gold flotation concentrate

gold flotation concentrate

The flotation method is a widely used technique for the recovery of gold from gold-containing copper ores, base metal ores, copper-nickel ores, platinum group ores and many other ores where other processes are not applicable. Flotation is also used for the removal of interfering impurities before hydrometallurgical treatment, for upgrading of low sulfide and refractory ores for further treatment. Flotation is considered to be the most cost-effective method for concentrating gold.

In this process of rock minerals that have been taken from the mine site and then destroyed by the machine to obtain a fine grain of sand to free metal-containing granules for further processing. In the destruction of mineral rocks of machine tools can use a stone crusher machine, so the minimum size of rock minerals can reach between 28 mesh.

At this stage after a mineral ore that is refined inserted into the machine agitator tank which is usually also called a flotation cell to produce a pulp slurry concentrate.Distilled water provision inserted into the flotation cell flotation machine is then run, examined the amount of initial pH and initial temperature. In the flotation tank, stirring with impellers, which are intended to produce turbulent motion of fluids (pulp), so that when inserted air flow will form air bubbles.In the pulp is then coupled collector-1,-2 collector and frother; flotation machine run back to the time varying adjustment, and examined the amount of the final pH and final temperature.

In the processflotation reagent which in use is a form of lime, bubble and collectors. Froth forming a bubble that is stable and that float to the surface as a froth flotation cell. Collector reagents react with the surface of the precious metal sulfide mineral particles making the surface is water repellent. surface of the mineral-bound water molecule is released and will be changed to hydrophobic.

Thus the collector end of the hydrophobic molecules will be bound hydrophobic molecules from the bubble, so the mineral ore can be adrift. Collector has a molecular structure similar to the detergent hydrophobic sulfide mineral grains are attached to the air bubbles that rise from the slurry zone into the froth that floats on the surface of cells.

In the flotation process of air bubbles formed initially has small size and some are attached to the surface of mineral particles. Furthermore, another air bubble formed next to join the existing air bubbles and form air bubbles with a larger size, so as to have sufficient lift to lift mineral particles to the surface. The mechanism of attachment of mineral particles in the air bubbles inside the tank during the flotation process flotation occurs when the hydrodynamic forces and the forces of interaction between mineral particles with air bubbles, resulting in collisions with air bubbles and mineral particles occurs attachment of mineral particles with air bubbles.

From the results ofbubblefrothflotation processthat resembles acolored foam detergent concentrate metallic orescarryinggold-coppermineral-ladenis thenuptothe tubshelter, and foam concentrate that has been lifted from the drain into the upper lip and into the trough flotation machine is in use as a valuable mineral collection.

In order fortheflotationprocesscan take placebyeithermeansof attachmentof particlestoairbubbleslasteduntilthetop edge of theflotationcell,it is necessary toconsiderthe followingmatters :

fine gold recovery by flotation

fine gold recovery by flotation

In view of the tendency of fine gold recover by floating and the dominance of heavy minerals, the prospects for flotation were worthy of re-investigation, particularly with the new tools of the air sparged hydrocyclone (ASH) and the flotation column. Although gold readily and naturally forms a hydrophobic film, dithiophosphates have been advanced as more selective for gold than xanthates and float gold up to 200 micron in size despite its density.

Our initial investigations in the mechanical flotation cell were with a wide variety of samples, including beach sands and alluvium. The more definitive work with the ASH and the flotation column has been on three samples. Some of the beach sands are cemented and can require vigorous attritioning to achieve liberation of the gold. All three samples had been scrubbed or attritioned, and were a product of either spiral or Knelson bowl concentration. There was no plus 0.5 mm material; 90% was between 50 and 200 microns and the gold was distributed fairly evenly through the size range.

Initially our interest arose from a need, in the laboratory, to pre-concentrate a sample prior to analysis, either to enhance the confidence in the analysis or to facilitate morphological characterisation. A Wemeco Fagergren machine with a 2 litre cell was used. Recoveries were somewhat inconsistent but it seemed more suitable for well- sorted (or de-slimed) feeds such as beach sands.

Subsequent client-funded work confirmed this, with consistently high recoveries (over 80%) and low mass yields (less than 5%). Rougher, scavenger and cleaner circuits appeared necessary to achieve the very high recoveries required for gold and sufficiently high concentration ratios to facilitate the final recovery step. It was becoming evident, however, that a conventional mechanical cell was inappropriate for such a coarse (minus 0.5 mm) and dense feed. The impeller had to turn at its maximum speed of 2600 rpm to maintain suspension. This was confirmed by our unsuccessful attempts to modify a two cell No. 5 Denver Sub-A mechanical flotation machine. The fixed speed impellors (1400 rpm) did not provide sufficient agitation. Modifications could only increase the speed to 2000 rpm, which was again not sufficient to maintain suspension. Higher speeds caused the onset of unacceptable stress. Whereas a mechanical cell could be designed for such a feed and the impellor speed could be minimised, the abrasive nature of the slurry could present tremendous wear problems.

The ASH embraces the cyclone principle but uses a long parallel porous body through which air is forced to carry hydrophobic minerals in a conditioned feed to the vortex and thus to overflow (Figure 1). Details of its design and fundamentals was developed for the flotation of fine particles.

However its abilityto recover gold certainly warranted its investigation. The gold recovery and mass yield were 81% and 0.5% respectively. Both the sand and the gold were fine and the sand less dense than the three beach sands. An ASH2 (i.e. 2 diameter) was installed in open circuit (Figure 1). A 250 litre reservoir of feed pulp was conditioned and maintained in suspension with a pump and closed loop, and which enabled a 5 minute run at 50 l/min (-0.3tph at 10% solids). Details on the ASH configuration, operating conditions, precedures and results are described and a selection of results in Table 1.

The conditions used to float the gold were used as a starting point, but were not the most appropriate. A lower pulp density and feed rate of 10% and 50 l/min respectively were necessary. Feed inlet pressure was by far the most sensitive parameter, at least in terms of mass yield. Low yields could only be obtained at pressures approaching the collapse of the vortex. Whereas the flotation reagents used were reasonably effective (xanthates and kerosene), those used in the work described above using the mechanical cell (dithiophosphate and MIBC) were quite ineffective.

The parameters examined were conditioning time (3 to 20 mins), pulp density (4 to 10%), inlet pressure (4 to 8 psi), pedestal diameter (3 sizes) and to a small extent reagent levels. Rarely did the gold recovery exceed 50% even with high mass yields. The gold remaining in the tailing was slightly coarser than that in the concentrate but could be recovered by re-passing. Cumulative gold recoveries of 80 to 90% could be obtained from three passes, but the cumulative mass yield was as high as 20%. Clearly therefore we have not been able to obtain the same success as for river sand. The coarser and denser nature of these sands is the likely cause.

The residence period is of the order of one second. Its higher capacity and therefore lower cost, relative to mechanical cells, highlight its potential. However long conditioning times (20 minutes in the case of the river sand), if indeed they are necessary, detract from the advantage.

The flotation column was developed for the improved recovery of fine particles too, and is finding commercial application. It embodies principles which are, with modification, appropriate for the flotation of gold from coarse dense sands. A small laboratory column, 1 m high by 50 mm in diameter (Figure 2), was constructed. It differed from other columns by having a stainless

steel sintered disc base through which fluidising water was passed to facilitate discharge by maintaining the sands in suspension. The external bubble generator was based on a design of the US Bureau of Mines (Anon 1987).

The feed rate to the column was as low as 2 l/min, which is a very low flow for reliable delivery where the feed is a slurry of coarse dense sand. It was achieved, in the system shown in Figure 2, by ensuring that the rising line velocity was well above the maximum terminal velocity and by using a distributor to split the pulp. Feed rates of up to 12 l/min were possible. The column was fed in open circuit from the 250 litre reservoir maintained in suspension by a pump and closed loop. In theory, runs of 2 hours duration were possible, but because of declining pulp density (max of 15% solids) they usually lasted about an hour. For this reason and because gold tended to concentrate in the reservoir, balances were based on what passed through the column.

Since there was close control of all inflows and outflows, the column could be readily balanced for the duration of the run, maintaining the froth/slurry interface (somewhat diffuse anyway) just above the feed port.

The reagent regime was similar to that used in the ASH, viz a xanthate collector (80 g/t), kerosene (50 g/t) and a propyleneglycol/methyl ether frother (100 g/t). The dithiophosphate was again not very effective.

The experimental programme was not exhaustive, since the material was being used to commission and evaluate the column. Many important parameters were not investigated, such as bubble size, air flow and feed rate. Nevertheless the results are very encouraging, with gold recoveries over 90% and mass yields less than 1% for a single pass. A selection of results is shown in Table 2.

There was evidence that even better results would be obtained under more quiescent conditions, i.e. at lower airflows. The specific slurry rates and specific solids rates per unit volume were considerably greater than those reported elsewhere but on different materials. In this sense, therefore, optimum performance can be expected to be even better.

gold flotation

gold flotation

Though the gold recovery methods previously discussed usually catch the coarser particles of sulphides in the ore and thus indirectly recover some of the gold associated with these and other heavy minerals, they are not primarily designed for sulphide recovery. Where a high sulphide recovery is demanded, flotation methods are now in general use, but in the days before flotation was known, a large part of the worlds gold was recovered by concentrating the gold-bearing sulphides on tables and smelting or regrinding and amalgamating the product.Though the modern trend is away from the use of tables, because flotation is so much more efficient.

