flotation machine falls

flotation-machine

Flotation machine (also called Flotation separator) is applicable for the separation of nonferrous metal and ferrous metal and nonmetal, such ascopper, zinc, lead, nickel, gold, silver, lead ,fluorite and talc. The impeller is driven by V-belts, which can bring the centrifugal effect to form the negative pressure. On the one hand, to inhale sufficient air to mix with ore pulp; on the other hand, to stir ore pulp and mix with medication to form the mineralized froth. To adjust the height of flashboard to control the liquid level and make the useful froth scraped by loam board. +86 +86 E-mail:[emailprotected] Contact Us

When flotation machine works, slurry is inhaled from the bottom of the trough to the space between impellers. Meanwhile, the low pressure air send by fan is sent to this area through the air distributor in the hollow shaft. After sufficient mixing, the slurry is pushed out by the impeller, and then goes to the whole trough. When the froth rises to the stable level, after the enrichment processing, froth overflows to the froth trough from the overflow weir. Another part of ore slurry flows to lower part of impeller for the remixing with air. The remained slurry flows to the next trough until it becomes residue.

how to solve the common problems of froth flotation process? - xinhai

As the common mineral processing process, the froth flotation process is widely used in the concentrator. Because the froth flotation process is strict with grinding fineness, pulp density, flotation reagent, the flotation plant operators must inspect the flotation conditions during the operation of the flotation machine. Here are some common problems of froth flotation process and their and solutions.

This situation is owing to the substandard grinding fineness of material in the flotation machine or the dosage problem of flotation reagent. If the grinding fineness is not up to the target, the grinding fineness should be appropriately increased, which can increase the number of grinding stages or stage grinding.

If the problem of flotation reagent dosage arises, the flotation reagent should be tested and the dosage and index of flotation reagent should be redetermined. For example, a copper and cobalt oxide mine in Congo, the main copper content is malachite, and the cobalt mineral is hydrocobalt. In order to improve the froth flotation process effect, the sodium thiohydride and ammonia sulfide was used to improve the effect of the froth flotation process in a 1:1 ratio.

When the flotation tank appears larger foam, and the foam has the reflective surface, the possible reason is that the flotation reagent is added too much. At this time, the flotation plant operator should adjust the flotation reagent dosage appropriately, and observe whether the flotation foam condition is improved.

There are four possible reasons for this situation, one is the overflow concentration is too dilute, two is the position of the liquid level adjustment plate is too high, three is the slurry overflow concentration is too thick, which may also lead to coarse size and serious sand sinking, four is the dosage of inhibitor and regulator is too large, resulting in the failure to produce the appropriate amount of foam. When this happens, the flotation operator should first observe the situation, control the pulp concentration and fineness, and pay attention to control the dosage of regulator and inhibitor to get the better index of the froth flotation process.

The possible reasons for the falling of slurry level include the grinding fineness is too thick, the processing capacity is increased, or the reagent dosage is insufficient. The flotation operator can adjust the grinding and classifying, reduce the amount of ore feeding, or adjust the reagent dosage.

Generally, this problem of froth flotation process has the following reasons: the ore feeding is too much, the inflating volume is too much, the middle box and the liquid level adjustment plate is blocked, the suction capacity of the suction tank is insufficient, the pressure of flushing water is too small. In these cases, the blockage of the liquid level adjustment plate is the most likely to cause a sand setting, the flotation operator can minimize the level adjustment plate, dredge clogs, open the bottom valve if necessary, to make the slurry flow and check whether the pipe is blocked. At the same time, the flotation operator should know the amount of ore feeding and overflow concentration, timely find out the reason, and adjust to ore feeding according to reasons.

When this phenomenon occurs, the flotation operator should first stop the ore feeding, then open the bottom flow valve at the bottom of the tank, and then start the ore feeding and block the bottom flow valve when the slurry flow returns normally. At the same time, pay attention to control the slurry concentration.

The above are seven common problems and their solutions in the froth flotation process. For the small and medium mineral processing plant, it is suggested to choose the turn-key mineral processing solution provider from the mineral processing test to the mine operation, so that the above problems can be solved effectively, the flotation operator can also be effectively guided, which reduces problems in the subsequent froth flotation process.

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 machine - an overview | sciencedirect topics

Industrial flotation machines can be divided into four classes: (1) mechanical, (2) pneumatic, (3) froth separation, and (4) column. The mechanical machine is clearly the most common type of flotation machine in industrial use today, followed by the rapid growth of the column machine. Mechanical machines consist of a mechanically driven impeller, which disperses air into the agitated pulp. In normal practice, this machine appears as a vessel having a number of impellers in series. Mechanical machines can have open flow of pulp between each impeller or are of cell-to-cell designs which have weirs between each impeller. The procedure by which air is introduced into a mechanical machine falls into two broad categories: self-aerating, where the machine uses the depression created by the impeller to induce air, and supercharged, where air is generated from an external blower. The incoming slurry feed to the mechanical flotation machine is introduced usually in the lower portion of the machine.Figure 7 shows a typical industrial flotation cell of each air delivery type.

The most rapidly growing class of flotation machine is the column machine, which is, as its name implies, a vessel having a large height-to-diameter ratio (from 5 to 20) in contrast to mechanical cells. The mechanism behind this machine to is provide a countercurrent flow of air bubbles and slurry with a long contact time and plenty of wash water. As might be expected, the major advantage of such a machine is the high separation grade that can be achieved, so that column cells are often used as a final concentrate cleaning step. Special care has to be exercised in the generation of fine air bubbles and controlling the feed rate to column cells.

Good mixing of pulp. To be effective, a flotation machine should maintain all particles uniformly in suspension within the pulp, including those of relatively high density and/or size. Good mixing of pulp is required for maximizing bubble-particle collision frequency.

Appropriate aeration and dispersion of fine air bubbles. An important requirement of any flotation machine is the ability to provide uniform aeration throughout as large a volume of the machine as is possible. In addition, the size distribution of the air bubbles generated by the machine is also important, but experience has shown that the proper choice of frother type and dosage generally dominates the bubble size distributions being produced.

Sufficient control of pulp agitation in the froth zone. As mentioned earlier, good mixing in the machine is important; however, equally important is that near and in the actual froth bed at the top of the machine, sufficiently smooth or quiescent pulp conditions must be maintained to ensure suspension of hydrophobed (collector coated) particles.

Efficient mass flow-mechanisms. It is also necessary in any flotation machine that appropriate provisions be made for feeding pulp into the machine and also for the efficient transport of froth concentrate and tailing slurry out of the machine.

Probably the most significant area of change in mechanical flotation machine design has been the dramatic increase in machine size. This is typified by the data ofFig. 8, which shows the increase in machine (cell) volume size that has occurred with a commonly used cell manufactured by Wemco. The idea behind this approach is that as machine size increases, both plant capital and operating costs per unit of throughput decrease.

The throughput capabilities of various cell designs will vary with flotation residence time and pulp density. The number of cells required for a given operation is determined from standard engineering mass balance calculations. In the design of a new plant, the characterization of each cell's volume and flotation efficiency is generally calculated from performing a laboratory-scale flotation on the same type of equipment on the ore in question, followed by the application of empirically derived design (scale-up) factors. Research work is currently under way to improve the understanding and performance of commercial flotation cells. Currently, flotation-cell design is primarily a proprietary function of the various cell manufacturers.

Flotation plants are built in multiple cell configurations (called banks), and the flow through various banks is adjusted in order to optimize plant recovery of the valuable as well as the valuable grade of the total recovered mass from flotation. This recovery vs grade trade-off is economically important in flotation, as increased recovery of the valuable is associated with decreased grade. For example, a 95% recovery of copper in the feed ore might give a concentrate grade of 18% Cu in the total recovered mass, while 80% Cu recovery might give a grade of 25% in the concentrate. Obviously, the higher the valuable recovery is, the higher the potential income, but if this higher recovery requires a great deal more grinding and/or expensive downstream processing (including further flotation) in order to upgrade the concentrate for metal refining such as smelting, the increase in potential recovery income may actually cause a net loss of total income. This grade-recovery optimization is generally worked out by individual flotation operators in each plant (and each mineral) and sets the operating philosophy of that plant.Figure 9 shows a typical industrial recovery vs grade trade-off curve for a copper sulfide ore containing pyrite. The higher the copper recovery is, the greater the amount of undesired pyrite contained in the concentrate.

The various banks of flotation cells in an industrial plant are given special names to denote the particular purpose of the banks. The rougher bank is the first group of cells that the pulp sees after size reduction. The goal of the roughers is to produce a concentrate with as high a recovery of valuable as possible with generally low grade of the valuable. The rejected gangue material from any bank of cells is commonly denoted as the tails or tailings. Usually, rougher tails are discarded so that valuable mineral not recovered in the rougher bank is lost. The concentrate of the rougher bank can be further concentrated, sometimes after additional grinding, in banks of cells called cleaners or recleaners. The tailings from the cleaners or recleaners can be recirculated to a bank of cells known as scavengers in order not to lose any valuable material in the upgrading process. Various banks of cells are also sometimes known by the particle size of the particular pulps being floated. Coarse particles, fine or slime particles, and middle-sized particles, denoted as middlings, can all be treated in separate banks.

