flotation machine 4 wood

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Woodgrove Technologies is the world leader in Flotation Technology and Advanced Process Control. The Direct Flotation Reactor is a paradigm shift and has taken flotation technology to a new level with an order of magnitude savings for clients on installation and operation and our Progressive Control solution will result in a significantly more stable and efficient circuit.

Woodgrove offers world-class mineral processing and ground-breaking optimization solutions to improve your performance. We serve our customers with the highest quality products and services. Come read more about us!

We have proven it (Woodgrove Technology) is a no brainer, and are the first company to use this on a broad basis. [] We will use it wherever we can as there are so many environmental and economic benefits to it.

The reason for choosing Portage (Woodgrove) associated to the Concentrator control system upgrade was purely about selecting a company that: listened to our needs could provide solutions to achieve results could integrate their system and controls seamlessly into our control system (DeltaV) had depth of knowledge and experience within their team associated to floatation (Metallurgical, Expert Control & DeltaV configuration) delivered on expectation

flotation machines

flotation machines

As pneumatic and froth separation devices are not commonly used in industry today, no further discussion about them will be given in this module. The mechanical machine is dearly the most common type of flotation machine currently used in industry, followed by the column machine which has recently experienced a rapid growth.

A mechanical machine consists of a mechanically driven impeller that disperses air into the agitated pulp. In normal practice this machine appears as a long tank-like vessel having a number of impellers in series. Mechanical machines can have open flow of pulp between the impellers or can be of cell-to-cell design with weirs between them. Below is a typical bank of flotation cells used in industrial practice.

The procedure by which air is introduced into a mechanical machine falls into two broad categories: self-aerating, where the machine uses the depression created by the impeller to induce air, and supercharged, where air is generated from an external blower. The incoming feed to the mechanical flotation machine is usually introduced in the lower portion of the machine. At the very below is shown a typical flotation cell of each air delivery type (Agitair & Denver)

The most rapidly growing class of flotation machine is the column machine, which is, as its name implies, a vessel having a large height-to-diameter ratio (from 5 to 20) in contrast tomechanical cells. This type of machine provides a counter-current flow of air bubbles and slurry with a long contact time and plenty of wash water. As might be expected, the major advantage of such a machine is the high separation grade that can be achieved, so that column cells are often used as a final concentrate cleaning step. Special care has to be exercised in the generation of fine air bubbles and the control of the feed rate to the column cell for such cells to be effective. Column cell use is often of limited value in the recovery of relatively coarse valuable particles; because of the long lifting distances involved, the bubbles can not carry large particles all the way to the top of the cell.

Probably the most significant area of change in mechanical flotation cell design has been the dramatic increase in machine cell volume with a single impeller. The idea behind this approach is that as machine size increases (assuming no loss of recovery performance with the larger machines), both plant capital and operating cost per unit of throughput decrease. In certain industrial applications today, cells of even a thousand cubic meters in volume (a large swimming pool) are being used effectively.

The throughput capabilities of various cell designs will vary with the flotation machines residence time and pulp density The number of cells required for a given operation is determined from standard engineering, mass balance calculations. In the design of a new plant, the characterization of each cells volume and flotation efficiency is generally calculated from data gathered on a laboratory scale flotation using the same type of equipment for the same material mixture in question. This procedure is then followed by the application of semi-empirically derived scale-up factors. Research work is currently under way to improve the understanding and performance of commercial flotation cells.

Currently, flotation cell design is primarily a proprietary material of the various cell manufacturers. Flotation plants are built in multiple cell configurations (called banks), and the flow through the various banks is adjusted in order to optimize plant recovery of the valuable as well as the grade of the total recovered mass from flotation. Up above is a typical flotation bank scheme. The total layout of a given flotation plant (including all of the various banks) operating on a given feed is called a flotation circuit.

The application of the air-lift to flotation is not new, but the first attempts to make use of the principle were not successful because the degree of agitation in the machine was insufficient to enable the heavy oils then in use as collecting reagents to function effectively. The advent of chemical promoters, however, made agitation of secondary and aeration of primary importance, with the result that the application of the air-lift principle became practicable and led to the introduction of the Forrester and the Hunt matless machines. South western Engineering Corporation are the owners in most countries of the rights to license and manufacture these and other types operating on the air-lift principle, and they have developed a machine based chiefly on the Welsh and Hunt patents which may be considered as representative of the type that is now most commonly used.

The Southwestern Air-Lift Machine, as it is called, consists of a V-shaped wood or steel trough of any length but of the standard cross-section shown in Fig. 40, the area of which is 9.85 sq. ft. and the interior depth 36 in. Low- pressure air is delivered from a blower through a main supply pipe to an air-pipe or header which runs longitudinally over the top of the machine. The air enters the trough itself through a seriesof vertical down-pipes , which are screwed into sockets welded tothe underside of the header at 4-in. intervals along its length and are open at their lower ends. They are from to 1 in. in diameter for roughing machines and from to in. for cleaners, and they reach to within 6 in. of the bottom. The air-lift chamber is formed by two vertical partitions, one on each side of the line of down-pipes, both of which extend from one end of the trough to the other, forming a compartment 6 in. wide. The lower edges of the partitions are an inch or two above the ends of the down-pipes and their upper edges are about level with the froth overflow lips at each side of the machine. A few inches above the top of the air-lift chamber is a deflector cap which serves to direct the rising pulp outwards and downwards against two vertical baffles. These extend the length of the trough parallel to and outside the partitions, their loweredges being several inches below the normal pulp level. The spacebetween the baffles and the sides of the machine forms two spitzkasten- shaped zones of quiet settlement where the froth collects.

The feed enters near the bottom of one end of machine and the tailing is discharged over an adjustable weir at the other end. The air, issuing in a continuous stream from the open ends of the down-pipes, carries the pulp up the central chamber on the principle of an air-lift pump. The air is subdivided into minute bubbles and more completely mixed with the pulp as the rising mass hits the cap at the top and is deflected and cascaded on to the baffles at each side, which direct it downwards, distributing the bubbles evenly throughout the pulp in the body of the machine and giving them ample opportunity to collect a coating of mineral. Rising under their own buoyancy, the bubbles enter the spitzkasten zones, up which they travel without interference, dropping most of the gangue particles mechanically entangled between them as they ascend. They collect on the surface of the pulp at the top as a mineralized froth, which is voluminous enough to pass over the lip into the concentrate launders without the need of scrapers. The pulp, on the other hand, continues its downward passage and enters the air-lift chamber again. In this way a continuous circulation of the pulp is maintained, its course through the machine being more or less in the form of a double spiral.

The aeration is generally controlled by a single valve in the header of each machine, but for selective flotation the machine is sometimes divided by transverse partitions into sections 4 ft. long, the header over each section being provided with a separate air-valve. The depth of the froth is regulated by means of the adjustable gate of the tailing weir. If difficulty is likely to be experienced in making a clean tailing with the normal amount of aeration, it is preferable to use two machines. The second one is run as a scavenger with an excess of air as compared with normal requirements, the low-grade froth so produced being pumped back to the head of the primary or roughing machine, in which the aeration is more normal in order that a comparatively clean concentrate may be produced. It is often possible to take a concentrate off the first few feet of the rougher rich enough to be sent to the filters as a finished product, the froth from the rest of the machine being pumped back to the head. When this method of flotation is adopted, it is an advantage to have the header divided into sections, each with its own valve, so that the aeration can be varied along the length of the machine. By increasing the volume of air at the discharge end the froth can be given a slight flow towards the head of the machine, with the result that the minerals are concentrated there to the exclusion of partially floatable gangue which might otherwise enter any bubbles not fully loaded with mineral.