The flotation process, which is today so extensively used for the concentration of base-metal sulphide ores and is finding increased use in many other fields. In1932flotation plants began to be installed for the treatment of gold and silver ores as a substitute for or in conjunction with cyanidation.

The principles involved and the rather elaborate physicochemical theories advanced to account for the selective separations obtained are beyond the scope of this book. Suffice it to say that in general the sulphides are air-filmed and ufloated to be removed as a froth from the surface of the pulp while the nonsulphide gangue remains in suspension, or sinks, as the expression is, for discharge from the side or end of the machine.

For more complete information reference is made to Taggarts Hand book of Mineral Dressing, 1945; Gaudins Flotation and Principles of Mineral Dressing; I. W. Warks Principles of Flotation; and the numerous papers on the subject published by the A.I.M.E. and U.S. Bureau of Mines.

Flotation machines can be classed roughly into mechanical and pneumatic types. The first employ mechanically operated impellers or rotorsfor agitating and aerating the pulps, with or without a supplementary compressed-air supply. Best known of these are the Mineral Separation, the Fagergren, the Agitair, and the Massco-Fahrenwald.

Pneumatic cells use no mechanical agitation (except the Macintosh, now obsolete) and depend on compressed air to supply the bubble structure and tohold the pulp in suspension. Well-known makes include theCallow and MacIntosh (no longer manufactured) the Southwestern, and the Steffensen, the last, as shown in the cross-sectional view in Fig. 47, utilizing the air-lift principle, with the shearing of large bubbles as the air is forced from a central perforated bell through a series of diffuser plates.

The number and size of flotation cells required for any given installation are readily determinedif the problem is looked upon as a matter of retention time for a certain total volume of pulp. The pulp flow in cubic feet per minute is determined from the formula

For ordinary ratios of concentration the effect on cell capacity of concentrate (or froth) removal can be neglected, but where a high proportion of the feed is taken off as concentrates, or where middlings are removed for retreatment in a separate circuit, due allowance should be made for reduced flow and, in consequence, increased detention time toward the tail end of a string of cells. Not less than a series of four cells and preferably six or more cells should be used in any roughing section in order to prevent short-circuiting.

It is not intended here to discuss the subject of flotation reagents in anydetail. The subject is a large one with a comprehensive technical and patent literature. Research leading to the development of new reagents and to our understanding of the mechanism involved has been largely in the hands of academic institutions and the manufacturers of chemical products.

Recent work reported by A. M. Gaudin on the use of Radioactive Tracers in Milling Research described, for instance, the use of a flotation reagents containing radioactive carbon to determine the extent of collector adsorption. The bubble machine devised to measure the angle of contact of air bubbles on collector-treated mineral surfaces has been extensively used for determining the theoretical value of various reagents as flotation collectors, but for the most part the actual reagent combination in use in commercial plants is usually the result of trial-and-error methods.

The following is a brief discussion of the reagents ordinarily used for the flotation of gold and silver ores prepared from notes submitted by S. J. Swainson and N. Hedley of the American Cyanamid Company.

Conditioning agents are commonly used, especially when the ores are partly oxidized. Soda ash is the most widely used regulator of alkalinity. Lime should not be used because it is a depressor of free gold and inhibits pyrite flotation. Sodium sulphide is often helpful in the flotation of partly oxidized sulphides but must be used with caution because of its depressing action on free gold. Copper sulphate is frequently helpful in accelerating the flotation of pyrite and arsenopyrite. In rare instances sulphuric acid may be necessary, but the use of it is limited to ores containing no lime. Ammo-phos, a crude monoammonium phosphate, is sometimes used in the flotation of oxidized gold ores. It has the effect of flocculating iron oxide slime, thus improving the grade of concentrate. Sodium silicate, a dispersing agent, is also useful for overcoming gangue-slime interference.

Promoters or Collectors. The commonly used promoters or collectors are Aerofloat reagents and the xanthates. The most effective promoter of free gold is Aerofloat flotation reagent 208. When auriferous pyrite is present, this reagent and reagent 301 constitute the most effective promoter combination. The latter is a higher xanthate which is a strong and non-selective promoter of all sulphides. Amyl and butyl xanthates are also widely used. Ethyl xanthate is not so commonly used as the higher xanthates for this type of flotation.

The liquid flotation reagents such as Aerofloat 15, 25, and 31 are commonly used in conjunction with the xanthates. These reagents possess both promoter and frother properties. When malachite and azurite are present, reagent 425 is often a useful promoter. This reagent was developed especially for the flotation of oxidized copper ores.

The amount of these promoters varies considerably. If the ore is partly oxidized, it may be necessary to use as much as 0.30 to 0.40 lb. of promoter perton of ore. In the case of clean ores, as little as 0.05 lb. may be enough. The promoter requirement of an average ore will usually approximate 0.20 lb.

The commonly used frothers are steam-distilled pine oil, cresylic acid, and higher alcohols. The third mentioned, known as duPont frothers, have recently come into use. They produce a somewhat more tender and evanescent froth than pine oil or cresylic acid; consequently they have less tendency to float gangue, particularly in circuits alkaline with lime. The duPont frothers are highly active frothing agents; therefore it is rarely necessary to use more than a few hundredths of a pound per ton of ore.

When coarse sulphides and moderately coarse gold (65 mesh) must be floated, froth modifiers such as Barrett Nos. 4 and 634, of hardwood creosote, are usually necessary. The function of these so-called froth modifiers is to give more stable froth having greater carrying power.

The conditioning agents used for silver ores are the same as those for gold ores. Soda ash is a commonly used pH regulator. It aids the flotation of galena and silver sulphides. When the silver and lead minerals are in the oxidized state, sodium sulphide is helpful, but it should not be added until after the sulphide minerals have been floated, because sodium sulphide inhibits flotation of the silver sulphide minerals.

Aerofloat 25 and 31 are effective promoters for silver sulphides, sulphantimonites, and sulpharsenites, as well as for native silver. When galena is present, No. 31 is preferable to No. 25 because it is a more powerful galena promoter. Higher xanthates, such as American Cyanamid reagent 301 and amyl and butyl xanthates, are beneficial when pyrite must be recovered. When the ore contains oxidized lead minerals, such as angle-site and cerussite, sodium sulphide and one of the higher xanthates may be used. In some instances reagent 404 effects high recovery of these minerals without the use of a sulphidizing agent.Silver ores require the same frothers as gold oresviz., pine oil, cresylic acid or duPont frothers.

Aero, Ammo-phos, and Aerofloat are registered trade-marks applied to products manufactured by this company. The Great Western Electro-Chemical Company, California, makes amyl xanthate, butyl xanthate, potassium xanthate, and sodium xanthate. In the United States these reagents are used on the gold ores of California and Colorado and in Canada on the gold ores and sulphides of Ontario and Quebec.

Flotation reagents of the Naval Stores Division of the Hercules Powder Company are as follows: Yarmor F pine oil, a frother for floating simple and complex ores; Risor pine oil, for recovering sulphides by bulk flotation; Tarol a toughener of froth, generally used in small amount with Yarmor F, but with some semioxidized ores where high recovery is essential yet the grade of concentrate not so important, Tarol does good work; Tarol a frother for floating certain oxide minerals, but it can be used in selective flotation of sulphide minerals and in bulk flotation where tough frothis desirable; Solvenol, for the floating of graphite in conjunction with Yarmor F.

The statement has come to the attention of the American Cyanamid Company that organic flotation reagents, such as xanthates, even in the small amounts used in flotation, cause reprecipitation of gold from pregnant cyanide solutions. The ore-dressing laboratory of this company is studying the question, and preliminary results indicate that this statement is unfounded. The addition of xanthate, in the amount usually found in flotation circuits, does not precipitate gold from a pregnant cyanide solution containing the normal amount of cyanide and lime.

Valueless slime, in addition to its detrimental effect in coating gold-bearing sulphide, thereby limiting or preventing its flotation, also becomes mixed with the flotation concentrate and lowers its value. Sometimes the problem in flotation is that, although the gold is floatable, the concentrate product is of too low grade. Talc, slate, clay, oxides of iron, and manganese or carbonaceousmatter in ores early form slime in a mill, without fine crushing. Such primary slime, according to E. S. Leaver and J. A. Woolf of the U.S. Bureau of Mines, interferes with the proper selectivity of the associated minerals and causes slime interference. The tendency of primary slime is to float readily or to remain in suspension and be carried over into the concentrate. Preliminary removal and washing of this primary slime before fine crushing is one method of dealing with it. At the Idaho-Maryland mill, Grass Valley, Calif., starch is regularly used as a depressant during flotation. Flotation tests using starch were made on a quartz ore containing carbonaceous schist from the Argonaut mine, Jackson, Calif.; a talcose ore from the Idaho-Maryland mine mentioned; a talcose-clayey ore from Gold Range, Nev.; a siliceous, iron and manganese oxide ore from the Baboquivari district, Nevada; carbonaceous and aluminous slime from the Mother Lode and some synthetic ores. The conclusions from the foregoing tests were in part as follows:

It acts first on the slime; then, if a sufficient excess of starch is present, it will cause some depression of sulphides and metallic gold, either by wetting out or by producing an extremely brittle froth. Therefore, care must be taken in regulating the amount of starch added to obtain the maximum depression of the slime commensurate with high recovery of the gold. In this, as in all other phases of flotation, each ore presents an individual problem and must be so studied.

