Solid residue from the second filtration stage is repulped with reclaim water to 35% solids, and is pumped to our EXAMPLE Conditioning and Argentite Flotation Circuit which is a process for the recovery of thisSilver Mineral.
A large portion of the silver in the ore will not have been leached in the initial grinding and leaching circuits. It is therefore the intent of this circuit to recover as much silver mineral (Argentite) as possible by producing a high grade silver flotation concentrate and delivering it to the Argentite Leach Circuit for further leaching.
Froth flotation separates different minerals by using the surface characteristics of the different minerals. The degree of preferential non-wetting or wetting of a mineral surface by water allows easily wetted minerals to separate from non-wettable minerals. Air bubbles, created by introducing air into a bubble diffusing mechanism, attach themselves by surface tension to the non-wettable minerals. The strength of the attachment is regulated by the mineral surface characteristics and can be enhanced or hindered by the use of chemical reagents.
For effective silver recovery in the flotation circuit, the flotation pulp must be thoroughly conditioned prior to entering the flotation cells. Two large flotation conditioning tanks provide this initial flotation pretreatment. The first conditioning stage is used to reduce the residual cyanide concentration in the pulp to less than 5 parts per million (ppm)total cyanide. This cyanide destruction process consists of oxidation of cyanide using a mixture of air and sulphur dioxide. The overall reaction being as follows:
The second conditioning stage is used to condition the flotation pulp with flotation reagents. The flotation reagents used in this circuit consist of: potassium amyl xanthate (PAX),as the collector; Aerofloat 208 (A-208), as the promoter; methyl isobutyl carbinol (MIBC), as the frother; lime as the pH modifier, copper sulphate as the activator; and sodium silicate as the slimes dispersant.
The Argentite Flotation Circuit is designed to provide one rougher stage and two cleaner stages of flotation. The rougher flotation circuit is designed to recovery the maximum possible amount of silver (and remaining gold) in the secondary filter residue and the cleaner flotation circuit is designed to maximize the silver grade of the final flotation concentrate.
Repulped slurry from the filtration circuit is pumped from the conditioning pump tank to the first conditioning tank by the 5 HP 4x3 conditioning feed pump at a rate of 48.7 cubic meters per hour and 35% solids.
The first conditioning tank is 4 meters diam x 4 meters and is equipped with a 50 HP agitator. This tank is primarily used to destroy cyanide according to the reaction given in the introduction of this section. Air is sparged into the bottom of the tank and the dual blade agitator dissolves oxygen into solution. Copper sulphate is added to the tank to assist the cyanide destruction reaction and sodium metabisulphite is added to provide the sulphur dioxide required to destroy the cyanide. Lime is added to this tank to maintain the protective alkalinity of the slurry, consequently neutralizing any acid generated from the destruction process. A hydrogen cyanide detector probe located in the vicinity of the first conditioning tank, will alarm at the mill control panel HCN controller if the level of toxic HCN gas near the tank reaches 30 ppm. An exhaust fan vents any gasses from the closed tank to the outside of the mill building. The size of the tank provides one hour retention time for the destruction reaction to occur. Slurry exits the first tank through a 6 riser pipe and flows into the second conditioning tank. The process air flow to both conditioning tanks is monitored by individual rotameters, and a single pressure gauge.
Silver Mineral Flotation reagents are added to the second conditioning tank: MIBC, A-208, and Xanthate, monitored by individualrotameters. Lime can be added to this tankto ensure a pH level of 8.0 respectively for flotation purposes. Sodium silicate can be added for a slime dispersant. An overload condition on either of the first or second conditioning tankagitators is alarmed at the mill control panelannunciator. The 50 HP agitator in the secondconditioning tank is two speed and requires gland seal water.This tank can be used for cyanide destruction purposes if necessary; consequently, it is equipped with air sparge lines, copper sulphate, and sodium metabisulphite lines identical to the first tank. Under normal conditions the agitator will be run on the lower speed, and during cyanide destruction it will be run at the higher speed. The discharge from the second conditioning tank is pumped by one of two 7.5 HP 4x3 flotation feed pumps to the rougher flotation cells. In order to prevent the second tank being pumped dry and the agitator from being exposed, this tank is maintained at a high slurry level with the conditioned slurry overflowing into a separate pumpdischarge compartment. A high level switch will alarm at the mill control panel annunciator on excessive overflow or high slurry level. The pH of the second tank is monitored and is calibrated to alarm at themill control panel on a pH of 8 or less. The instantaneous pHvalue is also monitored by a recorder located in the mill control panel. Any gases generated in the secondconditioning tank are vented outside by an exhaust fan.
The flotation feed pumps deliver the conditioned slurry through a PSA sampler to the feed box of the first bank of rougher flotation cells. A cyanide detector probe monitors the total cyanide concentration in the slurry, which should be approximately 2 ppm. A 5 10 ppm cyanideconcentration will alarm at the mill control panel annunciator.
The rougher flotation cells consist of two banks of four 100 cubic foot cells each. Each bank of rougher cells is equipped with two 7.5 HP motors and each motor drives two flotation mechanisms. The rougher cells are equipped with a feed box; prior to the first cell, a junction box; between the two banks of cells, and a discharge box; at the end of the last rougher cell.
The junction and discharge boxes are equipped with manually operated plug valves and are provided to control the overall pulp level in the rougher flotation cells. Reagents can be added to the feed box and/or the junction box, and feed rates can be controlled and monitored by rotameters.
Conditioned pulp is fed to the feed box of the rougher flotation cells and is drawn into the bottom of the first rougher cell by the flotation mechanism impeller. Process air from the flotation blower is delivered down the mechanism shaft to the impeller and is combined with the pulp. The air rate to each bank of rougher cells is adjustable and is monitored with rotameters. The impeller thoroughly mixes the air and pulp together and then expels this mixture against the diffuser blades which further mixes the air and the pulp. The aerated pulp is distributed over the bottom of the cell where the conditioned sulphide minerals attach to the air bubbles and are transported to the surface of the flotation cell; forming a stable froth layer on the surface of the pulp. The tailings material from this first rougher cell overflows a manually adjustable tailings weir and is drawn into the second rougher cell and the flotation process of mixing the air and the pulp is repeated. Reagent addition at the feed box and junction box of the rougher cells is provided to allow maximum operating flexibility. By the time the flotation pulp reaches the last rougher flotation cell, the majority of the silver and other sulphide minerals will have been recovered in the froth of each flotation cell. The froth from each bank of cells overflows into a common froth launder and flows by gravity to the 2 HP 2 vertical rougher concentrate pump. Water is added at the froth launder to aid in transporting the concentrate.
Silver Mineral Rougher concentrate is pumped to the feed box of the first stage of cleaner flotation cells at a rate of 2.6 cubic meters per hour and 25% solids. The first stage of cleaning consists of 3-30 cubic foot cells. The flotation process of the cleaner cells is identical to that of the rougher cells and likewise, flotation reagents can be added to the feed box of the first cleaner cells. The air rate to the cleaner cells is also adjustable and is monitored with rotometer. The froth from the first cleaner cells is swept off the pulp surface of each cell into the common froth launder with a rotating paddle. The first cleaner concentrate flows, with the help of spray water in the froth launder, directly into the single 30 cubic foot second stage cleaner cell. The concentrate from the second and final cleaning stage is collected and flows by gravity to the 2 HP 2 vertical Argentite leach feed pump.
