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silica ore crushing circuit

beneficiation of iron ore

beneficiation of iron ore

Beneficiation of Iron Ore and the treatment of magnetic iron taconites, stage grinding and wet magnetic separation is standard practice. This also applies to iron ores of the non-magnetic type which after a reducing roast are amenable to magnetic separation. All such plants are large tonnage operations treating up to 50,000 tons per day and ultimately requiring grinding as fine as minus 500-mesh for liberation of the iron minerals from the siliceous gangue.

Magnetic separation methods are very efficient in making high recovery of the iron minerals, but production of iron concentrates with less than 8 to 10% silica in the magnetic cleaning stages becomes inefficient. It is here that flotation has proven most efficient. Wet magnetic finishers producing 63 to 64% Fe concentrates at 50-55% solids can go directly to the flotation section for silica removal down to 4 to 6% or even less. Low water requirements and positive silica removal with low iron losses makes flotation particularly attractive. Multistage cleaning steps generally are not necessary. Often roughing off the silica froth without further cleaning is adequate.

The iron ore beneficiation flowsheet presented is typical of the large tonnage magnetic taconite operations. Multi-parallel circuits are necessary, but for purposes of illustration and description a single circuit is shown and described.

The primary rod mill discharge at about minus 10- mesh is treated over wet magnetic cobbers where, on average magnetic taconite ore, about 1/3of the total tonnage is rejected as a non-magnetic tailing requiring no further treatment. The magnetic product removed by the cobbers may go direct to the ball mill or alternately may be pumped through a cyclone classifier. Cyclone underflows usually all plus 100 or 150 mesh, goes to the ball mill for further grinding. The mill discharge passes through a wet magnetic separator for further upgrading and also rejection of additional non-magnetic tailing. The ball mill and magnetic cleaner and cyclone all in closed circuit produce an iron enriched magnetic product 85 to 90% minus 325 mesh which is usually the case on finely disseminated taconites.

The finely ground enriched product from the initial stages of grinding and magnetic separation passes to a hydroclassifier to eliminate the large volume of water in the overflow. Some finely divided silica slime is also eliminated in this circuit. The hydroclassifier underflow is generally subjected to at least 3 stages of magnetic separation for further upgrading and production of additional final non-magnetic tailing. Magnetic concentrate at this point will usually contain 63 to 64% iron with 8 to 10% silica. Further silica removal at this point by magnetic separation becomes rather inefficient due to low magnetic separator capacity and their inability to reject middling particles.

The iron concentrate as it comes off the magnetic finishers is well flocculated due to magnetic action and usually contains 50-55% solids. This is ideal dilution for conditioning ahead of flotation. For best results it is necessary to pass the pulp through a demagnetizing coil to disperse the magnetic floes and thus render the pulp more amenable to flotation.

Feed to flotation for silica removal is diluted with fresh clean water to 35 to 40% solids. Being able to effectively float the silica and iron silicates at this relatively high solid content makes flotation particularly attractive.

For this separation Sub-A Flotation Machines of the open or free-flow type for rougher flotation are particularly desirable. Intense aeration of the deflocculated and dispersed pulp is necessary for removal of the finely divided silica and iron silicates in the froth product. A 6-cell No. 24 Free-FlowFlotation Machine will effectively treat 35 to 40 LTPH of iron concentrates down to the desired limit, usually 4 to 6% SiO2. Loss of iron in the froth is low. The rough froth may be cleaned and reflotated or reground and reprocessed if necessary.

A cationic reagent is usually all that is necessary to effectively activate and float the silica from the iron. Since no prior reagents have come in contact with thethoroughly washed and relatively slime free magnetic iron concentrates, the cationic reagent is fast acting and in somecases no prior conditioning ahead of the flotation cells is necessary.

A frother such as Methyl Isobutyl Carbinol or Heptinol is usually necessary to give a good froth condition in the flotation circuit. In some cases a dispersant such as Corn Products gum (sometimes causticized) is also helpful in depressing the iron. Typical requirements may be as follows:

One operation is presently using Aerosurf MG-98 Amine at the rate of .06 lbs/ton and 0.05 lbs/ton of MIBC (methyl isobutyl carbinol). Total reagent cost in this case is approximately 5 cents per ton of flotation product.

The high grade iron product, low in silica, discharging from the flotation circuit is remagnetized, thickened and filtered in the conventional manner with a disc filter down to 8 to 10% moisture prior to treatment in the pelletizing plant. Both the thickener and filter must be heavy duty units. Generally, in the large tonnage concentrators the thickener underflow at 70 to 72% solids is stored in large Turbine Type Agitators. Tanks up to 50 ft. in diameter x 40 ft. deep with 12 ft. diameter propellers are used to keep the pulp uniform. Such large units require on the order of 100 to 125 HP for thorough mixing the high solids ahead of filtration.

In addition to effective removal of silica with low water requirements flotation is a low cost separation, power-wise and also reagent wise. Maintenance is low since the finely divided magnetic taconite concentrate has proven to be rather non-abrasive. Even after a years operation very little wear is noticed on propellers and impellers.

A further advantage offered by flotation is the possibility of initially grinding coarser and producing a middling in the flotation section for retreatment. In place of initially grinding 85 to 90% minus 325, the grind if coarsened to 80-85% minus 325-mesh will result in greater initial tonnage treated per mill section. Considerable advantage is to be gained by this approach.

