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

mineral processing, equipment manufacturers, ball mills, flotation, thickener - xinhai

mineral processing, equipment manufacturers, ball mills, flotation, thickener - xinhai

Xinhai devotes to providing Turn-key Solutions for Mineral Processing Plant (EPC+M+O), namely design and research - complete equipment manufacturing and procurement - commissioning and delivery - mine management - mine operation. The essence of EPC+M+O Service is to ensure sound work in every link. The model is suitable for most of the mines in the world.

Focusing on the research and development and innovation of mineral processing equipment, Xinhai has won more than 100 national patents, strives for perfection, strives to complete the combination of equipment and technology, improve productivity, reduce energy consumption, extend equipment stable operation time, and provide cost-effective services.

With Class B design qualifications in the metallurgical industry, rich in ore mining, beneficiation, smelting technology and experience, completed more than 2,000 mine design and research, not only can provide customers with a reasonable process, but also can provide customized equipment configuration.

The precious metal minerals are mainly gold and silver mines. Xinhai Mining has more than 20 years of experience in beneficiation for gold and silver mines, especially gold ore beneficiation technology. Gold craft and placer gold selection craft etc.

With Class B design qualification, it can provide accurate tests for more than 70 kinds of minerals and design a reasonable beneficiation process. In addition, it can also provide customized complete set of mineral processing equipment and auxiliary parts.

Xinhai can provide the all-round and one-stop mineral processing plant service for clients, solving all the mine construction, operation, management problems, devoting to provide modern, high-efficiency.

Through mineral processing experiment, the mineral processing flow is customized. Multiple tests are carried out in every link, and make sure the final processing flow to guarantee the successful mineral processing plant construction.

According to tailing processing technology, Xinhai has tailings reprocessing technology and tailings dry stacking. Tailings dry stacking is the self-launched tailings dewatering technology, which is the effective technology in green mine construction.

More than 2,000 mine design and research, equipment supply projects, more than 500 mining industry chain services (EPC+M+O) projects in more than 90 countries and regions around the world, we are always committed to providing you with one-stop, customized Chemical mine solution!

flotation reagents

flotation reagents

This data on chemicals, and mixtures of chemicals, commonly known as reagents, is presented for the purpose of acquainting those interested in frothflotation with some of the more common reagents and their various uses.

Flotation as a concentration process has been extensively used for a number of years. However, little is known of it as an exact science, although, various investigators have been and are doing much to place it on a more scientific basis. This, of course, is a very difficult undertaking when one appreciates how ore deposits were formed and the vast number of mineral combinations existing in nature. Experience obtained from examining and testing ores from all over the world indicates that no two ores are exactly alike. Consequently, aside from a few fundamental principles regarding flotation and the use of reagents, it is generally agreed each ore must be considered a problem for the metallurgist to solve before any attempt is made to go ahead with the selection and design of a flotation plant.

Flotation reagents may be roughly classified, according to their function, into the following groups: Frothers, Promoters, Depressants, Activators, Sulphidizers, Regulators. The order of these groups is no indication of their relative importance; and it is common for some reagents to fall into more than one group.

The function of frothers in flotation is that of building the froth which serves as the buoyant medium in the separation of the floatable from the non-floatable minerals. Frothers accomplish this by lowering the surface tension of the liquid which in turn permits air rising through the pulp to accumulate at the surface in bubble form.

The character of the froth can be controlled by the type of frother. Brittle froths, those which break down readily, are obtained by the alcohol frothers. Frothers such as the coal tar creosotes produce a tough bubble which may be desirable for certain separations.

Flotation machine aeration also determines to a certain extent the character of the froth. Finely divided air bubbles thoroughly diffused through the pulp are much more effective than when the same volume of air is in larger bubbles.

In practice the most widely used frothers are pine oil and cresylic acid, although, some of the higher alcohols are gradually gaining favor because of their uniformity and low price. The frothers used depends somewhat upon the location. For instance, in Australia eucalyptus oil is commonly usedbecause an abundant supply is available from the tree native to that country.

Frothers are usually added to the pulp just before its entrance into the flotation machine. The quantity of frother varies with the nature of the ore and the purity of the water. In general from .05 to .20 lbs. per ton of ore are required. Some frothers are more effective if added in small amounts at various points in the flotation machine circuit.

Overdoses of frother should be avoided. Up to a certain point increasing the amount of frother will gradually increase the froth produced. Beyond this, however, further increases will actually decrease the amount of froth until none at all is produced. Finally, as the excess works out of the system the froth runs wild and this is a nuisance until corrected.

Not enough frother causes too fragile a froth which has a tendency to break and drop the mineral load. No bare spots should appear at the cell surface, and pulp level should not be too close to the overflow lip, at least in the cells from which the final cleaned concentrate is removed.

A good flotation frother must be cheap and easily obtainable. It must not ionize to any appreciable extent. It must be an organic substance. Chemically a frother consists of molecules containing two groups having opposite properties. One part of the molecule must be polar in order to attract water while the other part must be non-polar to repel water. The polar group in the molecule preferably should contain oxygen in the form of hydroxyl (OH), carboxyl (COOH), carbonyl (CO); or nitrogen in the amine (NH2) or the nitrile form. All of these characteristics are possessed by certain wood oils such as pine oil and eucalyptus oil, by certain of the higher alcohols, and by cresylic acid.

