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simple structure flotation machine for copper ore

copper ore processing plant with parameter and quotation | fote machinery

copper ore processing plant with parameter and quotation | fote machinery

Note: In most cases, the gravity separation can only be used as a coarse concentration method to coarsely concentrate low-grade copper ore by discarding a large number of tailings before the flotation process. It cannot be used as a copper separation method for obtaining the final copper concentrates.

Five common flotation methods: sulphidizing flotation, fatty acid flotation, amine flotation, emulsion flotation and chelating agentneutral oil flotation. Flotation is the main method of copper ores separation.

The ore pulp discharged from the ball mill is transported to the mixing tank where flotation agents for rough selection are added. After the full stirring and activation, send it to the flotation machine for flotation.

During the flotation process, collectors, modifiers, and foaming agents are added. Useful minerals are enriched in the foam, which is scraped out by the scraper of the flotation machine and finally sent to the dehydration workshop for dehydration.

After coarse and fine crushing, transport the mined ores to the ball mill before the next stepclassification. Send them to the spiral classifier for classification and cleaning according to the weight of their particles, and finally transport them to the magnetic separator for rough and fine selection.

The copper ore of a copper mine in Zimbabwe is mainly copper-bearing pyrrhotite and high-copper skarn ore. The copper grade of the original ore is 1.25%, and its iron grade is 30.73%. The copper minerals in the ore are mainly chalcopyrite, followed by porphyry, and a small amount of chalcocite, covellite, malachite, etc.

Applicable copper ores: The ore that its copper is difficult to float in the form of chrysocolla, cuprite and copper ore impregnated by iron hydroxide and manganese aluminosilicate or in the form of combined copper.

For the treatment of refractory copper mineral materials, the selection of chemical processing for refractory copper ores mainly depends on their phase composition, structure and surrounding rock.

A copper mine in South Africa adopts the process of leaching extraction electrodeposition. The copper ore of the plant includes oxidized ore such as malachite and chrysocolla as well as sulphide ore, including chalcocite and bornite.

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copper sulfide flotation

copper sulfide flotation

Copper, due to the present world demand and price, is of foremost interest to the mining industry. Many new properties are either in the process of being brought into production or are being given consideration. Copper minerals usually occur in low grade deposits and require concentration prior to smelting. The method and degree of concentration depends on smelter location and schedules, together with the nature of the ore deposit. Sulphide copper ores generally occur with pyrite, pyrrhotite, arsenopyrite and molybdenite, and with gold and silver. A complete copper-iron separation may not always be essential for the maximum economic recovery and often is tied to the distribution of the gold and silver values.

The above flowsheet is designed for the treatment by flotation of copper as chalcopyrite with gold and silver values. The ore, ranging from 60-65% silica, with pyrite, arsenopyrite, and calcite with 3 to 4% copper. This flowsheet, though simple, is adequate for tonnages of 100 to 500 tons or more per day, depending on the size of equipment selected. It can be readily expanded by duplicating units for increased tonnages. By minor circuit changes, it provides the flexibility to treat a range of ore conditions which are often encountered in any mining operation. Generally in these small plants the recovery of molybdenum is disregarded unless it is present in considerable amounts. Larger plants generally will incorporate a circuit for molybdenum recovery from the copper concentrate by flotation. Sub- A Flotation is standard for this service.

Crushing Section. The crushing section with two-stage reduction is suitable for smaller tonnages, depending on the ore characteristics. Three-stage reduction in either an open or closed circuit, with screens for the removal of fines can be employed where conditions warrant. The fines are removed by a grizzly or screen ahead of each reduction stage for higher efficiency and for reduced wear on crushing surfaces.

Feed control is essential to efficient grinding and helps reduce surges and fluctuations throughout the entire plant. The Ball Mill in closed circuit with a Spiral Classifier discharges the pulp at about 60% minus 200 mesh. The Ball Mill is equipped with a Spiral Screen on the discharge for removal of any tramp oversize, worn grinding balls, and wood chips from the circuit.