As to overall capacities of flotation plants, the range is quite variable, depending on the type and value of the mineral being processed, the amount of valuable mineral in the feed ore to flotation, the degree and cost of size reduction involved, and the relative response of the valuable(s) to the flotation process. Smaller plants ranging in size from 500 to 5000 metric tons of feed per day are common, with feed materials having high amounts of valuable per ton of feed ore (>40%), such as coal, phosphate, and oxide ores. On the other hand, the sulfide minerals that are typically a small percentage of the ore (<10% and often less than 1%) require much greater capacity in order to achieve a reasonable economic return on investment. Thus, typical copper sulfide plants have capacities in the range of 20,000 to more than 60,000 metric tons of feed ore per day.

Conventional flotation machines house two functions in a single vessel: an intense mixing region where bubbleparticle collision and attachment occurs, and a quiescent region where the bubbleparticle aggregates separate from the slurry. The reactor/separator machines decouple these functions into two separate (or sometimes more) compartments. The cells are typically considered high-intensity machines due to the turbulent mixing in the reactor (see Section 12.9.5). The role of the separator is to allow sufficient time for mineralized bubbles to separate from the tailing stream which generally requires relatively short residence time (when compared to mechanical cells or columns).

Some of the earliest machine designs were of the reactor/separator-type. Figure 12.80 shows a design from a patent by Hebbard (1913). Feed slurry was mixed with entrained air in an agitation box (reactor) and flowed into the separation vessel where froth was collected as overflow. The design would be the basis for the Minerals Separation Corporation standard machine and early flotation cells used in the United States (Lynch et al., 2010).

The Davcra cell (Figure 12.81) was developed in the 1960s and is considered to be the first high-intensity machine. The cell could be thought of as a column or reactor/separator device. Air and feed slurry are contacted and injected into the tank through a cyclone-type dispersion nozzle, the energy of the jet of pulp being dissipated against a vertical baffle. Dispersion of air and collection of particles by bubbles occurs in the highly agitated region of the tank, confined by the baffle. The pulp flows over the baffle into a quiescent region designed for bubblepulp disengagement. Although not widely used, Davcra cells replaced some mechanical cleaner machines at Chambishi copper mine in Zambia, with reported lower operating costs, reduced floor area, and improved metallurgical performance.

Several attempts have been made to develop more compact column-type devices, the Jameson cell (Jameson, 1990; Kennedy, 1990; Cowburn et al., 2005) being a successful example (Figure 12.82). The Jameson cell was developed in the 1980s jointly by Mount Isa Mines Ltd and the University of Newcastle, Australia. The cell was first installed for cleaning duties in base metal operations (Clayton et al., 1991; Harbort et al., 1994), but it has also found use in coal plants and in roughing and preconcentrating duties. The original patent refers to the Jameson cell as a column method, but it can also be considered a reactor/separator machine: contact between the feed and the air stream is made using a plunging slurry jet in a vertical downcomer (the reactor), and the airslurry mixture flows downwards to discharge and disengage into a shallow pool of pulp in the bottom of a short cylindrical tank (the separator). The disengaged bubbles rise to the top of the tank to overflow into a concentrate launder, while the tails are discharged from the bottom of the vessel. Air is self-aspirated (entrained) by the action of the plunging jet. The air rate is influenced by jet velocity and slurry density and level in the separator chamber.

The Jameson cell has been widely used in the coal industry in Australia since the 1990s. Figure 12.83 shows a typical cell layout where fine coal slurry feeds a central distributor which splits the stream to the downcomers. Clean coal is seen overflowing as concentrate from the separation vessel. The major advantage of the cell in this application is the ability to produce clean concentrates in one stage of operation by reducing entrainment, especially when wash water is used. It also has a novel application in copper solvent extraction/electrowinning circuits, where it is used to recover entrained organic droplets from electrolyte (Miller and Readett, 1992).

The Contact cell (Figure 12.84) was developed in the 1990s in Canada. The feed slurry is placed in direct contact with pressurized air in an external contactor which comprises a draft tube and an orifice plate. The slurryair mixture is fed from the contactor to the column-type separation vessel, where mineralized bubbles rise to form froth. Contact cells employ froth washing similar to conventional flotation columns and Jameson cells. Contact cells have been implemented in operations in North America, Africa, and Europe.

The IMHOFLOT V-Cell (Figure 12.85(a)) was developed in the 19801990s and evolved from earlier designs developed in Germany in the 19601970s (Imhof et al., 2005; Lynch et al., 2010). Conditioned feed pulp is mixed with air in an external self-aeration unit above the flotation cell. The airslurry mixture descends a downcomer pipe and is introduced to the separation vessel via a distributor box and ring pipe with nozzles that redirect the flow upward in the cell. The separation vessel is fitted with an adjustable froth crowding cone which can be used to control mass pull. The concentrate overflows to an external froth launder, while the tailings stream exits at the base of the separation vessel. The V-Cell has been used to float sulfide and oxide ores with the largest operation being an iron ore application (Imhof et al., 2005).

The IMHOFLOT G-Cell (Figure 12.85(b)) was introduced in 2001 and employs the same external self-aerating unit as the V-Cell. The airslurry mixture which exits the aeration unit is fed to an external distributor box (located above the separation vessel) where pulp is split and fed to the separation vessel tangentially via feed pipes. The cell is unusual as an internal launder located at the center of the vessel collects froth. The centrifugal motion of the slurry enhances froth separation with residence times being ca. 30s.

The Staged Flotation Reactor (SFR) (Figure 12.86) is a recent development in the minerals industry. By sequencing the three processesparticle collection, bubble/slurry disengagement, and froth recoveryand assigning each to a purpose-built chamber, the SFR aims to optimize each of the three processes independently.

The SFR incorporates an agitator in the first (collection) chamber designed to provide high energy intensity (kWm3) and induce multiple particle passes through the high shear impeller zone, hence giving high collection efficiency. Slurry flows by gravity through the reactor stages, that is, there is no need to apply agitation to suspend solids, only for particle collection. As such, impeller speed can be adjusted online in correlation with desired recovery without sanding. The second tank is designed to deaerate the slurry (bubble disengagement) and rapidly recover froth to the launder without dropback. The froth recovery unit is tailored for use of wash water and for high solids flux. Efficient particle collection and high froth recovery translate into fewer, smaller cells, resulting in a smaller footprint and building height, with lower power consumption, and the potential for good selectivity in both roughing and cleaning applications.

Induced air flotation machines have gained a degree of popularity within certain sections of the minerals processing industry because of their ability to produce small bubbles at relatively high energy efficiency. The most common of such machines is the Jameson Cell. A downcomer protrudes out of the bubbly liquid in which is housed a plunging jet. Because this jet is at high velocity the pressure within the downcomer is low due to the Bernoulli equation, and air is induced into the downcomer creating a plume of bubbles within the liquid, which rise to form a foam. There are major problems with operating Jameson Cells because their high demand for surfactant causes downstream residual frother issues. (It is noted, as an aside, that frother strippers are being developed to remove residual frother in flotation circuits, and these are identical to foam fractionation units.) Notwithstanding that Jameson Cell technology has failed to live up to its promise, it has been successfully used as a pilot-scale foam reactor to effect the autothermal thermophilic aerobic digestion (ATAD) of high strength wastewater sludge produced at a chicken processing factory. The advantage that induced gas systems have over alternative pneumatic foam systems is their very high gasliquid surface area per unit volume of foam due to their small bubbles. This feature of the foams would also be an advantage in foam fractionation because it creates high flux of gasliquid surface. However, to the authors knowledge, no attempt has ever been made to use induced gas systems as foam fractionators.

The Denver DR flotation machine, which is an example of a typical froth flotation unit used in the mining industry, is illustrated in Figure 1.47. The pulp is introduced through a feed box and is distributed over the entire width of the first cell. Circulation of the pulp through each cell is such that, as the pulp comes into contact with the impeller, it is subjected to intense agitation and aeration. Low pressure air for this purpose is introduced down the standpipe surrounding the shaft and is thoroughly disseminated throughout the pulp in the form of minute bubbles when it leaves the impeller/diffuser zone, thus assuring maximum contact with the solids, as shown in Figure 1.47. Each unit is suspended in an essentially open trough and generates a ring doughnut circulation pattern, with the liquid being discharged radially from the impeller, through the diffuser, across the base of the tank, and then rising vertically as it returns to the eye of the impeller through the recirculation well. This design gives strong vertical flows in the base zone of the tank in order to suspend coarse solids and, by recirculation through the well, isolates the upper zone which remains relatively quiescent.

Froth baffles are placed between each unit mechanism to prevent migration of froth as the liquid flows along the tank. The liquor level is controlled at the end of each bank section by a combination of weir overflows and dart valves which can be automated. These units operate with a fully flooded impeller, and a low pressure air supply is required to deliver air into the eye of the impeller where it is mixed with the recirculating liquor at the tip of the air bell. Butterfly valves are used to adjust and control the quantity of air delivered into each unit.

Each cell is provided with an individually controlled air valve. Air pressure is between 108 and 124 kN/m2 (7 and 23 kN/m2 gauge) depending on the depth and size of the machine and the pulp density. Typical energy requirements for this machine range from 3.1 kW/m3 of cell volume for a 2.8 m3 unit to 1.2 kW/m3 for a 42 m3 unit.