If the froth from the feed end of the rougher is not of high enough grade, it must be re-treated in a separate cleaning machine, the length of which usually varies from one-quarter to one-half of the total length ofthe roughing and scavenging machines according to the amount of concentrate to be handled. Should still further cleaning be necessary, it is performed in a recleaner, which is generally of the same length as the cleaner. The tailings from these operations are often, but not necessarily, returned to the head of the rougher.

It is usual to prepare the pulp for flotation by adding the reagents to the grinding circuit or in a conditioning tank ahead of the flotation section, but soluble frothers such as pine oil and quick-acting promoters such as the xanthates can be added at the head of the machine if desired, since the air-lift provides enough agitation to emulsify and distribute them throughout the pulp. It is not as a rule advisable to introduce reagents into the air-lift chamber itself ; should it be necessary to do so to obtain a satisfactory recovery of the minerals, it is best to employ separate roughing and scavenging machines and to make the extra additions at the head of the scavenger.

Southwestern Air-Lift Machines are made of standard cross-section, as already stated, and in a series of lengths ranging, for ordinary purposes, from 4 to 48 ft. There is no limit to the possible length, however, and 100-ft. machines are in actual use. The tonnage capacities under different conditions will be found in Table 26. The pressure of air needed at the machine is from 1.6 to 1.7 lb. per square inch, which under normal conditions requires a pressure of about 2 lb. per square inch at the blower. It is usual to allow 75 to 100 cu. ft. of free air per minute at this pressure per foot of rougher and 45 to 70 cu. ft. per minute per foot of cleaner and recleaner. From these figures the approximate volume of air required for a machine or machines of any given length can be calculated. The power necessary to supply the air can then be found from Table 30.

The Callow Cell consists of a shallow horizontal trough, the bottom of which is covered with a porous medium, usually termed a blanket, consisting of a few layers of canvas or of a sheet of perforated rubber. Air is introduced at low pressure under the blanket, and, in passing through it, is split up into minute bubbles, which rise through the pulp in the cell, collecting a coating of mineral in the process.

Fig. 41 shows a section of the type of cell commonly employed. Its width is usually from 24 to 36 in., and its interior depth from 18 to 22 in. measured from the overflow lip ; the length varies according to requirements and is generally a multiple of the width. On the bottom are placed, side by side, the square open-topped cast-iron blanket frames or pans . The blanket covering the top of each pan is securely held in place by flat iron strips bolted round the edges, while one or two pipe grid-bars across the top prevent it from bulging. This arrangement allows a blanket to be changed in a few minutes should it becomedamaged. The air inlet to each pan projects through the bottom of the cell and is connected by a pipe and regulating valve to a header, which is provided with a main control valve.

The pulp enters one end of the cell through a feed opening and is discharged over an adjustable weir at the other end. There is no agitation, but the continuously rising stream of air bubbles keeps the particles of ore in suspension and induces a certain amount of circulation as the pulp passes along the cell. In this way the minerals are given many chances of becoming attached to the bubbles and thus of being carried over into the concentrate launder. The froth that forms on the surfaceof the pulp, usually to a depth of 8 to 10 inches, is voluminous enough to overflow the lips on each side of the cell without the use of mechanical scrapers.

For estimating purposes the average capacity of a Callow Cell may be taken as 2.5 tons of feed per square foot of blanket area per 24 hours and the air consumption as 9 cu. ft. of free air per minute per square foot of blanket at a pressure of 4 lb. per sq. in. A greater pressure is likely to be required if the blankets become blinded .

The Callow Cell has proved satisfactory for many types of ores, but it has the disadvantage that coarse or heavy sand settles on the blankets, and can only be kept in motion by flogging the latter with short rubber-buffered poles. Moreover, if lime is employed in the circuit, the blankets become impregnated and clogged with calcium carbonate, which necessitates periodical acid treatment for its removal. The use of perforated rubber sheets in place of canvas in the Callow Cell mitigates without entirely curing these difficulties, which at one time were thought to be inherent in the use of a porous medium. They have been overcome, however, by the development of the Callow-Maclntosh Machine.

The Callow-Maclntosh, or the Macintosh Machine, consists of a shallow trough or cell at the bottom of which is a hollow revolving rotor covered with a porous medium. Fig. 42 shows its construction. The pulp enters through a feed opening at one end, and is discharged at the other in much the same way as in a Callow Cell. The rotor, made of seamless steel tubing with a cast-steel ring welded in each end, is perforated with -in. holes at 7-in. centres; it is about 8 in. shorter than the length of the cell and is usually 9 in. in diameter. Its weight is taken by two hollow shafts, each fitted with a flange, which are bolted to the ends of the rotor by means of four studs. This method of attachment enables the rotor to be changed and a new one inserted with little loss of time, usually not more than 15 minutes. The shafts project through the ends of this cell and are supported on self-aligning ball and socket bearings outside, so placed that the rotor itself is a few inches clear of the bottom of the trough. A rubber gasket, shown in Fig. 43, seals the opening at each end by simple pressure on a cone-faced disc mounted on the shaft. The joint is not completely watertight and a slight leakage takes place through it at the rate of about one quart per minute. At the discharge end this escaping pulp gravitates to the tailing launder, while at the feed end it is usually led to one of the pumps returning a middling product to the roughing circuit. The gasket is preferable to a stuffing-box, as it contains no grease and requires no gland water.

The rotor covering consists of a canvas sock or of a single sheet of perforated rubber. The latter is now far more commonly employed, since it lasts five times as long as the other, its life generally exceeding 18 months ; moreover it seldom becomes blinded withcalcium carbonate, and requires an air pressure of only 2 lb. per square inch instead of the 3-lb. pressure needed for canvas. The rubber sheets are made of pure gum about 5/64 in. thick with 225 holes per sq. in., the holes being made so as to allow the air to pass through while preventing the percolation of the pulp into therotor in the event of a temporary shut-down. Two scraper bars of angle iron, 1 by 1 in., are bolted to opposite sides of the rotor on the top of the covering. They project 2 in. beyond the ends of the rotor, and their purpose is to keep in circulation any sand that settles on the bottom of the cell, at the same timeprotecting the porous medium from undue wear by contact withsuch material. Air is introduced into the rotor through one or bothof the hollow shafts, which are connected by special inlet joints with themain supply. When both ends are employed for the admission of air,the rotor is usually divided into two sections by a central partitionto enable each half to be controlled separately. The rotor is driven ata speed of about 15 r.p.m. by an individual motor connected with theshaft at one end of the cell; either a worm drive directly coupled to themotor or a chain drive coupled to the motor through a speed reducercan be employed.

The principle on which theCallow-Maclntosh Machine worksis very similar to that of a CallowCell. The air bubbles actuallyissue from the top of the rotor,where the hydraulic pressure islowest, and spread out as theyrise, their distribution throughthe pulp being quite as even andeffective as when a flat blanket isused. The cell never needs flogging since the movementof the rotor prevents sand fromsettling on it, and the scraperbars keep in circulation theheavy particles that would otherwise settle on the bottom. Themachine can, if necessary, handle ore as coarse as 20 mesh at a W/Sratio of 1/1 without choking.

The control of a pneumatic cell is different from that of a machine of the mechanically agitated type, of which each cell is capable of performing the function of a high-speed conditioner. Little conditioning takes place once the pulp has entered a pneumatic cell, and provision must therefore be made for its proper preparation when employing heavy oils or chemical reagents which need a long contact period. The froth is usually maintained at a depth of 8 to 10 in., giving an effective pulp depth of 18 to 20 in. The very large volume of air bubbles released enables flotation to be effected more rapidly than in any other type of machine, the actual time required depending mostly on the degree to which the minerals have been rendered floatable. The upward stream of bubbles is so voluminous that, under ordinary conditions, the froth overflows the lips on both sides of the cell without the need of scrapers. For the same reason a considerable quantity of gangue is often carried over into the concentrate launder by mechanical entanglement with the bubbles, and one, sometimes two, subsequent cleaning operations are generally necessary in consequence. This, however, is by no means therule ; a concentrate of high enough grade to be sent to the filtering section as a finished product can sometimes be made in a single rougher- cleaner cell. When the Callow-Macintosh Machine is run in this way (counter-current operation) a partitioned rotor is employed, since, by increasing the volume of air at the tailing-discharge end, the froth can be made to flow towards the head of the cell with the result that the minerals are concentrated there to the exclusion of gangue particles. The same effect can be obtained in a Callow Cell by regulating the admission of air to the individual pans in a similar way. If is often the practice, especially in counter-current operation, for the rougher to be followed by a scavenging cell, which is run with an excess of air as compared with the former, the froth being returned to the head of the first cell.

