industrial plant flotation machines in banks

industrial plant flotation machines in banks

A centrally located paddle-wheel type impeller generates a rotating pulp vortex which extends between two stationary elements: the sand-pipe located at the top of the cell, and the draft tube located at the bottom of the cell. The hydraulic action of this rotating vortex is to develop an internal cavity vacuum while simultaneously circulating the pulp from the bottom of the cell through the. draft tube into the rotor region. The suction developed in the vortex core draws air into the central region of the rotor which is mixed with the pulp circulated from the bottom to the cell.

Mineral flotation starts with grinding of the ore, with water and reagent, down to a chosen average grain size to secure liberation of the discrete mineral particles. This step raises a difficult mineralogical question; as perfect liberation is rarely achievable without excessive over-grinding, what defines the economic compromise?

The pulp (i.e. mineral-water slurry after grinding) is commonly passed through a hydrocylone to concentrate the sands and reject the slimes (the finest size fractions, say 30um), the slimes are objectionable in several respects: they are slow to float, the consume a disproportionate share of the reagents, and they may seriously spoil the selectivity of flotation of the sands by slime-coating them. There are sound arguments, in any case, for processing coarse and fine fractions separately.

The pulp is conditioned for a few minutes with reagents designed to accentuate differences of floatability of the various mineral species. An appreciable time is required to achieve good distribution of the reagents and to allow give-and-take competition between different mineral particles.

The conditioned pulp is run into a flotation cell, which is crudely a box with a stirrer, and a means of introducing air. Typically, the cell might contain 15-30% by volume of entrained air with bubbles ranging mainly from 0.1 to 5mm in diameter; the pulp density might be 25-40% of solids having a particle size ranging from 10-100um. The flotation grains are caught by bubbles and carried to the top of the cell, forming a froth, which is automatically scraped off over the lip of the cell, where it collapses and flows away in a launder. The frothing action is quite important. A moderate depth of froth is necessary to allow some back-drainage to take place, with release of non-floated particles which have been, unavoidably, entrained to some extent between the bubbles. Here is another reason why slimes are a nuisance they remain too long in the water between the bubbles and so reduce the grade of the floated product.

The first stage of flotation of the pulp amounts to a quite short average residence time in the rougher cell before it passes out, largely depleted, from the bottom of the cell. In a conventional flotation plant no attempt is made to engineer a perfect separation in one stage. Instead, both fractions leaving the rougher are re-treated at least once in cleaner and scavenger cells, respectively. Scavenger cells, in effect, prolong the flotation time, while competition for bubble surface is reduced. The net of the recycling and re-treatraent is improvements in the separation (grade) and proportion of valuable minerals obtained (recovery). As the latter is usually a minor component of the ore, it is preferable to float it, in preference to floating the much greater proportion of gangue; but in some cases the gangue is floated (reverse flotation).

It is a characteristic feature of flotation plants that the cells are comparatively small, but rows and rows of them are run in parallel to increase through-put and in series to improve grade. If more than one mineral is to be extracted, the pulp is re-conditioned with other reagents and further stages of flotation are operated. The engineering is simple, continuous, amenable to adjustment, and needs little operator attention. The concentrates from flotation are generally filtered, washed, dried and bagged for transport as powders.

PROBLEM 2 How many No. 18 Sp. (3232) Denver Sub-A Cells are required to treat 125 tons of load-zinc ore per day, with treatment time 14 minutes for the lead, dilution 3 to 1% and with treatment time 16 minutes for the zinc dilution 3 to 1. and sp. gr. 3.4 ?

Scale-up denotes the procedures for designing larger units of equipment for which smaller units are available with known operating characteristics. Such procedures are well established for chemical engineering equipment, particularly mixers, of which flotation cells are examples, although complex ones.

The bases for extrapolating design are the principles of similitude, including both geometrical and dynamical aspects. If full geometrical similarity is maintained in an equipment family, then once a single new size aspect is selected, for example, volume or impeller diameter, then all other dimensions are automatically established through the scale factor.

The dynamic aspect refers to the fluid motions involved and for mixers and flotation cells more specifically to the impeller speeds and energy inputs. It is here that complications arise. For relatively simple mixing problems, as in conditioners, where it has been established that power intensity is a scale-up factor, it can be shown that for contant Power Number and constant power intensity the relationship

However, in a flotation cell, the presence of air and the further necessity for balancing air and pulp flows to satisfy both flotation and suspension requirements simultaneously, introduces at least one other variable: air flow Qa; and at least one other dimensionless group the Air Flow Number Qa/ND. This may also be expressed as Qa/A/ ND

A further complication is that most families of cells have grown in size without rigorous scale-up procedures, (see Harris) Wemco does appear to have scaled its impellers closely to (cell area), i.e., D/L = constant, but this is not the case for Denver or Galigher. Cell depths now increase regularly for both Denver and Wemco, but not for Galigher.

Thus, at best, no automatic procedure for scale-up appears possible. Those published by Denver and Wemco illustrate the necessary mixture of rational and experimental steps. Wemco decided upon a 28 m (1,000 cu ft) cell and apparently also upon a .76 m (30 in) impeller diameter, which continues the D/(Area) ratio of 0. 227. Extrapolation of empirical relationships among air and liquid circulations, submergences, and HP, N, and D established ranges, but actual testing with a prototype was necessary to arrive at an impeller speed and submergence combination and to establish clearances. Finally capability of the selected combination for sand suspension was established experimentally.

For its largest cell, Denver first decided to use a 56 KW (75 HP) motor, as the largest compatible with a V-belt drive; with safety factors for start-up and overload this gave a 48 KW (65 HP) design criterion. Relationships were established from measured smaller cell operating characteristics among pumping rates, power draws, depths and cell volumes. These, with HP fixed and depth selected, resulted in fixing the volume at 36 m (1275 cu ft). The effects of clearances and independent variation of air flow on power draw were tested in a prototype as were the sand suspension capabilities.