The pulp from the Conditioner is treated in a 10-cell Sub-A Flotation Machine and a 4-cell Sub-A Flotation machine. Sometimes conditioners are not provided; however, their use insures that reagents are thoroughly mixed into the pulp ahead of flotation. This gives a more uniform feed and effective use of reagents plus improved flotation conditions. The 10-cell Sub-A Flotation Machine is of the free-flow type. Weirs for the control of pulp level through the machine are provided at the fourth, eighth and tenth cells. This free-flow type provides ample volume for normal fluctuations in the feed rate without cell level adjustment. Sand relief ports help extend the long life of the molded rubber wearing parts.

The first eight cells produce a rougher concentrate while the last two cells act as scavengers. The concentrate or middling product from these two cells is returned by gravity back to the fifth cell. The rougher concentrate from the first eight cells is cleaned in two stages in the four-cell standard Sub-A Flotation Machine, of the cell-to-cell type. No pumps are needed for the return of these flotation products for cleaning. This feature in Sub-A Flotation Machines gives added flexibility by enabling the operator to change cleaning circuits readily, should conditions require. The tailings from the cleaner flotation section are pumped back to the ball mill for regrind. To control dilution a cone classifier is placed in this circuit with the coarse solids going to regrind and the overflow used as dilution in the mill and classifier. It is possible to eliminate this classification in some cases but control is less positive. A separate regrind section could be provided if the quantity of middling products were enough to make this section feasible.

The final cleaned flotation concentrate flows or is pumped to a Spiral Rake Thickener. A Adjustable Stroke Diaphragm Pump, mounted on the thickener superstructure, meters the thickened concentrate to the Disc Filter. The Thickeners are often used to store concentrates for filtration at fixed intervals. These units have heavy duty construction throughout, overload indicators and positive rake lifting features. The Diaphragm Pump is used for concentrate recirculation purposes during such periods.

Lime is added to the Ball Mill by a Cone Type Dry Reagent Feeder. Other reagents, such as cyanide, xanthate, and a frother are fed and controlled by No. 12A Wet Reagent Feeders to the classifier and to the conditioner ahead of flotation.

This flowsheet stresses simplicity without sacrifice of efficiency. The factors of flexibility are essential to meet changing ore and market conditions. The unit arrangement which can be expanded by sections for increased capacity is an important feature. The equipment indicated has been proven for long life and low maintenance, and to give superior results. The Sub-A Flotation Machines are designed for high capacity and with features of flexibility to handle fluctuating conditions with a minimum of operating attention. Low final tailings and high grade concentrate are assured through the selective action of the Sub-A in the roughing, cleaning, and recleaning circuits.

Large scale mining operations, of which the porphyry coppers are typical, must resort to concentration. This is necessary as the ores are generally low grade and require flotation to produce a concentrate acceptable to the smelters.

These large scale milling operations handling low grade ore must provide very careful planning in the design of their plant flowsheet and selection of equipment. Milling circuits must be as simple as possible and for large tonnages, as few as possible. It is for this reason grinding mills and flotation circuits arenow designed to handle these large tonnages at low cost.

Sub-A Flotation Machines are a basic part of large tonnage operations and their use assures maximum economic recovery. Particular emphasis has been placed on the design and operation of these machines for roughing, scavenging and cleaning. Mechanisms have been greatly simplified and molded rubber wearing parts are standard for maximum abrasion resistance.

Three stage crushing is illustrated in the flowsheet; however, it is possible and practical to eliminate the third stage by incorporating a rod mill in the grinding section. This is a very practical arrangement and often a necessity when handling wet, sticky ore. There is evidence that this combination of crushing and grinding results in lower costs for reducing large tonnages of ore to flotation size.

The flowsheet illustrates a typical grinding circuit with a rod mill in open circuit. Its discharge, usually all 14 mesh, goes to a classifier for removal of finished material. The classifier sands are ground in a ball mill in closed circuit with the same classifier. High speed rod milling with speeds up to 80% of critical has shown definite improvement in efficiency and grinding capacity. Proper selection of mill density and grinding charges are also factors of importance. Usually the rod mill is operated at lower density so it acts partially as its own classifier for retaining oversize for further size reduction.

Some conditioning of the pulp ahead of flotation is usually very beneficial and will result in more uniform and rapid flotation of a selective high grade concentrate. For this service the (patented) Super Agitator and Conditioner is standard. Reagents added at this point are thoroughly mixed and reacted with the pulp. Any tendency of the pulp to froth prematurely is readily overcome by the patented standpipe arrangement which also assures positive pulp circulation.