Callow-Macintosh Machines are made in lengths of 10, 15, and 20 ft. and in widths of 24, 30, and 36 in. with a rotor 9 in. in diameter. The vertical distance from the centre-line of the rotor to the overflow lip is about 24 in. The design of the machine, however, lends itself to the construction of larger sizes for big scale operationsi.e., up to a 30-ft. cell 48 in. wide with one or two 9-in. rotors. The 30- and 36-in. cells are sometimes fitted with rotors up to 15 in. in diameter to meet special requirements.

The capacity of the standard machine varies considerably according to the grade and character of the ore. The average capacity of a rougher or rougher-cleaner cell is from 8 to 12 tons of dry feed per foot of rotor length per 24 hours. When cleaning is practised, the tonnage per foot of total rotor length (roughers, scavengers, and cleaners) may vary from 4 tons for a slow-floating ore needing double cleaning to 10 tons for an easily-floated ore with single cleaning, the average being about 6 tons per foot of total rotor length. The cleaning section usually amounts to between one-quarter and one-half of the combined length of the roughing and scavenging cells. The width of cell employed depends on the character of the ore, the time of treatment, and the tonnage.

The quantity of air necessary varies from 5 to 7 cu. ft. per minute per square foot of aerating surface at 2- to 2-lb. pressurethat is, from 12 to 16.5 cu. ft. per minute per linear foot of rotor. With a Roots type blower the power consumption in respect of the air supply is about 12 h.p. per 1,000 cu. ft. of free air per minute at a pressure of 2 lb. per square inch. The power needed to turn the rotor averages 0.5 h.p.

used flotation cells for sale. huatao equipment & more | machinio

used flotation cells for sale. huatao equipment & more | machinio

1. Self-absorption of air. 2. Good cycling property of pulp. 3. Flotation machine's suction volume is relatively stable. 4. Medium mixing strength and good suspension of solid particles. The Mineral Flotation / f...

Paper Mill Flotation Deinking Cell for Paper Recycling Line Application&Feature: 1.Mainlyappliedtodeinkrecycledpulp,itcaneffectivelyremovetheink,lightimpuritiesandstickiesetc. 2.Highefficientflo...

Recycle paper pulping Totally Enclosed Flotation Deinking pulp Cell Application&Feature: 1.Mainlyappliedtodeinkrecycledpulp,itcaneffectivelyremovetheink,lightimpuritiesandstickiesetc. 2.Higheffic...

Biobase Lab Small Coal Mining Mineral Iron Copper Ore Denver FrothFlotationCell TankSeparator Device With Floating Catalys Product Description Introduction BK-FD12 Laboratory multiple cell floatation device is...

Flotation cell accessory packages: - new induction motors - Blowers - Sheaves - Belts - Level control systems - Spare parts D'Angelo International can outfit your new or existing flotation cells with a range of a...

- New and unused flotation cells made to order in North America - Complete with mechanism, belt guards, and motor baseplate - All necessary cell and mechanism parts coated with protective 520 natural rubber (incl...

- 32 cells available; 1 bank of 12, 2 banks of 10 - 100 cubic feet per cell - Cleaned, painted, tested - Rubber and bearings in good condition - Motors, belts, and pulleys included - - Launder boxes required - Pr...

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atv woods trailer, pull behind trailers, single axle dump, single axle non-dumping, tandem, for hauling wood, soil, etc: outdoor solutions

atv woods trailer, pull behind trailers, single axle dump, single axle non-dumping, tandem, for hauling wood, soil, etc: outdoor solutions

Designed for those who need a bigger trailer than the galvanized version (above) but the budget wont support the tandem trailer. Sturdy single axle, non dumping, with a high clearance, it can go anywhere your ATV goes.

The wide floatation tires work better in soft terrain because they do not sink in the ground like highway tires do. The swivel hitch will eliminate damage to the ATV in the event of a trailer roll over.

flotation process - an overview | sciencedirect topics

flotation process - an overview | sciencedirect topics

Flotation processes are based on the different surface wettability properties of materials (Wang etal., 2015). In principle, flotation works very similarly to a sink and float process, where the density characteristics of the materials, with respect to that of the medium where they are placed are at the base of the separation. Sometimes a centrifugal field is applied to enhance separation. Flotation works in a different way in the sense that in a liquid medium, usually water, a carrier is introduced, air bubbles, responsible to float hydrophobic particles that adhere to the bubbles with respect to the hydrophilic ones that sink. According to surface plastic characteristics, this technique can be profitably applied, in principle, to separate waste polymers (Fraunholcz, 2004). To enhance or reduce plastic surface characteristics (i.e., hydrophobic or hydrophilic) appropriate collectors, conditioners (Singh, 1998; Shen etal., 2002), and flotation cell operative conditions (i.e., air flow rate, agitation) can be utilized. Usually plastic flotation is carried out in alkaline conditions (Takoungsakdakun and Pongstabodee, 2007). Once floated, hydrophobic polymers are recovered as well as the sunk ones (i.e., hydrophilic) at the bottom of the cell. This technique, even if it is well-known (Buchan and Yarar, 1995) and in principle quite powerful is not widely used mainly for three reasons: (1) it is a wet technique, this means that water has to be recovered and processed before reutilization, due to the presence of the reagents and contaminants, (2) polymer surface status (i.e., presence of dirtiness/pollutants and/or of physical/chemical alteration) can strongly affect floatability, and (3) large variation of waste plastics feed in terms of composition. Flotation allows to separate PS, PVC, PET, PC, and mixed polyolefins (MPO).

The flotation process depends on several design and operational variables. We consider a superstructure that includes the following three flotation stages: the rougher, which processes the feed; the cleaner, which generates the final concentrate; and the scavenger, which generates the final tailing, as shown in Fig. (1). This is a simple superstructure but is used here as an example.

The objective is to maximize the total income with respect to the operation conditions and process design. The decision variables to be optimized are divided into design and operating variables. The design variables include equipment dimensions, such as the cell volume and total number of cells for each stage. The operating variables correspond to operating times for each cell at each stage and the directions of tails and concentrate streams. In stochastic problems, the operating variables (second level variables) are able to adapt to each scenario to increase the total income. Moreover, the design variables (first level) are the same for all scenarios.

The flotation process depends on several design and operation variables. We consider a superstructure that includes three flotation stages: rougher, scavenger and cleaner stages, as is shown in figure 1. We allow for the consideration of multiple scenarios. The model consider constraints that enforce the kinetics of flotation and the mass balance on each flotation stage, the behavior at the splitters and mixers, the mass balance at the splitters and mixers, direction choice in the splitters, the penalty the seller must pay for arsenic content in the concentrate, cell volumes, and the costs associated with the flotation cells.

For the deterministic, model we have only a single scenario, and the model then simply maximizes the total income subject to the dynamic and economic constraints. In the stochastic models, we assume we have more than one scenario. Because of this, we need to replace the objective by the maximization of the expected total income. For this, we need the probability of a given scenario. In addition, we know that some of our decision variables can depend on the scenarios. This model corresponds to a stochastic MINLP.

