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flotation cell host

flotation mining equipment | apt trifloat appropriate process technologies | mineral processing plants

flotation mining equipment | apt trifloat appropriate process technologies | mineral processing plants

Host of benefits in the design; better mixing in slurry applications, resulting in greater efficiency for reactions and a greater uniformity within the tank. This results in fast flotation kinetics and better recoveries over conventional systems.

The cells can be arranged in various configurations depending on the mineral, throughput and float characteristics of the operation, allowing a TriFloat tank to be used as a complete rougher, cleaner and scavenger system, or for any portion of this circuit.

flotation cells

flotation cells

More ores are treated using froth flotation cells than by any other single machines or process. Non-metallics as well as metallics now being commercially recovered include gold, silver, copper, lead, zinc, iron, manganese, nickel, cobalt, molybdenum, graphite, phosphate, fluorspar, barite, feldspar and coal. Recent flotation research indicates that any two substances physically different, but associated, can be separated by flotation under proper conditions and with the correct machine and reagents. The DRflotation machine competes with Wemco and Outotec (post-outokumpu) flotation cells but are all similar is design. How do flotation cells and machinework for themineral processing industry will be better understood after you read on.

While many types of agitators and aerators will make a flotation froth and cause some separation, it is necessary to have flotation cells with the correct fundamental principles to attain high recoveries and produce a high grade concentrate. The Sub-A (Fahrenwald) Flotation Machines have continuously demonstrated their superiority through successful performance. The reliability and adaptations to all types of flotation problems account for the thousands of Sub-A Cells in plants treating many different materials in all parts of the world.

The design of Denver Sub-A flotation cells incorporates all of the basic principles and requirements of the art, in addition to those of the ideal flotation cell. Its design and construction are proved by universal acceptance and its supremacy is acknowledged by world-wide recognition and use.

1) Mixing and Aeration Zone:The pulp flows into the cell by gravity through the feed pipe, dropping directly on top of the rotating impeller below the stationary hood. As the pulp cascades over the impeller blades it is thrown outward and upward by the centrifugal force of the impeller. The space between the rotating blades of the impeller and the stationary hood permits part of the pulp to cascade over the impeller blades. This creates a positive suction through the ejector principle, drawing large and controlled quantities of air down the standpipe into the heart of the cell. This action thoroughly mixes the pulp and air, producing a live pulp thoroughly aerated with very small air bubbles. These exceedingly small, intimately diffused air bubbles support the largest number of mineral particles.

This thorough mixing of air, pulp and reagents accounts for the high metallurgical efficiency of the Sub-A (Fahrenwald) Flotation Machine, and its correct design, with precision manufacture, brings low horsepower and high capacity. Blowers are not needed, for sufficient air is introduced and controlled by the rotating impeller of the Denver Sub-A. In locating impeller below the stationary hood at the bottom of the cell, agitating and mixing is confined to this zone.

2) Separation Zone:In the central or separation zone the action is quite and cross currents are eliminated, thus preventing the dropping or knocking of the mineral load from the supporting air bubble, which is very important. In this zone, the mineral-laden air bubbles separate from the worthless gangue, and the middling product finds its way back into the agitation zone through the recirculation holes in the top of the stationary hood.

3) Concentrate Zone:In the concentrate or top zone, the material being enriched is partially separated by a baffle from the spitz or concentrate discharge side of the machine. The cell action at this point is very quiet and the mineral-laden concentrate moves forward and is quickly removed by the paddle shaft (note direct path of mineral). The final result is an unusually high grade concentrate, distinctive of the Sub-A Cell.

A flotation machine must not only float out the mineral value in a mixture of ground ore and water, but also must keep the pulp in circulation continuously from the feed end to the discharge end for the removal of the froth, and must give the maximum treatment positively to each particle.

It is an established fact that the mechanical method of circulating material is the most positive and economical, particularly where the impeller is below the pulp. A flotation machine must not only be able to circulate coarse material (encountered in every mill circuit), but also must recirculate and retreat the difficult middling products.

In the Denver Sub-A due to the distinctive gravity flow method of circulation, the rotating impeller thoroughly agitates and aerates the pulp and at the same time circulates this pulp upward in a straight line, removing the mineral froth and sending the remaining portion to the next cell in series. No short circuiting through the machine can thus occur, and this is most important, for the more treatments a particle gets, the greater the chances of its recovery. The gravity flow principle of circulation of Denver Sub-A Flotation Cell is clearly shown in the illustration below.

There are three distinctive advantages of theSub-A Fahrenwald Flotation Machines are found in no other machines. All of these advantages are needed to obtain successful flotation results, and these are:

Coarse Material Handled:Positive circulation from cell to cell is assured by the distinctive gravity flow principle of the Denver Sub-A. No short circuiting can occur. Even though the ore is ground fine to free the minerals, coarse materials occasionally gets into the circuit, and if the flotation machine does not have a positive gravity flow, choke-ups will occur.

In instances where successful metallurgy demands the handling of a dense pulp containing an unusually large amount of coarse material, a sand relief opening aids in the operation by removing from the lower part of the cell the coarser functions, directing these into the feed pipe and through the impeller of the flowing cell. The finer fraction pass over the weir overflow and thus receive a greater treatment time. In this manner short-circuiting is eliminated as the material which is bled through the sand relief opening again receives the positive action of the impeller and is subjected to the intense aeration and optimum flotation condition of each successive cell, floating out both fine and coarse mineral.

No Choke-Ups or Lost Time:A Sub-A flotation cell will not choke-up, even when material as coarse as is circulated, due to the feed and pulp always being on top of the impeller. After the shutdown it is not necessary to drain the machine. The stationary hood and the air standpipe during a shutdown protects the impeller from sanding-up and this keeps the feed and air pipes always open. Denver Sub-A flotation operators value its 24-hour per day service and its freedom from shutdowns.

This gravity flow principle of circulation has made possible the widespread phenomenal success of a flotation cell between the ball mill and classifier. The recovery of the mineral as coarse and as soon as possible in a high grade concentrate is now highly proclaimed and considered essential by all flotation operators.

Middlings Returned Without Pumps:Middling products can be returned by gravity from any cell to any other cell. This flexibility is possible without the aid of pumps or elevators. The pulp flows through a return feed pipe into any cell and falls directly on top of the impeller, assuring positive treatment and aeration of the middling product without impairing the action of the cell. The initial feed can also enter into the front or back of any cell through the return feed pipe.

Results : It is a positive fact that the application of these three exclusive Denver Sub-A advantages has increased profits from milling plants for many years by increasing recoveries, reducing reagent costs, making a higher grade concentrate, lowering tailings, increasing filter capacities, lowering moisture of filtered concentrate and giving the smelter a better product to handle.

Changes in mineralized ore bodies and in types of minerals quickly demonstrate the need of these distinctive and flexible Denver Sub-A advantages. They enable the treatment of either a fine or a coarse feed. The flowsheet can be changed so that any cell can be used as a rougher, cleaner, or recleaner cell, making a simplified flowsheet with the best extraction of mineral values.

The world-wide use of the Denver Sub-A (Fahrenwald) Flotation Machine and the constant repeat orders are the best testimonial of Denver Sub-A acceptance. There are now over 20,000 Denver Sub-A Cells in operation throughout the world.

There is no unit so rugged, nor so well built to meet the demands of the process, as the Denver Sub-A (Fahrenwald) Flotation Machine. The ruggedness of each cell is necessary to give long life and to meet the requirements of the process. Numerous competitive tests all over the world have conclusively proved the real worth of these cells to many mining operators who demand maximum result at the lower cost.

The location of the feed pipe and the stationary hood over the rotating impeller account for the simplicity of the Denver Sub-A cell construction. These parts eliminates swirling around the shaft and top of the impeller, reduce power load, and improve metallurgical results.

TheSub-A Operates in three zones: in bottom zone, impeller thoroughly mixes and aerates the pulp, the central zone separates the mineral laden particles from the worthless gangue, and in top zone the mineral laden concentrate high in grade, is quickly removed by the paddle of a Denver Sub-A Cell.

A Positive Cell Circulation is always present in theSub-A (Fahrenwald) Flotation Machine, the gravity flour method of circulating pulp is distinctive. There is no short circulating through the machine. Every Cell must give maximum treatment, as pulp falls on top of impeller and is aerated in each cell repeatedly. Note gravity flow from cell to cell.

Choke-Ups Are Eliminated in theSub-A Cell, even when material as coarse as is handled, due to the gravity flow principle of circulation. After shutdown it is not necessary to drain the machine, as the stationary hood protects impeller from sanding up. See illustration at left showing cell when shut down.

No Bowlers, noair under pressure is required as sufficient air is drawn down the standpipe. The expense and complication of blowers, air pipes and valves are thus eliminated. The standpipe is a vertical air to the heart of the Cell, the impeller. Blower air can be added if desired.

The Sub-A Flexibility allows it tobe used as a rougher, cleaner or recleaner. Rougher or middling product can be returned to the front or back of any cell by gravity without the use of pumps or elevators. Cells can be easily added when required. This flexibility is most important in operating flotation MILLS.

Pulp Level Is Controlled in each Sub-A Flotation Cell as it has an individual machine with its own pulp level control. Correct flotation requires this positive pulp level control to give best results in these Cells weir blocks are used, but handwheel controls can be furnished at a slight increase in cost. Note the weir control in each cell.

High Grade Concentrate caused by thequick removal of the mineral forth in the form of a concentrate increases the recovery. By having an adjustment paddle for each Sub-A Cell, quick removal of concentrate is assured, Note unit bearing housing for the impeller Shaft and Speed reducer drive which operates the paddle for each cell

Has Fewer Wearing Parts because Sub-A Cells are built for long, hard service, and parts subject to wear are easily replaced at low cost. Molded rubber wearing plates and impellers are light in weight give extra long life, and lower horsepower. These parts are made under exact Specifications and patented by Denver Equipment Co.

TheRugged Construction of theSub-A tank is made of heavy steel, and joints are welded both inside and out. The shaft assemblies are bolted to a heavy steel beam which is securely connected to the tank. Partition plates can be changed in the field for right or left hand machine. Right hand machine is standard.

The Minerals Separation or M.S. Sub-aeration cells, a section of which is shown in Fig. 32, consists essentially of a series of square cells with an impeller rotating on a vertical shaft in the bottom of each. In some machines the impeller is cruciform with the blades inclined at 45, the top being covered with a flat circular plate which is an integral part of the casting, but frequently an enclosed pump impeller is used with curved blades set at an angle of 45 and with a central intake on the underside ; both patterns are rotated so as to throw the pulp upwards. Two baffles are placed diagonally in each cell above the impeller to break up the swirl of the pulp and to confine the agitation to the lower zone. Sometimes the baffles are covered with a grid consisting of two or three layers each composed of narrow wood or iron strips spaced about an inch apart. The sides and bottom of the cells in the lower or agitation zone are protected from wear by liners, which are usually made of hard wood, but which can, if desired, consist of plates of cast-iron or hard rubber. The section directly under the impeller is covered with a circular cast-iron plate with a hole in the middle for the admission of pulp and air. The hole communicates with a horizontal transfer passage under the bottom liner, through which the pulp reaches the cell. Air is introduced into each cell through a pipe passing through the bottom and delivering its supply directly under the impeller. A low-pressure blower is provided with all machines except the smallest, of which the impeller speed is fast enough to draw in sufficient air by suction for normal requirements.

The pulp is fed to the first cell through a feed opening communicating with the transfer passage, along which it passes, until, at the far end, it is drawn up through the hole in the bottom liner by the suction of the impeller and is thrown outwards by its rotation into the lower zone. The square shape of the cell in conjunction with the baffles converts the swirl into a movement of intense agitation, which breaks up the air entering at the same time into a cloud of small bubbles, disseminating them through the pulp. The amount of aeration can be accurately regulated to suit the requirements of each cell by adjustment of the valve on its air pipe.

Contact between the bubbles and the mineral particles probably takes place chiefly in the lower zone. The pumping action of the impeller forces the aerated pulp continuously past the baffles into the upper and quieter part of the cell. Here the bubbles, loaded with mineral, rise more or less undisturbed, dropping out gangue particles mechanically entangled between them and catching on the way up a certain amount of mineral that has previously escaped contact. The recovery of the mineral in this way can be increased at the expense of the elimination of the gangue by increasing the amount of aeration. The froth collects at the top of the cell and is scraped by a revolving paddle over the lipat the side into the concentrate launder. The pulp, containing the gangue and any mineral particles not yet attached to bubbles, circulates to some extent through the zone of agitation, but eventually passes out through a slot situated at the back of the cell above the baffles and flows thence over the discharge weir. The height of the latter is regulated by strips of wood or iron and governs the level of the pulp in the cell. The discharge of each weir falls by gravity into the transfer passage under the next cell and is drawn up as before by the impeller. The pulp passes in this way through the whole machine until it is finally discharged as a tailing, the froth from each cell being drawn off into the appropriate concentrate launder.

