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twostage flotation cell of ore in ball mill

flash flotation with closed circuit grinding

flash flotation with closed circuit grinding

The reason why you need Flash Flotation in a Closed Grinding Circuit relates toRecovering your mineral as soon as free which has long been recognized in ore dressing practice. This not only applies to gravity treatment but also to flotation. For this application the FlashFlotation Cell was developed for use in the grinding circuit and has done a remarkable job in many plants.

A greater amount of granular higher grade concentrates can be produced and, in general, overall plant recovery is improved by reducing slime losses due to overgrinding and colliding of high specific gravity minerals.

Typical flowsheets are shown to indicate a few of the possible applications of FlashCells in grinding circuits. In recent years the successful application of hydraulic cyclones, rubber lined pumps, and two stage grinding circuits have enhanced the feasibility of unit cell applications. Cyclones in particular have increased the flexibility of such applications by permitting positive and continuous gravity flow of unit cell tailings to subsequent treatment steps.

Molded rubber wearing parts are used exclusively in unit cells. If wear is severe due to coarse abrasive solids, a special molded soft rubber compound is available which greatly extends the impeller and wearing plate life. Conical disk impellers and wearing plates are standard for unit cell applications.

Two stage classification is shown in this flowsheet with a Unit Cell between the classifiers. The primary classifier may overflow as coarse as 20 mesh and at densities up to 50% solids. This is ideal feed for the unit cell. Unit cell tailings are classified through a cyclone and the oversize returned to the ball mill for further grinding. The cyclone classifier overflow, 65 mesh or finer, is treated by regular bulk or selective flotation.

With this two stage classification system the unit cell can be conveniently located to deliver a positive gravity discharge of pulp to the pump feeding the cyclone. The pump sump box can be made a part of the unit cell tank if desired.

Two stage grinding and classification is provided in this flowsheet which is generally applicable to larger tonnage installations in which a substantial percentage of the values can be recovered directly from the grinding circuit. The primary Rod Mill in open circuit will reduce crushed ore to approximately minus 10 mesh.A Mineral Jig is recommended if coarse mineral and metallics are present.

The spiral or rake classifier overflow passes to the unit cell and on to classification and regrinding. High grade unit cell concentrates can be produced with this system, and on low ratio of concentration ores a substantial increase in mill capacity is possible. Slime losses are greatly minimized with this combination jig and unit cell circuit.

The trend in many of the large tonnage millingcircuits is to completely eliminate conventional rake or spiral classifiers by going to two stage grinding with a rod and ball mill in series. The rod mill discharge goes direct to the ball mill and then on to a pump and hydraulic cyclone classifier. These modern grinding and classification circuits are ideal for including the unit cell as the primary mineral recovery step.

One large copper operation with two stage grinding and cyclone classification, actually treats cyclone underflow, 20 plus 100 mesh, through Unit Cells. The unit cells recover a substantial percentage of the total copper in final concentrate form. The unit cell tailings at 55% solids return by gravity to the regrind ball mill feed. Since incorporating the unit cells and by careful checking between parallel circuits, it has been established by recovering the mineral as soon as free that final mill tailings were reduced by lb. copper per ton.

Unit Flash Cell flotation tests should be made before planning an installation. This will establish if the ore will respond to such treatment to advantage.A 100 lb. representative sample of the ball mill feed is sufficient for the unit cell flotation tests.

The simplest flotation circuit is a comparatively recent innovation. It consists of the introduction ofa flotation cell into the grinding circuit between ball mill and classifier as shown below.The discharge end of the mill is fitted with a trommel screen with openings about 4 mesh in size to separate out coarse material, which is laundered direct to the classifier ; the remainder of the pulp passes to the flotation cell where most of the mineral which has been released from the gangue is taken off as a concentrate, the necessary reagents having been added at some previous point in the circuit, usually at the mill feed box. Flotation can be carried out in a pulp containing as much as 65% of solids, but water is added in most cases to bring the W/S ratio to about 1/1. In such a thick, heavy pulp it is possibleto float particles as coarse as 10 mesh. At this size, however, only pure mineral will adhere to a bubble, for which reason the concentrate is generally of unusually high grade.

The advantage of this method of flotation is that the valuable minerals are removed from the circuit as soon as they have been released from the gangue, so that their accumulation in the classifier is prevented and the possibility of overgrinding them is reduced. Moreover, the granular nature of the particles floated assists considerably in the subsequent filtration of the combined flotation concentrate. It is a simple matter to instal the cell in the ball mill circuit, since it fits readily intothe space between mill and classifier and occasions no loss of head. The only disadvantage is the heavy wear to which the interior of the cell is subjected by the coarse material passing through it, but the modern method of lining both moving and stationary parts with rubber reduces this difficulty to minor proportions.

Mcintyre Porcupine Mines, Ltd., was one of the first companies to practise flotation in the grinding circuit. Their installation is described later in the paragraph headed Flotation of Gold and Silver Ores . Others soon also adopted the method. In their plant grinding is carried out in a Hardinge Ball Mill in closed circuit with a Dorr Classifier, and a Sub-A Cell is employed as the flotation unit between the two, the pulp being maintained at a density of 65% solids. Under normal operating conditions 60-70% of the copper and 40% of the nickel are recovered in the grinding circuit. Subsequent flotation in Sub-A Machines gives a total recovery of nearly 99% of the copper and over 94% of the nickel. No attempt is made to separate the copper from the nickel minerals.

The process can be adapted to the selective flotation of complex ores. Another mill for instance, where a lead-zinc ore is treated by two-stage selective flotation, the method being very similar to the standard procedure described in the paragraph entitled Flotation of Lead-Zinc Ores , the installation of a cell between ball mill and classifier has resulted in the removal of most of the galena before the pulp passes to the main flotation circuit. The sphalerite in the ore is so finely intermixed with a portion of the galena that, although a large proportion of the latter mineral is actually liberated at a comparatively coarse mesh, it is necessary to reduce the whole tonnage to 87% minus 200 mesh in order to separate the two minerals at all completely. Before flotation in the grinding circuit was tried, three stages of cleaning were required to make a high- grade lead concentrate, and 96% of the finished product would pass through a 325-mesh screen, only a trace remaining on 200 mesh. The introduction of a Sub-A Cell into the grinding circuit enabled over 70% of the lead to be recovered in one operation without cleaning in a concentrate running 65-70% lead. The concentrate contains only 53% of minus 325-mesh material, 25% remaining on 200 mesh, and, being more granular than that obtained in the main flotation circuit, it gives better filtration. The total recovery remains much the same as before. Reagents are added in the ball mill feed box, and the pulp is maintained in the cell at about 58% solids.