In flotation process, the gas or air bubbles are introduced through culture suspension, and the microalgal biomass get attached to gaseous molecules and accumulated on the liquid surface. This method is particularly effective for thin microalgae suspension that could be simply gravity thickening [38]. The basic variations of this process are dispersed air flotation, dissolved air flotation, electroflotation, and ozone flotation [55,56,57]. The ratio of gaseous molecules to microalgae is one of the most important factors affecting the performance of the flotation efficiency. Several researchers have confirmed that ozone flotation was more effective than other methods [58,59]. Also, ozoflotation could improve lipid recovery yields and modify fatty acid methyl ester (FAME) profiles. The ozone flotation could increase the cell flotation efficiency by modifying the cell wall surface and/or releasing the active agents from microalgal cells [60]. Moreover, the ozone flotation can also improve the quality of water by lowering the turbidity and organic contents of the effluent [58]. Flotation separation efficiency relates to bubble size [61]. Smaller size of gas bubbles has lower rise velocity and higher surface area to volume ratio. This enables their longer retention time and better attachment efficiency with the microalgae cells and leads to the increasing in harvesting efficiency by floatation [64]. Thus, one of the most efficient ways of achieving maximum attachment is by generating as many small bubbles as possible [61,62,63]. Combinations of flocculation with flotation have been also used to increase the harvesting efficiency [64,65,66].

In using these equations, however, one must use parameters with consistent units.(1-3)E=(Ci-Co)/Ci(1-4)E=K/(Qw-K')(1-5)E=(6Kpr2hqg)/(qwdb)whereE = efficiency per cellCi = inlet oil concentrationCo = outlet oil concentrationQw = liquid flow rate, BPDKp = mass transfer coefficientr = radius of mixing zoneh = height of mixing zoneqg = gas flow rateqw = liquid flow through the mixing zonedb = diameter of gas bubble

The froth flotation process is more than a century old and was developed over a long period of time [8]. It takes advantage of the surface chemistry of fine particlesif one particles surface is hydrophobic and another is hydrophilic, upon generation of air bubbles, the hydrophobic particles tend to attach to the air bubbles and float, allowing for a separation between particles in the froth and those in the main body of the liquid.

Typically three different types of chemicals are used in the froth flotation process: collector, frother, and modifier. First, the collector is added to the iron ore slurry to selectively coat the iron oxide particles, making the surface hydrophobic. The slurry then goes to a flotation cell, where air bubbles are generated using an impeller and aerator (Figure 1.2.4). At this step, the frother (for example, fuel oil) is added to the ore slurry to form stable froth and air bubbles. Iron oxide particles stick to the air bubbles and float. Floated and concentrated iron ore slurry is then skimmed from the surface of the bath, and water is removed using a filter press. If the desired iron content is not achieved, the process is repeated. A modifier is added in some cases to enhance the performance of the collector. Frother is the most important chemical that must always be present. Without the generation of stable air bubbles, hydrophobic particles will not have anything to attach to and will not separate from the bulk solution.

Depending on the type of collector, either iron oxide or silica particles can be floated. An anionic collector is added to float the iron oxide particles, a cationic collector for the silica particles [9]. Depending on the situation, the pH of the slurry can be adjusted by adding acid to the solution, which may also enhance the properties of the collector.

The basic objective of a flotation device is to keep the pulp in suspension and provide the air bubbles. The size of air bubbles matters as it controls flotation kinetics as well as the carrying capacity of the bubbles. The design technology determines the characteristics of the machine, resulting in concomitant factors like how the collision and contact between air bubbles and particles takes place. The two resultant products, concentrate and tails, need to be evacuated properly. The most widely used flotation machines can be broadly classified into mechanical and pneumatic depending on various factors. The former use impellers or rotors, which are absent in the latter.

The shape of a mechanical flotation tank is essentially rectangular, U-shaped, conical or cylindrical, according to the cell type and size. It is fitted with an impeller/rotor and stator/diffuser. Air enters into the device through a concentric pipe surrounding the impeller shaft either by self-aspiration or aided by a compressor. The function of the rotating impeller is to keep particles in suspension by thoroughly mixing the slurry and dispersing the injected air into fine bubbles through a diffuser. It also provides conditions for promoting particlebubble collisions.

There is a necessity for the generation of three different hydrodynamic zones for effective flotation. The region near the impeller comprises of a turbulent area required for solids suspension, dispersion of air into bubbles and bubbleparticle interaction. Above the turbulent region lies a quiescent zone where the bubbleparticle aggregates move up in a relatively less turbulent sector. This zone also helps in sinking the amount of gangue minerals that may have been entrained mechanically. The third region overhead the quiescent zone is the froth zone serving as an additional cleaning step, and improves the grade of the concentrate. Particles that do not attach to the bubbles are discharged out from the bottom of the cell (Vazirizadeh, 2015). Fig. 5.33 shows a typical schematic of a mechanical cell.

Mechanical cells are arranged in a series called a bank, having enough cells to assure the required particle residence time for adequate recovery, the subaeration cells are arranged in cell-to-cell flow, while the supercharged machines are placed in an open-flow design.

The strongly hydrophobic and optimised-sized particles are likely to float first in a bank of flotation cells. Sluggish flowing particles float in diminishing order, and so forth, giving rise to total recovery of about 100%. A minimum of four cells is required for coal flotation with a residence time of 5 minutes (Euston et al., 2012). The residence time, pulp volume and flotation kinetics play a vital role in determining the selection of the number of cells required in a flotation circuit. To prevent loss of floatable coal along with tailings, it is advisable to put cells in series. Fig. 5.34 indicates the coal recovery through multiple cells (in series) in a bank. Fig. 5.35 demonstrates arrangement of cells both in series and parallel, the series arrangement gives optimum recovery of combustibles.

The most common examples of pneumatic cells are the column cell and the Jameson cell. As shown in Fig. 5.36, a flotation column is typically a tall vertical cylinder. It is fed with coal pulp at the top third of column. It has no mobile parts or agitators. Air bubbles are injected either through external or internal spargers at the bottom. These bubbles rise up in countercurrent with the descending flow of the pulp. Hydrophobic particles attach to the air bubbles forming bubbleparticle aggregates and move upwards. The zone where this process takes place is called the collection zone. The ascending bubbleparticle aggregates accumulate in the upper part of the column called the cleaning or froth zone, and then overflow into a launder as a concentrate. Wash water is sprinkled at the top of the column to wash off entrained gangue (hydrophilic) particles, which are sent back into the collection zone. The application of wash water helps stabilise the froth and produce high-grade froth concentrates. The hydrophilic particles, along with misplaced hydrophobic particles, are finally released at the bottom of the column.

In spite of improved separation performance, low capital and operational cost, less plant space demand, low maintenance cost, ease of operation, lower energy consumption and adaptability to automatic control (Wills and Napier-Munn, 2006), axial mixing can significantly reduce the overall performance, particularly in larger-diameter columns. Axial mixing can be decreased by different methods (Kawatra and Eisele 1999, 2001):

A Jameson cell is schematically presented in Fig. 5.37. A high-pressure jet, created by pumping feed slurry through the slurry lens orifice, enters a cylindrical device called a downcomer. The downcomer acts as an air entrainment device which sucks air from the atmosphere. The jet of slurry disseminates the entrained air into very fine bubbles after plunging upon the liquid surface. Then, it creates very favourable conditions for collision of bubbles and particles, and their attachment. The particlebubble aggregates move down the downcomer to the cell and float to the top to form the froth. The hydrophilic minerals sink to the bottom and exit as tailings. Tailings recycling is practiced to reduce feed variations to the cell so that the downcomer can operate at a stable feed pressure and flow rate. This helps to ensure steady operation. The downcomer provides an ideal situation for particlebubble contact and minimises the residence time due to rapid kinetics and separate contact zone. Thus, the Jameson cell is of much lower volume compared to equivalent-capacity column or mechanical cells. There is also no requirement for agitators or compressors besides the feed pump.