For large tonnage circuits normally encountered in many of the copper operations the open or free flow type Sub-A Super Rougher Flotation Machine is recommended. Intermediate cell weirs are eliminated and circulation of pulp through the impeller is fixed to provide the desired agitation and aeration for rougher flotation conditions. Machines are usually arranged with up to six cells being open or free-flow without intermediate weirs. Two or more machines are always provided in series. This allows adequate volume for absorbing surges and fluctuation in feed without cell adjustment. Mineral and middlings in the teeter or quiescent zone of the cell are gradually forced upward to the froth removal zone. Only the coarser material in the agitation zone passes through the impeller for further conditioning and bubble attachment.

In the flowsheet each circuit consists of 16 or 18 cells in 4 or 6 cell units. These Sub-A Super Rougher Flotation Free-Flow Machines are in series. All of the mechanisms are of the single impeller type and are completely supported from the superstructure to facilitate maintenance. All heavy hoods and castings are eliminated and the impeller-diffuser clearance is pre-set and accurately maintained throughout the long life of the heavy duty moulded rubber wearing parts. The last two cells are the super scavenger type giving veryintense agitation and aeration to float the last trace of recoverable mineral or middling for re-treatment.

Rougher flotation concentrates are cleaned in a standard Sub-A Flotation Machine with cell to cell pulp level control. This arrangement for upgrading concentrates is universal in its acceptance by the ore dressing industry. Two or more stages of cleaning in the same machine are accomplished without auxiliary pumps and ideal flotation conditions for producing high-grade concentrates are easily maintained.

Cleaner flotation tailings are returned to the head of the rougher flotation circuit for retreatment. In many milling circuits, particularly if coarse grinding is used, the cleaner tailings will contain middlings or mineral with attached particles of gangue. In these cases it is necessary to thicken or classify and regrind this fraction. Centrifugal classifiers are being very successfully applied for the classification step although they do take considerable power and require more maintenance than a thickener with its underflow going to a regrind circuit.

The flowsheet incorporates thickening for both the concentrates and tailings for water reclamation and tailings disposal purposes. A Adjustable Stroke Diaphragm Pump on the concentrate thickener assures absolute control of the volumes delivered to the Disc Filter. When the filter is down temporarily for bag changes the concentrates may be recirculated to the thickener by this same pulp.

Flexibility and simplicity are the two most important points to design into any large tonnage flotation operation. The arrangement shown is flexible and will permit addition of extra milling sections up to the limit of the designed capacity of the crushing plant. Sub-A Flotation Machines are designed specifically for high tonnage installations and have been proven for all types of applications. Rugged construction will give years of service at lowest possible cost. This flowsheet is readily adaptable for the treatment of other ores. Note particularly the location and use of Automatic Sampler.

Copper, one of our most important minerals, is found in many parts of the world. One of the major sources of Copper is the so-called porphyry ores such as the large deposits in the west and southwestern United States, Mexico, South America and Europe.

Porphyry ores, with copper occurring in the form of Chalcocite and Chalcopyrite are normally low in grade and the copper minerals must be concentrated before smelting. In this flowsheet using Sub-A Cells the emphasis is on maximum economic recoveryhigh concentrating efficiency together with a premium smelter feed with a low alumina and magnesia content in the flotation concentrate.

To obtain lowest tailings from this ore usually requires scavenging of rougher flotation tails. This is performed ideally by the Sub-A Super Rougher Flotation Machine which was specially developed for this duty. This machine has a double impeller and gives tremendous aeration. The flowsheet in this study is designed to get the maximum recovery from a large tonnage of porphyry copper ore.

The crushing section consists of three-stage ore reduction with either a grizzly or vibrating screen between each crushing stage. Removing fines before putting the ore through a crusher increases the efficiency of the crusher as it is then only working on material that must be reduced, and is not hampered by fines already reduced in size. Electromagnets and magnetic pulleys are used to remove tramp iron from the ore, the former to remove the iron near the surface and the magnetic pulley to remove the tramp iron close to the conveyor belt.