No pipes are normally fitted for the transference of froth or other middling product back to the head of the machine or to any intermediate point. Should this be necessary, however, the material can be taken by gravity to the required cell through a pipe, which is bent at its lower end to pass under the bottom liner and project into the transfer passage, thus delivering its product into the stream of pulp that is being drawn up by the impeller

Particulars of the various sizes of M.S. Machines are given in Table 21. It should be noted that the size of a machine is usually defined by the diameter of its impeller ; for instance, the largest one would be described as a 24-inch machine.

The Sub-A Machine, invented by A. W. Fahrenwald and developed in many respects as an improvement in the Minerals Separation Machine, from which it differs considerably in detail, particularly in the method of aerating the pulp, although the principle of its action is essentially the same. Its construction can be seen from Figs. 33 and 34.

In common with the M.S. type of machine, it consists of a series of square cells fitted with rotating impellers. Each cell, however, is of unit construction, a complete machine being built up by mounting the required number of units on a common foundation and connecting up the pipes which transfer the pulp from one cell to the next. The cells are constructed of welded steel. The impeller, which can be rubber-lined,if required, carries six blades set upright on a circular dished disc, and is securely fixed to the lower end of the vertical driving shaft. It is covered with a stationary hood, to which are attached a stand-pipe, a feed pipe, and the middling return pipes. The underside of the hood is fitted with a renewable liner of rubber or cast-iron. The pulp, entering the first cell through the feed pipe and sometimes through the middling pipes, falls on to the impeller, the rotation of which throws it outwards into the bottom zone of agitation. The suction effect due to the rotationof the impeller draws enough air down the standpipe to supply the aeration necessary for normal operation. A portion of the pulp, cascading over the open tops of the impeller blades, entraps and breaks up the entrained air, the resulting spray-like mixture being then thrown out into the lower zone of agitation, where it is disseminated through the pulp as a cloud of fine bubbles. Should this amount of aeration be insufficient, air can be blown in under slight pressure through a hole near the top of the stand-pipe, in which case a rubber bonnet is fastenedto the lower bearing and clamped round the top of the stand-pipe so as to seal the supply from the atmosphere.

The bottom part of the cell is protected from wear by renewable cast-iron or rubber liners. Four vertical baffles, placed diagonally on the top of the hood, break up the swirl of the pulp and intensify theagitation in the lower zone. The pumping action of the impeller combined with the rising current of air bubbles carries the pulp to the quieter upper zone, where the bubbles, already coated with mineral, travel upwards, drop out many of the gangue particles which may have become entangled with them, and finally collect on the surface of the pulp as a mineralizedfroth. One side of the cell is sloped outwards so as to form, in conjunction with a vertical baffle, a spitzkasten-shaped zone of quiet settlement, where any remaining particles of gangue that have been caught and held between the bubbles are shaken out of the froth as it flows to the overflow lip at the front of the cell. The baffle prevents rising bubbles from entering the outer zone, thus enabling the gangue material released from the froth to drop down unhindered into the lower zone. A revolving paddle scrapes the froth past the overflow lip into the concentrate launder.

Should the machine be required to handle more than the normal volume of froth, it is built with a spitzkasten zone on both sides of the cell. For the flotation of ores containing very little mineral the spitzkasten is omitted so as to crowd the froth into the smallest possible space, the front of the cell being made vertical for the purpose.

Circulation of the pulp through the lower zone of agitation is maintained by means of extra holes at the base of the stand-pipe on a level with the middling return pipes. An adjustable weir provides for the discharge of the pulp to the next cell, which it enters through a feed-pipe as before. Below the weir on a level with the hood is a small sand holeand pipe through which coarse material can pass direct to the next cell without having to be forced up over the weir. The same process is repeated in each cell of the series, the froth being scraped over the lip of the machine, while the pulp passes from cell to cell until it is finally discharged as a tailing from the last one. The middling pipes make it an easy matter for froth from any section of the machine to be returned if necessary to any cell without the use of pumps.

Table 22 gives particulars of the sizes and power requirements of Denver Sub-A Machines and Table 23 is an approximate guide to their capacities under different conditions. The number of cells needed

Onemethod of driving the vertical impeller shafts of M.S. Subaeration or Denver Sub-A Machines is by quarter-twist belts from a horizontal lineshaft at the back of the machine, the lineshaft being driven in turn by a belt from a motor on the ground. This method is not very satisfactory according to modern standards, firstly, because the belts are liable to stretch and slip off, and, secondly, because adequate protection againstaccidents due to the belts breaking is difficult to provide without making the belts themselves inaccessible. A more satisfactory drive, with which most M.S. Machines are equipped, consists of a lineshaft over the top of the cells from which each impeller is driven through bevel gears. The lineshaft can be driven by a belt from a motor on the ground, by Tex- ropes from one mounted on the frame work of the machine, or by direct coupling to a slow-speed motor. This overhead gear drive needs careful adjustment and maintenance. Although it may run satisfactorily for years, trouble has been experienced at times, generally in plants where skilled mechanics have not been available. The demand for something more easily adjusted led to the development of a special form of Tex-rope drive which is shown in Fig. 35. Every impeller shaft is fitted at the top with a grooved pulley, which is driven by Tex-ropes from a vertical motor. This method is standard on Denver Sub-A Machines, and M.S. Machines are frequently equipped with it as well, but the former type are not made with the overhead gear drive except to special order.

The great advantage of mechanically agitated machines is that every cell can be regulated separately, and that reagents can be added when necessary at any one of them. Since, as a general rule, the most highly flocculated mineral will become attached to a bubble in preference to a less floatable particle, in normal operation the aeration in the first few cells of a machine should not be excessive ; theoretically there should be no more bubbles in the pulp than are needed to bring up the valuable minerals. By careful control of aeration it should be possible for the bulk of the minerals to be taken off the first few cells at the feed end of the machine in a concentrate rich enough to be easily cleaned, and sometimes of high enough grade to be sent straight to the filtering section as a finished product. The level of the pulp in these cells is usually kept comparatively low in order to provide a layer of froth deep enough to give entangled particles of gangue every chance of dropping out, but it must not be so low that the paddles are prevented from skimming off the whole of the top layer of rich mineral. Towards the end of the machine a scavenging action is necessary to make certain that the least possible amount of valuable mineral escapes in the tailing, for which purpose the gates of the discharge weirs are raised higher than at the feed end, and the amountof aeration may have to be increased. The froth from the scavenging cells is usually returned to the head of the machine, the middling pipes of the Denver Sub-A Machine being specially designed for such a purpose. The regulation of the cleaning cells is much the same as that of the first few cells of the primary or roughing machine, to the head of which the tailing from the last of the cleaning cells is usually returned.

A blower is sometimes required with the M.S. Subaeration Machine. Since each cell is fitted with an air pipe and valve, accurate regulation of aeration is a simple matter. The Denver Sub-A, Kraut, and Fagergren Machines, however, are run without blowers, enough air being drawn into the machines by suction.

In the Geco New-Cell Flotation Cellthe pneumatic principle is utilized in conjunction with an agitating device. The machine, which is illustrated in Fig. 44, consists of a trough or cell made of steel or wood, whichever is more convenient, through the bottom of which projects a series of air pipes fitted with circular mats of perforated rubber. The method of securing the mat to the air pipe can be seen from Fig. 45. Over each mat rotates a moulded rubber disc of slightlylarger diameter at a peripheral speed of 2,500 ft. per minute. It is mounted on a driving spindle as shown in Fig. 46. Each spindle is supported and aligned by ball-bearings contained in a single dust- and dirt-proof casting, and each pair is driven from a vertical motor through Tex-ropes and grooved pulleys, a rigid steel structure supporting the whole series of spindles with their driving mechanism. The machine can be supplied, if required, however, with a quarter-twist drive from a lineshaft over flat pulleys.

The air inlet pipes are connected to a main through a valve by which the amount of air admitted to each mat can be accurately controlled. The air is supplied by a low-pressure blower working at about 2 lb. per square inch. It enters the cell through the perforations in the rubber mat and is split up into a stream of minute bubbles, which are distributed evenly throughout the pulp by the action of the revolving disc. By this means a large volume of finely-dispersed air is introduced withoutexcessive agitation. There is sufficient agitation, however, to produce a proper circulation in the cell, but not enough to cause any tendency to surge or to disturb the froth on the surface of the pulp. All swirling movement is checked by the liner-baffles with which the sides of the cell are lined ; their construction can be seen in Fig. 44. They are constructed of white cast iron and are designed to last the life of the machine, the absence of violent agitation making this possible.The pulp must be properly conditioned before entering the machine. It is admitted through a feed box at one end at a point above the first disc, and passes along the length of the cell to the discharge weir without being made to pass over intermediate weirs between the discs. The height of the weir at the discharge end thus controls the level of the pulp in the machine. The froth that forms on the surface overflows the froth lip in a continuous stream without the aid of scrapers, its depth being controlled at any point by means of adjustable lip strips combined with regulation of the air.The Geco New-Cell is made in four sizesviz., 18-, 24-, 36-, and 48-in. machines, the figure representing the length of the side of the squarecell. Particulars of the three smallest sizes are given in Table 27. Figures are not available for the largest size.

flotation cell - an overview | sciencedirect topics

flotation cell - an overview | sciencedirect topics

The MAC flotation cell was developed by Kadant-Lamort Inc. It can save energy comparedto conventional flotation systems. The MAC flotation cell is mainly used in the flotation section of waste paper deinking pulping, for removal of hydrophobic impurities such as filler, ash,ink particles, etc. It can increase pulp whiteness and meet the requirements of final paper appearance quality. Table11.11 shows the features of MAC flotation cell. Kadants MAC flotation cell deinking system uses air bubbles to float ink particles to the cell surface for removal from the recycled material. The latest generation of the MAC cell deinking system incorporates a patented bubble-washing process to reduce power consumption and also fiber loss. It combines small, new, auto-clean, low-pressure injectors with a flotation cell. The function of injectors is to aerate the stock before it is pumped and sent tangentially to the top of the cell. The air bubbles collect ink particles in the cell and rise up to the top to create a thick foam mat that is evacuated because of the slight pressurization of the cell. The partially deinked stock then goes to a deaeration chamber and is pumped to the next stage. Here, the operation is exactly the same as for the first stage. This stage also has the same number of injectors and same flow (Kadant,2011). This operation is repeated up to five times for a high ink removal rate. Remixing of the air coming from downstream stages of the process helps the upstream stages and improves the overall cell efficiency. Adjustable and selective losses of fiberdepend on the application and technical requirements inks, or inks and fillers. The use of low-pressure injectors in the MAC flotation cell could save about 2530% of the energy used in conventional flotation systems (ECOTARGET,2009). The benefits of the MAC flotation cell are summarized in Table11.12.

Agitated flotation cells are widely used in the mineral processing industry for separating, recovering, and concentrating valuable particulate material from undesired gangue. Their performance is lowered, however, when part of the particulate system consists of fines, with particle diameters typically in the range from 30 to 100m. For example, it was observed difficult to float fine particles because of the reduction of middle particles (of wolframite) as carriers and the poor collision and attachment between fine particles and air bubbles; a new kinetic model was proposed [34].

As an alternative to agitated cells, bubble columnsused in chemical engineering practice as chemical reactorswere proposed for the treatment of fine particle systems. Flotation columns, as they came to be known, were invented back in the 1960s in Canada [35]. The main feature that differentiates the column from the mechanical flotation cell (of Denver type) is wash water, added at the top of the froth. It was thought to be beneficial to overall column performance since it helps clean the froth from any entrained gangue, while at the same time preventing water from the pulp flowing into the concentrate. In this way, it was hoped that certain cleaning flotation stages could be gained.

Let us note that the perhaps insistence here on mineral processing is only due to the fact that most of the available literature on flotation is from this area, where the process was originated and being widely practiced. The effect of particle size on flotation recovery is significant; it was shown that there exists a certain size range in which optimum results may be obtained in mineral processing. This range varies with the mineral properties such as density, liberation, and so on, but was said to be of the order of 10100m [36].

Regulating the oxidation state of pyrite (FeS2) and arsenopyrite (FeAsS), by the addition of an oxidation or reduction chemical agent and due to the application of a short-chain xanthate as collector (such as potassium ethyl xanthate, KEX), was the key to selective separation of the two sulfide minerals, pyrite and arsenopyrite [37]. Strong oxidizing agents can depress previously floated arsenopyrite. Various reagents were examined separately as modifiers and among them were sodium metabisulfite, hydrazinium sulfate, and magnesia mixture. The laboratory experiments were carried out in a modified Hallimond tube, assisted by zeta-potential measurements and, in certain cases, by contact angle measurements.

This conventional bench-scale flotation cell provides a fast, convenient, and low-cost method, based on small samples (around 2g), usually of pure minerals and also artificial mixtures, for determining the general conditions under which minerals may be rendered floatableoften in the absence of a frother (to collect the concentrate in the side tube) [38]. This idea was later further modified in the lab replacing the diaphragm, in order to conduct dissolved air or electroflotation testssee Section 3.

Pyrite concentrates sometimes contain considerable amounts of arsenic. Since they are usually used for the production of sulfuric acid, this is undesirable from the environmental point of view. However, gold is often associated with arsenopyrite, often exhibiting a direct relationship between Au content and As grade. There is, therefore, some scope for concentrating arsenopyrite since the ore itself is otherwise of little value (see Fig.2.2). Note that previous work on pyrites usually concentrated on the problem of floating pyrite [40].

