6. METALLURGICAL PROCESSING

6.5 Base Metals

6.5.1 Rosh Pinah

Original plant feed capacity was 570ktpa. This was increased to 670ktpa by plant modifications and upgrading in 1981 and during the latter part of the 1990s. These included the installation of column and tank flotation cells, and a secondary and a regrind mill. While the upgrades were motivated on efficiency improvements, some additional capacity was achieved. More recently, modifications were made to mill liners and mill ball load was increased, which increased mill power draw and hence capacity. Present plant capacity is around 750 – 800ktpa.

Ore is mined from a number of orebodies and a number of different ore types are identified including siliceous and non-siliceous carbonates, carbonaceous shales, etc. The ore contains about 2.5% lead (range 2% – 4%) as galena, about 10% zinc as sphalerite and 2.5% – 8% pyrite. It also contains minor copper and silver. Mined ore is crushed in three stages of crushing, milled and fed to a two-stage flotation circuit to recover lead and zinc. The lead and zinc concentrates are de-watered and stockpiled for dispatch to customers. The tailing is thickened and pumped to a slimes dam.

Mined ore is crushed to –150mm in an underground primary jaw crusher. The ore is brought to surface on a conveyor belt which discharges to a primary double-deck screen (45mm and 25mm top and bottom decks, respectively). Total oversize (+25mm) is fed to a secondary (Nordberg) cone crusher, in closed circuit with the primary screen. The screen undersize (–25mm) is conveyed to the primary stockpile, which has a live capacity of 6kt.

Ore is withdrawn from the stockpile by means of three vibrating feeders and conveyed to a double-deck secondary screen (19mm and 11mm top and bottom decks). In the case of both screens, the purpose of the top deck is simply to protect and reduce the load on the bottom deck. Plus 9mm oversize is fed to a tertiary crusher, a 66-inch Nordberg gyradisc, in closed circuit with the screen.

Screen undersize is conveyed to twin stockpiles with a total live capacity of 12,000t.

Ore is withdrawn from the stockpiles via vibrating feeders (three under each of the stockpiles) and conveyed to a primary, 12ft diameter. x 12ft Osborne Marcy ball mill. The mill is rubber-lined, has a grate discharge and is fitted with a 1,000kW motor. Cyanide (to depress zinc and pyrite in the lead flotation circuit) and a xanthate promoter (to float the lead) are added to the mill feed. Mill discharge is classified in two stages of cyclones. Primary cyclone underflow is returned to the mill, while secondary underflow is directed to the mill discharge sump. Secondary cyclone overflow is de-watered in a cyclone cluster, the overflow of which is also returned to the mill sump.

The cluster cyclone overflow may either be pumped to the secondary mill for further grinding or to the lead flotation conditioners. Present grind is about 80% passing 106m.

The secondary mill is a 5ft diameter x 10ft variable speed drive mill with a 100kW motor. The primary mill product is classified in a cyclone. Underflow is the secondary mill feed; overflow is directed to the mill discharge sump. Mill discharge is pumped to the lead conditioners.

The milling and flotation circuits are controlled by a Proscon PLC which incorporates a Minovex expert system for mill control, supported by an on-line Courier XRD analyser. The primary and secondary mills are wellinstrumented with a variety of mass, flow, density, pressure, power and level transmitters (as indicated on the flowsheet) which feed into the expert control system. The Courier analyser provides lead, zinc, iron, copper and silver analyses of the float feed, lead tail, lead and zinc concentrates and final tails. Samples for the Courier are taken by automatic samplers, which also provide samples for lab analysis. Operator inputs to the system include:

High and low limits for lead and zinc in feed (typically 4.5 and 1.5tph for lead and 18 and 10tph for zinc);
Plus 106m and –38m fractions in the milled product;
New mill feed, between 100 and 85tph; and
New mill feed, between 100 and 85tph; and

The expert system then controls the milling circuit to maximise tonnage (and consequently lead and zinc production) within the set-point limits and rules of the system. Typically with a low grade ore, tonnage would be increased towards the high limit, while with high grade ores metal input tonnage would become controlling.

