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.

The underground mine will be accessed by a 1.2 kilometres long (5 metres high by 5 metres wide) ramp into the deposit. The ramp will serve as the access to the mine for personnel and materials, the haulage of waste and ore, and for ventilation. Due to the high-grade nature of the ore body and the positive geotechnical conditions, the deposit will primarily be mined by the long-hole stoping method, with 20 metres wide, 25 metres high and 20 metres long stope sizes. Certain zones will utilize the drift and fill method where appropriate.

Initial daily ore production of 3,000 tpd is planned from primary and secondary stopes for the first four years, generating approximately 1,095,000 tonnes of ore annually. From year 5, daily average ore production of 3,400 tpd is planned to be achieved through plant optimization, generating 1,241,000 tonnes of ore annually. Ore will be trucked approximately 3.5 kilometres from the portal to the process facilities.

The process plant is designed to produce saleable separate copper and pyrite concentrates.

ROM ore from the underground mine is crushed on surface in a two-stage crushing (primary jaw crusher, secondary cone crusher) closed circuit to provide feed material suitable for milling in a single-stage ball mill. A 3,000-tonne live capacity stockpile provides an operating buffer between the crushing and milling circuits.

Material is withdrawn via tunnel reclaim from the stockpile and fed to the ball mill. The ball mill operates in closed circuit with a bank of hydro- cyclones. Milk of lime and low pH process water are added to the ball mill feed chute with the crushed ore to maintain a target mill pulp pH. Hydro-cyclone underflow material is returned to the grinding circuit. Hydro-cyclone overflow pulp reports to a grinding product thickener. Thickened underflow is conditioned with lime to raise the pH prior to copper rougher flotation in conventional tank cells.

Copper flotation feed is first conditioned with milk of lime and high pH process water in the copper rougher flotation conditioning tank. The conditioned slurry flows by gravity to the copper rougher feed box and into the first of four rougher flotation tank cells. Copper rougher tails are sent to the copper tails thickener. Copper rougher concentrate is reground in a vertically stirred media mill before copper cleaner flotation using two sequential stages of cleaning.

The first stage of copper cleaning consists of four conventional forced air flotation cells. Tails from the first cleaner cells flow to the copper first cleaner scavenger feed box and then to four conventional forced air flotation cells. Tails from the cleaner scavenger cells are directed to the copper tails thickener. Concentrate from the first cleaners is pumped to the second cleaners, whilst concentrate from the first cleaner scavengers is returned to the first cleaner feed box.

The second cleaner bank consists of three conventional forced air flotation cells. The second cleaner concentrate reports to final concentrate dewatering, and the second cleaner tails report to the first cleaner feed box.

Provision for the future addition of a bank of third cleaner cells and associated equipment has been allowed for in the process plant layout.

Final copper concentrate is thickened and filtered prior to being bagged for shipment on site.

The underflow from the copper tails thickener is pumped to the pyrite rougher flotation conditioning tank, where low pH process water is added. The conditioning tank overflows to the pyrite rougher feed box and discharges into the first of six pyrite rougher flotation tank cells.

Pyrite rougher concentrate is reground in a vertically stirred media mill before further cleaning in a single stage of flotation. The pyrite cleaning flotation circuit consists of six conventional forced air flotation cells. Tails from the pyrite rougher and first cleaner cells flow to the tailings thickener.

Concentrate from the first cleaners is final pyrite concentrate and reports to the pyrite concentrate thickener and pressure filters.

Provision in the future has been made for the installation of a bank of second cleaner cells and associated equipment.

Pyrite concentrate is dewatered and stored, in bulk, in the concentrate storage area prior to being shipped in containers.

Pyrite flotation tails are thickened in the tailings thickener. Thickener underflow is directed to the tailings filtration area for final dewatering. Filtered tails are hauled to either the FTSF for deposition or to the paste backfill plant at the mine as required.

Water recovered from the grinding circuit thickener, copper thickener and pressure filters are considered the source of high pH process water. Low pH water is recovered from the tailings
thickener and pyrite pressure filters.










Extensive metallurgical test work has demonstrated estimated overall gold, silver and copper recoveries into concentrate of 90%, 95%, and 96%, respectively.