to date no

minerals resources have been defined.

For detailed information on the mineralogy and structure of the veins and wallrock alteration,

refer to the Property Geology section of this report.



10 EXPLORATION

Exploration by previous operators on the property has been described in the history section of

this report.

In 2008 Eagle Peak Resources Inc. conducted soil geochemistry 3D induced polarization and

magnetic surveys over the Boulder Vein System and its projected extension to the east. These

surveys were designed as an orientation of the geochemical and geophysical signatures related

to the known veins and structures as an aid to further exploration. Drill hole DM09-046 was

drilled to target a coincident 3D induced polarization and zinc soil anomaly in the area of the

Chance Vein Minfile showing (Figure 5). The hole intersected 19.07 grams per tonne gold, 14.0grams per tonne silver and 6.70 % zinc over 0.6 metres from 49.4 to 50.0 metres down-hole

depth. The true thickness of the intersection is unknown.



11 DRILLING

The surface drilling database for the Dome Mountain Project is summarized in Table 3.

Additionally, thirty seven (37) short holes were drilled from underground. A complete listing of

the holes and collar information is attached as Appendix I. Eight holes without collar information

were removed from the database.

Table 3Surface Drilling Summary

Company Year Holes Metres Target

Noranda / Teeshin 1985 31 1,525 Exploration

Canadian United 1986 99 8,882 Boulder Vein

Canadian United / Teeshin / TotalErickson 1987 48 3,343 Boulder/Argillite Veins

Canadian United / Teeshin / TotalErickson 1988 10 1,339 Boulder/Argillite Veins

Teeshin / Total Energold 1989 20 2,457 Boulder/Argillite Veins

Teeshin 1990 18 2,326 Boulder/Argillite Veins

Habsburg / Timmins Nickel Inc. 1992 13 1,045 Boulder/Argillite Veins

Eagle Peak Resources Ltd. 2009 46 5,705 Boulder Vein

Totals 285 26,622

PRE-2009 DRILLING

The pre-2009 diamond drilling concentrated on the Boulder Vein System after the discovery of

the Boulder Vein in 1985. The information available from this work by various operators

consists of lithologic and structural drill logs, collar surveys, and assays. The collars were

surveyed relative to a mine grid but downhole surveys consisted of dip readings only. Assay

certificates show that gold and silver grades were determined by fire assay. This work pre

dates the implementation of NI 43-101 therefore no qualified persons are identified for the

programs.

DM09046DM09044

DM09043

DM09030

DM09020

DM09003

TS8728

TS8727

TS8719

TS8717

TS8716

TS8715

TS8706

TS8632

TS8624

TS8622

TS8620 TS8618

TS8616

TS8614

TS8610

TS8609

TS8608

TS8606

TS8601

TB8703

TB8702

TB8701

T8649

T8647T8646 T8645

T8644

T8643

T8642T8641

T8640T8639

T8637 T8633 T8631

T8629

T8627

T8624

T8622 T8620

T8616

T8613

T8610

T8609

T8607T8604

T8503

RP8824

RP8823

RP8822

RP8821

RP8709

RP8708

LM867

LM864

F8520

F8518

F8517

F8514 F8512

F8511F8508

DM8610DM8609

DM8608DM8607

DM8604DM8603

DM8601

D9213D9209

D9208

D9205D9204

D9203

D9202

D9201

D9018D9017

D9016

D9015

D9013

D9011

D9009

D9007D9006

D9001

D8913

D8911

D8909 D8907

D8906

D8905

D8904

D8903

D8901

C8533

C8531

Legend

Road

pre 2009 drill hole

2009 drill hole

0 200 m

652500 m 653000 m 653500 m

652500 m 653000 m 653500 m

6068500m

6069000m

6068500m

6069000m

Metal Mountain Resources Inc.

Figure 6

Drill Hole Location Map



In 2009 McElhanney Consulting Services Ltd. of Smithers BC converted the mine grid

coordinates to NAD83 UTM. The collar locations are plotted in Figure 6.

2009 DRILLING

During the period from July 28 to September 24, 2009, Eagle Peak Resources Inc. conducted a

drill program of 46 HQ holes totaling 5,705 metres. Most of the drilling was conducted on the

Boulder Vein system to in-fill the existing drill pattern and to confirm the results from the pre-

2009 drilling. The drill core is stored at the Metal Mountain warehouse in Smithers.

Driftwood Diamond Drilling Ltd. of Smithers, BC provided contract drill services with the

geological and field duties conducted by various hired independent consultants and contractors

working on behalf of Eagle Peak. The qualified person for the 2009 program was Daryl Hanson,

P.Eng.

The holes were logged under the supervision of the qualified person and stored in a digital

format. Metal Mountain Resources Inc. has a written procedure for logging, sampling, and

photographing core. The complete drill-hole logs with lithology, structure, alteration and assays

as well as the certificates of analysis are stored at the field office of Metal Mountain Resources

Inc. in Smithers, BC.

Down-hole surveys were conducted by the drill crew using a Reflex EZ-Shot hole survey tool.

The collar location of each hole was surveyed relative to the NAD83 UTM grid by McElhanney

Consulting Services Ltd of Smithers, B.C. using a Leica 803 Total Station instrument. The collar

locations are plotted on Figure 6.

North-south cross-sections were generated at 20 metre intervals along the Boulder Vein System

in order to test the along strike and down-dip continuity of the veins. Section 653,100 UTME is

presented in Figure 7 to demonstrate the down-dip continuity and extent of the Boulder Main

Vein.



12 SAMPLING METHOD AND APPROACH

PRE-2009 DRILLING

The details of the core sampling method and approach by previous operators are not in the

historic records.

2009 DRILLING

The intervals to be sampled were determined by the geologist at the time of logging and the

lengths of the intervals were adjusted to coincide with major lithological contacts. Samples were

taken from the quartz-carbonate-sulphide veins (QV) and the surrounding wall rock alteration

(Va and Vas). Minimum sample length was 0.2 metres. Half core samples were taken using a

diamond saw. A total of 932 sawn samples were taken for analysis.

A total of 70 specific gravity determinations were made from core samples covering all the major

rock types . The results are reported in the Mineral Resource Estimate section of this report.

Core recovery was consistently between 95 and 100% and there were no other drilling factors

that could materially affect the accuracy and reliability of the results.

There are no known sampling or geologic factors that could have contributed to sample bias.

A complete listing of intersection composites is provided in Appendix IV.

2009 UNDERGROUND SAMPLING

The vein exposures in the stopes accessed from the 1290 level were washed, mapped and chip

sampled. Samples were taken from the quartz-carbonate-sulphide veins (QV) and the

surrounding wall rock alteration (Va and Vas) at right angles to the vein. The minimum sample

width was 0.2 metres. A total of 193 chip samples were collected for analysis



13 SAMPLE PREPARATION, ANALYSES AND SECURITY

PRE-2009 DRILLING

The details of the drill core sample preparation, analysis and security by previous operators are

not complete in the historic records. Assay certificates indicate that samples were analyzed at a

variety of laboratories including Bondar-Clegg & Company Ltd., Chemex Labs Ltd., Acme

Analytical Laboratories Ltd., Rossbacher Laboratory Ltd., Min-En Laboratories Ltd., and

Kamloops Research & Laboratories Ltd. There is no record of Quality Assurance/Quality

Control (QA/QC) protocols.

A report dated March 12, 1990 by A. L'Orsa for Teeshin Resources Ltd. describes the results

obtained for 28 diamond drill holes. The holes, numbered D89-1 to D89-14 and D90-1 to D90-

14, were drilled on the Boulder Main and Argillite veins. All were NQ-size and the samples were

analyzed at Min-En Laboratories in North Vancouver, BC (L'Orsa, 1990). There is no mention

of assay QA/QC protocols.

2009 DRILLING

Samples were sent to Assayers Canada (Assayers) in Telkwa for preparation and then the

pulps were sent directly to Assayers Canada in Vancouver for gold and silver fire assays and for

ICP analysis. The certificates of analysis are stored at the Smithers field office and the

analytical procedure is documented in Appendix V. The preparation laboratory in Telkwa and

the analytical laboratory in Vancouver are certified by ISO 9001:2008.

Standards were included in each sample stream at a rate of 1 in 20 (5%) as a control on

laboratory accuracy, precision and bias. The ore reference standards used were CDN-GS-8A

and CDN-GS-11A prepared by CDN Resource Laboratories Ltd. of Langley, BC. GS-8A has a

recommended value of 8.25 ± 0.60 g/t gold while GS-11A has a recommended value of 11.21 ±

0.87 g/t gold (Appendix VI).

Blank samples of limestone aggregate were inserted into the sample stream randomly at a rate

of 5% as a check on laboratory contamination.



Twenty six (26) reject duplicate samples and sixty one (61) pulp duplicate samples were sent to

Acme Analytical Laboratories Ltd. in Vancouver for gold and silver analysis by fire assay. The

rejects were prepared at the Acme Preparation Laboratory in Smithers. The analytical

procedure is attached as Appendix VII. Both the Smithers preparation laboratory and the

analytical laboratory in Vancouver are certified by ISO 9001:2008.

Eighteen (18) quarter core field duplicate samples were submitted to Assayers for gold and

silver analysis by fire assay as a check on the sampling technique and variability of the

mineralization.

To check for possible bias due to the "nugget effect", eleven (11) coarse reject samples were

submitted to Assayers Canada for a metallic gold fire assay analysis. The analytical procedure

is presented in Appendix VIII.

A “chain of custody” was maintained from the drill to the preparation laboratory to ensure

sample security.

2009 UNDERGROUND SAMPLING

The sample preparation, analysis and QA/QC protocols for the underground samples was the

same as for the 2009 drill core samples.



14 DATA VERIFICATION

Original drill logs and assay certificates were used to verify and correct the historic digital

database which had been converted to MS Excel from PCExplore software. Original assay

certificates were not found for 991 (38%) samples.

In order to validate the pre-2009 sampling for the purpose of using the data in resource

estimation, cumulative frequency plots were made to compare the gold assay results from the

2009 drilling with the results from the 1985-93 drilling. The comparison was made within the

same volume between UTM coordinates 652940 E to 653150 E, 6068840 N to 6069000 N, and

1230 to 1370 metre elevation. This volume approach was chosen over hole twinning as it was

felt to be more appropriate for this type of deposit. The results of the comparison are presented

in the section on Mineral Resource Estimate.

STANDARDS

Figure 8 shows the Assayers Canada results for the gold standards as a percentage of their

expected value vs time.

Figure 8: Standards vs. Time Plot



In general the results are quite good but the trend of the line is rising over time, indicating a

slight positive bias of up to 4% starting in about mid-September..

QA/QC failures for the standards submitted are listed in Table 4 along with the action taken. The

criteria to flag a standard as a failure is based on a Z-score > ABS(3) AND %Diff > ABS(10).

Table 4QA/QC Standard Failures

Certificate Sample BHID Standard BADZ-score %Diff Action

9S0010RA A060053 DM-09-008 CDN-GS-11A -6.39 -24.8 Repeat A060049-A0600539S0014RA 60264 DM-09-020 CDN-GS-11A -4.00 -15.52 Repeat 60262-602669S0014RA 60286 DM-09-020 CDN-GS-8A +3.50 +12.73 Repeat 60284-60288

9S0021RA 60496 DM-09-027 CDN-GS-11A -4.92 -19.09Repeat 60493-60498 (thisincludes a bad lab dup for Ag in60493)

BLANKS

The analytical results for the limestone blank vs time are shown in Figure 9. There were no

blank failures.

Figure 9: Blank Limestone vs. Time Plot



PULP DUPLICATES

Of the sixty one pulps, twenty one were below the Acme detection limit of 0.17 g/t gold and

another 3 samples were below the practical detection limit of 0.85 g/t. The remaining thirty

seven sample pairs were used in the comparison. The average grade for Acme for these 37

assays is 9.74g/t, while Assayers Canada averaged 10.18g/t. Acme’s average was 4% lower.

The scatter plot for the thirty seven pairs (Figure 10) shows a tight scatter with Acme results

higher at the low end and lower on the high end. There were 27 cases where Acme assays

were lower, and only10 cases where they were higher, indicating a bias for lower Acme assays.

Figure 10 - Pulp Duplicate Scatter Plot

Dome Mountain CHECK ASSAYS from PULPS - Au, g/t - Above Detection Limit

20%

20%

-20%

-20%

y = 1.092x0.9391

R2 = 0.9792

0.100

1.000

10.000

100.000

0.100 1.000 10.000 100.000

ASSCAN Original Au, g/t

ACMERe

peatAu

,g/t



REJECT DUPLICATES

The scatter plot for gold check assays is shown below in Figure 11. Three detection limit

samples were excluded leaving 23 pairs for comparison. The plot shows a slight bias to lower

Acme assays at the high end. The scatter is quite tight, with R2 equal to 0.98. There are 21

data pairs above the Acme practical detection limit of 0.85 g/t. In fourteen cases the Acme

results were lower and only seven are higher. The average for Assayer’s Canada for these

samples is 14.61g/t gold, while Acme’s average grade is 12.70g/t gold. Acme is 13% lower than

Assayers Canada.

Figure 11 - Reject Duplicate Scatter Plot

Dome Mountain Project - Au Check Assays done on REJECTS from Core

20%

20%

-20%

y = 0.9536x0.9674

R2 = 0.961

0.10

1.00

10.00

100.00

0.10 1.00 10.00 100.00

Original Assayers Canada Au, g/t

AcmeRepeatAu

,g/t



FIELD DUPLICATES

Quarter core samples are not true duplicate samples. The scatter plot (Figure 12), however,

shows a reasonable correlation between the sample pairs. The half core data has a mean of

11.7 g/t gold and a standard deviation of 23 while the quarter core data has a mean of 12.6 and

a standard deviation of 20.0. Eight of the quarter core samples have gold values higher than

the half core samples while nine are lower and one has essentially the same value.

Figure 12 - Core Duplicate Scatter Plot

y = 1.5796x0.7434

R² = 0.7776

0.1

1

10

100

0.01 0.1 1 10 100

oneha

lfcore

original(g/t

Au)

one quarter core duplicate (g/t Au)

DomeMountain Project ‐ Gold Check Assays on Core Duplicates



METALLIC GOLD DUPLICATES

This is a very small sample and not representative of the total grade distribution as seven of the

samples had a calculated grade of less than 5 g/t. All eleven samples tested contained coarse

gold and seven samples had greater than 10% coarse gold as a percentage of the total gold

(Figure 13). More samples representing the complete range of the grade distribution are

required for the results to have a meaningful statistical significance.

Figure 13 - Metallic Gold Results

Even though there is a significant portion of coarse gold in the samples tested, a comparison of

metallic assays and fire assays shows excellent correlation (Figure 14). Again this may be due

to the small sample size and more work is required to prove this correlation.

0.0000

2.0000

4.0000

6.0000

8.0000

10.0000

12.0000

14.0000

16.0000

1 2 3 4 5 6 7 8 9 10 11

Fine vs Coarse Au in mg

Coarse_Au_mg

Fine_Au_mg



Figure 14 - Metallic Duplicate Scatter Plot

y = 1.0522x0.9509R² = 0.988

0.10

1.00

10.00

100.00

0.10 1.00 10.00 100.00

MA,Au

g/t

FA, Au g/t

Fire Assay vs Metallic Assay



15 ADJACENT PROPERTIES

Guardsmen Resources Inc. is the registered owner of several mineral tenures adjacent to the

Dome Mountain Project. These tenures contain the following Minefile occurrences:! 093L 024 Ascot prospect

Galena-sphalerite-barite showings were discovered near the head of Canyon Creek in

1951 but not explored at that time. The earliest significant work, in 1967-1969, was by

Texas Gulf Sulphur Company which acquired the property as a result of a

reconnaissance silt survey. Geological mapping, soil geochemistry and EM surveys were

done, followed by one drill hole. During the 1970s and 1980s the main showings were

re-staked and/or optioned numerous times.

Minor programs were conducted by prospectors and companies, including Geostar

Mining Corporation (1984), Noranda Exploration Company (1985) and Canadian United

Minerals Ltd. (1986). A comprehensive program was conducted in 1987 by Geostar

Mining Corporation. It included backhoe trenching which revealed several new mineral

occurrences. The most recent work was performed by Alliance Mining Inc. in 1996.

located at approximately 8 km northwest of the Boulder Vein System

! 093L 041 Ophir showing

The Ophir showing is host to a light grey medium-grained felsic tuff with quartz eyes that

contains 3 per cent pyrite as disseminations and in fracture filling along with minor

chalcopyrite. Quartz-carbonate and chlorite veinlets crosscut the felsic tuff.

Minor prospecting and geochemistry was conducted by Freemont Gold Corp. in 1985.

! 093L 332 Peggy (dome South) showing

The showing, a prospecting discovery in 2007, consists of silver-copper mineralization in

a zone of altered limestone cut by numerous quartz-sulphide stringers. Hand trenching

in the area of the showing exposed a barite-rich, siliceous sequence of volcanic

sediments with some areas displaying disseminated chalcopyrite. Chip samples

collected in 2007 graded up to 83.9 parts per million silver and 9,620 parts per million

copper over one metre



In 2009 Golden Odyssey Mining Inc. drilled nine holes on the Dome South project including six

on the Peggy showing, two on a 3D chargeability anomaly and one on a magnetic high

anomaly. Assays from the Peggy showing returned low concentrations of silver and base

metals.



16 MINERAL PROCESSING AND METALLURGICAL TESTING

The most recent metallurgical testing was initiated by Eagle Peak Resources Inc. in 2009. The

work is being conducted by PRA Metallurgical Division (PRA) of Inspectorate America Corp. in

Richmond, BC under the direction of Matt Bolu, P.Eng. The purpose of the testing was to

investigate the recovery of gold and silver by flotation, cyanide leaching and heap leaching at

the pre-feasibility level.

Approximately 1000 kg of sample was taken in equal amounts from the broken muck in six

underground drawpoints. The sample was collected under the direction of Daryl Hanson of IGS

and shipped to PRA for preparation, analysis and testing. The material from each drawpoint

was assayed separately and then combined in equal amounts to form a master composite

sample. The head grade of the master composite was 13.3 g/t gold and 61.7 g/t silver. The full

scope of the PRA work is presented in Appendix XI and a summary of the test results to-date is

presented below. These are preliminary results as the PRA report and the process flow sheet

have not been finalized at the present time.

MineralogyThis section on mineralogy is taken directly from the report titled "Petrographic Evaluation of a

Zinc, Copper, Lead, Silver and Gold-bearing Composite From Dome Mountain, British

Columbia" by Inspectorate America Corp. Petrographic Services dated November 23, 2009

(Appendix X).

