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By C. Van Den Heuvel, C. B. Van Buuren

BIOX® has been successfully demonstrated over the past 2 decades for the recovery of gold from refractory sulphide concentrates. To date 12 successful BIOX® plants have been installed in 8 countries spanning 4 continents with the 13th installation currently nearing completion in the Philippines. In 2010 BIOMIN introduced the ASTER™ process for improved water balances in commercial BIOX® applications by supporting tightening environmental legislation on water usage and availability in both arid and tropical regions. Previously, a common challenge for BIOX® plants was the inability to recycle thiocyanate containing water due to the low tolerance of the organisms. ASTER™, a biological thiocyanate and cyanide destruction process, renders a non-toxic solution for re-use in BIOX®. Two commercial ASTER™ plants are in operation with the 3rd ASTER™ plant under construction in the Philippines. A 3rd BIOMIN technology, specifically developed for double refractory ores is available for implementation. The HiTeCC process has shown considerable success on ores where competition from disseminated carbon has rendered ineffective gold recoveries. Through HiTeCC, gold is desorbed from the carbonaceous matter by manipulating the ionic strength and temperature of conventionally treated CIL product. One Australian commercial plant utilises HiTeCC while a 2nd installation is set for commissioning 4th Quarter 2014. BIOMIN’s premier solutions for refractory gold processing are described herein.

Jan 1, 2014

By Abid O. Adekunte

The design, construction and monitoring of the deep excavation support for the basement of a large-scale commercial development in Cavan, Ireland is presented. Excavation area was adjacent to existing structures and retained height H varied between 4.0m ? 11.0m. In areas where excavation depth H < 5.0m, wall was designed as cantilevered section. However in sections where 5.0m = H = 11m, additional wall supports were required for serviceability reasons. Traditionally, designers often provide additional wall support through the use of tie-back ground anchors, raking props and braces. On the current project, the main contractor was not in support of the use of temporary props so as to maximise working space, while owners of adjacent properties disapproved the use of tie-backs. The limited set-back between the wall line and site boundary also contributed to the complexity of the project. A rare alternative solution has been adopted. This novel approach involved buttressing the wall with steeply raking bored tension piles at regular intervals along the wall. Comparison of the monitored deflections in the cantilevered sections with those recorded at sections where tension piles were used shows the novel solution to be safe and quite effective. This approach is also considered to be more economical when compared with ground anchorage and propping.

Jan 1, 2008

By Elwyn King, Donald J. Dodds

"...Unfortunately, the research activities in soil mechanics had one undesirable psychological effect. They diverted the attention of many investigators and teachers from the manifold limitations imposed by nature on the application of mathematics to problems in earthwork engineering. As a consequence, more and more emphasis has been placed on refinements in sampling and testing and on those very few problems that can be solved with accuracy. Yet accurate solutions can be obtained only if the soil strata are practically homogeneous and continuous...." This statement about soil mechanics was made by Karl Terzaghi, thirty years ago and applies to a greater degree today to the infant science of rock mechanics. More uncontrolled variables are added to the already many existing in soil mechanics. Variables, such as Macro-discontinuity and in situ stress levels. In no other field of engineering are the problems as site oriented. Presently this site dependance demands the investigator or rock mechanics engineer to get his hands dirty rooting in the muck of a tunnel floor trying to properly install or readout test apparatus in a comprehensive site investigation. This paper discusses such a site investigation designed to determine the problems and solutions inherent in the driving of the Trans Koolau Tunnel. The paper will describe the various types of tests to be performed, how the results are to be used and the interrelation of the various tests themselves.

Jan 1, 1997

By D. Walton, W. C. Winegard, W. A. Tiller, J. W. Rutter

It is shown that the cellular substructure observed in binary alloy crystals grown from a melt of known concentration can be eliminated by the proper choice of growth conditions. For a given solute concentration, there exists a critical ratio of temperature gradient G in the melt to the rate of solidification R of the crystal, i.e., of G/R, which must be exceeded before the cellular structure can be eliminated. A secondary substructure is observed which is related to the cellular substructure. IN crystals grown from the melt, a prismatic substructure has been observed by Buerger,1 Pond and Kessler,2 and Rutter and Chalmers." This structure is columnar in the direction of growth and appears as a network of hexagonal cells in the plane of the solid-liquid interface. Rutter and Chalmers3 have shown that this substructure occurs as a result of the presence of a layer of high impurity concentration in the melt adjacent to the solid-liquid interface. The formation of the substructure is accompanied by a redistribution of this impurity in the neighborhood of the interface. The structure can be controlled by proper adjustment of the variables in the solidification process such as a rate of solidification and the temperature gradient in the liquid. The redistribution of solute during the solidification of metal crystals has been treated theoretically

