best flotation cell machiness for ballast

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].