There is no necessity to limit the size of the flotation machine between ball mill and classifier to a single cell, through the use of a multi-cell or a long pneumatic machine would involve changing the relative positions of the three units from the present standard arrangement. The trendof progress indicates that the flotation machine may become in some cases as important a factor in the proper classification of the ore during grinding as the classifier itself is at present.

Hydrocyclones are used in many grinding circuits to make a size separation which ideally sends the fine ore fraction to conventional flotation and the coarse fraction back to the mill for further grinding. The separation results not only from particle size but also from particle specific gravity. The result is a cyclone underflow which is higher in grade than the cyclone feed. When comparing sulphides and precious metals to silicates, the floatable size fraction in the cyclone underflow contains higher mineral values than gangue. This results because the lower specific gravity gangue particles tend to follow the water in the cyclone overflow.

Typical grinding mill circulating loads range from 200% to 500% with a large part being particles which should have reported to conventional flotation. These particles, mainly heavy sulphides and precious metals, are being reground sometimes several times before eventually making it to the conventional flotation circuit. With each pass through the mill, the particles are ground finer until they are overground. This produces slimes which are usually lost to tailings. The attempt has been made in the past to recover valuable minerals from grinding circuits by flotation. This method, referred to as a unit cell operation, treated grinding mill discharge with a conventional flotation machine. Until now the success of this method has been limited by the inability of a conventional machine to treat the very coarse, high density slurries associated with grinding mill discharges.

Outokumpu developed a specially designed tank to work in conjunction with its flotation agitation mechanism to recover valuable minerals from grinding and classification circuits before they are overground and lost as slimes. The Outokumpu Skim-air coarse flotation machine has been used successfully to recover values from both grinding mill discharge and hydrocyclone underflow.

The Skim-Air Flash Flotation machine floats only those liberated valuable particles which are quick floating. There is not sufficient residence time in the machine to float the slower floating middlings and gangue particles. Thus, each time a particle with ideal fast floating characteristics reports to the cyclone underflow, it is removed in the Skim-Air.

The feed to the Skim-Air machine is normally in the range of 65% to 85% solids. In most cases optimum results are achieved with little or no dilution water being added to the machine. The concentrate produced in the Skim-Air is normally sent as final concentrate. Because only quick floating mineral particles have time to float in the Skim-Air, the concentrate from the machine is usually higher grade than that produced from the conventional flotation circuit. The higher feed density allows coarser particles to be floated resulting in an overall coarser concentrate being produced in the Skim-Air than in conventional flotation.

The overall recovery of valuable minerals can be increased through reduced overgrinding. This results because the quick floating liberated valuable particles in the cyclone underflow are removed from the circuit by the Skim-Air and sent directly as final concentrate. If left in the circuit as part of the recirculating load, these particles will be further ground and reduced in size until they become part of the slow floating slimes fraction. At this point they can easily be lost to tailings. Overgrinding and slimes losses are particularly a problem when processing heavy sulphides such as copper, lead and zinc or precious metals such as gold and silver.

In operations where coarse fraction losses are a problem, the use of a Skim-Air allows the ore to be ground finer without fear of overgrinding the valuable mineral. This is particularly helpful in lead/zinc concentrators where a finer grind may liberate more zinc but also increase lead slimes losses.

Figure 2 shows the narrowing of the particle size range being sent to conventional flotation. The difference between the two curves corresponds to a decrease in relative losses of both the slimes and oversize fractions.

When being fed from the cyclone underflow, the Skim-Air is able to produce concentrate with a grade equal to or better than that being produced in conventional flotation. Much of the floatable size gangue in the hydrocyclone feed passes with the water to the cyclone overflow and on to conventional flotation. This results in a feed to the Skim-Air which is very low in floatable gangue. Coupling this with a particle residence time in the machine of 1-2 minutes which allows only liberated values to float, enables the Skim-Air to produce a high grade concentrate.

design flotation plant

design flotation plant

The flowsheet was based on laboratory tests wherein troublesome factors were eliminatedahead of design and construction. The flowsheet provides for unit arrangement of equipment and for added flexibility. Two-stage closed circuit crushing (with an apron feeder for control to the jaw crusher), provides ore for grinding circuit. Crushed material is conveyed to a Screen and the oversize is returned to the secondary cone crusher. The screened fraction drops to a reversible conveyor, thence to fine ore bins.

Adjustable stroke belt ore feeders regulate the feed to two 5 x 10 Steel Head Ball Mills in closed circuit with Cross-flow Classifiers. Each classifier discharge flows by gravity to two banks of 6 Cell No. 18 Sp. Sub-A Flotation Machines. The grind, 65 mesh is held as coarse as possible to reduce grinding costs and still attain maximum recovery. Each flotation section provides six cells for roughing and three cleaning stages with provisions for elimination of one or more stages for cleaning when type of ore permits (flexibility incorporated in Sub-A Machines).

The flotation concentrates are pumped to a regrind circuit to produce desired final size meeting specifications. The final ground concentrates are thickened and filtered. Filters are directly above concentrate storage bins.

The mill site near the mine is accessible to water, power, labor and supplies, and includes adequate space for expansion and tailings disposal. The topography makes gravity flow through the mill possible and allows for delivery of ore direct from mine cars to mill.

Equipment selected gives simplicity and flexibility in operation and allows for changes in tonnages and character of ore. Due to class of labor available, complicated controls and adjustments were eliminated where possible. The machinery selected and installed permits duplication of units for expansion.

As field welding was not available, bolted steel ore bins were installed. The machinery foundations, building foundations, retaining walls and floors are concrete. The flotation machines are mounted on low piers to allow for drainage. An operating platform of wood was installed between flotation machines to give clear working space. The flotation machine launders were designed to permit changes in cleaning stages without shut-downs or prolonged delays, and permits circuit adaptation to changes in ore characteristics. A tailings thickener permits partial reclaiming of water, in dry seasons.

The correct design of a milling plant can mean its success or failure when in operation, the difference between profit and loss. Maximum metallurgical results with low operating and maintenance costs requires thorough study and sound planning. Selection of equipment and construction features must be balanced with available finances and a minimum sacrifice in operating efficiency. Here is a typical small plant where proper design resulted in a successful operation.