The dissolved air flotation process takes advantage of the principles described above. Figure 7-104 presents a diagram of a DAF system, complete with chemical coagulation and sludge handling equipment. As shown in Figure 7-104, raw (or pretreated) wastewater receives a dose of a chemical coagulant (metal salt, for instance) and then proceeds to a coagulation-flocculation tank. After coagulation of the target substances, the mixture is conveyed to the flotation tank, where it is released in the presence of recycled effluent that has just been saturated with air under several atmospheres of pressure in the pressurization system shown. An anionic polymer (coagulant aid) is injected into the coagulated wastewater just as it enters the flotation tank.

The recycled effluent is saturated with air under pressure as follows: a suitable centrifugal pump forces a portion of the treated effluent into a pressure holding tank. A valve at the outlet from the pressure holding tank regulates the pressure in the tank, the flow rate through the tank, and the retention time in the tank, simultaneously. An air compressor maintains an appropriate flow of air into the pressure holding tank. Under the pressure in the tank, air from the compressor is diffused into the water to a concentration higher than its saturation value under normal atmospheric pressure. In other words, about 24 ppm of air (nitrogen plus oxygen) can be dissolved in water under normal atmospheric pressure (14.7 psig). At a pressure of six atmospheres, for instance (6 14.7 = about 90 psig), Henry's law would predict that about 6 23, or about 130 ppm, of air can be diffused into the water. In practice, dissolution of air into the water in the pressurized holding tank is less than 100% efficient, and a correction factor, f, which varies between 0.5 and 0.8, is used to calculate the actual concentration.

After being held in the pressure holding tank in the presence of pressurized air, the recycled effluent is released at the bottom of the flotation tank, in close proximity to where the coagulated wastewater is being released. The pressure to which the recycled effluent is subjected has now been reduced to one atmosphere, plus the pressure caused by the depth of water in the flotation tank. Here, the solubility of the air is less, by a factor of slightly less than the number of atmospheres of pressure in the pressurization system, but the quantity of water available for the air to diffuse into has increased by the volume of the recycle stream.

Practically, however, the wastewater will already be saturated with respect to nitrogen, but may have no oxygen, because of biological activity. Therefore, the solubility of air at the bottom of the flotation tank will be about 25 ppm, and the excess air from the pressurized, recycled effluent will precipitate from solution. As this air precipitates in the form of tiny, almost microscopic, bubbles, the bubbles attach to the coagulated solids. The presence of the anionic polymer (coagulant aid), plus the continued action of the coagulant, causes the building of larger solid conglomerates, entrapping many of the adsorbed air bubbles. The net effect is that the solids are floated to the surface of the flotation tank, where they can be collected by some means and thus be removed from the wastewater.

Some DAF systems do not have a pressurized recycle system, but, rather, the entire forward flow on its way to the flotation tank is pressurized. This type of DAF is referred to as direct pressurization and is not widely used for treatment of industrial wastewaters because of undesirable shearing of chemical flocs by the pump and valve.

The behavior of coal in the flotation process is determined not only by a coals natural floatability (hydrophobicity), but also by the acquired floatability resulting from the use of flotation reagents. The general classification of the reagents for coal flotation is shown in Table12.1 (Laskowski, 2001).

The use of liquid hydrocarbons (oils) as collectors in flotation of coal is characteristic for the group of inherently hydrophobic minerals (graphite, sulfur, molybdenite, talc, coals are classified in this group). Since oily collectors are water-insoluble, they must be dispersed in water to form an emulsion. The feature making emulsion flotation different from conventional flotation is the presence of a collector in the form of oil droplets, which must collide with mineral particles in order to enhance the probability of particle- to-bubble attachment. The process is based on selective wetting: the droplets of oil can adhere only to particles that are to some extent hydrophobic. The effect of emulsification on flotation has been studied, and its beneficial effect on flotation is known (Sun et al., 1955).

Coal flotation is commonly carried out with a combination of an oily collector (e.g. fuel oil) and a frother (e.g. MIBC). All coal flotation systems require the addition of a frother to generate small bubbles and to create a stable froth (Table 12.2). Typical addition rates for frothers are in the order of 0.050.3kg of reagent per tonne of coal feed. Depending on the hydrophobic character of the coal particles, an oily collector such as diesel oil or kerosene may or may not be utilized. When required, dosage rates commonly fall in the range of 0.21.0kg of reagent per tonne of coal feed, although dosage levels up to 2kg/t or more have been known to be used for some oxidized coals that are difficult to flotate.

PO stands for propylene oxide (CH2-CH2-CH2-O-), and BO for butylene oxide (CH2-CH2-CH2-CH2-O-) Cresylic acids (mixture of cresols and xylenols) that in the past were commonly used in coal flotation are not in use any more because of their toxicity.

PO stands for propylene oxide (CH2-CH2-CH2-O-), and BO for butylene oxide (CH2-CH2-CH2-CH2-O-) Cresylic acids (mixture of cresols and xylenols) that in the past were commonly used in coal flotation are not in use any more because of their toxicity.

The beneficial effect of a frother on flotation with an oily collector was demonstrated and explained by Melik-Gaykazian et al. (1967). Frother adsorbs at the oil/water interface, lowers the oil/water interfacial tension and hence improves emulsification. However, frother also adsorbs at the coal/water interface (Frangiskos et al., 1960; Fuerstenau and Pradip, 1982; Miller et al., 1983) and provides anchorage for the oil droplets to the coal surface. Chander et al. (1994), after studying various non-ionic surfactants, concluded that the flotation of coal can be improved in their presence because of the increased number of droplets, which leads to an increase in the number of droplet-to-coal particle collisions. While the use of oily collectors and frothers is the most common, also a group of flotation agents known as promoters have found application in coal flotation. In general, these are strongly surface-active compounds and are mostly used to enhance further emulsification of water-insoluble oily collectors in water.

Because of environmental concerns associated with tailing ponds, the method for disposing of fine refuse from coal preparation plants by underground injection has been gaining wide acceptance. Unfortunately, many common flotation reagents, including diesel oil, are not permitted when fine refuse is injected underground into old mine works. This is the main driving force for finding replacement for the crude-oil based flotation collectors (Skiles, 2003). An alternative to fuel oil may be biodiesel, a product created by the esterification of free fatty acids generally from soy oil, with an alcohol such as methanol, and subsequent transesterification of remaining triglycerides. Water, glycerol and other undesirable by-products are removed, to produce a product that has physical characteristics similar to diesel oil. The use of some vegetable oils was demonstrated to provide equivalent (and even superior) flotation results when compared with diesel fuel (Skiles, 2003). These are the results of commercial scale tests on a circuit that has 4.25m in diameter columns. The product concentrate ash was 13.5%. The consumption of the tested vegetable oil was about two times lower from the consumption of diesel oil in these tests.

It features both proven technology and the latest technical innovations at the same time. This flotation cell is highly efficient when it comes to costs and operation. It can be easily scaled to various production levels without compromising performance. In short, OptiCell Flotation enhances the performance of the deinking line cost efficiently and ensures a reliable flotation process. The heart of the flotation process is the injector. In the OptiCell system, this has beendesigned with special care, using the experiences of earlier flotation technologies, modern computational fluid dynamics calculations, and new image analysis methods. The combination of these approaches results in a unique injector that represents the latest technology. This injector differs from traditional injectors in following respects:

OptiCell flotation by Metso is based on computational fluid dynamics and uses new image analysis methods. It is designed to provide smooth-flow velocities that allow unobstructed transfer of bubbles to the surface of the pulp mixture or froth, which improves the efficiency of ink removal. The aeration injector ensures optimal bubble-size distribution. The injector is designed based on the experiences gained with earlier flotation technologies combined with modern computational fluid dynamics calculations and new image analysis methods.

