The Boumadine deposit is described as shallow mineralizing system similar to volcanic-hosted epithermal veins developed in subaerial resurgent caldera environments (Bouabdellah and Levresse, 2016; Abia et al., 2003).

Mineralization is structurally controlled, consisting of zones of vuggy quartz and veins, veinlets, and tectonic-hydrothermal breccias.

Three successive and overlapping stages of mineralization are distinguished (Ait Saadi, 1992; Abia et al., 2003; Bouabdellah and levresse, 2016). The origin of the mineralization is probably the mixing between ascending deep seated fluid and meteoric waters that have generated precipitation of precious- and associated base- metal mineralization.

The Boumadine mineralization is structurally controlled and comprises a system of subvertical veins, veinlets, cemented breccias within veins and filling of tension gashes veins. Mineralized structures occur exclusively within the rhyolitic sequence of the Tamerzaga Formation.

The orientation of the mineralized structures is related to Ediacaran transcurrent tectonics represented by N30 to N-S strike-slip fault zones (Freton, 1988; Ait Saadi, 1992; Abia et al., 1999, 2003). The mineralization and associated alterations are developped at the proximity of magmatic vents. Alteration, which appears around the mineralization, develops an assemblage of major quartz and white micas with minor quantities of chlorite and calcite (Ait Saadi, 1992).

At least five mineralized vein systems are documented at Boumadine deposit. IMARIREN zone, Tizi, NORD zone, CENTRE zone and SUD zone. The rich mineralization tends to occur where NNE, NNW and the textural features indicate that mineralization took place in open space with a relatively shallow mineralizing system (Abia et al., 1999).

All of the mineralized veins display more or less similar mineral assemblages. The mineral sequence established on the crosscutting relationships show three stages of mineralization (Abia et al., 2003; Bouabdellah and Levresse, 2016).

Stage I is characterized by the mineral assemblage dominated by massive pyrite with banded appearance, pyrrhotite, cassiterite and arsenopyrite emplaced under N160 E shortening.

The stage II is characterized by the precipitation of sphalerite, chalcopyrite, calcite, quartz, tetrahedrite-tennantite, silver, gold and Bismuth hosted in massive quartz veins, stock work veins and hydrothermal breccias. Gold is mostly distributed in pyrite and arsenopyrite crystals, and less frequently in galena and sphalerite. Galena is mainly present as a cement of the earlier brecciated sulfides and partially dissolving the arsenopyrite, pyrite and sphalerite. Quartz crystals are filling cavities within all pre-existing sulfides and mainly associated with small euhedral arsenopyrite crystals. Structural and textural data indicate that mineralization took place during the late Neoproterozoic as a result of open-space filling.

The stage III is resulting from the oxidation of primary sulphides and consists of minor amounts of goethite, and jarosite with traces of hematite.

These data further suggest that the Boumadine deposit represents a relatively shallow mineralizing system that was open to the surface, and is similar to volcanic-hosted epithermal veins developed in subaerial resurgent caldera environments (Abia et al., 2003).

It is schedule to start mining by open pit for the first two years the top of IMARIREN zone and mining underground the CENTRE and NORD zone, simultaneously. The financial analysis presumes that all IMARIREN is being extracted by underground method cost wise.

In order to minimize development requirements and take advantage of the polymetallic deposit geometry, the cut and fill mining method was selected for underground exploitation. This mining method is also considered ideal for steeply dipping high grade deposits and highly selective. It is also recommended to use the open long-hole mining method with sub-levels for the wider deposits of each zones.

The process plant will be designed to recover the lead, the zinc, the gold, the silver and the germanium from the fresh ore and the old tailings by a combination of flotation, pressure oxidation and cyanidation. For the fresh mine material the mill will incorporate the following sections : run-of-mine ore storage, a two-stage crushing plant, crushed ore storage, a double stages grinding bay with cyclone classification, lead, zinc and sulfides flotation, pressure oxidation of the sulfides concentrate, leaching (cyanidation) of the gold and the silver, refining and finally a water and reagents distribution system. For the old tailings, the process will be the same except that since the lead and the zinc were already floated and recovered during the 1989-1992 mining- milling operations, there will be no crushing, grinding or flotation. Process of old tailings will be done whenever there will be shortage of fresh ore from the mining operations. Even if for the first two years of operation the mill feed rate will be only 1500 tpd, the mill will be designed to process 2000 tpd since it will be the feed rate from 2023 to the depletion of the actual known resources. The mill process will be monitored at all time by a central Program Logic Controller (PLC).