It wasdescribe by the use of 600 series of flotation reagents which were developed primarily for the purpose of depressing carbonaceous and siliceous slimes in the flotation of gold ores. Carbonaceous material not only greatly increases the bulk and moisture content of a flotation concentrate, but its presence makes cyanidation of the concentrate difficult or impossible owing to reprecipitation of the gold during treatment.

In the treatment of an auriferous sulphide ore associated with carbonaceous shale from South Africa, up to 77 per cent of the carbon was eliminated by the use of 1 lb. per ton of reagent 637 with a 90.5 per cent gold recovery at 20.4:1 ratio of concentration.

A gold carbonaceous sulphide ore from California carrying free gold yielded a 93 per cent recovery into a concentrate at 14.4:1 to ratio of concentration after conditioning with 0.50 lb. per ton of reagent 645.

In each case the ore was ground to about 70 per cent minus 200 mesh and conditioned at 22 per cent solids with the reagents as indicated. Flotation reagents included reagents 301 and 208 and pine oil. In the second case some soda ash and copper sulphate where also used.

It is obvious that the most suitable treatment for ores carrying gold and silver associated with pyrite and other iron sulphides, arsenopyrite or stibnite, will depend on the type of association. Cyanidation is usually the most suitable process, but it often necessitates grinding ore to a fine size to release the gold and silver. Where it is possible to obtain a good recovery by flotation in a concentrate carrying most of the pyrite or other sulphides, it is often more economical to adopt this method, regrinding only the comparatively small bulk of concentrate prior to the leaching operation.

That the trend over the last 10 years has been in this direction will be noted from the numerous examples of such flow sheets in Canada and Australia (see Chap. XV). A number of plants formerly using all-cyanidation have converted to the combined process.

The suitability of the method involving fine grinding and flotation with treatment of the concentrate and rejection of the remainder should receive careful study in the laboratory and in a pilot plant. Mclntyre-Porcupine ran a 150-ton plant for a year before deciding to build its 2400-ton mill. Comparative figures given by J. J. Denny in E. and M. J., November, 1933, on the results obtained by the all-sliming, C.C.D. process formerly used and the later combination of flotation and concentrate treatment showed a saving of 12.1 cents per ton in treatment cost and a decrease of 15 cents per ton in the residue, a total of 27.1 cents per ton in favor of the new treatment.

Flotation may also prove to be the more economical process for the ore containing such minerals as stibnite, copper-bearing sulphides, tellurides,and others which require roasting before cyanidation, because this reduces the tonnage passing through the furnace.

Even when recovery of gold and silver from such ores by flotation is low, it may be advantageous still to float off the minerals that interfere with cyanidation, roasting, and leaching or possibly to smelt the concentrate for extraction of its precious metals. Cyanidation of the flotation tailing follows, this being simpler and cheaper because of prior removal of the cyanicides.

It is a good practice to recover as much of the gold and silver as possible in the grinding circuit by amalgamation, corduroy strakes, or other gravity means to prevent their accumulation in the classifier; otherwise gold that is too coarse to float may escape from the grinding section into the flotation circuit where it will pass into the tailing and be lost.

To prevent this, several companies including the Mclntyre-Porcupine at Timmins, Ontario, have inserted a combination of flotation cell and hydraulic cone in their tube-mill classifier circuits. At the Mclntyre- Porcupine, according to J. J. Denny in E. and M. J., November, 1933, this cell is a 500 Sub-A type. The total pulp discharged from each tube mill passes through 4-meshscreens which are attached to the end of the mills. The undersize goes to the flotation cell, and the oversize to the classifiers. Tailing from the cell flows to the classifiers, and the flotation concentrate joins the concentrate stream from .the main flotation circuit. The purpose of the hydraulic attachment is to remove gold that is too coarse to float, thus avoiding an accumulation in the tube-mill circuit. The cones have increased recovery from 60 to 75 per cent. Every 24 hr. the tube-mill discharge is diverted to the classifiers. Water is added for 15 min. to separate the gangue in the cells from the high-grade concentrate, after which a product consisting of sulphides and coarse gold is removed through a 4-in. plug valve equipped with a locking device. Each day approximately 400 lb. of material worth $2000 to $3000 is recovered. This is transferred to a tube mill in the cyanide circuit,with no evident increase in the value of the cyanide residue. The object of this arrangement is, of course, primarily to deplete the circulating load of an accumulation of free gold and heavy sulphides.

Flotation is used to recover residual gold-bearing sulphides and tellurides. The Lake Shore mill retreatment plant is an interesting example of this technique. The problem here was, of course, to overcome by chemical treatment the depressing action of the alkaline cyanide circuit on the sulphides. A full discussion of this and of the somewhat controversial subject as to whether flotation should in such an instance be carried out before, or after cyanidation will be found in J. E. Williamsons paper Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant elsewhere referred to. Summing up the specific considerations governing the choice oftreatment, the author says:

Incidental matters that influenced the choice of treatment scheme included the realization that preliminary flotation would have involved two separate treatment circuits with additional steps of thickening and filtration following the flotation. Furthermore, in the conditioning method evolved, as much as 60 per cent of the dissolved values in the cyanide tailings were precipitated and recovered.

There are, however, cases where flotation equipment was put in for the purpose of recovering the gold in a concentrate and rejecting the tailing only to find that the tailing was too valuable to waste and had finally to be cyanided before discarding.

It is generally true that cyanidation is capable of producing a tailing of lower gold content than flotation. At a price of $35 per ounce for gold this fact is of much greater importance than when gold was valued at $20.67 per ounce. The possible gold loss in the residue to be discarded will influence the choice of a method of treatment.

sulphide gold flotation concentrate cil leaching

sulphide gold flotation concentrate cil leaching

Gravity Gold Rougher Concentrate Flotation CIL Leaching & Refining: This mineral processing plant is to recover gold from sulphide ore deposits. Provided here are all major equipment for a plant arranged to recover gravity gold, float a gold rich sulphide concentrate, liberate the sulphide hosted gold with a light regrind and dissolved its precious metals using the Carbon-in-Leach process. The CIL process with activated carbon is best suited for ores with no or some levels of preg-robbing carbonaceous material. The initial gravity circuit will extract any coarse GRG gold and silver. More difficult (refractory) ores may need deeper ultra-fine regrind prior to leaching as well and more elaborate processing, not included here.

This simple metallurgical concentrator includes single stage crushing, conveying, primary grinding, spiral classification, gravity concentration, slurry pumping, rougher and cleaner (optional) flotation banks, regrind mill, cyanide leaching tanks, loaded carbon stripping and regeneration as well as refining facility.

This is a standard process plant which includes only the major components of the complete metallurgical flowsheet. A detailed engineering study is required to identify unforeseen omissions that may be required to design the optimum plant.

Process development testwork and detailed engineering are essential services 911Metallurgy Corp. offers separately. The equipment package described herein does not include any permitting, infrastructure, foundation, electrical, assembly, reagents/supplies or commissioning. These are all additional paid-for services we do offer if you need them.

Using gravity to supplement either flotation or cyanidation is a well established practice in the gold industry. It differs from other recovery circuits in that most of the gold recovered by gravity would be recovered in any case by the circuit downstream, be it flotation or cyanidation. The economic justification of gravity is therefore based on small margins (for example, a net smelter return of gravity gold of 99%, as opposed to 94% for flotation). It was easily demonstrated when either flotation or cyanidation were relatively inefficient processes, and labour costs low (as gravity can be labour intensive). Over the past fourty years, the introduction of better flotation machines (flash, column, high capacity), more effective collectors, and better control systems has increased flotations metallurgical performance, thereby decreasing the incentive for gravity recovery. Cyanidation technology has undergone similar changes, with the advent of activated carbon, oxygen and lead nitrate addition, and improved impeller design.

Today, gravity can remain an attractive option only inasmuch as it can be implemented with very low capital and operating costs. This has resulted in a relative shift away from gold gravity recovery (except for alluvial deposits), in the seventies and eighties. For example, as of the early nineties, gold gravity recovery has disappeared from the typical flow sheet. The advent of the iCON Concentrator, at the beginning of the eighties, foreshadowed a resurgence of gravity recovery, as golds grinding and classification behaviour makes it possible to achieve adequate gold recoveries with very simple, iCON based, gravity circuits.

bulk - monometallic flotation plant

bulk - monometallic flotation plant

This complete process plant is for recovering metal monometallic ore deposits. This applies, but is not limited to, the recovery by froth flotation of most base metals: Copper, Lead, Zinc, Cobalt, Nickel, Molybdenum, Pyrite, Pyrite-Gold, Silver-Sulphides. This process will also recover silver and/or gold associated with sulphide minerals. If your ore only contains one metal, you will be able to selectively float it away from the waste material. If you have more than one metal, you will be able to recover them in bulk into one single/common concentrate.