In the froth flotation cell used for coal washing, illustrated in Figure 1.48, the suspension contains about 10 per cent of solids, together with the necessary reagents. The liquid flows along the cell bank and passes over a weir, and directly enters the unit via a feed pipe and feed hood. Liquor is discharged radially from the impeller, through the diffuser, and is directed along the cell base and recirculated through ports in the feed hood. The zone of maximum turbulence is confined to the base of the tank; a quiescent zone exists in the upper part of the cell. These units induce sufficient air to ensure effective flotation without the need for an external air blower.

Most of the industrial flotation machines used in the coal industry are mechanical, conventional cells. These machines consist of a series of agitated tanks (usually 48 cells) through which fine coal slurry is passed. The agitators are used to ensure that larger particles are kept in suspension and to disperse air that enters down through the rotating shaft assembly (Fig. 11.13). Air is either injected into the cell using a blower or drawn into the cell by the negative pressure created by the rotating impeller. For coal flotation, trough designs that permit open flow between cells along the bank are more common than cell-to-cell designs that are separated by individual weirs.

Some of the major manufacturers of flotation equipment include Wemco (FLSmidth), Metso, Svedala, and Outokumpu. The commercial units are very similar in basic design and function, although some slight variations exist in terms of cell geometry and impeller configuration. Machines with large specific surface areas are generally preferred for coal flotation, due to the fast flotation kinetics of coal and the large froth solids loadings. Flotation machines with individual cell volumes of up to 28m3 are commonly used due to advantages in terms of capital, operating and maintenance costs. Some manufacturers also offer tank machines, which consist of relatively short cylindrical tanks equipped with conventional impellers. The simplified structural design, which allows these machines to be much larger, can offer significant savings in terms of capital and power costs for some installations. Tank cells with volumes as large as 100m3 are already in operation at coal plants in Australia.

Unlike conventional, mechanically agitated flotation machines, which tend to use relatively shallow rectangular tanks, column cells used in the coal industry are usually tall vessels with heights typically ranging from 7 to 16m depending on the application. Unlike conventional flotation machines, columns do not use mechanical agitation and are typically characterized by an external sparging system, which injects air into the bottom of the column cell. The absence of intense agitation promotes higher degrees of selectivity and can aid in the recovery of coarse particles.

In general, feed slurry enters the column at one or more feed points located in the upper third of the column body and descends against a rising swarm of fine bubbles generated by the air sparging system (Fig. 11.14). Hydrophobic particles that collide with, and attach to, the bubbles rise to the top of the column, eventually reaching the interface between the pulp (collection zone) and the froth (cleaning zone). The location of this interface, which can be adjusted by the operator, is held constant by means of an automatic control loop that regulates a valve on the column tailings line. Varying the location of the interface will increase or decrease the height of the froth zone. The froth is transported from the froth zone into the product launder via mass action.

Methods of sparging in columns are numerous and include air lances, porous tubes, eductors, static mixers, and Cavitation-TubesTM. The air rate used in a column is selected according to the feed rate and concentrate-production requirements. This parameter typically has the largest effect on the operating point of the column with respect to the ash/yield curve. The bubbles generated by the air sparging system are sized to provide the maximum amount of bubble surface area given a constant energy input. In other words, the designs of the various sparging devices are engineered to provide the smallest size and largest number of bubbles possible.

For an equivalent volumetric capacity, the cross-sectional surface area of a column cell is much smaller than that of a conventional cell. This reduced area is beneficial for promoting froth stability and allowing deep froth beds to be formed. This is an important aspect of column flotation, as a deep froth bed facilitates froth washing for the removal of unwanted impurities from the float product. Wash water, added at the top of the column, percolates through the froth zone displacing dirty process water and non-selectively entrained particles trapped between the bubbles. In addition, froth wash water serves to stabilize and add mobility to the froth. Sufficient water must be added to ensure that all of the feed water that would otherwise normally report to the froth product is replaced with fresh or clarified water. It has been reported that less than 1% of the feed pulp and associated clays will report to the froth in a well-operated column (Luttrell et al., 1999). The ability to maintain and wash a deep froth layer is the main reason cited for the improved product grades when comparing column cells to conventional cells.

In contrast, conventional mechanical cells do not operate with deep froths. Therefore, these devices allow some portion of the ultrafine mineral slimes to be recovered with the water that reports to the froth. Consequently product quality is reduced by this non-selective hydraulic conveyance (i.e., entrainment) of gangue into the product launder. In fact, fine particles (<0.045mm) have a tendency to report to the froth concentrate in direct proportion to the amount of product water recovered. As such, the flotation operator is often forced to make the decision to either pull hard on the cells to maintain yield (e.g., wet froth), or run the cells less aggressively to maintain grade (e.g., dry froth).

The primary advantage of utilizing wash water is the ability to provide a superior product grade when compared to conventional flotation processes. This capability is illustrated by the test data summarized in Fig. 11.15, which compares column flotation technology with an existing bank of conventional cells. As shown, the separation data for the column cells utilizing wash water are far superior to those obtained from the conventional flotation bank. In fact, the data for the column cells tend to fall just below the separation curve predicted by release analysis (Dell et al., 1972). A release analysis is an indication of the ultimate flotation performance and is often regarded as wash-ability for flotation. This figure suggests that columns provide a level of performance that would be difficult to achieve even after multiple stages of cleaning by conventional machines.

There are a significant number of full-scale column installations currently in commercial service around the world. The most popular brands of columns include the CPT CoalPro (Eriez), Jameson, and Microcel columns. Although the Jameson cell does not have the traditional column geometry, it is included since it typically uses wash water to improve ash rejection. Details related to the specific design features of the various column technologies are available in the literature (McKay et al., 1988; Finch and Dobby, 1990; Yoon et al., 1992; Manlapig et al., 1993; Davis et al., 1995; Rubinstein, 1995; Wyslouzil, 1997). The primary difference between the various columns used in the coal industry is the type of air sparging system employed. These include porous bubblers, static mixers, and dynamic air injectors. Details related to the features and operation of these systems have been discussed extensively in the literature (Dobby and Finch, 1986a; Xu and Finch, 1989; Huls et al., 1991; Groppo and Parekh, 1992; Yoon et al., 1992; Finch, 1995). Ideally, the spargers should produce small, uniformly sized bubbles at a desired aeration rate. Other factors, such as equipment costs, mechanical reliability, wear resistance, and serviceability also need to be carefully considered prior to selecting an industrial sparging system.

Due to economy of scale, recent trends in the coal industry have shifted away from the installation of large numbers of smaller units toward fewer, large units with diameters up to 5m or more. Although most column installations involve the treatment of particles finer than 0.150mm, several recent column operations have been installed to treat coarser particles, such as minus 1mm feeds or deslimed 0.1500.045mm feeds. Additionally, a move to more economical cells in terms of energy efficiency has been realized as manufacturers focus on the generation of the required air bubble dispersions while using significantly less power than traditional approaches. One such device is the Eriez StackCell, which utilizes both pre-aeration methods in conjunction with traditional froth washing (Davis et al., 2011) to maximize efficiency with regard to both installation and operating cost.

The two most important requirements of laboratory flotation machines are reproducibility and performance similar to commercial operations. These two criteria are not always satisfied. The basic laboratory machines are scaled down replicas of commercial machines such as Denver, Wemco and Agitair. In the scale down, there are inevitable compromises between simplification of manufacture and attempts to simulate full scale performance. There are scaling errors, for example, in the number of impeller and stator blades and various geometric ratios. Reproducibility in semi-batch testing requires close control of impeller speed, air flow rate, pulp level and concentrate removal.

Until now, deaeration tanks always had to be placed underneath the flotation machine and also frequently in the cellar of a facility in order to ensure a sufficient height difference for the conveyance of foam. In addition, the tanks are open on top and can overflow with excess foam. That is now a thing of the past with the Deaeration Foam Pump (DFP) 4000. The new pump can be linked directly to the deinking machine and forms a clean and closed disposal system. Because it can be placed at the same level as the flotation cells, the entire flotation system saves more space than previous systems. A cellar or an additional floor height for the flotation is no longer required. The deaeration results are very impressive with the DFP 4000 from Voith Paper. The air content of the foam mass is reduced when passing through the pump from 80% to an average of 8%. Conventional deaeration systems offer approximately 12%. In addition, by using the DFP 4000, upstream foam destroyers, downstream long piping as well as pumps with high head pressures to overcome the floor height can be dispensed. With the DFP 4000, it is possible to deaerate and convey the foam, which is loaded with inks and other impurities, within a single machine. As a compact unit, it fully replaces the foam destroyer, foam tank stirring unit, and pump of previous deaeration systems. This means a clear reduction in investment costs for the tank, stirring unit, pipes, pumps, and floor space.

The DFP 4000, developed by Voith, is a compact unit that integrates several elements of the flotation deinking system. This combines the pump and deaeration machine into one unit. The deaeration foam pump replaces the foam destroyer, foam tank, stirring unit, and pump and costs less than the current suite of equipment. The DFP 4000 achieves better deaeration of the foam than conventional systems.