Free-Flow Sub-A Flotation is a solution to the effective removal of silica from magnetic taconite concentrates. Present plants are using this method to advantage and future installations will resort more and more to production of low silica iron concentrate for conversion into pellets.

silica ore crushing circuit - binq mining

silica ore crushing circuit - binq mining

For calcite crushing, the two-phase closed-circuit crushing process is widely The mined calcite ores in lumps are conveyed by belt conveyer to the primary calcite cr. Our crusher is a silica sand crusher and screening machine manufacturer in

21 Nov 2012 Ore crushing equipment: Ore dressing plant generally adopts three phase size return to the mill for re-grinding, it is called closed-circuit grinding. Main iron ores are magnetite, hematite, limonite, iron silicate ore, iron

25 Mar 2013 For medium scale's concentrator, general zinc ore crushing plant for sale with Three sections + 1 closed circuit procedure. aggregate, ceramic raw supplies, iron ore, gold, copper, corundum, bauxite, silica and so on.

SBM designs and supplies complete iron ore,stone and mineral crushing and screening machines, building material, silicate, refractory material and fertilizer, aside from ore grinding, suitable for both open-circuit and closed-circuit plant.

Our crusher offer Ore Crusher,Mining crusher and Material crusher,We are a Introduction of Mica crushing,The mica group of sheet silicate (phyllosilicate) minerals Silver ore crusher involved,The major gear in a primary silver crushing circuit

These are some of the aluminum, magnesium and other water-based silicate After first crushing, the material will be transferred to impact clay ore crusher or ore parts will be returned to impact clay ore crusher, thus forming a closed circuit.

Operating approximately 17 hours a day, the coarse crushing circuit comprises Primary and. Secondary Crushers that deliver crushed ore at a P80 of in to the Coarse Ore Bins in the . Dextrin (a simple starch) is also added as a silicate

The chain mill, a size reduction crusher for various applications, is available from Stedman. Chain mill Sulpher Silicon Carbide Sylvite Scrap Tungsten Carbide Sandstone. Trona Ore Talc Tungsten Ore . Circuit: Open (material from Mexico)

After the lode ore is crushed, recovery of the valuable minerals is done by one, where water is used as a medium for washing the ore to remove the clays and silica jaw crusher and after that recycled to the screening and washing circuit

This is achieved by passing the finely crushed ore over a bath of solution containing the hematite will sink and the silicate mineral fragments will float and can be removed. This circuit assures sufficient reduction of the iron ore particles.

crushing circuit - an overview | sciencedirect topics

crushing circuit - an overview | sciencedirect topics

Ore mineralogy affects the shape and size distribution of iron ore particles. In the same crushing circuit, friable ore textures tend to generate more fine materials. Ore particles of different mineralogical characteristics also tend to have different physical characteristics and are subject to different breakage mechanisms when crushed. As a result, some ore particles, such as Ore A (hard and uniform ore types, which tend to undergo brittle fracturing) in Figure 14.4, are found to be platy with angular edges, while others (including microporous textures, which tend to undergo size reduction by abrasion), such as Ore B in Figure 14.4, are often equigranular or well rounded. Hida et al. (1982) and Roller (1982) pointed out the importance of particle shape in granulation. As shown in Figure 14.4, Ores B and C have equigranular morphology and are relatively coarse with low amounts of ultrafine and intermediate particles, so both ores granulate effectively with well-developed coatings. The platy, angular morphology of Ore A does not favor the embedding of adhering fines onto the surface of its particles, however, resulting in less distinct adhering fine coatings. While Ores A and C have a similar size distribution, Ore C also contains abundant goethitic ultrafine particles, which adhere effectively to the surface of large particles and bind other fine particles together.

The pore structure of ore particles plays an important role in determining granulation efficiency, because it influences their ability to absorb and hold moisture and the amount of moisture available for granulation at certain mix moisture contents. The porosity of iron ore particles is closely related to the mineralogical characteristics of iron ore, as evidenced in Figure 14.5a. The granulation efficiency is a strong function of the amount of moisture available for granulation, which is the total moisture present less the moisture absorbed and stored in the pores. The amount of moisture required for granulation was found to vary widely with ore type (Furui et al., 1977; Litster and Waters, 1988). The CSIRO moisture saturation measurement determines the capacity of an ore to hold water before dripping out, and therefore, it reflects the porosity of an ore. Based on the authors' studies in Figure 14.5b, it is clear that ores with a higher moisture saturation value need more water to achieve good granulation efficiency. This is because intraparticle pores need to be largely filled before surface water becomes available for interparticle adhesion. Therefore, the moisture saturation value of iron ore provides a good indication of the amount of water required for effective granulation.

The first step of physical beneficiation is crushing and grinding the iron ore to its liberation size, the maximum size where individual particles of gangue are separated from the iron minerals. A flow sheet of a typical iron ore crushing and grinding circuit is shown in Figure 1.2.2 (based on Ref. [4]). This type of flow sheet is usually followed when the crude ore contains below 30% iron. The number of steps involved in crushing and grinding depends on various factors such as the hardness of the ore and the level of impurities present [5].

Jaw and gyratory crushers are used for initial size reduction to convert big rocks into small stones. This is generally followed by a cone crusher. A combination of rod mill and ball mills are then used if the ore must be ground below 325 mesh (45m). Instead of grinding the ore dry, slurry is used as feed for rod or ball mills, to avoid dusting. Oversize and undersize materials are separated using a screen; oversize material goes back for further grinding.