In the aforementioned figure (shown as example), the following conditions were applied: (1) collector [2-coco 2-methyl ammonium chloride] 42mg/L, frother (EtOH) 0.15% (v/v), superficial liquid velocity uL=1.02cm/s, superficial gas velocity uG=0.65cm/s, superficial wash water velocity uw=0.53cm/s; (2) hexadecylamine, 45mg/L; pine oil, 50mg/L; EtOH, 0.025%; uL=0.84cm/s; uG=0.72cm/s; uw=0.66cm/s; (3) Armoflot 43, 50mg/L; pine oil, 50mg/L; EtOH, 0.025%; uL=0.84cm/s; uG=0.71cm/s; uw=0.66cm/s [39]. The pyrite (with a relatively important Au content of 21g/ton) was a xanthate-floated concentrate. The presence of xanthates, however, might cause problems in the subsequent cyanidation of pyrites when recovering their Au value, which perhaps justified the need to find alternative collectors. In general, the amines exhibited a behavior similar to that of the xanthates (O-alkyl dithiocarbonates). The benefit of the amine was in its lower consumption, as compared with the xanthate systems.

The arsenic content of the pyrite was approximately 9% (from an initial 3.5% of the mixed sulfide ore). The material was sieved and the75m fraction was used for the laboratory-scale cylindrical column experiments. The effect on metallurgical characteristics of the flotation concentrate of varying the amount of ferric sulfate added to the pulp was studied; three collectors were used and their performance was compared (in Fig.2.2). Both hexadecylamine and Armoflot 43 (manufactured by Akzo) exhibited an increased recovery but a very low enrichment, whereas 2-coco 2-methyl ammonium chloride (Arquad-2C) showed a considerable enrichment; a compromise had to be made, therefore, between a high-grade and a low recovery.

Electroflotation (electrolytic flotation) is an unconventional separation process owing its name to the bubbles generation method it uses, i.e., electrolysis of the aqueous medium. In the bottom of the microcell, the two horizontal electrodes were made from stainless steel, the upper one being perforated. The current density applied was 300 Am2. It was observed that with lime used to control pH, different behavior was observed (see Fig.2.3). Pyrite, with permanganate (a known depressant) also as modifier, remained activated from pH 5.0 to 8.0at 80% recovery, while it was depressed at the pH range from 9.0 to 12.0. A conditioning of 30min was applied in the presence of modifier alone and further 15min after the addition of xanthate. The pure mineral sample, previously hand collected, crushed, and pulverized in the laboratory, was separated by wet sieving to the45 to+25m particle size range.

Pyrite due to its very heterogeneous surface, consisting of a mosaic of anodic and cathodic areas, presents a strong electrocatalytic activity in the anodic oxidation of xanthate to dixanthogen. It is also possible that the presence of the electric field, during electroflotation, affected the reactions taking place. In order to explain this difference in flotation behavior thermodynamic calculations for the system Fe-EX-H2O have been done [41]. It was concluded that electroflotation was capable of removing fine pyrite particles from a dilute dispersion, under controlled conditions. Nevertheless, dispersed air and electroflotation presented apparent differences for the same application.

The size of the gas bubbles produced was of the order of 50m, in diameter [21]. Similar measurements were later carried out at Newcastle, Australia [42]; where it was also noted that a feature of electroflotation is the ability to create very fine bubbles, which are known to improve flotation performance of fine particles.

In fact, the two electrodes of a horizontal electrodes set, usually applied in electroflotation, could be separated by a cation exchange membrane, as only one of the produced gases is often necessary [43]. In the lower part/separated electrode, an electrolyte was circulated to remove the created gas, and in the meantime, increase the conductivity; hence having power savings (as the electric field is built up between the electrodes through the use of the suspension conductivity). Attention should be paid in this case to anode corrosion, mainly by the chloride ion (i.e., seawater).

Microorganisms have a tremendous influence on their environment through the transfer of energy, charge, and materials across a complex biotic mineralsolution interface; the biomodification of mineral surfaces involves the complex action of microorganism on the mineral surface [44]. Mixed cationic/anionic surfactants are also generating increasing attention as effective collectors during the flotation of valuable minerals (i.e., muscovite, feldspar, and spodumene ores); the depression mechanisms on gangue minerals, such as quartz, were focused [45].

Another design of a flotation cell which applies ultrasound during the flotation process has been developed by Vargas-Hernndez et al. (2002). The design consists of a Denver cell (Koh and Schwarz, 2006) equipped with ultrasonic capabilities of performing ultrasound-assisted flotation experiments. This cell is universally accepted as a standard cell for laboratory flotation experiments. In Figure 35.25, a schematic of the Denver cell equipped with two power transducers is shown operating at 20kHz. The ultrasonic transducers are in acoustic contact with the body of the flotation cell but are not immersed in the same cell. Instead, they are submerged in distilled water and in a thin membrane that separates the radiant head of the transducer from the chamber body. The floatation chamber has a capacity of 2.7l and is also equipped with conventional systems to introduce air and mechanical agitation able to maintain the suspension of metallurgical pulp. In the upper part of the cell there is an area in which the foam is recovered for analysis by a process called skimming. The block diagram of Figure 35.25 further shows that the experimental system was developed to do ultrasonic-assisted flotation experiments. The transducers operate at 20kHz and can handle power up to 400W. In the Denver cell an acoustic probe, calibrated through a nonlinear system and capable of measuring high-intensity acoustic fields, is placed (Gaete-Garretn et al., 1993, 1998). This is done in order to determine the different acoustic field intensities with a spatial scanner during the experimentation. Figure 35.26 shows the distribution of ultrasonic field intensity obtained by a spatial scanner in the central area of the flotation chamber. The Denver cell with ultrasonic capabilities, as described, is shown in Figure 35.27. The obtained results were fairly positive. For example, for fine particle recovery it worked with metallurgical pulp under 325mesh, indicating floating particles of less than 45m, and the recovery curves are almost identical to those of an appropriate size mineral for flotation. This is shown in Figure 35.28, where a comparison between typical copper recovery curves for fine and normal particles is presented. The most interesting part of the flotation curves is the increase in recovery of molybdenum with ultrasonic power, as shown in Figure 35.29. The increase in recovery of iron is not good news for copper mines because the more iron floating the lower grade of recovery. This may be because the iron becomes more hydrophobic with ultrasonic action. According to the experts, this situation could be remedied by looking for specific additives to avoid this effect. Flotation kinetics shown in Figure 35.30 with 5 and 10W of acoustic power applied also show an excellent performance. It should be noted that the acoustic powers used to vary the flotation kinetics have been quite low and could clearly be expanded.

Figure 35.28. Compared recovering percent versus applied power in an ultrasonic-assisted flotation process in a Denver cell: (a) fine and ultrafine particles recovering and (b) normal particles recovering.

These experiments confirm the potential of power ultrasound in flotation. Research on assisted flotation with power ultrasound has been also carried out by Ozkan (2002), who has conducted experiments by pretreating pulp with ultrasound during flotation. Ozkhans objective was to recover magnesite from magnesite silts with particles smaller than 38m. Their results show that under ultrasonic fields the flotation foam bubbles are smaller, improving magnesite recovery rates. When Ozkhan treated magnesite mineral with a conventional treatment the beneficial effect of ultrasound was only manifested for mineral pretreatment. The flotation performed under ultrasonic field did not show improvement. This was because power ultrasound improves the buoyancy of clay iron and this has the effect of lowering the recovery of magnesite.

Kyllnen et al. (2004) employed a cell similar to Jordan to float heavy metals from contaminated soils in a continuous process. In their experiments they obtained a high recovery of heavy metals, improving the soil treatment process. Alp et al. (2004) have employed ultrasonic waves in the flotation of tincal minerals (borax Na O710 B4 H2O), finding the same effects as described above, i.e., that power ultrasound helps in the depression of clay. However, the beneficial effect of ultrasound is weakened when working with pulps with high mineral concentration (high density), probably due to an increase in the attenuation of the ultrasonic field. Safak and Halit (2006) investigated the action mechanisms of ultrasound under different flotation conditions. A cleaning effect on the floating particles was attributed to the ultrasonic energy, making the particles more reactive to the additives put in the metallurgical pulp. Furthermore due to the fact that the solid liquid interface is weaker than the cohesive forces of the metallurgic pulp liquids, it results in a medium favorable to creation of cavitation bubbles. The unstable conditions of a cavitation environment can produce changes in the collectors and even form emulsions when entering the surfactant additives. In general, many good properties are attributed to the application of ultrasound in flotation. For example, there is a more uniform distribution of the additives (reagents) and an increase in their activity. In fact in the case of carbon flotation it has been found that the floating times are shortened by the action of ultrasound, the bubble sizes are more stable, and the consumption of the reagents is drastically lowered.

Abrego Lpez (2006) studied a water recovery process of sludge from industrial plants. For this purpose he employed a flotation cell assisted by power ultrasound. In the first stage he made a flotation to recover heavy metals in the metallurgical pulp, obtaining a high level of recovery. In the second stage he added eucalyptus wood cones to the metallurgical pulp to act as an accumulator of copper, lead, nickel, iron, and aluminum. The author patented the method, claiming that it obtained an excellent recovery of all elements needing to be extracted. zkan and Kuyumcu (2007) showed some design principles for experimental flotation cells, proposing to equip a Denver flotation cell with four power transducers. Tests performed with this equipment consisted of evaluating the possible effects that high-intensity ultrasonic fields generated in the cell may have on the flotation. The author provides three-dimensional curves of ultrasonic cavitation fields in a Denver cell filled with water at frequencies between 25 and 40kHz. A warming effect was found, as expected. However, he states that this effect does not disturb the carbon recovery processes because carbon flotation rarely exceeds 5min. They also found that the pH of tap water increases with the power and time of application of ultrasound, while the pH of the carbonwaterreagentsludge mixture decreases. The conductivity of the metallurgical pulp grows with the power and time of application of ultrasound, but this does not affect flotation. The carbon quality obtained does not fall due to the application of ultrasound and the consumption of lowered reagents. They did not find an influence from the ultrasound frequency used in the process, between 25 and 40kHz. They affirmed that ultrasound is beneficial at all stages of concentration.

Kang et al. (2009) studied the effects of preconditioning of carbon mineral pulp in nature by ultrasound with a lot of sulfur content. They found that the nascent oxygen caused by cavitation produces pyrite over oxidation, lowering its hydrophobicity, with the same effect on the change of pH induced by ultrasonic treatment. Additionally, ultrasound decreases the liquid gas interfacial tension by increasing the number of bubbles. Similar effects occur in carbon particles. The perfect flotation index increases 25% with ultrasonic treatment. Kang et al. (2008) continued their efforts to understand the mechanism that causes effects in ultrasonic flotation, analyzing the floating particles under an ultrasonic field by different techniques like X-ray diffraction, electron microscopy, and scanning electron microscopy techniques. In carbon flotation it is estimated that ultrasonic preconditioning may contribute to desulfurization and ash removal (deashing) in carbon minerals. Zhou et al. (2009) have investigated the role of cavitation bubbles created by hydrodynamic cavitation in a flotation process, finding similar results to those reported for ultrasonic cavitation flotation. Finally, Ozkan (2012) has conducted flotation experiments with the presence of hard carbon sludge cavitation (slimes), encountering many of the effects that have been reported for the case of metallurgical pulp with ultrasound pretreatment. This includes improved flotation, drastic reduction in reagent consumption, and the possible prevention of oxidation of the surface of carbon sludge. A decrease in the ash content in floating carbon was not detected. However, tailings do not seem to contain carbon particles. All these effects can be attributed to acoustic cavitation. However, according to the author, there is a need to examine the contribution of ultrasound to the probability of particlebubble collision and the likelihood of getting the bubbles to connect to the particles. The latter effects have been proposed as causes for improvements in flotation processes in many of the publications reviewed, but there is no systematic study of this aspect.

In summary, power ultrasound assistance with flotation processes shows promising results in all versions of this technique, including conditioning metallurgical pulp before floating it, assisting the continuous flotation process, and improving the yields in conventional flotation cells. The results of ultrasonic floating invariably show a better selectivity and an increase, sometimes considerable, in the recovery of fine particles. Paradoxically, in many experiments an increase has been recorded in recovering particles suitable for normal flotation. These facts show the need for further research in the flotation process in almost all cases, with the exception perhaps of carbon flotation. For this last case, in light of the existing data the research should be directed toward scale-up of the technology.

The concentrate obtained from a batch flotation cell changes in character with time as the particles floating change in size, grade and quantity. In the same way, the concentrate from the last few cells in a continuous bank is different from that removed from the earlier cells. Particles of the same mineral float at different rates due to different particle characteristics and cell conditions.

The recovery of any particular mineral rises to an asymptotic value R which is generally less than 100%. The rate of recovery at time t is given by the slope of the tangent to the curve at t, and the rate of recovery at time t1 is clearly greater than the rate at time t2. There is a direct relationship between the rate of flotation and the amount of floatable material remaining in the cell, that is:

The process is carried out in a flotation cell or tank, of which there are two basic types, mechanical and pneumatic. Within each of these categories, there are two subtypes, those that operate as a single cell, and those that are operated as a series or bank of cells. A bank of cells (Fig. 8) is preferred because this makes the overall residence times more uniform (i.e., more like plug flow), rather than the highly diverse residence times that occur in a single (perfectly mixed) tank.

