The deposit model guiding exploration targeting is mesothermal quartz vein style gold mineralization. Dr. Noel White (2011) and other consultants (e.g. Telluris, 2009) have also described the Ollachea deposit as a member of the class of orogenic gold deposits, with the possibility of local syngenetic gold enrichment playing a role in the location of the mineral deposit. This variety of gold deposit can also go by the name slate belt gold deposit and can be both very large and very rich.

Mineralization has been traced continuously for 900 m along strike. Gold mineralization has also been encountered to the west of the Minapampa Zone in a zone on the south side of the Oscco Cachi River that is referred to as Concurayoc, located some 400 m west of Minapampa. The known mineralized zone at Ollachea, including both the Minapampa and Concurayoc Zones as well as the intervening 400 m of strike between the two mentioned deposits, is approximately 1,900 m long, up to 200 m thick and has been traced in places to over 400 m below surface and remains open along strike as well as at depth.

An extensive shear zone hosts the gold mineralized horizons (Figure 7-4). The shear zone is characterized by a well-developed slaty cleavage, with quartz-sulphide veins and veinlets, broadly concordant with the slaty cleavage. Quartz-sulphide veins and veinlets vary from a few millimetres to centimetres wide, up to a maximum of 40 cm, but do not always contain gold mineralization. The gold mineralization is hosted both within as well as along contacts of the quartz -sulphide veins and veinlets. The veins can be strongly boundinaged, resulting in the development of packages of irregularly mineralized veins and veinlets within discrete mineralized horizons, incumbent to the sheared slate package.

Gold mineralization is associated with a sulphide assemblage consisting predominantly of pyrrhotite with minor pyrite, arsenopyrite and traces of chalcopyrite. Coarsely crystalline arsenopyrite and free gold are frequently observed in close association with one another within the central Minapampa zone. The occurrence of coarse pyrite without other sulphides is often a counter-indicator of gold mineralization.

At Concurayoc, as opposed to Minapampa, only six discrete gold mineralized horizons have been delineated by drilling. The mineralized structures strike essentially southwestnortheast and dip towards the north. Characteristics of the gold mineralization are very similar to those observed at Minapampa with a sole exception being that the economically mineralized horizons are not as wide as some of those at Minapampa. Mineralization has been traced continuously for 700 m along strike and up to 400 m in depth. The mineralized horizons remain open-ended in strike as well as at depth.

The mining method selected for the FS was long hole open stoping (LHOS) with paste backfill, which can also be referred to as bench stoping with paste backfill. Extraction occurs along the orebody strike direction on a retreat basis.

Stopes will be accessed longitudinally (along strike) on each level by, one, two or three strike ore drives dependent on lode thickness. Open stope strike length is dictated by geotechnical considerations and varies with lode width.

The direction of mining for the deposit will be from the bottom up. As each mining level is completed, the next level will start using the backfilled stope void as the mining platform.

The main access to the mineralisation will be via a 1.2 km-long exploration access incline (1.5%) which has its portal in a valley on the north-eastern side of Cerro Joropiña and the Oscco Cachi River valley. The drive is currently being excavated and will be used for exploration drilling as well as providing the primary access and haulage drive for the planned mine. This portal (lower) will be the main mine access portal and is located above the process plant area at 2765 mRL.

An incline drive and a decline drive will be excavated at a grade of one in seven from the main exploration incline, located at approximately 2782 mRL, to access the eastern part of the mine. The decline drive will extend to 2550 mRL to service the deepest planned mining level at 2565 mRL. The incline drive will extend to a mining level at 2865 mRL.

The main exploration incline will be extended as an incline drive at a grade of one in seven and will be developed to meet a decline drive that will be developed simultaneously from a second (upper) portal at 3060 mRL. These drives when connected will provide a second means of egress, access to all the mineralisation in the western part of the mine, and early establishment of the primary ventilation system.

The mine is split into two main production areas, east and west, with the western part of the mine providing approximately 71% of the life of mine production tonnage. All mining is completed using a bottom up mining direction.

To maximise mine extraction, the eastern part of the mine will be split into multiple mining panels consisting of four levels that can be mined simultaneously. The lowest level of each of these mining panels requires an artificial sill pillar to be created using high strength paste backfill to allow the mineralisation located directly below to be completely extracted. The western part of the mine has also been split to minimise the impact from the life of mine production tail. A sill pillar level has been located on 2940 mRL.

