The Morro do Ouro is a metamorphic gold system with finely disseminated gold mineralization hosted within metasedimentary rocks.

Gold mineralization was introduced syn-tectonically as the result of metamorphic alteration during thrusting of the Morro do Ouro Sequence over the rocks of the younger Vazante Formation. Structural interpretation suggests that mineralization was precipitated within a high strain zone where silica and carbonate were depleted from host phyllites, resulting in an increase in graphite content that may have acted as a chemical trap, precipitating out gold and sulfide mineralization remobilized during the metamorphic alteration of the Morro do Ouro Sequence.

The deposit has extraordinary lateral and longitudinal continuity. The majority of exploration efforts have sought to better define the continuous longitudinal continuity of mineralized phyllites at depth west of Rico Creek and the lateral limits of the economic mineralization.

In the property area, the rocks are phyllites and quartzite intensively altered by hydrothermal processes associated with regional metamorphism.

The entire mineralized system lies within a thick, heterogeneously deformed zone that contains both abundant NE vergent shallowly dipping shear fabrics and a strong planar or tabular (flattened) strain signature.

The Paracatu mineralization has two characteristic visual features. The first is the presence of a sulfide suite intimately associated with gold, comprising, in order of abundance, arsenopyrite, pyrite, pyrrhotite, sphalerite, galena, and chalcopyrite.

The sulfides occur in a variety of forms, predominantly within boudinaged quartz veins or on their edges, and in the necks of the boudins. Some sulfides (pyrite, arsenopyrite, and pyrrhotite, very rare chalcopyrite and sphalerite, no galena) are also present in the host shales as veinlets and disseminations, usually in the cm- to m-scale in the vicinity of the sulfide-bearing quartz veins. (Oliver et al., 2015).

The second characteristic feature of Paracatu mineralization is the occurrence of boudinaged quartz ± carbonate ± sulfide veins.

Boudins make up, on average, 8% to 10% by volume of the mineralized rock, though there are wide variations. The central part of the orebody may contain >20% boudin material, but at the margins volumes drop off to 1% to 2% or less. Boudins and their immediate host rocks contain >90% of the sulfides and gold, based on comparison of bulk ore analyses, boudin analyses, and spatial analysis of boudin/vein distribution.

Deformation has produced distinctly separated boudins, distributed along linear trains at a low angle to the bedding of the host stratigraphy. These structures were affected by predominantly oblate flattening strains, producing “chocolate-tablet” boudinage, with less common examples of simple log-like boudins expected from a plane strain deformation (Oliver et al., 2015).

Gold occurs either as free gold or electrum. Microscopic analysis indicates that 92% of the gold at Paracatu is free-milling with less than 8% encapsulated by sulfide grains or silica.

The Paracatu operation consists of an open pit mine, two process plants, two tailings facilities, and related surface infrastructure and support buildings.

At Paracatu, ore hardness increases with depth and, as a result, modelling the hardness of the Paracatu deposit is important for costing and process throughput parameters. KBM estimated that blasting of the Paracatu ore would be necessary for blocks with a BWI greater than 8.5 kWh/t.

As mining progresses to the southwest area of the pit, it is necessary to increase hauling capacity because of waste stripping. Currently the truck fleet consist of 25 CAT 793 and the life of mine peak is 35 trucks in 2024.

The open pit design criteria:
• Bench Height - 12 or 24 m;
• Bench Face Angle - 45 to 75°;
• Berm Width - 8 to 12 m;
• Berm Interval - 20 m;
• Inter-ramp Angles (Weathered Rock) - 38.8°;
• Inter-ramp Angles (Fresh Rock) - 49.2 to 56.4°;
• Inter-ramp Angles (Soil) - 26.6º.

Haul roads and in-pit ramps were designed to be 40 m wide with a gradient of 10%.

Paracatu has completed two waste dumps, Sul and Oeste. By the end of the mine life, another two dumps will be completed: the Central (in-pit) and the Ex-Pit dumps.

Paracatu operates 24 hours per day, 365 days per year, with 2 x 12h shifts to ensure continuous operation.

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Paracatu has two processing plants known as Plant I and Plant II. Plant I has operated continuously since 1987 and Plant II since 2008. Paracatu has been reprocessing tailings since 2015.

Flotation and Regrind (Plant I)
The grinding circuit product, cyclone overflow, feeds the rougher flotation circuit consisting of Wemco (10 cells of 42.5 m3 each); Outokumpu (4 cells of 16.5 m3 each); and Smartcells (4 cells of 127 m3 each). A portion of the rougher concentrate is fed to a Knelson Concentrator (QS48).

The cleaner flotation circuit, Outokumpu cells (5 cells of 16.5 m3 in operating) receives the concentrate from the rougher circuit.

The cleaner tails are refed to the rougher flotation feed, while the rougher tailings are discharged to the tailings dam.

