Gold mineralization in the Santa Rosa Gold Project has characteristics in common with mesothermal or orogenic and intrusion-hosted gold veins. Mineralization of these types consists commonly of quartz and quartz-carbonate veins located in moderately to steeply dipping, brittle-ductile shear zones and locally in shallow-dipping extension fractures. Veins commonly extend along strike and down dip over very significant distances and occur alone or, typically, in complex vein networks and shear zones. Vein minerals are mostly quartz and carbonates with minor native gold, pyrite, and base metals sulfides. Veins are usually massive or ribbon-textured, but vein breccias and drusy, crystalline quartz can also occur. Wall-rock alteration is zoned and consists of carbonate (often ankerite), sericite, and pyrite.

Hypogene gold mineralization within the Santa Rosa Gold Project is generally associated with the shear zones developed in homogeneous diorite country rock, with higher grades occurring in the associated sulfide-mineralized quartz veins or as steep high-grade quartz-sulfide veins. There are also related saprolitic gold deposits and colluvial gold deposits, both of which have been mined by artisanal miners underground and in hydraulically mined areas known locally as “baticiones.” The shear zones and veins are best exposed in adits and baticiones.

In the San Ramon deposit shear zone, mineralization occurs in fractured, brecciated, and ductily deformed rock. The lowest levels of mineralization (~0.035g Au/t to ~0.1g Au/t) consist of very weak to moderate ductile shearing that locally contains quartz veinlets and/or pyrite. At slightly higher grades (~0.1 to ~0.6 g Au/t), mineralization is characterized by moderate to strong ductile deformation that contains scattered quartz veinlets and sulfides, although the overall quantity of veinlets and sulfides is low. Some sericite, weak to moderate brittle overprinting, and quartz veins greater than 12 mm in thickness that are mostly, if not completely, barren of sulfides, may also be present. The boundary between the two grade ranges is likely gradational, and there are numerous instances where quartz veins or pyrite are not apparent in either. Shear zone intercepts containing these low grades of mineralization may be up to 80 m wide; the maximum true width of the mineralized shear zone is about 60 m.

Higher-grade mineralization, >0.6 g Au/t, within the San Ramon deposit shear zone is characterized by strong shearing, more intense sulfide mineralization, and quartz veins. At relatively lower grades within this higher-grade mineralization (~0.6 to 5.0 g Au/t), strong ductile shearing with abundant sulfide minerals and sericite is almost ubiquitous; quartz veins may or may not be present. Brittle overprinting is moderate to strong and may be genetically related to the high-grade mineralization. At grades in excess of ~5.0 g Au/t, massive and coarse-grained pyrite, pyrite stringers, medium-grained sphalerite, fine-grained galena, and traces of chalcopyrite are present in quartz veins and quartz vein fragments, although some massive pyrite and pyrite stringers occur independently of quartz veins. In the highest-grade intercepts (>50 g Au/t), relatively thick (=2 cm) massive pyrite veins are intermixed with quartz veins that contain coarsegrained and massive pyrite, coarse-grained sphalerite, fine-grained galena, and traces of chalcopyrite.

The mining method selected is Mechanized Shrinkage with Delayed Fill (“MSDF”). The method is similar to mechanized cut and fill but uses breast blasting of the back between lifts. Then, rather than mucking of ore and backfilling immediately, the MSDF method leaves the ore in place in the stope, only removing enough material from the stope to remove swell (similar to shrinkage stoping but removing the swell from the top instead of the bottom). Access to the stope is provided by establishing an attack ramp. For each
lift, the attack ramp back is blasted to establish access for the subsequent lift. Support is placed in the stope as needed during the drilling and blasting cycle of each lift. Once enough lifts have been drilled and blasted, the ore will be mucked out completely. The last cycle of mining for MSDF stopes is backfilling of the stope from the bottom up, progressively backfilling the attack ramp as well.

Backfill is to be used as required, but not all stopes will be fully backfilled. The backfill will consist of filtered tailings and development waste rock. Tailings will be placed in a single lift, and once the lift of tailings is placed, then a lift of waste rock from development will be placed on top of the tailings. The amount of waste placed on top of the tailings will be sufficient to allow equipment to drive on top of the backfill. It has been assumed that 60% tailings and 40% development waste rock will be a suitable mixture of tailings to waste to stabilize the fill.

The process plant design incorporates the following unit process operations:
- Primary crushing with a single toggle jaw crusher (1,100 x 700 mm) to produce a crushed product size of 80% passing (P80) 100 - 120 mm;
- A crushed ore surge bin (30 m3) with a nominal capacity of 1 hour process plant feed of 45 t;
- Single stage SAG mill (5.0 diameter x 3.5 m EGL 1,200 kW) in closed circuit with cyclones to produce a P80 grind size of 125 µm;
- Rougher scavenger flotation (6 x 8 m3 conventional cells) to produce a sulphides/gold concentrate;
- Tower mill (150 kW) for regrind of the concentrate to a P80 grind size of 15 - 20 µm;
- Pre-leach thickener (16 m dia) to minimise carbon in leach (CIL) tankage and reduce overall reagent consumption;
- A hybrid CIL circuit incorporating one leach tank and six adsorption tanks (430 m3 each) with 48 h total residence time;
- A 2 tonne AARL elution circuit with electrowinning and smelting to produce doré bars;
- Cyanide destruction using the SO2/Air process (120 min retention); and
- Tailings pressure filtration (445 m2 filter area) to 16% moisture content.

Ore is delivered by truck directly from the adjacent decline to a ROM stockpile. Sufficient area has been allocated to allow for ore blending and a total capacity of 3,000 tonnes (three days of plant feed). A front end loader (FEL) reclaims ore from the stockpile on the ROM pad and delivers ore to the ROM bin.

ROM ore enters the crushing circuit via a small ROM bin fed by the FEL.

A single stage semi-autogenous grinding (SAG) mill in closed circuit with hydrocyclones was selected to grind the ore and F80 of 100 mm to produce the target grind P80 size of 125 µm. The SAG mill is equipped with a 1,200 kW variable speed drive. The mill rotation speed range is between 60 – 80% of critical speed to handle varying ore competencies. In addition to standard charge weight and power control algorithms, the mill is equipped with noise monitoring to alarm low mill charge levels. This circuit provides stable and simple operation that is well understood and proven at many successful installations.

Cyclone overflow slurry gravitates to the flotation conditioning tank where the slurry is diluted to 35% w/w solids and flotation collector (PAX) will be added. The conditioned slurry gravitates to six conventional 8 m3 flotation cells arranged in a rougher – scavenger configuration with total residence time of 25 minutes. Flotation concentrate reports to the regrind circuit and flotation tails flow to the pre-leach thickener feed box.

A pre-leach thickener has been included ahead of the CIL circuit to allow for adequate slurry dilution ahead of flotation and provide high slurry density (50% solids) to the CIL circuit so both unit processes are optimised.

The leach characteristics identified during the metallurgical testwork program indicated that a 48 hour leach is required to achieve optimal gold extraction. To obtain this retention time a circuit configuration of one leach tank and six leach/adsorption tanks was selected. This is the minimum tankage required for acceptable stage efficiency and for achieving target solution grades.

Gold is electrowon from the pregnant solution within the security area of the goldroom.

After completion of the elution process, the barren carbon is transferred from the elution column to the carbon dewatering screen prior to entering the feed hopper of the horizontal carbon regeneration kiln.