The tailings from the second cleaner cell overflow into the first cell of the first cleaner stage and the tailings from the first cleaner stage are recirculated back to the feed of the rougher stage with the 2 HP 2 vertical cleaner tails pump.
The tailings from the rougher flotation stage flow by gravity from the rougher cells discharge box through the tailings sampler and into the mill tails box. The tailings from the Argentite Flotation Circuit are the main tailings from the processing plant.
The Froth Flotation Process is about taking advantage of the natural hydrophobicity of liberated (well ground) minerals/metals and making/playing on making them hydrophobic (water-repel) individually to carefully separate them from one another and the slurry they are in. For this purpose we use chemicals/reagents:
The froth flotation process was patented by E. L.Sulman, H. F. K. Pickard, and John Ballot in 1906, 19 years after the first cyanide process patents of MacArthur and the Forests. It was the result of the intelligent recognition of a remarkable phenomenon which occurred while they were experimenting with the Cattermole process. This was the beginning. When it became clear that froth flotation could save the extremely fine free mineral in the slime, with a higher recovery than even gravity concentration could make under the most favorable conditions, such as slime-free pulp, froth flotation forged ahead to revolutionize the nonferrous mining industry. The principles of froth flotation are a complex combination of the laws of surface chemistry, colloidal chemistry, crystallography, and physics, which even after 50 years are not clearly understood. Its results are obtained by specific chemical reagents and the control of chemical conditions. It not only concentrates given minerals but also separates minerals which previously were inseparable by gravity concentration.
This new process, flotation, whose basic principles were not understood in the early days, was given to metallurgists and mill men to operate. Their previous experience gave them little guidance for overcoming the serious difficulties which they encountered. Few of them knew organic chemistry. Those in charge of flotation rarely had flotation laboratories. Flotation research was done by cut and try and empirical methods. The mining industry had no well equipped research laboratories manned by scientific teams.
Froth flotation, as pointed out previously, was a part of the evolution of milling during the first quarter of the 20th centurya period during which the progress of milling was greater than in all of its previous history. It marks the passing of the stamp battery, after 400 years service to the mining industry, and the beginning of grinding with rod mills, ball mills, and tube mills without which neither the cyanide process nor the froth flotation process would have reached full realization. More than all of these, it was the time when custom and tradition were replaced by technical knowledge and technical control.
This volume, then, is dedicated to those men who, with limited means, made froth flotation what it is today. It is designed to record the impact of this great ore treatment development on the mining industry both present and future.
The single most important methodused for the recovery and upgrading ofsulfide ores, thats howG. J. Jameson described the froth flotation process in 1992. And its true: this process, used in several processing industries, is able to selectively separatehydrophobic fromhydrophilic materials,by taking advantage of the different categories of hydrophobicity that areincreased by using surfactants and wetting agents during the processalso applied to wastewater treatment or paper recycling.
The mining field wouldnt be the same without this innovation, considered one of the greatest technologies applied to the industry in the twentieth century. Its consequent development boosted the recovery of valuableminerals like copper, for instance. Our world, full of copper wires usedfor electrical conduction and electrical motors, wouldnt be the same without this innovative process.
During the froth flotation process, occurs the separation of several types ofsulfides,carbonatesandoxides,prior to further refinement.Phosphatesandcoalcan also be purified by flotation technology.
Flotation can be performed by different types of machines, in rectangular or cylindrical mechanically agitated cells or tanks, columns, aJameson Flotation Cellor deinking flotation machines. The mechanical cells are based in a large mixer and diffuser mechanism that can be found at the bottom of the mixing tank and introduces air, providing a mixing action.The flotation columnsuse airspargersto generate air at the bottom of a tall column, while introducing slurry above and generating a mixing action, as well.
Mechanical cells usually have a higher throughput rate, but end up producing lower quality material, while flotation columns work the other way around, with a lower throughput rate but higher quality material.The Jameson cell just combines the slurry with air in a downcomer: then, a high shear creates the turbulent conditions required for bubble particle contacting.
Advantages of froth flotation: first of all, almostallmineralscan be separatedbythis process. Then, the surface propertiescan be controlledandaltered by the flotationreagent. Finally, this technique is highly appropriate for the separation ofsulfideminerals.
To help towards an understanding of the reasons for the employment of specific types of reagents and of the methods of using them, an outline of the principal theoretical factors which govern their application may be of service. For a full discussion of the theory of flotation the various papers and text-books which deal with this aspect should be consulted.
The physical phenomena involved in the flotation of minerals, those, for example, of liquid and solid surface-tensions, interfacial tension, adsorption, flocculation, and deflocculation, are the manifestations or effects of the surface-energies possessed by all liquids and solids in varying degree. These, in turn, arise from the attractions which exist between the interior molecules of every substance and are responsible for their distinctive propertiesform, fluidity, cohesion, hardness, and so on. It follows, therefore, that every substance must exhibit some degree of surface-energy.
All the solids normally present in an ore i.e., metallic, non-metallic, and rock-forming mineralshave their particular contact-angle and hysteresis values and therefore tend to be wetted in varying degrees in accordance with such values. These differences, however, are not usually sufficient to allow of the effective separation of the mineral and gangue constituents from each other. It is the function of the flotation reagents employed to accentuate or magnify these differences to a degree which renders separation by flotation practicable. Some reagents (modifiers) are added with the object of decreasing the contact-angle and so increasing the degree of wetting of the unwanted particles, which are usually more prone to become wetted than the wanted minerals. Others (promoters) are added to increase the tendency toward non-wetting shown by the valuable minerals by coating them with a film of yet higher contact-angle value. Such films are said to be adsorbed in respect of the water.
In this connection reference to Fig. 28 will indicate that a reagent which decreases the surface-tension of water tends thereby to increase wetting of the solid, since, if the value of S1 and therefore of its horizontal component, is lessened, the water-edge, as at P, will tend to extend over the solid surface, making therewith a smaller contact-angle.
The reagents added to promote the separation of the wanted minerals by increasing the water/solid contact-angle consist of substances whose molecules or minute suspensions have a markedly lower attraction for water molecules than the latter exert between themselves. Finely divided oil emulsions in water, dissolved xanthates, and other promoters are typical of such reagents. Substances of such nature, when dissolved in or disseminated through water, are pre-eminently adsorbed, or thrust towards the water boundaries, where the intra-molecular attractions are less uniformly balanced. Normally, this would occur at the free or air/water surface. In a pulp, however, from which air surfaces are absent, but in which mineral particles are suspended, the same thing takes place at the water/solid boundaries, adsorption being most pronounced at those faces where the interfacial tension is greatest viz., those with the highest contact-angle value and lowest adhesion for water. The minute particles of oil or xanthate molecules are thus virtuallythrust into adherence with the more floatable solids, whose surfaces they therefore film, increasing the contact-angles to their own high values and so rendering the solid more floatable. Experimental work indicates that the film so formed is of the order of one molecule in thickness.
Adsorption can be both positive and negative. Substances whose molecules have less attraction for water than the water molecules have for each other are concentrated at the water boundaries as explained in the foregoing paragraph ; this is termed positive adsorption, but substances whose molecules have a greater attraction for water molecules than the latter have for each other will tend to be dragged away from the surface layers, at which their concentration thus becomes less than in the interior of the liquid ; this is negative adsorption. Substances that are negatively adsorbed are those which tend to form chemical compounds or definite hydrates with water, such as sulphuric acid. In froth flotation we are concerned more with positive than with negative adsorption.