This 75-ton, lead-zinc-gold-silver mill was based on a flowsheet developed through batch and continuous laboratory tests. These studies showed single stage crushing and grinding to 65 % minus 65 mesh was adequate for this operation. Tests indicated that over 70% of the gold, 40% of the silver and 60% of the lead was recoverable in the grinding circuit. Therefore Unit Cell and a Mineral Jig section were installed. Adequate flotation capacity to selectively float the lead and zinc was provided, together with a small concentrating table to visually show results of flotation. The zinc and lead concentrates are pumped v direct to a 4-foot by 4-disc Filter with two compartments. This filter was placed on top of the concentrate bin which location provides desirable operating room around the filter and could be seen from almostanywhere in the mill. The concentrate bins were of laminated wood constructionsturdy and inexpensive. Filtered concentrates drop directly into the bins.

Thickeners were eliminated due to initial installation expense and extra housing required because of climate. Low dilution, by keeping sprays in flotation launders to a minimum and ability of Vertical Pump to handle frothy pulps makes this possible.

The mill site was selected several miles from the mine at a point where water, power and tailing disposal area were available, and where it was accessible even during the heavy snow season. The site is on a natural slope, permitting gravity flow in the mill with minimum pumping requirements.

Equipment is arranged as compactly as possible without crowding and without sacrifice of working space, in order to keep the mill building to minimum size. This keeps capital investment low, and reduces heating costs during cold weather. The mill building, of wood construction with laminated wood roof trusses with one-quarter pitch, was covered with insulating material and corrugated sheet material. This construction was most suitable due to the climate and

low cost of lumber in the area. A small steam boiler and unit heaters were installed for heating. Mill and crushing buildings are on concrete foundations extending about four feet below the ground line. Concrete floors of 4 to 6 thick are sloped per foot, permitting slushing down with a hose.

Crushed ore conveyor is enclosed in a conveyorway for weather protection. Fine ore bin is housed within the mill building to prevent freezing. An inclined belt feeder with variable speed drive gains elevation to grinding mill. Feeder was designed with sloping hopper to reduce load on the belt of feed discharging from the bin.

Foundation for Ball Mill was made of reinforced concrete and cast in one section to prevent distortion and misalignment due to possible settlement or shifting of the foundation. Ball mill, unit flotation cell, jig, and classifier were arranged for easy access and operation. The ball mill was equipped with a spiral discharge screen to remove oversize material ahead of unit cell and jig. The 30 spiral classifier was equipped with a rotating motor-driven paddle to remove troublesome wood chips from classifier overflow screen.

The two six-cell No. 18 (2828) Sub-A Flotation Machines were elevated on timber bents with operating platforms between the machines; this gave space belowmachines for pipe lines, launders and concentrate pumps. Reagent feeders were grouped together above flotation machines and conditioners at elevation of filter floor for gravity flow of reagents, and for accessibility.

The mill control office located in the center of the mill, was designed with large windows so almost every machine in the mill could be seen. Operating floors most frequently used were kept as nearly as possible on the same level to reduce stair climbing for the operators. Mill was designed so two men per shift could handle this plant very well.

A gallery was provided in the trussed-roof section, the length of the building, for the installation of main electrical circuits, safety switches, and magnetic motor controls. This kept most of the electrical items away from splash and dirt. All wiring was selected oversize to reduce voltage drop, giving higher operating efficiency and reduced electrical maintenance. Totally enclosed motors were used for reduced maintenance. Push button start-stop controls were placed at the machines and in the mill control office, so that any machine could be controlled from either place.

A typical problem confronting a mining operation of moderate production is how to design a mill at a reasonable cost incorporating modern equipment and essential basic principles of materials handling with the minimum construction and mill costs.

The first step in mill design is the flowsheet based on reliable ore tests. The mill capacity and equipment sizes as shown has been selected as an example for treating 500-550 short tons of ore per 24 hours per day. Two-stage grinding is to all minus 65 mesh for an average ore. Sufficient flotation capacity is included for a medium to slow floating ore. Thickening and filter capacity is selected for a 10 to 1 ratio of concentration as would be the case when treating a 3% copper ore with the copper mineral being chalcopyrite. In such case it would be necessary to filter 50 to 55 tons of concentrates each day. The use of a mineral jig or flotation unit cell in the grinding circuit is recommended. A simple test in our laboratory can tell you whether a coarse product can be recovered easily in the mill circuit.

The general design will apply to other ores with slight modification. The arrangement provides for ultimate use of gravity flow as is noted by the absence of pumps and elevators. The basic machines in plan and elevation are shown along with a flowsheet of the crushing,grinding and mill recovery circuits.

Mine run ore is fed to the primary jaw crusher by a heavy duty apron ore feeder over a grizzly. Crushed ore from the primary crusher is fed over a vibrating screen ahead of the cone crusher to remove fines. The crushing plant is normally designed to crush the entire daily mill tonnage in one shift or, at the most, 2 shifts.

Two- stage grinding provides the grinding economies outlinedin DECO Bulletin B2-B13. In the wet grinding circuit, a rod mill takes the entire feed at and reduces it to approximately 14 to 20 mesh. This mill is normally operated in open circuit with the classifier and ball mill. Usually there is a power saving with this grinding arrangement and often a substantial saving in the cost of the entire mill can be effected by reducing to a minimum some of the requirements in the crushing plant due to this method.

The ground ore overflows the classifier at -65 mesh and approximately 25% solids and is shown being conditioned ahead of flotation. Two parallel banks of Sub-A Flotation Machines on the same floor level are shown for roughing, scavenging, cleaning and recleaning. This arrangement in the flotation circuit provides maximum flexibility in the flow of material, high grade selective concentrates, and low final tailings.

Normally 10 square feet of thickener area is provided for each ton of concentrates per 24 hours which gives reserve capacity to accommodate normal filter maintenance without shutting down the flotation circuit.

In thedesign of any milling operation, continuity of flow should be given first consideration and all weak links eliminated. The old saying an hours delay means no profits today is even more important in our modern milling circuits where labor costs are high.

Many typical design plans and flowsheets are available for your use. Templates of all basic machines, scaled to 1-foot in plan and elevation facilitate laying out these plants. Free tests are made by the Laboratory to check your grinding, thickening and filtering requirements.

If you have a mill design problem, large or small, it will pay you to consult with us. We want to help your engineers in their design work. This service will enable your engineers to lay out your mill at the millsite thus saving design, construction and operating expense. Your completely designed basic plant may already be available in our files with only minor changes necessary to modify it to fit your specific application.

silver lead zinc ore processing method using flotation

silver lead zinc ore processing method using flotation

Sulphide ore of lead and zinc containing considerable silver was submitted for testing with the purpose of determining a flowsheet for the production of separate lead and zinc concentrates for marketing at their respective smelters. It is necessary to recover as much silver as possible in the lead concentrate as a higher return for this silver is realized than for the silver in the zinc concentrate. The ore contained sphalerite, a portion of which was easily floatable but difficult to depress in the lead flotation circuit.