Porphyry copper ores usually are medium to medium hard and require grinding to about 65 mesh to economically liberate the copper minerals from the siliceous gangue. Sometimes a regrinding circuit is advantageous on the rougher concentrate and on the scavenger concentrate. This will liberate the mineral from the middling products and increase the recovery by putting those mineral particles into the concentrate. Rougher flotation may be accomplished at a relatively coarse grind and the subsequent regrind performed on a comparatively small tonnage.

Lime is usually added to the ball mill feed by a Dry Reagent Feeder. The frother and promoter are added in the classifier prior to flotation to realize the full effect of the reagent. Reagents can also be stage- added to the cells in the flotation circuit.

Standard Sub-A Flotation Machines are used for both the rougher and cleaner circuits, where their cell-to-cell principle gives both high recovery and a good grade of concentrate. The rougher concentration is accomplished in 6 or 8-cell flotation machines, with the concentrate from each goingto a separate bank for cleaning and re-cleaning. No. 30 Sub-A Flotation Machines are ideal for large tonnage operations, as each bank will handle from 1000 tons upward per day. Tails from the rougher circuit go to a scavenger circuit. Roughing, scavenging, cleaning and recleaning can be carried out in one bank of Sub-As. This is possible because of the distinctive gravity return of a product from any cell to any other cell of a bank without using pumps. In large installations, however, these steps are usually carried out in separate banks of cells. The scavenger flotation circuit consists of a 4-cell, Sub-A Super Rougher Flotation Machine with its super aeration. The concentrate from scavenger cells is returned to the head of the rougher cells and tails are sent to tailing pond. The new Sub-A Super Rougher Machine is designed especially to produce the lowest possible tailings in the mill circuit by scavenging off the last bit of recoverable and often difficult to float mineral. The Automatic Sampler is used on the flotation feed, concentrates and tailings to establish close mill control.

The flowsheet incorporates a thickener on the copper concentrates to thicken for optimum filtering. This also serves as a temporary storage space to accommodate operating requirements. The Adjustable-Stroke Diaphragm Pump on the thickener gives absolute control of volumes pumped to the filter. When the filter is shut down concentrates may be recirculated to the thickener by this same pump.

It is essential to have flexibility in any mill circuit, but particularly in large-tonnage operations such as this. Changing ore, changing market conditions and many other factors make this flexibility absolutely necessary. A slight change, easily made, in a flexible flowsheet may increase tonnage, improve recovery and lower grinding and reagent costs.

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.

flotation circuit - an overview | sciencedirect topics

flotation circuit - an overview | sciencedirect topics

Flotation circuits are a common technology for the concentration of a broad range of minerals and wastewater treatments. Froth flotation is based on differences in the ability of air bubbles to adhere to specific mineral surfaces in a solid/liquid slurry. Particles with attached air bubbles are carried to the surface and removed, while the particles that are not attached to air bubbles remain in the liquid phase. The concept is simple, but the phenomena are complex because the results depend on what happens in the two phases (froth and pulp phases) and other phenomena, such as particle entrainment. In flotation, several parameters are interconnected, which can be classified into chemical (e.g., collectors, frothers, pH, activators, and depressants), operation (e.g., particle size, pulp density, temperature, feed rate and composition, and pulp potential), equipment (e.g., cell design, agitation, and air flow), and circuit (e.g., number of stages and configuration). If any of these factors is changed, it causes (or demands) changes in other parts, and studying all of the parameters simultaneously is impossible. Conversely, there is not a model that includes all variables; most of the models are empirical and use only a few variables.

In the literature, various methodologies for flotation circuit design have been proposed, with most using optimization techniques. In these methodologies, the alternatives are presented using a superstructure, a mathematical model is developed, and an algorithm is used to find the best option based on an objective function. The differences between these methodologies depend on the superstructure, the mathematical representation of the problem, and the optimization algorithm. However, one problem with these methods is that the recovery of each stage must be modeled, and because the recovery of each stage is a function of many variables and is difficult to model, the results are usually debatable.