In some cases a chemical reaction between the solid and the reagent occurs at the interface ; for instance, in the activation of sphalerite by copper sulphate a film of copper sulphide is deposited on the mineral following adsorption of the copper salt at its surface. In many cases there is no evidence of any chemical change, but, whether chemical action takes place or not, there is no doubt that the filming of the mineral is due primarily to the adsorption property of the liquid itself, by virtue of which the promoting reagent dissolved or suspended in it is concentrated at the interface.
The chemical action of flotation reagents has been and still is the subject of a great deal of research work, which is bringing the various theories into common agreement, but there are still too many doubtful points and unexplained phenomena to make a simple explanation possible in these pages.
The foregoing paragraphs can be summarized by stating that the reagents employed in froth flotation can be classified into three general groups, comprising frothers, promoters, and modifiers, respectively, the purposes of each class being as follows :
The operation of flotation is not always confined to the separation of the valuable constituents of an ore in a single concentrate from a gangue composed of rock-forming minerals. It often happens that two classes of floatable minerals are present, of which only one is required. The process of floating one class in preference to another is termed selective or preferential flotation , the former being perhaps the better term to use. When both classes of minerals are required in separate concentrates, the process by which first one and then the other is floated is often called differential flotation , but in modern practice the operation is described as two-stage selective flotation .
Selective flotation has, therefore, given rise to two other classes of reagents, each of which may be regarded as falling within one of the classes already mentioned. They are known as depressing and activating reagents.
The use of these reagents has been extended in recent years to three- stage selective flotation. For example, ores containing the sulphide minerals of lead, zinc, and iron, can be treated to yield three successive concentrates, wherein each class of minerals is recovered separately more or less uncontaminated by the others.
Although the flotation of the commoner ores, notably those containing copper and lead-zinc minerals, has become standardized to some extent, there is nevertheless considerable variation in the amount and nature of the reagents required for their treatment. For this reason the running costs of the flotation section of a plant are somewhat difficult to predict accurately without some test data as a basis, more especially as the cost of reagents is usually the largest item. Tables 32 and 33 can therefore only be regarded as approximations. Table 32 gives the cost of the straightforward treatment in air-lift machines of a simple ore such as one containing easily floated sulphide copper minerals, and Table 33 that of the two-stage selective flotation of a lead-zinc or similar complex ore.
From Table 32 it will be seen that the reagent charge is likely to be the largest item even in the flotation of an ore that is comparatively easy to treat, except in the case of a very small plant, when the labour charge may exceed it. At one time the power consumption in the flotation section was as expensive an item as that of the reagents, but the development of the modern types of air-lift and pneumatic machines has made great economies possible in expenditure under this heading. As a ruleCallow-Maclntosh machines require less power than those of the air-lift type to give the same results, while subaeration machines can seldom compete with either in the flotation of simple ores, although improvements in their design in recent years have resulted in considerable reductions in the power needed to drive them. It should be noted that the power costs given in the table include pumping the pulp a short distance to the flotation machines, as would be necessary in an installation built on a flat site, and the elevation of the rougher and scavenger concentrates as in circuits such as Nos. 9 and 10.
The power costs decrease with increasing tonnage because of the greater economy of larger units and the lower price of power when produced on a large scale. The cost in respect of reagents and supplies also decreases as the size of the plant increases, due to better control and organization and to lower first cost and freight rates of supplies when purchased in bulk. The great disadvantage of a small installation lies in the high labour cost. This, however, shows a rapid reduction with increase of tonnage up to 1,000 tons per day, the reason being that with modern methods a flotation section handling this tonnage requires few more operators than one designed for only 200 tons per day. For installations of greater capacity the decrease is comparatively slight, since the plant then generally consists of parallel 1,000-ton units, each one requiring the same operating force ; the reduction in the cost of labour through increase of tonnage is then due chiefly to the lower cost of supervision and better facilities for maintenance and repairs. Provided that the installation is of such a size as to assure reasonable economy of labour, research work and attention to the technical details of flotation are generally the most effective methods of reducing costs, since improved metallurgy is likely to result in a lower reagent consumption if not in decreased power requirements.
The costs given in Table 33 may be considered as applying to a plant built on a flat site for the two-stage selective flotation of a complex ore in subaeration machines with a tank for conditioning the pulp ahead of each stage and one cleaning operation for each rougher concentrate. It is evident that the reagent charge is by far the largest item of cost. This probably accounts for the more or less general use of machines of the mechanically agitated type for complex ores in spite of their higher power consumption and upkeep costs, since the high-speed conditioning action of the impellers and provision for the accurate regulation of each cell offer the possibility of keeping the reagent consumption at a minimum. As in the case of single-stage flotation, the charge for labour falls rapidly as the capacity of the plant increases to 1,000 tons per day ; beyond this point the rate of decrease of this and all other items of cost with increase of tonnage is less rapid. The remarks in the previous paragraph concerning the importance of research work and attention to technical details apply with added force, because of the possibility through improved metallurgy of reducing the much higher reagent and power costs which a complex ore of the class in question has to bear.
How hard a ball mill operator has to work depends partly on himself, and partly on the kind of muck the mine sends over to the mill. In some plants, the ore may change two or three times a shift, and a ball mill operator has to keep on his toes.
Thats why it would be just as well for you, as a ball mill operator, to study out a few ways of doing your job easier and better, because there will be times, even in the best of mills, when youll run into a lot of trouble. Collected here you will find some practical suggestions, contributed by a number of good mill men, that might give you an idea or two that would help get around some of that trouble.
To be sure we understand each other, lets begin with the equipment. In a simple grinding circuit there will be a ball mill and a classifier. Some circuits, especially in large mills, have more units or two or three stages of grinding, but whatever is said here will apply to the complicated circuits as well as the simple ones.
The two types of ball mill in general use are the grate mill and the open-end mill. Most manufacturers make both kinds; the difference between them is that the grate mill has a steel grid clear across the discharge end, but the open-end mill has only an open trunnion at the discharge end, through which pulp flows freely. If you dont already know all about the inside of your ball mill and what it is supposed to do, it would be a good idea to ask the shift boss, the metallurgist, or the superintendent to tell you about it.
Mechanical classifiers make use of rakes, spirals, or a simple drag belt. For our purposes it doesnt matter which type you are working with, because you would handle them all in pretty much the same way.
In operation, you add water to the ball mill along with the ore. Flowing out of the ball mill, the ground ore pulp pours into the classifier pool, where the finished material is separated from the coarse sand. You do that by adding a lot more water to the pool, so that the finer sand overflows the weir and goes on to the next step (flotation perhaps), and the coarse sand settles to the classifier bottom and is raked back into the mill to be ground finer. You, the operator, aresupposed to control these actions in order to send on to the machines below you the right amount of ore, ground just fine enough, and with just the right amount of water with it.
To help you do this, and to make a. record of how things are going, you will have to take samples of the pulp regularly. Different mills have different ideas on sampling, but all of them take at least hourly samples of the classifier overflow. What it amounts to is weighing a certain volume of the pulp to determine its density. Higher density means thicker pulp and usually coarser sand. Lower density means thinner pulp and finer sand. The shift boss will tell you what the density ought to be, and it will be up to you to hold it there.