Typically, silica is the main gangue mineral that needs to be separated. Iron ore with high-silica content (more than about 2%) is not considered an acceptable feed for most DR processes. This is due to limitations not in the DR process itself, but the usual customer, an EAF steelmaking shop. EAFs are not designed to handle the large amounts of slag that result from using low-grade iron ores, which makes the BF a better choice in this situation. Besides silica, phosphorus, sulfur, and manganese are other impurities that are not desirable in the product and are removed from the crude ore, if economically and technically feasible.

Iron ore processing by Rio Tinto in the Pilbara region of Western Australia does not involve any chemical treatment. Flowsheets for the Brockman 2 and Paraburdoo processing plants are given in Figures 8.4 and 8.5 (Kinnel, 2013). The flowsheets are relatively simple. Dry processing involves up to three crushing circuits to produce lump and fines. Wet processing is primarily aimed at the removal of clays. Key processing variables include the following:

Crusher amps: crusher current (amps) is an indication of how hard the crusher is working. Alow bowl level, small gap setting, and high amps are a sign that the crusher could be working too hard. However, a high bowl level and low amps indicate the crusher is working too little. The crusher amps can also be affected by the hardness of the feed type.

Crusher product tons per hour: the crusher product tons per hour can be affected by the gap setting. A smaller gap setting lowers the tons per hour treated by the crusher but overall could increase the amount of material produced in the desired product size fraction. This could also minimize recirculation in the system, reducing the overall power consumption.

The sponge masses as produced by vacuum distillation have to be prepared before melting. The nine ton mass of sponge has to be crushed to about 12mm size pieces. The sponge in contact with retort wall and the push plates have a high likelihood of contamination with iron and nickel since these metals are soluble in titanium. The top of the mass may also have some contamination of iron and nickel from reaction with the radiation shield and substoichiometeric chlorides. To remove this contamination the outer skin of the sponge mass is removed by use of powered chisels. This material is downgraded from aerospace use and used in less critical applications. The sponge mass then is sliced radially to one to 5cm sections with a large guillotine or similar blade. The bottom section of the mass is removed first as this likely has the most amount of iron incorporated into the sponge. The sponge mass is removed from the working table, so this material can be segregated from the balance of the mass. At this point the mass is placed back on the table, sliced and then sent to a crushing circuit. Titanium sponge is malleable material, thus traditional mineral processing equipment such as roll or jaw crushers are not as effective as high shear shredding machines such as rotary shears or single rotor/anvil shears in preparing sponge with limited very fine particle generation.

Dust generation in the crushing process is a very important aspect of operation. Control of the dust by collection and washing of equipment on a periodic basis is very important to reduce the risks of fire in the processing of sponge. Care has to be taken to avoid working on equipment when dust present as titanium metal fires are difficult to extinguish; a class D extinguisher or rock salt are used to suppress the first. The high temperature of the fire and the low melting point of iron-titanium eutectic can result in melting of equipment, supports or piping in these plants if a fire does occur.

The core of the sponge mass has the lowest level of metal contamination. To harvest the material for applications that need low iron and low nickel levels, it is necessary to core the mass. This is done in several ways; the mass can be upended and the guillotine blade can be used to remove thick layers of outer skin, or chisels can be used to remove the outer layers. Control of the lot by separation during the crushing campaign is used to separate the high-purity products from the normal grades of sponge. Control of the nickel level in the magnesium used in the reduction is also important. Removal of as much stainless steel in piping, retorts and metal reservoirs is also important, as nickel in the magnesium will be incorporated into the sponge. Small concentrations of nickel in magnesium can take a long time to be purged from the process. Control of the quality of magnesium used for make up in the VDP process is as important, as some magnesium can be contaminated with nickel during production. Iron is not as significant an issue as its solubility in magnesium is low.

All significant characteristics of a lateritic ore likely to influence eventual prototype design must be identified prior to commencing a chemical leaching programme. Preliminary data on how the various materials are likely to respond to conventional methods of slurrying and desliming are provided by pilot-scale testwork conducted in association with drilling and sampling. Laboratory log sheets provide descriptions of the mineralogy and distribution of the other heavy minerals reporting with the coarse gold. Pre-concentration and roasting may be required for some primary ores; coarse (+200mm) gold liberated by size reduction in the crushing circuit can be recovered by conventional gravity means. Carbon in pulp (CIP) absorption by slow agitation is the most common method used in primary gold leaching circuits, however the following factors usually influence the selection of carbon in leach (CIL) processing methods for the treatment of most lateritic type gold ores:

Removal of the gold from solution at the time of dissolution minimises its exposure to organic material in the slurry, thus reducing the gold losses associated with those organics; residual gold deposits tend to be contaminated with significant quantities of organic matter and if this material is allowed to reach the leach circuit it will adsorb some of the dissolved gold and tend to block interstage screens; a high frequency screen will reject wood chips or other such debris that might remain in the slurry but regardless of what precautions are taken some contamination of the leach circuit is likely to occur.

The capital cost of a CIL process in which adsorption of gold occurs in the leach tanks is smaller than that of a CIP process; the CIP process requires an additional smaller series of CIP tanks although CIL operating costs may be slightly higher because of a larger carbon inventory.

