copper ore flotation cell manufacturer

flotation reagents

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

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

copper production: how is copper made?

Copper processing is a complex process that involves many steps as the manufacturer processes the ore from its raw, mined state into a purified form for use in many industries. Copper is typically extracted from oxide and sulfide ores that contain between 0.5 and 2.0% copper.

The refining techniques employed by copper producers depend on the ore type, as well as other economic and environmental factors. Currently, about 80% of global copper production is extracted from sulfide sources.

Regardless of the ore type, mined copper ore must first be concentrated to remove gangue or unwanted materials embedded in the ore. The first step in this process is crushing and powdering ore in a ball or rod mill.

Virtually all sulfide-type copper ores, including chalcocite (Cu2S), chalcopyrite (CuFeS2) and covellite (CuS), are treated by smelting. After crushing the ore to a fine powder, it is concentrated by froth flotation, which requires mixing the powdered ore with reagents that combine with the copper to make it hydrophobic. The mixture is then bathed in water along with a foaming agent, which encourages frothing.

Jets of air are shot up through the water forming bubbles that float the water repellent copper particles to the surface. The froth, which contains about 30% copper, 27% iron and 33% sulfur, is skimmed off and taken for roasting.

If economical, lesser impurities that may be present in the ore, such as molybdenum, lead, gold, and silver, can also be processed and removed at this time through selective flotation. At temperatures between 932-1292F (500-700C), much of the sulfur content remaining is burned off as sulfide gas, resulting in a calcine mix of copper oxides and sulfides.

Fluxes are added to the calcine copper, which is now about 60% pure before it is heated again, this time to 2192F (1200CC). At this temperature, the silica and limestone fluxes combine with unwanted compounds, such as ferrous oxide, and bring them to the surface to be removed as slag. The remaining mixture is a molten copper sulfide referred to as matte.

The next step in the refining process is to oxidize liquid matte in order to remove iron to burn off sulfide content as sulfur dioxide. The result is 97-99%, blister copper. The term blister copper comes from the bubbles produced by sulfur dioxide on the surface of the copper.

In order to produce market-grade copper cathodes, blister copper must first be cast into anodes and treated electrolytically. Immersed in a tank of copper sulfate and sulphuric acid, along with a pure copper cathode starter sheet, the blister copper becomes the anode in a galvanic cell. Stainless steel cathode blanks are also used at some refineries, such as Rio Tinto's Kennecott Copper Mine in Utah.

After crushing oxide-type copper ores, such as azurite (2CuCO3 Cu(OH)3), brochantite (CuSO4), chrysocolla (CuSiO3 2H2O) and cuprite (Cu2O), dilute sulfuric acid is applied to the surface of the material on leaching pads or in leaching tanks. As the acid trickles through the ore, it combines with the copper, producing a weak copper sulfate solution.

Solvent extraction involves stripping the copper from the pregnant liquor using an organic solvent, or extractant. During this reaction, copper ions are exchanged for hydrogen ions, allowing the acid solution to be recovered and re-used in the leaching process.

The copper-rich aqueous solution is then transferred to an electrolytic tank where the electro-winning part of the process occurs. Under electrical charge, copper ions migrate from the solution to copper starter cathodes that are made from high purity copper foil.

Other elements that may be present in the solution, such as gold, silver, platinum, selenium, and tellurium, collect in the bottom of the tank as slimes and can be recovered through further processing.

The development of SX-EW has allowed copper extraction in areas where sulfuric acid is not available or cannot be produced from sulfur within the copper ore body, as well as from old sulfide minerals that have been oxidized by exposure to air or bacterial leaching and other waste materials that would have previously been disposed of unprocessed.

This process involves drilling boreholes and pumping a leachate solution - usually sulfuric or hydrochloric acid - into the ore body. The leachate dissolves copper minerals before it is recovered via a second borehole. Further refining using SX-EW or chemical precipitation produces marketable copper cathodes.

Global mine production of copper is estimated to have exceeded 19 million metric tons in 2017. The primary source of copper is Chile, which produces approximately one-third of the total world supply. Other large producers include the US, China, and Peru.

Due to the high value of pure copper, a large portion of copper production now comes from recycled sources. In the US, recycled copper accounts for about 32% of annual supply. Globally, this number is estimated to be closer to 20%.