FIGURE 8. Flotation section of a 80,000t/d concentrating plant, showing the arrangement of the flotation cells into banks. A small part of the grinding section can be seen through the gap in the wall. [Courtesy Joy Manufacturing Co.]

The purpose of the flotation cell is to attach hydrophobic particles to air bubbles, so that they can float to the surface, form a froth, and can be removed. To do this, a flotation machine must maintain the particles in suspension, generate and disperse air bubbles, promote bubbleparticle collision, minimize bypass and dead spaces, minimize mechanical passage of particles to the froth, and have sufficient froth depth to allow nonhydrophobic (hydrophilic) particles to return to the suspension.

Pneumatic cells have no mechanical components in the cell. Agitation is generally by the inflow of air and/or slurry, and air bubbles are usually introduced by an injector. Until comparatively recently, their use was very restricted. However, the development of column flotation has seen a resurgence of this type of cell in a wider, but still restricted, range of applications. While the total volume of cell is still of the same order as that of a conventional mechanical cell, the floor space and energy requirements are substantially reduced. But the main advantage is that the cell provides superior countercurrent flow to that obtained in a traditional circuit (see Fig. 11), and so they are now often used as cleaning units.

Mechanical cells usually consist of long troughs with a series of mechanisms. Although the design details of the mechanisms vary from manufacturer to manufacturer, all consist of an impeller that rotates within baffles. Air is drawn or pumped down a central shaft and is dispersed by the impeller. Cells also vary in profile, degree of baffling, the extent of walling between mechanisms, and the discharge of froth from the top of the cell.

Selection of equipment is based on performance (represented by grade and recovery), capacity (metric tons per hour per cubic meter); costs (including capital, power, maintenance), and subjective factors.

Among all processing industries, only in the ore and mining industries is the accent more on wear resistance than corrosion. In mining industries, the process concerns material handling more than any physical or chemical conversions that take place during the refining operations. For example, in the excavation process of iron ore, conventional conveyer systems and sophisticated fluidized systems are both used [16,17]. In all these industries, cost and safety are the governing factors. In a fluidized system, the particles are transported as slurry using screw pumps through large pipes. These pipes and connected fittings are subjected to constant wear by the slurry containing hard minerals. Sometimes, depending on the accessibility of the mineral source, elaborate piping systems will be laid. As a high-output industry any disruption in the work will result in heavy budgetary deficiency. Antiabrasive rubber linings greatly enhance the life of equipment and reduce the maintenance cost. The scope for antiabrasive rubber lining is tremendous and the demand is ever increasing in these industries.

Different rubber compounds are used in the manufacture of flotation cell rubber components for various corrosion and abrasion duty conditions. Flotation as applied to mineral processing is a process of concentration of finely divided ores in which the valuable and worthless minerals are completely separated from each other. Concentration takes place from the adhesion of some species of solids to air bubbles and wetting of the other series of solids by water. The solids adhering to air bubbles float on the surface of the pulp because of a decrease in effective density caused by such adhesion, whereas those solids that are wetted by water in the pulp remain separated in the pulp. This method is probably the more widely used separation technique in the processing of ores. It is extensively used in the copper, zinc, nickel, cobalt, and molybdenum sections of the mineral treatment industry and is used to a lesser extent in gold and iron production. The various rubber compounds used in the lining of flotation cells and in the manufacture of their components for corrosive and abrasive duties are:

Operating above the maximum capacity can cause the performance of flotation cells to be poor even when adequate slurry residence time is available (Lynch et al., 1981). For example, Fig. 11.21 shows the impact of increasing volumetric feed flow rate on cell performance (Luttrell et al., 1999). The test data obtained at 2% solids correlates well with the theoretical performance curve predicted using a mixed reactor model (Levenspiel, 1972). Under this loading, coal recovery steadily decreased as feed rate increased due to a reduction in residence time. However, as the solids content was increased to 10% solids, the recovery dropped sharply and deviated substantially from the theoretical curve due to froth overloading. This problem can be particularly severe in coal flotation due to the high concentration of fast floating solids in the flotation feed and the presence of large particles in the flotation froth. Flotation columns are particularly sensitive to froth loading due to the small specific surface area (ratio of cross-sectional area to volume) for these units.

Theoretical studies indicate that loading capacity (i.e., carrying capacity) of the froth, which is normally reported in terms of the rate of dry solids floated per unit cross-sectional area, is strongly dependent on the size of particles in the froth (Sastri, 1996). Studies and extensive test work conducted by Eriez personnel also support this finding. As seen in Fig. 11.22, a direct correlation exists between capacity and both the mean size (d50) and ultrafines content of the flotation feedstock. The true loading capacity may be estimated from laboratory and pilot-scale flotation tests by conducting experiments as a function of feed solids content (Finch and Dobby, 1990). Field surveys indicate that conventional flotation machines can be operated with loading capacities of up to 1.52.0t/h/m2 for finer (0.150mm) feeds and 56t/h/m2 or more for coarser (0.600mm) feeds. Most of the full-scale columns in the coal industry operate at froth loading capacities less than 1.5t/h/m2 for material finer than 0.150mm and as high as 3.0t/h/m2 for flotation feed having a top size of 0.300mm feeds.

Froth handling is a major problem in coal flotation. Concentrates containing large amounts of ultrafine (<0.045mm) coal generally become excessively stable, creating serious problems related to backup in launders and downstream handling. Bethell and Luttrell (2005) demonstrated that coarser deslime froths readily collapsed, but finer froths had the tendency to remain stable for an indefinite period of time. Attempts made to overcome this problem by selecting weaker frothers or reducing frother dosage have not been successful and have generally led to lower circuit recoveries. Therefore, several circuit modifications have been adopted by the coal industry to deal with the froth stability problem. For example, froth launders need to be considerably oversized with steep slopes to reduce backup. Adequate vertical head must also be provided between the launder and downstream dewatering operations. In addition, piping and chute work must be designed such that the air can escape as the froth travels from the flotation circuit to the next unit operation.

Figure 11.23 shows how small changes in piping arrangements can result in better process performance. Shown in Fig. 11.23 is a column whose performance suffered due to the inability to move the froth product from the column launder although a large discharge nozzle (11m) had been provided. In this example, the froth built up in the launder and overflowed when the operators increased air rates. To prevent this problem, the air rates were lowered, which resulted in less than optimum coal recovery. It was determined that the downstream discharge piping was air-locking and preventing the launders from properly draining. The piping was replaced with larger chute work that allowed the froth to flow freely and the air to escape. As a result, higher aeration rates were possible and recoveries were significantly improved.

Some installations have resorted to using defoaming agents or high-pressure launder sprays to deal with froth stability. However, newer column installations eliminate this problem by including large de-aeration tanks to allow time for the froth to collapse (Fig. 11.24a). Special provisions may also be required to ensure that downstream dewatering units can accept the large froth volumes. For example, standard screen-bowl centrifuges equipped with 100mm inlets may need to be retrofitted with 200mm or larger inlets to minimize flow restrictions. In addition, while the use of screen-bowl centrifuges provides low product moistures, there are typically fine coal losses, as a large portion of the float product finer than 0.045mm is lost as main effluent. This material is highly hydrophobic and will typically accumulate on top of the thickener as a very stable froth layer, which increases the probability that the process water quality will become contaminated (i.e., black water).

This phenomenon is more prevalent in by-zero circuits, especially when the screen-bowl screen effluent is recycled back through the flotation circuit, either directly or through convoluted plant circuitry. Reintroducing material that has already been floated to the flotation circuit can result in a circulating load of very fine and highly floatable material. As a result, the capacity of the flotation equipment can be significantly reduced, which results in losses of valuable coal. Most installations will combat this by ensuring that the screen-bowl screen effluent is routed directly back to the screen bowl so that it does not return to the flotation circuit. The accumulation of froth on the thickener, which tends to be especially problematic in by-zero circuitry, is also reduced by utilizing reverse-weirs and taller center wells, as this approach helps to limit the amount of froth that can enter into the process water supply. Froth that does form on top of the clarifier can be eliminated by employing a floating boom that is placed directly in the thickener (Fig. 11.24b) and used in conjunction with water sprays. The floating boom can be constructed out of inexpensive PVC piping, and is typically attached to the rotating rakes. The boom floats on the water interface and drags any froth around to the walkway that extends over the thickener, where it is eliminated by the sprays.

Column cells have been developed over the past 30 years as an alternative to mechanically agitated flotation cells. The major operating difference between column and mechanical cells is the lack of agitation in column cells that reduces energy and maintenance costs. Also, it has been reported that the cost of installing a column flotation circuit is approximately 2540% less than an equivalent mechanical flotation circuit (Murdock et al., 1991). Improved metallurgical performance of column cells in iron ore flotation is reported and attributed to froth washing, which reduces the loss of fine iron minerals entrained into the froth phase (Dobby, 2002).

The Brazilian iron ore industry has embraced the use of column flotation cells for reducing the silica content of iron concentrates. Several companies, including Samarco Minerao S.A., Companhia Vale do Rio Doce (CRVD), Companhia Siderrgica Nacional (CSN), and Mineraes Brasileiras (MBR), are using column cells at present (Peres et al., 2007). Samarco Minerao, the first Brazilian producer to use column cells, installed column cells as part of a plant expansion program in the early 1990s (Viana et al., 1991). Pilot plant tests showed that utilization of a column recleaner circuit led to a 4% increase in iron recovery in the direct reduction concentrate and an increase in primary mill capacity when compared to a conventional mechanical circuit.

There are also some negative reports of the use of column cells in the literature. According to Dobby (2002), there were several failures in the application of column cells in the iron ore industry primarily due to issues related to scale-up. At CVRD's Samitri concentrator, after three column flotation stages, namely, rougher, cleaner, and recleaner, a secondary circuit of mechanical cells was still required to produce the final concentrate.

Imhof et al. (2005) detailed the use of pneumatic flotation cells to treat a magnetic separation stream of a magnetite ore by reverse flotation to reduce the silica content of the concentrate to below 1.5%. From laboratory testing, they claimed that the pneumatic cells performed better than either conventional mechanical cells or column cells. The pneumatic cells have successfully been implemented at the Compaia Minera Huasco's iron ore pellet plant.

This chapter presents a novel approach to establish the relationship between collector properties and the flotation behavior of goal in various flotation cells. Coal flotation selectivity can be improved if collector selection is primarily based on information obtained from prior contact angle and zeta potential measurements. In a study described in the chapter, this approach was applied to develop specific collectors for particular coals. A good correlation was obtained between laboratory batches and large-scale conventional flotation cells. This is not the case when these results are correlated with pneumatic cell trial data. The study described in the chapter was aimed at identifying reasons for the noncorrelation. Two collectors having different chemical compositions were selected for this investigation. A considerable reduction in coal recovery occurred at lower rotor speeds when comparing results of oxidized and virgin coal. The degree to which a collector enhances flocculation in both medium- and low-shear applications and also the stronger bubble-coal particle adherence required for high-shear cells must, therefore, all be taken into consideration when formulating a collector for coal flotation.

sulphide flotation - an overview | sciencedirect topics

sulphide flotation - an overview | sciencedirect topics

The Platsol process was originally developed in collaboration with the University of British Columbia, Kane Consultants Ltd., and Lakefield Research in Canada for the treatment of flotation sulfide concentrate for Polymet Mining Company in Minnesota and was tested on similar types of concentrate materials. This process involved dissolution in one step of the base metals (copper and nickel) as well as the gold and PGMs. This was followed by solidliquid separation, gold and PGM recovery, and conventional Cu SX/EW and recycling of the copper raffinate to the autoclave.

The fundamental difference between the Platsol process and the conventional high-temperature pressure-oxidation processes is that a small concentration of chloride ions is added to the autoclave with 25g/L sulfuric acid. The chloride favors the oxidation of gold and PGMs and stabilizes them as dissolved chloro-complexes. Grinding the ore with ceramic rather than iron balls was required to prevent cementation of gold chloride (Ferron etal., 2000).

The concentrate tested was a flotation concentrate from the Northmet project, USA, assaying 14.7% Cu, 3.05% Ni, 0.14% Co, 26.7% S, 1.4g/t Au, 2.2g/t Pt, and 9.9g/t Pd. Pressure-oxidation conditions were 225C, pulp density was 11%, retention time was 120min, and oxygen overpressure was 689kPa. The ore treated had a P80 of 15m. After solidliquor separation, the gold and PGMs were recovered by sulfide precipitation using NaHS or by activated carbon. The copper was recovered using conventional solvent extraction and electrowinning techniques. Overall recoveries were Cu 99.6%, Ni 98.9%, Co 96%, Pd 94.6%, Pt 96%, and Au 89.4% (Ferron etal., 2000). Studies in treating a variety of refractory gold concentrates under optimum conditions (225C, NaCl 1020g/L, 26h O2 at 700kPa) achieved gold extractions of the order of 9096% compared with direct cyanidation where gold extraction was less than 20% (Ferron etal., 2003).