"Pyrite was the dominant sulfide present in the 0905808-H Master Composite, comprising

5.15% of the sample by bulk modal analysis. Other sulfides present in the composite include

sphalerite (1.51%), chalcopyrite (1.21%), galena (0.19%) and chalcocite (0.02%). Trace

amounts (<0.01%) of bornite were observed. Sulfides were commonly multiply associated with

one another, with pyrite-sphalerite, pyrite-galena and pyrite-chalcopyrite associations the most

common, and sphalerite-chalcopyrite and sphalerite-galena associations also common.

However, gangue was typically non-opaque (91.3%), with silicates (mostly quartz, 44.5% with

accessory micas, 19.2% muscovite, clays, 5.6%, and feldspars, 2.8%) dominant. Carbonates

were also present, at 19.2%, while minor iron oxides were also detected (0.6%). Sulfates were

not detected above the detection limit of 0.5% by Xray diffraction; wet chemistry by PRA gave

0.03% SO4 2- concentration. Graphite was not observed, though graphitic carbon assay by Leco



by PRA indicated 0.05% graphite was present. Hosting minerals tended largely to be non-

opaque as well. Silver minerals were not observed, therefore it is likely that silver resides in

another mineral, possibly galena. It is also possible that silver minerals are submicroscopic.

The liberation and locking profile of the bulk sulfides indicate that about 84% of the sulfide

grains were liberated, and 16% were locked. Considering the mean sizes of liberated and

locked sulfides (42µm and 188µm, respectively), the weight percent sulfide liberation is 6% by

volume-weighted average.

Gold metal grains are primarily observed to be associated with pyrite (57% statistically and 98%

by volume weighted mean), though some of the gold-hosting pyrite is itself encapsulated by

other gangue minerals, primarily non-opaque gangue. The hosting minerals (encapsulating

grains which contained the gold and its tertiary, quaternary and quinternary associations) were

statistically non-opaque gangue (41%), and secondarily pyrite (34%). Other hosting minerals

included iron oxides, chalcopyrite and sphalerite. In all, 64% (statistically) of the gold

associations involved sulfides (99% by volume weighted mean). Small monomineralic gold was

also present at low levels. All of the gold observed was slightly pale, suggesting that some

electrum component may be present. The gold averages 51µm in diameter, and ranges from 1-

1060µm. One ~1mm gold nugget was observed, and was associated with non-opaque gangue.

As no other nuggets were observed, this occurrence was ignored for the purposes of calculating

volume-weighted mean associations, though it should be noted that a certain nugget effect may

cause data reproducibility problems, as nuggets may skew some test results towards free gold.

With the current 10-mesh crush, much of the gold, aside from random large nuggets, is

statistically leaching resistant. Moreover, some of the gold-hosting sulfides are themselves

encapsulated by large gangue grains. Grinding to free these grains should be prioritized. The

size distribution of the locked sulfides suggests that at least a 150µm crush would begin to free

the hosting sulfides. The gold itself would then require about a 50µm crush and leach. Preg-

robbing phases such as graphite and pyrrhotite are rare; the chief source of carbon in the

composite is the mineral ankerite. Overall, the sulfide association data shows that high gold

recoveries may be achieved by targeting pyrite. Leaching should also be employed to

sequester monomineralic gold and any remaining gold nuggets that remain."



Flotation (H.M. Bolu, personal communication)Four flotation tests were conducted at a nominal primary grind size 80% passing (P80) of 105

mm using conventional reagents such as potassium amyl xanthate (PAX), a sulphide mineral

collector; Methyl IsoButyl Carbinol (MIBC) a frothing agent; dithiophosphate promoter (Aerofloat

208), a collector for native gold and silver; and copper sulphate solution (CuSO4), an activator

of sulphide minerals with precious metal content. Optimization tests for reagent types and

consumption rates have not been carried out and may be done at a later project phase.

The results from these tests indicated that over 96% of the gold and over 98% of the silver could

be recovered into rougher concentrates at approximately 16% mass pull. The rougher

concentrate grades varied from 80-93 g/t Au and 426-470 g/t Ag. Cleaner flotation of the

rougher concentrates after regrinding to a nominal P80 size of 30 mm increased precious metal

grades to 224 g/t Au and 939 g/t Ag. The resulting recoveries were 92% for gold and 83% for

silver. These recoveries are expected to increase slightly in a closed circuit locked cycle test.

Cyanide Leaching (H.M. Bolu, personal communication)Whole ore cyanide leach tests were conducted using conventional process conditions as widely

practiced in the industry. Results and test conditions from the most relevant bottle roll leach

tests are summarized in Table 5.

Table 5Cyanide Leaching Test Results

Test No C2 C3 C4Sample ID Master Comp. Master Comp. Master Comp.Test Conditions

P80 Size mm 190 106 53% Solids 40 40 40NaCN g/L 1 1 1Leach time, h 72 72 72

Measured Headg/t Au 13.3 13.3 13.3g/t Ag 61.8 61.8 61.8

Calculated Headg/t Au 12.65 12.02 12.14g/t Ag 65.3 66.8 65.4

Extraction% Au 86.5 92.4 96.0% Ag 28.3 30.3 33.6

Residue Gradeg/t Au 1.71 0.91 0.48g/t Ag 46.8 46.6 43.4

Consumption (kg/t)NaCN 1.55 1.63 1.73Lime 0.4 0.4 0.4



The above test results indicate that the samples as tested were highly amenable to cyanidation,

producing high precious metal recoveries in leach solutions. These two-kg tests were

conducted at different grind sizes of 190 mm, 106 mm and 53mm, all at 80% passing particle

size (P80). Other test parameters such as leach slurry solids density, sodium cyanide

concentration and leach time were kept constant at 40% solids, 1.0 g/L and 72-h, respectively.

Gold recoveries in leach solutions varied from 86.5% for the coarse grind size of P80=190 mm

to 96.0% for the finest grind size of P80=53 mm. The intermediate grind size of P80=106 mm

resulted in 92.4% gold recovery. Silver recoveries on the other hand were lower and varied from

28.3% for the test with the 190 mm P80 grind size to 33.6% for the 53 mm P80 grind. As

expected, leach recoveries increased with increasing fineness of grind due to increased particle

liberation and exposure of precious metal surfaces to leaching. In subsequent tests, the

intermediate grind size of 106 mm P80 was used as the basis when investigating other process

conditions or variables.

Sodium cyanide consumption rates for the above tests were modest and varied from 1.55 kg/t of

feed solids for the 190 mm P80 grind size to 1.73 kg/t of feed solids for the 53 mm P80 grind

size. This is also consistent with metallurgical expectations that finer grind size results in

increased cyanide consumption rates due to increased exposed mineral surfaces for leaching.

The test with the intermediate 106 mm P80 grind size resulted in 1.63 kg sodium cyanide

consumption per tonne of feed solids. On the other hand, lime consumption rates for all three

tests were constant at a modest 0.4 kg/t of feed solids.

Extending leach residence times to 96-h along with increased cyanide concentration level of 2.0

g/L in leach solutions improved leach recoveries approximately 1% for gold and 9% for silver

while increasing sodium cyanide consumption rate to 3.0 kg/t of feed solids. Increasing leach

slurry density to 50% solids at 1.0 g/L cyanide concentration level and 96-h leach residence

time also resulted in increased leach recoveries of approximately 1% for gold and 4% for silver.

Sodium cyanide consumption rate for this test was 1.44kg/t of solids while lime consumption

was 0.5 kg/t.

A pre-manganese leach ahead of cyanidation, to free silver from silver-manganese minerals, did

not result in improved silver recoveries. Complete results of the tests are attached as Appendix

XII.



Cyanide leaching of flotation concentrates also showed promising results with unit recoveries

that ranged from 97% to nearly 99% for gold and 24%-23% for silver. However, overall

recoveries after flotation and including losses from flotation were 92-91% for gold and 22-19%

for silver. Sodium cyanide consumption rates varied from 1.36 kg/t to 2.02 kg/t of flotation feed

solids. Lime consumption rates were at less than 0.03 kg/t of flotation feed solids.

A 10-day diagnostics heap leach test on minus 1.5 inch crushed sample showed gold and silver

recoveries of 15% and 11% respectively. Due to high grade residue assays including the fine

size fractions, the heap leach tests were not pursued beyond this point.

Gravity (H. M. Bolu, personal communication)A single gravity separation test at a primary P80 grind size of 105 mm showed that 16% of the

gold and 3% of the silver in the feed can be recovered into a concentrate of 0.18% of the feed

weight. The concentrate grades were 1,180 g/t Au and 1,219 g/t Ag. Additional gravity testing

to develop a mass vs. grade/recovery relationship is in progress.



17 MINERAL RESOURCE ESTIMATES

There has been no attempt to confirm the historical geological resources that have been

reported from the property by earlier workers. Previous resource calculations predate the

implementation of National Instrument 43-101.

In March 2010 Gary Giroux, P.Eng. completed a technical review of all drill hole data and

produced a resource estimate for the Boulder Vein System. The estimate conforms to modern

requirements of the CIM Standards on Mineral Resources and Reserves – Definitions and

Guidelines (CIM, 2000). Gary Giroux is a consulting geological engineer with extensive

experience calculating mineral resources and qualifies as an Independent Qualified Person as

defined in National Instrument 43-101.

Details of the methodology are described in the following sub-sections written by Gary Giroux.

Data AnalysisThe provided data consisted of 330 drill holes with a total of 6,355 assays for gold and silver. A

total of 2,809 samples reported missing or -1 values for gold and were set to 0.005 g/t a value ½

the detection limit as were 6 samples reporting 0.000 g/t for Au. For silver 2,963 samples at -1

and 81 at 0.000 g/t were set to 0.05 g/t Ag a value ½ the detection limit. For the purpose of this

resource estimate only a portion of this total data was applicable. The resource was confined to

the limits described below.

The various identified veins were modelled by Moose Mountain Technical Services of Elkford,

BC using MineSight software (see Appendix IX) and are shown in Figure 15 below.



Figure 15: Three Dimensional Vein Solids

Assays were tagged with various flags designating mineralized Zone, lithology present and

colour of rock. Simple statistics for gold and silver were produced for each set of codes to

determine which coding method best outlined the mineralized zone.

AlphaZonecodes Alpha Description Zone CD

‐1 no code or zone assignment ‐1BLDR Boulder 100BFW Boulder footwall 150ARG Argillite 200EHW East Hanging‐wall 300WHW West Hanging‐wall 400

HWUncorrelated Hanging‐wallintersects 500

FW Uncorrelated Foot‐wall intersects 900

Assays from within the area of interest were sorted first by mineral zone and the statistics are

tabulated below in Table 6.



Table 6Assay Statistics Sorted by Zone

Zone Variable Number Mean StandardDeviation

Minimum Maximum CoefficientOf Variation

BLDR Au 324 14.14 25.42 0.005 331.83 1.80Ag 324 68.55 83.53 0.05 524.60 1.22

BFW Au 20 10.75 14.49 0.005 50.19 1.35Ag 20 40.96 55.54 0.05 210.10 1.36

ARG Au 122 16.38 24.87 0.005 142.67 1.52Ag 122 75.57 140.42 0.05 1028.40 1.86

EHW Au 13 24.68 22.82 1.82 72.37 0.92Ag 13 123.15 116.58 4.60 373.00 0.95

WHW Au 110 9.73 19.56 0.005 129.00 2.01Ag 110 46.57 71.34 0.05 350.30 1.53

HW Au 171 16.35 32.68 0.005 233.79 2.00Ag 171 91.76 174.86 0.05 853.90 1.91

FW Au 42 5.91 6.66 0.03 28.59 1.13Ag 42 29.02 34.73 0.05 141.90 1.20

Waste Au 3,693 0.24 1.23 0.005 50.39 5.15Ag 3,693 2.10 8.92 0.05 249.90 4.25

Clearly the 802 assays from all mineralized zone codes contain high average grades for gold

and silver. It is also apparent that significant grade exists in the un-coded or waste zones with

gold values up to 50.39 g/t and silver up to 249.90 g/t. Sorting the data by lithology, shown

below in Table 7, indicates the best grade occurs in material coded either quartz vein or fault

(Code = 95 or 85). However lithologies Altered Volcanic with quartz stringers (Code 81) and

Intercalated Argillites and Tuffs (Code 110) all contain average values of gold greater than 1 g/t

again indicating the presence of interesting mineralization outside and perhaps on the shoulders

of the quartz veins.



Table 7Assay Statistics Sorted by Lithology

AlphaLithologycodes Alpha Description Lith CD

Number Mean Au(g/t)

Number Mean Ag(g/t)

AB amygdaloidal basalt 12 444 0.23 444 1.04AR argillite 24 2 0.005 2 0.05AT Ash tuff 20 166 0.32 166 1.97BTm bedded maroon tuff 30 264 0.01 264 0.15CV Calcite vein 92 2 0.14 2 4.45Dk Dyke (intrusive?) 72 21 0.005 21 0.07Dkac rhyolite dyke 71 1 0.005 1 0.05DkinLT

andesite dyke and andesiteAndesite lapilli tuff

7015

13486

0.0070.19

13486

0.520.81

DT Cherty fine grained tuff (RD?) 21 12 0.009 12 0.09FP feldspar porphyry flow/tuff 60 192 0.18 192 1.02FT Fault 85 24 2.38 24 28.67NC ?? 75 1 0.005 1 0.05OB overburden 0 181 0.02 181 0.09QV Quartz vein 95 484 18.09 484 90.94RD Rhyodacite tuff? 90 2 0.005 2 0.05Va altered volcanic 80 1,383 1.06 1,383 6.40Vas altered volcanic w/ its stringers 81 543 2.30 543 13.00VS intercalated argillite and tuffs 110 203 1.44 203 7.55

XT Crystal tuff (FP?) 115 19 0.13 19 0.91

The colour codes while more ambiguous indicate the highest average grades in Colour White

(Code 80), which presumably describes the quartz veins. Again, however, significant

mineralization is indicated outside the veins, with average gold grades greater than 1 g/t in

rocks coloured bleached (Code 5), grey (Code 45) and un-coded (Code -1). The results are

summarized in Table 8.



Table 8Assay Statistics Sorted by Colour

AlphaColourcodes Alpha Description

ColourCD

Number Mean Au(g/t)

Number Mean Ag(g/t)

‐1 no code ‐1 138 2.02 138 11.91na no code 0 181 0.02 181 0.08BLE bleached 5 1,442 1.57 1,442 9.29BLK black 10 2 0.005 2 0.05G‐B green‐blue (buff?, beige?, bleached?) 25 14 0.44 14 0.72G‐G green‐grey or grey‐green 30 35 0.04 35 0.32G‐M green‐maroon 35 87 0.30 87 0.99GRN green 40 611 0.577 611 2.96GRY grey 45 490 1.82 490 11.11MAR Maroon, 55 874 0.16 874 0.77M‐G maroon‐green (grey?) 65 97 0.09 97 0.99RED red 70 73 0.06 73 0.59WHT white 80 451 18.06 451 89.57

The grade distributions for gold and silver were examined for each of the mineralized veins andwaste using lognormal cumulative frequency plots. In each case a capping level wasdetermined and samples above this level were reduced. The capping results are summarized inTable 9.

Table 9Capping Results Sorted by Zone

Zone Variable Cap Level (g/t) Number of AssaysCapped

BLDR Au 109.0 2Ag 400.0 3

BFW Au 77.0 0Ag 172.0 1

ARG Au 156.0 0Ag 548.0 2

EHW Au 101.0 0Ag 526.0 0

WHW Au 38.6 4Ag 606.0 0

HW Au 150.0 2Ag 1084.0 0

FW Au 12.0 3Ag 100.0 2

Waste Au 5.2 18Ag 88.0 6



The results of capping are compiled in Table 10.

Table 10Capped Assay Statistics Sorted by Zone



BLDR Au 324 13.25 17.59 0.005 109.00 1.33Ag 324 67.95 80.79 0.05 400.00 1.19

BFW Au 20 10.75 14.49 0.005 50.19 1.35Ag 20 39.06 50.09 0.05 172.00 1.28

ARG Au 122 16.38 24.87 0.005 142.67 1.52Ag 122 70.73 114.34 0.05 548.00 1.62

EHW Au 13 24.68 22.82 1.82 72.37 0.92Ag 13 123.15 116.58 4.60 373.00 0.95

WHW Au 110 7.68 10.24 0.005 38.60 1.33Ag 110 46.57 71.34 0.05 350.30 1.53

HW Au 171 15.37 26.85 0.005 150.00 1.75Ag 171 91.76 174.86 0.05 853.90 1.91

FW Au 42 4.97 4.16 0.03 12.00 0.84Ag 42 27.42 30.15 0.05 100.00 1.10

Waste Au 3,693 0.21 0.64 0.005 5.20 3.11Ag 3,693 1.99 7.05 0.05 88.00 3.55

Drilling on the Dome property has been completed between the years 1985 to 1993 where a

proper QA/QC program was not in place and again in 2009 when QA/QC procedures were used

as documented in earlier sections. To validate the earlier data a comparison was made for gold

grades from quartz veins (Code 95) within the same volume of rock assayed in different periods.

The comparison was made within the same volume contained between the coordinates 652940

E to 653150 E, 6068840 N to 6069000 N and 1230 to 1370 elevation. A lognormal cumulative

frequency plot is used to compare the two different gold grade distributions (see Figure 16).

The comparison is reasonable with older holes showing slightly higher grades above 7 g/t. This

is likely due to the fact that some of the higher grade material has been mined out and was not

available to be sampled in the 2009 drill holes.



Figure 16: Cumulative Frequency Plot for Au in Pre-2009 vs. 2009 Drilling

CompositesThe drill holes were “passed through” the domain solids with the entry and exit point marked for

each hole. Uniform down hole composites, 1 m in length, were formed to honour the domain

boundaries. Capped values for Au and Ag were used to form the composites. Small intervals,

less than 0.5 m at the domain boundaries, were combined with the adjoining sample to produce

a composite file of uniform support (1.0 ± 0.5 m). Table 11 gives the composite statistics sorted

by zone.