Of interest was Denvers use of averaged fluid rise velocities as design criteria for proper sand suspension. These decrease from about 4 m/min (13 ft/min) in their smaller cells to 2.8 m/min (9 ft/min) in the 36m cell. Figures are similar for the Wemco cells but increase rather than decrease. These velocities are a rough measure of the maximum particle mass which can be suspended, although suspension in the sense of uniformity is a misnomer. As both of the studies show, there is segregation of coarser sizes. Only the finer sizes move uniformly throughout the cell with water flow, while coarser sizes find more restricted flow paths in appropriately higher velocity streams in which they circulate. In the limit, if a high enough velocity region does not exist, sizes requiring such a velocity settle out.

The three principal U. S. cells have had decades of intense competition world-wide. In addition their innovations in increasing cell volumes by factors of twenty in recent years have paid dividends in much lower power intensities. If at the same time unit capacities remain reasonably close to those of smaller cells, then any new designs may have difficulties in breaking into the field. Exceptions may well be in coarse particle handling for which conventional cells definitely have limitations because of their need to be versatile over the broadest possible range, with some sacrifices necessary.

The main hope for significant improvements rests on further quantification of relationships among flow, aeration, and process kinetics as related to particle size particularly. It may be that adjustment of the cell control parameters either in separate circuits for different size ranges, or in different parts of the same circuit, with the air and impeller speed adjustments as part of the control network along with reagents, will be the trend of developments for the future.

This would follow from the decreased ratio of area to volume with increasing depth, and thereby the lowered volume flow requirements to obtain the same linear velocities for air and pulp. In the case of air, this suggests that increased depth of cell favors more efficient utilization of air because of the longer path per bubble and the greater probability thereby of maximum loading per bubble. However, for the same reason, there may be a depth limit beyond which no further increase would be obtained.

In brief outline, these include: lower investment costs, lower installation costs, lower labor costs, and lower costs for controls. If lower power intensities at constant capacity per unit volume are borne out, then power costs per tonne would also be lower. In addition the reduction in the number of units necessary and the higher capacities per unit of floor space should result in lower indirect costs for buildings. There is a probable present limit to further increases in size if only because there are no plants in sight above the 100,000 tonne/day capacity level. At this tonnage, four rows of 12 cells, or fewer rows of greater length, of the largest presently available cell sizes would be adequate.

The indicated savings in flotation equipment costs only vary for 2.8, 8.5, 14.2 and 36.1m (100, 300, 500, 1275 cu ft) cells as 1.00 /0.60 / 0.45 / 0.37 and for 1.7, 14.2 and 28.3 m (60, 500, 1000 cu ft) cells as 1.00 / 0.52 / 0.39. Even if further savings in power intensities with still larger cells do not materialize, costs per unit of capacity for equipment, installation, and building should continue to decrease, but at a lower rate with continued increases in size.

Although there has been the implication that the very large cells should be considered only for the largest plant capacities, this is not necessarily correct. Attention is being given even for smaller plants to splitting circuits between roughing and scavenging; using one or a few of the larger cells at the head of a bank, or even for the entire tonnage as roughers to take out the fastest floating 50 to 75 percent of the mineral; and the completing the scavenging with the appropriate number of smaller cells to provide the time and the appropriate conditions of aeration and agitation for the slower floating coarse, middling, or altered surface minerals. In this way controls could be centralised in roughing, particularly if the scavenger concentrate could be recycled to the rougher, and different cell conditions could be adapted to the different requirements.

Impellers are the main differences among older mechanical cells; of novel types few are entirely new; and several are revivals of concepts going back to the first decades of practice. They can be divided into the following groups: a) injection types with pressurized introduction of feed through nozzles, which simultaneously aspirate air and mix it with the feed; b) froth feed types, which add new feed into or close to the froth; and c) cells with different fluid moving mechanisms.

Three cells of this type have been in commercial development over the past two decades: (1) theHeyl and Patterson CycloCell; (2) the Deister Flotaire, originally developed by Hollingsworth for Borden Phosphate in Florida; and (3) the Davcra Cell, developed by Davis for Conzinc Rio Tinto in Australia. Although all three use nozzles for air ingestion and mixing, the CycloCell uses a pump with recycle of cell contents through the pump/nozzle; the Flotaire substitutes water-main pressure for pumping by injecting fresh water through the nozzle for air aspiration; and finally, the Davcra uses a pump/nozzle to inject feed through the side of its cell without recycle back through the pump.

The practicality of the CycloCell has been established through extensive applications to coal flotation, and it has been tested on Florida phosphate. The Davcra has been used for special applications at Palabora and Bougainville and at locations in Australia, but it is understood that its further development has been discontinued. The Flotaire Cell has only recently been taken over by Deister who are attempting to develop applications in fields other than phosphate. The injection principle appears interesting both to supply and mix air and to provide the fluid energy for particle suspension. The important question is whether external pressure supply with loss of head through external piping and a nozzle is more or less efficient than a submerged impeller whose kinetic energy is dissipated entirely within the cell and presumably for useful purposes.

Three examples of this principle are known: the Flotaire Cell, already mentioned; a cell designed in the USSR employing the principle, but calling it froth separation , and a cell designed by Paul Smith of Colorado SMR Institute All use the same arguments and general approach: a) that a major flotation problem is the recovery of the coarsest sizes; b) that much of the energy required by flotation is for suspending coarse particles; and c) that, therefore, by adding feed directly to the froth the coarse particle problem is largely eliminated.

The proponents of all three cells can show impressive results mainly with coarse phosphate feeds, although the Russian cell is apparently in operation on other applications also. But two questions must be raised. Since the principle obviously has its greatest potential for coarse feeds, how does it compare with other specialized processes such as belt flotation, spirals, and the Lang launder, all developed and in use for similar applications. A second question concerns the behavior of fines. For these sizes what is an advantage for coarse particles becomes a disadvantage: below some size limit particles will have insufficient residence time in the froth to drain out by gravity and will be carried out in the froth product. This effect could be more severe with this type of cell than with convential cells, suggesting that sized feeds may be necessary to cope with it.