The function of promoters in flotation is to increase the floatability of minerals in order to effect their separation from the undesirable mineral fraction, commonly known as gangue. Actuallywhat happens is that the inherent difference in wettability among minerals is increased and as a result the floatability of the more non-wettable minerals is increased to the point where they have an attraction for the air bubbles rising to the surface of the pulp. In practical operation the function of promoters may be considered two-fold: namely, to collect and select. Certain of the xanthates, for instance, possess both collective and selective powers to a high degree, and it is reagents such as these that have made possible some of the more difficult separations. In bulk flotation all of the sulphide minerals are collected and floated off together while the gangue remains unaffected and is rejected as tailing. Non- selective promoters serve very well for this purpose. Selective or differential flotation, on the other hand, calls for promoters which are highly selective or whose collecting power may be modified by change in pulp pH (alkalinity or acidity), or some other physical or chemical condition.

The common promoters for metallic flotation are xanthates, aerofloats, minerec, and thiocarbanilide. Soaps, fatty acids, and amines are commonly used for non-metallic minerals such as fluorspar, phosphate, quartz, felpsar, etc.

Promoters are generally added to the conditioner ahead of flotation to provide the time interval required for reaction with the pulp. Some promoters are slower in their action and in such case are added directly to the grinding circuit. Promoters which are fast acting or have some frothing ability are at times added directly to the flotation machine, as required, usually at several points. This practice is commonly known as stage addition of reagents.

The quantity of promoter depends on the character and amount of mineral to be floated, and in general for sulphide or metallic minerals .01 to .20 lbs. per ton of ore are required. Flotation of metallic oxides and non-metallic minerals usually require larger quantities of promoter, and in the case of fatty acids the range is from 0.5 to 2.5 lbs. per ton.

The function of depressants is to prevent, temporarily, or sometimes permanently, the flotation of certain minerals without preventing the desired mineral from being readily floated. Depressants are sometimes referred to as inhibitors.

Lime, sodium sulphite, cyanide, and dichromate are among the best known common depressants. Among organic depressants, starch and glue find widest application. If added in sufficient quantity starch will often depress all the minerals present in an ore pulp. Among the inorganic depressants, lime is the cheapest and best for iron sulphides, while zinc sulphate, sodium cyanide, and sodium sulphite depress zinc sulphide. Sodium silicate, quebracho, and also cyanide are commondepressants in non-metallic flotation.

Depressants are generally added to the grinding circuit or conditioner usually before addition of promoting and frothing reagents. They may also be added direct to the flotation cleaner circuit particularly on complex ores when it is difficult to make a clean cut separation or where considerable gangue may be carried over mechanically into the cleaning circuit as in flotation of fluorspar. Quantity of depressants required depends on the nature of the ore treated and should be determined by actual test. For instance, lime required to depress pyrite may vary from 1 to 10 lbs. a ton.

The function of activators is to render floatable those minerals which normally do not respond to the action of promoters. Activators also serve to render floatable again minerals which have been temporarily depressed in selective flotation. Sphalerite depressed with cyanide and zinc sulphate can be activated with copper sulphate and it will then respond to treatment like a normal sulphide. Stibnite, the antimony sulphide mineral, responds much better to flotation after being activated with lead nitrate.

The theory generally accepted on activation is that the activating substance, generally a metallic salt, reacts with the mineral surface to form on it a new surface more favorable to the action of a promoter. This also applies to non-metallic minerals.

Activators are usually added to the conditioner ahead of flotation and in general the time of contact should be carefully determined. Amounts required will vary with the condition of the ore treated. In the case of zinc ore previously depressed with zinc sulphate and cyanide, from 0.5 to 2.0 of copper sulphate may be required for complete activation. Quantities required should always be determined by test.

The most widely used sulphidizer is sodium sulphide, which is commonly used in the flotation of lead carbonate ores and also slightly tarnished sulphides such as pyrite and galena. In the sulphidization of ores containing precious metals careful control must be exercised as in some instances sodium sulphide has been known to havea depressing effect on flotation of metallics. In such cases it is advisable to remove the precious metals ahead of the sulphidization step.

Sulphidizers are usually fed into the conditioner just ahead of the flotation circuit. The quantity required varies with the characteristics of the ore and may range from .5 to 5 lbs. per ton. Conditioning time should be carefully determined and an excess of sulphidizing reagent avoided.

The function of regulators is to modify the alkalinity or acidity in flotation circuits, which is commonly measured in terms of hydrogen ion concentration, or pH. Modifying the pH of a pulp has a pronounced effect on the action of flotation reagents and is one of the important means of making otherwise difficult separations possible.

Soluble salts may have their source in the ore or water, or both, and in precipitating them out of solution they generally become inert to the action of flotation reagents. Soluble salts have a tendency to combine with promoters thus withdrawing a certain proportion of the reagents from action on the mineral to be floated. Removal of the deleterious salts therefore makes possible a reduction in the amount of reagent, required. Complexing soluble salts by keeping them in solution yet inert to the reagents is in some cases desirable.