Flotation circuits are commonly used for the concentration of a broad range of minerals; such circuits also used in wastewater treatment (Rubio et al. 2002). Froth flotation is based on the differences in the ability of air bubbles to adhere to specific mineral surfaces in a solid/water slurry. Then particles with or without air bubbles attached are either carried to the surface and removed or left in the liquid phase. The current method used to design these circuits is based on seven steps (Harris et al., 2002): 1) conduct a mineralogical examination in conjunction with a range of grinding tests; 2) conduct a range of laboratory scale batch tests and locked cycle tests; 3) create a circuit design based on scale-up laboratory kinetics; 4) perform a preliminary economic evaluation of the ore body; 5) pilot-plant test the circuit design; 6) evaluate the economy; 7) design a full-scale plant. This procedure has several problems: 1) the design of the circuit is made in step three based on rule-of-thumb scale-up from laboratory data; therefore, the design depends heavily on the designer's experience; 2) building laboratory and pilot plants are costly and time consuming; therefore, the designed circuit cannot be analyzed in depth; 3) other aspects, such as system dynamics, are not considered in the design process.

Froth flotation design and operation are complex tasks because various important parameters are interconnected. The parameters can be classified into four types of components, as shown in Fig. 1 If any of these factors is changed, other parts of the design will also be changed. It is impossible to study all of the parameters at the same time. For example, if six parameters are selected for study in a four-stage circuit, over eight million tests are required for a two-level fractional experiment design. In addition, for any given number of stages, there are several potential circuit configurations. For example, if six flotation stages are considered, more than 1,400 potential circuit configurations will be possible. Therefore, the design will not be analyzed in depth.

In this work, a new methodology that integrates the first five design steps outlined previously is presented. The objectives are 1) to better orient the goals of the laboratory tests; 2) to reduce the number of laboratory and pilot-plant testing, and achieving lower cost and execution times; 3) to design the flotation circuit systematically, and 4) to expedite the design process. The methodology, inspired by the work of d'Anterroches and Gani (2005), has three design steps: 1) definition and analysis; 2) process synthesis; and 3) final design. Four tools are used in this procedure: 1) laboratory testing, 2) process group contribution, 3) sensitivity analysis, and 4) reverse simulation (Fig. 2).

The example of flotation circuit without grinding considered the concentration of copper ore. The feed to the circuit corresponded to 6 t/h of chalcopyrite (33% of copper), 12 t/h of chalcopyrite slow (16% of copper), and 300 t/h of gangue. The superstructure considers five flotation stages. If all interconnection was allowed, there were over 3 million circuit structure alternatives. However, if origin-destination matrices were used to eliminate nonsense and redundant alternatives, the number of feasible flotation circuits was 6912. The procedure utilized for the postulation of a superstructure and the formulation of the mathematical programming model was the one utilized by Calisaya et al. (2016), which corresponded to a MINLP. The variables with uncertainties corresponded to the stage recoveries of the chalcopyrite, chalcopyrite slow, and gangue. The stage recoveries were difficult to model as there is not a model that can be used under all flotation circuit structures included in the superstructure. Here, the stage recoveries were represented by values obtained from the uniform distribution. Under these conditions, the design problem is a MILP.

After the optimal flotation circuit configuration is determined, the water integration problem is addressed. This problem has not been analyzed before because the main concern in mineral processing has been the recovery and the product grade. Water can be recycled from tail or concentrate dewatering operations. However, only some of the water recovered in these operations can be recycled because it affects the flotation behavior.

Recently, El-Halwagi et al. (2004) have used the concept of clustering for process design based on property integration. Property integration is defined as a functionality-based holistic approach to the allocation and manipulation of streams and processing units, which is based on tracking, adjusting, assigning, and matching functionalities throughout the process. Here the methodology developed by El-Halwagi et al. (2004) was used to design the water integration problem.

The overall problem definition given by El-Halwagi et al. (2004) can be adapted as follows: "Given a flotation circuit with certain sources (process streams and water streams) and flotation units along with their properties (pH, solid concentration, oxygen concentration) and constraints, it is desired to develop graphical techniques that identify optimum strategies for allocation and interception that integrate the properties of sources, sinks, and interceptors so as to optimize a desirable process objective (minimum usage of fresh water, maximum utilization of water recycled from tail and concentrate dewatering operations, minimum cost of slaked lime while satisfying the constraints on properties and flow rate for the sinks".