Examination of a variety of recovery options showed that loading of gold onto carbon from clear liquors and pulps was rapid and did not require prior neutralization. Zadra elution of the loaded carbon recovered more than 90% of the gold; however, further work in investigating carbon regeneration was required. Gold could easily be precipitated from acidic Platsol leach liquors with NaHS, but minimization of the co-precipitation of impurities, such as copper, needed to be addressed, given the co-leaching of base and precious metals. In addition, some tests on gold recovery by ion-exchange resins showed promise. Gold chloride can be precipitated using a synthetic covellite produced in residual copper recovery process. The Platsol process has undergone a detailed engineering phase following DFS and FEED studies for the Northmet project (Wardell-Johnson etal., 2009). More details may be found in Chapter 46.

Pyrite and arsenopyrite are the principal hosts of submicroscopic gold. In addition, gold minerals are often preferentially associated with these minerals; hence, their floatability is relevant to gold metallurgy. The following possibilities exist in the case of sulfidic gold ores:

Although the exact location of arsenic in the pyrite crystal structure is still being debated the general consensus is that arsenic is replacing one of the two sulfur atoms in the sulfur dipole, thus forming AsS2. Whatever the mechanism of arsenic incorporation in the pyrite structure, it is certain that with increasing arsenic content, pyrite becomes more readily oxidizable, which in turn affects its floatability. Pure pyrite floats without activation because of dixanthogen formation as a result of catalytic oxidation of xanthate on clean (fresh) pyrite surfaces. However, most floated pyrite is either copper or lead activated which means that incipient surface oxidation had to take place to form islands of pyrrhotite, which then became sites for activation by Cu or Pb. With increasing arsenic content, the surface oxidation of pyrite is greatly accelerated to the point where depression by surface oxidation overwhelms activation. This pyrite requires heavier collector loadings in order to float. If the oxidation is too fast, in the absence of activators irreversible depression takes place. Understanding the mechanism of arsenian pyrite flotation is particularly important given that it is the principal gold carrier in the very important submicron gold pyritic sedimentary-hosted gold deposits, also known as Carlin-type deposits (Thomas, 1997). To overcome the inadvertent oxidation of pyrite in the N2TEC process implemented at Twin Creek (NV, USA), grinding and flotation are carried out under nitrogen atmosphere using lead nitrate as the activating agent (Simmons, 1997). Barrick testwork on the Carlin ores demonstrated that the use of acidic flotation improved selectivity and kinetics. The success in maximizing recovery of gold-bearing arsenian pyrite fromCarlin-type ores lies in minimizing unwanted pyrite oxidation coupled with generous activation, while producinglower-grade sulfide concentrates by recovering middlings with finely disseminated gold-rich pyrite (see Figure5.16).

In the case of mesothermal pyritic gold ores, pyrite and arsenopyrite are sufficiently coarse grained, which allows for good liberation at modest grind fineness (P80 of 75120m) and results in high-grade concentrates with good recoveries. In some of these ores, submicroscopic gold is exclusively carried by arsenopyrite, which makes separation enticing from the barren pyrite (Donlin Creek, AK, USA), while in others submicroscopic gold is equally shared by arsenopyrite and arsenic-rich pyrite (Olympias, Greece).

Free gold recovery from the slimes fraction can be enhanced by adding some of the best-match collector in the regrind mill feed, to build up a heavier collector loading on already free tiny gold particulates as well as coat newlyliberated gold grains when their surfaces are still fresh. In ores with gold grains displaying a bimodal or broadsilverconcentration distribution, a matching collector will enhance flotation kinetics of the slower floating member.

Processes for the direct hydrometallurgical treatment of concentrates for recovery of PGMs, gold, and base metals have been developed but have not yet been applied commercially. This approach would avoid the need for conventional smelting, converting, and base-metal refining prior to refining of PGMs and gold.

The Platsol process was originally developed in Canada for the treatment of flotation sulfide concentrate for Polymet Mining Company in Minnesota and subsequently tested on other PGM concentrates. This process involved codissolution of base metals (copper and nickel), gold, and PGMs via addition of chloride to the autoclave. Grinding theore with ceramic rather than iron balls was required to prevent cementation of gold chloride (Ferron etal., 2000). Overall recoveries for the Northmet concentrate were Cu, Ni, Co >96%, Pd 95%, Pt 96%, and Au 89% (Ferron etal., 2003). The process underwent a detailed engineering phase following DFS and FEED studies for the Northmet project (Wardell-Johnson etal., 2009).

The Kell process comprises recovery of base metals by sulfuric acid pressure leaching, and a heat treatment may be applied to convert precious metalbearing minerals into forms that are soluble in a subsequent chlorination leach, at 9599% extraction efficiencies for the precious and base metals. The process has been patented (Liddell, 2003) and developed (Liddell etal., 2011; Liddell and Adams, 2012a,b), including integrated continuous pilot-scale testing on various PGM/Cu-Ni and polymetallic concentrates. Pallinghurst Resources is assessing potential construction of a full-scale Kell plant to extract base metals (copper, nickel, and cobalt) and PGMs as well as gold, at its Sedibelo Platinum Mines subsidiary (Seccombe, 2014).

Flotation circuit configuration on most gold mines can be divided into a number of categories, viz. open circuits with no cleaning at all, and open and closed circuits with single stage and two stages of cleaning. Open circuits have the advantage of no feedback from the effects of nonsteady-state operation and therefore are inherently more stable than the closed-circuit configuration. Closed- and open-circuit flotation cleaning is used on gold mines where high-grade concentrates are required for roasting and smelting. Under these conditions, it is difficult to maintain very high gold and sulfide flotation recoveries, while also producing an acceptable grade of concentrate. Where there is no constraint on concentrate quality, high gold and sulfide flotation recoveries are achievable to the extent that a discardable gold flotation tail is possible (Bax and Bax, 1993; O'Connor and Dunne, 1991). Cleaning-circuit configuration, either single or two stages of cleaning, and cleaner residence time are related to the particle size of the sulfides in the flotation feed and the presence or absence of floatable gangue components.

Unit flotation cells (Hasting, 1937; Taggart, 1945) and the more recent flash flotation cells (Kalloinen and Tarainen,1984) are installed in grinding circuits with the purpose of improving the overall flotation recovery of free gold (Taggart, 1945; Suttill, 1990; Laplante and Dunne, 2002). The aim is to remove as much as possible of the free gold contained in the circulating load of the grinding mill before it is overground or is covered with coatings of iron, sulfide, or other coatings that will lower flotation recoveries. Improved overall gold flotation recoveries of 210% have been quoted (Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990). Further, the inclusion of unit and flash flotation cells will generally provide better flotation stability and performance. Improved overall gold flotation recoveries from 3% to 10% have been quoted (Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990) for ores of variable gold and sulfide content (Taggart, 1945). Contrary to the belief of many, flash flotation does not recover particles coarser than that achieved by conventional flotation (Newcombe etal., 2012a,b). The particle size distribution of a flash flotation concentrate is typically coarser compared to conventional concentrates, with the fine particles being removed by a cyclone ahead of the flash flotation cell. The reasons for low coarse particle recovery are numerous, the major impediment being the relatively low concentrations of collector added in the flash flotation circuit. Much higher concentrations of collector are required to float very coarse particles (Bravo etal., 2005; Dunne, 2012); however, adding high levels of collector in a flash flotation circuit will compromise mineral separation selectivity in the downstream convention flotation circuit(s). Other items that impact coarse particle flotation include poor aeration capacity and bubble dispersion and solid residence times due to short circulating (Newcombe etal., 2013a,b).

Many coppergold concentrators have a combination of flash flotation and gravity separation in the milling circuit to enhance overall free gold recovery. The reason for the inclusion of gravity separation at these concentrators is that the coarse free gold is not recovered effectively in the flash flotation cell. Concentrators that have this circuit configuration have found that most of the fine gold is recovered in the flash flotation circuit, leaving the gravity circuit to recover the coarser gold. One of the dilemmas arising from laboratory testing for the combined circuit is that the standard gravity test procedure will overpredict plant gravity recovery, because fine free gold is captured in the laboratory test. To better predict both gravity and flash flotation free gold recovery in a concentrator a combined gravityflash flotation, a model has been developed by McGrath etal. (2013). The inputs for the model are obtained from a defined laboratory procedure for both the gravity and flotation separation steps.

Column flotation cells are used in roughing and cleaning duties on a number of mines treating gold ores (Lane and Dunne, 1987). A column cell typically provides higher concentrate grades compared to a mechanically agitated cell; however, losses of coarse gold may be higher in the column cell (Chryssoulis etal., 2003b). Testing combining high-intensity conditioning together with a three-product column produced high enrichment ratios of gold at acceptable recoveries when processing fine gold from low-grade tailings (Valderrama and Rubio, 2008).

Early work was carried out by Lakefield Research on thiosulfate leaching carbonaceous double refractory ores from Barrick Gold Corporation after pressure oxidation (POX) pretreatment (Thomas etal., 1998; Fleming etal., 2003). Up to 95% Au extraction was achieved from the finely divided gold left in the oxidized residue. In this process, POX residue leaves the autoclave at 35% solids and is directed to a leaching operation, where it is contacted with ammonium thiosulfate (5g/L) and copper sulfate (25ppm Cu) at 4060C and pH value 8. The slurry of gold-bearing leachate and solid residue leaving the leaching circuit contains in the range of 15ppm gold and is directed to an RIP circuit, where gold and copper are loaded onto a strong-base resin to 15kg/t Au and 1025kg/t Cu. Copper is eluted from the resin using ammonia thiosulfate (200g/L) and gold is eluted using potassium thiocyanate (200g/L). The copper-bearing eluate is returned to the leaching circuit, while the gold eluate is either electrowon or precipitated.

Jeffrey etal. (2008b) also reported gold recoveries of >95% from a pressure-oxidized flotation sulfide concentrate. In this work, an integrated RIP process was tested, using a gold thiosulfate process incorporating a novel sulfite-enhanced chloride-based elution of gold from the resin, together with electrowinning of gold, and recycle of eluate and resin. The favored thiosulfate leaching conditions were 5mmol/L copper sulfate and 50mmol/L ammonium thiosulfate at a pH value of 8.5, which allowed less generation of polythionates and their subsequent loading on the resin (0.1kg/t). Gold and copper loadings onto the strong-base resin were 2.5bkg/tAu and 2kg/tCu. Copper was eluted from the resin using 0.5 mol/L ammonia thiosulfate and gold was eluted using sodium chloride and sulfite mixture. The copper-bearing eluate was returned to the leaching circuit and the gold eluate was electrowon.

The ammonium thiosulfate leaching of pressure-oxidized carbonaceous ore and various sulfide concentrates has been recently evaluated by Breuer etal. (2014). A comparison cyanide leaching and thiosulfate of gold for various pressure-oxidized materials is shown in Figure28.10. Thiosulfate conditions, which favored gold extractions similar to or better than cyanide leaching, varied for different POX residues. Adding 10g/L of sodium chloride to the autoclave, which converts the metallic gold in the ore to an ionic form during pressure oxidation, enabled improved gold extraction, but thiosulfate leach conditions which favored maximum gold extraction was sensitive to pH conditions. The preparation of feed for pressure oxidation and ultimate products for various materials possibly influenced results. POX conducted in the absence of chloride produces a significant quantity of basic ferric sulfate which required neutralization with lime, whereas the addition of salt produces residues containing predominantly natrojarosite [NaFe3(SO4)2(OH)6]. The benefit of salt addition on gold extraction varied under alkaline and acid pressure oxidation treatment of carbonaceous sulfide ore residues. The results demonstrated that optimum POX/leaching conditions will vary depending on the residue being thiosulfate leached.

Figure28.10. Comparison of gold extraction for cyanide and thiosulfate leached pressure-oxidixed (POX) leached residues (a)without and (b)with NaCl for different ore and concentrate types (NaCN, 50mmol/L and air; ATS, 50mmol/L (NH4)2S2O3, 0.5mmol/L Cu(II) for pH 8.59.0; or 2mmol/L Cu(II) for pH 9.510.5).

Other recent studies have shown favorable gold extractions of up to 89% from a high-grade pressure-oxidized concentrate (Au 32g/t, Ag 12g/t, Fe 59%, S 21%, and As 19%) in 0.2mol/L thiosulfate and ammonia solution with 5% thiosulfate consumption; however, no comparison was made with cyanide to determine maximum gold recovery (Lampinen and Turunen, 2015).