Mineral surfaces may vary according to pulp pH conditions as many of the regulators appear either directly or indirectly to have a cleansing effect on the mineral particle. This brings about more effective action on the part of promoters and other reagents, and in turn increases selectivity.

pH control by action of regulators is in some cases very effective in depressing certain minerals. Lime, for instance, will depress pyrite, and sodiumsilicate is excellent for dispersing and preventing quartz from floating. It is necessary, however, to have a definite concentration of the reagents for best results.

The common regulators are lime, soda ash, and sodium silicate for alkaline circuits, and sulphuric acid for acid circuits. Many other reagents are used for this important function. The separation required and character of ore will determine which regulators are best suited. In general, from an operating standpoint, it is preferable to use a neutral or alkaline circuit, but in some instances it is only possible to obtain results in an acid circuit which then will require the use of special equipment to withstand corrosion. Flotation of non-metallic minerals is at times more effective in an acid circuit as in the case of feldspar and quartz. The pulp has to be regulated to a low pH by means of hydrofluoric acid before any degree of selectivity is possible between the two minerals.

Regulators are fed generally to the grinding circuit or to the conditioner ahead of flotation and before addition of promoters and activators. The amounts required will vary with the character of the ore and separation desired. In the event an excessive quantity of regulator is required to obtain the desired pH it may be advisable to consider removing the soluble salts by water washing in order to bring reagent cost within reason.

The tables on the following pages have been prepared to present in brief form pertinent information on a few of the more common reagents now beingused in the flotation of metallic and non-metallic minerals. A brief explanation of the headings in the table is as follows:

Usual Method of Feeding: Whether in dry or liquid form. A large number of reagents are available in liquid form and naturally are best handled in wet reagent feeders, either full strength or diluted for greater accuracy in feeding. Many dry reagents are best handled in solution form and in such cases common solution strengths are specified in percent under this heading. A 10% water solution of a reagent means 10 lbs. of dry reagent dissolved in 90 lbs. of water to make 100 lbs. of solution. Some dry reagents, because of insolubility or other conditions, must be fed dry. This is usually done by belt or cone type feeders designed especially for this service to give accurate and uniform feed rates.

Pasty, viscous, insoluble reagents present a problem in handling and are generally dispersed by intense agitation with water to form emulsions which can then be fed in the usual manner with a wet reagent feederor using a pump.

Price Per Lb.: Prices shown are approximate and in general apply to drum lots and larger quantities F.O.B. factory. This information is very useful whenmaking tests to determine the lowest cost satisfactory reagent combination for a specific ore. Some ores will not justify reagent expenditures beyond a certain limit, and in this case less expensive reagents must be given first consideration.

Uses: General use for each reagent as given is determined from experience by various investigators. Although the Equipment Company uses a large number of these reagents in conducting test work on ores received from all parts of the world, opinion, data, or recommendations contained herein are not necessarily based on our findings, but are data published by companies engaged in the manufacture of those reagents.

The ore testing Laboratory of 911metallurgist, in the selection of reagents for the flotation of various types of ores, uses that combination which gives the best results, irrespective of manufacturer of the reagents. The data presented on the following tables should be useful in selecting reagents for trials and tests, although new uses, new reagents, and new combinations are continually being discovered.

The consumption of flotation reagents is usually designated in lbs. per ton of ore treated. The most common way of determining the amount of reagent being used is to measure or weigh the amount being fed per. unit of time, say one minute. Knowing the amount of ore being treated per unit of time, the amount of reagent may then be converted into pounds per ton.

The tables below will be useful in obtaining reagent feed rates and quantities used per day under varying conditions. The common method of measurement is in cc (cubic centimetres) per minute. The tables are based on one cc of water weighing one gram. A correction therefore will be necessary for liquid reagents weighing more or less than water. Dry reagents may be weighed directly in grams per min. which in the tables is interchangeable with cc per min.

In the table on the opposite page the 100% column refers to undiluted flotation reagents such as lime, soda ash and liquids with a specific gravity of 1.00. Ninety-two per cent is usually used for light pine oils, 27 per cent for a saturated solution of copper sulphate and 14 per cent for TT mixture (thiocarbanilide dissolved in orthotoluidine). The other percentages are for solutions of other frequently used reagents such as xanthates, cyanide, etc.

The action of promoting reagents in increasing the contact-angle at a water/mineral surface implies an increase in the interfacial tension and, therefore, a condition of increased molecularstrain in the layer of water surrounding the particle. If two such mineral particles be brought together, the strain areas enveloping them will coalesce in the reduction of the tensionary system to a minimum. In effect, the particles will be pressed together. Many such contacts normally occur in a pulp before and during flotation, with the result that the floatable minerals of sufficiently high contact-angle are gathered together into flocks consisting of numbers of mineral particles. This action is termed flocculation , and obviously is greatly increased by agitation.

The reverse action, that of deflocculation , takes place when complete wetting occurs, and no appreciable interfacial tension exists. Under these conditions there is nothing to keep two particles of ore in contact should they collide, since no strain area surrounds them ; they therefore remain in individual suspension in the pulp.

Since substances which can be flocculated can usually be floated, and vice versa, the terms flocculated and deflocculated have become more or less synonymous with floatable and unfloatable , and should be understood in this sense, even though particles of ore often become unfloatable in practice while still slightly flocculatedthat is, before the point of actual deflocculation has been reached.