Jan 1, 1956

By J. B. Hiskey

The hydrometallurgical application of thiourea in processing precious metal ores and concentrates has reached the development stage. This paper reviews recent technical literature covering the process development of thiourea leaching. In addition, recent fundamental contributions are discussed. Most of the interest in thiourea leaching has involved treating high grade concentrates that have sufficient value to just the use of thiourea as a reagent. An important development in thiourea technology is the addition of SO2 to the leaching solution. This procedure effectively reduces the decomposition of thiourea and eliminates the problems of passive surface layers that limit the leaching efficiency. Results for the thiourea leaching of a bulk sulfide concentrate and a residue from the autoclave leaching of a smelter flue dust are also reported. As determined by a standard bottle roll cyanidation test, the bulk concentrate was extremely resistive to cyanide leaching yielding only 16% gold and 2% silver extraction in 24 hours. Leaching with a relatively concentrated thiourea solution (1MCS (NH2)2) resulted in 83% gold and 75% silver extraction in four hours. With the treated flue dust residue, gold extraction was about 60%. Gold and silver extraction with cyanide was sensitive to both cyanide concentration and oxidizing conditions. Cyanidation resulted in 59% gold extraction and 77% silver extraction in 24 hours.

Jan 1, 1985

By J de Cuyper, S. R. Rao, J. Leja, K. S. Moon, W. Freyberger

A prolonged grinding of sulphide ores in steel mills appears harmful to flotation recovery and selectivity whenever xanthates are used as collectors. Particles finer than about 10um show a pronounced decrease in floatability. A supposition was made that the cause of the reduction in floatability of ground sulphides is the presence of metallic iron abraded during grinding and deposited on the ground particles. Electro- chemical measurements were made of mixed potentials developed by different sulphide-mineral electrodes (galena, chalcopyrite, pyrite, marmatite, pyrrhotite) in xanthate solutions, without a direct con- tact or in contact with a metallic iron electrode. These measurements showed that without a galvanic contact between the sulphide and the iron, the values of mixed potentials were in the vicinity of the xanthate/dixanthogen redax potential. When contact with iron was established, the mixed potentials shifted towards the more negative reducing or cathodic values, The shift was greater the higher was the ratio of the surface areas: iron electrode to sulphide electrode. Further, the shift in mixed potentials was greater the more reducing were the conditions on the sulphide electrode. Chemical analysis of reaction products produced on the surfaces of sulphides and in the bulk solution showed that less thiosulphate was formed in all of the sulphide-xanthate systems and that more elemental sulphur was detect-