The linear structure of the flotation cells has a large surface area, which has reject separation and fiber loss. This flotation cell design also contributes to high sludge consistency (less water in the sludge) by ensuring smooth drainage of froth (Aksela,2008). The elliptical shape of the flotation cells in this technology is optimal for internal pulp circulation for improved ink removal. Moreover, the flatness of the cells intensifies the rise of air bubbles within the available volume. The first OptiCell Flotation system started operation in September 2008 at Stora Ensos Maxau mill in Germany, which has an approximately 1000 t/day deinking facility (Metso,2012b). According to Metso, the brightness from the complete flotation system has increased by two units. A brightness gain of 13 units from thick stock to accept was obtained with the OptiCell process. Reject ash content also improved and fiber losses decreased. As a result of the flotation performance and corresponding brightness improvement, peroxide consumption has decreased significantly in bleaching. In addition, the stickies content was reduced significantly. It was the lowest ever measured at the deinking line 1 at Maxau mill. The benefits of OptiCell flotation are summarized in Table11.8 (Aksela,2008; Metso,2012b).

"wood" flotation process & machine

In my opinion, the concentration of minerals by flotation is the most interesting problem in ore-dressing, and will command eventually far more consideration than it has at present. For many ores it furnishes a complete method of concentration absolutely independent of specific gravity, and thus simplifies the milling of many difficult combinations. Not having to recognize the influence of specific gravity, we are also able to dispense with some of the complications resulting from classifying-conditions.

Where flotation cannot be used for a complete separation of the minerals from the gangue, it will be found, in nearly all cases, to save those portions which are generally lost by reason of their fine state of division. For this reason, it should be looked upon mainly as an aid to general concentration, although there are many ores so well adapted to flotation that more than 90 per cent, of the valuable metals can be recovered from them in one operation by this process.

My earliest work with flotation was done in 1895, when I used various oils, grease, soap, etc., with results that were remarkable. I did not realize then that with water alone equally good separations could be made. Up to 1906, my endeavors were confined to attempts to keep the concentrates on the tables well wetted and away from exposure to the air. In common with others, I observed that when there was any obstruction in the wash-water on the table of sufficient size to divert the water to either side, the concentrates collected, and, on exposure to the air, would promptly float away. This convinced me that the sulphides were all good swimmers, and that if, instead of endeavoring to drown them, they chance, they would all float.

Having successfully demonstrated this fact, as well as another equally important, namely, that practically all the oxides are quickly wetted, become saturated like a sponge, and, sinking under the surface-film, become subject to gravity-conditions, I commenced my experiments.

I found that if dry-crushed ore, not necessarily very fine, was gently deposited upon a swiftly-moving sheet of water, separation of the sulphides from the gangue took place. Retarding the current permitted the gangue to sink, while a film of sulphides remained on the surface. To collect this film I have devised three types of machine, shown in Figs. 1, 2, and 3. Fig. 4 is a photographic view of the type C machine.

All three types are similar so far as the system of feed is concerned. In type A, a sectional view of which is presented in Fig. 1, the unsized ore, at whatever mesh has been decided upon, and ranging from 10-to 40-mesh or finer, is delivered to a feed-hopper, A, the outlet from which is under close control, so that a uniform amount of ore can be steadily dropped upon a plate, B, vibrating 500 times per minute. From this plate the ore falls direct to the sheet of water, which flows over a plane, C, inclined at 7. The film of sulphides crosses the tank, D, and in falling over the outlet is dewatered and split off by means of a nearly vertical screen, E. O, O, are guides or stretchers, and perform quite an important function when this type of machine is used. The elastic properties of the film are such that if these guides were not used, the sheet of water emerging from the tank would quickly assume a spiral shape, which would result in a wetting of the sulphides. These would then pass through the screen instead of over it.

The guides are so adjusted that the water intersects the screen in a line parallel to the surface of the tank. By slight adjustments one is then able to split off cleanly the film of concentrates from the water and the gangue in suspension. A separate patent was allowed for this feature. The water passing through the screen, containing more or less fine gangue in suspension, is caught at H and conducted to a launder, F, which also collects such coarse mineral and gangue as steadily issues from the bottom outlet of the tank, G, from which it passes to tables. I sometimes find it advantageous either to classify hydraulically or screen-size the tank contents en route to the table; but it is not essential in all cases, because, having removed the fine sulphides, as well as a large proportion of the larger sizes, by flotation, only the coarser mineral and the coarser gangue have to be considered, as the fine gangue in suspension passes off with the wash-water.

The feed of the type B machine, Fig. 2, is delivered over the inclined plane C to a conical tank, D, in which a rapidly-revolving motion with vortical effect is given to the surface of the water by means of a water-jet against the side of the tank. The feed falls upon an inclined sheet of water, The film of sulphides is guided by means of partly-submerged vertical screens, E, to a funnel, J, the outlet of which penetrates the side of the tank, where the concentrates are delivered to a launder, F.

The output of a tank 5 ft. in diameter will exceed 0.5 ton per hour on a 20-mesh ore that is adapted to flotation. Necessarily, a small amount of water is required to carry the film into the funnel. The film, having a definite thickness, is probably carried on an amount of water of two or more times the volume of the concentrate. As this water carries more or less gangue in suspension, it makes the concentrates siliceous. In certain sections of Mexico, where copper sulphides are in demand and silica is not objectionable, the conical-tank plan meets the requirements of the situation; but this form would not be satisfactory, for instance, in Utah, where a high-grade copper-concentrate is of more importance than the highest possible recovery. In Utah there is a penalty imposed by the smelters for silica above a certain amount, a condition which does not exist in Mexico.

In the type C machine, Fig. 3, the 20- to 40-mesh ore is dropped from the vibrating plate, B, to a corrugated rubber- or canvas-covered cylinder, K, which is revolving in water. Enough water is elevated above the water-level of the tank to wet the gangue and float the sulphides. This sulphide film passes from the cylinder over the tank-surface to an endless canvas belt, A, which emerges from the tank and is controlled by sprockets. Water-jets gently force the concentrates upon the canvas, but, as the canvas is elevated, the water drains back to the tank, carrying with it any gangue in suspension. The concentrates are deposited in another tank and collected in a launder, F. An automatic siphon, N, prevents an overflow of concentrates from the tank after they are discharged from the belt. G is the outlet-valve controlling the discharge of tailings after flotation. In most cases the tailings pass to a Wilfley table. Quite frequently, however, the recovery made by flotation makes any additional treatment unnecessary.

As clean wash-water is used at the delivery end of the take-off belt to remove the concentrates, they have a minimum silica- content. The difference between type B and type C in this respect was shown by comparative tests. The crude ore used in each test assayed 2.63 per cent, of copper. The B machine saved 88 per cent, and gave concentrates assaying 18 per cent, of copper, with 28 per cent, of SiO2, while the G machine saved 93 per cent., giving concentrates assaying 27 per cent, of copper, with 3.6 per cent, of SiO2, both tests being made by flotation alone.

As the corrugations of the rubber-covered cylinder are only adapted to comparatively fine ore, say 20-mesh and finer, they would not bring up enough water to saturate particles that were larger than the corrugations; hence, in treating 8- or 10- mesh ore, all of such particles would not reach the water, and other dry ore would fall on top of them, causing an unclean concentrate. To prevent such conditions in treating coarse ores, the cylinder is removed, and a rapid impetus is given to the surface of the water by means of water-jets from a pipe just in advance of the position formerly occupied by the cylinder.

An ore suitable for jigging, in which the sulphides are liberated at 4-mesh, can be fed, either sized or unsized, direct to the tank. The treatment of coarse ore, say from 4- to 12-mesh, is performed for the purpose of floating off fine particles of sulphides which mechanically adhere to the coarser sizes. I have carefully dry-screened ore from 4- to 20-mesh and then subjected these coarse sizes to flotation, the result being a high-grade product that would add 5 per cent, to the general recovery.