Apart from the external pump, and specialized impeller designs which claim better mixing or more effective aeration, the Outokumpu Cell being an example of the latter, a novel device is the Mekhanobr Vibromachine (USSR). This utilizes a flat paddle oscillating along a vertical axis at about 600 cycles/min with a few mm amplitude; air is admitted through perforations in the paddle during the up-stroke. The effect should resemble a high-frequency jigging action which should produce a fluidized bed-type suspension, possibly with less long-range circulation of pulp than in impeller type cells. Again the principle appears interesting and possibly requires less energy for flow induction, but in the absence of comparative data, no judgments are possible.

Air flow is actually a direct consequence of liquid flow for a self-aerating cell such as the Wemco. For all other mechanical cells air and liquid flows are independently variable, air through adjustment of supply, and liquid by impeller speed variation. Table contains a number of derived properties of Denver and Wemco cells covering the full ranges of cell volumes. Although availability of such data is too recent for any firm conclusions to be drawn, there are a number of trends evident.

The important aspect of these cell characteristics is not so much the actual values cited but that they are criteria of optimum cell operation. Too often, cell aeration and impeller speed are accepted as unchangeable instead of as potential control variables capable of manipulation, and in this respect have at least equal importance to reagent control. Thus, air flow is already being used to control concentrate grade at Mt. Isa. With the trend to larger and fewer cells, and coarser grinds, it may be advisable to control impeller speeds also, particularly for very coarse feeds, and to find the balance between these two controls for each operation.

As pointed out previously, the float machines were selected as 170 cubic foot Agitair machines for circuit flexibility. In our case we might have used 300 cubic foot machines or perhaps 500 cubic foot as the trend today is to go to larger and larger units which was considered, but eliminated for the flexibility factor. If we had selected 500 cubic foot units and utilized 5 units as roughers instead of 14 and 3 units in the scavenger area instead of 8 units, the cost savings in unit cost per cubic foot would be substantial with a resultant savings in building area. Due to the costs of building concentrators in these inflationary days, the larger units must be considered. 1,000 cubic foot units are now available and will soon be in operation.

Dilution water for the flotation circuit is derived from two sources. Concentrate thickener overflow water gravitates directly to the recleaner feed sump for dilution in recleaner flotation. Process water additions are made to each of the rougher, rougher scavenger, cleaner, recleaner and cleaner scavenger concentrate launders and to the regrind cyclone feed sump.

Vertical pumps are used for froth products and horizontal pumps for tailings and cyclone feeds. All seven banks of flotation machines are fitted with bubble-tube automatic pulp level controls and are Galigher 170 machines.

routing of cleaner scavenger tail direct to final tail. routing of cleaner concentrate to final concentrate . redirection of regrind cyclone overflow from cleaner feed to recleaner feed. by-pass of the regrind cyclones and hence the regrind section. routing of rougher concentrate to recleaner feed. by-pass of concentrate thickener overflow to the regrind cyclone feed sump and tailing thickener overflow sump.

Space has been provided in the flotation section layout for a talc flotation circuit, but no detailed design has been carried out. Its implementation will be dependent on the nature of the ores to be treated and is as follows:

The process flow diagram has been developed for the flotation of these talcy ores. The process depends on the rapid flotation of talc in this rougher flotation stage followed by three stages of cleaning of the rougher talc concentrate.

The talc rougher tail and three-stage cleaner tails are all directed to the No. 1 rougher scavenger bank which in this configuration becomes No. 1 copper rougher bank. The cleaned talc concentrate is discarded to final tail.

Flotation feed passes through one stage of rougher flotation, with rougher flotation tailing passing successively through three stages of rougher scavenger flotation. The unfloated product from the third rougher scavenger section is flotation tail. Rougher concentrate is refloated successively in cleaner and recleaner stages, with the recleaner concentrate leaving the flotation circuit as final concentrate. Cleaner tailing gravitates to cleaner scavenger flotation and recleaner tailing gravitates to cleaner flotation.

flotation process - an overview | sciencedirect topics

flotation process - an overview | sciencedirect topics

Flotation processes are based on the different surface wettability properties of materials (Wang etal., 2015). In principle, flotation works very similarly to a sink and float process, where the density characteristics of the materials, with respect to that of the medium where they are placed are at the base of the separation. Sometimes a centrifugal field is applied to enhance separation. Flotation works in a different way in the sense that in a liquid medium, usually water, a carrier is introduced, air bubbles, responsible to float hydrophobic particles that adhere to the bubbles with respect to the hydrophilic ones that sink. According to surface plastic characteristics, this technique can be profitably applied, in principle, to separate waste polymers (Fraunholcz, 2004). To enhance or reduce plastic surface characteristics (i.e., hydrophobic or hydrophilic) appropriate collectors, conditioners (Singh, 1998; Shen etal., 2002), and flotation cell operative conditions (i.e., air flow rate, agitation) can be utilized. Usually plastic flotation is carried out in alkaline conditions (Takoungsakdakun and Pongstabodee, 2007). Once floated, hydrophobic polymers are recovered as well as the sunk ones (i.e., hydrophilic) at the bottom of the cell. This technique, even if it is well-known (Buchan and Yarar, 1995) and in principle quite powerful is not widely used mainly for three reasons: (1) it is a wet technique, this means that water has to be recovered and processed before reutilization, due to the presence of the reagents and contaminants, (2) polymer surface status (i.e., presence of dirtiness/pollutants and/or of physical/chemical alteration) can strongly affect floatability, and (3) large variation of waste plastics feed in terms of composition. Flotation allows to separate PS, PVC, PET, PC, and mixed polyolefins (MPO).

The flotation process depends on several design and operational variables. We consider a superstructure that includes the following three flotation stages: the rougher, which processes the feed; the cleaner, which generates the final concentrate; and the scavenger, which generates the final tailing, as shown in Fig. (1). This is a simple superstructure but is used here as an example.