Here is a ListFlotation Reagents & Chemicals prepared to present in brief form pertinent information on a few of the more common reagents now being used in the flotation of metallic and non-metallic minerals. A brief explanation of the headings in the table is as follows:

Usual Method of Feeding: Whether in dry or liquid form. A large number of reagents are available in liquid form and naturally are best handled in wet reagent feeders, either full strength or diluted for greater accuracy in feeding. Many dry reagents are best handled in solution form and in such cases common solution strengths are specified in percent under this heading. A 10% water solution of a reagent means 10 lbs. of dry reagent dissolved in 90 lbs. of water to make 100 lbs. of solution. Some dry reagents, because of insolubility or other conditions, must be fed dry. This is usually done by belt or cone type feeders designed especially for this service to give accurate and uniform feed rates.

Pasty, viscous, insoluble reagents present a problem in handling and are generally dispersed by intense agitation with water to form emulsions which can then be fed in the usual manner with a wet reagent feeder.

The performance of froth flotation cells is affected by changes in unit load, feed quality, flotation reagent dosages, and the cell operating parameters of pulp level and aeration rates. In order to assure that the flotation cells are operating at maximum efficiency, the flotation reagent dosages should be adjusted after every change in feed rate or quality. In some plants, a considerable portion of the operators time is devoted to making these adjustments. In other cases, recoverable coal is lost to the slurry impoundment and flotation reagent is wasted due to operator neglect. Accurate and reliable processing equipment and instrumentation is required to provide the operator with real-time feedback and assist in optimizing froth cell efficiency.

This process of optimizing froth cell efficiency starts with a well-designed flotation reagent delivery system. The flotation reagent pumps should be equipped with variable-speed drives so that the rates can be adjusted easily without having to change the stroke setting. The provision for remotely changing the reagent pump output from the control room assists in optimizing cell performance. The frother delivery line should include a calibration cylinder for easily correlating pump output with the frother delivery rate. Our experience has shown that diaphragm metering pumps of stainless steel construction give reliable, long-term service. Duplex pumps are used to deliver a constant frother-to-collector ratio over the range of plant operating conditions.

In most applications, the flotation reagent addition rate is set by the plant operator. The flotation reagents can be added in a feed-forward fashion based on the plant raw coal tonnage. Automatic feedback control of the flotation reagent addition rates has been lacking due to the unavailability of sensors for determining the quality of the froth cell tailings. Expensive nuclear-based sensors have been tried with limited success. Other control schemes have measured the solids concentrations of the feed, product, and tailings streams and calculated the froth cell yield based on an overall material balance. This method is susceptible to errors due to fluctuations in the feed ash content and inaccuracies in the measurement device.

A series of simple math models have been developed to assist in the engineering analysis of batch lab data taken in a time-recovery fashion. The emphasis is to separate the over-all effect of a reagent or operating condition change into two portions : the potential recovery achievable with the system at long times of flotation, R, and a measure of the rate at which this potential can be achieved, K.

Such patterns in R and K with changing conditions assist the engineer to make logical judgements on plant improvement studies. Standard laboratory procedures usually concentrate on identifying some form of equilibrium recovery in a standard time frame but often overlook the rate profile at which this recovery was achieved. Study has shown that in some plants, at least, changes in the rate, K, are more important relative to over-all plant performance than changes in the lab measured recovery, R. Thus the R-K analysis can serve to improve the engineering understanding of how to use lab data for plant work. Long term plant experience has also shown that picking reagent systems having higher K values associated can be beneficial even when the plant, on the average, is not experiencing rate of mass removal problems. This is due to the cycling or instabilities that can and do exist in industrial circuits.

It is also important to note that the R-K approach does not eliminate the need for surface chemistry principles and characterization. Such principles and knowledge are required to logically select and understand potential reagent systems and conditions of change in flotation. Without this, reagent selection is quickly reduced to a completely Edisonian approach which is obviously inefficient. What the R-K analysis does is to provide additional information on a system in a critical stage of scale-up (from the lab to the plant) in a form (equilibrium recovery and rate of mass removal) which are interpretable to the engineer who has to make the change work.

The influence of operating conditions such as pH, temperature of feed water, degree of grind, air flow rate, degree of agitation, etc. have been characterized using the R-K approach with clear patterns evolving.

The effect of collector type and concentration on a wide variety of ore types have been studied with generally rather clear and sometimes rather significant patterns in R and K. The quantitative ability to analyze collector performance from the lab to the plant using the R-K profiles has been good.

The effect of frother type on various ores has also been undertaken with good success in differentiating between the qualitative directions and effects involved. However, the actual concentrations required in plants have not, in at least some tests, been accurately predicted. Thus further work remains in this area but in almost all cases the qualitative information on frothers that has been gained has proven very valuable in test work as a guide.

copper mining & extraction process flow chart

copper mining & extraction process flow chart

This flowchart made of machinery icons explains or expresses in simple but clear terms the step of theCopper Mining and Copper Extraction Process. Starting from either open-pit or underground mining and using a different relevant treatment method for oxide or sulphide copper mineral (ore).