Jan 1, 1976

By J. C. Cook

In solution-mining of underground salt and similar minerals, using drilled wells for access, it is desirable to monitor the lateral growth pattern of the resulting fluid-filled cavern. Therefore, a process of seismic surveying from the surface of the ground has been conceived in which the amplitudes of waves reflected from and transmitted through the soluble formation are measured. The large acoustic impedance contrast between solid and fluid should produce striking amplitude anomalies, especially if shear waves are employed, since thick fluid bodies are opaque to shear waves. Horizontally-polarized (SH) shear waves are best for preventing conversion to P waves at the numerous horizontal interfaces in the ground. Field tests to date have shown that a truck-mounted, half-ton hammer striking horizontally against the end of a trench produces usable SH-wave energy at lateral distances up to about 850 ft. Horizontally-directed explosive wave sources were effective to about 2000 ft. Conventional magnetic-tape recording and processing were used, but with the detecting geophones oriented to favor SH waves. An irregular solution cavity in bedded salt at 500-ft depth has apparently been located by SH-wave and SV-wave reflections. Further field work is planned to corroborate and extend this result. The Brine Cavity Research Group, an association of 11 chemical and salt producing companies, is supporting this work. Major deposits of salt in tabular beds lie beneath some 300,000 sq miles of land in the central and northeastern U.S. This salt is a basic source of soda ash and chlorine, and has been extracted as brine from drilled wells for about a century. During the past two decades, the U.S. solution-mining industry, following the lead of European operators, has greatly improved the extraction process through the application of engineering and science.&apos; In 1957, the Brine Cavity Research Group, an association of 11 chemical and salt producing companies, was formed. This group proceeded to attack certain common problems through the support of research. An outstanding problem has been that of determining the shape and location of the growing solution cavities in the underground salt, so that measures can be taken to maintain operating efficiency. The problem has been partially solved by the Dowel1 sonar mapping service, which employs a pulse-echo device lowered into the cavity through the well.2 However, the working range of this equipment is at present insufficient for large cavities, and echoes are not returned from highly sloping walls nor from behind such obstructions as rock debris. Therefore, an independent means of mapping the cavity, for example, from the surface without interfering with operation of the well, would be desirable. THEORY OF THE METHOD Seismic waves are the only physical agent known to be capable of sufficient resolution and penetration to define typical solution cavities from the surface of the ground. The geometry is unfavorable: cavity widths are generally less than half their depths below the surface; resolution and lateral location of boundaries and channels to within 50 ft at depths of 500 to 3000 ft is desirable. Conventional seismic surveying, as used for petroleum prospecting, is probably not the answer: isopach mapping, for example, is not thought accurate enough to define the cavity by the slight additional delay time it would introduce (of the order of 0.005 sec for a 50 ft-thick cavity in hard Paleozoic rocks). Refraction surveying has also been considered, but seismic specialists see little promise in it for this problem. In 1957, in correspondence with industry personnel, the writer suggested a seismic method based upon careful measurement of reflection amplitudes. As Table I illustrates, seismic reflection coefficients r for typical brine-rock interfaces are considerably higher than those for typical interfaces between different kinds of solid rock. This fact can be utilized in two ways, illustrated in Fig. 1: 1) If the cavity roof is reasonably flat (which it may sometimes be since the unsaturated top brine will be in contact with an insoluble rock stratum), extra-strong seismic reflections will be received from the salt stratum where the solid has been replaced by liquid.

Jan 1, 1964

The Manto Verde copper mine (Atacama Region, Chile) is due to go into production in 1996. It will produce 40 000 tonne/year of copper cathode by heap leaching/SX-EW. Mineralogical studies were undertaken in conjunction with the metallurgical testwork program to characterise the ore and its response to leaching by dilute sulphuric acid. The leachable copper mineralogy is dominated by the oxide copper minerals: brochantite and malachite with minor chrysocolla and atacamite. The gangue minerals include abundant orthoclase, specularite, quartz and chlorite with minor muscovite and trace smectite clay, calcite and limonite. The study showed that complete or partial decomposition of calcite, chlorite, smectite and wad consumed acid and generated solution impurities (Ca2+, Fe2+, Mg2+, A13+, K+, Si4+ and Mn4+2+ ). Although these were manageable at Manto Verde, their uncontrolled build-up could have affected the operation in several ways: the degradation of heap penneability and fonnation of crud in the SX plant from the precipitation of gypsum, goethite, colloidal silica and jarosite; and the entrainment of contaminated pregnant leach solution in the SX organic resulting in the transfer of impurities to the electrolyte and subsequent effects on cathode quality and electrowinning efficiency. In conclusion, process mineralogy played an important role in understanding and managing potential mineral-related problems in the heap leaching and SX-EW recovery of Manto Verde ore.

Jan 1, 1994

By R. Tipman, T. C. Burnett

"A general computer program has been written to calculate material balances using multiple regression techniques. The program can handle metal assays or size distribution data obtained in multiple surveys of the flotation circuit. The ""best fit"" weight split is calculated between the input and product streams by minimizing the sum of squares of the residual errors using all of the available data. The program includes an assay adjustment feature in which each assay is recalculated to fit the ""best fit"" mass balance. Thus assay or sampling errors can be easily identified. Also the assays can be statistically weighted so the most precise measurements are given the most influence on the material balance. The computer output reports the mass balance, the actual and corrected assays, the percent metal recoveries and the significance and confidence levels for the weight splits.The program enables a rapid and accurate evaluation of mass balances from large masses of data obtained in multiple surveys around a flotation circuit. The effects of changing process variables can be quickly evaluated with statistical confidence."