The objective of the flotation circuit used by the mining concentration plant in El Salvador is to obtain a concentrate rich in Cu and Mo. The influence of the input factors on the variability of the grade and recovery of Cu and Mo in the final concentrate is analyzed. The input factors to consider are the kinetic constants of each species in each stage, maximum recoveries of each species in each stage, residence time of the pulp in each stage, number of cells in the R and RS stages, solids concentration in each stage, froth depth height on the columns, and superficial air and water rates in the column. In total, there are 47 input factors, and uncertainty ranges were taken from Yianatos et al. (2005) and Yianatos and Henriquez (2006). Based on the current operating conditions of the El Salvador plant, variations of 15% and 3% for kinetic constants and maximum recoveries, respectively, were considered. In the case of the solids concentration and the froth depth, variations of 10% were considered. In the case of superficial air and water rates, variations of 11% and 12%, respectively, were considered. The variation in the number of cells in R and CS were 7-9 and 8-10, respectively. The variation of the residence time in the R and CS stages was 10%. Finally, the residence time in C varied by 7%.

Figure 2 shows that input factors 10 (kmax molybdenite in C), 11 (kmax of pyrite in C), 25 (surface flow), 27 (Rmax chalcopyrite in R), 30 (Rmax molybdenite in R), and 42 (Rmax molybdenite in CS) are the main sources of variability of Cu and Mo recoveries and grades. Input factor 10 affects the Mo grade and recovery, input factor 11 affects the Mo and Cu grades, input factor 25 affects both the Cu and Mo grades and recoveries, input factor 27 affects the Cu recovery, and input factors 30 and 42 affect the Mo grade and recovery. From these results, it is clear that six of the 47 input factors are the key variables for the Cu-Mo grade and recovery.

There are various methods available for cleaning fine coal, of which froth flotation has become the most common practice. Froth flotation depends on differences in surface properties between coal and shale. Air bubbles are generated within an aqueous suspension of fine raw coal with a solids concentration of less than 10%. The hydrophobic coal particles attach to the air bubbles and are buoyed to the top of the froth flotation cell where they are removed as froth. The hydrophilic shale particles remain as a suspension and are removed over the tailings weir. The property of hydrophobicity is imparted to coal particles by the addition of a collector like diesel oil. This facilitates the attachment of coal to air bubbles in preference to gangue particles. The aircoal attachment is made stable by the addition of a frothing agent like pine oil. Successful flotation is governed by different factors like oxidation and rank of coal, flotation reagents, agitation and aeration, particle size and pulp density, flotation machine, conditioning time, and pH of the pulp.

Conventional mechanically agitated flotation machines use relatively shallow rectangular tanks, whereas column cells are usually tall vessels with heights normally varying from 7 to 16m as per requirement. Column cells do not use mechanical agitation and are typically characterised by an external sparging system, which injects air into the bottom of the column cell. The absence of intense agitation promotes higher degrees of selectivity. Modern flotation machines are high-intensity equipment designed to create very small bubbles and higher flotation rates. Smaller bubbles are generated by intensive mixing of pulps with air so that fast collisions between particles and bubbles take place. Microcel machines work with forced air, whereas the Jameson cell works with induced air. These machines are particularly suitable for coal flotation (Lynch et al., 2010).

Hydrodynamic analyses have shown that the use of air bubbles smaller than typically generated by conventional flotation machines can improve fine coal recovery. The selectivity also increases as smaller bubbles rise more slowly through the pulp, leaving the high ash impurities at the bottom. One disadvantage of flotation is that efficiency reduces for size range below 100m. To overcome this constraint, a new stirrerless cone-shaped flotation cell was developed in Germany (Bahr, 1982), now called the Pneuflot. This cell uses a novel aeration technique in which minute bubbles are introduced into the slurry before it reaches the cell. The upper particle size is restricted to 300m.

Flotation circuits are simple for Indian coals. Concentrates can be produced in one stage of flotation, and recleaning of the products may not be generally necessary. In the case of highly oxidised coal, two-stage flotation may be required to be incorporated, with rougher cells and cleaner cells. The circuit consists of a number of banks depending upon the total quantity to be handled, whereas each bank can have four to eight cells. For smooth operation of the system, proper operation and control are necessary.