Table15 describes the major findings and recommendations from the research conducted related to LCA of these metals. Cobalt is a valuable metal found in the earths crust which is widely used in industrial applications. Cobalt mining has a notable impact on human health due to cancer-causing elements which may cause heart disease, vision problem, etc. Farjana etal. conducted the LCA of cobalt extraction process. According to their study, cobalt extraction is harmful to eutrophication and global warming. Cobalt extraction requires a large amount of electricity which is detrimental to global warming and also is the blasting (Farjana etal., 2019c). Cemented carbide has higher hardness and higher corrosion resistance, mostly used for drilling tools and cutting tools. China is the leading producer of cemented carbide. The cemented carbide ore is mined from extraction, crushing, milling, gravity method grinding, sulphide flotation and roasting. In the hydrometallurgy stage, the cemented carbide ore is digested, filtrated, precipitated, extracted using solvent and finally crystalised. In the pyrometallurgy stage, the ore goes through calcination, hydrogen reduction and carburisation. In the powder metallurgy stage, the ore goes through powder milling, granulation and sintering. Furberg etal. conducted a cradle-to-gate LCA of cemented carbide production with cobalt, while the geographic location was non-Chinese (Canada and United States). Their study stated that impacts were due to elements like kerosene, tailings, water and electricity. The highest impacts were on the category of TAP (terrestrial acidification), ODP (ozone depletion), FEU (freshwater eutrophication). And the lowest impact was on CC (climate change), PCOF (photochemical oxidant formation) and WD (water depletion) (Furberg etal., 2019). Manganese is an essential element for batteries, fertilisers and chemicals. Manganese is a widely used alloying element that comes in conjunction to make ferroalloys. The manganese alloy is produced using mineral extraction, hauling, ore preparation and beneficiation, sintering and transportation, smelting, crushing, screening and refining. Westfall etal. conducted an LCA study based on manganese alloys, where datasets were collected from 16 ore and alloy producers. The authors have conducted a cradle-to-gate LCA of silicomanganese, ferromanganese and refined ferromanganese. The impact categories considered were GWP, AP, POCP, water and waste. The analysis was done using CML 2001 method. According to their analysis, electricity demand, fuel consumption during smelting was the primary contributors for impact (Westfall etal., 2016). Magnesium oxide cement is widely produced in China, North Korea, Turkey, Russia and Australia. The magnesium oxide is produced from raw material acquisition, crushing, vertical shaft kiln, precipitation tank, screening, crushing, grinding and packaging. Ruan etal. analysed the LCA of magnesium oxide, where the functional unit was 1 tonne. They showed that MgO has a lower impact on the ecosystem and resources but a larger impact on human health. The analysis was done using EcoIndicator 99 method. They considered five different case scenarios based on fossil fuel and raw material consumption (Ruan and Unluer, 2016). Silver metal is most widely used for industrial purposes or for making jewellery. There are very few studies which addressed the environmental impact of silver mining processes. Farjana etal. analysed the environmental burdens associated with goldsilverleadzinccopper beneficiation process (Farjana etal., 2019b). In another study, they analysed the environmental impacts of goldsilver refining operations (Farjana etal., 2019d). They found that silver beneficiation and refining have the least environmental impacts than gold mining processes as they consume the least amount of electricity. However, there are some impacts on eutrophication, global warming and ecotoxicity (Farjana etal., 2019b,e). Titanium oxides are widely used for making high-performance metal parts, artificial body parts and engine elements. Ilmenite and rutile are the generally found form of titanium oxides. Ilmenite and rutile are extracted from mining site using heavy mineral concentration, rare-earth drum separation, electrostatic separation circuit and gravity separation circuit. Farjana etal. conducted a comparative LCA analysis of cradle-to-gate titanium oxides production. Ilmenite and rutile were considered where the geographic region considered was for Australia. The datasets were collected from the AusLCI database and SimaPro software. The study revealed that rutile had a significant environmental impact than for ilmenite due to higher energy consumption and electricity use. The GHG was 0.295kg CO2 eq/kg of ilmenite production and 1.535kg CO2 eq/kg of rutile production (Farjana etal., 2018c).

In sulfide flotation, recovery and selectivity are fundamentally dependent on the relative rate constants of various mineral phases (Boulton et al., 2003). Therefore, an evaluation of the hydrophobicity balance by mineral particles requires accurate selection of the mineral phase. The hydrophobichydrophilic (hydrophobicity) balance by mineral phases and the relative statistical average require determination of the main species contributing to each category in surface layers. This determination is not a simple procedure in a flotation pulp containing diverse mineral phases, various mineral sizes, adsorption of various reagents, different products oxidation, precipitations (often colloidal), and polysulfide Sn2-species(resulting from loss of metal ions, usually Fe2+) on mineral surfaces (Smart et al., 2003a,b, 2007).

Numerous studies have been conducted to evaluate the hydrophobichydrophilic (hydrophobicity) balance by mineral phases (Vickers et al., 1999; Piantadosi et al., 2000, 2002; Duan et al. 2003). For adsorption studies in mineral flotation, quantification of surface species by TOF-SIMS and simply using the peak intensities of adsorbed and substrate signals are unsuitable (It does not take into account many of the matrix effects of mineral phases) (Piantadosi et al., 2000). To generalize, in the case of adsorption, the ion ratio of interest can be expressed as:

where RPI is the relative peak intensity, Iads is the integrated peak area of the ion fragment characteristic of the adsorbed molecule, and Isub is the integrated peak area of the ion fragment characteristic of the substrate. In principle, RPI is the relative peak intensity measured by TOF-SIMS, or RPI is the ideal parameter for adsorption studies since it has the character of , the traditional measurement of uptake (Iads) function of monolayer capacity (Iads+Isub), and might be expected to vary regularly with the extent of coverage of the substrate adsorbent by the adsorbate (Vickers et al., 1999).

This method of quantification yields a clearer illustration of the differences between concentrates and tails (Piantadosi et al., 2002). It is required to use Eq. (2) for each index (Vickers et al., 1999). Piantadosi et al. (2000) investigated the coverage of potassium isobutyl xanthate (IBX) and sodium diisobutyldithiophosphinate (DBPhos) adsorbed on the surface of galena by TOF-SIMS. They developed models which fully described both hydrophilic and hydrophobic indices of recovery of particles by flotation. An example of an initial development is described below:

Development of a more extensive hydrophobic/hydrophilic index may involve the ratios of a number of these indices, as shown above. For instance, the DBPhos/SO3- indices may be chosen as a first attempt at a hydrophobic/hydrophilic ratio. An alternative hydrophobic/hydrophilic ratio has been chosen to form a more direct overall index (I), using the Iads/Isub ratios.

Piantadosi et al. (2002) demonstrated that statistically, particles in the concentrate are more hydrophobic and separable than particles in the tail when both hydrophobic (collectors) and hydrophilic (oxidation products) species are combined (Piantadosi et al., 2002). Piantadosi et al. (2002) continued their surface analysis by TOF-SIMS with the aim of investigation on the particle-by-particle statistics of hydrophilic and hydrophobic species on the surfaces of mixed samples (galena and pyrite) under flotation-related conditions. Using a similar procedure, they found that in the concentrate the surface of galena have less Ca/Pb, PbOH/Pb and oxy-sulphur species (SO3/S) compared top articles in the tail. In other words, they were less hydrophilic. These differences are statistically considerable. Statistical results obtained for other species, such as Mg/Pb species, did not show any significant difference. This technique identified the effective species that correlate with flotation. Using a similar method, Duan et al. (2003) predicted an advancing contact angle of 71+2 (degrees) for the chalcopyrite particles in the Mount Isa Mines ore using the DTP/SO3 ratio as measured by TOF-SIMS.

A great number of base metal sulphide flotation plants use saline water. Some of them are summarised in Table 1. The three nickel flotation plants (Mt Keith, Leinster mine and Kambalda Nickel Concentrator) in Western Australia operated by BHP Billiton use bore water with high ionic strength. Table 2 shows the elemental compositions of the bore water used in the Mt Keith Operation, the largest nickel flotation plant in Western Australia (Peng and Seaman, 2011). The salinity of this water is several times higher than that of sea water. In the Raglan mine operated by Xstrata Nickel (formerly Falconbridge) in Northern Quebec, Canada, salt levels range from 20,000 to 35,000ppm throughout the year. An apparent consequence of the high salt content in the Raglan mine is that the flotation circuit is able to operate without the addition of frother (Quinn et al., 2007).

In Chile, many copper flotation plants use seawater. One example is Las Luces, a copper-molybdenum plant in Taltal, owned by the Las Cenizas Mining Group (Grupo Minero Las Cenizas) of Chile. In Las Luces, seawater is brought from a distance of 7km, mixed with tailing dam water and then used in grinding and flotation circuits. During the last 15years the increase in the total dissolved solid content of the process water in Las Luces was from approximately 36.0g/L (seawater) to 46.4g/L (Moreno et al., 2011). Another large copper plant using seawater in Chile is the Esperanza Concentrator at Sierra Gorda (Antofagasta Minerals S.A AMSA). This plant is processing ore at 95,000tpd using sea water without any pre-treatment. Sea water is pumped 145km from the Pacific Ocean to a 60,000m3 pool at the mine site located at an altitude of 2300m above sea level (Castro, 2012).

Other important cases are the Batu Hijau Concentrator (Newmont, operating from 2000) located at the Indonesian island of Sumbawa, and ayeli Bakr letmeleri A.. (CBI) in Turkey. The Batu Hijau Concentrator uses sea water for processing a gold-rich phorphyry copper ore (chalcopyrite-bornite) (Castro, 2012), while CBI processes a complex CuZn sulphide ore using dissolved metal ions and sulphide ions, mainly in the form of SO42 and S2O32 (Bak et al., 2012).

In South Africa, the recovery of platinum group elements (PGE) through the selective flotation of base metal sulphides also uses saline water. Flotation operations can be found in different parts of the Bushveld Complex, e.g., Merensky reef and Platreef. The ionic concentration varies in all Merensky concentrators, South Africa (Corin et al., 2011; Miettunen et al., 2012; Shackleton et al., 2007; Wiese et al., 2005a; 2005b; Wiese et al., 2007).

Although hydrothermal technology has great advantages in solid waste treatment, the industrial application of hydrothermal processes suffers from various challenges because of the severe process conditions. For example, corrosion requires the use of expensive alloys, and the high operation pressures put tough requirements on process components such as feed pumps [105,136].

However, Pilot test of neutralization slag by hydrothermal sulfide flotation indicated that hydrothermal technology is expected to realize industrial application in the future [137]. Therefore, some challenges and questions that should be solved. The environment of high temperature and high pressure is relatively dangerous. In the future, more studies should be conducted on how to moderate the hydrothermal conditions. It is necessary to find a proper catalyst to lower the reaction temperature, especially for the high-temperature HT process [62,138]. How to produce value-added materials is an urgent problem to be solved, especially how to prepare nanomaterials from solid waste [83]. There are relatively few studies on the corrosion of reactors by alkaline additives. Therefore, the choice of additives will also be a hot research topic in hydrothermal reactions.

different types of flotation cells

different types of flotation cells

Flotation is both a science and an art. It brings together many complex variables. Such basic factors as knowledge of mineral structure, chemical reagents, pH of mill water, pulp density, temperature, technical skills of the operator, the dependability of the flotation machine, as well as a host of other factors which affect the flotation of each specific ore must be combined to produce economic metallurgy. Economic Metallurgy is the practical objective of all mineral processing that of securing the greatest possible profit from the operation. It incorporates the elements of greatest possible recovery, highest possible grade of concentrates, lowest possible milling, capital, operating and maintenance costs.

In 1961 the American Institute of Mining and Metallurgical Engineers commemorated the 50th Anniversary of Flotation in the U.S.A. with a special 700- page documentary report. The book, Froth Flotation, 50th Anniversary Volume, highlights some of the many changes to flotation machines that have taken place since the first flotation cells in the United States went on stream in 1911 at Basin, Montana.

In 1911 the only mineral recovered by flotation was sphalerite. However, today flotation is the principal method of mineral processing throughout the world. Capacity of the first flotation mill was only 50 tons of ore per day. Now more than 100 different minerals are commercially recovered by flotation and some mills process as much as 50,000 tons of ore every 24 hours.

One of the significant changes made to reduce operating costs has been the extensive use of pressure molded rubber parts to withstand abrasion and thus reduce maintenance. Design of flotation cells has been improved and simplified to handle increased tonnage. One such development was the free-flow tank design. Another change has been the use of flotation mechanisms which can be removed from the flotation machine quickly for the inspection or change of wearing parts. A simple change in the method of pre-mixing air with pulp as it enters the throat of the flotation impeller has made it possible to reduce power cost as much as 50%. Development of larger flotation cells means fewer flotation cells are needed to do the job. This simplifies maintenance and reduces construction costs.

The cell-to-cell flotation machine meets the needs for both mineral recovery and cleaning and recleaning of flotation concentrates. It incorporates simplicity and flexibility of adjustment that permits the flotation operator to use his skill in securing the exact flotation conditions required by his specific ore for economic metallurgy.

The cell-to-cell flotation machine is typified by a flotation mechanism suspended in an individual cell separated from the adjoining cells by an adjustable weir. A feed pipe conducts the flow of pulp from the weir of the preceding cell to the mechanism.

Cell-to-cell Flotation Mechanism showing how feed pipe conducts pulp to throat of the rotating impeller. Each cell has its own mechanism, adjustable overflow weirs and feed pipe. Molded rubber wearing parts are used. Free-Flow Flotation Mechanism showing how the pulp flows through the machine without interruption of weirs. Feed pipes are not used. Pulp enters the throat of the rotating impeller by flowing down the outer feed well. Air, under low pressure, is pre-mixed with the pulp and is thoroughly diffusedthroughout the cell by intense action of the impeller.

A typical modern fluorspar mill where cell-to-cell Flotation Machines are used to clean and reclean fluorspar concentrates to meet market specifications for acid-grade fluorspar. Note pipes on launders return froth for multiple cleaning without need for pumps.