Run-of-mine material is passed over a grizzly screen to remove oversize and trash; lateritic boulders larger than about 150mm are spalledto a size suitable for grinding. A drum type scrubber is usually preferred for breaking up lumps of clay associated with the lateritic nodules. Any residual organic trash in the product will be removed from the top deck of a double deck screen. Coarse (+2mm) solids will be discharged from the lower screen and transferred to a coarse ore stockpile as feed material to the grinding circuit. Fine undersize clayey material (2mm) scalped from the lower screen will be pumped to a bank of cyclone separators in the grinding circuit; the overflow comprising very fine material will bypass the grinding circuit and proceed directly to the leaching circuit.

Eliminating as much organic trash as possible away from the treatment plant and tailing pond water reclamation areas is a priority because of its deleterious effect on gold recovery. Although large trash is removed during the mining operations and most of the finer trash at the scrubbing-screening stage, the leach plant feed material will inevitably still contain a small amount of dilutionary material including any smaller plant fibres that have not been removed. Possible future problems such as fine screen blinding must be identified so that they can be compensated for in the leaching circuit design.

Screen sizing as an adjunct to separation in gold milling circuits is limited in its applications because vibrating screens have increasingly poor performance characteristics at smaller apertures and the flow rates are comparatively large. Mechanically operated screens are restricted to aperture sizes larger than 2mm and some form of hydraulic classification is preferred for the finer sizings. Vibrating screens are used mainly to remove trash ahead of gravity enhanced centrifugal separators and for the recovery of carbon in the leaching circuit.

Machines of the Dorr-thickener type and spiral classifiers are engineered to provide an effective pool area and overflow velocity for settling in accordance with the particular size separation requirements. However such traditional thickening device types have been replaced by metallurgical cyclones for sizing or desliming large flow volumes of slurry because they are comparatively cheap and occupy little headroom. As integral features of closed circuit grinding, metallurgical cyclones also provide the greatest opportunity for gravity concentration by recovering most of the gravimetrically recoverable gold in the cyclone underflow free of slime. The initial recovery of coarse gold minimises fragmentation and smearing onto other particles during multiple passes through the grinding circuit.

Chemical leaching testwork will usually involve continuous pilot plant testing by some selected chemical leaching process (e.g., carbon-in-leach) of batch samples of ore, that are as near as possible representative of future mill feed. Environmental protection measures and hazard costs have militated against the use of cyanides as a gold lixiviant and ammonium thiosulfate is being extensively tested as a substitute to replace cyanide in gold leaching. Bacteria are also used for the purpose but in the following treatment of the subject cyanidation processes have been preferred. Ancillary bond mill grindability testing, thickening and carbon stripping studies will be carried out as ancillary processes. The test work should indicate:

SAG mills are particularly suited to the first stage of breaking down residualtype gold ores because they always produce a very fine overflow, which can pass to the leaching circuit as a final product. In addition to the benefits provided by good operational control, SAG mills have the advantage of low capital cost, usually about 25% below that of a conventional crushing/grinding circuit. By eliminating the need for crushing, maintenance costs are also reduced. Steel ball and liner consumption is lower although these savings are somewhat offset by higher power consumption charges per tonne of ore milled.

Conventional ball or rod milling in closed circuit follows autogenous grinding. Fines from the SAG mill and the ball mill are combined with the scalp screen underflow and pumped to cyclone classifiers from which the coarse underflow will pass to jigs in a gravity concentration circuit. The presence of appreciable quantities of coarse gold in the SAG mill cyclone underflow will indicate the need for some form of gravity concentration (e.g. jigs and tables) ahead of the leaching circuit. Coarse gold requires long leach-retention times (up to 24h) to ensure complete dissolution and removing it will facilitate leach control by reducing fluctuations in the head feed. So called rusty or surface coated gold is highly refactory and may fail to dissolve completely. Metallurgical efficiency need not be high in regard to the recovery of the finer gold particles at this stage because all gravity rejects are returned to the ball mill circuit. Concentrates produced by the jigs are cleaned on a shaking table prior to smelting.

Cyclone overflow passes to the leach circuit surge tank where sodium cyanide solution and slaked lime is added to monitor and control slurry density, solution pH and cyanide concentration. A certain amount of gold is dissolved in the scrubbing and grinding circuit and will be present in the thickened overflow solution. Some recovery of this gold may be made by slurrying the ore using weak cyanide solution reclaimed from the tailing dam. Au stripping is carried out by washing the recovered carbon granules in a solution (cyanide+NaOH) at 135C, (retention time 68h). Carbon reactivating is done in a kiln at 125C to restore the active surfaces.

The kinetic activity of carbon is an important factor in determining the efficiency of carbon adsorption. The use of activated carbon for gold recovery depends initially upon the physical resistance of the carbon granules when submitted to abrasion forces within the pulp and impact forces against agitators, pipes and tank walls. The production of fines results in costly losses of both carbon and of gold carried away by the fines during operation. The adsorption characteristic of activated carbon is a compromise between gold loading capacity and adsorption characteristics on the one hand and hardness on the other. The adsorption of gold from solution onto activated carbon generally obeys the following empirical relationship:

The CIL process enables recovery of gold from slurry by mixing activated carbon granules of a coarse size with the slurry particles. Slurry is pumped from the surge tank to the first of a series of mechanically agitated leach tanks. Laboratory testwork will indicate an optimum leach retention time for the ground material. This retention time is provided by the feed surge tank and by the number of leach tanks in the series. Gold adsorbed from the solution by the carbon will then be recovered from the slurry by screening.