Also, the recovery of silver minerals occurring in a lead, zinc sulfide ore is efficiently accomplished using Flowsheet #2. The process consists of selective flotation to produce a mixed silver, lead concentrate for maximum smelter return and a separate zinc concentrate. Over-grinding of silver minerals is detrimental to efficient flotation recovery, so the Flash Flotation Unit-Cell is used in the grinding circuit to recover a large part of the silver and lead values as soon as liberated.The flowsheet is for a plant having a capacity in the range of 300 to 500-tons per day.

The crushing section of this 50-65 ton mill consists of a conventional layout of single stage crushing. The mine ore is fed from the coarse ore bin to a 9x 16 Forced Feed Jaw Crusher by means of a Apron Ore Feeder. The crushed ore is conveyed by a Belt Conveyor to the Bolted Steel Fine Ore Bin. A Adjustable Stroke Unit Flotation Cell are incorporated in the Belt Ore Feeder delivers the fine ore to the ball mill.

The Mineral Jig and the grinding circuit for immediate recovery of a substantial amount of the lead and silver at a relatively coarse grind. The 5 x 5 Steel Head Ball Mill discharges into an 8x 12 Selective Mineral Jig which in turn discharges into a small flashFlotation Cell. The tailings from the Unit Cell flow by gravity to the 30 Cross-Flow Classifier. The Mineral Jig and the flashcell treating an unclassified feed, produce high-grade concentrates of lead and silver with a minimum amount of zinc. Recovery of these important amounts of lead and silver at this point not only prevents detrimental sliming of the lead mineral and possible subsequent loss, but also increases the amount of new feed that can be fed to the ball mill. By taking advantage of recovering a clean product representing a high recovery of the lead leaves only a small amount of the lead to be recovered in the selective flotation section.

This section of the flowsheet uses two 6-cell (32 x 32) Flotation Machines. The classifier overflow is fed by gravity to the first rougher cell of the lead machine. Three rougher cells provide ample contact time for the flotation of the lead. This rougher lead flotation concentrate is then delivered by gravity to the cleaner cells. Three cleaner cells are used for triple cleaning of the lead concentrate. This triple cleaning was recommended because of the easily floatable zinc that could not be effectively depressed by conventional zinc depressant reagents. Roughing, plus triple cleaning in a 6-cell machine with no pumps or elevators is an example of flexibility a distinctive feature of Sub-A Flotation Machines.

The lead circuit tailing is then conditioned with reagents in a 6 x 6 Super-Agitator and Conditioner prior to zinc flotation. The conditioned pulp is then floated in a 6-cell No. 18 Special Sub-AFlotation Machine for the production of a cleaned zinc concentrate. This machine is arranged for four rougher cells and two cleanings of the rougher zinc concentrate.

Soda ash and zinc sulphate are fed to the ball mill by means of Cone Type Dry Reagent Feeders. Cyanide, sodium sulphite, MIBC frotherand xanthate (Z-3) are fed to the grinding circuit and lead flotation circuit using a multi-compartment Wet Reagent Feeder. Lime and copper sulphate (CuSO4) are added to the zinc conditioner and pine oil and xanthate (Z-5) are stage added to the zinc rougher circuit using Wet Reagent Feeders.

The Visual Sampler, consisting of a Suction Pressure Diaphragm Pump and a No. 13A Wilfley Concentrating Table, takes a portion of the final zinc tailing. This unit enables the operator to determine visually the results of flotation. Any necessary change of reagents is immediately indicated by observation of the concentrate streak shown on the table. Many installations of the Visual Sampler have proved this unit to be a money-saving necessity in any flotation plant.

Thickening, prior to filtration, was not recommended in this case because of the rapidity at which these concentrates filtered and the relatively small tonnage of this mill. Thickening is advisable on slower filtering ores and on larger tonnages.

The final lead-silver concentrates (including the Flash FlotationCell concentrate) are filtered on the 44-disc Filter, the filter cake discharging directly into concentrate bins. The dewatered Mineral Jig concentrate is combined with the filtered lead concentrate in the storage bin.

The above flowsheet incorporates the first rule of milling procedurerecover the mineral as soon as freedthis is accomplished by the Jig and Flash Unit Cell in the grinding circuit. Note that a high-grade lead product representing 2/3 of the total lead (very low in zinc), is recovered in the grinding circuit. This flowsheet successfully answers The Problem by recovering 84% of the total silver in the lead concentrate.

The recovery of silver minerals occurring in a lead-zinc sulfide ore is efficiently accomplished using the above flowsheet. The process consists of selective flotation to produce a mixed silver-lead concentrate for maximum smelter return and a separate zinc concentrate. Over-grinding of silver minerals is detrimental to efficient flotation recovery, so the Flash Flotation Unit-Cell is used in the grinding circuit to recover a large part of the silver and lead values as soon as liberated.The flowsheet is for a plant having a capacity in the range of 300 to 500-tons per day.

The crushing section consists of primary and secondary crushing with intermediate screening. Both crushers are located in the same building and conveniently attended by one operator. A minimum of conveying equipment is required by this arrangement. Dust collecting facilities are, likewise, limited to only one building.

The crushed ore after automatic sampling is subjected to two-stage grinding using a Rod Mill in open circuit and a Ball Mill in closed circuit with a Classifier. TheUnit Flotation Cell receives the discharge from the ball mill for recovery of a substantial amount of the granular silver minerals together with galena as soon as freed. Reagents are added to the ball mill. Tramp iron and occasional oversize gangue are removed from the circuit by the Spiral Screen attached to the ball mill and this prevents excessive wear or plugging of the unit cell. The classifier is of the latest design.

The Mineral Jig is not included in the flowsheet, but on many ores of this type it is applicable either alone or with the unit cell. The grade of jig concentrate is usually very high grade and ideal for blending with the flotation concentrate. If native silver or gold values are present, the jig is a very essential addition to the flowsheet and would be used on the rod mill discharge in this case.