The largest corporate producer of copper worldwide is the Chilean state enterprise Codelco. Codelco produced 1.84 million metric tonnes of refined copper in 2017. Other large producers include Freeport-McMoran Copper & Gold Inc., BHP Billiton Ltd., and Xstrata Plc.

pulp density - an overview | sciencedirect topics

The metallurgical cyclone is the most commonly used classifying device for fine particle sizing and desliming in both alluvial gold processing plant (Section8.3.4), and for closed circuit grinding in modern chemical leaching plant. Good cyclone separation depends upon control of pressure drop, pulp density and apex size. The pressure drop may vary but should not change rapidly, and is held at safe levels by keeping an adequate sump level. A falling sump level causes cavitation in the pump and reduction in feed rate; pressure drop in the cyclone falls and solids report increasingly to the overflow until the drop approaches zero and the entire slurry stream passes into the underflow. Additionally, while maintaining the required separation parameters, the pressure drop should always be minimised to minimise energy losses, thus reducing pump and cyclone wear.

The maximum pulp density is usually about 50% solids by weight; above that level small fluctuations in density will seriously affect separation. A coneshaped discharge of 2030 reduced angle usually produces optimum conditions for separation. Cyclone control is best obtained by optimising the feed density. With consistent ore types the cyclone feed density is a good indicator of cyclone overflow sizing. A ropy cyclone underflow indicates a very high-density state with a risk of plugging the apex. If control cannot be exercised either a larger apex is needed, or the addition of another cyclone. Striking an economic balance between high D50 and low D50 separation requires:

higher hydrocyclone pumping pressures with consequent higher power and maintenance costs; if the mill is unable to grind the ore at the given feed rate the final grind will not be any finer no matter what adjustments are made to the cyclone.

It is important to recover coarse gold as soon as possible to minimise fragmentation and smearing onto other particles during multiple passes through the grinding circuit. Unit capacity cost of an enhanced gravity concentration device is high and in order to minimise the capacity of an installed enhanced gravity concentrating device, conventional practice is to connect such a device to only a fraction of the cyclone underflow. It is considered that, statistically, any free gold escaping the grinding circuit via overflow processes will follow the laws of probability and eventually report to the gravity concentration device after several passes through the grinding mill. McAlister and Sprake (1999) propose that the incoming feed to the cyclone can be scavenged prior to cycloning. The cyclone feed pipeline is so arranged that an outlet under pressure can be installed on the bottom of a line at the end of a straight section that is either horizontal or inclined. Material taken off can be elevated to a sizing screen under its own pressure, the coarse fraction passing directly to gravity separation.

Figure 6.12 shows measured pulp density profiles in an industrial thickener operating normally and in an overloaded condition. The measured profile during normal operation agrees with that expected in an ideal thickener and the gradual increase in the pulp concentration between the lower conjugate concentrate CM and the discharge concentration CD is clearly evident. The data is from Cross (1963). The thickener was 22.9 m in diameter with a 3 m cylindrical section and a cone depth of 1.55 m.

Figure 6.12. Measured density profiles in an industrial thickener. The specific gravity of the feed was 1.116 and the underflow discharged at a specific gravity of 1.660 under normal conditions. Data is from Cross (1963)

Ideal Kynch thickening behavior does not describe the entire thickening process because it is not applicable when the sediment is under compression at the bottom of the thickener. Ideal Kynch thickening terminates when the slurry concentration is sufficiently high to allow individual flocs to touch and support each other. Consequently there is no natural settling of the individual particles relative to the water. However, as the floc bed increases in height, the weight of the accumulated flocs compresses the lower layers of the floc bed and the water is squeezed out. This water is forced upward through the floc bed and the upward drag on the flocs actually helps to support them. Even if the individual flocs are not themselves compressible, groups of flocs exhibit compressive behavior in that compression forces express water from the voids between the flocs. The touching flocs generate a structure that has internal strength which is a function of the solid concentration. This internal strength manifests itself as a normal stress on the solid phase and it is this stress that supports the upper layers of flocs in the compression zone (see Figure 6.13).

While the floc bed is being compressed, the interstitial water flows upward through the floc bed. The viscous drag generated by this upward flowing water helps to support the layers of flocs. The concentration at which the flocs just form supportive contacts is called the critical concentration CC. This concentration plays a pivotal role in determining the behavior of both batch and continuous thickeners. In the upper part of the thickener where the slurry concentrations are less than CC, ideal Kynch behavior occurs. In the lower regions of the thickener where slurry concentrations are greater than CC, a different model is required to describe the behavior of the sediment. In particular the settling velocity of the flocs in the sediment depends not only on the local concentration of the solids but on the gradient of the concentration as well, a condition which violates the basic Kynch postulate. A force balance over a horizontal slice of compressible floc bed will generate an equation for the rate of increase in the solid stress at increasing depth.