The expert system was only commissioned a few months ago, but appears to be functioning successfully.

Slurry from the lead conditioner, to which frother has been added, is fed to three banks of conventional rougher and scavenger flotation cells. Flotation pH is 8.5 – 9.0. Rougher concentrate is cleaned in a column cell. Scavenger concentrate and cleaner tail is returned to the conditioner. The first cleaner concentrate (second and third lead cleaner cells are not presently in operation) is thickened, de-watered on a belt filter and discharged to the lead drying pad. From where it is loaded for trucking to Aus and on by rail to port for export to Walvisbay.

Scavenger tails are thickened and pumped to the zinc float conditioners. Concentrate and tail thickener overflow water is recycled to the mills and the lead circuit for re-use.

The lead tail slurry is conditioned with a xanthate promoter, copper sulphate to reactivate the zinc, lime to a pH of 10.5 – 11.0 and frother. zinc rougher flotation is carried out in an Outokumpu tank cell. The rougher concentrate is cleaned in a first cleaner tank cell and recleaned in a second cleaner column cell. Both cleaner tails are recycled to the zinc conditioners. Rougher tails are pumped to a bank of conventional rougherscavenger cells. Concentrate is presently returned to the rougher tank cell. The regrind mill is not presently in circuit but may be used to regrind either or both of the rougher-scavenger concentrates.

The second cleaner concentrate is thickened, filtered on a belt filter and discharged to the zinc drying pad for dispatch to Aus by road and on to Zincor by rail.

Scavenger tails are thickened in one thickener and pumped to the tailings dam. Flocculant is used occasionally to aid settling.

Zinc concentrate and tailings thickener overflow is recycled and used for dilution in the zinc circuit. No water is returned from the tailings dam. The tailings are discharged at a L : S ratio of about 1 : 1, so plant fresh water consumption is about 1t per ore tonne treated. The ore contains small quantities of gold (0.1 – 0.3g/t) and 30 – 80g/t of silver. A Knelson concentrator was installed to investigate gold recovery, but this is not in operation. About 50% of the silver reports to the lead concentrate, 30% to the zinc concentrate and the balance reports to tails.

6.5.2 Zincor

The plant is split into three operating sections, comprising roasting (concentrate receiving, roasting and acid plants), the zinc plant (neutral leaching, purification, electrowinning and smelting and casting) and the recovery section (residue treatment and effluent treatment). Each section falls under a section manager who has reporting to him a plant superintendent responsible for production and a section engineer responsible for maintenance.

There are a number of service departments providing support to the production units. These include:

Engineering services, providing electrical and instrumentation services, workshop facilities and such functions as rigging, scaffolding and cranes;
Metallurgical services, who are engaged in research and development and project implementation, metal accounting and production forecasting;
Laboratory, which provides analytical services; and
SHEQ, which provides safety, health, environmental and quality assurance support services.

Concentrate Receiving: Concentrate from various sources is received by rail (some lots by road), off-loaded and stored separately in bunkers. The various feedstocks are blended and stored in bins, from which the roasters are fed. The prime reason for blending is to control Cu+Pb+SiO2 content of the feed to the roasters. Ideally this should be below 5% – 6%. Above this level, there is a danger of calcine sintering in the roaster bed. The 2003 and 2004 averages were 4.7% and 5.2%. In the early months of 2005, the average has been around 5.5%. A magnesium pre-leach plant for strong acid leaching of dolomite from concentrate ahead of roasting has been re-commissioned. Rosh Pinah concentrate is treated in this circuit because of its high dolomite (Ca-Mg carbonate) content. Magnesium increases electrolyte conductivity, adversely affecting electrowinning efficiency and capacity.

