The Citronen Deposit is interpreted as belonging to the SEDEX deposit class, forming syn-depositionally with sedimentation.

SEDEX deposits are formed in submarine environments by the precipitation of sulphides from metal bearing fluids introduced onto the seafloor through underlying fractures which act as metal-bearing fluid conduits. Large amounts of sulphur are precipitated principally as pyrite and focused around vent areas or ‘mounds’ on the sea floor. Base metal (Zn + Pb) bearing sulphides at Citronen are predominantly located within laminate horizons surrounding these larger sulphide accumulations.

Mineralisation at the Citronen Fjord Zn-Pb Deposit comprises several distinct sulphide mounds containing massive and net-textured pyrite-rich mineralisation, interpreted to represent the focal point of fluid influx, flanked by pyritic laminated sulphides that are locally sphalerite and galena-rich. These laminated sulphides host the majority of economic grade mineralisation.

The deposit consists of multiple sulphide mounds forming in three lateral positions (“vents”), defined as the Discovery, Beach and Esrum ore bodies.

The mineralisation is hosted within two fine- grained sedimentary units, separated by a mass carbonate debris flow. The majority of mineralisation is stratiform, with semi-massive net-textured to massive sulphides accumulating in the core of the mound structure. Structurally controlled stock-work style mineralisation is present within the carbonate debris flows, with the most notable termed the XX ore body. The stratiform mineralisation has been identified from outcrop to a depth of approximately 500 metres, with the mineralisation open at depth. Level 1 is located predominantly within the Discovery ore body, Level 2 is evident discontinuously across all three ore bodies and Level 3 contains the largest volume of sulphides, with a lateral extent of over 3,000 metres between the Beach and Esrum ore bodies.

The mineralisation is pyrite dominated with variable amounts of sphalerite ((Zn/Fe)S) and lesser galena (PbS) present as sulphide species. Minor chalcopyrite (CuFeS2) has been documented and interpreted as having formed during remobilisation and enrichment of primary stratiform-hosted mineralisation. No economically significant Cu or silver (Ag) has been identified to be associated with the sulphide mineralisation.

Primary mineralisation is generally fine to medium grained, weakly to moderately laminated and bedded parallel with regionally deposited sediments. Gangue mineralogy is primarily silt and clay from mudstones deposited contemporaneously with sulphide mineralisation.

The ore body’s nature and orientation is well understood and the room and pillar mining method has been selected as the most appropriate for this style of deposit.

With the mining method being room and pillar, a recovery factor of 100% recovery was assumed as ore loss should be negligible provided that a tele-remote bogging system is used for the recovery of the ore drive strip material. The likelihood of ore loss would increase if a line of site remote bogging system was used.

All mining has been planned for conduction via development jumbo and, as such, minimal overbreak (i.e. unplanned dilution) outside of the planned dilution skin is expected. As a result, no tonnage factors were applied to this design.

Open pit mining operations have been scheduled to commence at the end of the life of the underground mine in Year 11 (3.3 Mtpa scenario).

The open pit mining rate to continue to maintain a total ore feed rate of 3.3 Mtpa to the process plant until the open pit is depleted. Mining will utilise 10 x 60t haul trucks and excavators until the point where the project is closed out, subject to the discovery of additional resources.

A compact two stage crushing plant has been selected, with DMS testwork indicating acceptable metal recovery and mass rejection can be obtained with a primary crusher discharge P80 of 130 mm reduced to a P80 of 35 mm in the second stage.

The circuit equipment has been selected to achieve approximately 550 t/h average throughput levels at 18 h/d operation (6.5 d/wk, 365 d/a). The ore is delivered by haul truck and either direct dumped into the run of mine (ROM) bin or onto the ROM pad and later transferred by front-end loader (FEL) to the ROM bin. The ROM bin is protected by an 800 x 800 mm grizzly, and a rock breaker is mounted adjacent to break up oversize.

The DMS feed bin discharges at an approximate average of 375 t/h by belt feeders to a wet feed preparation screen that removes all -1.0 mm material, with the screen oversize feeding into the DMS cyclone feed hopper.

The DMS cyclone pump feeds -38+1 mm material combined with dense medium to the DMS cyclones. The DMS cyclone overflow is stripped of medium on a drain-and-rinse dual screening system with screen oversize conveyed to the external floats stockpile. The DMS cyclone underflow is stripped of medium on a drain rinse screen system, and the screen oversize is conveyed to the sinks (ball mill feed) bin.

The mill cyclone overflow of P80 size 45 µm gravity flows to a trash screen of 0.8 mm aperture. The trash screen undersize is pumped to a single conditioner that overflows to a pre-flotation circuit. The pre-flotation (pre- float) stage is necessitated by a carbonaceous population that would otherwise impact downstream grades and froth stability. Dextrin depressant and frother are dosed to the conditioner tank. Trash screen oversize is collected in a trash bunker for disposal.

The pre-flotation concentrate is discarded to the tailings thickener, and the intermediate tailing flows to two conditioner tanks in series (three minute residence time in each) in which lead circuit pyrite depressants and collector are dosed. The conditioners overflow in sequence to the lead rougher scavenger flotation cells at a design rate of approximately 310 t/h. Frother is dosed to the conditioner discharge and rougher cell junction boxes.

First rougher lead concentrates are diverted to final concentrate, cleaner feed or regrind feed depending on grade in the cleaner circuit. Second lead rougher and scavenger lead concentrates are pump-fed to 100 mm diameter dewatering cyclones and the cyclone underflow is reground to a P80 of 10 µm to 15 µm in two 185 kW stirred media detritors (SMDs) using 3 mm ceramic beads. Lime at 30% solids w/w concentration and ligno-sulphonate solution at 10% w/v strength are added to the lead regrind mills to depress pyrite.

The reground lead concentrate and first rougher lead concentrate are pumped to the first cleaner and first cleaner concentrate is pumped to the second cleaner feed. First cleaner lead tail is open circuited to the zinc flotation conditioners with the rougher lead scavenger tail.

Copper sulphate at 10% w/v strength is added to the zinc conditioner feed pump (Pb rougher tail) hopper for the re-activation of sphalerite.

The flotation of zinc is preceded by two conditioning tanks in series each of three minute residence time for the addition of lime and the flotation collector, respectively. Frother is added in stages to the zinc second conditioner discharge and rougher flotation cell junction boxes. The zinc conditioners overflow in sequence to the zinc rougher scavenger flotation cells.

Depending on grade, first rougher zinc concentrates can be diverted to final zinc concentrate, zinc cleaner feed or zinc regrind feed. The zinc second rougher and scavenger concentrates are pump-fed to 100 mm diameter dewatering cyclones and the cyclone underflow is reground to a P80 of 10 µm to 15 µm in five 355 kW SMDs. Lime and ligno-sulphonate are added to the zinc regrind mills to depress pyrite.

The reground zinc concentrate and the first rougher zinc concentrate with regrind zinc cyclone overflow are pumped to the first cleaner bank. The zinc cleaner circuit can be operated as either a three-stage cleaner, or a cleaner and re-cleaner system as required. The first zinc cleaner tail is pumped open circuit to the final tail thickener where it is combined with pre-flotation concentrate, DMS fines spiral lights, and rougher scavenger tailing.

Each concentrate is dewatered to a minimum of 60% solids by weight in a conventional thickener.