Crushing
Mining material grading approximately 1.03% Pb, 3.00% Zn, 1.67 g/t Au, 101.76 g/t Ag and Ge from 7.2 to 25.92 g/t will be hauled from the open pit and/or the underground openings by off the road trucks and whenever possible is dumped directly on a static grizzly. However, because the mining and crushing operations will not always be on the same time schedule, it is assumed that 25% of the time the haul trucks will proceed to a RoM stock pile and the rest of the time will dump directly onto the grizzly above the crusher feed hopper. The RoM stockpile area is sized to hold approximately 5,000 tonnes of ore. Secondary handling of the ore will be by a front-end loader that will, among other things, be used to feed the crusher hopper with stockpiled RoM ore as necessary.

Grinding
From the fine ore bins, the material is conveyed to a rod mill. Rod mill discharge is pumped to a set of cyclones. Cyclones underflow flows by gravity to a ball mill. Ball mill discharges into the same pump box as the rod mill and is pumped to same set of cyclones. Circulating load will be in the 300% range.

Cyclones underflow will flow by gravity back to the ball mill while cyclones overflow at a fineness of approximately D80 = 105 µm will flow by gravity to a reagents conditioner before flowing also by gravity to the first cell of the lead flotation rougher.

Flotation
Lead will be floated first. Even if because of the low head grade (0.41% Pb) of the sample tested at SGS failed to produce a smeltable lead concentrate while having a decent lead recovery, the author is confident that if the deposit Pb head grade is similar to the historic head grade, there will be no problem to obtain at least a concentrate grading 50% Pb along with a recovery of 60%. The weight recovery should be in the 1.3% range.

More than probably the lead circuit will require one stage of roughing, one stage of scavenging and two to three stages of cleaning.

Prior to floating the zinc, the lead rougher- scavenger tailings will be reground in a ball mill to a fineness of approximately D80 = 44 µm. The zinc circuit will comprise one rougher-scavenger stage followed by 3 cleaner stages.

According to some flotation tests made at SGS, it is anticipated that the zinc concentrate will be in the 55% range while the recovery will be around 80% for a weight recovery (mass pull) of approximately 2.0%.

It is anticipated the germanium will be recovered with the zinc.

In order to eliminate as much as possible of most of the non gold and silver bearing materials before the treatment of the gold and the silver (mainly silicates), all other sulfides left in the zinc scavenger tailings will be floated and concentrate sent to the gold and silver pressure oxidation + carbon in leach (POX/CIL) circuits. By doing so, some 25% of the non gold and silver bearing material will be eliminated for a very modest loss of 1.0% gold and 1.5% silver.

The sulfide concentrate will be thickened to 50% solid before being pumped to a pre-leach tank and from there pressure pumped to a POX autoclave.

Pressure oxidation (POX)
Pressure oxidation will last some 120 minutes at 50% solid in a 4-compartment autoclave reservoir at approximately 4,000 kPa (40 bars) and 230° C. A cryogenic oxygen plant capable of supplying some 12,000 m3 /h of oxygen per tonne of sulfide (one tonne O2 per tonne of sulfide) going to the autoclave will have to be incorporated to the circuit.

Cyanide leaching
After dissolution of the sulphates, the slurry will be allowed to cool down and filtered. The solid part will be repulped at 50% solid and pumped to a standard 2,000 tpd CIL process circuit. Cyanide leaching will be done for 24 hours. Cyanide (NaCN) and carbon concentrations will be maintained at 1.0 and 10.0 g/l throughout the whole leaching time. pH will be maintained at all time with addition of lime (CaO) between 10.5 and 11.

Carbon elution
The carbon elution circuit (ZADRA type circuit) comprises mainly a loaded carbon tank, an acid wash tank, a carbon strip vessel and a bank of electrowinning cells. Stripped carbon is reactivated in a horizontal kiln, quenched and classified. Classifier oversize is ready to be reused while very fine carbon particles are filtered in a filter press and kept aside to be eventually shipped to an outside smelter.

Refining
Because the whole mill will be serviced electrically by the Moroccan Office National d’Électricité (ONE), refining will be done with induction type furnaces. Since the CIP circuit will not discriminate between the gold and the silver, ingot bars will contain a mixture of both metals.