At Loma Larga, like in most typical high sulphidation epithermal systems, alteration is characterized by multiphase injections of hydrothermal fluids strongly controlled by both structure and stratigraphy. The alteration-mineralizing event is characterized by an early alteration phase caused by a strong inflow of volatile, acidic fluids which cooled progressively and were neutralized by their reaction with country rock, leading to the formation of silicified layers surrounded by alteration halos of clay minerals while the sulphides and gangue minerals associated with the mineralization were deposited by later fluids inside the silicified bodies (IAMGOLD, 2008).

The majority of the limited amount of outcrop exposed at Loma Larga exhibits silica alteration, due to its resistance to weathering. In epithermal environments the silica alteration displays evidence of hot acidic leaching. Multiple types of silica alteration occur at Loma Larga, including vuggy, sugary, banded, fracture fill, and hydraulic-breccia (MacDonald, 2010).

Alteration can be seen to be structurally controlled as it typically occurs as silica ribs mimicking fault locations and orientations. The most significant alteration zone, host to the deposit, is coincident with the north-trending Rio Falso fault, extending for over eight kilometres northsouth, along the eastern edge of the collapsed caldera. This long, linear zone contains multiple large pods of silica alteration ranging up to two kilometres in east to west width. The location of the Rio Falso fault suggests that it was coeval with or postdates the caldera collapse (MacDonald, 2010).

The silica alteration is surrounded by varying widths of a halo of argillic alteration, grading from higher to lower temperature mineral assemblages including pyrophyllite, alunite, dickite, kaolinite, illite, and smectite.

The high sulphidation epithermal gold-copper-silver mineralization in the Loma Larga deposit is also stratigraphically controlled as it occurs at lithological contacts between Quimsacocha Formation andesitic lavas and tuffs, and reaches greater thickness in the more permeable tuffs. The deposit is a flat lying to gently western dipping (less than ten degrees), north-south striking, cigar shaped body, which has a strike length of approximately 1,600 m north-south by 120 m to 400 m east-west and up to 60 m thick, beginning approximately 120 m below surface. It also dips slightly to the north, such that the mineralized zone is closer to surface at the south end. Resources are defined as a smaller, higher-grade subset within this mineralization.

Mineralized zones are characterized by multiple brecciation and open-space filling events and sulphides such as pyrite, enargite, covellite, chalcopyrite, and luzonite or, at lower sulphidation states, tennantite and tetrahedrite. Higher grade intervals typically coincide with increased amounts of enargite, minor barite, and intense hydraulic brecciation that contains subrounded to rounded silicified fragments. Visible gold is rare. Gold mineralization is found, for the most part, in one of the following mineralogical assemblages: (a) vuggy silica plus fine grained pyrite and enargite; (b) massive pyrite, including a brilliant arsenical pyrite; or (c) vuggy silica with grey silica banding, sulphide space-filling and banded pyrite. Very fine grained pyrite is dominant in semi-massive to massive zones, and is interpreted to have formed earlier than coarser fracture and vug-filling pyrite (MacDonald, 2010).

The focus of mineralization occurs at approximately 3,610 m (± 30 m) elevation, where structural feeder zone(s) intersected a permeable tuff horizon that was acid leached. There is an upper barren silicic lithocap, locally indicative of steam heated alteration, which is typically barren of mineralization, although there is an outcrop exposure of this zone that contains minor, fine visible gold (MacDonald, 2010). Silica textures within the upper zone range from sugary, to two-phase, massive, vuggy, and laminated, while the main body centred at 3,610 m exhibits either massive or vuggy silica with intense brecciation in the core and pervasive veinlet and vug infilling alunite alteration.

Loma Larga will use two mining methods to extract the ore, longhole and drift and fill. Longhole stoping using paste backfill provides good productivity, high extraction, and stable back support. Drift and fill mining provides extraction of smaller areas unsuitable for longhole stoping.

The longhole stope design parameters are 15 m W x 15 m H x 30 m L. The stoping height was increased in places to adjust for the undulating deposit. Approximately 62% of the stopes are 15 m in height, while 31% are 20 m in height, and 7% are 25 m in height. For stopes greater than 15 m in height, the strike length was reduced to maintain an appropriate hydraulic radius, based on geotechnical considerations. In areas that are less than 15 m, drift and fill mining will be used to extract the ore using 5 m W x 5 m H drifts.

The process design includes:

• Primary jaw crusher,

• Single stage semi-autogenous grinding (SAG) mill and classification,

• Sequential flotation
- Copper rougher flotation, copper concentrate regrinding, and copper cleaner flotation,
- Pyrite rougher flotation, pyrite concentrate regrinding, and pyrite cleaner flotation,

• Concentrate thickening, filtering, and loading,

• Tailings thickening and filtering, and

• Tailings loadout.

The undersize from the grizzly and the discharge from the jaw crusher will fall onto the primary crusher discharge conveyor that transfers ore to a transfer conveyor. From the transfer conveyor the crushed ore is deposited into the crushed ore stockpile, which has a capacity of 13.6 kt, or approximately 4.5 days of production.