Cumulative Frequency Plot comparing Au assays from Different Periods

10 -3

10 -2

10 -1

1

10 1

10 2

10 3AU(g/t)

Percent

10 -3

10 -2

10 -1

1

101

102

103

0.1

0.5

1.0510203040506070809095

99.0

99.5

99.9

0.1

0.5

1.0

51020304050607080909599.0

99.5

99.9

10 -3

10 -2

10 -1

1

10 1

10 2

10 3

AU(g/t)

Percent

10 -3

10 -2

10 -1

1

101

102

103

0.1

0.5

1.0510203040506070809095

99.0

99.5

99.9

0.1

0.5

1.0

51020304050607080909599.0

99.5

99.9

Au from Holes drilled1985-1993

Au from 2009 Holes



Table 111 m Composite Statistics Sorted by Zone



BLDR Au 242 14.11 16.63 0.005 108.07 1.18Ag 242 69.20 75.85 0.05 400.00 1.10

BFW Au 21 9.69 14.43 0.005 50.19 1.49Ag 21 33.85 42.42 0.05 140.20 1.25

ARG Au 86 13.66 18.75 0.005 93.64 1.37Ag 86 59.76 89.12 0.05 384.86 1.49

EHW Au 10 19.71 17.54 1.82 59.36 0.89Ag 10 94.19 89.98 4.60 291.90 0.96

WHW Au 91 8.42 10.76 0.005 38.60 1.28Ag 91 52.39 82.36 0.05 350.30 1.57

HW Au 142 12.96 22.42 0.005 150.00 1.73Ag 142 73.65 136.98 0.05 749.44 1.86

FW Au 32 4.62 4.19 0.030 12.00 0.91Ag 32 25.82 29.43 0.05 100.00 1.14

Waste Au 26,341 0.04 0.24 0.005 5.20 6.35Ag 26,341 0.41 2.93 0.05 88.00 7.23

VariographyPairwise relative semivariograms for gold and silver were produced for each of the veins with

sufficient data to model. Anisotropic nested structures were fit to the Bolder Vein (Code 100),

Argillite vein (Code 200) and the West Hanging Wall vein (Code 400). The HW and FW veins

(Codes 500 and 900 respectively) did not have sufficient geologic information to build 3

dimensional solids and as a result were not modeled. The Boulder FW vein (Code 150) was

similar enough to the Boulder Vein to apply the Boulder variography and the West Hanging Wall

model was used for the East Hanging Wall. For all veins modeled the principal directions were

along strike, down dip and across dip. The across dip direction had too little information to

model but the first pair was used to establish the nugget effect and nominal ranges of 5 and 10

m were assumed for the two structures. The semivariogram parameters for each vein are

summarized in Table 12 and the models are shown in Appendix II.



Table 12Summary of Semivariogram Parameters

Vein Variable Az. Dip Co C1 C2 Short Rangea1 (m)

Long Rangea2 (m)

BLDR Au 90 0 0.50 0.10 0.25 20 600 -35 0.50 0.10 0.25 5 10180 -55 0.50 0.10 0.25 20 50

Ag 90 0 0.40 0.20 0.32 28 600 -35 0.40 0.20 0.32 5 10180 -55 0.40 0.20 0.32 20 40

ARG Au 100 0 0.60 0.50 0.40 20 5010 -43 0.60 0.50 0.40 5 10190 -47 0.60 0.50 0.40 20 36

Ag 100 0 0.70 0.40 0.44 30 6010 -43 0.70 0.40 0.44 5 10190 -47 0.70 0.40 0.44 15 20

WHW Au 90 0 0.45 0.20 0.65 5 200 -42 0.45 0.20 0.65 5 10180 -48 0.45 0.20 0.65 10 20

Ag 90 0 0.15 0.70 0.54 10 200 -42 0.15 0.70 0.54 5 10180 -48 0.15 0.70 0.54 10 50

Waste Au Omni Directional 0.10 0.08 0.07 12 38Ag Omni Directional 0.08 0.12 0.07 10 40

Note: A nested semivariogram has a nugget effect (C0), combined with, in this case, two nestedstructures with structural component C1 over the range a1 for the short range structure andstructure C2 over the range a2 for the long range structure. The sill of the semivariogram is madeup of the combination of C0 + C1 + C2.

Block ModelA block model with blocks 5 x 5 x 5 m in dimension was superimposed over the vein solids with

the percent below surface topography and percent within each vein solid recorded for each

block. GEMCOM software was used to determine these percentages using a procedure called

“needling”. The block model parameters are shown below.

Lower left corner652810 East Column size - 5.0 m Number of Columns - 1336068775 North Row size - 5.0 m Number of Rows - 52

Top of Model1500 Elevation Level size - 5.0 m Number of Levels - 65

No Rotation.



Bulk DensityA total of 70 specific gravity determinations from pieces of drill core were made during the 2009

drill campaign. The methodology was the weight in air / weight in water procedure where SG =

(Weight in air) / (Weight in air – Weight in water).

The results are presented in Appendix III. In summary, specific gravities from a total of 20

samples in quartz veins ranged between 2.63 and 3.49 with an average of 2.86. A total of 50

samples in wall rock material had specific gravities ranging from 2.67 to 3.02 with an average of

2.82. Previous to this a composite of 10 core samples used for metallurgical purposes in 1986

were sampled and showed a specific gravity of 2.78. Historical estimates have used 2.75. For

this resource the composite 2.78 value was used for both vein and wall rock. In future drill

programs more measurements should be taken to better differentiate between vein material and

waste rock.

Grade InterpolationGrades for gold and silver were interpolated into blocks using ordinary kriging. In all cases the

kriging exercise was completed in a series of four passes with the search ellipse dimension tied

to the semivariogram range and expanded from ¼ the range in pass 1, to ½ the range in pass 2,

to the full range in pass 3 and finally to twice the range in pass 4. In each pass a minimum of 4

composites were necessary to estimate the block. If more than 12 composites were found in

any search the closest 12 were used. A maximum of 3 composites were allowed from any one

drill hole and as a result all blocks are estimated by a minimum of two drill holes.

The Boulder Vein (BLDR) had the most information. The strike and dip varied along strike from

west to east and as a result the kriging for this vein was completed in three stages. For the

western part of the vein (West of 652946 E) a strike of azimuth 81o and dip of -42o S was used

for the orientation of the search ellipse. Within the central part between East Coordinates of

652946 and 653200 the strike changed to azimuth 94o and dip -52o S. Finally for the eastern

part of the vein with East Coordinate > 653200 E the strike was along azimuth 91o and the dip

steepened to -81o S. A soft boundary for data was used in these three vein segments but the

orientation of the search ellipse tried to follow the strike and dip of the vein. Only composites

from within the Boulder vein were used to estimate grade.



For the Boulder Foot Wall vein (BFW) the general model for the Boulder vein was used but the

search ellipse was oriented along strike of azimuth 91o and dip of -81o S. Only composites

within the Boulder Foot Wall vein were used for this estimate.

The Argillite vein was also split into two segments for estimation. The western part west of

653190 E was estimated using a search ellipse oriented along strike at azimuth 115o and dip of

-47o S. For the eastern segment with east coordinate > 653190 E the strike was adjusted to

azimuth 97o and dip of -28o S. Only composites from within the Argillite vein were used for

these estimates.

The West Hanging Wall vein and East Hanging wall vein were estimated using composites from

within the WHW and EHW solids respectively.

ClassificationBased on the study herein reported, delineated mineralization of the Dome Deposit is classified

as a resource according to the following definitions from National Instrument 43-101 and from

CIM (2005):

“In this Instrument, the terms "mineral resource", "inferred mineral resource", "indicatedmineral resource" and "measured mineral resource" have the meanings ascribed to thoseterms by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIMDefinition Standards on Mineral Resources and Mineral Reserves adopted by CIMCouncil, as those definitions may be amended.”

The terms Measured, Indicated and Inferred are defined by CIM (2005) as follows:

“A Mineral Resource is a concentration or occurrence of diamonds, natural solidinorganic material, or natural solid fossilized organic material including base andprecious metals, coal and industrial minerals in or on the Earth’s crust in such formand quantity and of such a grade or quality that it has reasonable prospects foreconomic extraction. The location, quantity, grade, geological characteristics andcontinuity of a Mineral Resource are known, estimated or interpreted from specificgeological evidence and knowledge.”“The term Mineral Resource covers mineralization and natural material of intrinsic

economic interest which has been identified and estimated through exploration andsampling and within which Mineral Reserves may subsequently be defined by theconsideration and application of technical, economic, legal, environmental, socio-economic and governmental factors. The phrase ‘reasonable prospects foreconomic extraction’ implies a judgement by the Qualified Person in respect of thetechnical and economic factors likely to influence the prospect of economicextraction. A Mineral Resource is an inventory of mineralization that underrealistically assumed and justifiable technical and economic conditions might become



economically extractable. These assumptions must be presented explicitly in bothpublic and technical reports.”

Inferred Mineral Resource“An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which

quantity and grade or quality can be estimated on the basis of geological evidenceand limited sampling and reasonably assumed, but not verified, geological and gradecontinuity. The estimate is based on limited information and sampling gatheredthrough appropriate techniques from locations such as outcrops, trenches, workingsand drill holes.”

“Due to the uncertainty that may be attached to Inferred Mineral Resources, itcannot be assumed that all or any part of an Inferred Mineral Resource will beupgraded to an Indicated or Measured Mineral Resource as a result of continuedexploration. Confidence in the estimate is insufficient to allow the meaningfulapplication of technical and economic parameters or to enable an evaluation ofeconomic viability worthy of public disclosure. Inferred Mineral Resources must beexcluded from estimates forming the basis of feasibility or other economic studies.”

Indicated Mineral Resource“An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which

quantity, grade or quality, densities, shape and physical characteristics, can beestimated with a level of confidence sufficient to allow the appropriate application oftechnical and economic parameters, to support mine planning and evaluation of theeconomic viability of the deposit. The estimate is based on detailed and reliableexploration and testing information gathered through appropriate techniques fromlocations such as outcrops, trenches, pits, workings and drill holes that are spacedclosely enough for geological and grade continuity to be reasonably assumed.”

“Mineralization may be classified as an Indicated Mineral Resource by theQualified Person when the nature, quality, quantity and distribution of data are suchas to allow confident interpretation of the geological framework and to reasonablyassume the continuity of mineralization. The Qualified Person must recognize theimportance of the Indicated Mineral Resource category to the advancement of thefeasibility of the project. An Indicated Mineral Resource estimate is of sufficientquality to support a Preliminary Feasibility Study which can serve as the basis formajor development decisions.”

Measured Mineral Resource“A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which

quantity, grade or quality, densities, shape, and physical characteristics are so wellestablished that they can be estimated with confidence sufficient to allow theappropriate application of technical and economic parameters, to support productionplanning and evaluation of the economic viability of the deposit. The estimate isbased on detailed and reliable exploration, sampling and testing information gatheredthrough appropriate techniques from locations such as outcrops, trenches, pits,workings and drill holes that are spaced closely enough to confirm both geologicaland grade continuity.”

“Mineralization or other natural material of economic interest may be classifiedas a Measured Mineral Resource by the Qualified Person when the nature, quality,quantity and distribution of data are such that the tonnage and grade of themineralization can be estimated to within close limits and that variation from theestimate would not significantly affect potential economic viability. This category



requires a high level of confidence in, and understanding of, the geology and controlsof the mineral deposit.”

Geologic continuity has been established through diamond drilling over a number of drill

campaigns and underground mapping and mining. Grade continuity can be quantified by

semivariogram analysis for each variable.

In the better informed Boulder vein and Argillite vein, blocks estimated in Pass 1 and 2 for gold

were classified as Indicated. These blocks were estimated with composites from within ½ the

semivariogram range. The remaining blocks were classified as Inferred. For the Boulder FW

vein, WHW vein and EHW vein all blocks were classified as Inferred as the drill density is still to

sparse to allow for indicated.

Estimated blocks were trimmed to the overburden surface and volumes attributed to

underground development were subtracted from the block volumes (see Figure 17).

The resource contained within the mineralized veins assuming one could mine to the vein

boundaries and no external dilution is presented in Tables 13 and 14. The results are shown for

a variety of gold cut-off grades as there has been no economic evaluation completed at this time

to establish an economic cut-off. A “base case” 5.0 g/t cutoff has been highlighted as a possible

cut-off that would be reasonable for an underground gold deposit in north-western British

Columbia. However, it is quite likely that the entire vein will be mined as it will be very difficult to

mine to a grade cut-off in this style of narrow vein deposit.

DilutionTo establish the influence of a mining dilution on the resource and to determine the grade of this

dilution, two scenarios were produced for the Main Boulder Vein (Code 100).

The first extended the Main Boulder Vein solid 0.5 m into the foot wall of the vein. New

composites were created to cover this vein extension and the expanded vein solid was kriged in

a similar manner as described above.

The second exercise extended the Main Boulder Vein solid 0.5 m into the hanging wall of the

vein. Again new composites were created to cover this vein extension and the expanded vein

solid was kriged in a similar manner as described above.



The results are shown in graphical form (Figure 18) to show the kind of dilution that might be

expected with one half metre added to either the foot wall or hanging wall of the vein. As would

be expected the tonnage increases at the expense of grade. These runs will provide grades for

dilution in the mine planning stage.

Figure 17: Plan View of Veins Showing Underground Development



Table 13Indicated Resource within the Mineralized Veins

Au Cutoff(g/t)



1.00 159,000 13.55 68.58 69,300 350,6001.50 156,000 13.71 69.24 68,800 347,3002.00 154,000 13.87 69.83 68,700 345,8002.50 152,000 14.04 70.46 68,600 344,3003.00 149,000 14.24 71.15 68,200 340,8003.50 147,000 14.40 71.68 68,100 338,8004.00 143,000 14.76 72.83 67,900 334,8004.50 140,000 14.95 73.36 67,300 330,2005.00 138,000 15.10 73.93 67,000 328,0005.50 136,000 15.22 74.47 66,600 325,6006.00 133,000 15.46 75.36 66,100 322,2006.50 129,000 15.73 76.32 65,200 316,5007.00 126,000 15.98 77.21 64,700 312,8007.50 121,000 16.31 78.06 63,500 303,7008.00 115,000 16.78 79.42 62,100 293,600

Table 14Inferred Resource with the Mineralized Veins

Au Cutoff(g/t)



1.00 206,000 10.90 51.96 72,200 344,2001.50 202,000 11.08 52.43 72,000 340,5002.00 197,000 11.33 53.12 71,800 336,4002.50 194,000 11.45 53.62 71,400 334,5003.00 190,000 11.66 54.35 71,200 332,0003.50 181,000 12.08 55.74 70,300 324,4004.00 173,000 12.43 57.03 69,100 317,2004.50 166,000 12.80 58.50 68,300 312,2005.00 154,000 13.42 60.63 66,500 300,2005.50 144,000 13.96 62.77 64,600 290,6006.00 137,000 14.43 64.49 63,500 284,0006.50 128,000 15.01 66.27 61,800 272,7007.00 119,000 15.59 68.05 59,700 260,3007.50 110,000 16.30 70.09 57,600 247,9008.00 104,000 16.81 71.55 56,200 239,200



Figure 18: Grade-Tonnage Curves for Total Resource in the Undiluted, FW Diluted andHW Diluted Boulder Main Vein

0.00

2.00

4.00

6.00

8.00

10.00

12.00

14.00

16.00

18.00

20.00

0

50,000

100,000

150,000

200,000

250,000

300,000

350,000

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

5.0

5.5

6.0

6.5

7.0

7.5

8.0

Au(g/t)

Tonn

es

Au Cutoff (g/t)

TOTAL RESOUCE FOR BOULDER VEINSHOWING EFFECTS OF DILUTION

Tonnes (No Dilution)

Tonnage (FW Dilution)

Tonnage (HW Diulution)

Au Grades (no Dilution)

Au Grades (FW Dilution)

Au Grades (HW Dilution)



18 OTHER RELEVANT DATA AND INFORMATION

There is no other relevant data or information pertaining to the property known to the authors.



19 INTERPRETATION AND CONCLUSIONS

Resource EstimateThe following conclusions are made regarding the resource estimate:

! The 2009 drilling program successfully confirmed the results of pre-2009 drilling to a

degree that all holes could be used in the resource estimate.

! The 2009 in-fill drilling on sections 20 metres apart seems adequate to define indicated

resources for mine planning purposes.

! The indicated resource for the Boulder and Argillite Veins is not very sensitive to the cut-

off grade (see Table 15).

! There is insufficient drilling to classify the Boulder Footwall, East Hangingwall, and West

Hangingwall resources as indicated.

! The QA/QC procedures used in the 2009 drilling program were adequate to verify the

analytical data from Assayers Canada.

! More metallic assays are required to confirm the reliability of the fire assay technique.

! Additional specific gravity determinations are required for all rock types.

Table 15Sensitivity of Indicated Resource to Gold Cut-off Grade

Gold Cut-off Tonnage g/t gold g/t silver

5.00 138,000 15.10 73.931.00 159,000 13.55 68.58

Extension of Known StructuresThe most immediate targets for increasing the mineral resources of the Dome Mountain Project

are:

! The Boulder Main Vein and the Boulder Footwall Vein down-plunge to the east.

Interesting drill hole intercepts between sections 653140m E and 653460 m E are: hole

90-08 with 36.1 g/t gold over 0.7 m and 20.9 g/t gold over 1.6 m, hole DM09-09 with 12.5

g/t gold over 2.0 m, hole 92-11 with 16.3 g/t gold over 2.2 m, hole 92-08 with 23.3 g/t

gold over 1.0 m, hole 90-15 with 19.6 g/t gold over 1.8 m and 14.0 g/t gold over 1.9 m,

hole 90-18 with 31.7 g/t gold over 0.5 m, hole 92-07 with 41.7 g/t gold over 0.7 m, and

hole DM09-43 with 8.6 g/t gold over 1.5 metres. All intersection widths are apparent

thicknesses.



Most of the inferred mineral resource comes from these veins between 653140m E and

653400m E.

! The Argillite Vein below the 1290 level.

Exploration

The intersection in hole DM09-046 can be interpreted either as the Chance Vein or a new vein

discovery.

Based on the results of exploration to date and on the deposit model, there is significant

potential to develop mineral resources on the Forks, Elk, Free Gold and 9800 veins and to

discover new veins.



20 RECOMMENDATIONS

In-Depth Geological Services and Giroux Consulting Ltd. make the following recommendations

for the continued exploration and development of the Dome Mountain Project:

! A Preliminary Assessment (PA) should be commissioned to determine the potential

economic viability of re-opening the Dome Mountain underground mine on a direct

shipping (toll milling) basis.

! Additional in-fill diamond drilling is proposed to upgrade the inferred mineral resource.

! Exploration should be conducted in order to increase the Mineral Resource.

! The exploration work should be phased to allow for evaluation of results and planning of

subsequent work.