The DFP 4000 has two parts. In the upper part, foam is predeaerated by a mechanical foam destroyer. In the lower part, centrifugal force produced by a quick rotational movement further deaerates the foam. The resulting low-air-content suspension is brought to the required pressure so that it can be conveyed out of the machine to the next process stage. The air released during deaeration is conveyed out of the machine through a special air chamber on the side so that the airflow does not prevent the foam entering from above (Dreyer,2010).

The new pump can be linked directly to the deinking machine, forming a clean and closed disposal system. Because the deaeration pump can be placed at the same level as the flotation cells, the entire system requires less space than previous systems, so a cellar (or additional floor height) is no longer required to accommodate the system. When the foam mass passes through the DFP 4000, the foams air content is reduced from 80% to an average of 8% (Voith,2011a). Conventional deaeration systems reduce the air content to approximately 12%. The first DFP 4000 operating in a paper mill has been in service since September 2009 (Dreyer,2010). The benefits of the DFP 4000 are summarized in Table11.9 (Dreyer,2010; Voith,2011a).

Batch testing has been carried out using a specially designed 21 tumbler for mixing, and a standard Denver flotation machine for separation. A typical charge of the soil sample ranged from 200 to 600g, and the amount of coal varied depending on the contaminant concentration.

Figure 1 shows the block diagram of the 6T/day continuous unit. A slurry of contaminated soil and coal is fed at optimal solids concentration to a specially designed tumbler. In the front section of the tumbler, as a result of rotary motion, the solids are mixed and dispersed. In another section of the tumbler, layering, compaction and abrasion take place. After being discharged from the tumbler, the contents are screened into two streams. The 1mm particle size stream is directed to a high shear mixer where the oil-wetted coal particles are conditioned. The slurry is then transfered to flotation cells, where the coal microagglomerates, in the form of froth, are separated from clean soil. To facilitate dewatering and improve handleability of the combustible product, the froth can be subsequently fed into the low shear mixer for further agglomeration.

Flotation has progressed and developed over the years; recent trends to achieve better liberation by fine grinding have intensified the search for more advanced means of improving selectivity. This involves not only more selective flotation agents but also better flotation equipment. Since the froth product in conventional flotation machines contains entrained fine gangue, which is carried into the froth with feed water, the use of froth spraying was suggested in the late 1950s to eliminate this type of froth contamination. The flotation column patented in Canada in the early 1960s and marketed by the Column Flotation Company of Canada, Ltd., combines these ideas in the form of wash water supplied to the froth. The countercurrent wash water introduced at the top of a long column prevents the feed water and the slimes that it carries from entering an upper layer of the froth, thus enhancing selectivity.

The microbubble flotation column (Microcel) developed at Virginia Tech is based on the basic premise that the rate (k) at which fine particles collide with bubbles increases as the inverse cube of the bubble size (Db), i.e., k1/Db3. In the Microcel, small bubbles in the range of 100500m are generated by pumping a slurry through an in-line mixer while introducing air into the slurry at the front end of the mixer. The microbubbles generated as such are injected into the bottom of the column slightly above the section from which the slurry is with drawn for bubble generation. The microbubbles rise along the height of the column, pick up the coal particles along the way, and form a layer of froth at the top section of the column. Like most other columns, it utilizes wash water added to the froth phase to remove the entrained ash-forming minerals. Advantages of the Microcel are that the bubble generators are external to the column, allowing for easy maintenance, and that the bubble generators are nonplugging. An 8-ft diameter column uses four 4-in. in-line mixers to produce 56 tons of clean coal from a cyclone overflow containing 50% finer than 500 mesh.

Another interesting and quite different column was developed at Michigan Tech. It is referred to as a static tube flotation machine, and it incorporates a packed-bed column filled with a stack of corrugated plates. The packing elements arranged in blocks positioned at right angles to each other break bubbles into small sizes and obviate the need for a sparger. Wash water descends through the same flow passages as air (but countercurrently) and removes entrained particles from the froth product. It was shown in both the laboratory and the process demonstration unit that this device handles extremely well fine below 500-mesh material.

Another novel concept is the Air-Sparged Hydrocyclone developed at the University of Utah. In this device, the slurry fed tangentially through the cyclone header into the porous cylinder to develop a swirl flow pattern intersects with air sparged through the jacketed porous cylinder. The froth product is discharged through the overflow stream.

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.

flotation machines

As pneumatic and froth separation devices are not commonly used in industry today, no further discussion about them will be given in this module. The mechanical machine is dearly the most common type of flotation machine currently used in industry, followed by the column machine which has recently experienced a rapid growth.

A mechanical machine consists of a mechanically driven impeller that disperses air into the agitated pulp. In normal practice this machine appears as a long tank-like vessel having a number of impellers in series. Mechanical machines can have open flow of pulp between the impellers or can be of cell-to-cell design with weirs between them. Below is a typical bank of flotation cells used in industrial practice.

The procedure by which air is introduced into a mechanical machine falls into two broad categories: self-aerating, where the machine uses the depression created by the impeller to induce air, and supercharged, where air is generated from an external blower. The incoming feed to the mechanical flotation machine is usually introduced in the lower portion of the machine. At the very below is shown a typical flotation cell of each air delivery type (Agitair & Denver)

The most rapidly growing class of flotation machine is the column machine, which is, as its name implies, a vessel having a large height-to-diameter ratio (from 5 to 20) in contrast tomechanical cells. This type of machine provides a counter-current flow of air bubbles and slurry with a long contact time and plenty of wash water. As might be expected, the major advantage of such a machine is the high separation grade that can be achieved, so that column cells are often used as a final concentrate cleaning step. Special care has to be exercised in the generation of fine air bubbles and the control of the feed rate to the column cell for such cells to be effective. Column cell use is often of limited value in the recovery of relatively coarse valuable particles; because of the long lifting distances involved, the bubbles can not carry large particles all the way to the top of the cell.

Probably the most significant area of change in mechanical flotation cell design has been the dramatic increase in machine cell volume with a single impeller. The idea behind this approach is that as machine size increases (assuming no loss of recovery performance with the larger machines), both plant capital and operating cost per unit of throughput decrease. In certain industrial applications today, cells of even a thousand cubic meters in volume (a large swimming pool) are being used effectively.

The throughput capabilities of various cell designs will vary with the flotation machines residence time and pulp density The number of cells required for a given operation is determined from standard engineering, mass balance calculations. In the design of a new plant, the characterization of each cells volume and flotation efficiency is generally calculated from data gathered on a laboratory scale flotation using the same type of equipment for the same material mixture in question. This procedure is then followed by the application of semi-empirically derived scale-up factors. Research work is currently under way to improve the understanding and performance of commercial flotation cells.

Currently, flotation cell design is primarily a proprietary material of the various cell manufacturers. Flotation plants are built in multiple cell configurations (called banks), and the flow through the various banks is adjusted in order to optimize plant recovery of the valuable as well as the grade of the total recovered mass from flotation. Up above is a typical flotation bank scheme. The total layout of a given flotation plant (including all of the various banks) operating on a given feed is called a flotation circuit.

The application of the air-lift to flotation is not new, but the first attempts to make use of the principle were not successful because the degree of agitation in the machine was insufficient to enable the heavy oils then in use as collecting reagents to function effectively. The advent of chemical promoters, however, made agitation of secondary and aeration of primary importance, with the result that the application of the air-lift principle became practicable and led to the introduction of the Forrester and the Hunt matless machines. South western Engineering Corporation are the owners in most countries of the rights to license and manufacture these and other types operating on the air-lift principle, and they have developed a machine based chiefly on the Welsh and Hunt patents which may be considered as representative of the type that is now most commonly used.

The Southwestern Air-Lift Machine, as it is called, consists of a V-shaped wood or steel trough of any length but of the standard cross-section shown in Fig. 40, the area of which is 9.85 sq. ft. and the interior depth 36 in. Low- pressure air is delivered from a blower through a main supply pipe to an air-pipe or header which runs longitudinally over the top of the machine. The air enters the trough itself through a seriesof vertical down-pipes , which are screwed into sockets welded tothe underside of the header at 4-in. intervals along its length and are open at their lower ends. They are from to 1 in. in diameter for roughing machines and from to in. for cleaners, and they reach to within 6 in. of the bottom. The air-lift chamber is formed by two vertical partitions, one on each side of the line of down-pipes, both of which extend from one end of the trough to the other, forming a compartment 6 in. wide. The lower edges of the partitions are an inch or two above the ends of the down-pipes and their upper edges are about level with the froth overflow lips at each side of the machine. A few inches above the top of the air-lift chamber is a deflector cap which serves to direct the rising pulp outwards and downwards against two vertical baffles. These extend the length of the trough parallel to and outside the partitions, their loweredges being several inches below the normal pulp level. The spacebetween the baffles and the sides of the machine forms two spitzkasten- shaped zones of quiet settlement where the froth collects.