The objective is to maximize the total income with respect to the operation conditions and process design. The decision variables to be optimized are divided into design and operating variables. The design variables include equipment dimensions, such as the cell volume and total number of cells for each stage. The operating variables correspond to operating times for each cell at each stage and the directions of tails and concentrate streams. In stochastic problems, the operating variables (second level variables) are able to adapt to each scenario to increase the total income. Moreover, the design variables (first level) are the same for all scenarios.

The flotation process depends on several design and operation variables. We consider a superstructure that includes three flotation stages: rougher, scavenger and cleaner stages, as is shown in figure 1. We allow for the consideration of multiple scenarios. The model consider constraints that enforce the kinetics of flotation and the mass balance on each flotation stage, the behavior at the splitters and mixers, the mass balance at the splitters and mixers, direction choice in the splitters, the penalty the seller must pay for arsenic content in the concentrate, cell volumes, and the costs associated with the flotation cells.

For the deterministic, model we have only a single scenario, and the model then simply maximizes the total income subject to the dynamic and economic constraints. In the stochastic models, we assume we have more than one scenario. Because of this, we need to replace the objective by the maximization of the expected total income. For this, we need the probability of a given scenario. In addition, we know that some of our decision variables can depend on the scenarios. This model corresponds to a stochastic MINLP.

In flotation process, the gas or air bubbles are introduced through culture suspension, and the microalgal biomass get attached to gaseous molecules and accumulated on the liquid surface. This method is particularly effective for thin microalgae suspension that could be simply gravity thickening [38]. The basic variations of this process are dispersed air flotation, dissolved air flotation, electroflotation, and ozone flotation [55,56,57]. The ratio of gaseous molecules to microalgae is one of the most important factors affecting the performance of the flotation efficiency. Several researchers have confirmed that ozone flotation was more effective than other methods [58,59]. Also, ozoflotation could improve lipid recovery yields and modify fatty acid methyl ester (FAME) profiles. The ozone flotation could increase the cell flotation efficiency by modifying the cell wall surface and/or releasing the active agents from microalgal cells [60]. Moreover, the ozone flotation can also improve the quality of water by lowering the turbidity and organic contents of the effluent [58]. Flotation separation efficiency relates to bubble size [61]. Smaller size of gas bubbles has lower rise velocity and higher surface area to volume ratio. This enables their longer retention time and better attachment efficiency with the microalgae cells and leads to the increasing in harvesting efficiency by floatation [64]. Thus, one of the most efficient ways of achieving maximum attachment is by generating as many small bubbles as possible [61,62,63]. Combinations of flocculation with flotation have been also used to increase the harvesting efficiency [64,65,66].

In using these equations, however, one must use parameters with consistent units.(1-3)E=(Ci-Co)/Ci(1-4)E=K/(Qw-K')(1-5)E=(6Kpr2hqg)/(qwdb)whereE = efficiency per cellCi = inlet oil concentrationCo = outlet oil concentrationQw = liquid flow rate, BPDKp = mass transfer coefficientr = radius of mixing zoneh = height of mixing zoneqg = gas flow rateqw = liquid flow through the mixing zonedb = diameter of gas bubble

The froth flotation process is more than a century old and was developed over a long period of time [8]. It takes advantage of the surface chemistry of fine particlesif one particles surface is hydrophobic and another is hydrophilic, upon generation of air bubbles, the hydrophobic particles tend to attach to the air bubbles and float, allowing for a separation between particles in the froth and those in the main body of the liquid.

Typically three different types of chemicals are used in the froth flotation process: collector, frother, and modifier. First, the collector is added to the iron ore slurry to selectively coat the iron oxide particles, making the surface hydrophobic. The slurry then goes to a flotation cell, where air bubbles are generated using an impeller and aerator (Figure 1.2.4). At this step, the frother (for example, fuel oil) is added to the ore slurry to form stable froth and air bubbles. Iron oxide particles stick to the air bubbles and float. Floated and concentrated iron ore slurry is then skimmed from the surface of the bath, and water is removed using a filter press. If the desired iron content is not achieved, the process is repeated. A modifier is added in some cases to enhance the performance of the collector. Frother is the most important chemical that must always be present. Without the generation of stable air bubbles, hydrophobic particles will not have anything to attach to and will not separate from the bulk solution.

Depending on the type of collector, either iron oxide or silica particles can be floated. An anionic collector is added to float the iron oxide particles, a cationic collector for the silica particles [9]. Depending on the situation, the pH of the slurry can be adjusted by adding acid to the solution, which may also enhance the properties of the collector.

The basic objective of a flotation device is to keep the pulp in suspension and provide the air bubbles. The size of air bubbles matters as it controls flotation kinetics as well as the carrying capacity of the bubbles. The design technology determines the characteristics of the machine, resulting in concomitant factors like how the collision and contact between air bubbles and particles takes place. The two resultant products, concentrate and tails, need to be evacuated properly. The most widely used flotation machines can be broadly classified into mechanical and pneumatic depending on various factors. The former use impellers or rotors, which are absent in the latter.

The shape of a mechanical flotation tank is essentially rectangular, U-shaped, conical or cylindrical, according to the cell type and size. It is fitted with an impeller/rotor and stator/diffuser. Air enters into the device through a concentric pipe surrounding the impeller shaft either by self-aspiration or aided by a compressor. The function of the rotating impeller is to keep particles in suspension by thoroughly mixing the slurry and dispersing the injected air into fine bubbles through a diffuser. It also provides conditions for promoting particlebubble collisions.

There is a necessity for the generation of three different hydrodynamic zones for effective flotation. The region near the impeller comprises of a turbulent area required for solids suspension, dispersion of air into bubbles and bubbleparticle interaction. Above the turbulent region lies a quiescent zone where the bubbleparticle aggregates move up in a relatively less turbulent sector. This zone also helps in sinking the amount of gangue minerals that may have been entrained mechanically. The third region overhead the quiescent zone is the froth zone serving as an additional cleaning step, and improves the grade of the concentrate. Particles that do not attach to the bubbles are discharged out from the bottom of the cell (Vazirizadeh, 2015). Fig. 5.33 shows a typical schematic of a mechanical cell.