Havinga quick look now at how porphyry ores are treated and the metals extracted. There are two main process streams; one for sulfide ores and the other for ore that is being weathered to oxidize sulfides the so-called oxide ores. All ore in the pit is drilled and blasted and loaded into trucks and hauled for treatment if the ore is un-oxidized sulfidic ore then it needs to be crushed and milled to a fine slurry then it gets past through flotation cells in a concentrator to separate and concentrate the sulfides. The top picture shows the interior of a large concentrator with rows of individual flotation cells the floatation agent is added to the slurry and stirred. The floatation agent preferably sticks to the sulfide minerals rather than the waste minerals and then air is bubbled through the mixture and the floatation agent traps the fine bubbles which carry the sulfides to the surface of the cell where they are carried over aware and separated. From there they are dried to provide a concentrate which then goes on to a smelter. This is the same process for both copper and molybdenum porphyries. The smelter is basically a large furnace which melts the concentrate and drives off the sulfide to leave molten copper metal this is still contains impurities and it needs to be refined further to make it a salable product.

Returning to the overall process; that is the process for the sulfide ores and the oxide ore as I said are treated differently. Direct from the pit the oxide ore is piled onto large lined leach pads and the sulfuric acid. The top photo shows one of these leach pads with the new thick black plastic liner visible on the right of the pad. The copper oxide minimum minerals are dissolved by the acid to give a blue copper rich solution mainly of copper sulfate. This solution is tapped off from the bottom of the pad and placed into big tanks with steel plates an electrical current is passed from the tank to the steel which is then electroplated with pure copper. This process as the advantage of avoiding the smelting and refining stages required for sulfide ores.

dissolved air flotation - an overview | sciencedirect topics

dissolved air flotation - an overview | sciencedirect topics

DAF is particularly suited to water supplies with characteristics such as those containing low density particles and those yielding low density particles following chemical coagulation. Supply types include those containing 1) natural color, 2) algae, 3) low turbidity (< 10 NTU) and low TOC, and 4) moderate mineral turbidities (1050 NTU). Those containing color or algae are obvious choices. High quality supplies with low turbidity and TOC are good candidates for DAF treatment compared to direct filtration treatment because DAF clarification provides an additional process for particle and pathogen removal, and compared to sedimentation because high quality source waters produce low density floc difficult to settle without extensive flocculation and often the use of floc aid polymers. DAF performance is also less affected by low temperature compared to sedimentation so supplies in regions that have cold waters are also a good choice for DAF. To generalize, DAF is a good choice for reservoir supplies and should be considered for river supplies that fit the water quality types described.

DAF has several advantages. It is more efficient than settling in removing low density particles, even at the much higher hydraulic loadings compare 5 to 40 m h1 for DAF to 0.5 to 1 m h1 for conventional settling and 2 to 5 m h1 for high rate settling with plates and tubes The greater removal of particles (turbidity) by DAF versus settling means lower particle loading to filters is consequently found, which means granular media filters can be designed at higher rates or longer filter runs (higher water production) are achieved. It is common to find DAF effluent turbidities < 1 NTU and lower than 0.3 NTU when coagulation is optimum. The overall footprint for DAF plants can be small: smaller floc tanks because of shorter flocculation pretreatment time, smaller DAF tank areas compared to sedimentation tanks, and smaller filter area if designed at higher rates as explained above. DAF tanks that accumulate floated sludge and remove by scrapping with flight paddle or brush systems produce sludge with higher solids content (15%) than settled sludges, reducing sludge treatment.

In dissolved air flotation (DAF), bubbles are produced when the dissolution of air in water occurs under very high pressure. In this method, bubbles diameter typically ranges between 10 and 100m (Chen et al., 2011; Naghdi and Schenk, 2016). Some of the factors that influence the efficiency of this technique include bubbles size, saturator pressure, pH, hydraulic retention time, and recycle flow (Fuad et al., 2018). To promote aggregates formation and an increase in microalgal particles size (and thus improve the efficiency of the process), it is possible to use collectors (Pragya et al., 2013). This method is more effective than dispersed air flotation because the bubbles produced are smaller than those produced in dispersed air flotation. However, this method is more expensive, mainly because it requires pressurized air (Laamanen et al., 2016). Besson and Guiraud (2013) reported an efficiency of 90% when harvesting D. salina using DAF with sodium hydroxide (0.11M) as a surfactant. Zhang et al. (2014) harvested Chlorella zofingiensis using DAF and tested several surfactants at different concentrations. In this study, REs of 81%, 86%, 91%, and 87% were obtained when using as collectors chitosan (70mgg1), Al3+ (180mgg1), Fe3+ (250mgg1), and cetyltrimethylammonium bromide (CTAB, 500mgg1), respectively. Zhang et al. (2016) used DAF for 10min to harvest Nannochloropsis sp. and tested different concentrations of the surfactant magnesium, obtaining a flotation efficiency of 92% without extra addition of the surfactant, since this microalga was from marine water and presented a high concentration of this cation (1330mgL1) in the beginning of the experiment. The authors also harvested S. dimorphus, a freshwater microalga, that grows in a culture medium with low magnesium concentration (45.6mgL1), and obtained a flotation efficiency of 85%. Wiley et al. (2009), with the goal of comparing DAF and suspended air flotation, harvested a mixed culture (mainly composed by Chlorella and Scenedesmus) using DAF on batch mode, and reported an RE of 84.9% and an energy consumption of 0.76WhL1. Xia et al. (2017) used a combination of 40mgg1 of Al3+ as coagulant and 60mgg1 of CTAB as a collector to harvest Chlorella sp. XJ-445 through DAF. The experiment was carried out in batch mode for 15min with a gas flow rate of 50mLmin1, achieving an RE of 98.7%.

