The mineralization at El Domo shares most of the features of a VMS deposit (Franklin et al., 2005; Franklin et al., 1981; Large, 1992; Large et al., 2001; Lydon, 1996; Lydon, 1988a; Lydon, 1988b). VMS deposits are major sources of Zn, Cu, Pb, Ag, and Au, and can contain trace metals such as Co, Sn, Se, In, Bi, Te, Tl, Ga, and Ge.

Pratt (2008) was the first to document and describe a Kuroko-type VMS environment on the Project concessions. He established a lithostratigraphy for the Las Naves/El Domo area in which massive sulphide mineralization rests on a footwall sequence of rhyolite and dacitic autobreccias. He divided the sulphide mineralization into five types:
- Massive sulphides with indistinct texture. In some places, a fragmental texture can be seen within the sulphides, suggesting that they may be formed by the replacement of lapilli tuff.
- Sulphide-altered lapilli tuffs and peperites.
- Transported sulphide fragments within polymictic lapilli tuffs.
- Sulphide “pseudo”-fragments within polymictic lapilli tuffs. Rare, thinly laminated siliceous chert with banded sulphides.

The mineralised zone at El Domo is an intact, upright and only mildly disturbed Kuroko-type VMS deposit. As such, it displays the characteristic zoning of the model type from the underlying feeder pipe area through vertical and lateral variations upward to the abrupt termination of the massive sulphides against the characteristic hanging wall grainstone marker defined by Franklin et al. (2005). Over time, the evolution of the hydrothermal mineralizing system and the growth of the mineralised deposit account for the spectrum of mineralization types distinguished by Pratt.

Sphalerite, chalcopyrite, and pyrite are the principal sulphides in the mineralised rocks from the Curipamba prospect. Galena is less common, and tennantite/tetrahedrite and covellite are minor phases. Gold was identified within sphalerite + galena + barite mineralization, where it occurs as minute (5 µm to 50 µm) inclusions in sphalerite. The colloform banded sphalerite also contains an abundance of large, partly dissolved inclusions of skeletal galena. Careful microscopic examination revealed that gold was introduced to sphalerite via fractures with late chalcopyrite. Minute gold also occurs on the rim of some galena and is intergrown with chalcopyrite. The galena is partly replaced by tennantite, and it is rimmed and crosscut by chalcopyrite veinlets. Two small grains of gold were also identified in a late carbonate veinlet that crosscuts the sphalerite. The sphalerite is a pure zinc end-member with little or no iron content. In a number of samples, sphalerite is colloform banded, and just as some pyrite, often has framboidal texture. Textural evidence suggests that galena was largely contemporaneous with sphalerite, and both post-dated the pyrite. Tennantite and tetrahedrite represent a relatively minor phase and both crystallized at the expense of galena and less commonly, pyrite. Chalcopyrite was the last sulphide to crystallize in the polymetallic assemblage. In some samples, fragmented pyrite and sphalerite are “flooded” and partly replaced by massive chalcopyrite. Locally, chalcopyrite is stained to an unusual purple/blue colour. Microprobe analysis showed that in these domains the chalcopyrite has an unusual chemistry, and contains 2.2% to 3.7% bromine (by weight). Galena occurs as a skeletal inclusion in chalcopyrite and as replacement after pyrite. It contains inclusions of, and can be partly replaced by, tennantite and tetrahedrite. Covellite and chalcocyanite occur within sediments. Covellite forms a rim on detrital sphalerite and some pyrite, and chalcocyanite (anhydrous Cu-sulphate) is disseminated through the matrix. Barite is the principal gangue mineral (Schandl, 2009).

Mineralization at El Domo is broadly zoned with an upper “cap” of barite, enriched variably in silica sphalerite, galena, and gold. This cap is underlain by a massive sulphide zone with local zoning of zinc-rich mineralization along the hanging wall contact and a copper-rich base. This zonation, however, is not apparent throughout the massive sulphide zone. Zinc-rich mineralization consists of low iron sphalerite, some sulphosalts, barite, and pyrite. Copper mineralization is characterized by chalcopyrite and abundant pyrite. The base of the sulphide section is typically strongly silicified, with semi-massive pyrite and chalcopyrite as disseminations and stringer veins.