You may also have to take density samples of the ball mill discharge, which runs a lot thicker than the classifier overflow, and some mills also expect you to take measured samples of the ball-mill feed and weigh them.
Another sample you may take is one for pH, which is a term that takes a little explaining. You can find out exactly what pH means from a chemistry book if you want to; but for all practical purposes, it is enough to know that pH is a number that tells you how much acid or alkali there is in the pulp. A pH of 7 is alkaline.
If you add acid, the pH goes down below 7; if you add an alkali like lime, the pH will go up, say to 9 or 10, depending on how much lime you add. In any case, you can bank on it that if the brass hats want you to watch the pH at all, they have good reasons for wanting you to hold it steady.
You may also have the job of adding balls to the mill each shift. The shift boss tells you how many or what weight, and you put them in. Drop them into the scoop if you have a grate mill, or put them in through the discharge trunnion if it is an open-end mill.
The controls you will have to work with are given in Table I, and are also indicated in the drawing. As to which one of these controls is most important, mill men dont all agree. Probably it depends on what kind of ore you are grinding. Most good operators, though, say that the classifier water valve should be the first one to adjust, because it controls directly the kind of finished material you send on down the line to the next man.
The most important point is this: You cannot adjust any one of these controls without paying some attention to the other two. For example, if you change the feed rate, you will probably have to reset the two water valves. They all work together. In fact, the whole grinding circuit acts like a team of horses, and as thetime at first.
In Table II you will find some suggestions on what to look for to help you decide how to use these controls. In the column headed if you find, there are set down the things youll run into if something is wrong with the circuit. That is, if the ball-mill feed gets finer than it usuallyis, the top line tells you what to expect and what to do. But dont think you have to do all these things all the time. Do only as much as you are sure you have to do.
The classifier overflow is really the most important spot in the circuit, because whatever comes over that weir is out of your hands, and your work will be judged by how good a product it is. Most operators believe that if there is any change in setting to be made, the density of the classifier overflow is where you make it first. Remember, more water to the classifier means thinner pulp and finer overflow; less water means thicker pulp and coarser overflow.
The matter of feed to the ball mill brings up a point that is important in keeping you out of trouble. You can find out by asking the old- timers how each kind of ore is going to act when it hits the mill, and if youfor each change as it comes along.
For example, suppose you are working in a lead-zinc flotation mill where there are two kinds of ore one that is coarse and low grade, and another that is finer and higher grade. Keep the feed to the ball mill lower when the coarse stuff comes along, because it takes longer to grind and you dont want to overload the mill. Then when the fine muck shows up, increase the feed and also run the classifier density higher. That will throw the high grade over into the flotation cells where it belongs.
You see, the high-grade mineral is heavier than the low grade, and it takes a little higher density in the classifier to lift it out. If you carry a low density, too much lead and zinc mineral keeps going back into the ball mill, and eventually may be ground into slime and lost altogether. Doing extralittle things like this is what marks a really good operator, and you can learn these things only by study and asking questions.
Keep ahead of trouble is good advice for flotation operators, and it is just as good for bail-mill men. A good operator can take care of even big changes in muck so smoothly and easily that if you were watching him, youd never know anything was running differently.
On the other hand, consider Joe Blow, the Wonder Boy. Thats him down there sitting on the rail near the ball mill, swinging his heels and probably wondering whatever happened to that little blonde hasher over at the Greasy Spoon. Suddenly Joe looks up. He has heard a splashing sound that doesnt belong there. The ball mill is strangely quiet. Joe looks at the feed box, and finds pulp pouring out on the floor.
Joe can tell right away what has happened. The mill has been overloaded and the grate has plugged. Quick as a flash, Joe races around and shuts the feed off, then whips open the valve pouring water into the mill. Hes fast; he wants action.
He gets it. The mill comes unstuck with a vengeance and belches sand into the classifier like a tidal wave. Joe, the dope, flushes water into the classifier, too, and you can almost hear it groan as the rakes get buried. The flotation man down below is tearing his hair and spinning valves. What he says about Joe blisters the paint on the concentrate launders, but Joe cant hear him. Joe is up under the mill shovellingcleaned up before the shifter comes.
Watching the mill discharge (2) will tell you what goes on inside the mill. Some operators note how high on the side of the discharge flange the wave of pulp is carried when the mill is running right. Then if the wave runs higher or lower than that, they know something is wrong.
If the mill is low on muck, (3) it rattles and bangs like a boiler factory, and a lot of good steel goes to waste. But if the mill is too full of muck, you can hardly hear it. Keep your ear peeled for the sound of the mill that you know is right.
Many operators feel the classifier overflow (6) by nibbling their fingers together with their hands in the stream, and with a little experience, you can tell pretty accurately whether or not the overflow is fine enough.
The amp-meter (4) is really as good a guide to the condition of the mill as the sound or the discharge. It tells you how much power the mill drive motor is drawing, but remember that if you overload the mill, or if you underload it, it draws less power.
You check on the circulating load (7) by watching the height of the sand on the rakes or spiral flights as they push it back to the mill feed launder. The shift boss will tell you about how high the sand ought to come.
What was wrong? He shouldnt have let the mill plug in the first place. But suppose it plugged anyway, he should have cut off the feed all right, but he should also have shut down the classifier, and increased the head water only a little. Then he should have cut down on the classifier water and then increased it, little by little, when the mill opened up. He should have done his best to keep things balanced instead of slamming everything out of adjustment at once. Well, hell learn. He will, or the boss will murder him some dark night.
Now, just because all these things to look for and to do have been put down in a table, dont think you ought to walk your shift carrying this operating manual in one hand and a density sample can in the other. It is no use trying to run a mill out of a cookbook. But what we did want to do was to set these things down here so you could think about them, and keep thinking about them, while you are working.
Just go at the problem the way things are arranged in the table. When something in the circuit begins to change, make sure you know exactly what is happening; then ask yourself what is causing it. Then, when you have answered that question, decide what to do about it. Think out each thing you do, and dont do things in a rush or without knowing why you are doing them. Dont be a Joe Blow, in other words.
One thing more, and a very important thing: When you do make a change, allow a little time, say 15 or 20 minutes, for the effect to show up before you make another change. Dont over-control. For instance, if the density in the classifier is up a little and you add more water, dont expect the density to change right away, and dont go back and open the valve even wider just because nothing seems to have happened. It will; just wait a while. A superoperator who cant let well enough alone gets on everybodys nerves.
In starting the grinding circuit after anything but a very short shut spitars enough to clear the samepacked on the tank bottom, start the classifier overflow pump, then start the classifier, and after that, the ball mill. But dont throw in all these switches at once. Youll get the electricians down on you if you do. Keep the water fed to the circuit down low until the load builds up a little; then set the valve hand wheels at about thepoint they should be for normal operation. You can check on this setting by marking one spot on the rim of the wheel and counting turns, or by counting exposed threads on the valve stem. Dont forget to lower the classifier rakes again as the load builds up.
In shutting down the mill, cut off the feed a few minutes before the shutdown is due. That will give time to grind out some of the circulating load and will make starting easier. Then when you are ready to stop, shut down the mill, then the classifier (raise the rakes), then the pump for the classifier overflow, if there is one.