Phase 1Phase 1 of the proposed program will consist of the PA and continued in-fill drilling to upgrade

the inferred mineral resource of the Boulder Vein System. The PA should look at the economics

of mining the Boulder and Argillite veins by underground mining methods. Separately the PA

should look at the potential to mine the Argillite Vein by open-pit mining. The in-fill drilling

program should consist of 5300 metres on sections spaced 20 metres apart. As part of the in-fill

drilling, specific gravity and metallic assay samples should be taken at the rate of 1 in 20 assay

samples. Phase 1 will also include 3000 metres of exploration drilling down-plunge to the east

on the Boulder Vein System and around hole DM09-046 as well as soil geochemistry and 3D

induced polarization surveys. A budget for the proposed program is summarized in Table 16.

Table 16Phase 1 Budget

Item C$Preliminary assessment 80,000In-fill Diamond Drilling (all found) 583,000Line-cutting 50,0003D Induced polarization geophysics 70,000Soil geochemistry 50,000Mapping and trenching 15,000Drill pads, trails, trenching and reclamation 75,000Field office and core processing building 20,000Exploration diamond drilling (all found) 330,000

Sub-Total 1,273,000Contingency (15%) 191,000Total 1,464,000



Phase 2

Conditional upon the results of the Phase I exploration, 8000 metres of diamond drilling is

proposed for the Phase 2 work program. The budget for the proposed Phase 2 program is

presented in Table 17.

Table 17Phase 2 Budget

Item C$Exploration Diamond Drilling (all found) 880,000Planning, permitting, reporting 25,000Drill pads, trails, trenching and reclamation 75,000Mapping and trenching 25,000

Sub-total 1,005,000Contingency (15%) 151,000Total 1,156,000



21 REFERENCES

CIM (2003): Estimation of Mineral Resources and Mineral Reserves, Best Practice Guidelines;55 p.

L'Orsa, A. (March 12, 1990): Report on the Diamond November 1989 toJanuary 1990, DomeMountain, Omineca Mining Division, British Columbia; internal report to Teeshin Resources Ltd.

MacIntyre, D.G., Brown, D., Desjardins, P., and Mallet, P. (1987): Babine Project, BC Ministry ofEnergy, Mines and Petroleum Resources, Geological Fieldwork, 1986, Paper 1987-1, 22 p.

MacIntyre, D.G. and Desjardins, P. (1988): Geology of the Silver King - Mount Cronin Area, BCMinistry of Energy, Mines and Petroleum Resources, Open File 1988-20

Minestart Management Inc. (June 1986): Dome Mountain Gold - Smithers, British Columbia,internal report to Teeshin Resources Ltd. and Canadian-United Minerals Inc.

Roscoe Postle Associates Inc. (December 20, 1993): Report on the Dome Mountain Project forHapsburg Resources Inc.

Roscoe Postle Associates Inc. (March 31, 1998): Addendum to a Report on the Dome MountainProject - Prepared for DMR Resources Ltd.

Scott Wilson Roscoe Postle Associates Inc. (April 28, 2008): NI 43-101 Technical Report onDome Mountain Project, Smithers, British Columbia, Canada, Prepared for Metal MountainResources Inc.

Smit, H. (April 6, 1993): Dome Mountain Drill Proposal, internal memo to Hapsburg ResourcesInc., 4 p

Tipper , H. W., and Richards, T.A. (1976): Geology of the Smithers Area, Geological Survey ofCanada, Open File 351

Appendix I

List of Drill-holes Used in Resource Analysis

The holes used in the Resource estimate are highlighted in green

HOLE EASTING NORTHING ELEVATION HOLE LENGTH (m)1290-1 653328.00 6068883.50 1295.00 48.801290-2 653328.00 6068883.50 1295.00 61.001290-3 653328.00 6068883.50 1295.00 91.401290-4 653328.00 6068883.50 1295.00 56.401290-5 653328.00 6068883.50 1295.00 61.00BL85-21 653308.30 6068212.70 1326.50 20.73BL85-22 653298.70 6068216.60 1327.40 18.29BL85-23 653289.10 6068220.40 1332.20 22.10BL85-24 653329.90 6068233.40 1325.90 35.97BL85-25 653280.60 6068223.50 1334.10 21.45BL85-26 653252.50 6068251.40 1342.90 25.30C85-31 652439.70 6068939.30 1472.60 31.70C85-33 652504.80 6068963.30 1468.00 23.77D89-01 653385.40 6068762.90 1304.90 197.20D89-02 653384.90 6068761.30 1304.80 53.90D89-03 653469.10 6068895.00 1316.00 121.01D89-04 653538.80 6068782.60 1277.90 196.60D89-05 653416.30 6068738.00 1290.70 133.20D89-06 653359.00 6068839.20 1334.50 151.50D89-07 652871.50 6068956.30 1428.10 68.60D89-08 652871.50 6068956.30 1428.10 81.40D89-09 652845.20 6068961.00 1427.70 63.10D89-10 652845.00 6068960.30 1427.80 84.40D89-11 652859.00 6068901.60 1429.70 111.90D89-12 652858.90 6068901.10 1430.20 121.00D89-13 652835.80 6068911.10 1431.20 102.70D89-14 652835.70 6068910.80 1431.20 157.60D90-01 653349.80 6068812.90 1327.30 99.70D90-02 653349.80 6068812.90 1327.30 111.90D90-03 653340.50 6068848.00 1336.00 156.70D90-04 653340.50 6068846.80 1336.00 78.30D90-05 653375.60 6068837.80 1332.10 57.00D90-06 653289.80 6068854.80 1328.20 124.10D90-07 653136.10 6068842.80 1367.70 136.20D90-08 653136.20 6068842.30 1367.60 185.00D90-09 653157.10 6068893.40 1366.80 108.80D90-10 653156.90 6068892.90 1366.80 106.70D90-11 653137.80 6068868.30 1366.50 102.70D90-12 653148.60 6068871.40 1366.40 102.70D90-13 653175.50 6068845.30 1358.20 136.20D90-14 653175.00 6068844.70 1359.80 173.60D90-15 653278.80 6068824.60 1329.00 115.80D90-16 653276.70 6068786.60 1329.00 201.20D90-17 653381.80 6068833.80 1332.00 170.70D90-18 653455.90 6068838.50 1317.00 158.50

D92-01 653375.60 6068901.60 1337.50 91.40D92-02 653376.50 6068921.60 1337.20 82.30D92-03 653307.20 6068892.80 1337.80 91.40D92-04 653326.00 6068921.30 1340.40 67.10D92-05 653268.90 6068910.70 1337.30 45.70D92-06 653323.80 6068903.50 1338.50 45.70D92-07 653466.30 6068888.10 1316.60 106.70D92-08 653243.50 6068850.60 1339.80 109.70D92-09 653221.30 6068907.90 1347.90 61.00D92-10 653275.40 6068861.70 1329.90 94.50D92-11 653214.90 6068907.40 1348.50 103.60D92-12 653214.60 6068908.40 1348.30 85.30D92-13 653468.60 6068906.90 1316.30 61.00DM09-01 653099.45 6068923.41 1386.60 71.90DM09-02 653099.45 6068923.41 1386.60 78.30DM09-03 653099.45 6068923.41 1386.60 12.80DM09-04 653099.45 6068923.41 1386.60 85.00DM09-05 653100.37 6068836.95 1385.40 150.60DM09-06 653115.62 6068912.05 1377.90 72.20DM09-07 653115.62 6068912.05 1377.90 63.10DM09-08 653139.18 6068906.56 1370.30 75.30DM09-09 653139.18 6068906.56 1370.30 139.30DM09-10 653115.62 6068912.05 1377.90 114.90DM09-11 653079.70 6068845.78 1395.40 182.00DM09-12 653079.70 6068845.78 1395.40 185.00DM09-13 653079.54 6068924.61 1392.70 67.70DM09-14 653061.61 6068933.11 1399.20 66.10DM09-15 653061.61 6068933.11 1399.20 71.60DM09-16 653061.61 6068933.11 1399.20 81.40DM09-17 653044.28 6068949.96 1405.30 66.10DM09-18 653044.28 6068949.96 1405.30 121.00DM09-19 653044.50 6068948.95 1405.40 84.40DM09-20 653044.50 6068948.95 1405.40 118.00DM09-21 653020.59 6068951.55 1411.90 81.40DM09-22 653020.59 6068951.55 1411.90 84.40DM09-23 653020.59 6068951.55 1411.90 96.60DM09-24 653067.09 6068855.68 1402.30 137.50DM09-25 653067.09 6068855.68 1402.30 194.20DM09-26 653045.72 6068863.66 1409.30 185.00DM09-27 653045.72 6068863.66 1409.30 192.10DM09-28 653020.99 6068871.38 1415.80 169.80DM09-29 653020.99 6068871.38 1415.80 169.80DM09-30 652997.31 6068883.55 1420.70 160.60DM09-31 652997.31 6068883.55 1420.70 160.20DM09-32 652978.77 6068884.72 1425.70 145.40DM09-33 652978.77 6068884.72 1425.70 160.60DM09-34 652978.77 6068884.72 1425.70 160.60DM09-35 652953.77 6068884.94 1429.50 136.20

DM09-36 652953.77 6068884.94 1429.50 142.30DM09-37 652997.59 6068963.36 1416.80 84.40DM09-38 652997.59 6068963.36 1416.80 99.70DM09-39 652974.98 6068975.88 1420.00 90.50DM09-40 652952.02 6068971.30 1423.90 81.40DM09-41 652952.02 6068971.30 1423.90 84.40DM09-42 652952.02 6068971.30 1423.90 90.50DM09-43 653438.00 6068751.00 1288.00 422.80DM09-44 653650.00 6069133.00 1285.00 29.60DM09-45 653650.00 6069134.00 1285.00 26.80DM09-46 653756.00 6069146.00 1266.00 409.00DM86-01 652632.40 6068983.70 1453.70 67.00DM86-02 652632.30 6068981.80 1453.70 82.30DM86-03 652632.60 6068949.00 1450.90 94.20DM86-04 652525.70 6068935.60 1463.70 121.50DM86-05 652525.70 6068935.60 1463.70 111.90DM86-06 652470.40 6068940.50 1470.60 103.60DM86-07 652471.90 6068910.50 1465.30 121.60DM86-08 652525.40 6068911.50 1460.70 106.70DM86-09 652477.60 6068881.80 1455.70 115.80DM86-10 652524.10 6068880.90 1457.20 100.60F85-01 653175.50 6068389.10 1326.10 34.75F85-02 653188.00 6068373.80 1323.70 44.50F85-03 653204.80 6068367.00 1329.50 25.60F85-04 653202.30 6068311.60 1346.70 22.25F85-05 653247.80 6068383.80 1336.50 75.29F85-06 653186.20 6068301.20 1347.30 90.53F85-07 653186.10 6068300.30 1347.30 4.27F85-08 653205.60 6068410.20 1318.40 47.85F85-09 653206.30 6068410.80 1318.40 44.81F85-10 653187.60 6068300.30 1347.30 65.84F85-11 653216.90 6068425.80 1318.20 57.30F85-12 653197.10 6068437.60 1327.90 60.35F85-13 653198.70 6068439.30 1327.90 60.35F85-14 653159.10 6068447.90 1348.00 72.54F85-15 653233.90 6068411.60 1322.10 66.44F85-16 653234.20 6068412.30 1322.10 57.30F85-17 653202.40 6068489.20 1331.00 83.21F85-18 653234.60 6068465.30 1314.90 63.40F85-19 653150.40 6068401.50 1338.20 43.58F85-20 653267.80 6068439.50 1326.00 71.93LM86-1 653159.10 6068851.10 1363.80 90.50LM86-2 653159.10 6068851.10 1363.80 84.43LM86-3 653158.10 6068849.60 1363.20 114.90LM86-4 653254.20 6068822.10 1336.90 60.00LM86-5 653254.20 6068822.10 1336.90 44.80LM86-6 653254.20 6068822.10 1336.90 60.00LM86-7 653301.90 6068839.50 1327.00 94.50

RP87-01 651288.10 6069443.20 1665.10 79.24RP87-02 651259.00 6069623.90 1631.50 82.29RP87-03 651216.70 6069626.00 1633.40 97.53RP87-04 652179.70 6068112.30 1477.40 91.44RP87-05 652165.40 6068164.50 1483.20 60.96RP87-06 652122.20 6068247.70 1496.50 76.20RP87-07 652080.20 6068204.90 1498.30 27.43RP87-08 652388.50 6068877.80 1469.80 36.57RP87-09 652390.80 6068849.30 1468.90 109.72RP87-10 652358.80 6068848.60 1472.90 94.48RP87-11 652337.50 6068972.90 1479.40 91.44RP87-12 652294.40 6068962.10 1483.40 91.44RP87-13 652259.20 6068994.80 1485.80 124.00RP87-14 652232.70 6068957.10 1488.90 64.00RP88-15 652294.20 6068986.50 1483.70 150.00RP88-16 652231.00 6069000.30 1488.40 153.70RP88-17 652230.40 6069062.00 1488.30 128.62RP88-18 652251.00 6068868.60 1487.60 127.13RP88-19 652250.10 6068831.20 1486.20 150.30RP88-20 652159.70 6068874.80 1502.80 125.27RP88-21 652778.50 6068634.60 1402.20 91.46RP88-22 652783.40 6068584.50 1403.40 152.44RP88-23 652731.50 6068638.60 1408.50 99.97RP88-24 652731.70 6068576.90 1408.10 160.01RP89-26 651880.10 6068971.70 1547.50 190.81RP89-27 651869.80 6069101.40 1531.80 151.18RP89-28 651843.70 6068804.10 1570.30 166.42RP89-29 652165.60 6068830.10 1503.70 96.32RP89-30 651952.10 6068655.50 1558.80 90.22RP89-31 652224.40 6068485.80 1481.80 117.65T85-01 653207.30 6068370.90 1329.00 76.30T85-02 653234.40 6068410.40 1322.10 73.20T85-03 653204.90 6068465.00 1330.00 85.04T86-04 652998.20 6068992.90 1414.10 48.17T86-05 653013.50 6068991.30 1410.90 39.01T86-06 653013.50 6068991.30 1410.90 66.75T86-07 653036.20 6068978.90 1405.60 45.11T86-08 653036.20 6068978.90 1405.60 94.20T86-09 653031.80 6068957.60 1407.70 96.90T86-10 653030.00 6068926.60 1410.90 154.83T86-11 653055.30 6068980.20 1398.00 38.11T86-12 653055.30 6068980.20 1398.00 45.73T86-13 653077.50 6068963.10 1391.40 35.96T86-14 653077.50 6068963.10 1391.40 54.25T86-15 652998.20 6068992.90 1414.10 66.75T86-16 652972.00 6068998.80 1416.10 32.92T86-17 652972.00 6068998.80 1416.10 45.41T86-18 652949.20 6069001.50 1418.10 38.41

T86-19 652949.20 6069001.50 1418.10 42.67T86-20 652929.90 6069010.90 1418.70 26.83T86-21 652929.90 6069010.90 1418.70 42.67T86-22 652906.40 6069017.40 1419.60 26.83T86-23 652906.40 6069017.40 1419.60 33.53T86-24 653098.40 6068922.80 1386.80 69.80T86-25 653098.40 6068922.80 1386.80 78.96T86-26 653098.40 6068922.80 1386.80 111.50T86-27 653063.80 6068870.40 1400.80 176.46T86-28 653063.80 6068870.40 1400.80 243.90T86-29 653137.90 6068912.10 1372.10 48.17T86-30 653137.90 6068912.10 1372.10 72.56T86-31 652875.90 6068996.30 1422.80 61.57T86-32 652875.90 6068996.30 1422.80 54.56T86-33 652850.60 6068996.40 1423.30 51.20T86-34 652850.60 6068996.40 1423.30 67.97T86-35 652829.90 6068997.20 1424.10 75.59T86-36 652829.90 6068997.20 1424.10 62.18T86-37 652806.60 6068999.70 1425.00 38.10T86-38 652806.60 6068999.70 1425.00 42.40T86-39 652901.40 6068965.00 1425.70 74.98T86-40 652939.30 6068957.30 1427.30 82.00T86-41 652979.70 6068903.10 1426.10 139.90T86-42 652925.60 6068896.30 1434.70 158.54T86-43 652886.90 6068908.50 1429.50 142.94T86-44 652971.60 6068871.40 1427.00 184.76T86-45 653011.70 6068858.60 1419.10 197.86T86-46 652918.90 6068861.30 1433.10 191.10T86-47 652881.80 6068868.10 1429.20 243.60T86-48 653063.50 6068866.60 1401.80 152.09T86-49 653100.00 6068840.00 1384.80 197.81T86-50 653100.00 6068840.00 1384.80 216.46T86-51 653138.60 6068909.30 1371.30 142.68TB87-01 653010.00 6068961.60 1413.10 124.90TB87-02 652971.00 6068975.30 1420.30 72.50TB87-03 652980.20 6068927.40 1426.90 106.50TB87-04 652980.20 6068927.40 1426.90 134.00TS86-01 653221.00 6068891.00 1346.10 134.10TS86-02 653386.20 6068760.40 1305.60 51.80TS86-03 653270.60 6068781.80 1329.70 99.10TS86-04 653269.30 6068780.70 1329.80 85.30TS86-05 653236.40 6068817.50 1337.40 73.10TS86-06 653268.50 6068738.70 1326.80 73.10TS86-07 653268.50 6068738.70 1326.80 100.30TS86-08 653140.30 6068719.90 1358.40 112.80TS86-09 653177.40 6068656.80 1355.00 109.70TS86-10 653118.10 6068779.20 1371.90 155.40TS86-11 653149.30 6068872.90 1368.00 76.20