This simple metallurgical concentrator includes single stage crushing, conveying, primary grinding, spiral classification, slurry pumping, rougher flotation and 3 stages of concentrate dilution cleaning. A regrind circuit can be added upon request.

This is a standard process plant which includes only the major components of the complete metallurgical flowsheet. A detailed engineering study is required to identify unforeseen omissions that may be required to design the optimum plant.

Process development testwork and detailed engineering are essential services 911Metallurgy Corp. offers separately. The equipment package described herein does not include any permitting, infrastructure, foundation, electrical, assembly, reagents/supplies or commissioning. These are all additional paid-for services we do offer if you need them.

The flotation process is now used very extensively on gold ores, not only for concentrating the values without other processes, but also in conjunction with amalgamation, gravity concentration, or cyanidation leaching, to improve recoveries and to give lower treatment costs. In many instances, it is most profitable to recover the gold in the form of a concentrate, and either ship this product to a smelter, or amalgamate or cyanide the concentrates at the mine. In other cases, flotation has been utilized to recover the values remaining after amalgamation or gravity concentration. Today flotation is also being used to remove copper or other cyanicides prior to the cyanide treatment, and in many cases to recover a concentrate containing the values in the sulphides that remain in the tailings from the cyanide plant. These values that are contained in the sulphides can sometimes be recovered by further grinding and cyanidation or amalgamation treatment, or, if refractory, by roasting and cyanidation or by direct shipment to the smelter.

The present day success of many small gold mines is directly traceable to the use of flotation, as a flotation plant can be constructed for less than half the cost of a cyanide plant. The use of flotation has made it possible for many mines to operate where the ore bodies are complex and would not have justified the capital expenditure required for other methods of treatment.

Flotation Plants, illustrated in the following pages, show the compactness and simplicity of the modern milling plant. These have been designed to meet present day needs of standardized and economical milling plants with a flexible flowsheet that can be altered slightly so that a wide variety of ores can be efficiently treated. They are built to handle any tonnage desired, but 25-35 tons, 50-65 tons, and 100-125 tons per day are most common. Standard equipment is used throughout, and only machines that have proved successful in installations the world over, have been included. While the entire milling plants have been shown, individual machines can be easily incorporated in existing plants.

The selection of the equipment for these mills will naturally depend upon the flowsheet. Metallurgical tests are necessary to determine this flowsheet. This step should be considered, asthese ore tests will show the most economical treatment process, as well as indicating the recovery, grade of concentrates, and economic possibilities of the project.

ORE TREATMENT PROCESSES have been improving rapidly during the last few years, and this advancement is due primarily to the improvements made in the flotation process. The field for flotation has been broadened to include the concentration of most non-metallic minerals, and its efficiency has been increased on base metal ores.

Flowsheet 2, shown here, is widely used where straight flotation is the indicated method of treatment. This consists of a grizzly, followed by single stage crushing, and fine grinding in a closedball mill-classifier circuit. The classifier overflow passes to a Conditioner, and then to a Sub-A Flotation Machine, where a high grade concentrate and low tailing are made. The flotation tailing passes over pilot tables to act as an indicator for the operator, and to recover any oxidized or tarnished minerals that are not floated. A concentrate thickener and filter can be added to simplify the preparation of the concentrate for shipment. This flowsheet is used where a single concentrate is made and where the ore requires extreme fine grinding to liberate the values.

Where it is necessary to separate and recover two minerals from an ore to make marketable concentrates, Flowsheet 3B is generally used. The widest application of this circuit is on lead zinc ores, and the Unit Flotation Cell used in the fine grinding circuit is of great value in recovering the lead material at a coarse mesh as soon as freed, thus preventing overgrinding of the mineral particles and avoiding slime losses.

The classifier overflow passes to the first conditioner and then to the first series of Sub-A Flotation Cells where the first mineral is recovered in a high grade concentrate. The tailings from this operation pass to the second conditioner and second series of flotation cells where the second mineral is recovered. The tailings from this circuit pass over Pilot Tables so that the operator can continually check mill results. Often these tables are not only used for pilot work, but also to recover a pyrite concentrate when some precious metals are associated with this mineral.

It is often desirable to float out a coarse flotation product, regrind these middlings and then refloat in a second flotation unit at a fine mesh. The ability of the Sub-A Flotation Machine to handle a coarse feed as well as the fine feed makes this flexibility in Flowsheet 3B possible.

Flowsheet 2B, shown above, is universally used for treating gold and silver ores that are associated with any of the various base metals and in fact should be used on any flotation ore. Mill efficiency is always enhanced by removing the minerals as soon as grinding has freed them from the gangue. In this flowsheet is shown the FlashFlotation Cell where the minerals in the ball mill discharge are recovered when freed, and are not returned to the ball mill where they would be overground. In over-grinding the finer mineral values are coated by colloids and become difficult to float or to concentrate in a high grade product.

It is often preferable to remove a gravity concentrate from the grinding circuit. This is particularly true on oxidized or rusty gold ores where a reasonable percentage of the values are in the form of metallics which free readily during the grindingoperation. A coarse concentrating table may be substituted in the grinding circuit for the Sub- A FlashFlotation Cell, but a pump or elevator is required to complete this circuit, and great care must be used to prevent excessive dilution from the wash water necessary for the operation of the table.

Since its development, the Mineral Jig has been widely used to recover a high grade gravity concentrate, as the objections to the use of a concentrating table are overcome with this compact and efficient unit.

A (Selective) Mineral Jig may be installed in the grinding circuit in place of the unit cell illustrated in above Flowsheet 2B. As in the case with the Unit Flotation Cell, the Mineral Jig can be installed without pumps or elevators, and requires very little floor space. Much less dilution is required for the jig than is normally required for a concentrating table in the grinding circuit, so that there is not the drawback encountered in other types of gravity concentration machines at this point in the flowsheet. Additional water is added to the(Selective) Mineral Jig discharge before subsequent classification. The mills shown in the following drawings have been reproduced in exact proportion so that the comparative sizes can be readily visualized. While all of these drawings show the standard Flowsheet 2B, either a coarse concentrating table, or a (Selective) Mineral Jig, can be utilized with minor alterations. Larger Mills can be supplied, and as all of the equipment in thesevarious plants is standard, it can be used to advant-age if a larger tonnage mill is later warranted.

In these mill drawings are shown a Thickener, Diaphragm Pump, and Filter for concentrate dewatering. These are items that can be omitted to keep first cost at a minimum and added later when finances permit, for these machines will quickly pay for themselves.

The 50 to 65 Ton Mill is the most practical size for the average small milling plant. The Jaw Crusher (forced feed) produces enough crushed ore for the ball mill on one eight hour shift to run the balance of the mill for twenty-four hours. Oversize primary crushers are recommended for mosteconomical results. Note the flexibility of the flowsheet and the gravity flow through the mill which eliminates elevators, conveyors and pumps (except for middling products).

The 100 to 125 Ton Mill is arranged along the same standard lines as the smaller mills, and a large oversize forced feed crusher is recommended for primary crushing. An intermediate crusher can be installed later if necessary. Changes can be easily made according to your local conditions. Here too, a Selective Mineral Jig may be advisable on gold ores.

All machines have motor V Rope drive, but with slight changes belt drives can be furnished. Drives and flowsheet can be changed to meet your conditions, and by means of the sand pump, products can be returned to any part of the flowsheet.

laboratory flotation column

laboratory flotation column

The batch column flotation results presented here were previously published. These data, however, were important in establishing the conditions for optimum continuous column flotation of the Fish Creek fluorite ore. Batch column flotation tests were conducted on column length, feed injection location, tailing recirculation, froth depth, wash water additions, and particle size fractions.

Since the inception of flotation columns in the early 1960s, column length has been a concern to commercial mineral processing plants anticipating installation and operation of flotation columns. A column flotation cell is free from violent agitation. Feed and tailings slurry flow rates and particle settling rates affect the retention time of the particles in the column.

The collection zone has its upper boundary at the feed injection port and extends downward to the base of the column. This zone must have sufficient length to provide adequate retention time for the settling particles to attach to the rising bubbles. Column length design theory is based on this concept. Additional column length must be included for the upper three column zones as prescribed by the particular mineral system needs. Most work backing this theory has been performed on copper-molybdenum separations.

The effect of column length variations on the fluorite ore was studied by shortening the column, while maintaining a constant ratio of each column zone within physical limitations of the equipment and observing grade and recovery fluctuations. As the column flotation cell was shortened, recoveries of fluorite decreased (fig. B-1). Retention time was calculated for plug flow conditions based on the collection zone volume and tailings flow rate. Fluorite recovery decreased because particle retention time was not sufficient as the collection zone was shortened by decreasing the column length.

