The Fazenda Brasileiro gold deposit is an epigenetic, structurally controlled and hydrothermally altered Precambrian quartz vein hosted lode gold deposit that has been subjected to greenschist facies metamorphism. There is some suggestion of a partial syngenetic origin for the gold because of the anomalous gold content (0.05 g/t Au to 0.10 g/t Au) throughout visibly unmineralized CLX.

Gold mineralization is related to multi-phase quartz veining events. The gold occurs in at least three textural settings at MFB: as particles attached to sulphide grains, as particles within fractures in sulphide grains, and as particles within fractures in quartz/albite gangue. Gold grains typically contain less than 5% silver.

The bulk of the mineralization is hosted by quartz-albite-sulphide veins within the chlorite schist (CLX) unit. Individual veins vary from one to four centimetres thick, have irregular margins, and are typically oriented sub-parallel to the predominant east-west trend of the felsic and mafic schists. The veins occur in multiple vein sets which vary in true width from 1.5 m to 40 m and in horizontal mining width from a minimum 3 m to 40 m. The veins are generally arcuate in an east-west trend and south dipping at 40° to 70°, with a shallow to moderate east plunge. The plunge, however, is quite variable, with some zones plunging westerly.

Mineralization is also found stratigraphically below the CLX unit in the CLX2 and Canto horizons. Economic mineralization occurs in horizontal to sub-horizontal shoots, the locations of which are influenced by a combination of folding and shearing. Shoots range from tens of metres to hundreds of metres in length and tens of metres in height.

Many of the identified mineralized lenses outcrop to surface. Over the course of the operation’s history, several shallow open pits have been excavated to extract these near surface deposits. Currently, several small open pits are in operation, and mining is being completed using contractors. Typically, these small pits are 30 m to 50 m deep and employ air-track drills and backhoe excavators for mining, and highway-type trucks for haulage to the mill.

The main access to the underground operation is through a series of declines. Over the life of the operations, eight main declines - the B, C, D, E, F, G, EW, and EDEEP - have been developed along the strike of the mineralization. A main central shaft (470 m deep) exists, however, it is no longer used for hoisting as the portion of the deposit located within economic distance of the shaft has been mined out. The shaft is now utilized only as a part of the escapeway and ventilation circuit.

Underground mining employs blast hole stoping from sub-levels developed along the trend of the mineralization. The stoping areas are accessed initially from 5 m wide by 5.5 m high main haulage ramps developed at 12% road grade in the footwall, which leads to primary development crosscuts of 4.5 m wide by 5.1 m high, and secondary development drifts and crosscuts of 4.5 m wide by 4.7 m high. Sub levels are spaced at 25 m vertical intervals. Mined out stopes are not backfilled.

Sub-levels are developed into the stoping areas and fan drilling of blast holes into the mineralization is used to further define the boundaries of the mineralization and design the ultimate blast patterns. Remote-controlled 12 t Load-Haul-Dump (LHD) machines are used to load and haul the ore from the stoping areas to 25 t and 35 t articulated haulage trucks at loading points in the sub-levels.

The sub-horizontal plunge and approximate 45° dip of the orebody, combined with a thickness up to 40 m, provides for low development and operating costs. Maximum stope heights are 20 m. Future operations in the deeper areas of E Ramp will have higher haulage costs that will be partially offset by the shorter underground haulage in the F and G Ramps.

All bodies have a planned dilution of 14%, except for the F Body, which has a dilution of 20%. Planned mining recovery was estimated to be 91%, except for F Body, which is 89%. In a small number of areas, where weaker ground is anticipated, planned mining recovery was estimated between 83% and 85%.

The overall process flow sheet consists of:
• Three stage crushing;
• Ball mill grinding consisting of two mills in parallel closed with cyclones;
• Gravity concentration using centrifugal concentrators; treating the underflow of one cyclone in each of the grinding cyclone clusters;
• Thickening producing a leach feed of 50% solids;
• Cyanide leaching in two parallel circuits, Circuits 1 and 2;
• CIL in two parallel circuits, Circuits 1 and 2;
• Zadra pressure stripping of the carbon;
• Intensive cyanidation of the centrifugal concentrator concentrates;
• Electrowinning of the carbon eluent and gravity concentrate leach solution;
• Casting of gold bars in an induction furnace.

The ore flows from the feed hopper to a twin shaft Blake type 1,200 mm by 900 mm primary jaw crusher with a closed side setting of 89 mm.