Where K is the permeability of the flocculated sediment, the viscosity of the fluid and Vr the interstitial velocity of the water relative to the solid. The drag force can be expressed in terms of the terminal settling velocity of the floc bed by noting that if a slice of the bed were free to settle under its own weight (no support from below or pressure from above and no friction on the vertical walls) it would do so at its terminal settling velocity. Under these conditions

Equation 6.77 is a non-linear parabolic partial differential equation that can be solved only numerically using initial and boundary conditions that are appropriate to the physical problem that is to be solved. Brger et al. (1999) discuss solutions for a variety of important transient conditions in operating ideal thickeners.

Fundamental models for the effective solid stress in the compressible sediment have not yet been developed but the following empirical equations have been found to describe some typical compressible sediments.

where c is the critical volume fraction and uis the ultimate volume fraction of the sediment at which all compression stops. Adorjan (1976) has described an experimental apparatus that can be used for the direct measurement of the sediment compressive strength and has found values of a = 182.6 and m = 1.81.

The bacterial oxidation of sulfide minerals permits biodissolution of metal ions of interest such as zinc from sphalerite and, copper from chalcopyrite, covellite, and chalcocite. In case of precious metals, such as gold and silver, bacterial oxidation liberates the precious metal embedded in pyrite or arsenopyrite matrix by solubilizing the sulfide mineral and leaving behind liberated gold or silver in the leach residue which can be subsequently recovered by cyanidation. A. ferrooxidans and L. ferrooxidans are the most widely used organisms in the bioleaching of sulfide minerals. Important mechanisms involved in bacterial oxidation are illustrated in Chapter 4, Bioleaching Mechanisms.

The ore/concentrate samples are taken in conical flasks at a desirable pulp density in the presence and absence of A. ferrooxidans. Control flasks contain a bactericide (alcoholic thymol). The flasks are kept in orbital shakers at a constant temperature (2530C). The solution from inoculated and control flasks are analyzed for metal ion concentration and pH as a function of time.

Column leaching is used for coarse ore particles, which facilitate efficient percolation of the leach solution. Airlift percolators can be designed and fabricated easily. A typical arrangement is illustrated in Fig. 14.10.

The columns are designed for suitable L/D ratios. The columns are packed with the ore sample on glass wool (or porous frit), which is used as the filtering disc. The liquid level has to be maintained at a height of 1.5cm above the bed of the ore sample. Columns are operated under conditions simulating situations prevailing in actual heap leach sites. The test conditions are similar to those used in the agitation leaching tests. Air compressor is used to ensure supply of oxygen, CO2, and continuous recirculation of the leached liquor.

Scanning electron micrographs of pyrite mineral prior to leaching and after leaching in presence of A. ferrooxidans is depicted in Fig. 14.11. The attachment of A. ferrooxidans onto the mineral is illustrated in Fig. 14.12.

Gold ores are classified free-milling and refractory based on their response to cyanide dissolution. Bioprocessing can be used to treat refractory gold-bearing sulfide ores as well as carbonaceous gold ores. It is essential to characterize different types of gold ores to understand the extent and nature of gold-entrapment in different associated mineral phases before an appropriate microbiological or abiotic process can be adapted to enhance gold extraction. The following steps need to be undertaken to establish bioleaching potential of gold-bearing ores and concentrates.

Diagnostic leaching can be used as an analytical tool to establish the distribution of gold within different mineral phases such as pyrite, arsenopyrite, silicates, calcite, dolomite, and carbonates. Such a procedure will enable the choice of a suitable process for gold extraction. The first step should, however, be mineralogical evaluation of the ore sample.

Diagnostic leaching is based on the concept that the least stable mineral in the ore matrix will be the first to be solubilized. The leach residue is cyanided to extract liberated gold and the process repeated using still higher oxidative acid leach until all the nonrefractory and refractory gold are liberated and cyanided [2729].