Mechanical cells are arranged in a series called a bank, having enough cells to assure the required particle residence time for adequate recovery, the subaeration cells are arranged in cell-to-cell flow, while the supercharged machines are placed in an open-flow design.

The strongly hydrophobic and optimised-sized particles are likely to float first in a bank of flotation cells. Sluggish flowing particles float in diminishing order, and so forth, giving rise to total recovery of about 100%. A minimum of four cells is required for coal flotation with a residence time of 5 minutes (Euston et al., 2012). The residence time, pulp volume and flotation kinetics play a vital role in determining the selection of the number of cells required in a flotation circuit. To prevent loss of floatable coal along with tailings, it is advisable to put cells in series. Fig. 5.34 indicates the coal recovery through multiple cells (in series) in a bank. Fig. 5.35 demonstrates arrangement of cells both in series and parallel, the series arrangement gives optimum recovery of combustibles.

The most common examples of pneumatic cells are the column cell and the Jameson cell. As shown in Fig. 5.36, a flotation column is typically a tall vertical cylinder. It is fed with coal pulp at the top third of column. It has no mobile parts or agitators. Air bubbles are injected either through external or internal spargers at the bottom. These bubbles rise up in countercurrent with the descending flow of the pulp. Hydrophobic particles attach to the air bubbles forming bubbleparticle aggregates and move upwards. The zone where this process takes place is called the collection zone. The ascending bubbleparticle aggregates accumulate in the upper part of the column called the cleaning or froth zone, and then overflow into a launder as a concentrate. Wash water is sprinkled at the top of the column to wash off entrained gangue (hydrophilic) particles, which are sent back into the collection zone. The application of wash water helps stabilise the froth and produce high-grade froth concentrates. The hydrophilic particles, along with misplaced hydrophobic particles, are finally released at the bottom of the column.

In spite of improved separation performance, low capital and operational cost, less plant space demand, low maintenance cost, ease of operation, lower energy consumption and adaptability to automatic control (Wills and Napier-Munn, 2006), axial mixing can significantly reduce the overall performance, particularly in larger-diameter columns. Axial mixing can be decreased by different methods (Kawatra and Eisele 1999, 2001):

A Jameson cell is schematically presented in Fig. 5.37. A high-pressure jet, created by pumping feed slurry through the slurry lens orifice, enters a cylindrical device called a downcomer. The downcomer acts as an air entrainment device which sucks air from the atmosphere. The jet of slurry disseminates the entrained air into very fine bubbles after plunging upon the liquid surface. Then, it creates very favourable conditions for collision of bubbles and particles, and their attachment. The particlebubble aggregates move down the downcomer to the cell and float to the top to form the froth. The hydrophilic minerals sink to the bottom and exit as tailings. Tailings recycling is practiced to reduce feed variations to the cell so that the downcomer can operate at a stable feed pressure and flow rate. This helps to ensure steady operation. The downcomer provides an ideal situation for particlebubble contact and minimises the residence time due to rapid kinetics and separate contact zone. Thus, the Jameson cell is of much lower volume compared to equivalent-capacity column or mechanical cells. There is also no requirement for agitators or compressors besides the feed pump.

The dissolved air flotation process takes advantage of the principles described above. Figure 7-104 presents a diagram of a DAF system, complete with chemical coagulation and sludge handling equipment. As shown in Figure 7-104, raw (or pretreated) wastewater receives a dose of a chemical coagulant (metal salt, for instance) and then proceeds to a coagulation-flocculation tank. After coagulation of the target substances, the mixture is conveyed to the flotation tank, where it is released in the presence of recycled effluent that has just been saturated with air under several atmospheres of pressure in the pressurization system shown. An anionic polymer (coagulant aid) is injected into the coagulated wastewater just as it enters the flotation tank.

The recycled effluent is saturated with air under pressure as follows: a suitable centrifugal pump forces a portion of the treated effluent into a pressure holding tank. A valve at the outlet from the pressure holding tank regulates the pressure in the tank, the flow rate through the tank, and the retention time in the tank, simultaneously. An air compressor maintains an appropriate flow of air into the pressure holding tank. Under the pressure in the tank, air from the compressor is diffused into the water to a concentration higher than its saturation value under normal atmospheric pressure. In other words, about 24 ppm of air (nitrogen plus oxygen) can be dissolved in water under normal atmospheric pressure (14.7 psig). At a pressure of six atmospheres, for instance (6 14.7 = about 90 psig), Henry's law would predict that about 6 23, or about 130 ppm, of air can be diffused into the water. In practice, dissolution of air into the water in the pressurized holding tank is less than 100% efficient, and a correction factor, f, which varies between 0.5 and 0.8, is used to calculate the actual concentration.

After being held in the pressure holding tank in the presence of pressurized air, the recycled effluent is released at the bottom of the flotation tank, in close proximity to where the coagulated wastewater is being released. The pressure to which the recycled effluent is subjected has now been reduced to one atmosphere, plus the pressure caused by the depth of water in the flotation tank. Here, the solubility of the air is less, by a factor of slightly less than the number of atmospheres of pressure in the pressurization system, but the quantity of water available for the air to diffuse into has increased by the volume of the recycle stream.

Practically, however, the wastewater will already be saturated with respect to nitrogen, but may have no oxygen, because of biological activity. Therefore, the solubility of air at the bottom of the flotation tank will be about 25 ppm, and the excess air from the pressurized, recycled effluent will precipitate from solution. As this air precipitates in the form of tiny, almost microscopic, bubbles, the bubbles attach to the coagulated solids. The presence of the anionic polymer (coagulant aid), plus the continued action of the coagulant, causes the building of larger solid conglomerates, entrapping many of the adsorbed air bubbles. The net effect is that the solids are floated to the surface of the flotation tank, where they can be collected by some means and thus be removed from the wastewater.