The crushed ore is drawn from the stockpile by two vibrating feeders, installed in a tunnel under the stockpile, and conveyed to the secondary/tertiary crusher area. The ore is conveyed to the primary double deck, vibrating screen. The upper deck has 30 mm openings and the bottom deck has 12 mm openings. The oversized material from both decks feeds a Metso HP300 secondary cone crusher with a closed side setting of 19 mm. The secondary crusher product is conveyed to the secondary screen feed bin and then to two secondary double deck vibrating screens operating in parallel.

The crushed ore is drawn from the mill feed bins by belt feeders into the two overflow ball mills operating in parallel. One ball mill is 3.8 m diameter by 7.3 m long with a 1,500 kW drive and a capacity of 116 tph. The second mill is a 3.0 m diameter by 5.2 m long with a 635 kW drive and a capacity of 45 tph.

The smaller, 3.0 m diameter ball mill feeds a bank of four cyclones, with two cyclones operating and two in reserve. The cyclone overflow flows to the leach feed thickener. A portion of the cyclone underflow is diluted to 40% solids and feeds a centrifugal concentrator. The balance of the underflow slurry returns by gravity to the ball mill feed chute. The larger, 3.8 m diameter mill feeds two banks of cyclones, each with one cyclone operating and three in reserve. The cyclone underflow slurry from the first bank is diluted to 40% solids and flows to three centrifugal concentrators operating in parallel. The underflow slurry from the second bank returns by gravity to the 3.8 m ball mill feed. The tailings from the centrifugal concentrators are pumped to the second bank of cyclones. The underflow returns to the ball mill feed chute and the overflow will advance to the leach feed thickener.

The concentrate from the centrifugal concentrators is pumped to the intensive cyanidation, Acacia, reactor. The Acacia reactor is an automated system providing security for the processing of gravity gold concentrates. The concentrate is leached for twelve hours at 54°C, in a solution of 2.5% sodium cyanide and 1.5% sodium hydroxide to recover the gold. The pregnant Acacia leach solution is then pumped to a storage tank in the refinery area in preparation for electrowinning.

The cyclone overflow slurry is pumped to a horizontal vibrating screen for the removal of trash and other impurities. The screen underflow flows by gravity to a 32 m diameter leach feed thickener where it is diluted to 15% solids, flocculent is added and directed into the thickener feed well.

The thickener underflow slurry is pumped to two parallel trains of leach and CIL tanks.

Train 1 consists of six 350 m3 leach tanks in series, followed by six 110 m3 CIL tanks in series. Leach tanks 1 through 3 are being used for pre-aeration, to oxidize sulphides in order to reduce cyanide consumption. Lime is added and air is spared into the tanks during pre-aeration. Cyanide is added to tank 4 to begin leaching and carbon is being transferred to tanks 5 and 6 from the CIP circuit to operate as CIL tanks. The six CIP tanks are now operating as CIL tanks, though no air is sparged into the tanks. In summary, Train 1 consists of three aeration, one leach, two 350 m3 CIL, and six 110 m3 CIL tanks.

Train 2 consists of eight 500 m3 leach tanks in series followed by six 350 m3 CIL tanks. Similar to Train 1, leach tanks 1 through 3 are pre-aeration, tank 4 is a leach tank, where cyanide is being added, and leach tanks 5 through 8 are CIL tanks. The six CIP tanks are operating as CIL tanks. In summary, Train 2 consists of three pre- aeration tanks, one leach tank, four 500 m3 CIL tanks, and six 350 m3 CIL tanks.

The activated carbon is added to the last downstream tank in the CIL circuits and transferred from tank to tank counter current to the slurry flow. The carbon is loaded to approximately 800 g/t Au. Loaded carbon is pumped from the first CIL tank to the loaded carbon wash screens. The carbon passes over the screen and is collected in portable carbon transfer bins. The carbon is transferred to the carbon desorption and regeneration, ADR, circuit, while the screen underflow slurry returns to the first CIL tank.

The gold is eluted from the carbon using the pressure Zadra process. The process treats the carbon in four tonne batches.

The carbon eluent and the pregnant gold solutions from the Acacia intensive cyanidation reactor are processed by electrowinning. The solutions are circulated though electrowinning cells and the gold in solution is plated out onto stainless steel cathodes. Once the cathodes are loaded, the cathodes are removed from the cells and the gold is washed from the cathodes with high pressure water. The resulting gold mud is filtered and then dried in a drying oven. The dried mud is then mixed with fluxes and melted in an electric induction furnace at 1,200 degrees C to produce gold doré bars for sale. The doré from the leaching circuit typically assays 86% gold, and the doré from the gravity and intensive cyanidation circuit typically assays 90% gold.