The feed enters near the bottom of one end of machine and the tailing is discharged over an adjustable weir at the other end. The air, issuing in a continuous stream from the open ends of the down-pipes, carries the pulp up the central chamber on the principle of an air-lift pump. The air is subdivided into minute bubbles and more completely mixed with the pulp as the rising mass hits the cap at the top and is deflected and cascaded on to the baffles at each side, which direct it downwards, distributing the bubbles evenly throughout the pulp in the body of the machine and giving them ample opportunity to collect a coating of mineral. Rising under their own buoyancy, the bubbles enter the spitzkasten zones, up which they travel without interference, dropping most of the gangue particles mechanically entangled between them as they ascend. They collect on the surface of the pulp at the top as a mineralized froth, which is voluminous enough to pass over the lip into the concentrate launders without the need of scrapers. The pulp, on the other hand, continues its downward passage and enters the air-lift chamber again. In this way a continuous circulation of the pulp is maintained, its course through the machine being more or less in the form of a double spiral.

The aeration is generally controlled by a single valve in the header of each machine, but for selective flotation the machine is sometimes divided by transverse partitions into sections 4 ft. long, the header over each section being provided with a separate air-valve. The depth of the froth is regulated by means of the adjustable gate of the tailing weir. If difficulty is likely to be experienced in making a clean tailing with the normal amount of aeration, it is preferable to use two machines. The second one is run as a scavenger with an excess of air as compared with normal requirements, the low-grade froth so produced being pumped back to the head of the primary or roughing machine, in which the aeration is more normal in order that a comparatively clean concentrate may be produced. It is often possible to take a concentrate off the first few feet of the rougher rich enough to be sent to the filters as a finished product, the froth from the rest of the machine being pumped back to the head. When this method of flotation is adopted, it is an advantage to have the header divided into sections, each with its own valve, so that the aeration can be varied along the length of the machine. By increasing the volume of air at the discharge end the froth can be given a slight flow towards the head of the machine, with the result that the minerals are concentrated there to the exclusion of partially floatable gangue which might otherwise enter any bubbles not fully loaded with mineral.

If the froth from the feed end of the rougher is not of high enough grade, it must be re-treated in a separate cleaning machine, the length of which usually varies from one-quarter to one-half of the total length ofthe roughing and scavenging machines according to the amount of concentrate to be handled. Should still further cleaning be necessary, it is performed in a recleaner, which is generally of the same length as the cleaner. The tailings from these operations are often, but not necessarily, returned to the head of the rougher.

It is usual to prepare the pulp for flotation by adding the reagents to the grinding circuit or in a conditioning tank ahead of the flotation section, but soluble frothers such as pine oil and quick-acting promoters such as the xanthates can be added at the head of the machine if desired, since the air-lift provides enough agitation to emulsify and distribute them throughout the pulp. It is not as a rule advisable to introduce reagents into the air-lift chamber itself ; should it be necessary to do so to obtain a satisfactory recovery of the minerals, it is best to employ separate roughing and scavenging machines and to make the extra additions at the head of the scavenger.

Southwestern Air-Lift Machines are made of standard cross-section, as already stated, and in a series of lengths ranging, for ordinary purposes, from 4 to 48 ft. There is no limit to the possible length, however, and 100-ft. machines are in actual use. The tonnage capacities under different conditions will be found in Table 26. The pressure of air needed at the machine is from 1.6 to 1.7 lb. per square inch, which under normal conditions requires a pressure of about 2 lb. per square inch at the blower. It is usual to allow 75 to 100 cu. ft. of free air per minute at this pressure per foot of rougher and 45 to 70 cu. ft. per minute per foot of cleaner and recleaner. From these figures the approximate volume of air required for a machine or machines of any given length can be calculated. The power necessary to supply the air can then be found from Table 30.

The Callow Cell consists of a shallow horizontal trough, the bottom of which is covered with a porous medium, usually termed a blanket, consisting of a few layers of canvas or of a sheet of perforated rubber. Air is introduced at low pressure under the blanket, and, in passing through it, is split up into minute bubbles, which rise through the pulp in the cell, collecting a coating of mineral in the process.

Fig. 41 shows a section of the type of cell commonly employed. Its width is usually from 24 to 36 in., and its interior depth from 18 to 22 in. measured from the overflow lip ; the length varies according to requirements and is generally a multiple of the width. On the bottom are placed, side by side, the square open-topped cast-iron blanket frames or pans . The blanket covering the top of each pan is securely held in place by flat iron strips bolted round the edges, while one or two pipe grid-bars across the top prevent it from bulging. This arrangement allows a blanket to be changed in a few minutes should it becomedamaged. The air inlet to each pan projects through the bottom of the cell and is connected by a pipe and regulating valve to a header, which is provided with a main control valve.

The pulp enters one end of the cell through a feed opening and is discharged over an adjustable weir at the other end. There is no agitation, but the continuously rising stream of air bubbles keeps the particles of ore in suspension and induces a certain amount of circulation as the pulp passes along the cell. In this way the minerals are given many chances of becoming attached to the bubbles and thus of being carried over into the concentrate launder. The froth that forms on the surfaceof the pulp, usually to a depth of 8 to 10 inches, is voluminous enough to overflow the lips on each side of the cell without the use of mechanical scrapers.

For estimating purposes the average capacity of a Callow Cell may be taken as 2.5 tons of feed per square foot of blanket area per 24 hours and the air consumption as 9 cu. ft. of free air per minute per square foot of blanket at a pressure of 4 lb. per sq. in. A greater pressure is likely to be required if the blankets become blinded .

The Callow Cell has proved satisfactory for many types of ores, but it has the disadvantage that coarse or heavy sand settles on the blankets, and can only be kept in motion by flogging the latter with short rubber-buffered poles. Moreover, if lime is employed in the circuit, the blankets become impregnated and clogged with calcium carbonate, which necessitates periodical acid treatment for its removal. The use of perforated rubber sheets in place of canvas in the Callow Cell mitigates without entirely curing these difficulties, which at one time were thought to be inherent in the use of a porous medium. They have been overcome, however, by the development of the Callow-Maclntosh Machine.

The Callow-Maclntosh, or the Macintosh Machine, consists of a shallow trough or cell at the bottom of which is a hollow revolving rotor covered with a porous medium. Fig. 42 shows its construction. The pulp enters through a feed opening at one end, and is discharged at the other in much the same way as in a Callow Cell. The rotor, made of seamless steel tubing with a cast-steel ring welded in each end, is perforated with -in. holes at 7-in. centres; it is about 8 in. shorter than the length of the cell and is usually 9 in. in diameter. Its weight is taken by two hollow shafts, each fitted with a flange, which are bolted to the ends of the rotor by means of four studs. This method of attachment enables the rotor to be changed and a new one inserted with little loss of time, usually not more than 15 minutes. The shafts project through the ends of this cell and are supported on self-aligning ball and socket bearings outside, so placed that the rotor itself is a few inches clear of the bottom of the trough. A rubber gasket, shown in Fig. 43, seals the opening at each end by simple pressure on a cone-faced disc mounted on the shaft. The joint is not completely watertight and a slight leakage takes place through it at the rate of about one quart per minute. At the discharge end this escaping pulp gravitates to the tailing launder, while at the feed end it is usually led to one of the pumps returning a middling product to the roughing circuit. The gasket is preferable to a stuffing-box, as it contains no grease and requires no gland water.

The rotor covering consists of a canvas sock or of a single sheet of perforated rubber. The latter is now far more commonly employed, since it lasts five times as long as the other, its life generally exceeding 18 months ; moreover it seldom becomes blinded withcalcium carbonate, and requires an air pressure of only 2 lb. per square inch instead of the 3-lb. pressure needed for canvas. The rubber sheets are made of pure gum about 5/64 in. thick with 225 holes per sq. in., the holes being made so as to allow the air to pass through while preventing the percolation of the pulp into therotor in the event of a temporary shut-down. Two scraper bars of angle iron, 1 by 1 in., are bolted to opposite sides of the rotor on the top of the covering. They project 2 in. beyond the ends of the rotor, and their purpose is to keep in circulation any sand that settles on the bottom of the cell, at the same timeprotecting the porous medium from undue wear by contact withsuch material. Air is introduced into the rotor through one or bothof the hollow shafts, which are connected by special inlet joints with themain supply. When both ends are employed for the admission of air,the rotor is usually divided into two sections by a central partitionto enable each half to be controlled separately. The rotor is driven ata speed of about 15 r.p.m. by an individual motor connected with theshaft at one end of the cell; either a worm drive directly coupled to themotor or a chain drive coupled to the motor through a speed reducercan be employed.

The principle on which theCallow-Maclntosh Machine worksis very similar to that of a CallowCell. The air bubbles actuallyissue from the top of the rotor,where the hydraulic pressure islowest, and spread out as theyrise, their distribution throughthe pulp being quite as even andeffective as when a flat blanket isused. The cell never needs flogging since the movementof the rotor prevents sand fromsettling on it, and the scraperbars keep in circulation theheavy particles that would otherwise settle on the bottom. Themachine can, if necessary, handle ore as coarse as 20 mesh at a W/Sratio of 1/1 without choking.