TS86-12 653149.30 6068872.90 1368.00 61.00TS86-13 653149.30 6068872.90 1368.00 83.20TS86-14 653172.30 6068823.70 1355.10 79.30TS86-15 653172.30 6068823.70 1355.10 115.80TS86-16 653216.00 6068782.30 1341.00 82.30TS86-17 653214.90 6068781.10 1341.00 79.30TS86-18 653296.60 6068747.70 1324.50 103.60TS86-19 653296.60 6068747.70 1324.50 72.20TS86-20 653248.30 6068758.10 1330.50 77.40TS86-21 653248.30 6068758.10 1330.50 97.50TS86-22 653332.80 6068672.00 1314.60 118.90TS86-23 653332.10 6068671.00 1314.80 61.00TS86-24 653274.70 6068689.40 1326.90 131.10TS86-25 653288.80 6068785.20 1324.90 89.30TS86-26 653288.80 6068785.20 1324.90 82.30TS86-27 653199.30 6068821.10 1349.80 76.20TS86-28 653199.30 6068821.10 1348.20 91.44TS86-29 653250.50 6068783.20 1334.80 61.00TS86-30 653243.90 6068837.70 1339.10 48.70TS86-31 653243.90 6068837.70 1339.10 51.80TS86-32 653316.90 6068851.50 1335.80 61.00TS86-33 653316.90 6068851.50 1335.80 15.20TS86-34 653387.10 6068826.90 1329.30 42.70TS87-01 653292.80 6068791.00 1325.20 67.05TS87-02 653292.80 6068791.00 1325.20 45.72TS87-03 653292.80 6068791.00 1325.20 57.90TS87-04 653281.90 6068853.30 1327.80 31.80TS87-05 653391.90 6068763.70 1304.70 85.34TS87-06 653256.00 6068799.60 1335.20 57.90TS87-07 653256.80 6068800.80 1336.80 64.00TS87-08 653254.70 6068798.20 1335.40 73.15TS87-09 653243.50 6068816.50 1337.50 67.05TS87-10 653244.30 6068817.20 1337.60 57.91TS87-11 653242.50 6068815.80 1337.70 88.39TS87-12 653229.10 6068829.60 1341.10 79.24TS87-13 653229.10 6068829.60 1341.10 56.40TS87-14 653229.10 6068829.60 1341.10 58.21TS87-15 653512.60 6068567.00 1262.40 54.86TS87-16 653537.60 6068633.90 1268.30 41.15TS87-17 653397.60 6068673.30 1299.60 66.34TS87-18 653399.50 6068678.90 1299.50 42.67TS87-19 653181.90 6068860.90 1359.00 54.86TS87-20 653181.90 6068860.90 1359.00 100.27TS87-21 653204.10 6068861.20 1348.90 60.96TS87-22 653203.60 6068860.60 1348.90 57.91TS87-23 653239.80 6068842.50 1339.90 36.57TS87-24 653246.50 6068299.20 1343.30 39.01TS87-25 653219.00 6068317.90 1343.20 35.05

TS87-26 653262.50 6068403.60 1336.70 47.24TS87-27 653293.90 6068440.20 1336.00 45.72TS87-28 653351.00 6068452.70 1321.80 70.10TS87-29 653344.70 6068262.80 1324.80 88.39TS87-30 653252.30 6068392.60 1336.10 47.10UG-01 652925.60 6068992.00 1370.00 8.23UG-02 652926.40 6068996.00 1370.00 4.50UG-03 653079.20 6068946.30 1368.00 37.80UG-04 653080.00 6068946.30 1368.00 33.83UG-05 653080.00 6068946.30 1368.00 39.93UG-06 653080.00 6068946.30 1368.00 29.26UG-07 653079.20 6068946.30 1368.00 43.90UG-08 653093.30 6068913.80 1368.00 59.80UG-09 653093.30 6068913.80 1368.00 70.87UG-10 652977.60 6068992.20 1368.00 13.41UG-11 652946.20 6068989.50 1368.00 9.45UG-12 652948.00 6069006.50 1390.00 9.14UG-13 652948.00 6069006.50 1390.00 7.62UG-14 652996.40 6068990.80 1368.00 6.09UG-15 653009.60 6068997.30 1368.00 14.32UG-16 653019.40 6068995.30 1368.00 27.43UG-17 653039.60 6068992.60 1368.00 20.13UG-18 653025.80 6069004.70 1390.00 10.26UG-19 653025.80 6069004.70 1390.00 20.43UG-20 653047.90 6068986.60 1368.00 20.12UG-21 653079.00 6068940.50 1368.00 52.14UG92-06 653123.40 6068930.40 1298.80 13.41UG92-07 653123.40 6068930.40 1299.10 13.72UG92-08 653200.90 6068948.10 1298.10 49.68UG92-09 653107.30 6068930.50 1298.80 8.84UG92-10 653107.60 6068933.20 1299.10 15.24UG92-11 653089.20 6068923.30 1297.50 9.14UG92-12 653093.20 6068923.50 1297.50 13.11UG92-13 653100.20 6068922.50 1297.50 15.54UG92-14 653143.00 6068934.60 1299.10 22.56UG93-15 653156.70 6068931.50 1299.10 28.65UG93-16 653125.80 6068918.80 1297.50 33.22

Appendix II

Semivariogram Plots for Gold and Silver

Appendix III

Specific Gravity Measurements

DDH ID SampleFrommeters

Tometers

RockCode Description

Weightin Air(g)

Weightin water(g)

Wt(air)minusWt(water) S.G.

QV Standard 1 0.6582 0.4326 0.2256 2.92

QV Standard 1 0.6582 0.4327 0.2255 2.92

QV Standard 1 0.6583 0.4327 0.2256 2.92

DM09-01 A060004 51.30 51.45 Va serc-gm 0.5396 0.3513 0.1883 2.87

DM09-01 55.70 55.85 Va (+Vas) serc-chl 1.2334 0.8053 0.4281 2.88

DM09-01 60.22 60.32 LT 0.9984 0.6387 0.3597 2.78

DM09-01 DM-1 69.95 70.10 LT 0.6200 0.3950 0.2250 2.76

DM09-06 49.57 49.66 Va serc-gm-hem 0.6832 0.4464 0.2368 2.89

DM09-06 A060037 50.50 50.63 Vas serc-gm 1.1322 0.7380 0.3942 2.87

DM09-06 A060040 53.35 53.57 Vas serc-gm; hi-su 0.9809 0.6506 0.3303 2.97

DM09-06 A060044 56.40 56.55 Vas serc-ep-hem; hi-su 0.5495 0.3669 0.1826 3.01

DM09-06 A060044 56.68 56.80 QV hi-su 0.6767 0.4714 0.2053 3.30

DM09-06 58.51 58.61 Va chl-serc 0.8797 0.5667 0.3130 2.81

DM09-06 59.88 60.00 CV partial QV 1.0975 0.6915 0.4060 2.70

DM09-06 65.40 65.50 LT maroon bx 1.0272 0.6569 0.3703 2.77

DM09-19 68.89 69.00 Va serc-gm-hem; dism py 0.9691 0.6371 0.3320 2.92

DM09-19 A060238 70.00 70.10 QV (+Va) mod su 0.3513 0.2281 0.1232 2.85

DM09-19 A060240 71.00 71.10 QV lo su 0.4773 0.2994 0.1779 2.68

DM09-19 A060243 74.22 74.30 Va serc-clay 0.3663 0.2365 0.1298 2.82

DM09-23 A060345 70.06 70.17 Vas serc-gm; hi-su 0.4982 0.3331 0.1651 3.02

DM09-23 A060345 70.17 70.26 QV mod su 0.3853 0.2489 0.1364 2.82

DM09-23 A060346 71.05 71.20 Va serc-gm-ep 0.6492 0.4215 0.2277 2.85

DM09-23 A060349 73.10 73.24 Vas serc-gm-hem-ep; dism py 0.5708 0.3728 0.1980 2.88

DM09-23 A060351 78.45 78.56 QV lo su 0.4206 0.2615 0.1591 2.64

DM09-23 DM-5 81.60 81.70 AT 0.5060 0.3248 0.1812 2.79

DM09-26 A060438 120.80 121.00 QV bx, mod py 0.7580 0.4822 0.2758 2.75

DM09-26 A060441 127.36 127.46 Va hem(chl) 0.3675 0.2329 0.1346 2.73

DM09-26 DM-7 134.20 134.30 AT maroon 0.4745 0.3039 0.1706 2.78

DM09-26 A060450 142.87 143.00 QV 4% py 0.4109 0.2654 0.1455 2.82

DM09-26 A060456 148.60 146.75 LT wk alt: chl-hem-clay 0.6065 0.3928 0.2137 2.84

DM09-26 A060460 151.83 151.96 QV bdd; 5% su 0.5481 0.3556 0.1925 2.85

DM09-26 A060462 153.50 153.60 Va serc; cubic py 1% 0.5351 0.3462 0.1889 2.83

DM09-31 A060640 130.36 130.50 QV bdd; 10% su 0.5955 0.3891 0.2064 2.89

DM09-31 A060643 136.20 136.35 Vas serc; py 1% 0.6542 0.4274 0.2268 2.88

DM09-31 145.05 145.14 AT maroon 0.8542 0.5456 0.3086 2.77

DM09-32 A060582 112.95 113.03 QV bdd; 25% su 0.3822 0.2727 0.1095 3.49

DM09-32 A060587 116.10 116.20 QV msv/bx lo su 0.7598 0.4712 0.2886 2.63

DM09-32 A060590 118.65 118.80 QV bx; 2% su 0.5480 0.3520 0.1960 2.80

DDH ID SampleFrommeters

Tometers

RockCode Description

Weightin Air(g)

Weightin water(g)

Wt(air)minusWt(water) S.G.

DM09-32 128.30 128.40 AT maroon, cb stwk 0.8226 0.5278 0.2948 2.79

DM09-34 A060676 123.05 123.18 QV bdd; 15% su 0.6044 0.4186 0.1858 3.25

DM09-34 A060677 123.39 123.47 Va intense serc-sil 0.3179 0.1988 0.1191 2.67

DM09-34 144.13 144.22 FP cb stwk 0.8643 0.5511 0.3132 2.76

DM09-34 149.00 149.10 FP wk alt 0.8574 0.5398 0.3176 2.70

DM09-35 A060691 120.34 120.50 QV bdd; 3% su 0.6582 0.4174 0.2408 2.73

DM09-35 A060696 122.10 122.20 QV msv; lo su 0.4219 0.2629 0.1590 2.65

DM09-35 A060697 123.20 123.30 Va chl (serc) 0.4339 0.2770 0.1569 2.77

DM09-35 DM-13 129.40 129.50 LT wk chl-hem 0.9064 0.5794 0.3270 2.77

DM09-36 A060708 113.40 113.50 Va serc-clay; 2% py 0.4054 0.2673 0.1381 2.94

DM09-36 A060716 116.60 116.70 Va serc(hem); 1% py 0.4386 0.2864 0.1522 2.88

DM09-36 A060723 119.55 119.70 QV msv; lo su 0.3800 0.2373 0.1427 2.66

DM09-36 A060726 120.65 120.80 Va bdd; 2% py 0.6379 0.4150 0.2229 2.86

DM09-38 A060772 78.00 78.12 QV stwk su 10% 0.5837 0.3892 0.1945 3.00

DM09-38 A060777 79.80 79.95 QV patchy su 2% 0.6705 0.4259 0.2446 2.74

DM09-38 A060782 82.80 82.90 Vas chl (serc) 0.3813 0.2468 0.1345 2.83

DM09-38 88.75 88.85 FP 0.7180 0.4592 0.2588 2.77

TS87-13 20.50 20.65 V/S wk serc alt'd 0.3627 0.2294 0.1333 2.72

TS87-14 27.00 27.10 V/S wk chl alt'd 0.3081 0.1932 0.1149 2.68

TS87-12 44.00 44.15 V/S py strs 1% 0.3991 0.2561 0.1430 2.79

D90-14 144.80 144.95 AT chl (hem) + qz strs 0.4210 0.2706 0.1504 2.80

T86-28 102.00 102.15 LT cb strs 0.4475 0.2840 0.1635 2.74

T86-28 82.00 82.20 BTm strong cb alt'd 0.4295 0.2761 0.1534 2.80

D90-13 63.00 63.10 AB strong cb, wk hem alt'd 0.2742 0.1760 0.0982 2.79

TS87-13 55.00 55.15 AB mod cb alt'd 0.3453 0.2209 0.1244 2.78

DM09-02 45.50 45.60 AB 0.8969 0.5774 0.3195 2.81

DM09-44 14.20 14.30 AR gritty 0.7039 0.4614 0.2425 2.90

DM09-44 17.40 17.50 AR calcareous, debris flow 0.6594 0.4236 0.2358 2.80

DM09-44 25.50 25.60 QV bx, graphitic 0.8655 0.5588 0.3067 2.82

DM09-45 22.20 22.30 QV bx, graphitic 0.4154 0.2651 0.1503 2.76

DM09-43 87.10 87.20 AB maroon 0.8586 0.5604 0.2982 2.88

DM09-43 95.50 95.55 BTm calcareous 0.2997 0.1926 0.1071 2.80


DM09-43 102.56 102.63 AB maroon 0.7124 0.4575 0.2549 2.79


Appendix IV

2009 Drill-hole Intersection Composites

BHID FROM TO LENGTH AU AG CU ZN PBm m m g/t g/t % % %

DM‐09‐001 36.80 37.10 0.30 0.23 0.9 ‐ ‐ ‐DM‐09‐001 52.10 53.20 1.10 54.61 398.1 1.60 8.08 1.40DM‐09‐002 58.00 60.00 2.00 2.22 8.2 ‐ 1.12 ‐DM‐09‐002 60.00 61.70 1.70 51.68 146.8 ‐ ‐ ‐DM‐09‐004 69.50 72.00 2.50 34.37 80.4 ‐ 2.58 1.58DM‐09‐005 83.50 84.20 0.70 1.33 0.2 ‐ ‐ ‐DM‐09‐005 112.00 112.50 0.50 0.83 0.9 ‐ ‐ ‐DM‐09‐005 114.00 114.40 0.40 0.23 0.5 ‐ ‐ ‐DM‐09‐005 124.00 126.50 2.50 12.85 73.4 ‐ ‐ ‐DM‐09‐006 47.40 48.60 1.20 1.70 6.8 ‐ 1.49 ‐DM‐09‐006 51.30 57.10 5.80 15.67 74.7 1.31 4.27 1.47DM‐09‐007 50.40 52.80 2.40 0.05 1.0 ‐ ‐ ‐DM‐09‐007 61.70 63.10 1.40 11.80 82.6 ‐ 16.90 ‐DM‐09‐008 34.30 36.30 2.00 0.87 3.7 ‐ ‐ ‐DM‐09‐008 57.70 58.40 0.70 10.50 36.8 ‐ ‐ ‐DM‐09‐009 50.30 51.30 1.00 27.97 95.5 ‐ 1.48 ‐DM‐09‐009 84.70 86.40 1.70 7.94 26.3 ‐ 3.02 ‐DM‐09‐009 93.00 95.00 2.00 2.39 24.2 ‐ ‐ ‐DM‐09‐009 126.90 128.00 1.10 2.27 9.5 ‐ ‐ ‐DM‐09‐009 133.00 134.00 1.00 18.27 141.9 1.65 ‐ ‐DM‐09‐010 69.60 70.00 0.40 1.37 13.5 ‐ ‐ ‐DM‐09‐010 90.50 96.00 5.50 5.16 48.5 ‐ ‐ ‐DM‐09‐010 108.40 109.50 1.10 6.23 34.6 ‐ ‐ ‐DM‐09‐011 125.80 127.30 1.50 0.02 0.9 ‐ ‐ ‐DM‐09‐011 152.20 153.90 1.70 8.37 61.9 1.36 ‐ ‐DM‐09‐011 162.00 163.00 1.00 4.71 9.7 ‐ ‐ ‐DM‐09‐012 155.20 156.20 1.00 0.27 2.6 ‐ ‐ ‐DM‐09‐013 66.00 66.40 0.40 17.67 124.0 ‐ ‐ ‐DM‐09‐014 44.50 44.80 0.30 1.77 5.4 ‐ ‐ ‐DM‐09‐014 60.00 61.00 1.00 4.26 51.5 ‐ 3.84 ‐DM‐09‐014 62.00 66.10 4.10 22.59 106.1 ‐ 4.49 1.20DM‐09‐015 49.00 50.30 1.30 12.69 50.9 ‐ 4.48 ‐DM‐09‐017 45.50 45.80 0.30 2.80 11.3 ‐ ‐ ‐DM‐09‐017 57.00 58.00 1.00 0.13 1.8 ‐ ‐ ‐DM‐09‐018 49.40 49.90 0.50 6.20 22.4 ‐ ‐ ‐DM‐09‐018 52.70 53.00 0.30 32.57 79.8 ‐ 11.40 ‐DM‐09‐018 63.60 64.60 1.00 4.14 32.0 ‐ 4.13 ‐DM‐09‐019 49.50 50.10 0.60 6.57 9.2 ‐ ‐ ‐DM‐09‐019 56.00 56.60 0.60 4.63 8.3 ‐ ‐ ‐DM‐09‐019 63.00 63.50 0.50 0.87 42.9 ‐ 1.47 ‐DM‐09‐019 69.70 72.00 2.30 8.54 56.1 ‐ 2.83 ‐DM‐09‐020 67.70 68.00 0.30 3.79 64.7 ‐ 3.37 ‐DM‐09‐020 74.80 76.50 1.70 0.02 0.4 ‐ ‐ ‐DM‐09‐020 83.30 84.00 0.70 0.77 1.8 ‐ ‐ ‐DM‐09‐020 95.30 96.20 0.90 0.33 3.5 ‐ ‐ ‐