For a complex AuAg ore, direct cyanide leaching resulted in only about 47% and 19% extraction of gold and silver even after very fine grinding at <38m. Diagnostic leaching was carried out to assess the nature of refractoriness [29]. Depending on the results of the diagnostic test results, extraction strategies for gold and silver can be worked out. For example, for adaptation of biooxidation of refractory pyritearsenopyrite containing gold ores and concentrates, diagnostic leaching in presence of FeCl3 / HNO3 would establish the amount of precious metal tied-up with the above sulfides [29].

In the leaching of any metal from an insoluble ore two interrelated physical parameters can be critical, regardless of the extraction methods involved. These are pulp density and specific surface area of the mineral substrate to be leached. Pulp density describes the mass of mineral in unit volume of liquid available, usually expressed as a % (w/v). This term has most meaning in experimental systems where relatively small amounts of crushed or ground ore are suspended with continuous agitation in a large volume of liquid medium, as in laboratory scale tests of direct leaching by bacterial cultures (Torma, 1971; Torma et al, 1970, 1972, 1973, 1974). In such systems, using lead, zinc, nickel, cobalt and cadmium sulphides, chalcopyrite and pentlandite, the rate of extraction of metal by cultures of Thiobacillus ferrooxidans was limited by the pulp density up to a saturating value. The pulp density allowing maximum extraction rate was dependent on the particle size employed for any particular sulphide. Thus for a single pulp density, extraction rate increases as the particle size is reduced. This indicates that the rate of leaching in bacterial cultures is limited by the surface area available for attack by the bacteria or their products. In general, therefore, leaching is most rapid and most complete when the mineral being leached is ground to very small particle size and the greatest possible surface area (e.g. up to 27 m2/g) is exposed per unit volume of culture. With minerals whose sulphates are insoluble, or concentrates containing lead sulphide with Cu, Cd and Zn, the total amount of metal extracted may decrease with higher particle size (i.e. lower specific surface area) because the mineral surface becomes coated with insoluble lead sulphates. Such problems can be overcome by (a) smaller initial particle size; or (b) regrinding of the mineral after an initial leaching of larger particles. The latter may be economically more realistic.

This work, particularly of Torma's group, has shown that a highly controllable and predictable process can be operated to give rapid, high extraction of metals from sulphides in contact with pure cultures of Thiobacillus ferrooxidans. Under such conditions adequate nutrient supply (O2, CO2, ammonia, Mg, P etc) for the bacteria is essential.

The use of extreme thermophilic archaea (Sulfolobus acidocaldarius like) at temperatures of the order of 65C in the extraction of zinc from complex sulfides has been demonstrated. Higher pulp densities of the order of 15%20% can be used. Pilot tests with ores containing 3% Zn, 0.15% Cu, 0.6% As, 29% Fe, 25% S besides gold and silver resulted in 91% zinc recovery at pH 1.6. Gold and silver can be recovered after cyanidation of neutralized leached residues [2528]. Biooxidation of a fine-grained, complex zinc and gold-containing sulfide ore has been performed in a series of experiments at a bench scale with 20L leaching volume in a series of three continuously stirred reactors. A mixed culture of moderate thermophilic bacteria was used for bioleaching at 45C, and a mixed culture of extreme thermophilic archaea used for bioleaching at 65C. The leaching rates for zinc were in the range 80%87% with the moderate thermophilic bacteria and 96%98% with the extremely thermophilic microorganisms. It was found that to obtain a high zinc recovery with a low-degree of pyrite oxidation, a fine particle size was essential.

The step-by-step operations were as follows: calcinations of the initial material (barium hydroxide/fluorescent powder equal to 0.4); leaching of REEs with sulfuric acid (10% pulp density, 2M sulfuric acid, 2h, and 52.5C); Precipitation of yttrium and europium with oxalic acid (1h, room temperature); calcination of the oxalates (1h, 600C); wastewater treatment.

One of the most important problems with spent FL disposal is represented by the presence of mercury. The proposed process reduced the dangerousness of the waste by volatilization of mercury during the thermal pretreatment. From both industrial and environmental points of view, the pretreatment section must be properly designed to treat the hazardous emissions. The residual outputs of the process were: the leaching cake and the residual solution after precipitation. The cake did not contain mercury; hence this is less dangerous in respect to the initial material. However, it needs to be properly disposed of to avoid the dispersion of materials that contain metal sulfates and other impurities. The residual solution that contains metals, sulfates, and oxalates can be partly recycled, and part could be purified, e.g., with lime. The treated water could be recycled in the process.