The classifier overflow is treated in a conventional manner using Sub-A Flotation Machines of cell-to-cell design which enables double cleaning of the silver-lead and zinc concentrates without the need of pumps. For large tonnage operations the Sub A Free Flow Machine is optional for roughing and scavenging, but the cell to cell type is always used in the cleaner circuits where high selectivity is essential. The two flotation banks are arranged so that the banks face one another and can be conveniently controlled by one operator from a single aisle. Operation of the Conditioner can also be observed from this aisle. A Sampler is used on the zinc tailing to provide an instant means for the operator to evaluate plant results. Some plants find it beneficial to use a visual sampler on the lead tailing ahead of the zinc circuit. The Sampler is also useful for evaluating the lead or zinc concentrate.General view of the flotation section at a modern silver-lead-zincmill. The lead circuit is on the left and the zinc circuit ison the right.

The silver-lead concentrate (including the unit cell concentrate) and the zinc concentrate are separately treated through wet cyclones to remove the coarse sulfides as thick underflow products suitable for direct filtration. The cyclone overflow products are ideally suited for thickening and subsequent filtration with their respective cyclone underflows. This procedure avoids any overload of heavy sulfides in the thickeners and, therefore, simplifies the operation of the thickeners. SRL Pumps are engineered for use with wet cyclones and give trouble-free service.

In addition to the feed sample, which is cut by means of a Type C Automatic Sampler, the final silver-lead and zinc flotation concentrates are sampled using Type B cutters. The final plant tailing is also sampled in the same manner.

This flow-sheet incorporates all features of a modern day mill for optimum efficiency and general simplicity for ease of operations. Instrumentation devices can be included to facilitate automatic control of the plant circuits if desired.

Many factors affect the metallurgical results of every plant. However, in a study of this type it is interesting to note the recoveries and grades that are actually being made at successful mills. The figures of these two plants are included for their value in making economic studies of new deposits.

how antimony is processed by flotation

how antimony is processed by flotation

The problem discussed in this antimonyprocess study is limited to a concentrator capable of beneficiating 150 tons per day of antimony ore. The antimony in this study occurs as the mineral stibnite (Sb2S3) in association with small amounts of pyrite, arsenopyrite, galena and lead sulfantimonides. The gangue is composed largely of quartz but contains, in addition, a small amount of talc. The talc mineral is particularly trouble some since it tends to float with the stibnite and hence lowers the grade of the final antimony concentrate.

The crushing section with two-stage reduction is suitable for tonnages of from 100 to 300 tons of ore per 24 hours. It is designed to permit crushing sufficient ore during one 8 hour shift to operate the mill during a full 24 hours. Removing the fines with a grizzly before the jaw crusher and with a vibrating screen ahead of the secondarycrusher increases the efficiency of size reduction since thecrushers are working only on material that must be reduced. Removal of tramp iron to prevent damage to the secondary crusher is accomplished with an electromagnet and a magnetic head pulley.

The crushed ore is fed from the fine ore storage bin at a controlled rate by a Adjustable Stroke or Variable Speed Ore Feeder. Fine grinding is accomplished in a Ball Mill operating in closed circuit with a Spiral Classifier. To reduce overgrinding of the friable stibnite it is generally desirable to carry a relatively large circulating load in the ball mill-classifier circuit. If necessary sufficient lime or soda ash should be added to the ball mill feed to provide the proper pH for subsequent conditioning and flotation operations.

A head-ore sample is cut from the classifier overflow to use as the flotation circuit control and metallurgical balance. A Automatic Sampler is used. Normally the sampling interval is one cut each 15 minutes which is collected each shift as a composite sample. A Unit Flotation Cell has not been included in the ball mill-classifier circuit of this study. However, in certain instances a Unit Flotation Cell may prove particularly effective in that it is capable of recovering a significant portion of the total stibnite as a relatively high grade grannular concentrate. Thus, overall plant recovery may be increased by reducing slime losses due to overgrinding of the soft, high specific gravity stibnite.

The classifier overflow adjusted to a pH of approximately 7.5 to 7.8 is conditioned with a stibnite activator such as copper sulfate or lead acetate and a collector such as Z-11 or a liquid type Aerofloat. Depending upon the mineralogy of the ore, other auxiliary reagents may also be required to depress sulfides which tend to float with and thus lower the grade of the stibnite concentrate. For example, bleaching powder has been used effectively for the depression of arsenopyrite and cyanide for the depression of pyrite.

The conditioner overflow passes to an 8-cell Sub-A Flotation Machine, arranged to provide four rougher cells, two scavenger cells and one cell for each of two stages of cleaning. The distinctive cell-to-cell design of Sub-A Flotation Machines permits the transfer of the various pulp streams within the flotation circuit without the use of pumps. Extremely fine talcose slimes which floated with the stibnite in the rougher cells are effectively depressed in the cleaners with the aid of a small quantity of yellow dextrine.

The cleaned antimony concentrate is pumped to a Spiral Rake Thickener to remove excess water prior to filtering. A Adjustable Stroke Diaphragm Pump meters the thickened flotation concentrate to a Disc Filter.

For the efficient control of milling operations, reliable sampling is required. For this purpose Automatic Samplers are used on the heads, concentrates and tailings. A metallurgical balance is run for each shift. Sampling interval of one cut each 15 minutes has proved adequate except where quality of ore is spotty or unusual conditions are encountered.

The beneficiation of stibnite ores by flotation can frequently be a difficult problem because of the complex nature of the mineral associations often encountered. It is essential, therefore, that a comprehensive laboratory test program be carried out on a representative sample of the ore before planning a milling installation.

While antimony concentrates assaying less than 50% antimony may be salable, such low-grade concentrates require special arrangements with thesmelters. Smelter returns for antimony concentrates are based on a sliding scale which provides for 65% Sb content as the basis. Antimony payment is based on a short ton unit containing 20 lbs. antimony. A 1963 quotation for 65% antimony concentrate was $4.25 per short ton unit.

copper flotation

copper flotation

Although basic porphyry copper flotation and metallurgy has remained virtually the same for many years, the processing equipment as well as design of the mills has continually been improved to increase production while reducing operating and maintenance costs. Also, considerable attention is paid to automatic sensing devices and automatic controls in order to assure maximum metallurgy and production at all times. For simplicity in this study most of these controls are not shown.Many of the porphyry copper deposits contain molybdenite and some also contain lead and zinc minerals.

Even though these minerals occur in relatively small amounts they can often be economically recovered as by-products for the expense of mining, crushing, and grinding is absorbed in recovery of the copper.

Because the copper in this type of ore usually assays only plus or minus 1% copper, the porphyry copper operations must be relatively large in order to be commercial. The flowsheet in this study illustrates a typical 3,000 ton per day operation. In general most operations of this type have two or more parallel grinding and flotation circuits. For additional capacity, additional parallel circuits are installed.