Jan 1, 1978

By S. J. Pearce

There are a number of methods for determining relatively small quantities of sulfur dioxide in air. However, those methods require fairly elaborate apparatus and laboratory facilities. Frequently in industrial hygiene surveys it is desirable to have a simple arrangement to obtain "on the spot" determinations. This has been accomplished for concentrations of sulfur dioxide of hygienic significance by using the midget impinger and decolorizing a starch-iodine solution.

Jan 1, 1948

By L. G. Austin

If S (x) is the specific rate of breakage of size x and B (x, y) (see Table 1 for Nomenclature) is the cumulative breakage distribution function, the Herbst Fuerstenau2 assumption is that Inserting this assumption into first-order grinding equations it has been shown" that there are then simple analytical solutions to the grinding equations, as shown in Table 2. In this context "plug flow" is defined as all material passing through the mill in the same length of time, the residence time t. The term "n fully-mixed" means that the mill has a residence time distribution" which can be treated as if the mill were n equal fully mixed sections in series. In addition it has been shown that the assumption of Eq. 1 also leads to simple analytical solutions for closed-circuit, steady-state grinding with an ideal classifier at the mill exit returning oversize to the mill feed. The solutions are also given in Table 2, with µ as the ideal classifier cut size, C the circulation ratio, and Q (x) the size distribution of the circuit product. The specific surface area of these size distributions is where the density-shape factor k. is assumed to be constant with size. In practice, most methods of size analysis or area [ ] measurement have a finite lower limit, so that the integration is really performed from xm, to xm, For a realistic degree of grinding the value of So is dominated by the fine sizes because of the 1/x factor. Consider batch grinding of a feed of size y to y + dy. The batch grinding solution given in Table 2 is then [ ] The fresh surface area produced is the area of the fines plus the area of the original material remaining minus the area of the original material, [ ] Little error is introduced by neglecting the last term on the right-hand side since at t = 0 it is zero, and when t is significant the first term dominates because of the fine sizes produced. In addition, the form of P (x, t) for [ ] realistic values of S (x) 2t ax° is such that the major fraction of fresh surface produced is contained within the bottom 10% of weight of broken material, that is, for sizes for which P(x, t) < 0.1 (providing y is not near x.,). Thus for surface area purposes Eq. 3 can be approximated by [ ] becomes providing again that the feed sizes are coarse relative to xmi. Then the fresh area production is and W is the mass in the mill. The rate of surface production is WK, and the specific surface is Kt. From Table 2, the specific energy is mt/W so surface production per unit energy is [ ] Surface produced per unit energy input = KW/m (8) where m is the energy input of the mill. Clearly, identical reasoning applies to open-cycle, plug-flow grinding, with t replacing t, and Eq. 8 is obtained again. Considering open-cycle, fully mixed grinding, similar reasoning gives P(x) - t S(x) for surface area purposes, and

Jan 1, 1974

By Bryan C. Smith, Duane R. Skidmore

Volumetric mass transfer coefficients and gas holdups were measured in a concentric tube airlift reactor designed for the microbial desulfurization of coal. The slurries were comprised of an acidified basal salts solution containing thirteen different weight percents of coal (74 µm Ohio #1) at three temperatures. An enhancement in the mass transfer coefficient, KLa, with respect to that measured in pure solution was found at weight percents of 5 (volume fraction solids, ES = 0.035) or less, the greatest enhancement being as great as 37% at approximately 2 weight percent (ES = J 0.014). At higher solids fractions, the mass transfer coefficient decreased with further solids additions, decreasing by as much as 91% at 40 weight percent coal (ES = 0.315). KLa increased with temperature at all solids concentrations, the greatest increase being 392% at 72°C vs that measured at 30°C. Gas holdup, EG decreased with all increases in solids loading by as much as 65% and increased by as much as 67% with temperature. KLa and EG were correlated with operating conditions separately, and the separate correlations combined to yield overall correlations for the mass transfer coefficient and for the gas holdup. The correlations are suitable for inclusion in system models devised earlier for microbial coal desulfurization processes, and may be used for design, operation and cost estimation for such systems.