In addition to the conventional DAF process, there is a modified version of this process, called PosiDAF. In this process, bubbles produced are positively charged, due to the addition of chemicals in the saturator. The chemicals used in the saturator can be surfactants, coagulants, or polymers that have a hydrophobic and hydrophilic part, to promote the bonding between cells and bubbles (Fuad et al., 2018).

Dissolved air flotation operates on the principal of the transfer of floc to the surface of water through attachment of air bubbles to the floc. The floc accumulated on the surface, known as the float, is skimmed off as sludge (Section 7.19). The clarified water is removed from the bottom and is sometimes called the subnatant or floated water. Since rain, snow, wind, freezing could cause problems with the float, flotation tanks must be fully enclosed in a building; some users enclose the flocculation tanks as well. The process is particularly suited to treatment of eutrophic, stored lowland or otherwise algae laden waters and soft, low alkalinity upland coloured waters (Longhurst, 1987; Rees, 1979). Like all clarification processes flotation performance depends on the effectiveness of coagulation and flocculation. Polyelectrolyte dosing is often included to compensate for reduced performance at low water temperature or if the floc is fragile. Although the process has been successfully used for some directly abstracted waters other clarification methods tend to be more suitable for treatment of such waters especially when the turbidity consistently exceeds about 100 NTU (Gregory, 1999). Table 7.4 shows some typical results when treating algal laden waters.

There is, however, some experience with eutrophic waters with very high counts of algae where dissolved air flotation has not been successful, so that caution is necessary when choosing the process. It should be noted that sedimentation can achieve degrees of removal comparable to flotation, if algae are first inactivated by chlorination. This would however result in the formation of DPBs by the action of chlorine on algal metabolic products.

Flotation is preceded by a flocculation stage of the hydraulic or mechanical type usually dedicated to each flotation cell. The flocculation tank should have at least two compartments in series (Section 7.12). Flotation is normally carried out in rectangular tanks designed with surface loading rates between 812 m3/h.m2 but rates as low as 5 m3/h.m2 or as high as 1520 m3/h.m2 have been used on some plants (Pfeifer, 1997; Nickols, 1997). With such high rates there is a risk of air entrainment in the clarified water causing problems such as negative head due to air binding in downstream filtration processes (Section 8.2). This can be overcome by installing lamellas in the clarified water section as in DAFRapide. The use of the lamellas enhances the physical separation air bubbles (Edzwald, 2007). A similar effect can be achieved by minimising the high velocity that can cause bubble entrainment at the DAF outlet.

In flotation the solids loading can vary in the range 415 kg dry solids/h.m2. Typical tank depth is 23 m and the preferred length:width ratio is 1.332.5:1 with lengths up to 15 m using end-feed of air or 20 m with centre-feed of air. Width is limited to about 6 m for scarped tanks. The retention time in the flotation tank is between 1020 minutes. The velocity in the subnatent opening should not exceed 0.05 m/s. The flow over the clarified water discharge weir should be less than 100 m3/h per m of weir length.

For effective flotation the quantity of air required is about 610 g/m3 or 46 l/m3 of water treated and requires a recycle flow rate of about 615% (typically 810%) depending on temperature and dissolved oxygen concentration of the incoming water (Edzwald, 1992).. The recycle flow should be included in the flow used for the computation of the rates for the flotation unit and downstream filters. Recycle water should preferably be filtered water. Clarified water if used should be strained to prevent recycle nozzle blockage. Oil-free compressors are preferred but not essential for the air supply. Air is dissolved in recycle water under pressure either in pressure vessels equipped with an eductor on the inlet side for adding air or in a packed column; the operating pressures of the two respective saturator systems are 67 bar and 3.56 bar. In packed columns a packing depth of 0.8 to 1.2 m of 2537.5 mm Pall or Rashig rings of polypropylene (unsuitable for chlorinated water) or PVDF are used. The hydraulic loading rate of the air dissolving units lies in the range 5090 m3/h.m2. Saturator efficiency for packed column type is about 9095% whilst that for unpacked type is about 6575% (Amato, 1997). Saturator efficiency is 100 times the amount of air measured in the recycle water divided by the amount of air that could be dissolved theoretically. Air saturated water is returned to the flotation tank through a series of nozzles or needle valves to give a sudden reduction in pressure and release of air bubbles in a white water curtain. Typically bubble size ranges from 10 to 100 m with a mean diameter of 40 m (Zabel, 1984). The outlets are usually spaced at 0.30.6 m for needle valves and 0.1 to 0.3 m for nozzles (Dhalquist, 1997). A typical nozzle density is about 10 per m2 provided in 2 or 3 manifolds which could be isolated independently to facilitate greater turndown of recycle flow without loss of pressure. The contact time in the riser section should be about 100120 seconds.