Carbon returning from the stripping circuit is introduced into the final leach tank and is moved through the leach circuit semi-continuously through five of the six leach tanks in a counter-current direction to the direction of the slurry flow. Carbon will advance through each stage to the second leach tank where it will be removed and transferred back to the stripping circuit. The first tank will not contain carbon and will be used only for leaching to ensure that solution gold tenor in the feed to the second leach tank is sufficiently high to enable optimum gold loading on the carbon. This in turn reduces the carbon inventory and the size of the subsequent stripping circuit.

Leached slurry leaving the final leach tank passes to a vibrating emergency screen, which is placed in the circuit to prevent catastrophic loss due to screen failure in the main launders. The aperture of the final vibrating screen will be marginally smaller than that of the launder screens in order to minimise gold losses associated with carbon fines. The carbon fines so retained will be treated separately to recover their gold content. Screened slurry will report to an agitated tailing surge tank with a 30-minute or so residence time, before being transferred to the tailing pond.

Despite environmental constraints, cyanide is usually chosen for leaching because of its almost universal applicability to all ore types. The use of other leachants (e.g., thio-urea, ammonia and chlorine) could be considered if the testwork does not respond favourably to cyanidation, or if the location is subject to environmental restraints. Gold is stripped from carbon by passing a caustic cyanide solution preheated to 90C through two or more high stripping columns containing loaded carbon on a batch process. The carbon is treated with dilute hydrochloric acid in the working column to remove carbonates that may build up due to the addition of lime to the leach circuit.

Other acid-insoluble fouling agents such as organics may also build up on the carbon and lower its activity; provision is thus made to regenerate the carbon in a small vertical oil-fired furnace at 700C. Regenerated carbon will then be air cooled and screened on a small vibrating screen to remove fine carbon prior to recycling to the leach circuit. The screen is marginally coarser than those in the leach circuit to ensure that carbon particles are effectively retained in each leach tank.

Loaded carbon recovered in the leach circuit is eluted for a pre-determined period in a series of columns, with hot 1% caustic solution containing 0.1% NaCN at atmospheric pressure. Stripped carbon is then washed with dilute hydrochloric acid and regenerated in an indirect oil-fired furnace to remove any impurities such as carbonate or organics that may build up on the carbon during leaching. Regenerated carbon is cooled and screened to remove degraded fine carbon prior to recycling to the last leach tank in the CIL process. Pregnant gold- bearing eluate from the carbon stripping section will be pumped to a pregnant solution storage tank to provide surge capacity ahead of the electrowinning circuit.

Amalgamating barrels, centrifugal separators and wet shaking tables are used to upgrade gravity plant concentrates prior to smelting. Size classification is essential for efficient separation and supplementary equipment will normally include a small vibrating screen and a cyclosizer. Vibrating screens provide more precise size separations in the dry state than any other screen type. Dimensions of width and length, which determine the main separating characteristics of conventional screens, are matched to suit the characteristics of the plant concentrates. Screen capacity is a function of width; efficiency is a function of length.

The associated heavy minerals (e.g., magnetite and ilmenite) are unlikely to cause any screening problems. The main difficulties are found with odd-shaped gold particles, angular fragments of some rock-forming minerals such as tourmaline, and fine rock particles that have held back with the heavies. Such fine particles may give rise to dust problems if dry, and clogging when damp. The Gemini Table is one type of dry shaking table for recovering gold smaller than 1mm in size. Conventional screening standards usually require surface moisture to be less than 3% by volume.

Slurries of particles larger than 200 microns may be sized on small vibrating screens to provide feed to wet shaking tables at the final stage of coarse gold concentration. Smaller sizings are made using some form of hydraulic sizing. Hydrosizing is a common, if rather inefficient procedure in dressing shed operations where sized fractions are required for feed to such units as shaking tables, belt separators and centrifugal separators. A hydrosizer comprises a series of compartments increasing in size in the direction of flow. The velocity of the surface flow is highest in the first, smallest compartment where only the largest, heaviest particles can settle out. The velocity reduces across each successive compartment thus allowing the solids to be sorted preferentially according to their settling rates in these conditions. Each compartment has a spigot discharge that may be operated normally or mechanically, depending upon the level of sophistication of the unit.

Gold-bearing eluate is pumped from the carbon strip column to a pregnant solution storage tank. The gold is recovered by passing the pregnant eluate through an electrowinning cell, which houses cathode baskets each measuring, say, 0.9m0.9m1.5m. Gold will plate out on steel wool contained within each cathode basket up to a maximum loading of about 1300g Au/kg of steel wool. The maximum flow rate through the cell may be about 15g/min. Barren solution leaving the electrowinning cell is returned to the barren solution storage tank and recycled to the carbon-strip section.

To ensure maximum recovery of gold, cathodes will be progressively moved from the back to the front end of the electrowinning cell in a counter-flow direction to solution flow. When the steel wool is fully loaded with the gold, it is removed from the cathode baskets and smelted with suitable fluxes in a small reverberatory furnace to produce a final product gold Dore bar.

grinding circuit - an overview | sciencedirect topics

grinding circuit - an overview | sciencedirect topics

Grinding circuits are fed at a controlled rate from the stockpile or bins holding the crusher plant product. There may be a number of grinding circuits in parallel, each circuit taking a definite fraction of the feed. An example is the Highland Valley Cu/Mo plant with five parallel grinding lines (Chapter 12). Parallel mill circuits increase circuit flexibility, since individual units can be shut down or the feed rate can be changed, with a manageable effect on production. Fewer mills are, however, easier to control and capital and installation costs are lower, so the number of mills must be decided at the design stage.