The cleaner concentrate feeds dewatering thickener, which is subsequently fed to a regrinding circuit consisting of two ball mills (220 kW and 130 kW) operated in closed circuit with hydrocyclones. A portion of the cyclone underflow is treated by a Knelson concentrator (XD30). The regrinding circuit product at a P90 of 45 µm is directed to a dewatering thickener. Dewatered slurry is directed to the Hydro (CIL) circuit for further processing.

Concentrates from the Knelson concentrator are processed in an Acacoa intensive leaching system. The thickener overflow water returns to the Plant I grinding circuit.

Flotation, Gravity and Regrinding (Plant II)
The flotation rougher circuit receives fresh feed plus circulating load from cleaner flotation cells.

The rougher circuit consists of 24 rougher flotation cells, arranged in four rows of six cells each, which are fed by cyclone overflow from the grinding circuit. The cells are 160 m3 tanks fitted with self-aspirating mechanisms. Reagents (collectors and frother) required for rougher flotation are supplied from reagent tanks to head tanks in the grinding circuit, fitted with metering pumps adding reagents to the grinding hydrocyclone overflow launder.

First rougher concentrate is processed by three Knelson concentrators (QS48). The rougher concentrate is treated in a cleaner circuit consisting of two rows of five selfaspirated 60 m3 tank cells. Cleaner tailings flow to a pump box and are pumped by a horizontal slurry pump to the rougher flotation feed distribution box. Cleaner concentrate is collected in a single pump box and pumped by horizontal slurry pumps to the dewatering thickener. Water from thickener overflow returns to the ball mill discharge boxes and the dewatered concentrate is then fed to a vertimill (13.5 m high, 931 kW vertical stirred type) where it is ground to 90% passing 45 microns. The mill operates in closed circuit with ten 254 mm diameter hydrocyclones (GMAX26). A Knelson concentrator (Model XD40) treats the circulating load (underflow) of the vertimill.

The vertimill product, hydrocyclone overflow, reports to the dewatering thickener. Dewatered slurry is directed to the Hydro circuit for further processing.

The gravity concentrates are transported to the Hydro (CIL) Plant for further processing in an Acacia reactor.

Hydro
Carbon in Leach
Hydro receives the flotation concentrate from Plant I and Plant II. The typical concentrate gold grade is 10-15 g/t and total sulfur content is 20-25%. The concentrate is fed to an agitated pre-aeration tank where lime is added to adjust the pH to approximately 10.5. Pre-aeration residence time is approximately four hours and CIL residence time is 30- 40 hours. The CIL circuit consists of 8 tanks with a volume of 750 m3 each.

Oxygen is injected into the pre-aeration tank using a fillblast system to increase the dissolved oxygen concentration to approximately 9-10 mg/L. The slurry flows by gravity through an eight stage CIL circuit. Cyanide is added into the first CIL tank and is adjusted to a concentration of approximately 700-750 mg/L NaCN. The cyanide concentration is allowed to decrease down the CIL circuit, reaching 150-200 mg/L NaCN in the final tank.

The leached slurry passes out of CIL 8 into the 2x cyanide treatment tanks (250 m3 each).

The tailings slurry is treated for cyanide destruction in an agitated tank using ammonium bisulfite and oxygen.

Carbon Elution and Regeneration
Loaded carbon is transferred out of CIL twice a day. The loaded carbon is screened and flows by gravity to an acid washing column. The carbon is treated with 5% hydrochloric acid for 4 hours to remove calcium carbonate deposits and other inorganic contaminants. Spent acid is neutralized with sodium hydroxide before discarding it to the tails pump box. From the acid wash vessel (1 column, 14t capacity; 36 m3) carbon is pumped to the elution columns (2 columns, 14t capacity; 36 m3). The elution cycle operates with a 0.2% sodium cyanide and 2-3% sodium hydroxide solution at a temperature of 140°C and a pressure of 300 kPa for approximately eight hours. After elution is complete, carbon is pumped to the regeneration kilns. Two electrically powered kilns with 600 kg/h of capacity each, regenerate the carbon at 700°C. Regenerated carbon is screened to remove fines before being reintroduced to the CIL circuit.

Intensive Cyanidation
Gravity concentrates produced from the Plant I and Plant II gravity circuits are treated by intensive cyanidation. Plant I gravity concentrate is treated in an Acacia CS2000, while Plant II concentrate is treated in an Acacia CS8000. The pregnant leach solution from the Acacia systems are pumped to the electrowinning circuit. The solids residue from the Acacia system is sampled and pumped to the CIL circuit.

Electrowinning and Refining
Pregnant solution from elution is combined with the solution from the gravity circuit and is pumped to four electrowinning cells, which are sludging cathode type and were fabricated by Summit Valley; Model 125 EC33; 3.5 m3 each. Periodically, the cathodes are cleaned using a high pressure washer and the gold-bearing sludge is recovered by a filter press. The resulting filter cake is dried, mixed with fluxes, usually borax, soda ash and occasionally sodium nitrate and fed to electric induction furnaces. The doré metal and slag separate in the furnace, and the slag is poured off into slag pots. The doré metal is then poured into bars.