If the power fails suddenly, shut off the water valves and raise the classifier rakes. And for goodness sake, dont forget to shut off any drip cans or siphon feeders of pine oil or other reagent you may have running somewhere in the circuit.
So far as mechanical trouble goes, there will probably be little of that if the equipment is reasonably good. Ball mills spring leaks from time to time because the bolts holding in the liner plates work loose. If a leak develops near the discharge end of the mill, shut down right away and fix it. This is especially true of an open-end mill. The point is that you dont want sand getting into theout in short order.
Now a word about safety, a subject that I am putting last because I want to leave it first in your minds. Whatever else you do, dont go poking aimlessly around the mill or the classifier, sticking your nose or your fingers in here and there to see how the machinery works. I wouldnt make that statement if I hadnt seen a man or two doing just that. Nor have I forgotten the time I was routed out of bed at two a.m. to help bandage a man whose right-hand fingers had just been taken off by the ball-mill scoop as effectively, though not as neatly, as a surgeon could have done it.He had been just poking around, too. Remember your company, and your country, need you on that ball-mill floor, and you wouldnt be happy holding down a hospital bed these days. So just be careful.
This Public DomainRobert Ramsey article is based in large part on experiences and opinions generously supplied by the following mill men: Clyde Simpson, Bagdad Copper Co., Hillside, Ariz.; E. J. Duggan and M. E. Kennedy, Climax Molybdenum Co., Climax, Colo.; John Palecek, Keystone Copper Corp., Copperopolis, Calif.; Frank M. McKinley, Bunker Hill & Sullivan M. & C. Co., Kellogg, Idaho; Malcolm Black, Wright-Hargreaves Mines, Ltd., Kirkland Lake, Ont.; and the concentrator staffs at Hudson Bay Mining & Smelting Co., Flin Flon, Manitoba, and Sherritt Gordon Mines, Ltd., Sherridon, Manitoba.https://archive.org/details/malozemoffmining00platrich
A hydrocyclone is a high-throughput gravity separation device used for separating slurry particles based on particle weight. For example, particles of similar size but different specific gravity, or particles of different size but identical specific gravity. Cyclones are also commonly used for dewatering of slurries given that they are much faster throughput than the alternatives, such as the traditional filter press. The operating principle of the hydrocyclone is that of a conical vortex-generating chamber. Liquid slurry enters at the top of the conical wall of the hydrocyclone through a vortex finder which creates tangential flow and therefore a strong vortex in the hydrocyclone. The slurry in the cyclone spins in a high velocity vortex. The fines capable of remaining in suspension exit out the top central pipe (the overflow), the coarse particles spiral down to fall through the underflow at the narrow vortex base, as shown in Fig. 12.12. Each particular cyclone is designed for a specific separation threshold: the cutoff between oversize and undersize (overweight and underweight) particles. In some cases total solids removal (slurry dewatering) is the objective.
Fig. 12.12. Hydrocyclone schematic. Slurry enters, at a high flow rate, at the feed point (1) as shown, into the vortex finder, and forms a powerful vortex inside the conical chamber. The coarse particles settle out and exit at the underflow (2) and the fines remain in suspension and exit out the overflow (3).
Hydrocyclones are very widely used for high-throughput coarse/fine particle separation in the mineral processing industry using a vortex effect, or simply slurry dewatering. They dont suffer from the rate-limiting steps of the alternatives:
A hydrocyclone can handle very high slurry throughput rates, and often an entire battery of hydrocyclones can be found at a mineral processing facility. The advantage of fast processing by a hydrocyclone is somewhat offset by the associated high impingement wear rate of the hydrocyclone internal surface lining by the fast-moving abrasive slurries. For this reason, wear-resistant linings of the pipes (inlet and overflow), vortex finder, cyclone chamber, and exit spigot (underflow) need to be lined with wear-resistant materials, usually alumina ceramics. Moreover, the impingement angle varies for different areas within the hydrocyclone. For practical servicing reasons, it is desirable that the wear rate is uniform, which means that ceramics of higher wear resistance might be used for high-impingement-angle areas such as the top wall near the vortex finder, and the underflow spigot.
Refurbishing of hydrocyclones represents one of the highest volume uses of alumina in the mineral processing industry. Ideally, it should be done as infrequently as possible due to the huge cost of production downtime in a mineral processing plant, which can run into the $millions per day range in large plants. Cyclone components of the cyclone assembly that can be manufactured in alumina ceramics include:
Linings of the curved interior wall are a large-volume ceramic requirement given the size of the vortex chamber of an industrial hydrocyclone. This can be either a single monolithic liner (small cyclone), or tiles. If tiles, ideally they should be curved to match the contours of the hydrocyclone interior. When flat tiles are used, this leaves a series of flats radially around the internal surface of the cyclone which interrupts material flow due to changing impingement angle and increases impingement wear on the highest impingement-angle areas of the tiled surfaces. Obviously the spigot and piping of the hydrocyclone also require ceramic linings.
Alumina is the most common material used for hydrocyclone linings. Some examples of this are shown in Figs. 12.13 and 12.14. In rare cases silicon carbide is used, however it is much more expensive than alumina, in the order of three times the cost, depending on the respective grades, for only a marginally improved service life. So silicon carbide is only justifiable in the cost-benefit equation in special cases. It is unwise to use ceramic-coated metals in a hydrocyclone. The wear-resistant coatings are soon gone and then very rapidly the fast-moving slurry erodes the unprotected metal wall of the hydrocyclone.
Fig. 12.14. Freshly refurbished hydrocyclone interior, small, and large examples shown. Note the smaller hydrocyclone on the left has a monolithic liner at the base (the region of highest wear). Wear resistance is significantly enhanced when joint lines are absent, as is the case with a monolithic liner.
In general, alumina gives an unbeatable combination of low price and high wear life for hydrocyclones and this represents one of the highest volume uses of alumina ceramics in the world, in tonnage terms, though not in dollar terms. In dollar terms, industrial wear-resistant alumina ceramics are at the bottom end of the alumina price scale with armor ceramics and refractories, where biomedical alumina is at the top end, and electrical and specialty wear components are in the middle.
The alumina supplier offers a cyclone-refurbishment service. Frequently mine sites do not have the facilities or expertise for cyclone refurbishment. The cyclone is trucked to the cyclone-refurbishment workshop of the alumina supplier.
OEM engineering facilities offer a cyclone-refurbishment service. The hydrocyclone is trucked to their facility. They order in the alumina parts from an alumina supplier, and refurbish the hydrocyclone (Fig. 12.14).
Hydrocyclones can also be used for plastic film washing and separation. They serve the same purpose as float-sink tanks, where pieces of plastic films are separated from heavier contaminants, however, in a more efficient manner with a much higher separation effect, about 20 times earth's gravity.
A hydrocyclone utilizes a centrifugal force to separate pieces of plastic film according to size, shape, and gravity (specific weight). The plastic waste is fed into the hydrocyclone in a suspension. Pressure jets excel the water mixture of films and contaminants within a cylindrical apparatus. The cyclone generated pushes the lighter plastic films outward and upward, while heavier contaminants move inward and downward to the bottom of the hydrocyclone. The hydrocyclone's sensitivity and selectivity can be adjusted by choosing the nozzle sizes of the exiting outlets at both the top and the bottom.