The high unit capacity SAG mill/ball mill circuit is dominant today and has contributed toward substantial savings in capital and operating costs, which has in turn made many low-grade, high-tonnage operations such as copper and gold ores feasible. Future circuits may see increasing use of high pressure grinding rolls (Rosas et al., 2012).

Autogenous grinding or semi-autogenous grinding mills can be operated in open or closed circuit. However, even in open circuit, a coarse classifier such as a trommel attached to the mill, or a vibrating screen can be used. The oversize material is recycled either externally or internally. In internal recycling, the coarse material is conveyed by a reverse spiral or water jet back down the center of the trommel into the mill. External recycling can be continuous, achieved by conveyor belt, or is batch where the material is stockpiled and periodically fed back into the mill by front-end loader.

In Figure 7.35 shows the SAG mill closed with a crusher (recycle or pebble crusher). In SAG mill operation, the grinding rate passes through a minimum at a critical size (Chapter 5), which represents material too large to be broken by the steel grinding media, but has a low self-breakage rate. If the critical size material, typically 2550mm, is accumulated the mill energy efficiency will deteriorate, and the mill feed rate decreases. As a solution, additional large holes, or pebble ports (e.g., 40100mm), are cut into the mill grate, allowing coarse material to exit the mill. The crusher in closed circuit is then used to reduce the size of the critical size material and return it to the mill. As the pebble ports also allow steel balls to exit, a steel removal system (such as a guard magnet, Chapters 2 and 13Chapter 2Chapter 13) must be installed to prevent them from entering the crusher. (Because of this requirement, closing a SAG mill with a crusher is not used in magnetic iron ore grinding circuits.) This circuit configuration is common as it usually produces a significant increase in throughput and energy efficiency due to the removal of the critical size material.

An example SABC-A circuit is the Cadia Hill Gold Mine, New South Wales, Australia (Dunne et al., 2001). The project economics study indicated a single grinding line. The circuit comprises a SAG mill, 12m diameter by 6.1m length (belly inside liners, the effective grinding volume), two pebble crushers, and two ball mills in parallel closed with cyclones. The SAG mill is fitted with a 20MW gearless drive motor with bi-directional rotational capacity. (Reversing direction evens out wear on liners with symmetrical profile and prolongs operating time.) The SAG mill was designed to treat 2,065t h1 of ore at a ball charge of 8% volume, total filling of 25% volume, and an operating mill speed of 74% of critical. The mill is fitted with 80mm grates with total grate open area of 7.66m2 (Hart et al., 2001). A 4.5m diameter by 5.2m long trommel screens the discharge product at a cut size of ca. 12mm. Material less than 12mm falls into a cyclone feed sump, where it is combined with discharge from the ball mills. Oversize pebbles from the trommel are conveyed to a surge bin of 735t capacity, adjacent to the pebble crushers. Two cone crushers with a closed side set of 1216mm are used to crush the pebbles with a designed product P80 of 12mm and an expected total recycle pebble rate of 725t h1. The crushed pebbles fall directly onto the SAG mill feed belt and return to the SAG mill.

SAG mill product feeds two parallel ball mills of 6.6m11.1m (internal diameterlength), each with a 9.7MW twin pinion drive. The ball mills are operated at a ball charge volume of 3032% and 78.5% critical speed. The SAG mill trommel undersize is combined with the ball mills discharge and pumped to two parallel packs (clusters) of twelve 660mm diameter cyclones. The cyclone underflow from each line reports to a ball mill, while the cyclone overflow is directed to the flotation circuit. The designed ball milling circuit product is 80% passing 150m.

Several large tonnage copper porphyry plants in Chile use an open-circuit SAG configuration where the pebble crusher product is directed to the ball mills (SABC-B circuit). The original grinding circuit at Los Bronces is an example: the pebbles generated in the two SAG mills are crushed in a satellite pebble crushing plant, and then are conveyed to the three ball mills (Mogla and Grunwald, 2008).

Hydrocyclones have come to dominate classification when dealing with fine particle sizes in closed grinding circuits (<200m). However, recent developments in screen technology (Chapter 8) have renewed interest in using screens in grinding circuits. Screens separate on the basis of size and are not directly influenced by the density spread in the feed minerals. This can be an advantage. Screens also do not have a bypass fraction, and as Example 9.2 has shown, bypass can be quite large (over 30% in that case). Figure 9.8 shows an example of the difference in partition curve for cyclones and screens. The data is from the El Brocal concentrator in Peru with evaluations before and after the hydrocyclones were replaced with a Derrick Stack Sizer (see Chapter 8) in the grinding circuit (Dndar et al., 2014). Consistent with expectation, compared to the cyclone the screen had a sharper separation (slope of curve is higher) and little bypass. An increase in grinding circuit capacity was reported due to higher breakage rates after implementing the screen. This was attributed to the elimination of the bypass, reducing the amount of fine material sent back to the grinding mills which tends to cushion particleparticle impacts.