Roaster-Acid Plant: This plant has two streams, each comprising two fluid bed roasters and an acid plant. Line 1 has the smaller roasters (noted earlier) and the smaller acid plant. The relative capacities of the two lines are about 1/3 : 2/3. The concentrate, containing up to 10% moisture, is fed via conveyors, rotary feeders and slingers which feed the roasters. Roasting takes place at ~950°C. Oxygen is injected into the roasters to increase kinetics and therefore capacity. The SO₂ gas produced exits the roasters, is cleaned and cooled and then fed to the acid plants where it is converted to sulphuric acid. Acid production capacity is about 550tpd. The calcine is cooled, dry milled and transported to the calcine storage tanks.

Neutral Leach: The zinc oxide (contained in the calcine) is leached with spent electrolyte in the ‘neutral’ leach circuit (pH 4.5 – 5). Purchased oxide concentrate is also leached in this circuit, depending on its composition. If halide content is excessive, the oxide would enter via the roasters to remove the halides. The leached slurry is thickened and the thickener underflow is pumped to the residue treatment (recovery) plant for more intensive leaching and zinc recovery. The thickener overflow solution, containing the leached zinc, is purified ahead of electrowinning.

Solution Purification: The impure zinc sulphate solution is subjected to two stages of purification. In the first, copper, cobalt and nickel is removed by precipitation with zinc dust and As₂O₃. In the second stage, cadmium is removed, also by way of precipitation with zinc dust and copper sulphate. The hot, purified zinc sulphate solution (steam heated at various points in the process to keep gypsum in solution) is pumped to storage ahead of electrowinning. The Cu-Co-As precipitate and the cadmium precipitate after further upgrading, are stored for sale.

Electrowinning: The hot purified solution is passed through cooling towers in order to precipitate gypsum, which is removed in a thickener and dewatered on a sieve bend. The gypsum is disposed of via the effluent treatment plant. Gypsum removal is critical to prevent scaling of pipelines as the solution cools after purification. It is also important to maintain temperature in the impure hot solution, through purification, until it reaches the cooling towers, for a similar reason. The cold solution is fed to the electrowinning circuit via storage/surge tanks where zinc is plated onto aluminum cathodes. Lead alloy anodes are used. A strontium carbonate salt is added to the electrolyte to precipitate impurities, in particular manganese. Sludge removed from the cells, rich in MnO₂, is sent to neutral leach. The cell house contains 42 banks of cells, each bank comprising 12 cells. Zinc is manually stripped from the cathodes once per day. These are transported to the melt house. Spent electrolyte is circulated to the neutral leach as well as to the residue treatment plant. Excess solution is recirculated to electrowinning. Spent electrolyte may also be bled from the circuit to maintain the overall solution balance and for Mg control.

Smelting and Casting: Cathode zinc is melted in four induction furnaces with an ammonium chloride flux, to increase the dross layer fluidity, which improves furnace performance. One and 2t jumbo ingots and 25kg ingots are produced. In addition, aluminium and lead-zinc alloy ingots are produced to meet customer requirements. Dross skimmed off the furnaces is sold; excess is returned to the roasters. Metallics from the dross and scrap zinc is converted to zinc dust and used in the solution purification circuit.

Residue Treatment: The residue treatment plant retreats the neutral leach residue to recover residual zinc and to separate dissolved iron. The neutral leach residue is subjected to a series of successive leaches with spent electrolyte at higher temperatures and stronger acid conditions. The leached slurry from each stage is thickened. Thickener overflow from each stage is returned to the previous stage, while underflow advances to the following leach stage. Final thickener underflow is filtered and washed on a belt filter and the final Pb/Ag residue disposed of on the tailings dam. The first stage overflow is treated with ZnO (calcine) to precipitate iron (as an iron-hydroxy-sulphate). The resulting iron rich slurry is thickened. The zinc rich thickener overflow solution is returned to the neutral leach while the underflow is filtered and washed on a second belt filter. The diluted belt filter filtrate is recycled within the circuit, the iron residue is disposed of to the slimes dam.