Some DAF systems do not have a pressurized recycle system, but, rather, the entire forward flow on its way to the flotation tank is pressurized. This type of DAF is referred to as direct pressurization and is not widely used for treatment of industrial wastewaters because of undesirable shearing of chemical flocs by the pump and valve.

The behavior of coal in the flotation process is determined not only by a coals natural floatability (hydrophobicity), but also by the acquired floatability resulting from the use of flotation reagents. The general classification of the reagents for coal flotation is shown in Table12.1 (Laskowski, 2001).

The use of liquid hydrocarbons (oils) as collectors in flotation of coal is characteristic for the group of inherently hydrophobic minerals (graphite, sulfur, molybdenite, talc, coals are classified in this group). Since oily collectors are water-insoluble, they must be dispersed in water to form an emulsion. The feature making emulsion flotation different from conventional flotation is the presence of a collector in the form of oil droplets, which must collide with mineral particles in order to enhance the probability of particle- to-bubble attachment. The process is based on selective wetting: the droplets of oil can adhere only to particles that are to some extent hydrophobic. The effect of emulsification on flotation has been studied, and its beneficial effect on flotation is known (Sun et al., 1955).

Coal flotation is commonly carried out with a combination of an oily collector (e.g. fuel oil) and a frother (e.g. MIBC). All coal flotation systems require the addition of a frother to generate small bubbles and to create a stable froth (Table 12.2). Typical addition rates for frothers are in the order of 0.050.3kg of reagent per tonne of coal feed. Depending on the hydrophobic character of the coal particles, an oily collector such as diesel oil or kerosene may or may not be utilized. When required, dosage rates commonly fall in the range of 0.21.0kg of reagent per tonne of coal feed, although dosage levels up to 2kg/t or more have been known to be used for some oxidized coals that are difficult to flotate.

PO stands for propylene oxide (CH2-CH2-CH2-O-), and BO for butylene oxide (CH2-CH2-CH2-CH2-O-) Cresylic acids (mixture of cresols and xylenols) that in the past were commonly used in coal flotation are not in use any more because of their toxicity.

PO stands for propylene oxide (CH2-CH2-CH2-O-), and BO for butylene oxide (CH2-CH2-CH2-CH2-O-) Cresylic acids (mixture of cresols and xylenols) that in the past were commonly used in coal flotation are not in use any more because of their toxicity.

The beneficial effect of a frother on flotation with an oily collector was demonstrated and explained by Melik-Gaykazian et al. (1967). Frother adsorbs at the oil/water interface, lowers the oil/water interfacial tension and hence improves emulsification. However, frother also adsorbs at the coal/water interface (Frangiskos et al., 1960; Fuerstenau and Pradip, 1982; Miller et al., 1983) and provides anchorage for the oil droplets to the coal surface. Chander et al. (1994), after studying various non-ionic surfactants, concluded that the flotation of coal can be improved in their presence because of the increased number of droplets, which leads to an increase in the number of droplet-to-coal particle collisions. While the use of oily collectors and frothers is the most common, also a group of flotation agents known as promoters have found application in coal flotation. In general, these are strongly surface-active compounds and are mostly used to enhance further emulsification of water-insoluble oily collectors in water.

Because of environmental concerns associated with tailing ponds, the method for disposing of fine refuse from coal preparation plants by underground injection has been gaining wide acceptance. Unfortunately, many common flotation reagents, including diesel oil, are not permitted when fine refuse is injected underground into old mine works. This is the main driving force for finding replacement for the crude-oil based flotation collectors (Skiles, 2003). An alternative to fuel oil may be biodiesel, a product created by the esterification of free fatty acids generally from soy oil, with an alcohol such as methanol, and subsequent transesterification of remaining triglycerides. Water, glycerol and other undesirable by-products are removed, to produce a product that has physical characteristics similar to diesel oil. The use of some vegetable oils was demonstrated to provide equivalent (and even superior) flotation results when compared with diesel fuel (Skiles, 2003). These are the results of commercial scale tests on a circuit that has 4.25m in diameter columns. The product concentrate ash was 13.5%. The consumption of the tested vegetable oil was about two times lower from the consumption of diesel oil in these tests.

It features both proven technology and the latest technical innovations at the same time. This flotation cell is highly efficient when it comes to costs and operation. It can be easily scaled to various production levels without compromising performance. In short, OptiCell Flotation enhances the performance of the deinking line cost efficiently and ensures a reliable flotation process. The heart of the flotation process is the injector. In the OptiCell system, this has beendesigned with special care, using the experiences of earlier flotation technologies, modern computational fluid dynamics calculations, and new image analysis methods. The combination of these approaches results in a unique injector that represents the latest technology. This injector differs from traditional injectors in following respects:

OptiCell flotation by Metso is based on computational fluid dynamics and uses new image analysis methods. It is designed to provide smooth-flow velocities that allow unobstructed transfer of bubbles to the surface of the pulp mixture or froth, which improves the efficiency of ink removal. The aeration injector ensures optimal bubble-size distribution. The injector is designed based on the experiences gained with earlier flotation technologies combined with modern computational fluid dynamics calculations and new image analysis methods.