Heavier plastic components can be separated from the polyolefin, which is the desired material of a film washing line. A further advantage of the hydrocyclone separation step is the high amount of water present in the water circuit, ensuring together with the revolving forces arising due to the hydrocyclone a very good washing result of the films. Deposits of organic substances, a frequent feature of film from household waste, are easily removed by washing. In contrast, films from supermarkets often have a high percentage of paper in the form of affixed labels. It is a real challenge to separate this paper from the film as these linear low-density polyethylene (LLDPE) films from supermarkets are ideal as feeding material for recyclate used for the production of new film.
DE2900666 A1 (1980, BAHR ALBERT) discloses a method for the separation of mixed plastic waste composed of polymers of different densities (e.g., a mixture of granulates of polyethylene, polystyrene, and films of plasticized PVC) by reducing their size to a maximum 20mm (preferably 5mm long), washing, and suspending the polymers in a liquid. Successive hydrocyclones are arranged in stages, with upper cylindrical parts through which the suspended materials are dropped; the ratio of the diameter of the overflow/underflow outlets is 14 (preferably 12.5) and the ratio of the length/diameter of the top cylindrical part is 110 but increases with decrease in difference in density between the polymers and liquid. If the polymer's density exceeds that of the liquid, and if they are mainly plastic films, the angle of the cyclone cone is 120180 degrees; if the density of the polymers is only partly above that of the liquid, this angle is 540 degrees.
According to EP0791396 A2 (1997, DEUTZ AG), a disadvantage of the above method is the unsatisfactory separation effect, which leads to loss of solid material. Also, the required separation of the liquid from the solid matter fraction suspensions must be implemented downstream. These disadvantages are largely avoided in the separating centrifuge disclosed in EP0553793 A2 (1993, KLOECKNER HUMBOLDT DEUTZ AG), in which the separation ensues in the generated centrifugal field of a rotating container.
EP0791396 A2 (1997, DEUTZ AG) discloses a method for the separation of mixed plastic waste containing heavy contaminants, e.g., sand or metal residues, according to their density, wherein some of the heavy impurities are separated off in a hydrocyclone (15), and the plastic mixture (e.g., polyethylene and PVC films) is separated from the rotating suspension in a following sorting centrifuge (20). The method apparatus, shown in Fig.8.18, comprises a mixing vessel (10) for combining the solids (12) and a separating liquid (11), connected via a pump (24) to a hydrocyclone (15) and a sorting centrifuge (20) whose feed (14) is connected to the overflow (16) from the hydrocyclone. Wear on components of the separating centrifuge caused by these heavy materials can be avoided, and in some applications, a second separating stage at a higher separation density can be eliminated.
Figure8.18. Flow diagram with direct successive connection of the hydrocyclone and the separating centrifuge (1997, DE19606415 A1, DEUTZ AG). 10, Mixing vessel; 11, Separating liquid; 12, Solid materials; 14, Inlet; 15, Hydrocyclone; 16, Overflow; 17, Underflow,; 18, Heavy goods; 19, Light goods; 20, Sorting centrifuge; 24, Conveyor pump; 26, Mixing apparatus; 27, Sluice; and 28, Line.
Herbold Meckesheim GmbH assists its customers retrofit their film washing lines by introducing a hydrocyclone separation step in place of the common separation tank . The new step improves the quality of film flakes extrusion. Herbold Meckesheim GmbH has installed a model line for Rodepa Plastics B.V. in the Netherlands that was launched at the beginning of 2018. High-quality granulate for film thicknesses lower than 30m is produced from a mixture of plastic waste. This mixture consists of commercial films and LDPE film waste from sorting postconsumer packaging waste as is the case with automatic waste sorting plants . This plant can cope with highly contaminated films as well as with very thin gauge films. The wet shredder integrated into the washing plant and the hydrocyclone separation technology are the outstanding construction features of Herbold Meckesheim GmbH (see Fig.8.19). To separate the contamination from the film, in the washing line as early as the presize reduction step, a wet shredder especially designed for this purpose is used with Rodepa Plastics B.V. The feeding materials consist of a mixture of different plastics.
Hydrocyclones are cono-cylindrical in shape, with a tangential feed inlet into the cylindrical section and an outlet at each axis. The outlet at the cylindrical section is called the vortex finder and extends into the cyclone to reduce short-circuit flow directly from the inlet. At the conical end is the second outlet, the spigot. For size separation, both outlets are generally open to the atmosphere. Hydrocyclones are generally operated vertically with the spigot at the lower end, hence the coarse product is called the underflow and the fine product, leaving the vortex finder, the overflow. Figure 1 schematically shows the principal flow and design features of a typical hydrocyclone: the two vortices, the tangential feed inlet and the axial outlets. Except for the immediate region of the tangential inlet, the fluid motion within the cyclone has radial symmetry. If one or both of the outlets are open to the atmosphere, a low pressure zone causes a gas core along the vertical axis, inside the inner vortex.
The operating principle is simple: the fluid, carrying the suspended particles, enters the cyclone tangentially, spirals downward and produces a centrifugal field in free vortex flow. Larger particles move through the fluid to the outside of the cyclone in a spiral motion, and exit through the spigot with a fraction of the liquid. Due to the limiting area of the spigot, an inner vortex, rotating in the same direction as the outer vortex but flowing upward, is established and leaves the cyclone through the vortex finder, carrying most of the liquid and finer particles with it. If the spigot capacity is exceeded, the air core is closed off and the spigot discharge changes from an umbrella-shaped spray to a rope and a loss of coarse material to the overflow.
The diameter of the cylindrical section is the major variable affecting the size of particle that can be separated, although the outlet diameters can be changed independently to alter the separation achieved. While early workers experimented with cyclones as small as 5mm diameter, commercial hydrocyclone diameters currently range from 10mm to 2.5m, with separating sizes for particles of density 2700kg m3 of 1.5300m, decreasing with increased particle density. Operating pressure drop ranges from 10 bar for small diameters to 0.5 bar for large units. To increase capacity, multiple small hydrocyclones may be manifolded from a single feed line.
The ASH was invented by Professor Jan Miller of the University of Utah, and was patented in the USA in 1981. It makes use of the centrifugal forces which arise when air is sparged through the walls of a hydrocyclone. The device consists of a cylinder with a porous wall enclosed in an external chamber (Figure 1). The feed slurry enters tangentially through a conventional hydrocyclone header at the top of the cyclone, to form an annular liquid layer on the inner surface of the porous wall. The slurry moves downwards through the cylinder with a strong swirling motion. Bubbles are generated at the surface of the porous wall and, because of the swirling motion, bubbles which are produced on the porous surface experience an inwardly directed centrifugal force which carries them away from the wall, to pass quickly through the annular layer, collecting floatable particles on the way, forming a froth layer in the core of the cyclone. The froth leaves through the vortex finder in the top of the cylinder, while the tailing particles whose density is greater than that of water move towards the wall and are discharged through an annular gap in the bottom of the vessel.
An important feature of the ASH is the froth pedestal in the base. This stabilizes the froth and prevents it from passing out in the tailings. The froth zone is forced to move upwards through the vortex finder, carrying the hydrophobic particles. The hydrophilic particles are carried out in the tailings slurry.
The performance of the ASH is dictated by the fluid motion in the swirl layer adjacent to the porous wall, which in turn is controlled by the kinetic energy in the inflowing slurry, and the physical dimensions of the header and the vertical cylinder.