Effluent Treatment and Disposal: A number of effluent streams are treated in the effluent section before disposal:

Spent electrolyte bleed is treated with slaked lime (to a pH of 6.0 – 7.0) to precipitate zinc. The resulting slurry is filtered. Filtrate containing magnesium and halides is pumped to the effluent treatment plant. The solids are repulped with spent electrolyte (at a pH of ~4.5) to leach zinc and added to the residue treatment process stream. The neutral leach plant run-off water is collected in an emergency dam and treated in the effluent treatment plant as is weak acid bleed.
In effluent treatment, the various waste streams are neutralised (at pH 9.0 – 10.0) with limestone and slaked lime to precipitate other heavy metals. The resulting slurry is thickened. The thickened solids are disposed of to a demarcated section on the tailings dam. Thickener overflow is pumped to an evaporation system on the tailings dam. The lead/silver and iron residues are also disposed of separately on the tailings dam. Dam decant runs into a “penstock return water dam” and is returned to the effluent treatment plant. Some penstock return water is also pumped to a spray evaporation area adjacent to the penstock return water dam.

6.5.3 Chifeng

The nominal design capacity of the Chifeng smelter originally was 21ktpa slab zinc with 36ktpa sulphuric acid. Since 1998 the smelter has consistently exceeded the design capacity for both zinc slab and sulphuric acid production. The capacity of Phase I is now stated to be 23.5ktpa zinc. This highlights the robustness of the chosen process, which is supported by an original conservative design.

Baiyinuoer Mine started with the construction of a roaster and sulphuric acid plant at Lindong. With the intervention of Kumba this facility was taken over to form part of Chifeng.

The Lindong design as proposed by the Yangzhou Chemical Engineering Company is conservatively based on producing the equivalent of 30ktpa of contained zinc in calcine and 47ktpa of sulphuric acid.

Roasting: Concentrate roasting is performed via fluidized bed roasters roasting according to the standard ENFI design, or variants of it, and the Lurgi fluidised bed roaster concept. The major difference in the above roasting concepts being in the hearth and tuyere design, where the ENFI concept makes use of a bubble cap design whilst the Lurgi design makes use of small annular perforations in the bed of the hearth. All the concepts however basically adhere to the fluidisation velocity criterion for zinc concentrates of 0.4 – 0.7m/s in the hearth area, and 0.25 – 0.4m/s in the freeboard zone. The above figures imply air rates per hearth area of between 500 and 550Nm³/m²/hr. Typical dry solids design throughputs for ENFI roasters are 5.5 dry t/m²/day, which for the latter air fluidising rates equates 90 to 100 dry t/Nm³ air.

Zinc concentrate roasting for Chifeng and Lindong is preferentially conducted at temperatures in the region of 870º – 920ºC. This is a precautionary step governed by the high silica levels of the concentrates treated. Temperature control is executed by cooling coils and feed control. As a backup plan and when problems occur, direct water injection into the bed is used. At the fluidising air rates and operating temperatures that the roasters operate at, the typical ratio of calcine to bed material discharge is 60% – 65% and calcine to fume dust carryover to waste heat boiler are between 40% – 35%.

Fume/Particulate removal from the gas stream exiting the roaster is executed via conventional methods. Coarser particulates are first recovered in the waste heat boiler, cyclones and electrostatic precipitators. Finer air-borne particulates are scrubbed from the gas in a humidifying tower or venturi-scrubber, followed by a gas cooling tower and electric demister (mist precipitator). In this gas scrubbing circuit a weak acid solution is continuously circulated across the humidifying/venturi-scrubber and gas cooling towers to prevent solid and impurity carry-over to the sulphuric acid. This circulating weak acid stream undergoes heating in the process and requires to be cooled with cold water through a plate heat exchanger. A constant weak acid bleed stream is removed to the wastewater treatment plant in order to control a build-up of impurities.

Calcine Handling: For both Chifeng and Lindong the roaster calcine has to be cooled before being transferred to a storage or loading facility.