DM‐09‐021 58.00 59.00 1.00 0.07 1.0 ‐ ‐ ‐DM‐09‐021 64.60 68.60 4.00 14.15 102.6 1.18 1.68 ‐DM‐09‐022 69.60 70.50 0.90 0.02 0.7 ‐ ‐ ‐DM‐09‐022 72.50 73.50 1.00 5.37 13.6 ‐ ‐ ‐DM‐09‐022 74.10 75.30 1.20 11.53 84.0 1.06 ‐ ‐DM‐09‐023 65.90 66.10 0.20 0.03 0.2 ‐ ‐ ‐DM‐09‐023 71.40 73.60 2.20 0.93 2.7 ‐ ‐ ‐DM‐09‐023 78.40 79.60 1.20 21.90 89.6 ‐ ‐ ‐DM‐09‐024 116.00 117.00 1.00 0.03 0.7 ‐ ‐ ‐DM‐09‐024 120.70 122.10 1.40 8.89 57.6 ‐ 4.41 ‐DM‐09‐025 150.00 150.40 0.40 0.17 2.3 ‐ ‐ ‐DM‐09‐025 154.50 155.00 0.50 6.83 62.1 ‐ ‐ ‐DM‐09‐025 162.00 162.60 0.60 1.07 5.5 ‐ ‐ ‐DM‐09‐026 25.30 25.70 0.40 8.00 26.9 ‐ 1.34 ‐DM‐09‐026 116.40 117.50 1.10 0.03 3.6 ‐ ‐ ‐DM‐09‐026 119.80 121.90 2.10 13.06 57.6 ‐ ‐ ‐DM‐09‐026 134.60 135.60 1.00 0.17 4.3 ‐ ‐ ‐DM‐09‐026 150.00 151.30 1.30 3.17 25.2 ‐ ‐ ‐DM‐09‐027 119.60 120.30 0.70 0.47 1.1 ‐ ‐ ‐DM‐09‐027 126.70 127.40 0.70 18.04 166.2 ‐ 1.83 ‐DM‐09‐027 143.50 144.20 0.70 0.23 0.2 ‐ ‐ ‐DM‐09‐027 157.20 157.80 0.60 8.72 84.0 ‐ ‐ ‐DM‐09‐028 38.00 39.00 1.00 3.00 10.6 ‐ ‐ ‐DM‐09‐028 115.10 115.50 0.40 0.37 1.9 ‐ ‐ ‐DM‐09‐028 121.80 122.40 0.60 2.40 386.6 ‐ ‐ ‐DM‐09‐028 156.40 156.90 0.50 0.03 0.8 ‐ ‐ ‐DM‐09‐029 119.60 120.20 0.60 23.46 84.5 ‐ ‐ ‐DM‐09‐029 120.20 128.00 7.80 3.01 13.7 ‐ ‐ ‐DM‐09‐029 128.00 129.10 1.10 14.31 141.5 ‐ ‐ ‐DM‐09‐030 119.50 120.00 0.50 9.23 102.9 ‐ ‐ ‐DM‐09‐030 121.50 122.00 0.50 16.93 30.8 ‐ ‐ ‐DM‐09‐030 123.00 125.50 2.50 15.61 70.1 ‐ ‐ ‐DM‐09‐030 125.50 127.50 2.00 6.79 55.0 ‐ ‐ ‐DM‐09‐031 118.80 119.70 0.90 0.07 0.1 ‐ ‐ ‐DM‐09‐031 127.40 130.70 3.30 4.38 47.6 ‐ ‐ ‐DM‐09‐032 112.30 113.60 1.30 11.90 122.4 ‐ ‐ ‐DM‐09‐032 115.00 119.10 4.10 25.52 73.8 ‐ ‐ ‐DM‐09‐033 113.20 114.00 0.80 0.10 0.1 ‐ ‐ ‐DM‐09‐033 118.00 118.40 0.40 0.77 14.3 ‐ ‐ ‐DM‐09‐033 122.00 123.00 1.00 2.10 1.1 ‐ ‐ ‐DM‐09‐034 115.00 116.00 1.00 4.33 4.1 ‐ ‐ ‐DM‐09‐034 122.80 123.50 0.70 18.86 72.4 ‐ ‐ ‐DM‐09‐035 120.00 122.90 2.90 7.37 14.3 ‐ ‐ ‐DM‐09‐036 112.30 112.60 0.30 0.93 8.3 ‐ ‐ ‐DM‐09‐036 119.40 120.50 1.10 7.54 171.8 ‐ ‐ ‐


DM‐09‐037 63.50 64.10 0.60 0.02 0.2 ‐ ‐ ‐DM‐09‐037 66.10 68.00 1.90 16.12 114.1 1.22 1.44 ‐DM‐09‐038 72.00 72.50 0.50 0.07 3.5 ‐ ‐ ‐DM‐09‐038 77.70 78.60 0.90 11.84 71.7 ‐ ‐ ‐DM‐09‐038 79.80 80.20 0.40 8.17 42.4 ‐ ‐ ‐DM‐09‐039 55.30 56.90 1.60 13.68 35.9 1.56 ‐ ‐DM‐09‐039 56.90 57.60 0.70 15.20 19.2 ‐ ‐ ‐DM‐09‐040 57.00 58.00 1.00 2.60 18.4 ‐ ‐ ‐DM‐09‐040 59.90 61.50 1.60 18.72 69.2 ‐ ‐ ‐DM‐09‐041 64.50 65.00 0.50 3.13 4.7 ‐ ‐ ‐DM‐09‐041 68.00 69.40 1.40 12.24 42.4 ‐ ‐ ‐DM‐09‐042 74.60 76.20 1.60 2.68 27.5 ‐ ‐ ‐DM‐09‐042 76.70 77.30 0.60 6.26 68.4 ‐ ‐ ‐DM‐09‐043 116.20 116.60 0.40 0.73 2.4 ‐ ‐ ‐DM‐09‐043 215.50 217.00 1.50 8.55 25.0 ‐ ‐ ‐DM‐09‐043 293.50 293.70 0.20 1.20 6.0 ‐ ‐ ‐DM‐09‐044 24.00 24.70 0.70 2.07 2.2 ‐ ‐ ‐DM‐09‐045 22.00 22.30 0.30 0.27 1.7 ‐ ‐ ‐DM‐09‐046 49.40 50.00 0.60 19.07 14.0 ‐ 6.70 ‐

Appendix V

Assayers Canada Fire Assay and ICP Analytical Procedure

Appendix VI

Ore Reference Standards

CDN Resource Laboratories Ltd. #2, 20148 – 102 Avenue, Langley, B.C., Canada, V1M 4B4, 604-882-8422, Fax: 604-882-8466 (www.cdnlabs.com)

GOLD ORE REFERENCE MATERIAL: CDN-GS-8A

Recommended value and the "Between Laboratory" two standard deviations

Gold concentration: 8.25 ! 0.60 g/t

PREPARED BY: CDN Resource Laboratories Ltd. CERTIFIED BY: Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia INDEPENDENT GEOCHEMIST: Dr. Barry Smee., Ph.D., P. Geo. DATE OF CERTIFICATION: July 15, 2009 ORIGIN OF REFERENCE MATERIAL: Standard CDN-GS-8A was prepared using ore supplied by Comaplex Minerals Corporation. The ore is from the 1100 lode of the Tiriganiaq Gold Deposit north of Rankin Inlet in Nunavut. It is a banded magnetite iron formation zone with gold in quartz shears with accessory pyrrhotite, pyrite, and arsenopyrite. The gold is free milling although there may be a small refractory component. METHOD OF PREPARATION: Reject ore material (640 kg of Comaplex ore plus 160 kg of blank granitic ore) was dried, crushed, pulverized and then passed through a 270 mesh screen. The +270 material was discarded. The -270 material was mixed for 6 days in a double-cone blender. Splits were taken and sent to 13 commercial laboratories for round robin assaying. Round robin results are displayed below:

Lab 1 Lab 2 Lab 3 Lab 4 Lab 5 Lab 6 Lab 7 Lab 8 Lab 9 Lab 10 Lab 11 Lab 12 Lab 13Sample Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t) Au (g/t)

GS8A-1 8.50 8.46 8.68 7.65 8.01 8.24 8.69 8.29 8.07 8.59 8.35 8.55 8.02GS8A-2 8.04 8.54 8.01 8.10 7.96 8.21 7.59 8.18 8.31 8.17 8.06 8.35 8.29GS8A-3 8.01 8.21 8.52 7.82 8.16 8.25 8.91 8.70 7.84 8.64 8.12 8.38 7.89GS8A-4 8.89 8.2 8.17 8.62 8.27 8.18 8.25 8.54 8.90 7.80 8.18 8.15 8.00GS8A-5 8.31 8.15 8.09 8.60 8.84 8.19 7.87 8.35 8.71 8.60 8.08 8.41 8.22GS8A-6 8.83 8.28 8.07 8.10 7.58 8.20 7.96 8.86 8.36 8.24 7.60 8.48 8.13GS8A-7 7.88 8.61 8.57 8.33 7.42 8.21 8.20 8.29 8.40 8.21 8.07 8.16 8.10GS8A-8 8.02 8.34 8.58 8.10 7.76 8.21 7.75 8.12 8.91 8.60 7.77 8.56 8.12GS8A-9 7.94 8.32 8.40 8.78 8.09 8.27 8.39 7.97 8.82 8.54 7.67 8.54 8.32GS8A-10 8.78 8.09 8.58 7.52 7.54 8.18 8.21 7.85 8.40 8.52 7.81 8.17 7.78

Mean 8.32 8.32 8.37 8.16 7.96 8.21 8.18 8.32 8.47 8.39 7.97 8.38 8.09Std. Dev. 0.399 0.171 0.255 0.424 0.419 0.031 0.411 0.314 0.359 0.275 0.244 0.165 0.169%RSD 4.79 2.05 3.05 5.19 5.27 0.38 5.02 3.78 4.24 3.28 3.05 1.97 2.09 Assay Procedure: all assays were fire assay, ICP finish on 30g samples APPROXIMATE CHEMICAL COMPOSITION:

Percent Percent SiO2 70.6 Na2O 1.4

Al2O3 8.6 MgO 1.7 Fe2O3 6.9 K2O 1.7 CaO 3.5 TiO2 0.4 MnO 0.1 LOI 3.6

S 1.0

GOLD ORE REFERENCE MATERIAL: CDN-GS-8A

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The mean and standard deviation were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

Participating Laboratories: (not in same order as table of assays) Acme Analytical Laboratories Ltd., Vancouver, Canada Activation Laboratories, Ancaster, Ontario, Canada Activation Laboratories, Thunder Bay, Ontario, Canada ALS Chemex, North Vancouver, Canada Alaska Assay Laboratories, Alaska, USA

Assayers Canada Ltd., Vancouver, Canada Eco Tech, B.C., Canada Genalysis Lab.Services, Australia Labtium Inc., Finland Omac Laboratory, Ireland SGS Toronto, Canada TSL Laboratories Ltd., Saskatoon, Canada Ultra Trace Pty. Ltd., Australia

Legal Notice: This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. nor Barry Smee accept any liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by _____________________________________ Duncan Sanderson, Certified Assayer of B.C.

Geochemist _____________________________________

Dr. Barry Smee, Ph.D., P. Geo.

CDN Resource Laboratories Ltd. #2, 20148 - 102nd Avenue, Langley, B.C., Canada, V1M 4B4, Ph: 604-882-8422 Fax: 604-882-8466 (www.cdnlabs.com)

GOLD ORE REFERENCE STANDARD: CDN-GS-11A

Recommended value and the "Between Laboratory" two standard deviations

Gold concentration: 11.21 ! 0.87 g/t

PREPARED BY: CDN Resource Laboratories Ltd. CERTIFIED BY: Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia INDEPENDENT GEOCHEMIST: Dr. Barry Smee., Ph.D., P. Geo. DATE OF CERTIFICATION: February 10, 2009 ORIGIN OF REFERENCE MATERIAL: Standard CDN-GS-11A was prepared using ore supplied by Comaplex Minerals Corporation. The ore is from the 1100 lode of the Tiriganiaq Gold Deposit north of Rankin Inlet in Nunavut. It is a banded magnetite iron formation zone with gold in quartz shears with accessory pyrrhotite, pyrite, and arsenopyrite. METHOD OF PREPARATION: Reject ore material was dried, crushed, pulverized and then passed through a 200 mesh screen. The +200 material was discarded. The -200 material was mixed for 6 days in a double-cone blender. Splits were taken and sent to 12 commercial laboratories for round robin assaying. Round robin results are displayed below:

Lab 1 Lab 2 Lab 3 Lab 4 Lab 5 Lab 6 Lab 7 Lab 8 Lab 9 Lab 10 Lab 11 Lab 12Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t Au g/t

GS11A-1 10.81 11.05 11.07 12.20 10.76 10.9 10.1 10.50 11.30 11.64 11.29 11.0GS11A-2 11.37 11.20 11.80 10.95 10.34 11.3 10.0 11.00 11.70 12.08 11.43 10.9GS11A-3 10.60 11.40 10.90 11.25 10.94 11.1 10.6 11.61 11.30 11.40 11.66 11.3GS11A-4 11.91 11.15 11.10 10.60 10.77 11.8 10.4 11.60 11.00 10.84 10.90 10.8GS11A-5 11.71 11.35 10.80 12.10 10.66 11.4 10.2 10.04 11.70 12.48 11.18 11.1GS11A-6 11.77 11.25 11.50 11.90 11.06 11.5 10.5 10.19 12.00 12.20 11.34 11.2GS11A-7 11.33 10.75 11.43 13.10 10.64 11.6 10.9 10.55 10.30 11.81 10.90 10.6GS11A-8 11.16 10.90 11.57 10.90 11.12 11.6 10.9 10.93 11.00 11.56 11.50 11.3GS11A-9 11.56 11.20 11.30 11.10 10.83 11.2 10.8 11.51 12.30 11.56 10.86 10.7GS11A-10 11.53 11.05 11.73 10.15 10.38 10.8 10.1 10.76 11.00 11.16 11.84 11.2

Mean 11.38 11.13 11.32 11.43 10.75 11.32 10.45 10.87 11.36 11.67 11.29 11.01Std. Dev. 0.419 0.199 0.344 0.884 0.261 0.322 0.344 0.569 0.582 0.491 0.334 0.251%RSD 3.68 1.79 3.04 7.74 2.43 2.85 3.29 5.23 5.12 4.21 2.96 2.28 Assay Procedure: all assays were fire assay, gravimetric finish on 30g samples except for labs 8 and 12 which used ICP finish. APPROXIMATE CHEMICAL COMPOSITION:

Percent Percent SiO2 70.8 Na2O 0.9

Al2O3 9.1 MgO 1.8 Fe2O3 6.8 K2O 2.6 CaO 3.3 TiO2 0.4 MnO 0.1 LOI 4.0

S 0.9 C 1.1

GOLD ORE REFERENCE STANDARD: CDN-GS-11A

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The mean and standard deviation were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

Participating Laboratories: (not in same order as table of assays) Acme Analytical Laboratories Ltd., Vancouver, Canada Activation Laboratories, Ontario, Canada ALS Brisbane, Australia

ALS Chemex, North Vancouver, Canada Assayers Canada Ltd., Vancouver, Canada Alex Stewart (Assayers) Argentina Ltd. Genalysis Lab.Services, Australia Labtium Inc., Finland Omac Laboratory, Ireland Skyline Assayers & Laboratories Ltd, Arizona, USA

TSL Laboratories Ltd., Saskatoon, Canada Ultra Trace Pty. Ltd., Australia Legal Notice: This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. nor Barry Smee accept any liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by _____________________________________ Duncan Sanderson, Certified Assayer of B.C.

Geochemist _____________________________________

Dr. Barry Smee, Ph.D., P. Geo.

Appendix VII

Acme Fire Assay Analytical Procedure

CARECOMMITMENTPERFORMANCE

CARECOMMITMENTPERFORMANCE ISO 9001:2008

FM 63007

1020 Cordova Street East, Vancouver BC V6A 4A3 Phone (604) 253 3158 Fax (604) 253 1716 e-mail: [email protected]

Sample Preparation version 2.2 Revision Date: September 3, 2009

GENERAL SAMPLE PREPARATION METHODS Soil and Rocks and Sediments Drill Core Comments A Yes Yes A No No

Samples arrive from Client

Do pulp samples

need additional

processing?

Samples arrive from Client

Receiving: Samples arrive via courier, post or by client drop-off; shipment inspected for completeness.

Sorting and Inspection: Samples sorted and inspected for quality of use (quantity and condition). Pulp samples inspected for homogeneity and fineness. Coarse pulps are screened or pulverized after getting client’s approval.

Drying: Wet or damp samples are dried at 60°C (40°C if specified by the client).

Sieving: Soil and sediment sieved to -80 mesh ASTM (-180 microns) unless client specifies otherwise. Sieve cleaned by brush and compressed air between samples. Reference material G-1 (pulp made of granite blank) is carried as first sample in sequence (sieve›weigh›digest›analyse) to monitor background noise.

Crushing and Pulverizing: Rock and Drill Core crushed to 80% passing 10 mesh (2 mm), homogenized, riffle split (250 g subsample) and pulverized to 85% passing 200 mesh (75 microns). Crusher and pulverizer are cleaned by brush and compressed air between routine samples. Granite wash scours equipment after high-grade samples, between changes in rock colour and at end of each file. Granite is crushed and pulverized as first sample in sequence and carried through to analysis to monitor background noise.

Compositing: Equal weights of crushed, pulverized or sieved material from 2 or more samples are combined and pulverized for 60+ seconds to produce a homogeneous mixture.

Storage: Pulp samples (up to 100g for soils or sediments and up to 250 g for rock and drill core) are archived for 3 months at no cost. Soil and sediment rejects are discarded immediately. Rock and drill core rejects are stored for 3 months at no charge. Client may request additional storage, return or disposal of pulps and rejects after initial free storage period.

Samples sorted and inspected

Samples dried

Samples sieved to desired mesh

Samples sorted and inspected

Samples dried

Samples crushed and pulverized to desired

mesh

Pulps retained for analysis and storage

(3 months free)

Pulps retained for analysis. Pulps and rejects stored for 3

months at no charge

Appendix VIII

Assayers Canada Metallic Gold Analytical Procedure

Filename.doc Rev. 01/2002

Procedure Summary:

F141, F 142: Metallic Gold Fire Assay

Elements Analyzed:

Metallic Gold

Procedure:

A 500 g sample is pulverized to 95%-150 mesh, sieve out to +150 mesh and -150 meshportions. The whole sample is fire assayed for +150 and two times at 1 AT charge for the – 150.The total gold content is calculated based on the +150 and -150 gold concentration as

Total gold(mg)=(Concentration of +150(g/t) x Wt of +150) +(Concentration of -150(g/t) x Wt of -150)

Detection Limit: 0.01g/tone

8282 Sherbrooke Street,Vancouver, B.C.

Canada V5X 4R6Tel: 604 327-3436Fax: 604 327-3423

Appendix IX

Memo from Moose Mountain Technical ServicesRe: Dome Mountain Vein Modeling

DomeModel description (Mar-5).doc ! Page 1

MemoFrom: Mike Takkinen

To: Daryl Hanson

Cc: Jim Gray; Gary Giroux

Date: February 28, 2010

Re: Dome Mountain Vein Modeling

Moose Mountain Technical Services (MMTS) was tasked with the responsibility to create threedimensional solid models of mineralized zones in the vicinity of previously developed mining levels atthe Dome Mountain property. The veins of interest have historically been referred to as the Boulder andArgillite veins. Preliminary review of the available drill hole data indicate additional distinct and relativelycontinuous mineralization zones in limited areas of the hanging-wall of the primary targets.Furthermore, local knowledge indicates the westernmost panel developed by previous owners waslocated in the foot-wall of the Boulder horizon. Thus, in addition to the Boulder (BLDR) and Argillite(ARG) veins, 3d solid models were also created for theWest Hanging-wall (WHW), East Hanging-wall(EHW), and Boulder Foot-wall (BFW) zones.

Figure 17.1: Three dimensional vein solids modeled for Dome Mountain

DomeModel description (Mar-5).doc ! Page 2

The lower elevation limit of modeled solids is defined by the lower-most mineralized assays intersectedby current drilling.To create the 3d solids, historical drilling (including 2009) and mining data, as provided by Eagle PeakResources (EPR), are compiled in MineSight! (MS) using standard drill hole, survey, and translatedAutocad! files. Preliminary mineralization zones, indicated by assay data, were assigned by EPR andthen checked and rationalized by MMTS in the MS 3d workspace. The Domemodel zone horizons arenormally defined by the occurrence of quartz veins but also include non-quartz intersections wheresignificant mineralization is indicated and the intersection generally conforms to vein geometry.The following figure portrays the concepts utilized in the development of the Dome Mountain veinmodels.

