As documented by Lewis and Murahwi (2013) and Munroe (2014), the San Sebastián del Oeste silver-gold district comprises classic, high grade silvergold, epithermal vein deposits, characterized by low-sulphidation mineralization and adularia-sericite alteration. The veins are typical of most other epithermal silver-gold vein deposits in Mexico in that they are primarily hosted in volcanic flows, pyroclastic and epiclastic rocks, or sedimentary sequences of mainly shale and their metamorphic counterparts.

Low-sulphidation epithermal veins in Mexico typically have a well-defined, subhorizontal ore horizon about 300m to 500m in vertical extent where the bonanza grade mineralization shoots have been deposited due to boiling of the hydrothermal fluids. Neither the top nor the bottom of the mineralized horizons at the Terronera Project has yet been established precisely.

Low-sulphidation deposits are formed by the circulation of hydrothermal solutions that are near neutral in pH, resulting in very little acidic alteration with the host rock units. The characteristic alteration assemblages include illite, sericite, and adularia that are typically hosted by either the veins themselves or in the vein wall rocks. The hydrothermal fluid can either travel along discrete fractures where it may create vein deposits or it can travel through permeable lithology such as a poorly welded ignimbrite flow, where it may deposit its load of precious metals in a disseminated deposit. In general terms, this style of mineralization is found at some distance from the heat source.

Geologic information and field observations indicate that the hydrothermal system at the Terronera vein is preserved over an elevation difference of 600m. Regionally, the known deposits contain polymetallic sulphide mineralization in wide vein structures. The veins at higher elevations may represent the tops of ore shoots containing significant silver and gold mineralization at depth.

The Mineral Reserve extends from the 1,610m to 1,260m elevations, a vertical distance of approximately 350m, and has a lateral extent of approximately 1,400m. A conceptualized mechanized cut and fill mining method plan has been laid out to extract the deposit using trackless underground equipment, including scooptrams, haulage trucks, and electric-hydraulic drill jumbos. A small amount (7% of total Mineral Reserve) of mechanized long hole mining is also required for the recovery of sill pillars.

Primary access to the deposit will be via a 526m long -12% 5m by 5m trackless haulage ramp, from the portal, at the 1469 m Level to the 1410m Level. The -12% haulage ramp will be driven an additional 283m to the 1380m Level haulage drift. The 1380m Level haulage drift will be driven at +1% for an addition 892m, during the next 3 quarters of a year to the end of the 1380m Level haulage drift.

Access to the mining zones will be via a series of five +/-12% up-and-down spiral ramps, access cross-cuts, and stope attack cross-cut ramps. This development will have cross-sectional dimensions of 4.5m by 4.5m.

The principal mining method envisaged is mechanized non-captive cut and fill using trackless underground equipment, including scooptrams, haulage trucks and electric-hydraulic drill jumbos. The average mining width is estimated to be 4.4m.

Initial access to the stope lifts will be via cross cut attack down slope ramps. The initial bottom cut and fill lift will be excavated 4m high by the width of the mineralized domain (up to 6m wide), as drift development. Stope lifts will be mined from the bottom up. The backs of the attack ramps will be slashed to access higher lifts. Stope lift strike lengths will be up to 300m in length and 150m in each direction from the attack ramp access. Mining breasts will be drilled using horizontal blast holes. Where more than one vein occurs in the same mining block, adjacent veins will not be mined) simultaneously at the same elevation.

A brief description of the processing circuits included in the design is provided in the following paragraphs.

• The crushing circuit is comprised of a (24 tonne) dump ore pocket fitted with a stationary Grizzly with a 10 inch by 10 inch opening. The Grizzly oversize is broken with a hydraulic breaker.
• An apron feeder sends the ore to a primary jaw crusher to reduce the material to minus 85 mm.
• The crushed material and the dust fines from the feeder are transported to subsequent stages of screening and crushing in closed circuit to further reduce the material to minus 9 mm. Conveyor belts are used to transport the intermediate and fine crushed materials throughout the entire crushing circuit.
• The crushing circuit design provides weigh scales, crushed ore sampling system and magnetic separators to protect the cone crusher from iron coming from underground mining operations. The finely crushed product is transported to two fine ore bins (A and B) with 750 metric tons live capacity each.
• A variable speed belt feeder transports the crushed material from the fine ore bin to the primary grinding mill. The material is ground to 80 percent minus 150 mesh (100 microns) in closed circuit with a battery of cyclones.
• Some flotation reagents will be added into the grinding mill to allow for conditioning. Ground slurry (cyclone overflow) at approximately 23 percent solids is sent to a trash screen for removal of debris produced in the underground mining operation. After trash removal the clean slurry is directed to a conditioning tank where flotation reagents are added for conditioning prior to the flotation process.
• The flash flotation cell is installed in the grinding circuit area. This cell is fed with a portion of the cyclone underflow slurry. Dilution water is added to allow flotation of liberated values. The flash flotation tail product is directed to the mill feed box. The flotation concentrate is sent directly to the final concentrate thickener.
• The flotation circuit consists of two lines (A and B) banks of Rougher followed by Scavenger cells to achieve maximum precious metal recovery. The Rougher and Scavenger concentrates are sent to a regrind circuit to achieve optimum liberation of precious metal values. The cyclone overflow from the regrind circuit feeds a two stage cleaning circuit to achieve the highest possible gold and silver grade in the final concentrate. The first cleaner tailing product is returned to the head of Rougher flotation.
• The second cleaner concentrate reports to the concentrate thickener. The final concentrate is filtered and the filter cake with moistures ranging from 15 to 20 percent is stored and air dried in a warehouse prior to shipment.
• Each concentrate shipment will be sampled and analyzed for precious metal and moisture contents. Impurities present in the concentrate will be quantified and evaluated prior to shipment.
• Flotation tails will be sent to a thickener. The higher density slurry produced will be filtered using two ceramic disk filters. The filter cake produced will be stockpiled at the dry tailings filter plant prior to sending the solids to the dry tailings storage facility (DTSF). After sedimentation and filtration, the flotation tailings will be transported to the dry tailings stacking area by means of trucks. The filtered tailing material will be placed in a stockpile by means of a radial stacker. Front end loaders and compaction equipment are ultimately used to load, spread and compact the tailing material as required in the TSF.