It is observed that the existing flotation circuits in India are not working well. There is substantial loss of coal substance along with tailings. The causes of poor performance can be attributed to the following major factors as stated by Haldar (2007):

In addition, the quality of coal fines has deteriorated and other parameters have changed. Lower seam coals of inferior quality are now supplied. The concept of treating fine coal may need changes. The floatability tests are illustrated in Fig. 8.8; it is possible to produce clean coal with ash% of 17%18% ash with sufficient yield.

A model for the design of flotation circuits under uncertainty has been presented. Uncertainty is represented by scenarios that include changes in the feed grade and in the metal price. The model allows the operating conditions (residence time and mass flows of each stream) and flow structure (tail and concentrate stream of cleaner and scavenger stage) to be changed for each scenario while the fixed design (size of cells in flotation stages) for all scenarios is maintained. The model can be modified to include other uncertainties and other adaptive variables.

To solve the two-stage stochastic model, two solution strategies were proposed. The results show that the use of average values for the stochastic parameters leads to an inefficient design and hence a decrease in the profits made in the process.

Finally, it can be concluded that the use of stochastic programming can be a beneficial tool in the design of a metallurgical process, specifically the copper flotation process. The optimal configuration is capable of adapting to uncertainty, leading to an increase in the company profits

The simplest way of smoothing out grade fluctuations and of providing a smooth flow to the flotation plant is by interposing a large agitated storage tank (agitator) between the grinding section and the flotation plant:

Any minor variations in grade and tonnage are smoothed out by the agitator, from which material is pumped at a controlled rate to the flotation plant. The agitator can also be used as a conditioning tank, reagents being fed directly into it. It is essential to precondition the pulp sufficiently with the reagents (including sometimes air, Section 12.8) before feeding to the flotation banks, otherwise the first few cells in the bank act as an extension of the conditioning system, and poor recoveries result.

Provision must be made to accommodate any major changes in flowrate that may occur; for example, grinding mills may have to be shut down for maintenance. This is achieved by splitting the feed into parallel banks of cells (Figure 12.53). Major reductions in flowrate below the design target can then be accommodated by shutting off the feed to the required number of banks. The optimum number of banks required will depend on the ease of control of the particular circuit. More flexibility is built into the circuit by increasing the number of banks, but the problems of controlling large numbers of banks must be taken into account. The move to very large unit processes, such as grinding mills, flotation machines, etc., in order to reduce costs and facilitate automatic control, has reduced the need for many parallel banks.

Some theoretical considerations have been introduced (Section 12.11.2), but there is a practical aspect as well: if a small cell in a bank containing many such cells has to be shut down, then its effect on production and efficiency is not as large as that of shutting down a large cell in a bank consisting of only a few such cells.

Flexibility can include having extra cells in a bank. It is often suggested that the last cell in the bank normally should not be producing much overflow, thus representing reserve capacity for any increase in flowrate or grade of bank feed. This reserve capacity would have to be factored in when selecting the length of the bank (number of cells) and how to operate it, for example, trying to take advantage of recovery or mass pull profiling. If the ore grade decreases, it may be necessary to reduce the number of cells producing rougher concentrate, in order to feed the cleaners with the required grade of material. A method of adjusting the cell split on a bank is shown in Figure 12.54. If the bank shown has, say, 20 cells (an old-style plant), each successive four cells feeding a common launder, then by plugging outlet B, 12 cells produce rougher concentrate, the remainder producing scavenger concentrate (assuming a R-S-C type circuit). Similarly, by plugging outlet A, only eight cells produce rougher concentrate, and by leaving both outlets free, a 1010 cell split is produced. This approach is less attractive on the shorter modern banks. Older plants may also employ double launders, and by use of froth diverter trays cells can send concentrate to either launder, and hence direct concentrate to different parts of the flowsheet. An example is at the North Broken Hill concentrator (Watters and Sandy, 1983).

Rather than changing the number of cells, it may be possible to adjust air (or level) to compensate for changes in mass flowrate of floatable mineral to the bank. To maintain the bank profile at Brunswick Mine, total air to the bank was tied to incoming mass flowrate of floatable mineral so that changes would trigger changes in total air to the bank, while maintaining the air distribution profile (Cooper et al., 2004).