The linear structure of the flotation cells has a large surface area, which has reject separation and fiber loss. This flotation cell design also contributes to high sludge consistency (less water in the sludge) by ensuring smooth drainage of froth (Aksela,2008). The elliptical shape of the flotation cells in this technology is optimal for internal pulp circulation for improved ink removal. Moreover, the flatness of the cells intensifies the rise of air bubbles within the available volume. The first OptiCell Flotation system started operation in September 2008 at Stora Ensos Maxau mill in Germany, which has an approximately 1000 t/day deinking facility (Metso,2012b). According to Metso, the brightness from the complete flotation system has increased by two units. A brightness gain of 13 units from thick stock to accept was obtained with the OptiCell process. Reject ash content also improved and fiber losses decreased. As a result of the flotation performance and corresponding brightness improvement, peroxide consumption has decreased significantly in bleaching. In addition, the stickies content was reduced significantly. It was the lowest ever measured at the deinking line 1 at Maxau mill. The benefits of OptiCell flotation are summarized in Table11.8 (Aksela,2008; Metso,2012b).

flotation machine factory, buy good quality flotation machine products from china

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circularflotationmachine-yantai jinpeng mining machinery, ore dressing process & equipment, ore flotation & beneficiation

circularflotationmachine-yantai jinpeng mining machinery, ore dressing process & equipment, ore flotation & beneficiation

The structure of forced-air circular flotationcell is shown in the following figure.Thosecells can be adopted as rougher, scavenger and cleaner. The capacity ofa singleflotationcellcan be up to 680m3.

flotation feed - an overview | sciencedirect topics

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].

advanced flotation technology | eriez flotation division

advanced flotation technology | eriez flotation division

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

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

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

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

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

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

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

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

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

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

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

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

phosphate beneficiation process

phosphate beneficiation process

Large scale mining and processing of phosphate is essential for operating at a profit. In the Florida area, Phosphate Beneficiation by flotation unlocked the door to vast tonnages of ore which in the past could not be recovered by conventional washing methods which saved only the coarser pebble phosphate. Many of the areas now being mined contain very little or no pebble phosphate, so the main recovery is from the fine sands.

In the treatment of phosphatic shales for recovery of phosphate, a simple low cost flexible flowsheet is highly desirable. Since all grades of ore from low to high P2O5 content may occur in a deposit it is important to consider the possibility of either mixing the ores or to segregate them into grades for separate treatment. Laboratory and pilot plant work is very valuable in establishing the final treatment which will give the maximum net economic return.

Electrically operated drag lines strip off the overburden from the mining area and deposit the phosphate matrix around a pump pit. Here it is sluiced with streams of high pressure water to the suction of a large centrifugal pump which transports the matrix slurry to the washing plant which may be a mile or two away.

At the washer the water-matrix slurry discharges into a surge receiving bin or tub and is screened for removal of clay, sand and fine phosphate from the mud balls. The screen oversize passes through a hammer mill to break down the mud balls and occasional large pebbles. The screen undersize and disintegrated mud balls pass to a pebble washing and screening section consisting of a trommel screen and log washer. Here, clean washed pebble phosphateusually plus 14 mesh, depending upon the character of matrixbeing minedis removed. Tramp oversize is recycled to the hammer mill for further disintegration, and 14 mesh matrix slurry passes to the fine recovery section.

Sub-A Flotation Machine (Phosphate Type). Please note free-flow of pulp for fast-floating ore and high tonnage operation. Shaft assembly is removable as one unit. Repeated pulp circulation (as indicated) assures proper agitation and aeration. Supercharged air can be added down standpipe or down hollow shaft, or both, as conditions require.

Following removal of the pebble, the balance of the ore or matrix flows by gravity to a large hydroclassifier for separation of the sands from the slimes at 150 mesh. The 150 mesh slimes containing the colloidal clay, very fine silica and phosphate are discarded.

Fines are removed from the storage bins by a centrifugal pump, fed through pinch valves and regulated by high pressure water jets. The pump discharges into a spiral or rake classifier, the overflow going to a secondary hydroclassifier, and the sands to a hydraulic classifier. This classification system splits out a +20 mesh pebble phosphate, a 20 +35 mesh fraction for agglomerate separation, and 35 mesh which is the feed to flotation. All classifier overflow and excess water accumulated in the washing steps are diverted back to the large primary hydroclassifier for retreatment. Wet cyclones can be successfully applied in the classification systems to eliminate slime and excess water from the fines.

The 20 +35 mesh phosphate-sand mixture is further deslimed and conditioned with reagents at high density, 70-75% solids in a rotary drum type conditioner. The phosphate particles are filmed with the fuel oil, fatty acid, caustic mixture which renders them non- wettable. This coarse pulp then is subjected to tabling, spiraling, or belt agglomerate treatment which separates the phosphate from the silica particles. This usually produces a finished, high grade phosphate concentrate and a clean waste product. In some cases it may be necessary to clean the phosphate product by silica flotation, as indicated by the alternate flow lines.

The 35 mesh matrix is fed from the fine recovery bins to a rake or spiral classifier for further desliming, and the sands at high density, 65 to 70% solids, are introduced in a Heavy Duty Open Type Duplex Phosphate Conditioner. Two or more conditioning tanks are generally used in series. Caustic soda, fuel oil, and tall oil are metered to the conditioner feed and agitated to thoroughly film all the phosphate particles.

The discharge from the conditioning circuit is diluted down to about 25% solids and fed directly to a Sub-A Flotation Machine. Since tonnages are high, usually in excess of 100 long tons per hour to each circuit, the No. 30 (5656) Sub-A Flotation Machinehaving 100 cubic ft. of volume per cellis standard for the fatty acid flotation separation. Phosphate is removed in the froth product and the silica passes out the end of the machine to waste. Flotation is very rapid when the feed is properly conditioned and metered to the machine. Usually at least 4 cells are used in series for each circuit. In some cases, one stage of cleaning is necessary to produce an acceptable grade in the fatty acid flotation section. Grade is usually 68 to 72% BPL (Bone Phosphate of Lime).

Neoprene wearing parts are necessary to withstand the action of the fatty acid-fuel oil flotation reagents and large adjustable sand reliefs are standard in the Cells for phosphate flotation. Because of the high percentage of phosphate to be removed in the froth product, a double overflow machine with froth paddles is standard. Supercharging improves performance and reduces power requirements for flotation.

Usually, when it is necessary to produce concentrates having 72-78% BPL, the fatty acid flotation concentrate is subjected to reverse flotation in which the phosphate is depressed and the silica contaminant activated and floated in a separate circuit.