The process typically consists of cooling of roaster bed calcine in a water-cooled cooling drum before it is transferred to a pneumatic transfer vessel, in the case of Chifeng, and to a storage shed in the case of Lindong. The calcine/fume recovered in the waste heat boiler, cyclone and electrostatic precipitators is collected and conveyed via redler/conveyors to a common collecting water-cooled redler conveyor. All conveyors are enclosed units in order to prevent material spillages and losses.

At the Lindong site calcine is conveyed to a covered storage facility where it is bagged and loaded onto trucks for transport to the Chifeng site. On arrival at the Chifeng site the trucks are offloaded via an overhead crane and the bagged calcine stored in a new storage shed ahead of the main calcine storage silos.

The calcine collected from the Chifeng roasting facility is pneumatically transferred to the main calcine storage silos at the head of the leaching section, whilst the Lindong calcine is first emptied and milled in a dedicated milling circuit before being pneumatically conveyed to the same common calcine storage silos.

From the main storage silos the Chifeng and Lindong calcine mix is transferred to intermediate calcine storage silos located within the leach building from where it is added normally only at neutral leach stage and sometimes to stabilise the process also at the pre-neutralisation stage.

Acid Production: Acid production typically follows the double contact – double absorption process across a converter with vanadium pentoxide catalyst beds. A variant of this process is a possibility where a triple contact – double absorption system may be adopted. In the latter, the gas passes through three beds of V₂O₅ catalyst (versus two for the former) before passing through the inter-pass absorption tower. From the inter pass absorption tower the gas then passes through two beds of catalyst before being contacted with 98% H₂SO₄ in the final absorption tower.

Gas temperatures into the converter are maintained at the required and optimum temperatures for SO₂ conversion by contacting hot and cold gas streams counter currently in insulated vertical shell and tube heat exchanger arrangements. For cold start-ups an in-line electrical primary pre-heater is found in the ducting prior to the first entry point of the converter. This pre-heater is assisted by a secondary pre-heater to aid falling gas temperatures prior to entry of the last two contacting stages. The secondary pre-heater is generally also utilised for cold start-ups but may also be required during normal operations to raise falling converter temperatures.

Air and gas transport across the roasting to contact sections of the acid plant may be accomplished by various blower/fan arrangements. Typically a “push-pull” concept is adopted across the gas-train with the roaster blower “pushing” and an intermediate fan located prior to the contacting section of the Acid plant “pulling” the gas through the waste heat boiler, electrostatic precipitator and gas scrubbing sections and then “pushing” it on to the contact plant. This operation may be conducted with a roaster blower working either in tandem with a hot gas fan (located after the electrostatic precipitator) and a main SO₂ blower located in front of the contacting circuit (drying tower), or only with a main SO₂ blower (without a hot gas fan). Only a main SO₂ blower without a hot gas fan in both current acid plants is used.

Leaching: Leaching of the calcine is conducted using the conventional neutral leach and hot acid leaching stages with an intermediate pre-neutralisation step located between the two for acidity control prior to iron precipitation.

The Chifeng process operates a neutral leach at typical pH levels encountered in the industry. Its single hot acid leach differs from the norm in that is a single step operating at moderate terminal acidities (60 – 80 g/l H₂SO₄) and temperatures (80º – 85ºC).

The main reason for this is to prevent unmanageable circulating loads of silica building up within the leach circuit due to the high silica input from the raw materials. Furthermore this operating regime requires very little to no calcine addition to the pre-neutralisation stage for acidity control. The lack of process control and instrumentation complicates the operation of this pre-neutralisation stage and periodically leads to troublesome physical liquid/solid separation and throughput problems.

A secondary reason the chosen flow sheet is pursued is that there is no need to upgrade the Pb/Ag residue for silver recovery, as the silver input in raw materials is low. A slight sacrifice in zinc extraction is made at the option of operating a more forgiving and operator friendly circuit. At this stage the option of expanding the hot acid leaching stages is not justified as the operating costs and process throughput problems will outweigh the potential extra zinc recovery.