Bucket-wheel dredgers are generally preferred to suction cutter dredgers for production dredging. Bucket wheels are more able to cut harder materials; they clean up more effectively at bedrock and deliver the slurry at a higher pulp density to the treatment plant. Their use is currently limited to a dredging depth of about 30m because of the weight of the wheel and ladder. Although suffering the same clogging and blockage problems as suction cutter dredgers the bucket wheel is more suited to overcoming them. For example, the bucket-wheel configuration shown in Fig.7.25 may be fitted with clearance fingers to cut through roots and hard clayey fragments.

In the bucket wheel mining operation described diagrammatically in Fig.7.25, the dredger pumps the spoil through floating slurry pipelines to a gravity treatment plant floating in the same pond. Manoeuvring of both dredger and treatment plant unit is usually effected through a combination of spuds and anchor-lines or by crossed bow, side and stern lines. This particular dredger is manoeuvred using side-slewing winches. The method of advancing an operating dredger is an important factor affecting its efficiency. In Fig.7.26 the Ellicott Company compares the dredging efficiencies of (a) conventional walking spuds and (b) the Ellicott spud carriage system.

Particle size and frequency are determining factors in pipeline transportation and manufacturers normally supply separate pump performance curves and tables for silts, sands and gravels. These charts offer general solutions for specific physical relationships to assist in preliminary studies. They do not however, offer an unambiguous means of predicting the performance of pumps that are called upon to handle heterogeneous and constantly changing mixtures of sediments from a dredging face. The final pump selection is a more or less safe compromise for the particular set of conditions based upon the results of detailed screen analyses and the manufacturers experience.

The density of the slurry is measured automatically and continuously in the nucleonic density gauge (Figure 3.23) by using a radioactive source. The gamma rays produced by this source pass through the pipe walls and the slurry at an intensity that is inversely proportional to the pulp density. The rays are detected by a high-efficiency ionization chamber and the electrical signal output is recorded directly as pulp density. Fully digital gauges using scintillation detectors are now in common use. The instrument must be calibrated initially on-stream using conventional laboratory methods of density analysis from samples withdrawn from the line.

The mass-flow unit integrates the rate of flow provided by the flowmeter and the pulp density to yield a continuous record of tonnage of dry solids passing through the pipe, given that the specific gravity of the solids comprising the ore stream is known. The method offers a reliable means of weighing the ore stream and removes chance of operator error and errors due to moisture sampling. Another advantage is that accurate sampling points, such as poppet valves (Section 3.2.4), can be incorporated at the same location as the mass-flow unit. Mass-flow integrators are less practicable, however, with concentrate pulps, especially after flotation, as the pulp contains many air bubbles, which lead to erroneous values of flowrate and density. Inducing cyclonic flow of slurry can be used to remove bubbles in some cases.

India is a large consumer of gold while in terms of production of the precious metal, India produces only about 46 tons of the metal per year. The Hutti Gold Mines Company Limited (HGML) in the state of Karnataka is the only primary gold producer in the country. Most of the ore is mined by underground mining, and gold is extracted from free milling ores by cyanidation using carbon-in-pulp method.

HGML has identified several newer deposits containing gold and silver. G.R.Halli and Ajjanahalli deposits in Karnataka are sulfidic ore deposits containing about 34g of gold per ton of ore. Recovery of gold from such ores or their flotation concentrates by conventional cyanidation has not been encouraging. The concentrate needs to be biooxidized before it is cyanided to enhance gold and silver recovery. HGML and Indian Institute of Science, Bangalore, undertook a joint project to develop bioreactor technology for biooxidation of the refractory G.R.Halli concentrate with financial support from the Department of Biotechnology, Government of India, during 19972002 [25,26].

Typical as mined gold- and silver-bearing pyritearsenopyrite ore, samples from G.R.Halli mines were initially characterized, crushed, and ground for flotation beneficiation. The flotation concentrate contained 45%50% pyrite, 20%25% arsenopyrite, 23g per ton of gold, and 3000g per ton of silver. A. ferrooxidans, isolated and cultured from the mine site itself, were preadapted to As(III), As(V), and the concentrate. Bacterial adaptation to dissolved arsenic was achieved through serial subculturing with increasing concentrations as shown in Fig. 8.8.

Adaptation to concentrates was done through serial bacterial growth under systematic increase in pulp densities from 1% to 20%. Growth was seen to be impacted beyond 15% pulp density. Benefit of using preadapted strains on iron dissolution is illustrated in Fig. 8.9.