In plants where there is a need for raw water ozonation and flotation, the two processes could be combined with air in the flotation process being replaced by an ozoneair or ozoneoxygen mixture (Boisdon, 1994).

High rate flotation processes are finding application as they require smaller footprint. These include proprietary designs DAFRapide (see above) AquaDAF (Plate 12(a)) and Clari-DAF. AquaDAF comprises a pre-fabricated perforated false floor with distribution of holes of different sizes across the floor, designed for uniform withdrawal of flow over the whole area of the tank which is said also to maintain a deeper bubble blanket throughout the float area. The holes are also thought to act as bubble collectors and allow for bubble coalescence preventing carry over to the filters. The combination of these effects is believed to provide performance comparable to conventional flotation (where clarified water is collected at one end and the bubble blanket is concentrated at the inlet end and grows shallower along the length of the tank) but at much higher rates. The surface loading rates are between 2550 m3/h.m2. The width of the tank is greater than the length in the ratio 1.52:1 and the depth is about 4 m. The other design parameters (such as flocculation requirements, bubble size, recycle ratio and air dose) are similar to conventional flotation. Clari-DAF tank geometry is similar to conventional flotation design, but deeper and clarified water is removed through a pipe lateral system located on the floor of the tank. The surface loading rates up to 50 m3/h.m2 are claimed.

Since the clarified water is taken from the bottom of the tank in the flotation process it could be combined with rapid gravity filtration in the same tank with the filtration section placed underneath (DAFF) e.g. Flofilter. Therefore the surface loading rates of the two processes need to be the same and should include the recycle flow. COCO DAFF (counter-current dissolved air flotation filtration) is an innovative combined flotationfiltration design in which air and water flow counter-current as against co-current in the conventional dissolved air flotation process (Fig. 7.6 and Plate 12(b)). Air is introduced with recycle water across the total tank sectional area below the flotation zone and therefore only the filter surface loading rate should include the recycle flow. COCO DAFF gives more efficient particlebubble interaction, and therefore increase in turbidity during desludging is minimized. (Officer, 2001) The process combines flotation and gravity filtration in one tank and uses a group of flocculation tanks common to all of the flotation cells. Flocculation is usually hydraulic and continues within the bubble blanket. Since the recycle flow is dissipated into the clarified water and not to the flocculated water as in conventional DAF, floc damage is minimized. The process requires far fewer recycle nozzles.

The flotation process is suitable for stop/start operation and has a flow turndown of about 2:1 or greater depending on the design of aeration manifolds. The former is one of its advantages when dealing with a water subject to high algal loadings; a plant can be switched in as and when needed and will give a steady quality treated water within 45 minutes (Rees, 1979). Apart from the drawbacks common to all high rate clarifiers, the flotation process has high energy requirements (about 0.050.075 kWh/m3 of water treated).

In the dissolved-air flotation system, a liquid stream saturated with pressurized air is added to the dissolved-air flotation unit where it is mixed with the incoming feed. As the pressure returns to the atmosphere, the dissolved air comes out of the liquid forming fine bubbles bringing fine particles with them. These rise to the surface and are then removed by a skimmer.

The production of fine air bubbles in the dissolved-air flotation process is based on the higher solubility of air in water as pressure increases. Saturation at pressures higher than atmospheric and flotation under atmospheric conditions was examined and used for algae separation [59]. It was suggested that algae separation by dissolved-air flotation should be operated in conjunction with chemical flocculation [25,60]. The clarified effluent quality depends on operational parameters such as recycling rate, air tank pressure, hydraulic retention time, and particle floating rate [25,59], while slurry concentration depends on the skimmer speed and its overboard above the water surface [19].

Algae pond effluent containing a wide range of algae species may be clarified successfully by dissolved-air flotation achieving thickened slurry up to 6%. The solids concentration of harvested slurry could be further increased by a downstream second-stage flotation [18,19,25,61]. High reliability of dissolved-air flotation algae separation can be achieved after optimal operating parameters have been ascertained. Autoflotation of algae by photosynthetically produced dissolved oxygen (DO) following flocculation with alum or C-31 polymer was examined [62]. Algae removal of 80%90% along with skimmed algal concentrations averaging >6% solids was achieved at liquid overflow rates of up to 2m/h. It was reported that the autoflotation was subject to dissolved oxygen concentration. No autoflotation was observed below 16mg DO/L.

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 effects of powdered activated carbon on the performance of a dissolved air flotation unit were investigated [4]. Refinery wastewater of different pollutant concentrations was treated and the effects of different operating parameters on the removal efficiency of pollutants in terms of biological oxygen demand (BOD) and COD were studied.

It was found that for doses of activated carbon in the range of 50150mg L1, the removal efficiencies for BOD increased from 2770% to 7694%, while those for COD increased from 1664% to 7292.5% for inlet values of 4595mg L1 and 110200mg L1 for BOD and COD, respectively [4].

Heavier than water particles can also be made to float. Dissolved air flotation, the most common approach, works by attaching small bubbles of air to suspended solids. The bubbles are generated by saturating a recycled stream of water with air under pressure, then releasing the pressure rapidly to produce clouds of microbubbles. Attaching the bubbles to the solids requires a reduction in charge of the particles and the production of hydrophobic spots on the surface of the solids via chemical/physical pretreatment.

