The San Martín mine is considered to be a typical example of a low sulfidation epithermal deposit, and the geological model used for exploration as well as the mineral resource estimation is that of a low sulfidation vein type deposit. Epithermal deposits form at shallow depths in volcanic-hydrothermal and geothermal environments, typically at temperatures between 160°C and 300°C (White and Hedenquist, 1995). They define a spectrum with two end members, low and high sulfidation (Hedenquist et al., 1998).

Fluid inclusion microthermometry carried out in sphalerite, quartz and fluorite for the Zuloaga vein indicates average homogenization temperatures and salinities of 297°C and 4.1 wt% NaCl eq (Albinson et al., 2001). The homogenization temperatures and low salinities (diluted fluids) in the Zuloaga vein and the alteration-mineral assemblage (interstratified illite/smectite in Zuloaga and Rosario) are consistent with low sulfidation vein-type deposits described by White and Hedenquist (1995) and Hedenquist et al. (1998). The presence of epidote in Zuloaga, Rosario and Other Veins is also consistent with the homogenization temperatures around 300°C determined for Zuloaga by Albinson et al (2001).

In the opinion of the QP, the deposits in the San Martín mine area are considered to be examples of low sulfidation epithermal deposits. The Miocene age of the mineralization and its association with volcanic rocks of the Upper Volcanic Series is also found for other low sulfidation deposits in Mexico. Additionally, structural-textural features, such as hydrothermal breccias cemented by quartz-calcite, stockworks and cymoid loops, are also common in other low sulfidation epithermal vein-type deposits in Mexico.

Mineralization in the San Martín mine occurs in east–west, northwest–southeast, northeast southwest and north–south fault structures in the form of stockworks, sheeted veinlets, veins, and breccias. The veins in the San Martín mine can be described as fault veins or mineralized faults, given that the amount of gangue minerals such as quartz, calcite, fluorite, epidote, ankerite and adularia are very limited, i.e., they do not form massive or banded veins typical of open space-filling veins.

San Martín veins and deposits are hosted on the side of a mountain range. Access to the workings is through adits developed horizontally, followed by ascendant and descendent ramps developed in waste.

All mine workings in San Martín are located above the water table, and no evidence of water bodies have been found during exploration. There are water inflows in the workings close to surface, mainly during the rainy season, but these inflows are managed by pumping.

Geotechnical studies have been completed in support of design parameters for the excavations, as well as ground support requirements.

San Martín currently uses cut-and-fill mining using resue to extract the mineralization. Resue is a mining variation that implements a two-phased process where the ore is extracted first and then the mining section is extended to allow access to mining equipment for subsequent cuts. A combination of jumbo and conventional (hand-held pneumatic) drills are used and the type of drill used depends on mining widths and availability of the jumbos.

The current minimum mining width used at site for cut-and-fill mining is 0.8 m, and 2.5 m for equipment access. After the resue portion is mined (typically the mineralization), additional waste is mined to allow for equipment access. Mined waste either reports to the surface waste storage facility or is used as fill for subsequent lifts. When mineralization that is greater than 2.5 m in width is mined, no additional waste is mined. Each drift is mined 3 m high where six drifts are mined to extract 18 m of a 20-m-high panel. Updated designs incorporate a minimum stand-off distance of 20 m to locate ramps away from mineralization. Planned development includes: access drifts; sills (development on mineralization); operating waste development (sills mining material below cut-off); sumps; escapeways and accesses to the escapeways; return airways and accesses to the return airways; stockpiles; and ore passes and access to the ore passes, where required. Vertical development will primarily be completed by conventional mining techniques up to a size of 1.5 m by 1.5 m. Large diameter raises will be excavated either by a raisebore machine (contract) or by longhole raising.

Where necessary, all future production voids will be backfilled. As the operation uses sill pillars to separate active mining blocks, the backfill is uncemented waste rock.

The existing load-and-haul fleet currently handles up to 900 tpd (27 kt per month), with haulage requirements met by the onsite contractor through the provision of conventional haulage trucks. The mine plan uses development rates and productivities based on the existing fleet.

Gold and silver are extracted at the processing plant which operates 24/7, normally processing 860 tpd. The plant has a name plate capacity of 1,300 tpd, which provides ample operational flexibility.

The processing plant comprises several areas: crushing, grinding, leaching, tailings dam, Merrill-Crowe circuit (auto jet and precipitation), and refinery. A project is currently underway to build a new tailings filter dry stack tailings area.

The crushing area normally operates 18 hours per day, allowing six hours a day for maintenance and housekeeping tasks. Run of Mine (ROM) ore is transported to the crushing area where it is stockpiled in a yard near the primary (jaw) crusher. With the help of a front-end loader, ore is fed through a chute into the primary (24” x 36”) crusher, which is capable of handling a maximum rock size of 14”. The crushing circuit is equipped with a stationary hydraulic hammer to break oversized ROM material when present.