The control of a pneumatic cell is different from that of a machine of the mechanically agitated type, of which each cell is capable of performing the function of a high-speed conditioner. Little conditioning takes place once the pulp has entered a pneumatic cell, and provision must therefore be made for its proper preparation when employing heavy oils or chemical reagents which need a long contact period. The froth is usually maintained at a depth of 8 to 10 in., giving an effective pulp depth of 18 to 20 in. The very large volume of air bubbles released enables flotation to be effected more rapidly than in any other type of machine, the actual time required depending mostly on the degree to which the minerals have been rendered floatable. The upward stream of bubbles is so voluminous that, under ordinary conditions, the froth overflows the lips on both sides of the cell without the need of scrapers. For the same reason a considerable quantity of gangue is often carried over into the concentrate launder by mechanical entanglement with the bubbles, and one, sometimes two, subsequent cleaning operations are generally necessary in consequence. This, however, is by no means therule ; a concentrate of high enough grade to be sent to the filtering section as a finished product can sometimes be made in a single rougher- cleaner cell. When the Callow-Macintosh Machine is run in this way (counter-current operation) a partitioned rotor is employed, since, by increasing the volume of air at the tailing-discharge end, the froth can be made to flow towards the head of the cell with the result that the minerals are concentrated there to the exclusion of gangue particles. The same effect can be obtained in a Callow Cell by regulating the admission of air to the individual pans in a similar way. If is often the practice, especially in counter-current operation, for the rougher to be followed by a scavenging cell, which is run with an excess of air as compared with the former, the froth being returned to the head of the first cell.

Callow-Macintosh Machines are made in lengths of 10, 15, and 20 ft. and in widths of 24, 30, and 36 in. with a rotor 9 in. in diameter. The vertical distance from the centre-line of the rotor to the overflow lip is about 24 in. The design of the machine, however, lends itself to the construction of larger sizes for big scale operationsi.e., up to a 30-ft. cell 48 in. wide with one or two 9-in. rotors. The 30- and 36-in. cells are sometimes fitted with rotors up to 15 in. in diameter to meet special requirements.

The capacity of the standard machine varies considerably according to the grade and character of the ore. The average capacity of a rougher or rougher-cleaner cell is from 8 to 12 tons of dry feed per foot of rotor length per 24 hours. When cleaning is practised, the tonnage per foot of total rotor length (roughers, scavengers, and cleaners) may vary from 4 tons for a slow-floating ore needing double cleaning to 10 tons for an easily-floated ore with single cleaning, the average being about 6 tons per foot of total rotor length. The cleaning section usually amounts to between one-quarter and one-half of the combined length of the roughing and scavenging cells. The width of cell employed depends on the character of the ore, the time of treatment, and the tonnage.

The quantity of air necessary varies from 5 to 7 cu. ft. per minute per square foot of aerating surface at 2- to 2-lb. pressurethat is, from 12 to 16.5 cu. ft. per minute per linear foot of rotor. With a Roots type blower the power consumption in respect of the air supply is about 12 h.p. per 1,000 cu. ft. of free air per minute at a pressure of 2 lb. per square inch. The power needed to turn the rotor averages 0.5 h.p.

flotation feed - an overview | sciencedirect topics

In suspension, it is essential that the impeller or air jet of the machine is capable of keeping the solids in the pulp in suspension. If the degree of agitation is inadequate then solids, particularly the largest particles, will tend to settle out. Some settling out, for example in the corners of the cell, is not serious but significant sanding of the cell floor will upset pulp flow patterns within the cell and prevent proper contact between suspended particles and air bubbles. Particles not in suspension cannot make effective contact with air bubbles.

Effective aeration requires that the bubbles be finely disseminated, and that the air rate is sufficiently high, not only to provide sufficient bubbles to make contact with the particles but also to provide a stable froth of reasonable depth. Usually, the type and amount of frother will be able to influence the froth layer, but the frother and air rate can both be used as variables.

The difficulty facing the flotation designer is that the cell performance is a strong function of the size of the particles to be floated, and that flotation feeds contain a wide range of particle sizes. For any given particle size, the effects of impeller speed and bubble diameter can be summarised as follows [1]:

If the bubble size is too large, the fewer will be the number of bubbles created for a constant air flowrate. Since the overall rate of flotation depends on the number as well as on the size of the bubbles, the recovery will drop.

This sets the boundaries for the optimum conditions of impeller speed and bubble size for flotation of any feed. If the feed size range is broad, then the optimum conditions for flotation of the coarse particles may be considerably different to the optimum conditions for the flotation recovery of the fine particles.

The pressure near the centre of the rotating impeller is lower than the ambient pressure at the same point if the rotating impeller were not present. This is due to the centrifugal pressure gradients induced by the rotation. The pressure near the impeller may be so low as to be less than the hydrostatic pressure in the pulp so that a pipe placed near the impeller and open to the atmosphere may suck air into the impeller region. This is known as induced air and the practice of introducing air into the impeller region is called sub-aeration. Common practice in coal flotation is to use this induced air as the only aeration mechanism. In mineral flotation it is common to supercharge the air to provide a slight excess pressure to give a greater amount of air per unit volume of pulp.

Flotation impellers would be expected to follow a similar equation, although a slightly different constant may be found. The circulation rates are very high. For example, a 14.2 m3 cell with an impeller of diameter 0.84m, rotating at 114rpm, would have an internal circulation of 51 m3 per minute, thus circulating the cell contents between three and four times a minute. The interaction of the liquid circulating in the cell due to the impeller and the air introduced into the impeller generates the size and distribution of bubbles found in the cell.

At very low rates (QVA/D3<0.02), the air enters the core of the vortices formed behind the tips of the blades, with a strong outwards velocity component due to the pumping action. The bubble size and number are small.

At higher rates (0.02

As the air rate continually increases, the power consumption decreases, because an increasing proportion of the space in the impeller is occupied by air. Increasing the air rate leads to a lower liquid circulation rate to the extent that the suspended particles may settle out. The general behaviour of the power ratio (the ratio of power consumed in the cell to the power consumed with no air flow) versus the air-flow number is shown in Figure18.5.

The onset of flooding coincides with a sudden drop in the power consumption, and is influenced somewhat by impeller design. For best operation a cell should operate well below the flooding gas velocity. Flooding results in very large bubbles, which are of little value for flotation. For example, it is found that a reduction in air flow to an induced air flotation cell by closing off part of the air intake can substantially improve the recovery.

This chapter is concerned with the processing of the coarse and small (also called intermediate) feed size fractions, larger than about 1.0mm, up to typically 50mm or even as large as 300mm. Standard practice in Australia is to break the feed to pass 50mm before using DMCs. The lower particle size set for the DMC is often about 1.0mm. However, a finer size might be used, e.g. 0.5mm, if flotation is the only other separation method employed.

There appears, however, to be a growing realisation that fines classification at 0.5mm wedge wire to generate a flotation feed is in many instances not the best choice. The wedge wire permits elongated particles larger than 0.5mm through and, given screen wear, relatively large particles~1mm are sent to flotation cells. The flotation recovery of these relatively large particles is often very poor; hence vast quantities of fine coal are lost in flotation tailings.

Thus there is an argument that 0.5mm is too coarse for efficient flotation, which means that a fine stream becomes applicable, say between 0.20 and 1.0mm. With parallel circuits, the goal is to run them all at constant incremental ash in order to maximise overall plant yield (Luttrell et al., 2003). There is also new interest in running the classifying screen above 1.0mm. The argument is that less classifying screen area is required, and hence capital investment can be reduced. A further argument is that DMC performance breaks away below 4.0mm, and increasingly below 2.0mm. This effect is believed to be greater in large DMCs, but this issue is contested (see Section10.3.3). This again increases the need for an intermediate stream, now perhaps between 0.25 and 2.0mm.

Particle agglomeration by coagulation and flocculation is used for thickeners and filters to assist in dewatering. Coagulation through salts reduces the surface potential of the solids, and thus enables agglomeration through van der Waals forces. Coagulation results in micro-flocs. It is particularly important for coal tailings containing high clay content. Flocculation utilising synthetic polymeric chemicals as bridging flocculation is used for flotation coal dewatering in vacuum filters, and also introduced into screen bowl centrifuges to assist the separation within them. Sufficient floc conditioning (Bickert and Vince, 2010) by appropriate mixing energy and mixing time after adding the flocculant is important. Modern simple plants, which gravity feed flotation concentrate directly from the flotation cell launder onto filters, usually do not provide sufficient shear for floc mixing, or residence time for floc formation. This is believed to lead to over-flocculation.

Coal beneficiation requires the fractioning of the ROM coal, and these different size fractions are effectively dewatered by different equipment. Most SLS equipment is maximally effective for a particular (narrow) size fraction. While this is the case for coarse and fine coal centrifuges, the addition of coarse aids ultrafines filtration, in particular when the packing density is maximised and a homogeneous isotropic cake structure can be achieved (Anlauf, 1990).

Thickening flotation concentrate and tailings prior to filtration reduces the amount of water to be removed by filtration, and thus increases the capacity but also dampens fluctuations within the thickener, resulting in a more consistent, stable filter operation. This is particularly beneficial for throughput increase on high capacity vacuum disc filters while the capacity increase for pilot and full-scale filters is as per prediction by filtration theory (Bickert, 2006).

Vibrating screens are used to remove most of the water on coarse coal after wet beneficiation prior to centrifugation. They can be used as final dewatering devices, either for coarse reject or very coarse product such as from jigs and baths. Screens are also used extensively in other duties for sizing and desliming within coal preparation plants.

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.