Fluorite grades increased with decreasing column length because only the particles with sufficient hydrophobicity to achieve bubble attachment were reported to the concentrate stream. As the column was shortened, the particle retention time decreased causing smaller fractions of the more liberated fluorite to be collected, while also reducing the amount of gangue that was either entrained or collected to the froth (fig. B-2).

Fluorite recovery gradually decreased as the feed injection port was moved closer to the base of the column (fig. B-3). In essence, feed injection location is directly linked to particle retention time in the collection zone of the column. As the feed injection location was moved towards the base of the column, the length of the collection zone decreased, reducing the particle retention time, and decreased fluorite recovery resulted (fig. B-4).

Fluorite concentrate grades increased as feed injection location approached the base of the column. Lowering the vertical position of the feed slurry injection port is equivalent to shortening the length of the collection zone. This accounts for the similar response of fluorite recovery and grade to variations in either column length or feed slurry injection location. The increased fluorite grade observed as feed injection location moved towards the base of the column were due to the added concentration effects of the upper three column zones. Unlike the column length variation tests, the upper two zones remained at constant length, while the pulp phase cleaning zone increased in length as the feed injection location approached the column base.

Fluorite grade and recovery variations were investigated as a portion of the tailings stream was recirculated at different rates during column flotation. The tailings recirculation flow rates were converted to superficial pulp velocities based on plug flow conditions in the collection zone volume. The resultant trend is given in figure B-5.

Fluorite recovery suffered and the grade was enhanced as the recirculation velocity was increased. Since short circuiting of some of the feed slurry, caused by axial mixing, increases as recirculation velocity increases, these trends were expected. Furthermore, trending of data from this investigation versus superficial liquid velocity showed no correlation, therefore, the reduced recoveries and increased grades were concluded to result from the decrease in the particle retention time distribution. Under these conditions, only those particles with the strongest adsorption energies have sufficient time to attach to the rising bubbles.

These results show that the collection zone must be lengthened to provide sufficient particle retention time as the particle retention time distribution is broadened and to compensate for increase in pulp mixing.

A series of tests were conducted on the fluorite ore to determine the effect of froth depth on mineral grades and recoveries. Froth depth had a discernible effect on fluorite concentrate grades; as froth depth increased, fluorite grades increased (fig. B-6). Fluorite upgrading occurred in the froth phase cleaning and pulp-froth interfacial zones. The froth phase is more efficient than the pulp phase in upgrading the concentrate. At greater froth depths, more fluorite cleaning took place and higher grade concentrates were obtained.

Fluorite recovery data were scattered. No adequate trend could relate the recovery data with greater accuracy than a line with a slope of nearly zero. Although the source of these fluctuations could not be identified, no direct connection with froth depth variations could be made. It was concluded that froth depth had no primary correlation with fluorite recovery.

Since froth depth enhanced fluorite grades without hindering recoveries, froth depth should be maintained at as great a depth as possible, while maintaining sufficient column length for the collection zones to provide the necessary particle retention time to maintain mineral recoveries.

The principal reason for using wash water in column flotation systems is to increase the grade of the recovered concentrate by displacing entrained hydrophilic (gangue) particles that report to the froth phase. Wash water additions have also been observed to aid in stabilizing the froth bed. Tests were performed to determine the effect of changes in wash water addition rates on column

flotation grades and recoveries of the Fish Creek fluorite ore. Wash water was introduced through a spray nozzle located 1 in above the top of the column. The addition rates were normalized by dividing by the volumetric feed slurry flow rates to the column.

Wash water additions affected fluorite grade and recovery during column flotation in a complex manner (fig. B-7). Increasing wash water additions from 0 to 6 pet of the volumetric feed slurry flow rate increased fluorite grade, but decreased recovery. Increasing wash water additions from 6 pct to approximately 35 pct improved fluorite recovery and decreased fluorite grade. Above 35 pct wash water addition, fluorite grade again increased, while recovery remained approximately constant.

The complex effect of wash water additions on fluorite flotation in the column may be attributed to the twofold nature of column wash water in removing gangue material reporting to the froth, while fluidizing the froth bed to prevent mineral overloading. As detailed in figure B-7, the optimum wash water addition rate was 6 pct. However, wash water flow rates in excess of 40 pct produced grades and recoveries that approached those at 6 pct. These additions may not be feasible due to increased water consumption, system dilution, and downstream materials handling problems.

To quantify the separation efficiency of column flotation compared with conventional flotation, particle size distribution analyses were performed on products obtained from each method. The study was conducted by extracting a portion of the conditioned column feed slurry and sending it to a conventional batch flotation process. The column and conventional flotation products were sized using Tyler 48-, 65-, 80-, 100-, 150-, 200-, 270-, 325-, and 400-mesh screens.

Column rougher flotation produced substantially higher grade concentrates than did conventional flotation of the fluorite ore (fig. B-8). Column flotation fluorite grades were greater than those of conventional flotation for all size fractions.

Conventional flotation recoveries of fluorite were slightly higher than those of column flotation, except particles between 65 and 100 mesh (fig. B-9). Although conventional flotation provides a small increase in fluorite recovery for most particle size fractions, column flotation still holds the advantage because conventional fluorite recoveries would fall well below column fluorite recoveries in cleaning stages that would be necessary to meet the fluorite grades achieved in column flotation.

A 5.5-m (18-ft) by 6.1-cm (2.5-in) diam plexiglass flotation column with 0.64-cm (0.25-in) walls was used for testwork (Fig. 1). The column was composed of two 0.3-m (1-ft) and eight 0.6-m (2-ft) flanged sections, with 2.9-cm (1.125-in) diam ports located at 0.3 m (1 ft) intervals along its length. Solenoid-actuated sample valves with stainless steel spring loaded plungers were designed and constructed. Sampling valves were fitted into the 2.9-cm (1.125-in) diam ports located at 0.3~m (1-ft) intervals along the column length. These valves permitted rapid, consistent extraction of a representative pulp or froth sample. A concentric cylindrical overflow weir was mounted on the top section of the column. A conical bottom with a 1.3-cm (0.5-in) diam port was attached to the bottom section of the column.

The versatile column design allowed parameters such as column length, feed injection, tailings removal and recirculation, air and wash water injection, and sampling locations to be readily varied. The clear plexiglass construction allowed visual monitoring; consequently, response to flotation parameter changes could be observed immediately.

The conditioned feed material was pumped through flexible tubing from the conditioner to a port located 3.4 m (11 ft) from the column base using a peristaltic pump. Mineralized froth was continuously removed from the top of the column while tailings were pumped from the column base to filtration.

The fine bubble generation system was used to aerate the flotation column. The bubble generator was a 15.2-cm (6-in) tall by 5.1-cm (2-in) diam clear cylindrical plexiglass chamber with 2.5-cm 1-in) thick walls (Fig. 2). The generator had a removable plexiglass top attached to the generator body with an O-ring seal. A spacer supported a fritted glass disk in the chamber to prevent short circuiting of air. The remaining chamber volume was filled with 1-mm (0.04-in) diam glass beads to improve air-water contact. A section of 28-mesh screen was placed over each orifice to prevent loss of glass beads. House air (110 psig) was regulated to 60 psig and introduced through the side port of the bubble generator. Water, pressurized to 60 psig by a turbine blade pump, was introduced through the top bubble generator orifice. Air and water were mixed in the contact chamber; the pressurized mixture exited the chamber through the bottom port and was injected into the column through an aluminum tip with a 1-mm (0.0f-in) diam orifice. The generator design allowed for bubble size control from less than 0.1- to over 3-mm average bubble diameter by adjusting air and water flow rates and by adding Dowfroth 100 frother. Previous investigation showed coarse bubbles (3- to 5-mm diam) produced the best results on the Fish Creek fluorite ore, and they were therefore used for all parameter testing. An air flow rate of 4,500 cm/min and a water flow rate of 800 mL/min were used to generate the 3- to 5-mm-diam bubbles for this testwork.

gold flotation production line,gold flotation plant,gold flotation technology-beijing hot mining tech co ltd

gold flotation production line,gold flotation plant,gold flotation technology-beijing hot mining tech co ltd

The flotation method is a widely used technique for the recovery of gold from gold-containing copper ores, base metal ores, copper-nickel ores, platinum group ores and many other ores where other processes are not applicable. Flotation is also used for the removal of interfering impurities before hydrometallurgical treatment, for upgrading of low sulfide and refractory ores for further treatment. Flotation is considered to be the most cost-effective method for concentrating gold.

In this process of rock minerals that have been taken from the mine site and then destroyed by the machine to obtain a fine grain of sand to free metal-containing granules for further processing. In the destruction of mineral rocks of machine tools can use a stone crusher machine, so the minimum size of rock minerals can reach between 28 mesh.

At this stage after a mineral ore that is refined inserted into the machine agitator tank which is usually also called a flotation cell to produce a pulp slurry concentrate.Distilled water provision inserted into the flotation cell flotation machine is then run, examined the amount of initial pH and initial temperature. In the flotation tank, stirring with impellers, which are intended to produce turbulent motion of fluids (pulp), so that when inserted air flow will form air bubbles.In the pulp is then coupled collector-1,-2 collector and frother; flotation machine run back to the time varying adjustment, and examined the amount of the final pH and final temperature.