The primary crusher reduces the ore to 4" top size. Particles finer than 3/8” in the primary crusher discharge stream are separated in a vibrating screen (Sandvik LF1240D) and then sent to a fine ore bin (final crushing circuit product). The plus 3/8” material is processed in a secondary (4.5’ Symons) cone crusher where ore is reduced to 5/8". Particles finer than 3/8” in the secondary crusher discharge stream are separated in a second vibratory screen (Sandvik LF1240D) and then sent to the fine ore bin. The plus 3/8” material in the secondary crusher discharge stream is processed in a tertiary cone crusher (Sandvik CH430) which operates in closed circuit.

The grinding circuit comprises three balls mills (10’ x 10’, 9’ x 12’ and 9’ x 9’) operating in closed circuit with hydrocyclones. To promote metal extraction, semi-pregnant solution containing 1,000 ppm of sodium cyanide is added to the grinding circuit.

The three mills are primary mills; however, only mill #1 (10’ x 10’) and mill #2 (9’ x 12’) are currently in operation as the installed capacity of these two mills is sufficient to process the typical tonnage of 860 tpd. A key feature of the grinding circuit is that mill #1 and #2 can operate as primary and secondary (grind–regrind) mills. This provides the capacity to achieve very fine (85% passing 200 mesh) feed to the leaching circuit, significantly improving metal recovery due to the higher particle liberation.

The leaching circuit consists of two 50’ diameter primary thickeners (one in operation and one in standby) which are fed with the product from the grinding circuit and with semi-pregnant solution. The dirty pregnant solution in the primary thickeners is sent to the Merrill-Crowe plant, while solids are agglomerated and settled with the help of flocculants and limewater.

The thickened solids are pumped to the first agitated tank, and from there by gravity into 14 additional tanks in series (15 in total). The nominal residence time in the tanks is 120 hours. Leaching takes place in the presence of cyanide and air. Typical conditions in the leaching tanks are: 43–45% solids, at least 7 ppm oxygen concentration, pH of 10–11, and sodium cyanide at a concentration of 900–1,100 ppm. From the last agitator, the pulp is sent to the Counter-Current Decantation (CCD) circuit.

The CCD circuit consists of six thickeners, with thickener #6 installed at the highest elevation and #1 at the lowest. It is worth noting that Flocculant and limewater may be added when needed to increase the efficiency of sedimentation and clarification. The pulp from the last agitated tank (#15) is sent by gravity to thickener #1, from this to #2, and so on until reaching #6, after which the remaining pulp is sent to the tailings dam.

Barren solution is added to thickener #6 and then by gravity to thickener # 5, and so on until reaching thickener #1 (i.e., in counter current relative to the flow of pulp). This process enriches the barren solution, and it becomes semi-pregnant. This semi-pregnant solution is fed to the mills and to the primary thickeners, as mentioned above.

There are two tailing dams currently in operation. These dams are fed with the discarded material from thickener #6. With the aid of a hydrocyclone, the pulp is separated into coarse and fine particle fractions. The coarse material forms the lip of the dam, and fines are deposited close to the center of the dam from where solution is recycled.

The Merrill-Crowe plant comprises two sub areas: auto jets and precipitation. The plant is fed with a clean pregnant solution coming from the primary thickeners; this passes through the auto jet filters for clarification, lowering the turbidity from 30 ntu to below 5 ntu. Once filtered, the solution passes to the deoxygenation towers (2 towers that work in parallel), where the oxygen is withdrawn from the solution with a vacuum pump and zinc powder is added to precipitate the gold and silver. The solutions are then pumped through filters presses where the gold and silver are recovered in the form of a precipitate. There are three presses which work alternately, operating at maximum two per day to achieve the required pumping rate of for recovery of values 7,200 m3 per day, while the other press remains in standby.

In the refinery, the zinc precipitate is treated with an acid solution to eliminate zinc and other metals. It is then dried in an electrical dryer prior to melting. A mixture of borax, sodium carbonate, silica and coal is used as flux. Once the precipitate has been mixed with the reagents, it is deposited in one of two induction furnaces (two furnaces provide sufficient time for maintenance). The precipitate is melted into the induction furnaces at temperatures close to 1,200°C. The material is poured into doré bars. The quality of the doré bars average more than 90% Ag and 9% Au. The bars are delivered to various refineries based on annual tendering, for final refining to deliver a product of 99.99% Ag ounces and 99.9% Au ounces. Since 2013, there has been no penalty related to deleterious elements as per the current commercial terms.