Initial biooxidation tests were carried out in batch mode in a laboratory bioreactor assembly. The concentrate was first acid stabilized before addition of adapted cells. Cell growth and iron oxidation were monitored as shown in Fig. 8.10. Solid-leached residues were neutralized, lime-treated, and subjected to cyanidation to establish gold recovery as a function of percent iron sulfide oxidation (Fig. 8.11). Without biopretreatment, only about 50% gold recovery could be obtained. Gold recovery increased steadily with biooxidation of iron sulfides and about 85%90% recovery could be obtained corresponding to 85% oxidation.

Biooxidation tests were subsequently carried out on a continuous mode. Initially the tests were carried out at 1% pulp density and the iron leached estimated. When the iron leached out approximately attained the theoretical amount in the concentrate, more concentrate was added and stepwise leaching continued with varying pulp densities until the pulp density reached 10%15%. Further increase in pulp density did not promote the leaching rate.

Higher iron and sulfur content in the concentrate introduced a lag period, impairing leaching rates. Stepwise leaching was found to be more effective than starting at a 10% higher pulp density in the beginning itself. Most of the solubilized iron was in the form of ferric ions. The leaching rate of iron was commensurate with reported results for pyrite and arsenopyrite bearing gold ores. The reactor was then operated in a continuous mode with a feed rate of 1.5L/day and a residence period of 4 days. On an average, 90% of gold and 95% of silver could be extracted from the biooxidized concentrate. The bioreactor was operated for about 15 days in continuous mode and consistently good recoveries could be achieved.

After extensive laboratory testing for about a year, it was decided to demonstrate the technology on-site at HGML. A demonstration pilot facility was designed with a capacity to process 100kg of concentrate per day. Three bioreactors having a total capacity of 6m3 was used. The concentrate was fed from a feed tank to the first two bioreactors, arranged in parallel. Discharged products from Reactors 1 and 2 were fed to Reactor 3 from where the biooxidized slurry was sent to a settling tank for subsequent cyanidation. The reactors were provided with thermostated water jackets for controlling the temperature at 30C. The bacteria (A. ferrooxidans) were cultured in the bioreactors itself in the presence of the concentrate by the step-leaching technique with all the necessary nutrients for bacterial growth without added ferrous sulfate. The important process variables such as Eh, pH, temperature, pulp density, iron concentration, and bacterial population were monitored at regular intervals. Flowchart for the demonstration of biooxidation plant is shown in Fig. 8.12.

Using a mixed culture of A. ferrooxidans and A. thiooxidans, 87% of copper could be recovered from the flue dust. Using stirred tank bioreactors, copper recovery could be further increased to 91%. Because the smelter flue dust contains secondary and primary sulfides of copper, leaching rates may vary. Continuous tests in two stages of air lift bioreactors were used to establish effect of pulp density, retention time, and temperature. Overall copper extraction at 2%, 4%, and 7% pulp densities were 90%, 89%, and 86% with mean retention times of 2, 7, and 4 days, respectively [6771].

Copper converter slags contain 1%9% Cu, 1%2% Ni, and 0.4%0.8% Co depending on the grade of the feed concentrates. In the presence of a mixture of A. ferrooxidans and A. thiooxidans, 66% Cu, 64% Co, and 50% Ni could be recovered from finely ground slag at pH 2 [7174].

The BRISA process, which involves dynamic leaching with ferric sulfate coupled with biogeneration of the lixiviant, was used for leaching copper converter slag using A. ferrooxidans. Of note, 93% Cu could be dissolved in 4hours at 60C, 20g/L pulp density, and pH of 1.7 from finely ground slag. A fungus, A. niger could leach 47% Cu, 50% Ni, and 23% Co from the slag. Biogenerated organic acids could complex metals present in slag particles [7577].

Bioleaching of tailings and slag from a flotation and smelting plant to recover copper and nickel was also tried using A. ferrooxidans [78]. From flotation tailings about 16% Cu and 45% Ni could be extracted. The use of a mixed culture containing A. ferrooxidans and L. ferrooxidans could leach most of copper and nickel from a dump slag at pH 1.5 [79]. A mixed culture containing thermophilic and mesophilic organisms was found to be more efficient in bioleaching of metals from a Pb-Zn smelter slag at 65C and pH of 1.5 [8082]. Bioleaching of a filter residue (58% Zn, 11% Pb, 6% Sn, and 0.5% Cu) from a copper smelter in the presence of Penicillium sp. could dissolve 93% Zn. Up to 90% of Zn could be dissolved in a bioreactor in the presence of Penicillium simplicissimum in the pH range 27 [8284].