Consider that the circuit discussed in section 2 corresponds to a flotation circuit in which a material composed of gangue (specie 2) and a valuable species (species 1). The operating conditions of this circuit are such that the transfer functions for species 1 are, TR1=0.75, TS1=0.85, TC1 =0.73, and for specie 2 are TR2=0.25, TS2=0.40, TC2=0.30. Each stage corresponds to flotation bank with 5, 8 and 7 cells in the stages R, S and C, respectively. The transfer function for each stage is given by (Cisternas et al., 2006)

Under the conditions of Figure 2, we have thatR1 ( TR1, 0. 86 ,0.73 ) TR1 and R2 ( TR2, 0 40, 030) TR2 . Also, as TR1=0.75 for species 1 and TR1 =0.25 for species 2, we have the following overall recoveriesR1(0.75, 0.86, 0.73)=0.94 and R2(0.25, 0.40, 0.30 =0.143 (i.e., 94% for species 1 and 14.3% for species 2.) A sensitivity analysis can help us to improve the operational conditions of the stages involved in the process, and, therefore, reduce the percentage of gangue recovery without affecting too much the percentage recovery of useful material. The recovery and sensitivity of the above circuit are given by the equations 1 to 4. Then, setting the transfer functions TSj and TCj in the eq.2, the behavior of global recovery sensitivity with respect to the transfer function TRjcan be studied as shown in figure 3. Figure 3a shows Rj/TRj versus TRj for values of (TSj,TCj) close to the species 2. It can be seen that sensitivity increases as the value of TRjincreases. The opposite behavior is shown in Figure 3b which corresponds to values of (TSj,TCj) near species 1. This means that the behavior of the sensitivity is inversely between the valuable specie and the gangue.

Figure 4a shows RjTijversus Tijfor values of (TRj,TSj,TCj)close to species 2.We can observe that for species 2, the highest sensitivity to the global recovery is given in ( 0.25 , 0.4, TC2) and ( TR, 0.40, 0.3 ), which indicates that it is more sensitive to the transfer functions in the stage C (TC2 ) and R (TR2 ). Looking at the graphs in Figure 4b, we can observe that for species 1, the highest sensitivity in the global recovery is given in (0.75,TS1, 0.73), which indicates that it is more sensitive to the transfer function in the stage S (TS1). The same analysis can be performed for the other derivatives in equations 3 and 4. Performing the sensitivity analysis, it is possible to reach the following conclusions: 1) for species 1, the highest sensitivity to the global recovery for the stages R, S and C are given in ( 0.75 , TS1, 0.73 ),( 0.75 ,TS1, 0.73 )-( 0.75 , 0.86 , TC)and (0.75,TS 1, 0.73)-(0.75,0.86,TS 1), respectively. On the other hand, for species 2, the highest sensitivity to the global recovery for the stages R, S and C are given in (0.25, 0.4,TC2)-(TR2, 0.4,0.3),(0.25, 0.4,TC2)-(TR2, 0.4,0.3) and (TR2, 0.4, 0.3), respectively.

With this information about the behavior of global recovery, the sensitivity with respect to the transfer functions, we can intuitively change the values of the transfer functions of species 1 and 2 in the stages where they are most influential. Then, by reverse simulation with equation 5, it is possible to determine new designs (N) and/or operating conditions(kijand/or ) to achieve a better system performance. Thus, by changing the number of cells at 11 and 9 for stages S and C respectively, it is possible to obtain the following results: R1(0.75, 0.80, 0.68)=0.91 and R2(0.25, 0.40, 0.20)=0.077.

The self-tuning control algorithm has been developed and applied on crusher circuits and flotation circuits [2224] where PID controllers seem to be less effective due to immeasurable change in parameters such as the hardness of the ore and wear in crusher linings. STC is applicable to non-linear time-varying systems. It however permits the inclusion of feed forward compensation when a disturbance can be measured at different times. The STC control system is therefore attractive. The basis of the system is

The disadvantage of the set-up is that it is not very stable and therefore in the control model a balance has to be selected between stability and performance. A control law is adopted. It includes a cost function CF, and penalty on control action. The control law has been defined as

A block diagram showing the self-tuning set-up is illustrated in Figure20.26. The disadvantage of STC controllers is that they are less stable and therefore in its application, a balance has to be derived between stability and performance.

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