Mineralization in the grainstone shows evidence of at least two mineralization events. The sulphides include breccias which appear to have been caused by some form of collapse (possible anhydrite dissolution), while interstitial spaces were infilled by sphalerite and, in some cases, chalcopyrite. This brecciation replacement texture is common in many massive sulphide mounds as they grow by “displacement” or expansion. During this process, the core of a mound is constantly impregnated by high temperature hydrothermal fluid, displacing the lower temperature minerals outwards and leading to a constant zone refinement of the mineralization.

The sulphide and precious metal compositions have numerous unusual features, usually associated with high temperature systems that have achieved boiling just prior to their expulsion on or near the seafloor. The exceptionally high gold recorded in many of the upper zones in all of the occurrences, together with the anomalous antimony, arsenic, mercury, and bromine contents of some of the minerals, can only be achieved by this process, which enables exceptionally efficient gold precipitation. Gold is conserved in the vapour phase of a hydrothermal fluid, and thus may be deposited over a much wider area than the base metals. At the Project, gold is generally associated with baritic exhalite.

The Project is planned as a conventional open pit operation with haul trucks, hydraulic excavators and loaders. The ore will be transported from the pit to its destination by truck; ore material will be loaded preferentially by wheel loaders and waste material will be loaded preferentially by hydraulic excavators. Ore material will either be sent directly to the plant ROM pad or to the ore stockpile which then would to be rehandled and sent to the mill at a later time. At the stockpile, the different ore material types (High Zn, Mixed Zn/Cu, and High Cu) will be stockpiled separately to enable blending. The mill has a capacity of 666 kt/y, of which a maximum of 15% High Cu material is accepted, until there is no longer sufficient supply of the other ore materials for blending.

Waste material will be sent to one of four destinations: the overburden stockpile, the saprolite stockpile, the tailings storage facility, or the waste rock facility. Overburden will be sent to the overburden stockpile, saprolite will be sent to the saprolite stockpile, and other waste material will be divided as required to stabilise the overburden and saprolite stockpiles, or to build the tailings storage facility. Additional waste material will be sent to the waste storage facility.

The mine will operate year-round, seven (7) days per week, twenty-four (24) hours a day with two (2) 12-hour shifts a day. Fifteen (15) days of weather delays are considered in the mine plan.

Total tonnage to be mined from this pit is 61.82 Mt, which will be mined over a 10-year mine life with an additional 1.5 years of pre-production.

There will be three (3) ore stockpiles, one for each of the ore types. The stockpiles will be located near the entrance of the pit along the road to the process plant, near or on top of the saprolite stockpile, depending on the year.

The mine plan was based on feeding the mill 666 kt ore per year. High Copper material was limited to 15% of the total material sent to the mill until the end of the mine life. The mine plan was estimated on a monthly basis for the pre production period and the first year of production. The remaining nine (9) years of production were generated on a yearly basis.

Five (5) phases in order to mine the Curipamba deposit and ensure access to the mining faces throughout the life of mine. Other considerations were also taken into account, such as ensuring enough andesite could be extracted for the tailings dam construction. In addition, the main ramp was designed to act as both a ramp and a geotechnical berm, to reduce waste. Benches were designed to be either 5 m or 10 m in ore, depending on the area, to maximise ore extraction, while they were designed to be 10 m in waste to reduce the amount of waste mined.

During pre-production, waste material is required to build the tailings storage facility prior to the start of production. As such, the pre-production period has a higher mining rate than the production period. During the pre-production period, the total material movement was capped at 1 Mt per month while it was capped at 6.5 Mtpy during production to ensure a feasible schedule considering the difficulty of mining an area with steep terrain. Additionally, the vertical advance rate was limited during preproduction to ensure the schedule is feasible.