The concept of processing fine coal in spirals is not new. Innovation in design and construction materials has improved the performance of spirals. High levels of processing efficiency can be realised at comparatively low capital and operating cost. Plant capacity can be increased by diverting part of HM cyclones and froth flotation feed in an existing plant to spirals. The efficiency levels can be further improved by rewashing the middlings of primary spirals in secondary units. The drawback of the spiral circuits is their inability to make a low density cut at below 1.5 sp. gr. (Bethel, 1988). Spirals are the largest amongst the fine coal-cleaning technologies due to the following advantages (Honaker et al., 2013):

Spirals are used extensively to process fine (<10.15mm) coal. A spiral consists of a corkscrew-shaped conduit (Fig. 8.13) with a modified semicircular cross-section (Luttrell and Honaker, 2013). The slurry is fed at the top of the spiral, usually from a constant head tank. The design of the device imposes a centrifugal force in addition to the flowing-film separation. The combination of these actions forces the low-density particles outward, while the high-density particles are driven inward. The coal and refuse particles are separated at the bottom of the trough by splitting the flow into clean coal, refuse and usually middling streams (Klima, 2013). Adjustable diverters (called splitters) are used to control the proportion of particles that report to the various products. Conventional spirals have 5.25 turns around the vertical shaft, whereas compound spirals have seven turns. Compound spirals combine a two-stage operation into a single unit.

Since spirals have low unit capacity (24t/h), several units (two or three) are intertwined along a single central axis to increase the capacity for a given floor space. Multiple spirals are usually combined into a bank fed by an overhead common radial distributor each having a separate feed point or start. Spirals have been successfully utilised in combination with water-only cyclones to improve the efficiency of separating fine coal (Honaker et al., 2007), such as:

Lack of uniformity in feeding results in substantial falls in operating efficiency and can lead to severe losses in recovery, this is especially true with coal spirals (Holland-Bat, 1993). This is due to the creation of differences in RD cut-off points between different spiral units.

Control of dry solids tonnage, slurry flow rate, feed solids content, distributer level, oversized particles, sanding/beaching and good operating practices with effective maintenance programmes (Luttrell, 2014).

As in coal flotation, oil agglomeration takes advantage of the difference between the surface properties of low-ash coal and high-ash gangue particles, and can cope with even finer particles than flotation. In this process, coal particles are agglomerated under conditions of intense agitation. The following separation of the agglomerates from the suspension of the hydrophilic gangue is carried out by screening.

The amount of oil that is required is in the range of 510% by weight of solids. Published data indicate that the importance of agitation time increases as oil density and viscosity increase, and that the conditioning time required to form satisfactory coal agglomerates decreases as the agitation is intensified. Because the agitation initially serves to disperse the bridging oil to contact the oil droplets and coal particles, higher shear mixing with a lower viscosity bridging liquid is desirable in the first stage (microagglomeration), and less intense agitation with the addition of higher viscosity oil (macroagglomeration) is desirable in the second stage. Viscous oil may produce larger agglomerates that retain less moisture. With larger oil additions (20% by weight of solids), the moisture content of the agglomerated product can be well below 20% and may be reduced even further if tumbling is used in the second stage instead of agitation.

The National Research Council of Canada developed the spherical agglomeration process in the 1960s. This process takes place in two stages: First, the coal slurry is agitated with light oil in high shear blenders where microagglomerates are formed; then the microagglomerates are subjected to dewatering on screen and additional pelletizing with heavy oil.

Shell developed a novel mixing device to condition oil with suspension. Application of the Shell Pelletizing Separator to coal cleaning yielded very hard, uniform in size, and simple to dewater pellets at high coal recoveries. The German Oilfloc Process was developed to treat the high-clay, 400-mesh fraction of coal, which is the product of flotation feed desliming. In the process developed by the Central Fuel Research Institute of India, coal slurry is treated with diesel oil (2% additions) in mills and then agglomerated with 812% additions of heavy oil.

It is known that low rank and/or oxidized coals are not a suitable feedstock for beneficiation by the oil agglomeration method. The research carried out at the Alberta Research Council has shown, however, that bridging liquids, comprising mainly bitumen and heavy refinery residues are very efficient in agglomeration of thermal bituminous coals. Similar results had earlier been reported in the flotation of low rank coals; the process was much improved when 20% of no. 6 heavy oil was added to 2 fuel oil.

This is another oil agglomeration process that can cope with extremely fine particles. In this process, fine raw coal, crushed below 10cm, is comminuted in hammer crushers to below 250m and mixed with water to make a 50% by weight suspension; this is further ground below 15m and then diluted with water to 15% solids by weight. Such a feed is agglomerated with the use of Freon-113, and the coal agglomerates and dispersed mineral matter are separated over screen. The separated coal-agglomerated product retains 1040% water and is subjected to thermal drying; Freon-113, with its boiling point at 47C, evaporates, and after condensing is returned liquified to the circuit. The product coal may retain 50ppm of Freon and 3040% water.

Various coals cleaned in the Otisca T-Process contained in most cases below 1% ash, with the carbonaceous material recovery claimed to be almost 100%. Such a low ash content in the product indicates that very fine grinding liberates even micromineral matter (the third level of heterogeneity); it also shows Freon-113 to be an exceptionally selective agglomerant.

In some countries, for example in Western Canada, the major obstacles to the development of a coal mining industry are transportation and the beneficiation/utilization of fines. Selective agglomeration during pipelining offers an interesting solution in such cases. Since, according to some assessments, pipelining is the least expensive means for coal transportation over long distances, this ingenious invention combines cheap transportation with very efficient beneficiation and dewatering. The Alberta Research Council experiments showed that selective agglomeration of coal can be accomplished in a pipeline operated under certain conditions. Compared with conventional oil agglomeration in stirred tanks, the long-distance pipeline agglomeration yields a superior product in terms of water and oil content as well as the mechanical properties of the agglomerates. The agglomerated coal can be separated over a 0.7-mm screen from the slurry. The water content in agglomerates was found to be 28% for metallurgical coals, 615% for thermal coals (high-volatile bituminous Alberta), and 723% for subbituminous coal. The ash content of the raw metallurgical coal was 18.939.8%, and the ash content of agglomerates was 815.4%. For thermal coals the agglomeration reduced the ash content from 19.848.0 to 512.8%, which, of course, is accompanied by a drastic increase in coal calorific value. Besides transportation and beneficiation, the agglomeration also facilitates material handling; the experiments showed that the agglomerates can be pipelined over distances of 10002000km.

Flotation has progressed and developed over the years; recent trends to achieve better liberation by fine grinding have intensified the search for more advanced means of improving selectivity. This involves not only more selective flotation agents but also better flotation equipment. Since the froth product in conventional flotation machines contains entrained fine gangue, which is carried into the froth with feed water, the use of froth spraying was suggested in the late 1950s to eliminate this type of froth contamination. The flotation column patented in Canada in the early 1960s and marketed by the Column Flotation Company of Canada, Ltd., combines these ideas in the form of wash water supplied to the froth. The countercurrent wash water introduced at the top of a long column prevents the feed water and the slimes that it carries from entering an upper layer of the froth, thus enhancing selectivity.

The microbubble flotation column (Microcel) developed at Virginia Tech is based on the basic premise that the rate (k) at which fine particles collide with bubbles increases as the inverse cube of the bubble size (Db), i.e., k1/Db3. In the Microcel, small bubbles in the range of 100500m are generated by pumping a slurry through an in-line mixer while introducing air into the slurry at the front end of the mixer. The microbubbles generated as such are injected into the bottom of the column slightly above the section from which the slurry is with drawn for bubble generation. The microbubbles rise along the height of the column, pick up the coal particles along the way, and form a layer of froth at the top section of the column. Like most other columns, it utilizes wash water added to the froth phase to remove the entrained ash-forming minerals. Advantages of the Microcel are that the bubble generators are external to the column, allowing for easy maintenance, and that the bubble generators are nonplugging. An 8-ft diameter column uses four 4-in. in-line mixers to produce 56 tons of clean coal from a cyclone overflow containing 50% finer than 500 mesh.

Another interesting and quite different column was developed at Michigan Tech. It is referred to as a static tube flotation machine, and it incorporates a packed-bed column filled with a stack of corrugated plates. The packing elements arranged in blocks positioned at right angles to each other break bubbles into small sizes and obviate the need for a sparger. Wash water descends through the same flow passages as air (but countercurrently) and removes entrained particles from the froth product. It was shown in both the laboratory and the process demonstration unit that this device handles extremely well fine below 500-mesh material.

Another novel concept is the Air-Sparged Hydrocyclone developed at the University of Utah. In this device, the slurry fed tangentially through the cyclone header into the porous cylinder to develop a swirl flow pattern intersects with air sparged through the jacketed porous cylinder. The froth product is discharged through the overflow stream.

Coal flotation is a separation process performed mainly based on differences in surface hydrophobicity between coal and gangue. The flotation reagent can improve the hydrophobicity of coal, and also the adhesion between coal and air bubbles. Kerosene and light diesel oil are widely used as collectors in coal flotation. However, collectors are not completely dispersing in water due to their chemical stability, hydrophobicity and symmetric structure. Besides, flotation feeds contain massive fine and even ultra-fine clay, resulting in poor selectivity of the reagent toward coal particles and even requirement of higher dosages of reagent. However, the ultrasound can tremendously improve the properties of flotation reagent [5357]. So emulsified collectors by ultrasound are beneficial to improve the adsorption speed of collector over coal surface and selectivity of the collector toward coal particles in coal flotation.