Three reclaim feeders will draw ore out of the stockpile and transfer it to the SAG mill feed conveyor. The feed conveyor transfers the ore into a SAG mill. Process water is also added to the SAG mill feed to produce a slurry density in the mill of approximately 75% solids by weight. The slurry discharges from the mill onto the SAG mill discharge screen. Underflow from the screen discharges into the primary cyclone feed pump box. From the pump box the primary cyclone feed pumps (one operating and one standby) pump the slurry to the primary cyclones. The cyclone underflow flows by gravity to the SAG mill feed for further grinding. The oversize from the SAG mill discharge screen is conveyed to a pebble crusher which is included to reduce the critical size particles that discharge from the SAG mill. The crushed pebbles are conveyed to the SAG mill feed chute. The overflow from the cyclones is the final product from the grinding circuit. It has a nominal particle size distribution of 80% passing (P80) 75 µm. The slurry flows across a trash screen prior to advancing to the copper flotation area.

The underflow from the trash screen flows by gravity into the copper flotation agitated conditioning tank. Milk of lime, sodium sulphite (Na2SO3), collectors (Aero 407 and Aerophine 3418A), and methyl isobutyl carbonol (MIBC) frother are added to the conditioning tank. The target pH is greater than 11.8.

The slurry overflows from the copper flotation conditioning tank into the copper rougher flotation circuit. The concentrate from the copper rougher flotation circuit is pumped to the copper regrind mill by the copper rougher concentrate pump. The regrind mill is an IsaMill that is designed to operate in open circuit. The target regrind size is P80 15 µm.

The ground slurry from the copper regrind mill discharges into the copper first cleaner flotation circuit. Additional milk of lime, methyl isobutyl carbine (MIBC), and 3418A are added to the circuit. The concentrate from the copper first cleaner flotation circuit is pumped to the copper second cleaner flotation circuit. The tailings from the copper first cleaner flotation circuit are recirculated to the copper flotation conditioning tank. The tailings from the copper second cleaner circuit are pumped to the feed of the copper first cleaner flotation circuit. The concentrate produced by the copper second cleaner flotation circuit is the final copper concentrate. The mass recovery to the final copper concentrate is approximately 0.8% at the average feed grade. The estimated concentrate grade is 30.0% Cu. For this Study, it is assumed that the recoveries and the concentrate grade of copper will remain constant. Correspondingly, the mass of concentrate and the gold, silver, and arsenic will change as the feed grades to the plant fluctuate. The average copper concentrate grade is 111 g/t Au, 1,577 g/t Ag, and 10% As.

The tailings from the copper rougher flotation circuit are the feed to the pyrite rougher flotation circuit. They are pumped to the pyrite flotation agitated conditioning tank. Sulphuric acid (H2SO4) and potassium amyl xanthate (PAX) are added to the pyrite flotation conditioning tank. The pH is modified to approximately 7.0.

The slurry overflows from the pyrite flotation conditioning tank into the pyrite rougher flotation circuit. Tailings from the pyrite rougher flotation circuit are the final tailings from the plant. Concentrate from the pyrite rougher flotation circuit is pumped to the pyrite regrind mill by the pyrite rougher concentrate pump. The regrind mill is an Isamill M500 225 kW mill. The target regrind size is P80 30 µm. The slurry from the pyrite regrind mill discharges to the pyrite cleaner flotation circuit. Additional sulphuric acid and PAX are added to the circuit. The tailings from the pyrite cleaner flotation circuit are recycled to the pyrite flotation conditioning tank.

The concentrate from the pyrite cleaner flotation circuit is the final pyrite concentrate. For this Study, it is assumed that the recoveries and the concentrate grades for gold will remain constant (i.e., 37 g/t Au). The silver grade and the mass of concentrate will change as the feed grade to the plant fluctuates. The average mass recovery is 9.8% and the average silver grade is 142 g/t.

The final copper concentrate is pumped to the copper concentrate feed agitated tank. From the tank the copper concentrate is pumped to the copper concentrate recessed plate and frame pressure filter. The filter cycles include filling the filter, applying pressure, and blowing air through the filter to achieve a final moisture content of less than 10% water. The filtrate from the filter press is returned to the process water tank for reuse. At the end of the cycle, the filtered copper concentrate drops onto copper concentrate feeder which transfers the copper concentrate onto the copper concentrate bin feed conveyor. The copper concentrate discharges into the copper feed bin. From the cone bottomed bin, the copper concentrate can be loaded into bulk bags for transport.

The final pyrite concentrate is pumped to the pyrite concentrate thickener. The underflow from the thickener is pumped by the pyrite thickener underflow pumps to the pyrite concentrate feed tank. The slurry density is estimated to be 56% solids by weight. From the tank the pyrite concentrate is pumped to the pyrite concentrate recessed plate and frame pressure filters. The filter cycles include filling, applying pressure, and blowing air through the filter to achieve a final moisture content of less than 10% water. At the end of the cycle, the filtered pyrite concentrate drops into the pyrite concentrate loadout bin. From the bin, the pyrite concentrate can be loaded into containers for transport.

The bags of copper concentrate and bins of pyrite concentrate can be stored in the covered concentrate storage area while awaiting trucks that will transport the bags to the port for storage and overseas shipment to smelters.