The proportion of High Cu material sent to the mill was limited to 15% of the total mill feed in any given year to optimise the mill productivity and metal recoveries, as determined through process design. To ensure this constraint was met, the different ore types were stockpiled and rehandled as required. At the end of the open pit mine life, when there is no longer sufficient High Zn and Mixed Zn/Cu material to supply the mill at 0.67 Mtpy, a larger proportion of High Cu material was allowed.

Recommended inter-ramp angles vary between 46.1° and 54°, based on wall orientation, overall wall height, geotechnical domain, and controls on slope stability. Inter-ramp slope heights are limited to 70 m, after which a geotechnical berm (or ramp) with a minimum width of 12 m or 14 m, depending on the area, is required.

Benches were designed to be either 5 m or 10 m in ore, depending on the area, to maximise ore extraction, while they were designed to be 10 m in waste to reduce the amount of waste mined.

PRIMARY AND SECONDARY CRUSHING
Run-of-mine (ROM) ore is dumped into a feed bin by the front-end loader (FEL). The mining operation should target minimising the quantity of oversize material that cannot pass through the 600 mm vibrating grizzly.

The ore crushing circuit uses skid built crushing equipment and skid-built screen and conveyors as a cost-efficient option for this size of the concentrator plant.

Primary crushing is completed using a mobile jaw crushing skid. A vibrating pan feeder delivers ore to the vibrating grizzly. Oversize material gravity feeds to the primary 110 kW 1050 x 800 mm CJ411 or equivalent jaw crusher, assisted with a rock breaker used for the larger rocks. The jaw crusher product discharges onto the belt conveyor where it is combined with grizzly undersize material. The material is then conveyed to secondary crushing which operates in a closed circuit.

Primary crushed ore combines with screen oversize in a surge bin allowing for choke feed to the cone crusher. Feed to the 220 kW CH440 or equivalent cone crusher is controlled via a vibrating pan feeder which discharges material onto a feed conveyor. Secondary crushed material is conveyed to a double deck vibrating screen for classification.

Over-size material passes over the screen and is conveyed to the secondary crusher surge bin. Screen undersize of 18 mm D80 is the final product conveyed to a radial stacker which stockpiles material within the crushing area. The stockpile provides a buffer for at least 8 hours of the concentrator plant operation.

Crushed ore is reclaimed from the stockpile by means of a FEL used to load the crushed ore hopper. Crushed ore is reclaimed by an apron feeder and fed to the grinding circuit using a ball mill feed conveyor. The crushed ore can also be directly fed onto the ball mill feed conveyor through the bypass hopper by means of the FEL. The radial stacker has the capability to feed into the crushed ore hopper directly.

GRINDING
Grinding is performed using a 1.55 MW VFD (Variable Frequency Drive) driven 13' x 19.5' grate discharge ball mill operating in closed-circuit with a cluster of hydrocyclones. Fresh ball mill feed (crushed ore from the stockpile) is combined with cyclone underflow in a mill feed chute. The grinding circuit reduces feed material to a product size P80 of 125 µm which is the target grind size for the bulk flotation process. Process water is added to the circuit to maintain solids density in the mill. Mill discharge is pumped to the cyclone cluster for classification. Cyclone underflow material returns to the mill feed chute for further grinding. Cyclone overflow is fed to a linear trash screen prior to bulk flotation conditioning.

The Curipamba concentrator operation was designed to process 666,000 tonnes per year of ore into copper, lead and zinc concentrates.

The process consists of the following unit operations:
- Primary and Secondary Crushing;
- Grinding and Classification;
- Bulk Flotation and Regrind;
- Copper and Lead Flotation;
- Zinc Flotation;
- Copper Concentrate Dewatering;
- Lead Concentrate Dewatering;
- Zinc Concentrate Dewatering;
- Concentrate Storage and Loadout;
- Reagent Preparation and Distribution;
- Tailings Thickening and Handling;
- Water Services;
- Air Services; and
- Process Consumables.

BULK FLOTATION AND REGRIND
The objective of the bulk flotation circuit is to obtain maximum recovery of metal sulphides and reject gangue material prior to separation in the subsequent copper, lead , and zinc flotation circuits. Cyclone overflow (post trash screen) from the grinding and classification cir ........