If, after grinding a given ore to the right size, it is at once floated, from 30 to 85 per cent, by weight of its sulphide particles are likely to float. If another portion, prepared at the same time and under the same conditions, is set aside for several days and then treated, less than one-third of what floated before will remain on the surface. This applies particularly to the copper sulphides, and must be due to partial oxidation, although such a change is not visible to the eye. This behavior of copper-ores is more pronounced if steam-heat be applied in drying a previously-made concentrate for flotation. It is therefore not advisable to regrind wet jig-tailings and dry them for flotation. To secure the best possible flotation from copper-ores, they must be ground directly to the required mesh for separation of the mineral from the gangue, and promptly milled.

Curiously enough, this state of affairs does not exist in the treatment of some other complex sulphides, for instance, those containing galena and the non-magnetic resin blend. At 40- mesh a 10-per cent, lead-ore, with 20 per cent, of zinc, will produce a flotation-concentrate accounting for 80 per cent, of the lead, assaying from 56 to 60 per cent, of Pb and 7 per cent, of Zn. If the tailings, after flotation, are passed over a Wilfley table, a high-grade coarse lead-concentrate is easily cut out from the zinc and silica. The removal of the silica from the zinc-middlings enables us to obtain a high-grade zinc-concentrate. Now, if this blende, which sank in the tank during flotation, be dried and re-treated by flotation, nearly all of it will float, while the accompanying silica will sink, leaving a high-grade product. This particular effect is not so pronounced when the blende is of the magnetic or so called black-jack variety.

I have noted in the treatment of copper-ores that the tendency to float decreases as the sulphur-content diminishes. More than 90 per cent, of the copper of chalcopyrite ore with a hard white quartz gangue can be saved with less than 5 per cent, of silica in the concentrate by flotation alone; but if bornite or chalcocite is present, the proportions of these minerals saved are much smaller than of chalcopyrite. If chalcocite contains a small percentage of chalcopyrite, practically all of the latter will float and some of the fine chalcocite with it. This leaves the coarse chalcocite in good shape for final table-separation.

Most of the magnetic minerals will sink, so that a good flotation-concentrate can sometimes be made from a mixture of chalcopyrite associated with a magnetic blende. Magnetite will sink and leave suspended on the surface any fine sulphides that may be with it, as well as the metals, such as gold, silver, platinum, etc. To obtain the best effects for the fine metals, we sometimes modify the film by using a minute amount of oil. The oil being used purely as a conveyor, and not as an enveloping element, my patents do not conflict with the oil-systems. My main idea is the use of water alone. I rarely use oil and never acid or gas.

Judging from the appearance of a 20-mesh flotation-concentrate, the inexperienced observer assumes, from the smooth surface of the film, that only the extremely fine sulphides are being recovered. A screen-analysis of the concentrate, however, shows that more than 10 per cent, of the concentrate is between 20- and 40-mesh. Apparently the elasticity of the film is such that it supports the coarser particles, but not rigidly, so that only the apex of each piece of sulphide is shown, and as it lies along with the fines, a uniform, smooth surface is presented.

The strength of the film is illustrated in another way. If particles of dry ore are placed upon the film of sulphides, they will be thrown off by the rebound of the film as it passes over the curve of the dam, the effect being practically the same as the repulsion of highly-conductive minerals in passing over an electrostatically-charged roller.

The elastic properties of the film are plainly shown during the operation. The sulphide film spreads out rapidly over the tank-surface. If its passage is restricted, it will immediately be compressed into ridges, which in time become sufficiently heavy to sink of their own weight. Also, before sinking, the film will creep up vertically on the corrugated roller.

As long as the feed is continuous, the sulphides on the film easily slip over to the take-off belt. This action is assisted by a gentle water-pressure directly in front of the belt. A gentle current of air also produces the same result.

I have alluded to oxidizing influences preventing or interfering with flotation-results in the treatment of copper sulphides. With several other sulphides this is not the case, so that flotation can frequently be applied, after amalgamation and concentration, as a more economical method of recovery than cyanidation of the tailings.

For instance, I had a $20 gold-ore, of which $15, or 75 per cent., was removed by amalgamation and concentration. Of the $5 left in the tailings, 93 per cent, was extracted by cyanide, but at a cost of $2.75, by reason of the high cyanide-consumption, caused by values still locked in the fine sulphides. The same tailings, assaying $5, were dried and flotated, producing a 10-oz. gold-concentrate and leaving 80 cents per ton in the final tailings. The cost was far less than that of cyanide treatment.

Nearly all the average free-gold ores of California contain arsenical pyrites, and this, in turn, carries gold that is not free. Such ores, if crushed dry and treated by flotation, yield high- grade concentrates, accounting for from 10 to 20 per cent, of the total valuable metals in the ore. After flotation, the free gold can be amalgamated as usual. I have found that all that is necessary is to feed over an amalgamated plate, which has an inclination of 7 and is covered by water.

Such flotation-concentrates of arsenical pyrites frequently assay many thousands of dollars per ton, and constitute an important item in the total recovery. A considerable percentage of the values in some varieties of wet-crushed ores, particularly of copper and arsenical sulphides in a hard gangue, will float. While we cannot do anything with sulphides in suspension, we can readily save all floating slimes from wet-crushing by this same general system, and with a simple device placed in any launder below the mill, provided the surface-conditions are not rough.

All over the world, valuable ores, such as sulphides of silver, gray copper, galena, etc., are frequently found in a matrix of baryta, or associated with heavy garnet. The specific gravity of such gangues practically prohibits either jigging or table- separation ; but when this method of ore-dressing is applied to them, from 80 to 90 per cent, of their values can be recovered at one simple operation. A recent example of this was a 1,200-oz. silver-concentrate made from a Mexican ore containing 78 per cent, of BaSO4. Another, from Canadaa crystallized barite containing 10 per cent, of Pbgave a 90-per cent, recovery in a 70-per cent, lead-product. Many mines containing ores of such composition have been abandoned because of the closeness of the specific gravities.

We have had several ores of nearly clean pyrrhotite containing a small percentage of chalcopyrite. Most of the latter will float by itself. If the remainder is concentrated, the ratio of concentration is too low to make a valuable copper-product; but if this concentrate is dried and passed through a magnetic separator, such as the Wetherill or Dings, at a low amperage, the pyrrhotite is lifted off and the non-magnetic tailings will be a clean chalcopyrite. This, combined with the flotation-concentrate, gives a total high extraction.

The capacity of the series of flotation-machines here shown is variable, depending upon several conditions, among which are the nature of the gangue, the size of the ore, and the concentration-ratio. A standard machine treating a 20-mesh quartz ore, using a 3-ft. width of feed and having a 4-ft. take-off belt, will vary in capacity from 1.000 to 2,000 lb. per hour, unless the ratio of concentration is low, in which case the capacity will be smaller. Some ores that possess an easily wetted gangue and call for a high concentration-ratio can be fed rapidly at 20-, 30-, or 40-mesh.

For instance, a 1 or 2 per cent, molybdenite ore in a quartz gangue will give a clean concentrate, even if the ore is fed several times faster than an ordinary sulphide ore; and the same is true of graphite. In a special form of the machine, designed for the concentration of molybdenite and graphite, one or two partly-submerged cylinders, covered with either canvas or fine wire screen, are inserted in the tank. These cylinders revolve in the same direction and at the same rate of speed as the main take-off belt or belts. The film-concentrate is lifted from the tank, and passes over the rollers down to the tank-level again. With each elevation partly-saturated gangue becomes thoroughly wetted, and drops off into the tank without any apparent disturbance or wetting of the concentrate. Such action naturally raises the grade of the concentrate. In leaving the take-off belt, it is not washed off by direct water-jets, as is done with other ores, but the take-off belt is allowed to discharge its load at water-level in a shallow compartment, and the concentrate gently floats off to a collecting-launder.