The bubble contact time in the hydrocyclone is of the same order as that of the pulp residence time, around 10 s. There is a correspondingly high capacity per unit volume, which is of the order of 100600 tons day1 ft3 of cell volume (360021 500 tonne day1 m3) as against 12 tons day ft3 (3570 tonnes day1 m3) for mechanical cells and columns. To date, the cells are not very large, but the capacity is quite high. Thus an ASH of diameter 5cm and height 50cm has a capacity of 318 tpd of solids.
The feed enters at conventional hydrocyclone pressures of 525psi (35170kPa) and the air is supplied at a relatively high pressure of around 65psi (440kPa), which is necessary to force the air through the porous wall at the required flow rate.
An important parameter which limits the performance of flotation cells is the superficial velocity Jg, which is the volumetric flow rate of the flotation air divided by the cross-sectional area of the pulp normal to the direction of the air flow. A high Jg will lead to a high concentrate production rate, other things being equal. In conventional cells, the only force acting on the liquid in the froth is that of gravity. Because of the centrifugal field in the ASH, the drainage force on the liquid in the froth is enhanced, and high Jgs are possible. Thus, the typical air velocity in an ASH is around 1 standard L min1 cm2 of cylinder wall, which corresponds to a superficial velocity Jg of 17cm s1. This figure may be compared with typical values for flotation columns, which are of the order of 0.54cm s1, and mechanical cells where the figure is generally lower still around 1cm s1. The consequence is that the ratio of air-to-pulp flow rates can be very high, leading to high recoveries despite the short residence time. Reported values of the air-to-pulp ratio are as high as 16:1. In mechanical cells and flotation columns, the ratio is usually 1:1.
As far as contact between particles and bubbles is concerned, the ASH is clearly a very intensive flotation device. However, it is not so effective at handling the froth-phase requirements. Ideally, to obtain high grades, it is necessary to be able to apply clean washwater, which can drain through the froth and flush the gangue into the tailings stream, while leaving the hydrophobic material attached to the bubbles. For this to occur, the velocity at which water can drain through the froth under gravity must be greater than the superficial upward froth velocity in the core. Using published data, it is possible to calculate that the axial upward velocity of the froth core in an ASH is in the range 1801300cm s1. The diameter of flotation columns is fixed to allow for froth washing and the maximum working superficial air velocity Jg is about 4cm s1 far below the values attained in the ASH. Evidently it is not possible to design an ASH which can allow both intensive contact between bubbles and particles, and effective control of the froth to obtain high grades. Accordingly, the ASH is most effective in applications where grade is unimportant, and where high recovery is desired. It is not surprising that the first large scale applications have appeared in the paper industry, for the removal of toner particles from recycled paper.
This is one of the simplest types of solids separators. It is a high-efficiency separation device and can be used to effectively remove solids at high temperatures and pressures. It is economical because it has no moving parts and requires little maintenance.
The separation efficiency for solids is a strong function of the particle-size and temperature. Gross separation efficiencies near 80% are achievable for silica and temperatures above 300C, while in the same temperature range, gross separation efficiencies for denser zircon particles are greater than 99% .
Cross-flow filters behave in a way similar to that normally observed in crossflow filtration under ambient conditions: increased shear-rates and reduced fluid-viscosity result in an increased filtrate number. Cross-microfiltration has been applied to the separation of precipitated salts as solids, giving particle-separation efficiencies typically exceeding 99.9%. Goemans et al.  studied sodium nitrate separation from supercritical water. Under the conditions of the study, sodium nitrate was present as the molten salt and was capable of crossing the filter. Separation efficiencies were obtained that varied with temperature, since the solubility decreases as the temperature increases, ranging between 40% and 85%, for 400 C and 470C, respectively. These workers explained the separation mechanism as a consequence of a distinct permeability of the filtering medium towards the supercritical solution, as opposed to the molten salt, based on their clearly distinct viscosities. Therefore, it would be possible not only to filter precipitated salts merely as solids but also to filter those low-melting-point salts that are in a molten state.
Prior to hydrocyclone, the process-sorting steps are the same as the flotation method involving size reduction, cleaning, and sieving. Once completed, the waste is mixed with water in order to transport the solution. The water and waste mix is passed through a hydrocyclone chamber, which is more commonly used in modern recycling and separation facilities, and can be an effective material separation technique (Menad, 2016).
The mix of waste and water enters the chamber at the inlet and rotates into a vortex, which generates a centrifugal acceleration several times higher than gravity. As a result, lighter materials move toward the center, where there is an air interface and a vortex finder that draws off this material at the top of the unit. Heavier materials move toward and along the outer surface and run along a tapered sidewall; these particles take a downward path and finally drop out of the bottom as shown in Fig.10.8. Often a number of hydrocyclones are used in series to carry out multiple density separations of the WEEE polymer material. Each process stage of the hydrocyclone refines the purity of the material (Brandrup, 1996; Coates and Rahimifard, 2009).
The hydrocyclone separation process has similar issues to the flotation technique as both systems use density properties as a means to make a distinction. Separation capability is better than that of the flotation method. The system is able to identify heavier particles from lighter ones, and the transportation medium does not need to be altered by means of chemical modification. Processing the WEEE polymers multiple times will refine the waste until an acceptable level of purity is achieved. Up to a 99% separation efficiency can be achieved, although the technique will still have difficulty in separating two materials with the same density characteristics (Gent etal., 2009; Kikuchi etal., 2008).
Cleaners or hydrocyclones remove contaminants from pulp based on the density difference between the contaminant and water. These devices consist of conical or cylindrical-conical pressure vessel into which pulp is fed tangentially at the large diameter end (Figure 6). During passage through the cleaner the pulp develops a vortex flow pattern, similar to that of a cyclone. The flow rotates around the central axis as it passes away from the inlet and toward the apex, or underflow opening, along the inside of the cleaner wall. The rotational flow velocity accelerates as the diameter of the cone decreases. Near the apex end the small diameter opening prevents the discharge of most of the flow which instead rotates in an inner vortex at the core of the cleaner. The flow at the inner core flows away from the apex opening until it discharges through the vortex finder, located at the large diameter end in the center of the cleaner. The higher density material, having been concentrated at the wall of the cleaner due to centrifugal force, is discharged at the apex of the cone (Bliss, 1994, 1997).
Cleaners are classified as high, medium, or low density depending upon the density and size of the contaminants being removed. A high density cleaner, with diameter ranging from 15 to 50cm (620in) is used to remove tramp metal, paper clips, and staples and is usually positioned immediately following the pulper. As the cleaner diameter decreases, its efficiency in removing small sized contaminants increases. For practical and economic reasons, the 75-mm (3in) diameter cyclone is generally the smallest cleaner used in the paper industry.
Reverse cleaners and throughflow cleaners are designed to remove low density contaminants such as wax, polystyrene, and stickies. Reverse cleaners are so named because the accepts stream is collected at the cleaner apex while the rejects exit at the overflow. In the throughflow cleaner, accepts and rejects exit at the same end of the cleaner, with accepts near the cleaner wall separated from the rejects by a central tube near the core of the cleaner, as shown in Figure 7.
Continuous centrifuges used in the 1920s and 1930s to remove sand from pulp were discontinued after the development of hydrocyclones. The Gyroclean, developed at Centre Technique du Papier, Grenoble, France, consists of a cylinder that rotates at 12001500rpm (Bliss, 1997; Julien Saint Amand, 1998, 2002). The combination of relatively long residence time and high centrifugal force allows low density contaminants sufficient time to migrate to the core of the cleaner where they are rejected through the center vortex discharge.
