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

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

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.

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.

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.

Whole ore cyanide leach tests were conducted using conventional process conditions as widely practiced in the industry.

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.

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.

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.

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.

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.