Circulation of material occurs in several parts of a mineral processing flowsheet, in grinding and flotation circuits, for example, as well as the crushing stage. In the present context, the circulating load (C) is the mass of coarse material returned from the screen to the crusher relative to the circuit final product (or fresh feed to the circuit), often quoted as a percentage. Figure 8.2 shows two closed circuit arrangements. Circuit (a) was considered in Chapter 6 (Example 6.1), and circuit (b) is an alternative. The symbols have the same meaning as before. The relationship of circulating load to screen efficiency for circuit (a) was derived in Example 6.1, namely (where all factors are as fractions):

The circulating load as a function of screen efficiency for the two circuits is shown in Figure 8.3. The circulating load increases with decreasing screen efficiency and as crusher product coarsens (f or r decreases), which is related to the crusher set (specifically the closed side setting, c.s.s.). For circuit (a) C also increases as the fresh feed coarsens (n decreases), which is likely coming from another crusher. In this manner, the circulating load can be related to crusher settings.

In industrial grinding process, in addition to goal of productivity maximization, other purposes of deterministic grinding circuit optimization have to satisfy the upper bound constraints on the control variables. We know that there lies a tradeoff between the throughput (TP) and the percent passing of midsize classes (MS) from the previous work of Mitra and Gopinath,2004. In deterministic optimization formulation, there are certain parameters which we will assume them as constant. But, in real life that may not be case. There are such six parameters in our industrial grinding process which are R, B, R, B are the grindability indices and grindability exponents for the rod mill (RMGI) and the ball mill (BMGI); and P, S are the sharpness indices for the primary (PCSI) and secondary cyclones (SCSI). These parameters are treated as constant in deterministic formulation. As they are going to be treated as uncertain parameters in the OUU formulation. These parameters are assumed uncertain because most of them are obtained from the regression of experimental data and thus are subject to uncertainty due to experimental and regression errors. In the next part of the section, we consider them as fuzzy numbers and solve the OUU problem by FEVM. In FEVM formulation, the uncertain parameters are considered as fuzzy numbers and the uncertain formulation is transformed into the deterministic formulation by expectation calculations for both objective function and constraints. So, the converted deterministic multi-objective optimization problem is expressed as:

Another spinning batch concentrator (Figure 10.27), it is designed principally for the recovery of free gold in grinding circuit classifier underflows where, again, a very small (<1%) mass pull to concentrate is required. The feed first flows up the sides of a cone-shaped bowl, where it stratifies according to particle density before passing over a concentrate bed fluidized from behind by back-pressure (process) water. The bed retains dense particles such as gold, and lighter gangue particles are washed over the top. Periodically the feed is stopped, the bed rinsed to remove any remaining lights and is then flushed out as the heavy product. Rinsing/flushing frequency, which is under automatic control, is determined from grade and recovery requirements.

The units come in several designs, the Semi-Batch (SB), Ultrafine (UF), and i-Con, designed for small scale and artisanal miners. The first installation was at the Blackdome Gold Mine, British Columbia, Canada, in 1986 (Nesset, 2011).

These two batch centrifugal concentrators have been widely applied in the recovery of gold, platinum, silver, mercury, and native copper; continuous versions are also operational, the Knelson Continuous Variable Discharge (CVD) and the Falcon Continuous (C) (Klein et al., 2010; Nesset, 2011).

To liberate minerals from sparsely distributed and depleting the ore bodies finer grinding than generally obtained by the conventional Rod Mill Ball Mill grinding circuits is needed. Longer grinding periods in the conventional milling processes prove too expensive mainly due to large power consumption. Stirrer mills have been tried in mineral industry with considerable success and have therefore been increasingly used. In this chapter, the theories involved in the design and operation of these mills, as established till now, are explained. Further theoretical studies and designs of the mills are still in progress for a better understanding and improved operation. Presently, the mills have been proved to be economically viable and the mineral of interest conducive to improved recovery and grade.

IMP Technologies Pty. Ltd. has recently tested a pilot-scale super fine crusher that operates on dry ore and is envisaged as a possible alternative to fine or ultra-fine grinding circuits (Kelsey and Kelly, 2014). The unit includes a rotating compression chamber and an internal gyrating mandrel (Figure 6.13). Material is fed into the compression chamber and builds until the gyratory motion of the mandrel is engaged. Axial displacement of the compression chamber and the gyratory motion of the mandrel result in fine grinding of the feed material. In one example, a feed F80 of 300m was reduced to P80 of 8m, estimated to be the equivalent to two stages of grinding. This development is the latest in a resurgence in crushing technology resulting from the competition of AG/SAG milling and the demands for increased comminution energy efficiency.

The iron oxide crystal grains in most iron ores are not evenly distributed in size. Spiral separators can therefore be used to take out the coarser iron oxide grains in the primary grinding circuit to save grinding energy and help achieve a higher iron recovery. Figure 9.14 presents a typical flow sheet for processing an oxidized ore containing about 30% Fe using a combination of spiral and SLon magnetite separators and reverse flotation. This ore is mainly composed of hematite, magnetite, and quartz, and the iron oxide crystals range in size from 0.005 to 1.0mm with an average size of about 0.05mm. The average size of the quartz crystals is approximately 0.085mm.

In the primary grinding stage of the flow sheet in Figure 9.14, the ore is first ground down to about 60% -75m and then classified into two size fractions, a coarse size fraction and a fine size fraction. The coarse size fraction is treated with spiral separators to recover part of the final iron ore concentrate. Then, drum LIMS and SLon magnetic separators are used to reject some of the coarse gangue minerals as final tailings. The magnetic products from the LIMS and SLon are sent back to the secondary ball mill for regrinding, and the milled product returns to the primary cyclone classifier.