In the process flotation reagent which in use is a form of lime, bubble and collectors. Froth forming a bubble that is stable and that float to the surface as a froth flotation cell. Collector reagents react with the surface of the precious metal sulfide mineral particles making the surface is water repellent. surface of the mineral-bound water molecule is released and will be changed to hydrophobic.

Thus the collector end of the hydrophobic molecules will be bound hydrophobic molecules from the bubble, so the mineral ore can be adrift. Collector has a molecular structure similar to the detergent hydrophobic sulfide mineral grains are attached to the air bubbles that rise from the slurry zone into the froth that floats on the surface of cells.

In the flotation process of air bubbles formed initially has small size and some are attached to the surface of mineral particles. Furthermore, another air bubble formed next to join the existing air bubbles and form air bubbles with a larger size, so as to have sufficient lift to lift mineral particles to the surface. The mechanism of attachment of mineral particles in the air bubbles inside the tank during the flotation process flotation occurs when the hydrodynamic forces and the forces of interaction between mineral particles with air bubbles, resulting in collisions with air bubbles and mineral particles occurs attachment of mineral particles with air bubbles.

From the results ofbubblefrothflotation processthat resembles acolored foam detergent concentrate metallic orescarryinggold-coppermineral-ladenis thenuptothe tubshelter, and foam concentrate that has been lifted from the drain into the upper lip and into the trough flotation machine is in use as a valuable mineral collection.

In order fortheflotationprocesscan take placebyeithermeansof attachmentof particlestoairbubbleslasteduntilthetop edge of theflotationcell,it is necessary toconsiderthe followingmatters :

sulphide flotation - an overview | sciencedirect topics

sulphide flotation - an overview | sciencedirect topics

The Platsol process was originally developed in collaboration with the University of British Columbia, Kane Consultants Ltd., and Lakefield Research in Canada for the treatment of flotation sulfide concentrate for Polymet Mining Company in Minnesota and was tested on similar types of concentrate materials. This process involved dissolution in one step of the base metals (copper and nickel) as well as the gold and PGMs. This was followed by solidliquid separation, gold and PGM recovery, and conventional Cu SX/EW and recycling of the copper raffinate to the autoclave.

The fundamental difference between the Platsol process and the conventional high-temperature pressure-oxidation processes is that a small concentration of chloride ions is added to the autoclave with 25g/L sulfuric acid. The chloride favors the oxidation of gold and PGMs and stabilizes them as dissolved chloro-complexes. Grinding the ore with ceramic rather than iron balls was required to prevent cementation of gold chloride (Ferron etal., 2000).

The concentrate tested was a flotation concentrate from the Northmet project, USA, assaying 14.7% Cu, 3.05% Ni, 0.14% Co, 26.7% S, 1.4g/t Au, 2.2g/t Pt, and 9.9g/t Pd. Pressure-oxidation conditions were 225C, pulp density was 11%, retention time was 120min, and oxygen overpressure was 689kPa. The ore treated had a P80 of 15m. After solidliquor separation, the gold and PGMs were recovered by sulfide precipitation using NaHS or by activated carbon. The copper was recovered using conventional solvent extraction and electrowinning techniques. Overall recoveries were Cu 99.6%, Ni 98.9%, Co 96%, Pd 94.6%, Pt 96%, and Au 89.4% (Ferron etal., 2000). Studies in treating a variety of refractory gold concentrates under optimum conditions (225C, NaCl 1020g/L, 26h O2 at 700kPa) achieved gold extractions of the order of 9096% compared with direct cyanidation where gold extraction was less than 20% (Ferron etal., 2003).

Examination of a variety of recovery options showed that loading of gold onto carbon from clear liquors and pulps was rapid and did not require prior neutralization. Zadra elution of the loaded carbon recovered more than 90% of the gold; however, further work in investigating carbon regeneration was required. Gold could easily be precipitated from acidic Platsol leach liquors with NaHS, but minimization of the co-precipitation of impurities, such as copper, needed to be addressed, given the co-leaching of base and precious metals. In addition, some tests on gold recovery by ion-exchange resins showed promise. Gold chloride can be precipitated using a synthetic covellite produced in residual copper recovery process. The Platsol process has undergone a detailed engineering phase following DFS and FEED studies for the Northmet project (Wardell-Johnson etal., 2009). More details may be found in Chapter 46.

Pyrite and arsenopyrite are the principal hosts of submicroscopic gold. In addition, gold minerals are often preferentially associated with these minerals; hence, their floatability is relevant to gold metallurgy. The following possibilities exist in the case of sulfidic gold ores:

Although the exact location of arsenic in the pyrite crystal structure is still being debated the general consensus is that arsenic is replacing one of the two sulfur atoms in the sulfur dipole, thus forming AsS2. Whatever the mechanism of arsenic incorporation in the pyrite structure, it is certain that with increasing arsenic content, pyrite becomes more readily oxidizable, which in turn affects its floatability. Pure pyrite floats without activation because of dixanthogen formation as a result of catalytic oxidation of xanthate on clean (fresh) pyrite surfaces. However, most floated pyrite is either copper or lead activated which means that incipient surface oxidation had to take place to form islands of pyrrhotite, which then became sites for activation by Cu or Pb. With increasing arsenic content, the surface oxidation of pyrite is greatly accelerated to the point where depression by surface oxidation overwhelms activation. This pyrite requires heavier collector loadings in order to float. If the oxidation is too fast, in the absence of activators irreversible depression takes place. Understanding the mechanism of arsenian pyrite flotation is particularly important given that it is the principal gold carrier in the very important submicron gold pyritic sedimentary-hosted gold deposits, also known as Carlin-type deposits (Thomas, 1997). To overcome the inadvertent oxidation of pyrite in the N2TEC process implemented at Twin Creek (NV, USA), grinding and flotation are carried out under nitrogen atmosphere using lead nitrate as the activating agent (Simmons, 1997). Barrick testwork on the Carlin ores demonstrated that the use of acidic flotation improved selectivity and kinetics. The success in maximizing recovery of gold-bearing arsenian pyrite fromCarlin-type ores lies in minimizing unwanted pyrite oxidation coupled with generous activation, while producinglower-grade sulfide concentrates by recovering middlings with finely disseminated gold-rich pyrite (see Figure5.16).

In the case of mesothermal pyritic gold ores, pyrite and arsenopyrite are sufficiently coarse grained, which allows for good liberation at modest grind fineness (P80 of 75120m) and results in high-grade concentrates with good recoveries. In some of these ores, submicroscopic gold is exclusively carried by arsenopyrite, which makes separation enticing from the barren pyrite (Donlin Creek, AK, USA), while in others submicroscopic gold is equally shared by arsenopyrite and arsenic-rich pyrite (Olympias, Greece).

Free gold recovery from the slimes fraction can be enhanced by adding some of the best-match collector in the regrind mill feed, to build up a heavier collector loading on already free tiny gold particulates as well as coat newlyliberated gold grains when their surfaces are still fresh. In ores with gold grains displaying a bimodal or broadsilverconcentration distribution, a matching collector will enhance flotation kinetics of the slower floating member.

Processes for the direct hydrometallurgical treatment of concentrates for recovery of PGMs, gold, and base metals have been developed but have not yet been applied commercially. This approach would avoid the need for conventional smelting, converting, and base-metal refining prior to refining of PGMs and gold.

The Platsol process was originally developed in Canada for the treatment of flotation sulfide concentrate for Polymet Mining Company in Minnesota and subsequently tested on other PGM concentrates. This process involved codissolution of base metals (copper and nickel), gold, and PGMs via addition of chloride to the autoclave. Grinding theore with ceramic rather than iron balls was required to prevent cementation of gold chloride (Ferron etal., 2000). Overall recoveries for the Northmet concentrate were Cu, Ni, Co >96%, Pd 95%, Pt 96%, and Au 89% (Ferron etal., 2003). The process underwent a detailed engineering phase following DFS and FEED studies for the Northmet project (Wardell-Johnson etal., 2009).

The Kell process comprises recovery of base metals by sulfuric acid pressure leaching, and a heat treatment may be applied to convert precious metalbearing minerals into forms that are soluble in a subsequent chlorination leach, at 9599% extraction efficiencies for the precious and base metals. The process has been patented (Liddell, 2003) and developed (Liddell etal., 2011; Liddell and Adams, 2012a,b), including integrated continuous pilot-scale testing on various PGM/Cu-Ni and polymetallic concentrates. Pallinghurst Resources is assessing potential construction of a full-scale Kell plant to extract base metals (copper, nickel, and cobalt) and PGMs as well as gold, at its Sedibelo Platinum Mines subsidiary (Seccombe, 2014).