Emulsification, i.e. intimate mixing of two immiscible liquids was one of the first applications of ultrasound, which has begun as early as in the 1927 [58,59]. Ultrasound is a very efficient emulsification technology than others such as mechanical agitation. [22]. As a result of cavitation, the excess energy for creating the new interface decreases the interfacial tension, which breaks the large oil droplets into small droplets [6062]. It is shown that emulsions produced by ultrasound are stable [63] with smaller droplets [64,65], and consumes lower energy [66] than mechanical action for producing emulsions [67,68]. It is beneficial to improve adhesion between reagent and mineral particles as well as flotation efficiency [56,60,69].

The properties of emulsified reagents by ultrasound are affected by many factors, such as surfactants, temperature, pressure, ultrasonic parameters, emulsifier concentration, and viscosity [22]. Bondy and Sllner [70] qualitatively analyzed the ultrasonic emulsification process. They found an optimum in emulsification efficiency occurred at an absolute pressure of about two atmospheres. Li and Fogler [71,72] considered the ultrasonic time as the key parameter in ultrasound emulsification since a very short time (a few seconds) only produce coarse emulsions (e.g. 70m droplets), whereas longer time can produce submicron emulsions. In addition, they also proposed a two-step mechanism of acoustic emulsification, as shown in Fig. 8.

The first step involves a combination of interfacial waves and Rayleigh-Taylor instability, leading to the conversion of dispersed phase droplets into the continuous phase. The second step is the breakup of droplets through cavitation near the interface. Therefore, it can be considered that the intense effects such as disruption and mixing of shock waves are the key for producing very small droplet size.

Generally, the less viscous liquid (e.g. water) undergoes cavitation more easily [73] and becomes the emulsion continuous phase (oil-in-water, O/W, or direct emulsion). The cavitation threshold decreases with liquid viscosity. The cavitation threshold is lower in the less viscous liquid, which can better disperse the oil droplets in the water and form the emulsion continuous phase. A reverse emulsion (water-in-oil, W/O type) can be obtained by changing type of emulsifier, oil-water ratio as well as ultrasonic field conditions [22]. Currently, the W/O preparation by ultrasound has been rarely reported. This is because the W/O emulsions are unstable and their properties such as droplet size are difficult to be directly analyzed compared with the O/W emulsions [74]. In addition, Wood and Loomis [58] proposed that ultrasonic emulsification was the unique phenomenon that one liquid incompletely permeated into other liquid to further form small droplets. The kinetics of ultrasonic emulsification were investigated by Rajagopal based on experimental data and theory analysis [75]. A working model to determine the rate of ultrasonic emulsification was proposed, considering the dispersion at the interface and the coagulations of the emulsion. The results are important to understand the individual contributions of dispersion and coagulation to emulsion formation.

During the process of ultrasonic cavitation, cavitation bubbles rapidly collapse in a short time by high-frequency oscillation, which produce shock waves and microjets in liquid accompanying with local high pressure (>100MPa) and high temperature (5000K) [41,76]. Therefore, collector, such as kerosene and light diesel oil, can be dispersed to form smaller oil droplets with uniform distribution in suspension. Kang et al. [54] investigated flotation performance of bituminous coal using ultrasonic emulsified kerosene. The results showed that the droplet size after ultrasonic emulsification of kerosene gradually increased with a decrease in the ratio of oil-water. However, the relationship of wetting heat of emulsified kerosene and ratio of oil-water showed a positive correlation. The decrease in the ratio of oil-water can weaken the stability of the droplet after ultrasonic emulsification. In addition, the average contact angle of coal slime was improved using emulsified kerosene, leading to the improvement of efficiency and selectivity of coal flotation. Ruan et al. [53] reported that the stability of emulsified diesel was impacted by some factors such as the dosage of emulsifier, ratio of oil-water, ultrasonic time and the dosage of butyl alcohol as shown in Table 1. Table 1 shows the influence of different factors on emulsion stability: dosage of emulsifier>ratio of oil-water>ultrasonic time>dosage of butyl alcohol. Under the same reagent consumption, emulsified diesel can obtain a lower concentrate ash content compared with unemulsified kerosene whereas the concentrate yield had no change.

Sahinolu and Uslu [77] also found that size of the oil droplets decreased with an increase of ultrasonic power and treatment time. At oil agglomeration tests of oxidized coal fines, ash and pyritic sulfur rejections without ultrasonic emulsification were 50.38% and 85.28%, respectively and were increased to maximally 56.89% and 88.69% respectively by using ultrasonic emulsification before agglomeration, as shown in Fig. 9. Increasing ultrasonic power didn't affect ash and pyritic sulfur rejections considerably whereas increasing ultrasonic treatment time at higher power levels had positively affected for them. In addition, Fig. 9 also shows that both ultrasonic power and treatment time affected the combustible recovery adversely, which may be caused by small size of oil droplets limiting growth of agglomerates. Coal particles of 0.5mm size fraction may be too coarse and removed from the agglomerate structure due to effect of gravitational force, resulting in destruction of the stability of the agglomeration. In order to eliminate this adverse effect of ultrasonic emulsification on combustible recovery, they proposed a method for reducing the particle size of coal particles.

Letmathe et al. [78] found that the application of ultrasound during emulsified reagents could achieve an improvement in separation efficiency. Therefore, the purity and ash content of the graphite at a constant solids recovery were improved and decreased, respectively. Sun et al. [79] found that with the aid of emulsifiers, intense high-frequency sound waves were effective in emulsifying any collector in water. The ultrasonic emulsified collectors were more effective in the flotation of bituminous coal than the non-emulsified collectors, particularly for the insoluble and slightly soluble ones.

In addition, the dispersive effects also lead to the formation of an emulsion when ultrasound is applied to a pulp containing stabilizers such as surfactant. The use of ultrasound in this way can improve the efficiency of the reagents and decrease the consumption of reagent [80]. This is resulted from the reagents more uniform distribution in the suspension after ultrasonic treatment and also in enhancement of the activity of the chemicals [81]. Dyatlov [82] reported that ultrasonic conditioning of the reagents promoted the formation of fine dispersed emulsions in the coal flotation. The yield of concentrate and ash content of the tailings were both improved using these ultrasonic emulsified reagents. Though the ultrasonic emulsified reagents had a positive effect in improving concentrate yield, the selectivity of coal particles in flotation seemed to be decreased. Oyama and Tanaka [83] investigated that the frothers, as a flotation reagent (50g/ton), were emulsified in water using ultrasound with a power of 4W/cm2. Even if the reagents were hardly mixing with water, the emulsions produced by ultrasound can be easily intermixed into the pulp. The recovery of galena was improved from 58% to 93% and the recovery of chalco-pyrite was increased from 73.4% to 88.9% using emulsified reagents when duration of flotation was 5min. Using ultrasonic emulsified reagents can significantly improve the selectivity of minerals flotation. In addition, lowest economical consumption of the reagents was achieved using ultrasonic emulsions.

The ultrasonic emulsified reagents may be not stable because of the high energy input in a range of small volume pulp, near the emitting surface of the ultrasonic probe or transducers. The interaction forces involved in physical adsorption of a reagent molecule are weaker than forces involved in chemical adsorption. The physical bond between reagent molecule and mineral surface can be easily broken by hydrodynamic. Hence, the stabilizer or surface-active reagent is added to prevent mergers of collector droplets and improve stability of ultrasonic emulsified reagents. The amount of surfactant required to give a stable ultrasonic emulsion is generally lower than other techniques such as mechanical agitation. Besides, in actual ultrasonic emulsified process, breaking a planar interface requires a large amount of ultrasonic energy, hence it may be more advisable to first prepare a coarse emulsion (e.g. by gentle mechanical stirring) before applying acoustic power. It is also possible to add the second liquid (dispersed phase) to the first liquid (continuous phase) progressively or feed into a continuous reactor with both phases. This way, ultrasonic treatment can make the reagent more homogeneous in suspension, and further improve the activity and stability of the emulsified reagent by adding chemical reagent. The ultrasonic emulsification can also bring substantial cost savings [8487].

professional flotation column micro-bubble counter-flow jet | y&x

The flotation column is an aerated flotation machine that inflates and agitates the ore slurry through compressed air through a porous medium (aerator). There are micro-bubble counter-flow flotation column and micro-bubble jet flotation column.

Compressed air is filled into the column through the vertical air tube through the plenum at the lower end of the column. A large number of small bubbles formed are evenly distributed on the whole section.

Compressed air is filled into the column through the vertical air tube through the plenum at the lower end of the column. A large number of small bubbles formed are evenly distributed on the whole section.

In the convection movement, due to the action of the agent, the selected minerals will adhere to the surface of the rising bubbles, forming a layer of mineralized foam on the upper part of the column, which will be scraped into or sprinkled by the scraper into the concentrate tank.

Stable flotation dynamics, relatively small bubbles, more uniform distribution, sufficient bubble-particle flotation interface, large enrichment ratio, High recovery rate, suitable for the selection of fine-grained minerals and easy to realize automatic control and large-scale;

processes | free full-text | flotation in water and wastewater treatment | html

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