Figure 17.2 Concept diagram of Dome Mountain vein modelQuartz intersections (1.) provide the key indicator for the mineralized zone. Assays indicatingmineralization (2.) further define the extent of the vein (3.). Intersections are assigned to respectiveveins based on lithological features and conformity to the overall strike and dip of the zone. (Note:Quartz and mineralized areas 1a. and 2a. are shown for illustrative purposes. Only the overall veinshape (3.) has been modeled for volumetric purposes.)Mineralized intersections that do not conform to the predominant geometry of the zone are excludedfrom the modeled 3d solids (6.). Of the 802 assigned assays in the area of interest, 213 intersectionsare not included in the present model and therefore represent an area of future model developmentand potential additional resources. The area upsection of the Argillite zone contains the greatestdensity of unassigned, mineralized assays.In addition to drill hole intersections, ‘as mined’ survey data (EPR provided) of the vein on the 1290 andestimates on the1370 level, are utilized to define the extent of the Boulder and Argillite zones in areasof previous development. Similarly, stope surveys above the 1290 level are utilized to conform themodeled solids to known previous development work. The previous mining on the 1290 & 1370 levelssignificantly adds to the credibility and continuity of the geological interpretation of Bolulder veinPreliminary zone bottom surfaces for each zone are created by triangulating drill hole and survey datapoints assigned to a given zone. To prevent the overestimate of zone volume, minimal thicknessintersections (0.01m) are assigned to drill holes intersecting the predicted mineralization surface butthat show no (or little) indication of mineralization at the zone horizon. Zone top surfaces are then

2.a5.

2. Assaysindicatingmineralization

1. Quartzlogged &assayed

3. Mineralizationzone limit indicated byzone assignments

4. No HW orFW assaysavailable

5. Footwall &hanging-wall ‘skins’ at0.5m (vertical)increments for dilutionanalysis.

1 a.

2.a

2.a

6.6.

6.

6.

6. Discontinuous mineralized intersections not included invein models at this time.

1 a.

1 a.

! Page 3

created by triangulating drill hole, survey, and minimal thickness data points. The solid model is thencomprised of the union of bottom and top surfaces for each zone..Traditional modeling techniques have included drawing vein shapes on two dimensional sections andprojecting the shape midway to the bordering sections. This method however, fails to honour theprecise location of the vein intercepts, and, when drill holes are not perfectly parallel to the section, canintroduce apparent thickness error. Overestimates of volume due to apparent thickness are thereforecontrolled in the Domemodel by using surfaces defined only by actual 3d drill and survey data points.Furthermore, rationalization and validation of the DomeMountain zone solids has been completed byensuring that overlaps and subsurface intersections (ie/ near zone merges and at the bedrock /overburden horizon) have been eliminated. Finally, zone solid volumes were estimated in Minesightand corroborated by GCL Consulting.

Appendix X

Petrographic Evaluation of a Zinc, Copper, Lead, Silver andGold-Bearing Composite from Dome Mountain, British Columbia

Inspectorate America Corporation Metals & Minerals Division 101 Woodland Highway Belle Chasse, LA 70037 Tel: 504-392-7660 Fax: 504-393-5248

www.inspectorate.com

Inspectorate America Corp. Petrographic Services

Petrographic Evaluation of a Zinc, Copper, Lead, Silver and Gold-bearing Composite From Dome Mountain, British Columbia

For

Process Research Associates

Eagle Peak Resources

November 23, 2009

Wesley Mathews Petrographic Services Manager

Karla Clayton Petrologist

2 Petrographic Services Introduction

The following analysis was performed on a copper, zinc, lead, gold and silver-bearing master composite: A bulk head composite was examined to determine mineral bulk modal percentages and mineral associations in support of gold processing efforts. Bulk locking and liberation with degree of locking analysis was performed on the sulfide minerals. The methods used in the modal analysis include reflected light microscopy (E400 Petrographic Microscope) for ore minerals and X-ray diffraction (XRD). Sulfide mineral abundances and associations were collected via statistical modal analysis, and quantified by normalization to assay, while non-sulfide gangue data was normalized to semi-quantitative XRD. Liberation and locking data for sulfides was calculated as statistical grain percentages. A gold associations study was also undertaken. The purpose was to determine the appropriate method or methods of processing the gold. The methods used in the study include reflected light microscopy (E400 and Nikon Optiphot Petrographic Microscope). Associations for sulfides were likewise calculated as statistical grain percentages. Photomicrographs were taken in support of the analyses. The following details the samples submitted for this study.

Sample Identification

Sulfur Samples (Bulk Modal Analysis, Mineral Associations, Locking and Liberation, with Degrees of Locking) 1. 0905808-H Master Composite

Summary of Results Optical Results Pyrite was the dominant sulfide present in the 0905808-H Master Composite, comprising 5.15% of the sample by bulk modal analysis. Other sulfides present in the composite include sphalerite (1.51%), chalcopyrite (1.21%), galena (0.19%) and chalcocite (0.02%). Trace amounts (<0.01%) of bornite were observed. Sulfides were commonly multiply associated with one another, with pyrite-sphalerite, pyrite-galena and pyrite-chalcopyrite associations the most common, and sphalerite-chalcopyrite and sphalerite-galena associations also common. However, gangue was typically non-opaque (91.3%), with silicates (mostly quartz, 44.5% with accessory micas, 19.2% muscovite, clays, 5.6%, and feldspars, 2.8%) dominant. Carbonates were also present, at 19.2%, while minor iron oxides were also detected (0.6%). Sulfates were not detected above the detection limit of 0.5% by X-ray diffraction; wet chemistry by PRA gave 0.03% SO4

2- concentration. Graphite was not observed, though graphitic carbon assay by Leco by PRA indicated 0.05% graphite was present. Hosting minerals tended largely to be non-opaque as well. Silver minerals were not observed, therefore it is likely that silver resides in another mineral, possibly galena. It is also possible that silver minerals are submicroscopic. The liberation and locking profile of the bulk sulfides indicate that about 84% of the sulfide grains were liberated, and 16% were locked. Considering the mean sizes of liberated and locked sulfides (42µm and 188µm, respectively), the weight percent sulfide liberation is 6% by volume-weighted average. Gold metal grains are primarily observed to be associated with pyrite (57% statistically and 98% by volume weighted mean), though some of the gold-hosting pyrite is itself encapsulated by other gangue minerals, primarily non-opaque gangue. The hosting minerals (encapsulating grains which contained the gold and its tertiary, quaternary and quinternary associations) were statistically non-opaque gangue (41%), and secondarily pyrite (34%). Other hosting minerals included iron oxides, chalcopyrite and sphalerite. In all, 64% (statistically) of the gold associations involved sulfides (99% by volume weighted mean). Small monomineralic gold was also present at low levels. All of the gold observed was slightly pale, suggesting that some electrum component may be present. The gold averages 51µm in diameter, and ranges from 1-1060µm. One ~1mm gold nugget was observed, and was associated with non-opaque gangue. As no other nuggets were observed, this occurrence was ignored for the purposes of calculating volume-weighted mean associations, though it should be noted that a certain nugget effect may cause data reproducibility problems, as nuggets may skew some test results towards free gold.

3 Petrographic Services Processing Notes With the current 10-mesh crush, much of the gold, aside from random large nuggets, is statistically leaching-resistant. Moreover, some of the gold-hosting sulfides are themselves encapsulated by large gangue grains. Grinding to free these grains should be prioritized. The size distribution of the locked sulfides suggests that at least a 150µm crush would begin to free the hosting sulfides. The gold itself would then require about a 50µm crush and leach. Preg-robbing phases such as graphite and pyrrhotite are rare; the chief source of carbon in the composite is the mineral ankerite. Overall, the sulfide association data shows that high gold recoveries may be achieved by targeting pyrite. Leaching should also be employed to sequester monomineralic gold and any remaining gold nuggets that remain.

Inspectorate America Corporation Metals & Minerals Division 101 Woodland Highway Belle Chasse, LA 70037 Tel: 504-392-7660 Fax: 504-393-5248

www.inspectorate.com

Data and Figures

Size (µm) Cu Assay Au, ppm Chalcopyrite Covellite Bornite Chalcocite Pyrite Iron Oxides Non-Opaque Sphalerite Galena Other Lib LockedTotal 0.43 13.79 1.21 0.00 Trace 0.02 5.15 0.6 91.3 1.51 0.19 Trace 84 16

*Other minerals include Malachite, Ilmenite, and Cuprite

Modal Analysis of Size Fractions, wt. %

PercentTotal Cu-Sulfide (ch,cov,bor,cc) 1.2Total Fe-Sulfide (pyr & mar) 5.1Total Zn-Sulfide (sphalerite) 1.5Total Mo-Sulfide (molybdenite) 0.0Total Pb-Sulfide (galena) <0.5Total Fe-Oxide (mag & hem) 0.6Total Carbonate (cal,dolo,sid,ank,mag) 19.2Total Sulfate (anhy,gyp,bass) <0.5Total Quartz 44.5Total Feldspar (alb,ortho,micro) 2.8Total Andalusite 0.0Total Garnet (andradite, grossular) 0.0Total Olivine (forsterite) <0.5Total Mica (mus,bio,phlog) 19.2Total Clay (kao,chl,ill,mont) 5.6Total Talc 0.0Total Pyroxene (diopside, augite) 0.0Total Amphibole/Serpentine 0.0

normalized sum: 100

Mineral SummaryXRD Analysis - Bulk Sample

5 Petrographic Services

0905808 Master CompositeOre Mineralogy

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

Total

Perc

ent

GalenaSphaleriteChalcociteBorniteCovelliteChalcopyrite

0905808 Master CompositeGangue Mineralogy

0

20

40

60

80

100

Total

Perc

ent Other

Non-OpaqueIron OxidesPyrite

Size (µm)Iron Oxides Non-Opaque Other

Bulk 2 98 Trace42µm / 1-1800µm 188µm / 4-1910µm

Non-Sulfide Hosting MineralsLiberation Profile of Bulk Sulfides

Liberated LockedMean Size / Size Range ( µm) Mean Size / Size Range ( µm)

0905808 Master CompositeLiberated Bulk Sulfide Size Range

1-1800µm

2-810µm

2-53µm

1-1800µm

2-1713µm

10-218µm

0 500 1000 1500 2000

Bulk

Chalcopyrite

Bornite

Chalcocite

Pyrite

Sphalerite

Galena

Size (µm)

0905808 Master CompositeLocked Bulk Sulfide Size Range

4-1910µm

4-1210µm

13-250µm

2-1100µm

5-1910µm

4-1770µm

13-540µm

0 500 1000 1500 2000 2500

Bulk

Chalcopyrite

Bornite

Chalcocite

Pyrite

Sphalerite

Galena

Size (µm)


0905808 Master CompositeLocked Sulfide Distribution

Renormalized to 100%

5 71916

19

34

66 100 86

67

41

98

138 6

2 22

1

512

0

20

40

60

80

100

Chalcopyrite Bornite Chalcocite Pyrite Sphalerite Galena Bulk Sulfide

Perc

ent

GalenaSphaleriteNon-OpaqueIron OxidesPyriteChalcopyrite

Size (µm) Size (µm)Liberated 0-50% 50-70% 70-90% 90-100% Mosaic Dissem. Liberated 0-50% 50-70% 70-90% 90-100% Mosaic Dissem.

Bulk 81 10 2 2 2 0 5 Bulk 0 0 0 0 Trace 0 Trace


Bulk 27 18 9 9 9 0 27 Bulk 89 4 1 2 Trace 0 4


Bulk 71 14 Trace 4 7 0 4 Bulk 67 17 8 Trace 8 0 Trace

Size (µm)Liberated 0-50% 50-70% 70-90% 90-100% Mosaic Dissem.

Bulk 84 7 1 2 2 0 4

Chalcocite, Statistical Grain %

Chalcopyrite, Statistical Grain %Locked

Bornite, Statistical Grain %Locked

Pyrite, Statistical Grain %Locked Locked

Bulk Sulfide, Statistical Grain %Locked

Degrees of LockingLocking and Liberation Study

Locked LockedSphalerite, Statistical Grain % Galena, Statistical Grain %


0905808 Master CompositeSulfide Degrees of Locking

0

20

40

60

80

100

Chalcopyrite Bornite Chalcocite Pyrite Sphalerite Galena Bulk Sulfide

Perc

ent

Dissem.Mosaic90-100%70-90%50-70%0-50%Liberated

SizeFraction Pyrite Iron Oxides Non-Opaque Sphalerite Galena Pyrite Iron Oxides Non-Opaque Sphalerite Galena Chalcopyrite Covellite Bornite Chalcocite

TotalChalcopyrite 16 2 66 13 2 0.50 2.50 0.11 1.97 6.25 0.00 0.00 Trace

Covellite 0 0 0 0 0 0 0 0 0 0 0.00 0.00 0.00Bornite Trace 0 Trace Trace 0 0 0 0 0 0 0.00 0.00 Trace

Chalcocite 0 0 100 0 0 Trace 0 Trace 0 0 Trace 0.00 Trace

% Liberated % Locked % in Pyrite % in Iron Oxide % in Non Opaque % in Sphalerite % in GalenaPyrite 89 11 0 9 1 0Iron Oxides* 15 85 3 78 3 0Non-Opaque* 99 1 1 0 0 0Sphalerite 71 29 5 0 19 1Galena 67 33 11 0 14 2

Cu-S Binary Associations% Cu-S Mineral Locked in Other Mineral % Cu-S Mineral with Cu-S Mineral% Other Mineral Locked in Cu-S Mineral

Cu-S / Gangue Associations

Other Mineral Locking and Associations


Size (µm) Mean Min. Max Mean Min. Max Mean Min. Max Mean Min. MaxBulk 25 2 810 0 0 0 21 2 53 36 1 1800

Size (µm) Mean Min. Max Mean Min. MaxBulk 108 2 1713 58 10 218

Size (µm) Mean Min. Max Mean Min. Max Mean Min. Max Mean Min. MaxBulk 118 4 1210 80 13 250 193 10 1100 303 5 1910



Pyrite

Sphalerite Galena

Chalcopyrite Bornite ChalcociteSize Data for Liberated Sulfides

Size Data for Locked SulfidesChalcopyrite Bornite Chalcocite Pyrite

Sphalerite Galena

Iron Oxides Non-OpaqueSize Data for Non-Sulfide Gangue Hosting Sulfides

9 Petrographic Services Gold Associations Study

Hosting Mineral %Gold 11

Dual Pyr Only 18 Pyr & N-O 5 Pyr & Gal 2

Pyr & Born 2 Pyr, N-O, & Gal 2

Pyr, Gal, & C 2Quinternary Pyr, Sphal, Gal, & C 2

Dual Fe Only 9Ternary Fe & Pyr 4

Quaternary Fe, Cc, & N-O 2Dual N-O Only 13

N-O & Pyr 13 N-O & Fe 4

N-O, Pyr & Sphal 2 N-O, Pyr & Born 2 N-O, Sphal & Gal 2

Quinternary N-O, Pyr, C & Cc 2Ternary Sphal & N-O 2

Quaternary Sphal, Pyr & C 2Quaternary C, Gal, & Cc 2Quinternary C, Born, Pyr, & Sphal 2

Pyrite Ternary

Quaternary

Sphalerite

Chalcopyrite

Gold Associations, Statistical Grain %

MonomineralicAssociations

Fe Oxides

Non-OpaqueTernary

Quaternary

Mean Min MaxBulk Gold 51 1 1060Bulk Gangue 580 20 2800Pyrite 468 20 1800Fe Oxides 232 30 530Non-Opaque 916 60 2800Sphalerite 60 25 95Chalcopyrite 198 130 265

Sizes (µm)


0905808 Master Composite - Gold Metal Hosting Minerals, Renormalized to 100% (Statistical Grain %)

Sphalerite4%

Chalcopyrite4% Pyrite

34%

Fe Oxides17%

Non-Opaque41%

PyriteFe OxidesNon-OpaqueSphaleriteChalcopyrite

0905808 Master CompositeSize Ranges of Associations

20-2800µm

135-265µm

25-95µm

60-2800µm

30-530µm

20-1800µm

1-1060µm

0 500 1000 1500 2000 2500 3000

Gangue

Chalcopyrite

Sphalerite

Non-Opaque

Fe Oxides

Pyrite

Gold

Size (µm)

11 Petrographic Services Photomicrographs

PHOTO 1 PHOTO 2

PHOTO 3 PHOTO 4


PHOTO 5 PHOTO 6

The interpretations or opinions expressed represent the best judgment of Inspectorate America Corp. and it assumes no responsibility and makes no warranty or representations, as to the profitability or liability of any substance, material and/or natural resource. These analyses, opinions or interpretations are based on observations and materials supplied by the client for whom this report is made.

Appendix XI

Proposal on Metallurgical Testing - Dome Mountain Project

Prepared for: Eagle Peak Resources Inc.

Suite 413 – Bentall 3 595 Burrard Street, P.O. Box 49096 Vancouver, BC V7X 1G4

Attention: Mr. Mr. Lloyd Tattersall, Vice President cc: David Pow, Director cc: Mr. Matt Bolu, Bolu Consulting Engineering

Prepared by: Inspectorate America Corporation PRA Metallurgical Division 11620 Horseshoe Way Richmond, B.C. V7A 4V5

PRA Proposal No.: P0907408

Date: August 10, 2009

PROPOSAL ON METALLURGICAL TESTING FOR GOLD AND SILVER RECOVERY, AS REQUESTED DOME MOUNTAIN PROJECT CENTRAL BRITISH COLUMBIA

Eagle Peak – Dome Mountain - 1 -

!Inspectorate America Corporation - PRA Metallurgical Division

TABLE OF CONTENTS

Page No.

1.0 INTRODUCTION ........................................................................................ 2

2.0 SCOPE OF SERVICES .............................................................................. 3 2.1 Sample Preparation .................................................................................. 3 2.2 Head Assays ............................................................................................. 3 2.3 Heap Leach Exploration .......................................................................... 4 2.4 Flotation Tests .......................................................................................... 5 2.5 Gravity Tests ............................................................................................ 5 2.6 Agitated Cyanide Leach ........................................................................... 5 2.7 Alternative Leach Tests ........................................................................... 6 2.8 Thickening and Filtration ......................................................................... 6 2.9 Grindability Tests ..................................................................................... 6 2.10 Detoxification Test ................................................................................... 6 2.11 Bulk Production Leach Tests .................................................................. 7

2.12 Summary Overview .................................................................................. 7

3.0 COSTS AND INVOICING .......................................................................... 8

4.0 SCHEDULE ............................................................................................. 10

5.0 KEY PERSONNEL ................................................................................... 11 APPENDIX A – Task and Cost Details: Table 1

APPENDIX B – Corporate Profile



1.0 INTRODUCTION

A Request for Proposal (RFP) was issued on behalf of Eagle Peak Resources Inc.