The crushing section consists of two or three crushing stages with the second or third stages in either closed or open circuit with vibrating screens. Generally, size of the primary crusher is not determined by capacity but by the basic size of the mine run rock. The mine-run ore is normally relatively large as most of the porphyry mines are open pit.The crushing section illustrated is designed to handle the full tonnage in approximately 8 to 16 hours thus having reserve capacity in case of expansion.

Many mills store not only the coarse ore but also the fine ore in open stockpiles using ore as the side walls and drawing the live ore from the center. During prolonged periods of crusher maintenance the ore walls can be bulldozed over the ore feeders to provide an uninterrupted supply of ore for milling.

As it is shown in this study the or 1 crushed ore is fed to a rod mill operating in open circuit and discharging a product approximately minus 14-mesh. The discharge from this primary rod mill is equally distributed to two ball mills which are in closed circuit with SRL Rubber Lined Pumps and two or more cyclone classifiers. The rod mill and two ball mills are approximately the same size for simplified maintenance.

Porphyry copper ores, usually medium to medium hard, require grinding to about 65-mesh to economically liberate the copper minerals from the gangue. Although a clean rougher tailing can often be achieved at 65-mesh the copper mineral is not liberated sufficiently to make a high grade copper concentrate, thus some form of regrinding is necessary on the rougher flotation copper concentrate. It is not unusual to grind the rougher flotation concentrate to minus 200-mesh for more complete liberation of mineral from the gangue.

The cyclone overflow from each ball mill goes to a Pulp Distributor which distributes the pulp to two or more parallel banks of Flotation Cells. These distributors are designed so that one or more flotation banks can be shut down for maintenance or inspection and still maintain equal distribution of feed to the remaining banks.

In some cases it is beneficial to have conditioning before flotation, but this varies from one operation to another and it is not shown in this flowsheet. Ten or more Free-Flow Flotation Cells are used per bank and these cells are divided into groups of four or six cells with an intermediate step-down weir between groups. Free-Flow Flotation Cells are specified, as metallurgy is extremely good while both maintenance and operating expenses are traditionally low. One or more Free-Flow Mechanisms can be stopped for inspection or even replaced for maintenance without shutting down the bank of cells.

The concentrates from rougher flotation cells are sent directly to regrind. Often the grind is 200-mesh. After regrind is flotation cleaning. In some cases the concentrate from the first three or four rougher flotation cells can be sent directly to cleaning without regrinding.

After the rougher flotation concentrate is reground it is cleaned twice in additional Free-Flow Flotation Machines with the recleaned concentrate going to final concentrate filtration or, as the metallurgy dictates, to a copper-moly separation circuit.

The thickening and filtering is similar to other milling operations, however, as the porphyry copper installations are often in arid areas, the mill tailing is usually sent to a large thickener for water reclamation and solids go to the tailings dam.

Automatic controls are usually provided throughout modern plants to measure and control pulp flow, pH and density at various points in the circuit. Feed and density controls are relatively common and the newer installations are using automatic pulp level controls on flotation machines and pump sumps. Automation is also being applied to the crushing systems.

The use of continuous on stream X-ray analysis for almost instantaneous metallurgical results is not shown in thus study but warrants careful study for both new and existing mills. Automatic sampling of all principal pulp flows are essential for reliable control.

The flowsheet in this study illustrates the modern approach to porphyry copper treatment throughout the industry. Each plant will through necessity have somewhat different arrangements or methods for accomplishing the same thing and reliable ore test data are used in most every case to plan the flowsheet and design the mill.

In most plants engaged in the flotation of ores containing copper-bearing sulphide minerals with or without pyrite, pine oil is employed as a frother with one of the xanthates or aerofloat reagents or a combination of two or more of them as the promoter. Lime is nearly always used for maintaining the alkalinity of the circuit and depressing any pyrite present. The reagent consumption is normally within the following limits

While good results are often obtained with ethyl xanthate alone as a promoter, the addition of a small quantity of one of the higher xanthates is frequently found to improve the recovery of those minerals that are not readily floated by the lower xanthate, especially those that are tarnished or oxidized, but since the action of a higher xanthate is, as a rule, more powerful than that of the ethyl compound, it is usually best to add no more of the former reagent than is necessary to bring up the less readily floatable minerals, controlling flotation with the less powerful and more selective lower xanthate. Better results are obtained with some ores by replacing the higher xanthate with one of the dithiophosphates, flotation being controlled, as before, with ethyl xanthate. Sometimes a dithiophosphate can be effectively used without the xanthate, although the dual promotion method is more common. A rule of thumb system for the selection of these reagents cannot be laid down as the character of the minerals differs so widely in different ores ; the best combination can only be found by experiment.When aerofloat is employed alone as the promoter, the reagent mixture is somewhat different from that given above. A reliable average consumption is difficult to determine as the plants working on these lines are few in number, but the following is what would normally be expected.If this combination of reagents gives results equal to those obtainable with a xanthate mixture, its employment has these advantages over the latter method: The control of flotation is not so delicate as with xanthates, it has less tendency to bring up pyrite, and, if selectivity is not required, the circuit may be neutral or only slightly alkaline.

When the ore is free from pyrite, the function of the lime, whatever the reagent mixture, is to precipitate dissolved salts and to maintain the alkalinity of the pulp at the value which has been found to givethe best results ; soda ash is seldom employed for this purpose. When pyrite is present, lime performs the additional function of a depressor, the amount used being balanced against that of the promoterthat is, no more lime should be added than is required to prevent the bulk of the pyrite from floating, as any excess tends to depress the copper minerals, and no more of the promoter should be employed than is needed to give a profitable recovery of the valuable minerals in a concentrate of the desired grade, since any excess tends to bring up pyrite. In many cases a more effective method of depressing pyrite is to add a small quantity of sodium cyanidee.g., 0.05-0.10 lb. per tonin conjunction with lime, less of the latter reagent then being necessary than if it were used alone.

It is not often that a conditioning tank has to be installed ahead of the flotation section in the treatment of sulphide copper ores, as the grinding circuit usually provides suitable points for the introduction of the reagents. The normal practice is to put lime into the primary ball mills and to add xanthates at the last possible moment before flotation, while aerofloat and di-thio-phosphates are preferably introduced at some point in the grinding circuit, since they generally need an appreciable time of contact as compared with xanthates. There is no special place for the addition of pine oil, but care should be taken if it is put into the primary ball mills, as a slight excess may cause an undue amount of froth to form in the classifiers.

In a plant where the primary slime is by-passed round the grinding circuit, it is necessary to ensure that this portion of the pulp receives its correct proportion of and contact time with the reagents.