Early-day Sub-A Flotation Machine. Note the wood tank and flat-belt drive. Machines of this type were used in the 1920s and 1930s. They were the principal type of flotation machine used throughout the world. Experience and continual improvement are behind modern Flotation Cells.

The need for a flotation machine to handle larger tonnages in bulk flotation circuits led to the development of the Free-Flow type flotation machine. These units are characterized by the absence of intermediate partitions or weirs between cells. Individual cell feed pipes have been eliminated. Pulp is free to flow through the machine without interference. Flotation efficiency is high, operation is simple and the need for operator attention is minimized. Most high tonnage mills today use the free-flow type of flotation machine. Many are equipped with automatic devices for control of pulp density, pulp level, and other variable factors.

Just as modern flotation machines have evolved from the past they will change to meet future needs of the industry. Larger, more efficient flotation cells, automatic control of grinding circuits, flow meters, continuous on-stream sampling, direct reading density, pH, and pulp level devices, new reagents as well as instantaneous X-ray analysis will make possible almost completely automated flotation circuits and new achievements in economic metallurgy.

leachox process for flotation concentrates | maelgwyn mineral services

leachox process for flotation concentrates | maelgwyn mineral services

The majority of Aachen reactors are employed on Run of Mine (ROM) feed applications primarily to increase gold recovery through increasing the kinetics. In the Leachox process the role of the Aachen reactor is somewhat different in that the reactor is used to facilitate partial oxidation of the sulphide minerals encapsulating gold particles and is therefore normally used to improve the gold recovery on sulphide flotation concentrates.

Processes such as roasting, pressure oxidation, and bacterial oxidation are all aimed at breaking down the sulphide matrix to liberate gold. Ultra-fine grinding performs the same function particularly where gold is locked in silicates or other minerals. Many of these processes are well developed and can yield very high gold recoveries but unfortunately all tend to have inherent issues.

Roasting for instance is an environmentally unfriendly process and presents permitting issues in many countries; pressure oxidation requires a fairly high degree of operator skill and control not to mention often exotic materials of construction. Bacteria used in bacterial oxidation whether heap of tank leaching are susceptible to changes in environmental conditions and require careful control. Pressure oxidation works well but requires high levels of technical expertise and exotic materials of construction

Whilst it is not possible here to go into the relative merits and demerits of each process the biggest issue that they all have in common to some degree is the very high associated capital and operating costs limiting their application to large high grade deposits which can accommodate the high capex and opex

For many deposits where the resource base is too small to justify Pressure oxidation etc then Leachox can be used. Leachox is a partial oxidation process for refractory ores centred around the Aachen reactors. Whilst it does not yield as high a gold recovery as POX and Roasting it is order of magnitudes cheaper resulting in an overall improved project economics

Often when gold is locked in sulphides and silicates it is necessary to reduce the particle size to a size where gold is liberated or partially liberated. For refractory ores this generally translates into grinding below 10 microns and often to as low as 3-4 microns. Historically this was cost prohibitive as the only mills that were available to do this were tumbling mills which become highly inefficient at these low sizes particularly below 20 microns. The last 10-15 years however, have seen the development of a number of grinding mills specifically designed for ultra-fine grinding in the minerals industry. This has lead to a commensurate reduction in cost to grind fine and has been the catalyst for the development of many refractory ore treatment processes.

One of the drawbacks of grinding finer is that in addition to liberating the desired mineral it also increases the surface area of other host minerals .This can result in order of magnitude increases in reagent consumption particularly cyanide for the subsequent leaching process unless cognisance of this is taken in the final process design.

Whilst the installation is similar to that used for pre-oxidation or Aachen assisted leaching in the Leachox process the flotation concentrate is pumped through the reactor multiple times perhaps as many as 30 passes in contrast to 1-2 passes for the pre-oxidation role .Depending upon the mineralogy ultrafine grinding of the flotation concentrate may be required prior to cyanidation

In Leachox rather than just raising the DO level and providing shear to remove passivating species to enhance cyanidation kinetics the sulphide matrix is partially oxidised (60-70%) resulting in the liberation of gold particles from the host sulphide matrix.

Unfortunately the breakdown of the sulphide matrix along is associated with a significant increase in the surface area of the various cyanide consuming minerals and can result in order of magnitude increases in cyanide consumption often as high as 20-30kg/t which in itself is uneconomic. In addition, significant passivation also can occur rendering the mineral surface refractory to cyanide

The Aachen reactor solves both of these problems by accelerating the leach kinetics such that gold is able to dissolve into solution prior to the cyanide being consumed and also continuously removing the passivating layers forming on the gold particles (See information on Aachen reactors for pre-oxygenation and Aachen assisted Leaching)

Whilst as mentioned previously oxygen lances might be suitable for readily cyanidable oxide ores their limitations become apparent with the more demanding refractory leaching applications and their use is associated with very high cyanide and oxygen consumptions and poor gold recoveries

Imhoflot G-Cells are used to produce the flotation concentrate. The reason for this is that the patented G-Cell is able to produce a higher grade flotation concentrate than conventional mechanical flotation cells resulting in a small concentrate mass and so reduced treatment costs (see under Imhoflot flotation for more information)

As previously mentioned ultra fine grinding is often required on the flotation concentrates. Grinding is by its very nature an expensive process and more so for fine grinding where particle size reduction is through abrasion rather than impact. Whilst there are a number of commercial ultrafine grinding mills available Maelgwyn specifically designed its own mill the Ro-Star mill to reduce the capex and opex costs for ultra- fine grinding where this is required

The cyanide consumption in refractory gold leaching is generally much higher than in ROM ore cyanidation circuits and so efficient cyanide destruction is important. The Aachen reactors can be used to enhance the well-known sulphur dioxide based cyanide detox process

flotation '21

flotation '21

The 10th International Flotation Conference (Flotation '21) is organised by MEI in consultation with Prof. Jim Finch and is sponsored by Promet101, Maelgwyn Mineral Services, Magotteaux, Gold Ore, CiDRA Minerals Processing, Hudbay Minerals, Senmin, Clariant, BASF, Eriez, Nouryon, Festo, Newmont,Cancha, Zeiss,FLSmidthand Kemtec-Africa.

forced-air flotation cell | flsmidth

forced-air flotation cell | flsmidth

Flotation is about creating the proper energy dissipation rate in the cells to obtain optimal contact between the air bubbles and the particles for extracting the minerals. The function of the rotor/stator is to make bubbles from the forced air, suspend the particles, and create an environment for bubbles and particles to make contact and rise to the top as froth for concentration and collection.

Our forced-air flotation design features a streamlined, high-efficiency rotor that works as a very powerful pump. Working together the stator, these components generate an energy-intensive turbulence zone in the bottom of the cell. The forced-air design allows for control of the air flow. The well-defined turbulence zone results in multiple passes of unattached particles through the highest energy dissipation area of the cell where fine particles are driven into contact with the air bubbles.

The stator design, in addition to providing good separation of the cell zones, also serves to redirect the rotor jet uniformly across the tank. This allows dispersion, or distribution, of the maximum amount of air into the cell without disturbing the surface an important consideration for fine particle recovery. The air dispersion capabilities of our Dorr-Oliver cell design exceed all competitive forced-air designs.

By containing the intense circulation energy at the bottom of the cell, the upper zones of the cell remain quiescent, or passive, to maximise recovery of marginally attached coarse particles and minimise the carriage of undesired material.

We have equipped our forced-air flotation tank cells with a uniquely designed, high-efficiency radial launder system that accelerates froth removal as it reaches the surface. Bubble-particle aggregates travel vertically through the froth lattice. The high-efficiency radial launder is shaped to receive the froth uniformly from the cell surface, as well as from the typically heavy-loaded area near the centre of a forced-air machine. On passing over the lip, the froth accelerates to the perimeter of the cell. This unique design rarely requires launder water.

The two factors having the strongest impact on a flotation circuits performance are metallurgical recovery and flotation cell availability. Our forced-air flotation machines provide superior performance in both of these important areas, while offering additional, distinct advantages.

Superior metallurgical performance: Intense recirculation in a well-defined mixing zone multiplies the chances of contact between mineral particles and air bubbles, providing for greater mineral recoveries and higher concentrate grades.

Greater availability: Non-clogging design of the rotor reduces maintenance requirements, minimising failure, and increases availability. Our flotation mechanisms also can be removed for maintenance without process interruption.

Low reagent costs: Air is a natural reagent in the flotation process. Having a wide air dispersion capability permits you to fine-tune your flotation plant to deliver the optimum value for your process.

FLSmidth provides sustainable productivity to the global mining and cement industries. We deliver market-leading engineering, equipment and service solutions that enable our customers to improve performance, drive down costs and reduce environmental impact. Our operations span the globe and we are close to 10,200 employees, present in more than 60 countries. In 2020, FLSmidth generated revenue of DKK 16.4 billion. MissionZero is our sustainability ambition towards zero emissions in mining and cement by 2030.

flotation cell control - international mining

flotation cell control - international mining

The difficulty in process plants comes from the constantly changing feed characteristics, stringent product quality requirements and the economic need to maximize the recovery of a finite resource. A key part of successful plant control is the operation of the flotation circuit.

Flotation cells have three main control parameters (1) reagent dosing rate (2) froth depth and (3) air addition rate. Many other parameters may vary such as feed rate, particle size distribution and head grade, however these are the output of upstream processes and are not controlled in the flotation circuit itself.

The selection of reagent type and dosing rate is critical to successful processing of a given ore. It offers a coarse control mechanism as it is difficult to determine the impact of changes in either dosing rate or reagent type unless significant change in flotation performance is observed. In a relatively stable operation, the addition rate of reagents does not vary greatly. The operator seeks to ensure that a slight excess of reagent is available for the flotation process. Too much reagent, however, results in wastage and economic loss whilst too little results in either reduced grade or recovery and again economic loss. Thus, in a situation where the ore changes and marginally less reagent could be used, the operator generally should not chase this small reduction as it is difficult and time-consuming to optimize. The exception to this is where the ore change is expected to last for a long time.

Froth depth is fundamentally used to provide concentrate grade control. This occurs in two ways firstly, the depth determines the residence time in the froth phase and thus the time available for froth drainage.

Generally the greater the froth depth, the more drainage of entrained gangue (waste) and the richer the concentrate grade. There is a limit to the froth depth that a given flotation situation will support. If the froth gets too deep it begins to collapse on itself. The depth at which collapse begins is determined by the structure of the froth. Froth structure is driven by factors such as reagent type, reagent dose rate and the quality/level of mineral in the ore. Froth depth also plays a role in the recovery rate of the concentrate from the cell. As the froth gets deeper, the rate of froth removal reduces at a constant air addition rate. It is important to note that froth depth relationships are not linear in nature.

Once a froth depth has been established for a particular flotation duty (i.e rougher, cleaner etc), changes are generally small and infrequent. A flotation circuit where the slurry level is subjected to large or frequent changes is usually going to be in a constant state of flux as the changes in one cell will impact other cells in the circuit. Pump hoppers overflowing and flotation cell pulping are common symptoms of this.

Air addition rate offers the finest control of flotation cells. Small changes in concentrate recovery rate and grade can be achieved via changes in air addition rate. The impacts of changes in air addition rate are observed quickly in the plant providing a good source of operator feedback. Changes to air addition rate may be made several times in a normal shift as operators seek to optimise concentrator performance. As air addition represents a fine control method, changes should be small and one needs to wait several minutes before these results can be seen. Sudden large changes in air addition rate can create issues with level control as the pulp in the flotation cell will experience a rapid expansion and may overflow the cell launders. The ability to make regular changes to air addition rate in a convenient manner has led to automatic air control being the norm in modern concentrators. Changes in the concentrate grade that result from changes in air addition rate can be observed rapidly by utilising an on stream analysis system.

Leading-edge minerals processing plants incorporate automatic process control through some form of PID-driven system. In the case of flotation plants, the ideal system uses the three parameters discussed above to control a single parameter such as froth speed. Instruments such as FrothMaster use vision technologies to measure the speed of the froth over the lip. The desired froth speed can then be determined by monitoring the concentrate grade via an on stream analysis system. This type of control system automates the minute-to-minute running of the flotation circuit, which is driven by the desired concentrate grade. In plant trials, this approach has seen a significant improvement in recovery when compared to a manually monitored plant.

In figure 1, above, the variation between automatic control (Line 1) and manual control (Line 2) can easily be seen. In Line 2, the air addition rate is manually set so has to regularly monitor concentrate grade and recovery rate. This is time-consuming and is also not necessarily the best means of achieving the targeted set point. In Line 1, where the circuit was completely automatic, the control system constantly monitors and responds accordingly to any variations from the set point goal. In this particular trial, fully automated control brought real dollar benefits increasing overall recoveries, substantially reducing deviations from targets and reducing use of reagents in the cell.