(EPR) by Bolu Consulting Engineering, on August 5th, 2009. Objectives are to

optimize the potential of gold and silver recovery for the Dome Mountain project in

central BC, by cyanide-leaching mainly. Inspectorate America’s PRA Metallurgical

Division (PRA) is pleased to respond with this Proposal, building on its recent

testing experience on two of the client’s major projects, and numerous programs

for similar undertakings.

Inspectorate America Corporation (see www.inspectorate.com) is a global provider

of inspection, analyses and testing services for a wide range of commodity sectors

and consumer goods. The PRA Metallurgical Division has served the mining

industry for over 15 years and, along with its iPL Analytical Division, is located in a

modern 30,000 sq. ft. research facility situated in Richmond BC, Canada. Both

PRA and iPL are wholly owned subsidiaries of Inspectorate America Corporation.

The scope of work as outlined in the RFP is quite specific, and the following task

descriptions will reflect 11 items of interest that were listed, assuming that 900 kg

of representative sample will consist of six level composites of differing grades. A

separate environmental test program will require generation of sufficient amounts

of representative process tailings. Coordination of details and close supervision of

all work through the client’s consultant is to be expected.

Cost estimates for each task are broken down in labor and assay components and

tabulated in Appendix A. Corporate information including recent project histories

are provided in Appendix B.



2.0 SCOPE OF SERVICES

2.1 SAMPLE PREPARATION

Arrival of samples will trigger project initiation and in-depth confirmation against

client provided lists. Next, a set of detailed instructions are issued for sorting into

sub-lots (A, B, C, D, E, F) of ~150kg each, to be air-dried if needed and separately

jaw-crushed at a 1½” setting. Riffle splitting will produce representative halves

one of which is archived. The other split is crushed to ½”, halved again with only

one split (¼ of original sub-lot) further reduced to 10-mesh, prior to splitting into

2kg test charges and head assay aliquots.

Included in this group of tasks are careful labeling and standard storage of unused

samples for future retrieval, the eventual preparation of a coarse-crushed Main

Blends and shipping of one each of representative splits for Rod Mill Index work

(Hazen, Colorado) and optical Mineralogy (CTI, New Orleans).

2.2 HEAD ASSAYS

Aliquots of each of the six 10-mesh level composites will be assayed by standard

Inductively Coupled Plasma with Mass Spectroscopy (ICP-MS) following multi-acid

digestion, augmented by quantitative determinations of Au and Ag by fire assay,

As by hydride generation and Hg by cold vapor attachments. Based on the

findings, blending of the Main Composite will follow client-input and a confirmatory

head assay will be required.

A mineralogical determination of mineral associations, impurities as well as pay

metal occurrences can be conducted by optical means at Inspectorate’s CTI

Laboratory in New Orleans, or any alternate provider if preferred by the client. As

soon as the head assays are verified by the client, testing on the Main Composite

can proceed as follows.



2.3 HEAP LEACH EXPLORATION

The RFP specified a 3-day coarse bottle-roll test on a 1½”-crushed Main Blend

with a residue screen size assay to evaluate the top-size for a single column leach

simulation test of 9-weeks duration. It is recommended that whole-ore leaching be

first conducted on a 5kg charge with 0.5g/L NaCN, pH 10.5 at ~40% solids and a

low RPM tumbling rate to avoid attrition. Kinetic solution samples will be taken at

the 2, 8, 24, 48 and 72-hour marks. The bottle roll test is ended with filtration and

displacement washing of the leach residues, which are then screened into 5 size

fractions for complete pulverization and assaying for Au and Ag only.

Whilst levels of values in coarser particle sizes of samples tend to be variable,

rational evaluation criteria for the 3-day bottle roll leach tests will need to be

defined. Particle size targets and major conditions for short column testing are

then derived. Should bottle roll results be discouraging enough, the budget for

column leaching better be used for remedial heap-leach improvement ideas as will

be discussed with the client. For now it is assumed that column leaching at ½” top

size is possible and a separate estimate (Table 2) is provided on that basis.

The appropriate amount of Main Blend material at the selected size is prepared for

loading into a suitable PVC column without agglomeration. A calculated dosage of

lime would have been added to the feed charge, and tap water is circulated over

the first few days to ensure that the target application and percolation rates, and a

safe pH level for cyanide leaching are attainable. A carbon adsorption column is

inserted on-line, and cyanide leaching is started with daily monitoring and barren

solution adjustments prior to recycling to the top of the column. Regardless of the

indicated levels of extractions, the column test will be closed out in the 9th week of

operation, with a tap water displacement rinse of two bed volumes. Inventories

are then taken for mass balance calculation and evaluation of standard heap leach

potentials.



2.4 FLOTATION TESTS

Main Blend 10-mesh test-charges will be ground at 65% solids in a laboratory SS

rod mill, at different durations to achieve three different grind sizes for bulk sulfide

flotation testing using reagents and conditions that are to be discussed. Tests will

include a rougher and scavenger stage, followed by a regrind and cleaning in 2

mechanical stages. Selected products (maximum of three) are to be tested for

baseline cyanide leaching. Au and Ag will be followed, with S and ICP checks

where needed only.

2.5 GRAVITY TESTS

Centrifugal gravity concentration tests will indicate the recoverable free gold levels

at various primary grinds mainly, whereas reliable grade and yield estimates could

come from larger scale plant simulations only. Given the feeble amounts of bowl

concentrates recovered, an indication of the cleaning potential by hand panning is

normally preferred. The RFP specifies cleaning by MMS using the C800 Mozley

Mineral Separator, which is more sophisticated than the Super-panner available at

PRA. Other options such as mini-tabling or hydro-separation at no additional cost

can be discussed if the project is awarded to PRA.

In total four gravity tests are planned to cover 3 grind sizes and one alternative

mass pull strategy for grade-recovery sensitivity evaluation. Cleaner concentrates

will be assayed to extinction; only Au and Ag will be followed.

2.6 AGITATED CYANIDE LEACH

Standard 72-hour leach tests will be conducted at 3 grind sizes, 2 NaCN control

levels, and whole-ore at 2 pulp densities; room for extra tests to assess the kinetic

effects of oxygen-sparging, flotation product leaching or infill confirmation should

be provided. Solution samples are taken at 2, 4, 8, 24 and 48-hour marks; Au and

Ag will be followed, whilst reagent consumptions and dO2 levels are tracked.



2.7 ALTERNATIVE LEACH TESTS

Where indicated, measures to enhance the leach performance by pre-aeration,

lead nitrate or carbon addition, and common alternatives may need to be tested.

The RFP specifies that room for 4 CIL tests should be budgeted for using nominal

carbon-in-leach test conditions, as derived from standard whole-ore leach results.

2.8 THICKENING AND FILTRATION

Ground head and leach residue pulps need to be tested for flocculation, settling

and vacuum filtration characteristics for plant design purposes. The screening of

common Flocculants will be tested in small graduated cylinders on leach residues.

Two settling tests are then conducted in raked 2-L cylinders and the underflow is

subjected to a standard leaf filtration test.

2.9 GRINDABILITY TESTS

A 15kg split of feed crushed to ½” will be sent out for determination of the Bond

rod-mill work index (RMI) at the Hazen Research Institute in Denver, Colorado. A

10kg split of the same feed is crushed to 6-mesh and will be tested in-house for

the Bond ball-mill work index (BMI) at a standard 150-mesh closing screen size. A

series of test grinds on 2kg charges for 3 selected durations in the dedicated

laboratory mill will allow reliable achievement of target sizes for the test program.

2.10 DETOXIFICATION TEST

Three SO2/air ratios for cyanide destruction will be tested in batch mode on fresh

undiluted leach pulp from a 25kg bulk leach test to confirm optimized conditions.

The product solutions are assayed for free cyanide only at intermediate intervals,

and the final barren product will be analyzed by ICP and CN-speciation. Details of

testing will be defined in consultation with the client.



2.11 BULK PRODUCTION LEACH TESTS

Tentatively, 100kg of leach residue and 60L of supernatant solution are needed for

environmental testing elsewhere. Bulk leach tests in !25kg feed batches will be

used to confirm bench leach performances and to generate the leached pulp for

the detoxification test as well. Final products of each bulk test will be assayed as

requested, including Acid-Base Accounting (ABA) on two selected products to

evaluate prospective fresh plant tailings.

2.12 SUMMARY OVERVIEW

The work that is outlined in items 2.1 to 2.11 above will be coordinated by PRA in

conjunction with the client and their representative (Bolu Consulting) and could be

modified to suit any as yet unforeseen budgetary or technical constraint. Results

of testing will be presented as drafts after cursory internal reviews for QA/QC

purposes, including comments and observations pertaining to follow-up testing. A

set of detailed draft procedures will be issued for pre-approval of each test and an

ongoing dialogue about the interpretation of results is welcomed.

Budget estimates for the likely body of work have been organized under different

headings in Table 1 of Appendix A. With several uncertainties yet to be resolved

for a 9-week column leach campaign, a separate quotation is provided in Table 2.

Essentially all items listed in the RFP have been considered in this proposal, whilst

the limited body of data available at PRA and on the internet has been reviewed.

Rather elevated total sulfur and zinc levels of 7.7% and 4.1%, respectively, were

noted in the composite tested. This would suggest that the flotation option might

be a viable solution that could produce sulfide-depleted barren tailings suitable for

disposal, whilst shrinking the required cyanide leach circuit to a fraction of the size

needed for whole-ore leaching.



3.0 COSTS AND INVOICING

Except where noted above, the test work will be performed at PRA at standard

rates, and all cost estimates are based on recently updated Inspectorate and sub-

contractor fees. The detailed breakdown into tasks and associated costs for major

metallurgical categories is shown in Tables 1 and 2 of Appendix A.

Charges for labor and equipment maintenance versus fees and analyses are given

as a separate cost items. Applied labor rates for sample preparation are $50/h,

rising to $85/h for testing by technicians, and $150/h for project management that

includes supervision and reporting. The program is priced without overtime

charges and assumes work will be performed during regular scheduled hours of

8:30 am to 4:30 pm, Monday to Friday. Technical instruction and supervision,

data tabulation and entry, as well as client liaison is charged at 15% of the total

laboratory estimate.

A report including an executive summary, introduction, procedures, results,

conclusions and recommendations, along with all detailed test results will be

provided for work performed through PRA. An electronic copy and three bound

hard copies of the final report will be issued within 3 weeks of the conclusion of the

test program. Interim progress reporting which will include detailed spreadsheets

of the results as generated. Interim and final reporting is charged at 10% of the

total laboratory charges, as outlined in Tables 1 and 2 of Appendix A.

Shipping charges and storage charges are not yet included. There are no storage

fees during the test program (except when cold storage is requested). Following

completion of the project, applicable storage charges for less than 1 tonne are

based on a sliding scale by weight. Nominal bulk storage charges (>1 tonne) are

applied at $500/tonne/month. Alternatively, any residual material can be disposed

of (at cost) or returned to the client.



An advance payment of $15,000 is required to initiate the project. Outstanding

balances will be invoiced net 30 days along with actual services provided during

the billing period, as detailed on each invoice.

The cost estimates are in Canadian dollars and remain valid for 6 months from the

date of this proposal. Totals quoted as outlined in Tables 1 and 2 of Appendix A

are $43,000 + $12,230 = CA$55,230, not including any applicable taxes. The

number of tests, and methods proposed can be modified to suite the client’s

technical and budgetary requirements following the award of the project, based on

detailed task cost schedules as provided.



4.0 SCHEDULE

Sample preparation can begin as soon as the sample has been received with

prepayment. It is anticipated that the sample preparation, head analyses, size by

size analysis and mineralogy could be completed in two weeks.

The remaining schedule will depend on the total of number of tests to be

performed and on the sequencing requested. Bench scale work including gravity

separation, agitated leaching and rougher flotation can be completed with data

compilation within five weeks. Column leaching would only start towards the end

of that period and extend for another 10-weeks for completion.

The main test schedule as outlined in Table 1 would require seven weeks to

complete and, if column leaching is by then initiated, an overall delivery time of at

most 16 weeks is estimated. The PRA schedule can be modified (or fast-tracked

at additional cost) to suit the client’s requirements. Should PRA be awarded the

test program, details of testing, scheduling and task executions shall be refined

and discussed in-depth with the client.



5.0 KEY PERSONNEL

The staff members involved with the project have experience with all mineral

process and assay procedures proposed. The key project personnel and client

contacts are as follows:

! Technical Manager: Mr. Gie Tan, PhD. Is a Senior Metallurgist who has

been with PRA since 1998. He will act as Project Manager on behalf of

PRA and would be the primary client contact. He will coordinate the task

requirements and interpret the consistency and rational of all findings.

! General Manager: Alice Shi, PhD. Ms. Shi has been with PRA for 7 years

and will assist Dr. Tan in providing technical instruction, as well as with the

scheduling and logistic coordination of the program.

! Laboratory Manager: Boja Grcic, BSc. Ms. Grcic has been with PRA for

over 10 years and oversees the detailed execution of the work. She will

supervise and direct a dedicated team of well-trained technicians and

review the correct entry of all laboratory data.

Appendix XII

Cyanidation Test Reports

Client: Metal Mountain Resources Inc. Date:Test: C2 Project: 0905808

Sample: Master Comp.

CYANIDATION TEST REPORT

19-Dec-09


Objective: Baseline bottle roll cyanide leach in 1 g/L NaCN at a target grind of 80% passing 200 microns

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,024 g - repulped to 40% solidsSolution: 3,000 g - adjusted to and maintained pH 10.5

Solids: 40 % - adjusted to and maintained at 1.0g/L NaCNGrind Size - P80: 190 µm - sampled at 2,7,24,32,48 and 56 hours

Initial NaCN: 1.0 g/L - test ended after 72 hoursTarget pH: 10.5 - filtered and displacement washed with hot cyanide solut

Test Duration: 72 hours followed by two hot water displacement washes- solution and solids assayed for Au, Ag content

HEAD GRADEAu Ag

Calculated Total: 12 6 g/t 65 3 g/tCalculated Total: 12.6 g/t 65.3 g/tMeasured Total: 13.3 g/t 61.8 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 SlurryWeight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg)0 1.00 3.00 0.60 7.8 10.8 5,024 3,0002 0.80 0.61 10.7 5,048 3,024 30 4.03 12.2

Solution

, ,4 0.90 0.30 10.7 5,048 3,024 30 4.64 14.27 1.00 10.7 7.1 5,046 3,022 30 5.26 16.2

24 0.82 0.55 0.10 10.3 10.7 7.8 5,032 3,008 30 5.87 18.132 0.90 0.30 11.0 5,042 3,018 30 6.02 18.848 0.80 0.61 0.10 10.3 10.6 8.0 5,024 3,000 30 6.38 19.956 0.92 0.24 11.0 5,044 3,020 30 6.54 20.872 0.82 10.7 8.2 5,034 3,010 6.97 22.1

Total 5.61 0.80

SOLIDSSOLIDS

TimeWeight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)72 2,024 1.71 3.46 46.8 94.7

CYANIDATION RESULTS

Time Reducing Power

Total Residue

Distribution Reagent ConsumptionAu Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)2 47.6 17.3 0.297 63.3 20.2 0.44

24 70.8 23.2 0.7132 73.5 24.5 0.8648 77.9 25.8 1.1656 81.1 27.172 86.5 28.3 1.55 0.40

Residue 13.5 71.7

1.28265

Residue 13.5 71.7Total 100.0 100.0




19-Dec-09






Initial NaCN: 1.0 g/L - test ended after 72 hoursTarget pH: 10.5 - filtered and displacement washed with hot cyanide soluti


HEAD GRADEAu Ag


LEACH TEST DATA



Solution

, ,4 0.90 0.30 0.05 10.5 10.7 5,022 3,014 30 4.74 14.47 1.00 10.7 7.6 5,024 3,016 30 5.36 16.5

24 0.80 0.60 0.10 10.3 10.7 7.9 5,006 2,998 30 5.97 18.432 0.90 0.30 11.0 5,006 2,998 30 6.23 19.348 0.80 0.60 0.10 10.3 10.6 7.7 5,006 2,998 30 6.58 20.656 0.90 0.30 11.0 5,004 2,996 30 6.56 20.772 0.82 10.7 8.0 5,000 2,992 7.06 22.3

Total 5.73 0.73

SOLIDSSOLIDS

TimeWeight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)72 2,008 0.91 1.83 46.6 93.6

CYANIDATION RESULTS

Time Reducing Power

Total Residue


(hours) (%) (%) (kg/t) (kg/t) (mL)2 51.6 16.6 0.307 68.3 20.1 0.45

24 76.2 24.0 0.7632 80.1 25.2 0.9148 85.2 26.9 1.2156 85.7 28.672 92.4 30.3 1.63 0.36

Residue 7.6 69.7

1.36290

Residue 7.6 69.7Total 100.0 100.0




19-Dec-09






Initial NaCN: 1.0 g/L - test ended after 72 hoursTarget pH: 10.5 - filtered and displacement washed with hot cyanide solut


HEAD GRADEAu Ag


LEACH TEST DATA



Solution

, ,4 0.88 0.36 0.10 10.5 10.7 4,998 2,988 30 5.11 15.47 1.00 10.7 7.4 5,000 2,990 30 5.73 17.5

24 0.80 0.60 0.10 10.2 10.6 7.8 4,988 2,978 30 6.34 19.432 0.90 0.30 11.0 5,004 2,994 30 6.59 20.448 0.84 0.48 0.10 10.1 10.7 7.6 4,974 2,964 30 6.95 21.556 0.85 0.45 11.0 4,998 2,988 30 7.10 22.372 0.80 10.6 7.8 4,982 2,972 7.46 23.4

Total 5.85 0.84

SOLIDSSOLIDS

TimeWeight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)72 2,010 0.48 0.96 43.4 87.2

CYANIDATION RESULTS

Time Reducing Power

Total Residue


(hours) (%) (%) (kg/t) (kg/t) (mL)2 55.2 16.5 0.337 71.6 20.7 0.51

24 79.5 25.3 0.8132 83.8 27.1 0.9648 88.0 29.9 1.2156 91.4 31.872 96.0 33.6 1.73 0.42

Residue 4.0 66.4

1.42360

Residue 4.0 66.4Total 100.0 100.0

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