Flotation circuit configuration on most gold mines can be divided into a number of categories, viz. open circuits with no cleaning at all, and open and closed circuits with single stage and two stages of cleaning. Open circuits have the advantage of no feedback from the effects of nonsteady-state operation and therefore are inherently more stable than the closed-circuit configuration. Closed- and open-circuit flotation cleaning is used on gold mines where high-grade concentrates are required for roasting and smelting. Under these conditions, it is difficult to maintain very high gold and sulfide flotation recoveries, while also producing an acceptable grade of concentrate. Where there is no constraint on concentrate quality, high gold and sulfide flotation recoveries are achievable to the extent that a discardable gold flotation tail is possible (Bax and Bax, 1993; O'Connor and Dunne, 1991). Cleaning-circuit configuration, either single or two stages of cleaning, and cleaner residence time are related to the particle size of the sulfides in the flotation feed and the presence or absence of floatable gangue components.

Unit flotation cells (Hasting, 1937; Taggart, 1945) and the more recent flash flotation cells (Kalloinen and Tarainen,1984) are installed in grinding circuits with the purpose of improving the overall flotation recovery of free gold (Taggart, 1945; Suttill, 1990; Laplante and Dunne, 2002). The aim is to remove as much as possible of the free gold contained in the circulating load of the grinding mill before it is overground or is covered with coatings of iron, sulfide, or other coatings that will lower flotation recoveries. Improved overall gold flotation recoveries of 210% have been quoted (Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990). Further, the inclusion of unit and flash flotation cells will generally provide better flotation stability and performance. Improved overall gold flotation recoveries from 3% to 10% have been quoted (Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990) for ores of variable gold and sulfide content (Taggart, 1945). Contrary to the belief of many, flash flotation does not recover particles coarser than that achieved by conventional flotation (Newcombe etal., 2012a,b). The particle size distribution of a flash flotation concentrate is typically coarser compared to conventional concentrates, with the fine particles being removed by a cyclone ahead of the flash flotation cell. The reasons for low coarse particle recovery are numerous, the major impediment being the relatively low concentrations of collector added in the flash flotation circuit. Much higher concentrations of collector are required to float very coarse particles (Bravo etal., 2005; Dunne, 2012); however, adding high levels of collector in a flash flotation circuit will compromise mineral separation selectivity in the downstream convention flotation circuit(s). Other items that impact coarse particle flotation include poor aeration capacity and bubble dispersion and solid residence times due to short circulating (Newcombe etal., 2013a,b).

Many coppergold concentrators have a combination of flash flotation and gravity separation in the milling circuit to enhance overall free gold recovery. The reason for the inclusion of gravity separation at these concentrators is that the coarse free gold is not recovered effectively in the flash flotation cell. Concentrators that have this circuit configuration have found that most of the fine gold is recovered in the flash flotation circuit, leaving the gravity circuit to recover the coarser gold. One of the dilemmas arising from laboratory testing for the combined circuit is that the standard gravity test procedure will overpredict plant gravity recovery, because fine free gold is captured in the laboratory test. To better predict both gravity and flash flotation free gold recovery in a concentrator a combined gravityflash flotation, a model has been developed by McGrath etal. (2013). The inputs for the model are obtained from a defined laboratory procedure for both the gravity and flotation separation steps.

Column flotation cells are used in roughing and cleaning duties on a number of mines treating gold ores (Lane and Dunne, 1987). A column cell typically provides higher concentrate grades compared to a mechanically agitated cell; however, losses of coarse gold may be higher in the column cell (Chryssoulis etal., 2003b). Testing combining high-intensity conditioning together with a three-product column produced high enrichment ratios of gold at acceptable recoveries when processing fine gold from low-grade tailings (Valderrama and Rubio, 2008).

Early work was carried out by Lakefield Research on thiosulfate leaching carbonaceous double refractory ores from Barrick Gold Corporation after pressure oxidation (POX) pretreatment (Thomas etal., 1998; Fleming etal., 2003). Up to 95% Au extraction was achieved from the finely divided gold left in the oxidized residue. In this process, POX residue leaves the autoclave at 35% solids and is directed to a leaching operation, where it is contacted with ammonium thiosulfate (5g/L) and copper sulfate (25ppm Cu) at 4060C and pH value 8. The slurry of gold-bearing leachate and solid residue leaving the leaching circuit contains in the range of 15ppm gold and is directed to an RIP circuit, where gold and copper are loaded onto a strong-base resin to 15kg/t Au and 1025kg/t Cu. Copper is eluted from the resin using ammonia thiosulfate (200g/L) and gold is eluted using potassium thiocyanate (200g/L). The copper-bearing eluate is returned to the leaching circuit, while the gold eluate is either electrowon or precipitated.

Jeffrey etal. (2008b) also reported gold recoveries of >95% from a pressure-oxidized flotation sulfide concentrate. In this work, an integrated RIP process was tested, using a gold thiosulfate process incorporating a novel sulfite-enhanced chloride-based elution of gold from the resin, together with electrowinning of gold, and recycle of eluate and resin. The favored thiosulfate leaching conditions were 5mmol/L copper sulfate and 50mmol/L ammonium thiosulfate at a pH value of 8.5, which allowed less generation of polythionates and their subsequent loading on the resin (0.1kg/t). Gold and copper loadings onto the strong-base resin were 2.5bkg/tAu and 2kg/tCu. Copper was eluted from the resin using 0.5 mol/L ammonia thiosulfate and gold was eluted using sodium chloride and sulfite mixture. The copper-bearing eluate was returned to the leaching circuit and the gold eluate was electrowon.

The ammonium thiosulfate leaching of pressure-oxidized carbonaceous ore and various sulfide concentrates has been recently evaluated by Breuer etal. (2014). A comparison cyanide leaching and thiosulfate of gold for various pressure-oxidized materials is shown in Figure28.10. Thiosulfate conditions, which favored gold extractions similar to or better than cyanide leaching, varied for different POX residues. Adding 10g/L of sodium chloride to the autoclave, which converts the metallic gold in the ore to an ionic form during pressure oxidation, enabled improved gold extraction, but thiosulfate leach conditions which favored maximum gold extraction was sensitive to pH conditions. The preparation of feed for pressure oxidation and ultimate products for various materials possibly influenced results. POX conducted in the absence of chloride produces a significant quantity of basic ferric sulfate which required neutralization with lime, whereas the addition of salt produces residues containing predominantly natrojarosite [NaFe3(SO4)2(OH)6]. The benefit of salt addition on gold extraction varied under alkaline and acid pressure oxidation treatment of carbonaceous sulfide ore residues. The results demonstrated that optimum POX/leaching conditions will vary depending on the residue being thiosulfate leached.

Figure28.10. Comparison of gold extraction for cyanide and thiosulfate leached pressure-oxidixed (POX) leached residues (a)without and (b)with NaCl for different ore and concentrate types (NaCN, 50mmol/L and air; ATS, 50mmol/L (NH4)2S2O3, 0.5mmol/L Cu(II) for pH 8.59.0; or 2mmol/L Cu(II) for pH 9.510.5).

Other recent studies have shown favorable gold extractions of up to 89% from a high-grade pressure-oxidized concentrate (Au 32g/t, Ag 12g/t, Fe 59%, S 21%, and As 19%) in 0.2mol/L thiosulfate and ammonia solution with 5% thiosulfate consumption; however, no comparison was made with cyanide to determine maximum gold recovery (Lampinen and Turunen, 2015).

Table15 describes the major findings and recommendations from the research conducted related to LCA of these metals. Cobalt is a valuable metal found in the earths crust which is widely used in industrial applications. Cobalt mining has a notable impact on human health due to cancer-causing elements which may cause heart disease, vision problem, etc. Farjana etal. conducted the LCA of cobalt extraction process. According to their study, cobalt extraction is harmful to eutrophication and global warming. Cobalt extraction requires a large amount of electricity which is detrimental to global warming and also is the blasting (Farjana etal., 2019c). Cemented carbide has higher hardness and higher corrosion resistance, mostly used for drilling tools and cutting tools. China is the leading producer of cemented carbide. The cemented carbide ore is mined from extraction, crushing, milling, gravity method grinding, sulphide flotation and roasting. In the hydrometallurgy stage, the cemented carbide ore is digested, filtrated, precipitated, extracted using solvent and finally crystalised. In the pyrometallurgy stage, the ore goes through calcination, hydrogen reduction and carburisation. In the powder metallurgy stage, the ore goes through powder milling, granulation and sintering. Furberg etal. conducted a cradle-to-gate LCA of cemented carbide production with cobalt, while the geographic location was non-Chinese (Canada and United States). Their study stated that impacts were due to elements like kerosene, tailings, water and electricity. The highest impacts were on the category of TAP (terrestrial acidification), ODP (ozone depletion), FEU (freshwater eutrophication). And the lowest impact was on CC (climate change), PCOF (photochemical oxidant formation) and WD (water depletion) (Furberg etal., 2019). Manganese is an essential element for batteries, fertilisers and chemicals. Manganese is a widely used alloying element that comes in conjunction to make ferroalloys. The manganese alloy is produced using mineral extraction, hauling, ore preparation and beneficiation, sintering and transportation, smelting, crushing, screening and refining. Westfall etal. conducted an LCA study based on manganese alloys, where datasets were collected from 16 ore and alloy producers. The authors have conducted a cradle-to-gate LCA of silicomanganese, ferromanganese and refined ferromanganese. The impact categories considered were GWP, AP, POCP, water and waste. The analysis was done using CML 2001 method. According to their analysis, electricity demand, fuel consumption during smelting was the primary contributors for impact (Westfall etal., 2016). Magnesium oxide cement is widely produced in China, North Korea, Turkey, Russia and Australia. The magnesium oxide is produced from raw material acquisition, crushing, vertical shaft kiln, precipitation tank, screening, crushing, grinding and packaging. Ruan etal. analysed the LCA of magnesium oxide, where the functional unit was 1 tonne. They showed that MgO has a lower impact on the ecosystem and resources but a larger impact on human health. The analysis was done using EcoIndicator 99 method. They considered five different case scenarios based on fossil fuel and raw material consumption (Ruan and Unluer, 2016). Silver metal is most widely used for industrial purposes or for making jewellery. There are very few studies which addressed the environmental impact of silver mining processes. Farjana etal. analysed the environmental burdens associated with goldsilverleadzinccopper beneficiation process (Farjana etal., 2019b). In another study, they analysed the environmental impacts of goldsilver refining operations (Farjana etal., 2019d). They found that silver beneficiation and refining have the least environmental impacts than gold mining processes as they consume the least amount of electricity. However, there are some impacts on eutrophication, global warming and ecotoxicity (Farjana etal., 2019b,e). Titanium oxides are widely used for making high-performance metal parts, artificial body parts and engine elements. Ilmenite and rutile are the generally found form of titanium oxides. Ilmenite and rutile are extracted from mining site using heavy mineral concentration, rare-earth drum separation, electrostatic separation circuit and gravity separation circuit. Farjana etal. conducted a comparative LCA analysis of cradle-to-gate titanium oxides production. Ilmenite and rutile were considered where the geographic region considered was for Australia. The datasets were collected from the AusLCI database and SimaPro software. The study revealed that rutile had a significant environmental impact than for ilmenite due to higher energy consumption and electricity use. The GHG was 0.295kg CO2 eq/kg of ilmenite production and 1.535kg CO2 eq/kg of rutile production (Farjana etal., 2018c).