In the flotation-treatment of molybdenite it frequently happens that copper sulphides, iron pyrite, and other impurities, are present, and must be removed to make a salable product. If copper and iron are present, they are re-concentrated on a Wilfley table, so handled that the tailings will contain the molybdenite. Another way is to dry the flotation-concentrate, give it a gentle roast, and pass it through a magnetic separator, to remove the copper and iron. Molybdenite- and graphite- concentrates are frequently contaminated by mica. This in some cases can also be taken out by the magnetic separator, or by re-flotation with the water at 110 F. It can also be removed electrostatically.

Objections are made to the drying of flotation-concentrates on accouut of their supposed fineness, but this is a mistake, as mentioned previously. There should be no more difficulty in drying concentrates of this nature than those produced by vanners or ordinary slime-tables. By alternating settling- tanks, which can be siphoned, or by using filter-presses, they can be constantly dewatered. They can be briquetted if necessary.

There are, in Mexico and elsewhere, dry-crushing mills that use magnetic separators to remove large amounts of gangue in the shape of garnets, siderite, etc., in order to get more favorable conditions for gravity-separations on the tables for the lead and zinc. If a flotation-machine be placed so that it will take the non-magnetic tailings before they reach the table, a high-grade lead-product results, and aids in the general separation of the coarse lead and zinc, improving each from 15 to 20 per cent.

Flotation will be found to be of the greatest benefit when applied to ores that have the largest losses in milling, as well as to certain ores that are virtually not subject to wet concentration on account of small differences in specific gravity.

The proper preparation of ores for flotation is essential. For instance, if a small sample is pulverized unnecessarily flue, in such a state as would result from the use of an ordinary muller, poor returns may be expected, but if the same ore is shattered by pounding on an iron plate, most favorable products will ensue, as the crystals retain more of their original shape.

For this reason, I have found that ores should be prepared in ball-mills, or by rolls, rather than in any machine that grinds much of the gangue to an impalpable dust, in which condition both the gangue and the mineral of many ores will float. Flotation-treatment of definitely-sized ore will materially increase the capacity of the large sizes by themselves, but mechanical difficulties are encountered in feeding the smaller sizes by themselves, and the capacity is also much reduced; hence it is advantageous, after having decided upon the proper mesh to liberate the sulphides from the gangue, to feed them all at once without previous sizing.

I wish, also, to place special emphasis upon the property of easy saturation or wetting of nearly all of the oxides. The only one that has surprised us and floated in appreciable proportions is the specular hematite in certain New Mexico zinc-ores, which floats up to 10 or 20 per cent, by weight. This is largely due to the tabular shape of its crystals. After magnetic removal of the iron, the non-magnetic tailings assay from 45 to 50 per cent, of zinc.

Simple as these flotation-devices are, it has taken several years to secure patents, both domestic and foreign, as well as to develop the general standard type of machine now on the market. It weighs 1,000 lb. net; measures 4 by 5 by 6 ft. Ready for mule-back transportation, its weight is 1,100 lb. It requires 0.25 h-p. and less water than is used in wet concentration.

Sulphides of the following metals are particularly favorable for flotation methods: Copper, lead, silver, antimony, bismuth, mercury, arsenic, molybdenum, zinc, iron, nickel, etc., as well as tellurides, graphite, gold, and platinum.

When tellurides are present, a high-grade flotation-concentrate is obtained and the tailings are left in excellent shape for cyanidation. I have proved this commercially on ores from Boulder county, Colo.

The concentration of molybdenite by flotation is decidedly interesting, in that it furnishes illustrations of several of the peculiarities of surface-tension. For instance, in the recent flotation of a Canadian ore of this nature, with a quartz-schist gangue, which assayed 1.87 per cent, of MoS2, I obtained an apparently pure film of MoS2, which, when assayed, was found to contain only 45 31 per cent, of MoS2, and 41.97 per cent, of insoluble residue. On examination, the under side of the film was seen to be composed of fine mica scales, which must have adhered to the flakes of MoS2. My first conclusion, that the mica was simply enveloped in a coating of MoS2, was not correct. The recovery in this particular case was 79.08 per cent. The test was made with water having a temperature of 56 F.

The above examination was followed by that of an 8.65-per cent, molybdenite ore from Alaska. This was a complex ore, containing pyrrhotite and magnetite, and low in silica. The flotation-concentrate, at 40-mesh, amounted to 23.4 per cent, by weight. As it assayed only 30 per cent, of MoS2, a recovery of 79 per cent., it was evidently contaminated by other minerals that had floated.

Another sample of the same ore, lightly roasted, gave a flotation-concentrate assaying 81.45 per cent, of MoS2, the ratio of concentration being 35.71 : 1, but the saving was only 26.3 per cent. This indicates that either some of the molybdenum is present in the original ore as an oxide, or that the roasting was carried too far and oxidized it. In the first test the saving was 79 per cent., assaying 30 per cent, of MoS2, while the second trial recovered but 26 per cent., assaying 81.45 per cent. It is possible, therefore, that by obtaining the right roasting-conditions, such an ore could be profitably treated.

The concentrates from one test of this ore assayed 1.86 per cent, of copper. After a gentle roast, this was removed by the Wetherill separator, leaving only a trace of copper, and with no material loss of MoS2.

A certain Alaskan ore recently treated by me in commercial quantities furnishes an illustration of the extreme sensitiveness of chalcopyrite to oxidizing influences. It contained 4 per cent, of copper, closely associated with a heavy pyrrhotite in a siliceous gangue. Careful work on this ore by jigging, tabling, regrinding of jig-tailings, all followed by canvas slime-table treatment, gave a 15 per cent, copper-concentrate, with a recovery of 62 per cent. Direct table-concentration made a 16 per cent, copper-concentrate, saving 47 per cent. The same ore, by flotation alone, made a 25-per cent, copper-concentrate, accounting for 69 per cent, of the copper. To this was added 11 per cent, from the table-work, a total of 80 per cent. In making this examination the ore was dry-crushed at 8-mesh, and the from 8- to 20-mesh product (dry-screened) was passed through the flotation-machine, resulting in a high-grade concentrate, which added 4.25 per cent, to the total recovery. This proportion came from the mechanically adhering fine sulphides, and would be lost in ordinary jig-work. The from 8- to 20- mesh jig-tailings of this test were dried and reground for flotation ; but on account of the oxidation not a particle floated.

Table I. gives the results obtained by flotation alone. All of the ores tried were difficult ones to treat, and several of them were not susceptible to gravity-separation. The recovery by flotation alone, in many instances, exceeded the returns from the best possible treatment in the wet way.

Alaska.30 mesh, 4.08 per cent, of copper. Close classification, jigging, table-concentration, with regrinding of jig-tailings, gave 65 per cent, of the copper, assaying 15.66 per cent, of Cu. Ratio, 5.9 tons to 1.

By flotation at 30-meshthe same orethe concentrate as-sayed 25.03 per cent, of Cu, recovering 73 per cent, at a ratio of 8.4 into 1. The Wilfley concentrate added to this 7.31 per cent, more, assaying 6.32 per cent, of Cu, ratio 21.19 to 1, a total recovery of 80.31 per cent. When treated at 40-mesh by flotation alone, the recovery exceeds the above total.

A second test of this ore gave 91 per cent. of its gold-content, 96.3 per cent. of the silver, 93.9 per cent. of the lead, and 84.41 per cent. of the copper. Several other combination methods applied to this ore failed to produce satisfactory commercial products.

used planers 4 sided moulders for sale. jiasheng equipment & more | machinio

used planers 4 sided moulders for sale. jiasheng equipment & more | machinio

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