Being able to successfully manage the different control parameters in a flotation circuit is a critical exercise in minerals recovery. Whilst the air addition rate offers the finest means of control, other parameters such as reagent type, reagent dosage rate and froth depth are also important controls for an operator to understand. It is also vital for an operator to understand the cause and effect relationships of these controls. Automated technologies such as on-stream analysers, along with newer developments such as FrothMaster or froth imaging systems, take this control to the next level. Not only do these automated systems monitor and analyse froth characteristics highly efficiently, but they can also optimise recovery, reduce reagent use and free up operator-time.

column flotation parameters

column flotation parameters

Investigation of coarse bubble column flotation resulted in identification of four distinct beneficiation zones of a flotation column. Comparison of average normalised percent mineral upgrading per foot values (zkn) indicated the relative degree to which these zones contribute to overall column mineral grades and recoveries. The pulp-froth interfacial zone produced the most critical Zkn values of fluorite upgrading and silica rejection.

Retention time in the column zones caused mineral grade and recovery variations as column parameters such as froth height, feed injection location, and column length were varied. The basis for column length design was determined to be particle retention time in the collection zone of the column.

Axial mixing problems encountered when scaling from a small- to a large-diameter column were simulated by recirculating a portion of the tailings stream to a point just below the feed injection port at different rates. This effectively broadened the particle retention time distribution in the collection zone. Extra column length is needed to compensate for this problem.

Coarse bubble column flotation proved superior to conventional flotation; column flotation produced much higher grade concentrates for all particle size fractions, with recoveries averaging slightly lower than those of conventional flotation. Column fluorite recoveries were increased over conventional fluorite recoveries for all size fractions after fine tuning of column parameters and bubble size.

Column flotation is rapidly gaining attention in the minerals industry as a potentially profitable beneficiation technique because it holds several distinct advantages over conventional flotation, as follows:

The effects of basic parameters such as column length, froth depth, feed injection location, wash water addition rate, and flotation behavior within the column have not been extensively investigated or documented. This report focuses or the key operating parameters affecting column flotation of fluorite and explores the beneficiation behavior of the column to identify apparent trends and correlations.

Since the inception of flotation columns in the early 1960s, the question of column length has been a consistent concern to commercial mineral processing plants anticipating installation and operation of flotation columns. The proposed theory is based on particle retention time. Column flotation is free from violent agitation and relies not only on feed slurry flow rate but also on particle free-settling rates, which are a function of the specific gravity of the ore being processed. Sufficient column length is necessary to equate particle settling time with nominal retention time.

The collection zone has its upper boundary at the feed injection port and extends downward to the base of the column. This zone must have sufficient length to provide adequate retention time for the settling particles to attach to the rising bubbles. This is the basis of column design. Additional column length is included as prescribed by particular mineral system needs when considering the upper three column zones. Host work backing this theory has teen performed on copper-molybdenum separations.

A test series was conducted to determine the effect of column length on fluorite grade and recovery by shortening the sectional column in 0.6-m (2-ft) increments from 5.5 to 1.2 m (18 to 4 ft) while maintaining the same ratio of cleaning zone to collection zone as nearly as possible within the physical limitations of the system. As the column flotation cell was shortened, recoveries decreased consistently (Fig. 5). This reduction in recovery substantiates the theory that column length is based on particle retention time. For our purposes, retention time was calculated as plug flow retention time based on collection zone volume and tailings flow rate. Recovery decreased because particle retention time was not sufficient as the collection zone was shortened by decreasing the column length (Fig. 6).

Fluorite grades increased with decreasing column length because only the particles with sufficient hydrophobicity and retention time to achieve bubble attachment reported to the concentrate stream. As the column was shortened the particle retention time decreased, causing smaller fractions of the more liberated fluorite to be collected while also reducing the amount of gangue that was either entrained or collected to the froth (Fig. 5 and 6).

Column length should be chosen to allow sufficient retention time in the collection zone for particle-bubble attachment with additional column length provided for the upper three zones. However, it should be noted that retention time required in a flotation column is usually much less than that required in a conventional flotation cell as evidenced by results achieved by a number of investigators including Foot (1985), McKay (1985), and Wheeler (1985). At present, no methodology exists for scaling up from conventional flotation data to commercial flotation columns. Laboratory column testing is an essential step in column scaleup. Dobby (1985) has developed a most extensive procedure for column scaleup based on laboratory column testing.

Experimentation was performed to determine the effect of the vertical location of the feed slurry injection port on fluorite grades and recoveries. Feed injection location was varied from 5.3 to 1.2 m (17.5. to 1 ft) from the base of the column flotation cell.

Fluorite recovery gradually decreased as the feed injection port was moved closer to the base of the column (Fig. 7). In essence, feed injection location is directly linked to retention time in the collection zone. As the feed injection location was moved toward the base of the column, the length of the collection zone decreased, reducing the particle retention time and resulting in decreased fluorite recoveries (Fig. 8).

Fluorite concentrate grades increased as feed location was moved to lower points on the column. Lowering the vertical position of the feed slurry injection port is equivalent to shortening the length of the column. This accounts for the similar response of fluorite grades and recoveries to variations in either column length or feed slurry injection location.

Feed injection location is chosen at the closest vertical position to the base of the column that provides sufficient column length necessary for adequate particle retention time to achieve bubble-particle attachment.

Unlike conventional flotation cells with limited froth depth control, column froth depths may be varied from a few inches to several feet. This provides an additional operating variable which may be used to control the flotation separation process. It was hypothesized that froth depth variations may affect product grade and possibly mineral recovery. Therefore, a series of tests was conducted to determine the effect of froth depth on fluorite grade and recovery. Variations in froth depth ranged from 0.2 to 3.2 m (0.5 to 10.5 ft).

Froth depth had a discernible effect on fluorite concentrate grades; as froth depth increased, fluorite grades increased (Fig. 9). Fluorite upgrading occurred in the froth phase cleaning and pulp-froth interfacial zones. The results of this investigation indicate that the froth phase is much more efficient than the pulp phase for mineral upgrading, and this improved separation efficiency becomes increasingly Important as the difference in hydrophobicity between the minerals to be separated decreases.

Fluorite recovery data were scattered. No trend was found to adequately describe the recovery data, and a best fit least squares linear regression equation shown in Fig. 9 (poor fit) through the data points had a flat slope. Although the source for these fluctuations was not found, it was apparent that they are not directly connected to changes in froth depth. It was concluded that froth depth had no primary correlation with fluorite recoveries.

Since increased froth depth enhanced fluorite grades without hindering recoveries, froth depth should be maintained at as great a depth as possible while still maintaining sufficient column length for the collection zone to provide the proper particle retention time to maintain mineral recoveries.

The principal reason wash water additions have been employed in column flotation systems is to increase the grade of the recovered concentrate by displacing entrained hydrophilic (gangue) particles that have reported to the froth phase. Typically, wash water addition rates from 1 to 15 pct of the volumetric feed slurry flow rate have been utilized in industry. A series of tests was performed to determine the effect of changes in wash water addition rates on fluorite rougher column flotation grades and recoveries. Wash water was introduced at addition rates from 0 to 1200 mL/min through a spray nozzle located 1 in above the top of the column. The teat points were normalized as to the volumetric feed slurry flow rates to the column, producing values ranging from 0 to 50 pct, respectively.

Wash water additions affected fluorite grades and recoveries during coarse bubble column flotation in a complex manner (Fig. 10). Increasing wash water addition from 0 to about 6 pct of the volumetric feed slurry flow rate increased column flotation fluorite grades but decreased fluorite recoveries. Increasing wash water addition from 6 to approximately 35 pct improved fluorite recovery but decreased fluorite grade. Above 35 pct wash water addition, fluorite grade again increased; however, recovery remained approximately the same. The complex effect of wash water additions on fluorite grades and recoveries may be attributed to the twofold function of column wash water; i.e., for removing gangue material reporting to the froth and fluidlzing the froth bed to prevent mineral overloading. As detailed in Fig. 10, the optimum wash water addition rate was recorded at 6 pct of the volumetric feed slurry flow rate. Wash water flow rates in excess of 35 pct of the volumetric feed slurry flow rate produced grades and recoveries that approached those at 6 pct; however, these additions increased water consumption, system dilution, and downstream materials handling problems.

The most hazardous problem in scaling from a small-diameter test column to a large-diameter industrial column, identified by Dobby (1981), is short circuiting of feed material caused by the change from plug flow to axial mixing conditions. Axial mixing broadens the particle retention time distribution, which results in mineral short circuiting and reduces recovery. This problem requires compensation by lengthening the collection zone in large-diameter columns. Mixing conditions were simulated in the 6.4-cm (2.5-in) diam test column by recirculating a portion of the tailings stream at various rates to a point just below the feed injection port. Tailings recirculation rates were varied from 0 to 7400 mL/min. These flow rates were converted to superficial velocities ranging from 1.47 cm/s with no recirculation to 5.37 cm/s at 7400 mL/min.

The degree of axial mixing was dependent on the rate of recirculation. Short circuiting of some of the feed slurry occurred as the degree of axial mixing increased because of the increased tailings recirculation rate. This resulted in decreased fluorite recoveries (Fig. 11).

Fluorite concentrate grades were enhanced by the presence of axial mixing. The increase of fluorite concentrate grade is characteristic of the shortened particle retention time caused by short circuiting of the feed slurry, which permitted only the fluorite particles with sufficient time and energy to attach to the rising bubbles and be concentrated. This phenomenon was also observed as the collection zone was shortened, as mentioned previously, and experimentally substantiates the theory of Dobby and Finch that the collection zone must be lengthened to provide sufficient particle retention time under increased axial mixing conditions.

Conventional rougher flotation was compared with coarse bubble rougher column flotation for separation efficiency of various particle size fractions of the Fish Creek fluorite ore. A single continuous column and two batch conventional flotation tests were performed with the products sized using Tyler 48-, 65-, 80-, 100-, 150-, 200-, 270-, 325-, and 400- mesh screens.

Column rougher flotation produced substantially higher grade concentrates than did conventional rougher flotation (Fig. 12). Column flotation fluorite grades were greater than conventional flotation grades for all size fractions. On the average, column fluorite concentrate grades were 34 pct higher than conventional fluorite concentrate grades.

Conventional flotation recoveries averaged 8 pct greater than those achieved using coarse bubble column flotation, but column flotation produced slightly higher recoveries for the minus 65-, plus 150-mesh material,. Recovery of the plus 48-mesh fluorite was drastically lower for column flotation than for conventional flotation, while that for the minus 150-mesh fluorite was only slightly lower (Fig. 13). According to Peterson (1986), fine tuning of column parameters and bubble size of a continuous column flotation unit showed increased column recoveries over conventional recoveries for all size fractions.

advanced flotation technology | eriez flotation division

advanced flotation technology | eriez flotation division

Eriez Flotation is the world leader in column flotation technology with over 900 installations. Columns are used for floating well-liberated ores. Typically they produce higher grade and have lower power costs than conventional cells. Applications include Roughers Scavengers Cleaners

Eriez Flotation is the world leader in column flotation technology with over 900 installations. Columns are used for floating well-liberated ores. Typically they produce higher grade and have lower power costs than conventional cells. Applications include

The HydroFloat fluidized bed flotation cell radically increases flotation recoveries of coarse and semi-liberated ores. Applications include: Split-feed flow-sheets Flash flotation Coarse particle recovery

The StackCell uses a 2-stage system for particle collection and froth recovery. Collection is optimized in a high shear single-pass mixing canister and froth recovery is optimized in a quiescent flotation chamber. Wash water can be used.

The StackCell uses a 2-stage system for particle collection and froth recovery. Collection is optimized in a high shear single-pass mixing canister and froth recovery is optimized in a quiescent flotation chamber. Wash water can be used.

The CrossFlow is a high capacity teeter-bed separator, separating slurry streams based on particle size, shape and density. Applications include: Split-feed flow-sheets with the HydroFloat Density separation Size separation

The rotary slurry-powered distributor (RSP) is used to accurately and evenly split a slurry stream into two or more parts, without creating differences based on flow, percent solids, particle size or density. Applications include Splitting streams for feeding parallel lines for any mineral processing application

The rotary slurry-powered distributor (RSP) is used to accurately and evenly split a slurry stream into two or more parts, without creating differences based on flow, percent solids, particle size or density. Applications include

Eriez Flotation provides advanced engineering, metallurgical testing and innovative flotation technology for the mining and minerals processing industries. Strengths in process engineering, equipment design and fabrication positionEriez Flotation as a leader in minerals flotation systems around the world.

Applications forEriez Flotation equipment and systems include metallic and non-metallic minerals, bitumen recovery, fine coal recovery, organic recovery (solvent extraction and electrowinning) and gold/silver cyanidation. The company's product line encompasses flotation cells, gas spargers, slurry distributors and flotation test equipment.Eriez Flotation has designed, supplied and commissioned more than 1,000 flotation systems worldwide for cleaning, roughing and scavenging applications in metallic and non-metallic processing operations. And it is a leading producer of modular column flotation systems for recovering bitumen from oil sands.

Eriez Flotation has also made significant advances in fine coal recovery with flotation systems to recover classified and unclassified coal fines. The group's flotation columns are used extensively in many major coal preparation plants in North America and internationally.

Eriez Flotation provides advanced engineering, metallurgical testing and innovative flotation technology for the mining and minerals processing industries. Strengths in process engineering, equipment design and fabrication positionEriez Flotation as a leader in minerals flotation systems around the world. Read More

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