As regards flotation installations, the present tendency is to employ machines of the air-lift or Callow-Maclntosh rather than of the subaeration type. While two stages of cleaning (circuits 10 and 11) are sometimes essential to the production of a clean final concentrate, circuits 8 and 9 comprising a single stage of cleaning are probably the most widely used. Occasionally the primary machines can be run as rougher-cleaner cells (circuit No. 5), particularly when they are of the air-lift or subaeration type. This method, however, is not often employed, although its use is more common in the flotation of copper sulphide minerals than of any other class of ore ; a stage of cleaning is preferable as providing greater lattitude of control.

Two variations of normal procedure are worth notice. In one or two plants employing two-stage grinding, improved results have been obtained by separating the slime from the primary ball mill circuit and sending it direct to a special flotation section. This method is useful when the feed to the flotation plant contains an appreciable quantity of fines, which, due generally to oxidation through exposure, require different treatment from the unweathered part of the ore. Such fines are usuallyfriable and can be separated as slime from the primary grinding circuit without the inclusion of an undue proportion of unoxidized material, the bulk of which thus passes to the secondary grinding circuit and thence to its own division of the flotation plant.

The second variation consists of grinding the rougher concentrate before cleaning. The method is applicable to an ore in which the copper- bearing minerals are so intimately associated with pyrite that very fine grinding is necessary to liberate them completely. It is often possible, after grinding such an ore to a comparatively coarse mesh, to make a profitable recovery of the copper in a low-grade concentrate which does not represent too large a proportion, say 30% or less, of the total weightof the feed. The concentrate can then be reground and refloated with the production of a high-grade copper concentrate together with a low- grade pyritic tailing suitable for return to the roughing circuit. This method is likely to be less costly than one involving the fine grinding of the whole ore. No standard system can be given for handling the various products as their disposal depends so much on the occurrence of the minerals and the efficiency of the regrinding operations, but a typical flow sheet is illustrated in circuit No. 12 (Fig. 60). It is diagrammatic to the extent that the thickener and regrinding unit may receive its feed from several roughing machines and deliver its discharge to a number of cleaning cells. It is usual to dewater the rougher concentrate and return the water to the primary circuit for two reasons : First, to supply the regrinding mill with a thick enough pulp for efficient operation, and, secondly, as far as possible to prevent the reagents used in the roughing circuit from entering the cleaning section.

In normal practice a recovery of over 90% of the copper which is present as a sulphide is generally possible, whatever the flotation process or circuit employed. As regards the average grade of concentrate, no more can be said than that it depends on the class of the copper-bearing minerals present and their mode of occurrence and on the character of the gangue. It usually contains over 20% of copper, but a difficult chalcopyritic ore may yield a concentrate with less than that percentage, while it is theoretically possible to obtain one running over 75% should the mineral consist entirely of pure chalcocite.

The flotation of native copper ores is nearly always preceded by gravity concentration in jigs and tables not only because the combined process is more economical as regards costs, but also because the copper often occurs as large grains which flatten out during grinding and cannot be broken to a size small enough for flotation. The flow sheet depends on the mode of occurrence of the mineral. The tailings from some of the gravity concentration machines may be low enough in value to be discarded, but those products which still contain too much copper to be sent to waste are thickened and reground, should either operation be necessary, and then floated with pine oil and a xanthate or aerofloat reagent in a neutral or slightly alkaline circuit. The reagent consumption is approximately the same as that given for the treatment of copper- bearing sulphides. While a pine oil, lime, and ethyl xanthate mixture has proved satisfactory, better results have sometimes been obtained by the substitution of aerofloat and sodium di-ethyl-di-thio-phosphate, soda ash being used instead of lime on account of its gangue deflocculating properties. On the average 0-12 lb. per ton of aerofloat and 0.03 lb. of the di-thio-phosphate are substituted for 0.1 lb. of xanthate.

Since a high-grade concentrate is desired in order to keep smelting costs as low as possible, the circuit usually comprises two stages of cleaning. In most plants flotation is carried out in mechanically agitated machines.

The problem of the flotation of oxidized copper ores has not yet been solved. One or two special processes are in operation for the flotation of malachite and azurite, but none of them has more than a limited application; nor has any method been worked out on a large scale for the bulk flotation of mixed oxidized and sulphide copper minerals when the former are present in the ore in appreciable quantity.

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.

home - crushers, ball mills and flotation cells for mining and mineral beneficiation

home - crushers, ball mills and flotation cells for mining and mineral beneficiation

Founded in 1987, ZJH is mainly focus on producing and supply crushers,ore grinding equipment, mineral beneficiation equipment, laboratory and pilot scale ore dressing equipment for Mines and Mineral Beneficiation Plants.Our aim is to work together with the Mining and Mineral Processing Industry for helping to carry on the production technical innovation, to reduce the operating cost ,to improve the operating efficiency.

Mining thickener is mainly used for dewatering the wet concentrate during the ore dressing process. Our thickener is mostly located between cleaning beneficiation process and filtration equipment. Thickener is applied to both the concentrate and tailings to recover water. The thickener could be used to recover immediately reusable water back to mineral processing plant, as []

Complete set of Graphite Beneficiation Equipment usually includes jaw crusher, ball mill, classifier, rod mills, agitation tanks, flotation machine, rotary dryers etc. ZJH minerals as more than 30 years of professional mineral beneficiation equipment manufacturers, according to customers requirementsand actual situation, can provide ore testing, ore dressing experiments, process design, a complete set of equipment []

Asphalt from old road surface is elastic and include hard stone. sizing crusher is best choice for crushing Asphalt from old road surface. 1. The crushing work conditions: Raw material: asphalt from old road surface The feeding size: 400*400*100mm The discharged size: less than 16mm Capacity: 400t/h 2. Solutions with 2 stage crushing Primary sizing: []

Why the mineral beneficiation plant shall build a mineral processing laboratory? The ore characteristics is always changed as the mining stage different, so the present meniral beneficiation method or process do not recover the aimed mineral well. At this time, the mineral beneficiation plant shall observe the structure of the ore and analyze the nature []

Our semi-industrial flotation plant with capacity 1-3 t/d is mainly designed for semi-industrial scale test of continuous flotation. The features of mobile flotation pilot plant 1.beneficiation reagent dosing machine+ agitation tanks + flotation cells are formed a pilot flotation plant, which installed in a container for easy moving and transportation by truck. 2. easy operation. []

we recommend you adopt ourour lab jaw crushersZJEP-10060withzirconia ceramic jaw plates. itisapplicable for crushing high purity materials, avoid mixing other elements, no metal ion pollution, ensure the high purity and cleanliness of materials crushing 1, Technical data : 2, Application: our lab jaw crushers with zirconia jaw plates is applicable for crushing high purity materials, []