Due to the non-visual nature of the orebody, grade control diamond drilling is planned on a minimum of a 15 m by 15 m grid. In the eastern part of the mine, this will be completed from dedicated hanging wall drives that will provide coverage for four production levels. The western part of the mine will be grade control drilled on each level from the main hanging wall access drive. Mineralised zones will be re-interpreted from the grade control program; ore drives will then be driven primarily on survey control and backed by face and wall channel sampling. An onsite laboratory is planned and is this has been designed to provide a 24 hour turnaround of samples.

Production from the eastern part of the mine will start on 2790 mRL and 2805 mRL for the western part of the mine. The primary ventilation system will be fully established prior to the start of stope production.

Stope size will be controlled by the nature of the lodes (dip and width variability) and interpreted geotechnical conditions. Stope sublevel spacing will be 15 m vertically floor to floor. Planned stope strike length is based on geotechnical interpretation and varies between 13 m and 23 m dependent on lode width. To control the stability of the longitudinal stopes and minimise dilution, the length of open voids can be altered based on local ground conditions.

Production drilling will be medium diameter (76 mm or 89 mm) down holes with some requirement for up-holes when mining below an artificial sill pillar. Up-holes will also be used where lodes pinch out and there is no requirement for development above. Stope blast initiation (void) will be via the use of drop raise slots as the distance from the floor of the top cut to the back of the bottom cut will be approximately 10 m vertically or 14 m on dip.

To minimise dilution, maintain stability and maximise open stope strike length, cable bolts will be installed in the hanging wall of the stopes. A dedicated cable-bolter (drill and install) is planned to complete this activity. Development ground support installation will be completed by development jumbos.

Stopes will be backfilled using paste derived from mine tailings to maximise the resource extraction, provide long term mine stability and reduce the surface area required for waste and tailings disposal. Small quantities of waste rock will be used as a capping for tramming purposes on all paste filled stopes.

Testwork suggests that crushing and grinding of ore to P80 of 106 µm with gravity concentration (GRG recovery), high mass pull gravity (HMPG) concentration, separate carbon-in-leach (CIL) treatment of HMPG concentrates, followed by CIL treatment of the re-combined whole of the ore stream can be used to achieve gold recovery of over 91% from the Ollachea mineralization.

The flowsheet applied will comprise two stages of open circuit crushing followed by overflow ball milling, with the mill circuit in closed circuit with hydrocyclones. A gravity circuit employing a centrifugal concentrator will treat a split of the hydrocyclone underflow, the concentrate from which will be further upgraded using a Gemeni table, prior to smelting.

Hydrocyclone overflow will undergo continuous gravity concentration in the HMPG circuit. HMPG concentrates will be pre-aerated prior to being leached in a dedicated Carbon in Leach (CIL) circuit for approximately 24 hours. HMPG tails will be preaerated prior to being recombined with gravity tails and leached for approximately 36 hours in another CIL circuit.

Use of CIL and blanking reagents are techniques that will be employed to reduce the influences of preg-robbing minerals in the ore, such that almost all of the leachable gold will be recovered by the activated carbon in the CIL. To ensure high activated carbon quality, the circuit will include acid washing and Anglo American Research Laboratories (AARL) elution followed by thermal carbon regeneration. Final doré production will be achieved on-site by electrowinning of the AARL eluate and smelting. CIL tailings will be treated to detoxify cyanide using the SO2/air process, followed by iron (cyanide) precipitation by zinc sulphate. After iron precipitation, tailings will be thickened using a high-rate thickener and filtered using press filters. The filter cake will be transported to a paste plant at Minapampa to produce pastefill, when backfill is required in the underground mine. When backfill is not required, the filter cake will be stacked on a load-out platform for reclaim and haulage to a dry-stack tailings storage facility.

Plant design includes a water treatment plant (pre-aeration followed by neutralization and thickening) that will treat mine drainage, site run-off, and process plant excess water prior to discharge to the environment. Reagents including hydrated lime for pH control, sodium cyanide for leaching, kerosene for blanking, hydrochloric acid for carbon washing, copper sulphate and sodium metabisulfite for cyanide detoxification, and zinc sulphate for iron precipitation will be prepared and distributed by reagent preparation circuits considered in the FS plant design.