Iron Removal: Iron removal is in the form of ammonium jarosite based on the low contaminant jarosite process proposed in the late 1970s by Electrolytic Zinc Australia (Pasminco). The low contaminant claim comes from the fact that no calcine is added to the stage during iron precipitation, hence the precipitate should be very low in contaminants such as zinc, cadmium, copper and lead. Trace levels of heavy metal contaminants would then be associated with co-precipitated species and occluded soluble species.

The Chifeng process makes use of ammonium hydro carbonate as the source of alkali for both acid neutralization as well as ammonium jarosite precipitation and operates the process as close to 95ºC as possible and within the prescribed acidity levels.

The process operates smoothly with the required amount of iron being removed easily within the prescribed residence time. An easily filterable product is produced at the end of the reaction period.

The major weakness the process suffers from is the precipitate contamination caused though solids carryover in the preceding pre-neutralisation stage.

Purification: Purification technology adopts a batch purification technique for the removal of copper and cobalt with traces of nickel in a common precipitate, followed by cadmium removal in low-grade cement. For the removal of cobalt the process adopts the hot zinc dust/arsenic technique with the arsenic being added as sodium arsenate after dissolving arsenic trioxide with caustic soda solution. In the cadmium scavenging process, zinc dust additions are made together with copper and ferrous sulphate, and potassium permanganate solution. The emphasis in the batch processing approach is placed on the ability to produce a first-pass high quality solution with extremely low levels of cobalt and cadmium. This quality comes at a high price in terms of reagent consumptions – particularly zinc dust powder. The lack of proper process control together with the high zinc dust dosage renders low-grade primary cakes in terms of cobalt, copper and cadmium. These primary cakes also contain high levels of basic and metallic zinc salts. Both the primary purification cakes undergo a mild acid leach/wash to recover the bulk of the zinc as well as to upgrade the by-product metal contents in a subsequent secondary treatment outside of the main process flow for zinc production.

By-Products Production: The by-products production process entails the production of three upgraded products of which two can be sold (containing copper and cobalt), and the third (cadmium), which is stockpiled. In the treatment process the combined primary purification filter cakes are firstly exposed to a mildly acidic leach/wash, which renders most of the cadmium and cobalt soluble but leaves the bulk of the copper precipitate intact. The solids recovered from this first stage represent the upgraded copper residue. In subsequent processing the dissolved cadmium is first cemented out with zinc dust powder under controlled zinc dust additions to yield high-grade cadmium cement. This step is followed by the controlled precipitation of the dissolved cobalt with arsenic trioxide and zinc dust powder to produce an upgraded cobalt residue.

Electrowinning: In the process of electrolysis, standard industry operating techniques are generally used within a cell house operating with a 24-hour plating cycle with manual cathode stripping. A conservative design in terms of operational current density is applied upfront.

The cell house configuration adopts tight cell electrode packing with small inter-electrode gap, which enables the operation to take full advantage of low electrolyte resistances and resultant low cell voltages and power consumption. There are two tank houses with two banks each of 57 cells containing 46 cathodes and 47 anodes. The current efficiency is 86%.

Casting: Casting of the zinc cathode into zinc slab adopts standard induction furnace and continuous casting belt technology.

Zinc Metal Supply: There are currently 25 different mines supplying concentrate to Chifeng. No formal toll treatment contracts are in place. There is a format on which payment is made. The compensation to each mine is negotiated every month. Based on current market conditions, a base line zinc price equivalent to ±65% of the zinc price is determined. Concentrates with >50% zinc content are fully compensated for with a RMB20/t deduction for every 1% below 50%. For concentrates containing 45% and less zinc RMB50/t is subtracted for every 1% less than 45% zinc. Concentrates containing 40% and less zinc are generally not accepted but if accepted RMB100/t is subtracted for every 1% less than 40% zinc. A penalty is imposed for CaO content of more than 0.5% at a rate of RMB60/t.