Quartz sand is a kind of non-metallic mineral.Its main mineral composition is silica (SiO2). Quartz sand is made by quartz ore crushing, screening, washing and other processes. its hard, wear-resisting, chemical properties of stability is widely used in glass, casting, ceramics and refractory materials, smelting ferrosilicate, metallurgical flux, metallurgy, construction, chemical, plastic, rubber, abrasive and []

Polymetallic sulfide ore dressing by flotation The copper, lead, zinc polymetallic sulfide ore has a wide variety of minerals, copper, lead, zinc sulfide. The minerals are closely symbiosis. The dissemination particle size is very uneven. It is easy to float and difficult to separate ore. According to the characteristics of the metal ore, the grinding []

The main component of bauxite is alumina. Bauxite beneficiation by flotation process can be roughly divided into washing primary beneficiation crushing grinding classification separation concentrate concentrationfiltration,several processes. Bauxite beneficiation crushing commonly used three-stage one-closed circuit crushing process;Grinding and Classification: Grinding often adopts grid type ball mill with closed circuit []

Magnetite Beneficiation or ore dressing production line combined by vibrating feeder, jaw crusher, vibrating screen, ball mill, classifier, magnetic separator, thickener and dryer and other main equipment. With the feeder, hoist, conveyor can form a complete ore dressing production line.The Magnetite Beneficiation production line has the advantages of high efficiency, low energy, high handling capacity []

The Lead and Zinc ore dressing, according to the different types of ore, then choose different ore dressing methods, also need different lead-zinc ore beneficiation equipment. Sulfide ores are usually flotation method.Oxidized ore is beneficiated by flotation or gravity separation combined with flotation, or flotation after curing roasting, or flotation after gravity separation with sulfuric []

Froth Flotation Beneficiation process is one of the important processes in the application of ore dressing, and it is widely used. Froth Flotation Beneficiation is widely used in copper, nickel ore, iron ore, gold ore, lead and zinc ore, potassium feldspar, graphite and other metal and non-metal ore concentrate selection, with high efficiency, low energy, []

Complete set of potassium feldspar beneficiation equipment usually includes jaw crusher, ball mill, classifier, magnetic separator, agitation tanks, flotation machine, high gradient magnetic separation machine etc. ZJH minerals as more than 30 years of professional mineral beneficiation equipment manufacturers, according to customers requirements and actual situation, can provide ore testing, ore dressing experiments, process design, []

Froth flotation of gold is a widely used beneficiation method for treating rock gold ore in gold concentrating plant, which is often used to treat gold ore containing sulfide minerals with high floatability. The froth flotation process can concentrate the gold in sulfide minerals to the maximum extent, and the tailings can be directly discarded. []

The production line for beneficiation of fluorite is up to the aim of fluorite concentration by the work of a series of equipment with a clear division of job. The main machine for beneficiation of fluorite includes jaw crusher, grate ball mill, overflow ball mill, spiral classifier,agitation tank, flotation cells, thickner, filter,etc. The froth flotation []

we supply Potash feldspar grinding mill. The Potash feldspar grinding mill includes the bin, belt feeder, ball mill, air classifier and bag filter. The Potash feldspar grinding mill has the follow features: high working efficiency environmental friendly The flow chart of Potash feldspar grinding mill The technical specifications of Potash feldspar grinding mill capacity: 4500kg/h []

Slag is a common raw materials for cement industry. The hardness of slag is around 6-7 in Mohs scale (harder than cement clinker). For the aim to grind slag (15-20 mm) to a final fineness of 30 microns. After the slag dry, the slag less than 50mm fed into the rolling mill, the slag will []

Clay is sticky. sizing crusher is best choice for crushing the sticky material. 1. The crushing work conditions: Raw material: clay with 10% of stones The feeding size: 200 The discharged size: less than 100mm Capacity: 300-400t/h Solutions Model: FP 63AS Power: 200KW Weight: 22 ton Dimension: refer to the attached CAD drawing the []

The basic data on the grinding system Raw material1.Calcined Alumina2.Alumina Trihydrate Feeding size200 mesh, Output sizeAlumina Trihydrate:D50=9m-1m;D97=36m-4m Calcined Alumina:D100=325mesh-800mesh Capacity0.5-0.7 T/h(on the basis of Calcined AluminaD100=800mesh),this system could produce other size product. The capacity is different according to different size. This system is also for iron avoidance superfine grinding for other hard materials, for example: []

PE 4080 Mini pollution free Jaw Crusher is mainly for crushing the geological and geochemical rock samples. It can strictly control the contamination of other elements except silicon and aluminum. It worked with our anti-pollution disc grinding mill, by which the whole process of sample preparation can be realized to prevent pollution.

ZJH mainly focus on producing and supply crushers, ore grinding equipment, mineral Beneficiation equipment, laboratory and pilot scale ore dressing equipment for Mining and Mineral Processing Industry. Our aim is to work together with Mines, Mineral Beneficiation Plantsfor helping to reduce the operating cost ,to improve the operating efficiency.

flotation cells - pineer mining machinery

flotation cells - pineer mining machinery

The impeller is rotated by V-belt of motor and produces negative pressure by centrifugal function. Enough air is sucked to mix slurry, and slurry mixes drug at the same time. Mineral sticks on bubble completely and floats on the surface of slurry to form mineralized bubble. Useful bubble is scraped out by adjusting flash-board height and controlling liquid surface.

ball mills - page 2 of 2 - crushers, ball mills and flotation cells for mining and mineral beneficiation

ball mills - page 2 of 2 - crushers, ball mills and flotation cells for mining and mineral beneficiation

Our Ball mills are designed for mining and mineral processing industry, for grinding the ore to the expected size for the next stage beneficiation process. We also supply the laboratory and pilot scale ball mills for mineral beneficiation laboratory test.

JM series stirred ball mill have adopted by the gold ore, copper ore, silver ore,molybdenum ore, lead zinc ore, manganese ore, iron ore, nickel ore, such ore dressing plant for fine grinding or regrinding operations.

ZJH mainly focus on producing and supply crushers, ore grinding equipment, mineral Beneficiation equipment, laboratory and pilot scale ore dressing equipment for Mining and Mineral Processing Industry. Our aim is to work together with Mines, Mineral Beneficiation Plantsfor helping to reduce the operating cost ,to improve the operating efficiency.

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