A rod mill was integrated with a hydrocyclone and produced a grind at the rate of 250t/h. Samples taken simultaneously from the discharge of the rod mill and the overflow and underflow streams from the hydrocyclone gave the following results. Calculate the circulating ratio and circulating load.
Though the solidliquid hydrocyclone has been established for most of the 20th century, satisfactory liquidliquid separation performance did not arrive until the 1980s. The offshore oil industry had a need for compact, robust and reliable equipment for removing finely divided contaminant oil from water. This need was satisfied by a significantly different type of hydrocyclone, which of course had no moving parts.
After explaining this need more fully and comparing it with solidliquid cyclonic separation in mineral processing, the advantages that the hydrocyclone conferred over types of equipment installed earlier to meet the duty are given.
Separation performance assessment criteria are listed prior to discussing performance in terms of feed constitution, operator control and the energy required, i.e. the product of pressure drop and flowrate.
The environment for petroleum production sets some constraints for materials and this includes the problem of particulate erosion. Typical materials used are mentioned. Relative cost data for types of oil separation plant, both capital and recurrent, is outlined, though sources are sparse. Finally, some pointers to further development are described, as the oil industry looks to equipment installed on the sea bed or even at the bottom of the wellbore.
During steady operation the products from a hydrocyclone has a definite cut point. However due to variations in the feed slurry characteristics and changes in the hydrocyclone geometry, especially the diameter of the apex due to abrasion, the cut point changes during operation. It is necessary to hold the performance at the desired d50C value for down stream operations. The control strategy could be to monitor the deviation of the cut point. The alteration in cut point was obviously due to change in feed characteristics and additionally to changes in cyclone geometry due to abrasion. Working with a D6B Krebs hydrocyclone and quartz suspension of known particle size distribution Gupta and Eren  indicated that for a constant pressure differential, the relative effect on d50C was:
The logic of the control program adopted was to calculate the d50C value during a steady state condition using a mathematical model. When the cut-point was altered due to any change in the variables the computer sequentially searched for the offending variable and restored it to the original value. The restoration was done by iteration using perturbance technique. The advantage of the technique was to predict changes using the previous reading as the initial value. Thus a variable (Du, , Q or T) was chosen by the computer and the established model was considered as f(x) and the step changes in d50C calculated using the expression:
Fig. 18.25 shows the set up and instrumentation for automatic control of a Krebs D6B-hydrocyclones. The apex of the cyclone was fitted with a rubber sleeve which could be pneumatically squeezed to alter its diameter. The vortex finder was specially designed to travel up and down. The centrifugal pump was fitted with a frequency controller. The control strategy is illustrated in Fig. 18.26.
Laboratory trials suggested that when a hydrocyclone variable was subjected to a step change and the d50C value deviated, the operation of the hydrocyclone could be restored such that the d50C could be maintained to within 5% of the calculated value. The mathematical model used for calculations was derived to suit specific slurry conditions. The conclusion was that such techniques could be developed for automatic control of the cut point of hydrocyclones.
Environmental pollution in Nigeria presents an urgent need to assess wastewater treatment facilities in various industries. This article presents an assessment of dissolved air flotation (DAF) operation in a dairy industry. The industry was visited, wastewater treatment facilities were assessed (based only on efficacy to remove selected environmental health-related pollutants) and measurements of essential design and characterization parameters were taken. The study revealed that the averages of flow rate, biochemical oxygen demand at 5days (BOD5), chemical oxygen demand (COD), suspended solids (SS) and total solids (TS) of the influent wastewater into the plant (DAF) were 3.45L/s, 1652.37, 3304.67, 2333.82, and 4396.10mg/L compared to effluent quality of 560.37, 1127.33, 172.33, and 1866.67mg/L for BOD5, COD, SS, and TS, respectively. The pH of the wastewater is being adjusted by addition of lime before the effluent equalization tank and individual efficacies of the system were 66.09, 65.89, 65.89, 57.54, 8.68, and 94.49% for BOD5, COD, SS, TS, DS, and total nitrogen, respectively, with overall efficacy of 38.10%. It was concluded that failure (lower overall efficacy) of the system can be attributed to setting of lime in the oversized equalization tank (50m3 instead of 16.82m3 per 8h shift), the lack of application of standardized engineering code and practices (provision of underground tank in the process, lack of complete coagulation processes, coagulation and flocculation units), lack of adequate aeration unit and lack of reliable systems for automatically adjusting dosage of coagulant and flocculant. Although, DAF unit is the centerpiece of a DAF-based system design, there are several other supporting systems important to optimal DAF operation. These observations, coupled with the analysis in this report, demonstrate that the facilities necessary to minimize continuous environmental pollution are lacking. Pollution will become an increasing problem unless pollution preventing codes and standards are developed; incorporated into government regulations and the regulations are enforced.
Rubio, J.: Environmental applications of the flotation process. In: Castro, S.H., Vergara, F., Sanchez, M. (eds.) Effluent Treatment in the Mining Industry, pp. 335364. University of ConcepcinChile, Chile (1998)
Rubio, J., Capponi, F., Matiolo, E., Nunes, D., Guerrero, C.P., Berkowitz, G.: Advances in flotation of mineral fines. In: Proceedings of the XXII International Mineral Processing Congress, Cape-Town, frica do Sul, pp. 10021014 (2003)
Ross, C.C., Valentine, G.E., Smith, B.M., Pierce, J.P.: Recent advances and applications of dissolved air flotation for industrial pretreatment. Presented at the industrial water/wastewater program North Carolina AWWA/WEA conference, Greensboro, North Carolina (2003)
Dai, Z., Fornasiero, D., Ralston, J.: The attachment efficiency of particles to gas bubbles in flotation. In: Ralston, J., Miller, J., Rubio, J. (eds.) Flotation and Flocculation from Fundamentals to Applications, Hawaii (2003)
Solari, J.A., Gochin, R.J.: Fundamental aspects of microbubbles flotation. In: Ralston, J., Laskowski, J.S. (eds.) Colloid Chemistry in Mineral Processing. Development in Mineral Processing, vol. 12, pp. 395418. Elsevier, Amsterdam (1992)
Oke, I.A.: Development and performance-testing of an electrochemical process for selected industrial wastewaters. Unpublished PhD thesis (Civil Engineering Department), Obafemi Awolowo University, Ile-Ife (2007)
The pollutants that are removed by the treatment plant can be divided into two parts of a and b; a represents the turbidity removed and b represents microorganisms reduced specifically bacteria. Under normal condition or operation of wastewater treatment pollutants are expected to be removed or reduce to FEPA  standard limits. Let F a denote the concentration of pollutant in the raw water (F), O a denote the fraction of a removed from the raw wastewater but in the floc (O), and U a denote the fraction of a present in the treated wastewater (U). Then the pollutant removed from the raw wastewater can be expressed as follows:
Babatola, J.O., Oladepo, K.T., Lukman, S. et al. Failure Analysis of a Dissolved Air Flotation Treatment Plant in a Dairy Industry. J Fail. Anal. and Preven. 11, 110122 (2011). https://doi.org/10.1007/s11668-010-9430-z