The fatty acid froth product, in this case, flows by gravity to another set of Heavy Duty Acid Proof Conditioners. These are neoprene lined. Sufficient concentrated sulphuric acid is added to the pulp to produce a low acid pH. This cuts and removes the fatty acid reagents from the phosphateparticles. The reagent and acid water is removed by thorough washing and desliming. The classifier sand product containing the de-activated phosphate and silica impurity passes to a second Sub-A Flotation Circuit for removal of silica in a froth product. Cationic reagents, such as amine acetate, are used to activate and float the silica.

In some circuits, the phosphate concentrate from the coarse agglomerate separation section, if not high enough grade, is introduced into the silica flotation section along with the product from the fatty acid flotation section. This is done when a high purity product, usually 76-78% BPL, is required. The flotation circuits are set up for maximum flexibility to accommodate changes in tonnage and character of feed as well as requirements on the finished products.

In the treatment of phosphatic shales for recovery of phosphate, a simple low cost flexible flowsheet is highly desirable. Since all grades of ore from low to high P2O5 content may occur in a deposit it is important to consider the possibility of either mixing the ores or to segregate them into grades for separate treatment. Laboratory and pilot plant work is very valuable in establishing the final treatment which will give the maximum net economic return.

Extensive tests have established that where mining can be controlled the ores should be selectively mined and treated as two distinct operations, namely: to produce ore with a low or medium phosphate content and an ore with high phosphate content. For the low grade and medium grade ores, containing 18 to 22% P2O5, generally a coarse waste product can beproduced and rejected. Treatment of the high grade ores containing about 28% P2O5 will generally result in the production of a coarse phosphate product. This is in addition to production of granular fines as an acceptable phosphate and a slime waste product.

In case only one crusher is used for both types of ore, the crushing period can be divided to accommodate the respective tonnages. Usually crushing is confined to one or two shifts per day, depending upon the size of the operation.

Water is added to give a feed density of approximately 67% solids. Retention time in the scrubber is important to thoroughly break down and scour the slime from the ore. Retention time will vary from 5 to 30 minutes depending upon the grade and character of the ore. The scrubbed ore passes the trommel section of the scrubber where the + oversize is removed. Sprays are applied to give a clean oversize product.

In the case of low and medium grade phosphatic shale ores the + fraction is low grade and is rejected as a waste product. This usually amounts to 9-10% of the feed tonnage. The high grade ores, on the other hand, when treated in this manner, give an oversize product + sufficiently high in P2O5 content to be a finished product. About 8% of the weight represents this fraction of acceptable product + size.

In the case of the low grade ores, the screening is done at 35 mesh. The screen oversize +35 mesh is usually not a finished product and therefore requires further grinding. High grade ores will generally produce an acceptable product at +20 mesh without further treatment.

Low grade - to 35 mesh screened and washed oversize is reduced to minus 35 mesh in a pheripheral discharge Rod Mill which is in closed circuit with the vibrating screen. An elevator or SRL Pump can be used to transfer the mill discharge to the screen. The rod mill action polishes off the softer shale fraction from phosphate particles and keeps sliming of phosphate to a minimum. The feed to the rod mill will approximate 12% of the initial feed tonnage.

Extensive pilot plant tests have shown that wet cyclones provide a very efficient method for removal of slimes which are largely minus 400 mesh. The slimes, so produced, are low in P2O5 content and are discarded to waste. Two stages of cones are required and the underflow from the secondary stage constitute a final product ready for filtration.

In the case of low and medium grade ores, the 35 mesh cone feed represents about 90% of the plant tonnage and will reject 35 to 36% as slimes to waste. The 20 mesh feed for high grade ores represents about 85% of the plant tonnage and the overflow to waste will be about 15 to 16%. Underflow from the primary cones constitutes feed to the secondary cones.

Secondary cones in the case of low and medium grade ores, receive 62% of the initial plant feed. The overflow is recycled back to the primary cone feed. The overflow fraction amounts to 6-7% of the original plant feed. The underflow, representing about 55% of the plant feed at 65% solids, is ready for filtration. In the case of high grade ores, the secondary cone feed is about 70% and the recycled overflow is about 4% The underflow product representing 65 to 66% at 65% solids is fed to the filter.

No water needs to be added to the primary cone pump sump. Water is necessary in the secondary cone feed pump and the resulting cone overflow then becomes dilution for the primary cone circuit. Primary cone overflow will vary between 4 and 8% depending upon the type of ore being treated.

The discharge from the secondary cones is a slime free granular product containing about 65% solids. This is ideal filter feed, but it is necessary to use a top feed or horizontal filter for efficient de-watering.

The slime overflow from the primary cones at 4 to 8% solids is fed to a thickener for water reclamation. The slimes will settle to about 25% solids and are pumped from the underflow to tailing ponds. Special precautions should be taken when impounding this material due to the almost complete absence of sand. Several separate ponds may be necessary to store this waste.

With reclamation of plant water from the slime tailing in a thickener, the amount of new water for low and medium grade ores will be about 300 gallons per ton of ore treated. High grade ores require about 150 gallons per ton. This is well within the range of average mill water requirements when reclamation is a part of the system.

Phosphate rock being a low-priced material is produced as near the fertilizer market as possible and haulage costs determine production. Flotation of fine sand from the pebble mines in Florida is economical because the material has been mined and presents a disposal problem if not salvaged.

Generally a fatty-acid reagent combination is employed to float the phosphate from the silica to produce a grade of 70 to 72 per cent BPL. Reagents are caustic soda, fuel oil, and Tall oil added to the pulp and conditioned at high density, 70 to 75 per cent solids, before flotation. For producing premium grade phosphate, the fatty-acid-floated product is conditioned with sulphuric acid to neutralize and cut off the oil film, washed, repulped, and the silica floated with a cationic reagent such as amine acetate. Flotation is very rapid. It is very important to employ the proper high-density conditioning technique to bring about thorough activation and selectivity in both the fatty acid and cationic flotation steps.

Flow-sheet of a Florida phosphate plant recovering 35+150 mesh phosphate by flotation. De-sliming and conditioning at 65 to 70 percent solids with reagents is essential for proper separation by flotation.

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