Slags from smelting and other pyrometallurgical metal extraction processes contain valuable metals. Smelter slag from Boliden, Harjavalta, Finland, contained about 41% Fe, 2.8% Zn, 0.4% Cu, 0.05% Ni, 0.04% Co, 0.14% S, and almost 30% SiO2. Fayalite (Fe2SiO4) constituted 49% and magnetite 22%. Among the copper sulfides were bornite, chalcocite, and chalcopyrite, with 0.07% metallic copper, 0.23% ZnS, and 0.16% PbS. Ground slag samples (73% 45m) was used in bioleaching. Acidophilic iron-and sulfur-oxidizing bacteria enriched from the slag discharge pipe as well from water and sediment samples were used. The culture contained A. ferrivorans, Alicyclobacillus tolerans, Alicyclobacillus cycloheptanicus, and Alicyclobacillus herbarium [85]. The samples were acid preleached for acid stabilization to maintain pH at 1, 1.5, and 2.0. Bioleaching was carried out in shake flashes and in 3-L stirred tank reactors. Acid consumption ranged from 400900g H2SO4 at pH 12. About 80% Cu and 27% Zn were dissolved at 10% pulp density in shake flasks when elemental sulfur was provided for bacterial generation of H2SO4. Bioleaching in stirred tank reactor at 10% pulp density with a mixed culture resulted in 44% Cu, 14% Zn, 4.5% Ni, and 14% Co dissolution after 37 days.

In yet another bioleaching of the above copper smelter slag using an enrichment culture containing A. ferrooxidans, A. thiooxidans, A. caldus, Leptospirillum ferrooxidans, and Sulfobacillus thermotolerans in a continuous stirred tank reactor for 29 days at 5% pulp density, 41% Fe, 62% Cu, 35% Zn, and 44% Ni could be solubilized. Using an effluent from a sulfidogenic fluidized bed reactor, over 98% of both copper and zinc could be precipitated at pH values in the range 34. The reactor grown precipitating sulfate-reducing bacteria (SRB) culture contained Desulfovibrio, Desulfotomaculum, Desulfobulbus, Desulfurispora, and Desulfobacca [86,87].

More than 80%85% of copper, zinc, and arsenic could be bioleached from a Pb-Zn smelter slag using indigenous thermophile bacteria at pH 1.5, 10% pulp density, and a particle size of 0.8mm after 6 days [88].

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The flotation machine is driven by V-belt drive motor rotating impeller to create negative pressure by centrifugal vacuum. The flotation cell is a very important part of the flotation machine.The flotation cell, on the one hand, absorbs enough air mixed with plenty of pulp.

On the other hand, stirring the pulp mixed with drugs, the flotation machine refines foam to adhere to the mineral and float to the surface and then form the slurry bubble mineralization. You can adjust the height of the flashboard to control surface so that the useful foam could be scraped.

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Our automatic production line for the grinding cylpebs is the unique. With stable quality, high production efficiency, high hardness, wear-resistant, the volumetric hardness of the grinding cylpebs is between 60-63HRC,the breakage is less than 0.5%. The organization of the grinding cylpebs is compact, the hardness is constant from the inner to the surface. Now has extensively used in the cement industry, the wear rate is about 30g-60g per Ton cement.

Grinding Cylpebs are made from low-alloy chilled cast iron. The molten metal leaves the furnace at approximately 1500 C and is transferred to a continuous casting machine where the selected size Cylpebs are created; by changing the moulds the full range of cylindrical media can be manufactured via one simple process. The Cylpebs are demoulded while still red hot and placed in a cooling section for several hours to relieve internal stress. Solidification takes place in seconds and is formed from the external surface inward to the centre of the media. It has been claimed that this manufacturing process contributes to the cost effectiveness of the media, by being more efficient and requiring less energy than the conventional forging method.

Because of their cylindrical geometry, Cylpebs have greater surface area and higher bulk density compared with balls of similar mass and size. Cylpebs of equal diameter and length have 14.5% greater surface area than balls of the same mass, and 9% higher bulk density than steel balls, or 12% higher than cast balls. As a result, for a given charge volume, about 25% more grinding media surface area is available for size reduction when charged with Cylpebs, but the mill would also draw more power.