A first-principles design would involve predicting how much air per kg of water could be dissolved in water at a given temperature and pressure using Henrys law, then working out the required recycle water flowrate based upon an amount of air per incoming solids load. It is this second factor which cannot be determined from first principles. It can be measured experimentally, or more commonly estimated based on experience.

Dissolved air is the most common type of flotation gas used in potable water treatment. The dissolved air flotation (DAF) process mixes a clarified stream from the outlet of the unit with air at 39bar, to produce a supersaturated (compared with saturation at atmospheric pressure) solution of air in water. This is rapidly depressurized at the inlet of the unit to produce a mass of microbubbles which attach to the solids present, floating them to the surface.

Its 15-min HRT and 15m/h surface loading to give 95% solids removal makes DAF a compact alternative to settlement tanks for drinking water treatment, favored by many designers since the 1980s for treatment of upland waters and more recently for algae removal. Generally the higher the feed solids, the higher the %removal as the effluent quality is substantially constant.

More recently, high-intensity DAF has been used to pretreat seawater to protect against algal blooms prior to its desalination for drinking or industrial water. In this application, surface loadings of 50m/h are common as the algal cells have densities close to or lower than seawater.

An additional advantage of DAF is that it yields sludge at maybe 5% dry solids content as opposed to approximately 1% dry solids from a settlement tank, which means that sludge pumping and dewatering are cheaper. 5% sludge may however be hard to remove and operators may feel they need hosepipes running to wash it away! Consequently, for new build it is sometimes thought better to use hydraulic desludging (also known as flooding) and use a separate thickener, which also overcomes some problems with level control. It does however remove a key benefit of DAF.

Designers should bear in mind that it is important to mix the DAF sludge well with any thinner sludges which are being cotreated prior to subsequent sludge treatment. A suitable mixed buffer tank is recommended to avoid problems caused by variable sludge solids content on downstream processes and for degassing DAF sludge.

In low-rate drinking water treatment duties, around 9g air per m3 of recirculated water are required. A rough rule of thumb is that the air compressor should deliver a volumetric flow of air (measured at atmospheric conditions) equivalent to 25% of recycle water flow rate. The recycle water flow rate should be at least 10% of the unit throughput. An HRT of at least 15min is required, and a surface loading of 15m/h.

For high-rate DAF, sludge treatment, and industrial applications, required recycle rates may be several hundred percent of the throughput. Air solubility in water is temperature-dependent, which may also be a factor in selecting a DAF process, since sludges and industrial effluents may be warm, as may seawater used by those countries which desalinate for drinking water treatment.

Many methods have been reported for the dewatering and harvesting of microalgae, such as, filtration, centrifugation, dissolved air flotation, electro floatation and flocculation (bio flocculation, chemical flocculation and pH induced flocculation) (Milledge and Heaven 2013), but no dewatering technology has actually been tested for non-destructive milking. Therefore, for the milking process, the dewatering and harvesting technologies were screened based on the biological characteristics of B. braunii (forming large colonies and self-floating ability). The following parameters were used as the screening criteria for selecting the potentially suitable methods for non-destructive dewatering and harvesting.

Centrifugation is considered as the more effective dewatering method for most of the microalgal species than filtration due to very low cell size causing difficulties for filtration (Molina Grima et al. 2003). However, for microalgal species with larger cell size, filtration is a more suitable method (Milledge and Heaven 2013). Due to the colonial structure of B. braunii, filtration may be the more suitable option for B. braunii dewatering used in milking process. However, filtration technologies involving high pressure or vacuum may destruct the algal colonies.

Due to low operational cost, flocculation is considered as one of the most accepted methods for conventional microalgae dewatering. However, this method may not be a suitable for B. braunii dewatering when the aim is not to damage the cells. Flocculation involves the external chemicals (chemical flocculation) or the bio-organisms (bio flocculation) to induce the flocculation (Milledge and Heaven 2013). The contact of chemicals with the B. braunii will harm the cell viability and no or incomplete separation of these chemicals from the culture after harvesting will lead to accumulation of them in the recycled media. Similarly, bio flocculation will lead to bio contamination in the culture which is not favourable for the process. Adjustment of pH of the culture also causes flocculation (pH induced flocculation), however, it is an unreliable method and can cause algal death (Milledge and Heaven 2013). Moreover, as the average particle size of B. braunii (due to colonies) is higher than other algal species, flocculation is not necessarily required.

Dissolved air flotation is also usually combined with chemical flocculation and is one of the most energy consuming methods of harvesting, so, it has not been considered here. Again, the ability of at least some strains of B. braunii in self-floating eliminates the requirement of the air flotation step. Also, the technologies reported in the literature with very high energy consumption such as, decanter bowl centrifuge (energy consumption of 8 KWh/m3) or poor reliability such as, hydro cyclones (Molina Grima et al. 2003) are not considered in this analysis. The technologies considered for dewatering of B. braunii for milking process in this study, the input assumptions of final concentration achieved after harvesting, and the energy consumption for these technologies are shown in Table1.

different types of flotation cells

different types of flotation cells

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

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

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

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

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

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

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

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

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

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

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

flotation cell - an overview | sciencedirect topics

flotation cell - an overview | sciencedirect topics

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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].

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