The fine size fraction is about 90% -75m and is processed using drum LIMS separators and SLon magnetic separators in series to take out the magnetite and hematite, respectively. The magnetic products from the magnetic separators are mixed to generate the feed for reverse flotation to produce another component of the final iron ore concentrate.

The key advantage of this flow sheet lies in the fact that the spirals and SLon magnetic separators take out about 20% of the mass of the final iron concentrate and about 20% of the mass of the final tailings, respectively, from the coarse size fraction. This greatly reduces the masses being fed to the secondary ball mill and reverse flotation, thereby greatly reducing the total processing cost. From the plant results for this flow sheet, an iron concentrate containing 67.5% Fe could be produced from a run-of-mine ore containing 30.1% Fe, at a mass yield to the iron concentrate of 34.9%, an iron recovery of 78.0%, and a tailings grade of 10.2% Fe.

The first step of physical beneficiation is crushing and grinding the iron ore to its liberation size, the maximum size where individual particles of gangue are separated from the iron minerals. A flow sheet of a typical iron ore crushing and grinding circuit is shown in Figure 1.2.2 (based on Ref. [4]). This type of flow sheet is usually followed when the crude ore contains below 30% iron. The number of steps involved in crushing and grinding depends on various factors such as the hardness of the ore and the level of impurities present [5].

Jaw and gyratory crushers are used for initial size reduction to convert big rocks into small stones. This is generally followed by a cone crusher. A combination of rod mill and ball mills are then used if the ore must be ground below 325 mesh (45m). Instead of grinding the ore dry, slurry is used as feed for rod or ball mills, to avoid dusting. Oversize and undersize materials are separated using a screen; oversize material goes back for further grinding.

Typically, silica is the main gangue mineral that needs to be separated. Iron ore with high-silica content (more than about 2%) is not considered an acceptable feed for most DR processes. This is due to limitations not in the DR process itself, but the usual customer, an EAF steelmaking shop. EAFs are not designed to handle the large amounts of slag that result from using low-grade iron ores, which makes the BF a better choice in this situation. Besides silica, phosphorus, sulfur, and manganese are other impurities that are not desirable in the product and are removed from the crude ore, if economically and technically feasible.

While used sometimes on final concentrates, such as Fe concentrates, to determine the Blaine number (average particle size deduced from surface area), and on tailings for control of paste thickeners, for example, the prime application is on cyclone overflow for grinding circuit control (Kongas and Saloheimo, 2009). Control of the grinding circuit to produce the target particle size distribution for flotation (or other mineral separation process) at target throughput maximizes efficient use of the installed power.

Continuous measurement of particle size in slurries has been available since 1971, the PSM (particle size monitor) system produced then by Armco Autometrics (subsequently by Svedala and now by Thermo Gamma-Metrics) having been installed in a number of mineral processing plants (Hathaway and Guthnals, 1976).

The PSM system uses ultrasound to determine particle size. This system consists of three sections: the air eliminator, the sensor section, and the electronics section. The air eliminator draws a sample from the process stream and removes entrained air bubbles (which otherwise act as particles in the measurement). The de-aerated pulp then passes between the sensors. Measurement depends on the varying absorption of ultrasonic waves in suspensions of different particle sizes. Since solids concentration also affects the absorption, two pairs of transmitters and receivers, operating at different frequencies, are employed to measure particle size and solids concentration of the pulp, the processing of this information being performed by the electronics. The Thermo GammaMetrics PSM-400MPX (Figure 4.18) handles slurries up to 60% w/w solids and outputs five size fractions simultaneously.

Other measurement principles are now in commercial form for slurries. Direct mechanical measurement of particle size between a moving and fixed ceramic tip, and laser diffraction systems are described by Kongas and Saloheimo (2009). Two recent additions are the CYCLONEtrac systems from CiDRA Minerals Processing (Maron et al., 2014), and the OPUS ultrasonic extinction system from Sympatec (Smith et al., 2010).

CiDRAs CYCLONEtrac PST (particle size tracking) system comprises a hardened probe that penetrates into the cyclone overflow pipe to contact the stream and effectively listens to the impacts of individual particles. The output is % above (or below) a given size and has been shown to compare well with sieve sizing (Maron et al., 2014). The OPUS ultrasonic extinction system (USE) transmits ultrasonic waves through a slurry that interact with the suspended particles. The detected signal is converted into a particle size distribution, the number of frequencies used giving the number of size classes measured. Applications on ores can cover a size range from 1 to 1,000m (Smith et al., 2010).

In addition to particles size, recent developments have included sensors to detect malfunctioning cyclones. Westendorf et al. (2015) describe the use of sensors (from Portage Technologies) on cyclone overflow and underflow piping. CiDRAs CYCLONEtrac OSM (oversize monitor) is attached to the outside of the cyclone overflow pipe and detects the acoustic signal as oversize particles (rocks) hit the pipe (Cirulis and Russell, 2011). The systems are readily installed on individual cyclones thus permitting poorly operating units to be identified and changed while allowing the cyclone battery to remain in operation. Figure 4.19 shows an installation of both CiDRA systems (PST, OSM) on the overflow pipe from a cyclone.

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