In sulfide flotation, recovery and selectivity are fundamentally dependent on the relative rate constants of various mineral phases (Boulton et al., 2003). Therefore, an evaluation of the hydrophobicity balance by mineral particles requires accurate selection of the mineral phase. The hydrophobichydrophilic (hydrophobicity) balance by mineral phases and the relative statistical average require determination of the main species contributing to each category in surface layers. This determination is not a simple procedure in a flotation pulp containing diverse mineral phases, various mineral sizes, adsorption of various reagents, different products oxidation, precipitations (often colloidal), and polysulfide Sn2-species(resulting from loss of metal ions, usually Fe2+) on mineral surfaces (Smart et al., 2003a,b, 2007).

Numerous studies have been conducted to evaluate the hydrophobichydrophilic (hydrophobicity) balance by mineral phases (Vickers et al., 1999; Piantadosi et al., 2000, 2002; Duan et al. 2003). For adsorption studies in mineral flotation, quantification of surface species by TOF-SIMS and simply using the peak intensities of adsorbed and substrate signals are unsuitable (It does not take into account many of the matrix effects of mineral phases) (Piantadosi et al., 2000). To generalize, in the case of adsorption, the ion ratio of interest can be expressed as:

where RPI is the relative peak intensity, Iads is the integrated peak area of the ion fragment characteristic of the adsorbed molecule, and Isub is the integrated peak area of the ion fragment characteristic of the substrate. In principle, RPI is the relative peak intensity measured by TOF-SIMS, or RPI is the ideal parameter for adsorption studies since it has the character of , the traditional measurement of uptake (Iads) function of monolayer capacity (Iads+Isub), and might be expected to vary regularly with the extent of coverage of the substrate adsorbent by the adsorbate (Vickers et al., 1999).

This method of quantification yields a clearer illustration of the differences between concentrates and tails (Piantadosi et al., 2002). It is required to use Eq. (2) for each index (Vickers et al., 1999). Piantadosi et al. (2000) investigated the coverage of potassium isobutyl xanthate (IBX) and sodium diisobutyldithiophosphinate (DBPhos) adsorbed on the surface of galena by TOF-SIMS. They developed models which fully described both hydrophilic and hydrophobic indices of recovery of particles by flotation. An example of an initial development is described below:

Development of a more extensive hydrophobic/hydrophilic index may involve the ratios of a number of these indices, as shown above. For instance, the DBPhos/SO3- indices may be chosen as a first attempt at a hydrophobic/hydrophilic ratio. An alternative hydrophobic/hydrophilic ratio has been chosen to form a more direct overall index (I), using the Iads/Isub ratios.

Piantadosi et al. (2002) demonstrated that statistically, particles in the concentrate are more hydrophobic and separable than particles in the tail when both hydrophobic (collectors) and hydrophilic (oxidation products) species are combined (Piantadosi et al., 2002). Piantadosi et al. (2002) continued their surface analysis by TOF-SIMS with the aim of investigation on the particle-by-particle statistics of hydrophilic and hydrophobic species on the surfaces of mixed samples (galena and pyrite) under flotation-related conditions. Using a similar procedure, they found that in the concentrate the surface of galena have less Ca/Pb, PbOH/Pb and oxy-sulphur species (SO3/S) compared top articles in the tail. In other words, they were less hydrophilic. These differences are statistically considerable. Statistical results obtained for other species, such as Mg/Pb species, did not show any significant difference. This technique identified the effective species that correlate with flotation. Using a similar method, Duan et al. (2003) predicted an advancing contact angle of 71+2 (degrees) for the chalcopyrite particles in the Mount Isa Mines ore using the DTP/SO3 ratio as measured by TOF-SIMS.

A great number of base metal sulphide flotation plants use saline water. Some of them are summarised in Table 1. The three nickel flotation plants (Mt Keith, Leinster mine and Kambalda Nickel Concentrator) in Western Australia operated by BHP Billiton use bore water with high ionic strength. Table 2 shows the elemental compositions of the bore water used in the Mt Keith Operation, the largest nickel flotation plant in Western Australia (Peng and Seaman, 2011). The salinity of this water is several times higher than that of sea water. In the Raglan mine operated by Xstrata Nickel (formerly Falconbridge) in Northern Quebec, Canada, salt levels range from 20,000 to 35,000ppm throughout the year. An apparent consequence of the high salt content in the Raglan mine is that the flotation circuit is able to operate without the addition of frother (Quinn et al., 2007).

In Chile, many copper flotation plants use seawater. One example is Las Luces, a copper-molybdenum plant in Taltal, owned by the Las Cenizas Mining Group (Grupo Minero Las Cenizas) of Chile. In Las Luces, seawater is brought from a distance of 7km, mixed with tailing dam water and then used in grinding and flotation circuits. During the last 15years the increase in the total dissolved solid content of the process water in Las Luces was from approximately 36.0g/L (seawater) to 46.4g/L (Moreno et al., 2011). Another large copper plant using seawater in Chile is the Esperanza Concentrator at Sierra Gorda (Antofagasta Minerals S.A AMSA). This plant is processing ore at 95,000tpd using sea water without any pre-treatment. Sea water is pumped 145km from the Pacific Ocean to a 60,000m3 pool at the mine site located at an altitude of 2300m above sea level (Castro, 2012).

Other important cases are the Batu Hijau Concentrator (Newmont, operating from 2000) located at the Indonesian island of Sumbawa, and ayeli Bakr letmeleri A.. (CBI) in Turkey. The Batu Hijau Concentrator uses sea water for processing a gold-rich phorphyry copper ore (chalcopyrite-bornite) (Castro, 2012), while CBI processes a complex CuZn sulphide ore using dissolved metal ions and sulphide ions, mainly in the form of SO42 and S2O32 (Bak et al., 2012).

In South Africa, the recovery of platinum group elements (PGE) through the selective flotation of base metal sulphides also uses saline water. Flotation operations can be found in different parts of the Bushveld Complex, e.g., Merensky reef and Platreef. The ionic concentration varies in all Merensky concentrators, South Africa (Corin et al., 2011; Miettunen et al., 2012; Shackleton et al., 2007; Wiese et al., 2005a; 2005b; Wiese et al., 2007).

Although hydrothermal technology has great advantages in solid waste treatment, the industrial application of hydrothermal processes suffers from various challenges because of the severe process conditions. For example, corrosion requires the use of expensive alloys, and the high operation pressures put tough requirements on process components such as feed pumps [105,136].

However, Pilot test of neutralization slag by hydrothermal sulfide flotation indicated that hydrothermal technology is expected to realize industrial application in the future [137]. Therefore, some challenges and questions that should be solved. The environment of high temperature and high pressure is relatively dangerous. In the future, more studies should be conducted on how to moderate the hydrothermal conditions. It is necessary to find a proper catalyst to lower the reaction temperature, especially for the high-temperature HT process [62,138]. How to produce value-added materials is an urgent problem to be solved, especially how to prepare nanomaterials from solid waste [83]. There are relatively few studies on the corrosion of reactors by alkaline additives. Therefore, the choice of additives will also be a hot research topic in hydrothermal reactions.

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