Jan 1, 1989

By Mohammad Noaparast, Babak Fotoohi

"Due to certain advantages in (semi-)Autogenous-Ball mill circuits in most of the high-throughput ore grinding units of mineral processing plants, much attention is paid to optimize the mill performance. Despite the maturity of the grinding mill mechanical design, there had been made less attention to the efficiency of the grinding media in the past. In this research, the possibility of establishing a relationship between liner/lifter wear rate and variations in the composition of the iron ore (from Chadormalu mine) is investigated.Measurements and determination of liner/lifter wear rates and patterns concluded in recognition of critical liner segments exposed to much abrasion during the ore grinding. After metallurgical analysis of liner materials and correlation of wear rate with the ore composition, the homogeneity in the metallurgical microstructure of the alloyed steel (such as the uniform distribution and amounts of the constituent phases) was determined as the most influential factor in the response of the liner/lifter to the variations in the ore composition. Magnetite with greater abrasion index than hematite, in finer particles, is responsible for additional abrasion of liners; hematite with greater work index exists in the mill with larger particle sizes, thus it could produce wear on liner parts with less tolerance to impact as is the situation for martensitic steels.INTRODUCTIONThe mineralization in Chadormalu plutonic iron ore deposit is consisted of mostly magnetite, hematite, martite, apatite, quartz and carbonates. The average composition of the ore is presented in Table 1. In the processing circuit, the gyratory crushed ore reduced to 30 cm would finally be the feed to three similar 32’ semi-autogenous mills in parallel processing lines. The pulp in the mill would have the characteristics mentioned in Table 2. The quality of the mill output has its direct effect on the downstream processes either physically or chemically (e.g. in cycloning or flotation respectively). In order to take advantage of optimum performance for these circuits, a close control of the ore grinding conditions is necessary; because any shortages in throughput, liner/lifter damage or changes in ore composition could bring the unexpected casualties. In this work, it has been attempted to find (possibly) a reasonable relationship between variations in ore characteristics together with liner quality and the amount of wear rate in different liner/lifter segments of a semi-autigenous mill. Liner/lifter segments are supplied by three different domestic steel making companies. The alloy structure of the liners/lifters is made up of pearlitic chromoly steel."

Jan 1, 2005

By C. Villaseca, F. Penna, R. Villarroel

Acid mist is the undesired result of copper electro-winning (EW) process, generating a contaminated and hazardous environment in the working areas, as well as promoting high corrosion conditions for structures and equipment in the EW tank-house due to the buoyancy of sulfuric acid in aerosol form. Likewise, organic entrainments that occasionally occur in this process contaminate the cathodes, as well as increasing the risk of fire. LateralVex® and LateralFlow® are technologies that effectively and efficiently solve these problems, without affecting the usual activities of the operators, thus presenting a significant distinction with respect to other technologies that seek to solve these problems. Minera Antucoya, an AMSA group company, which produces 80,000 tons of electrolytic Cu per year, decided to install both technologies for its EW tank-house, LateralVex® in commercial cells and LateralFlow® in scavenger cells to keep both the acid mist and the organic entrainments under control. This paper presents the effects of applying these technologies in Minera Antucoya’s electro-winning plant, the challenges addressed, the difficulties encountered, the problems and opportunities that their implementation involved, as well as the surprising results obtained at the time of measuring the acid mist— less than one hundred’s part of the Allowed Weighted Limit—in samples taken with the plant operating at maximum current capacity. Additionally, the paper includes some of the improvements identified during the period, more than 3 years, in which these technologies have been in operation in Minera Antucoya. INTRODUCTION Electrowinning is the most used electro-chemical process in the hydrometallurgical industry for the recovery of highly pure metallic copper as its final product, which is formed in previous stages in an aqueous medium that goes through a selective process using organic solvents to separate impurities and concentrate the metal in a sulfuric acid solution. In the last stage, this solution goes through a process in which high- density current at a low voltage is applied, so that the metal in the solution will deposit in solid state on a plate, usually a stainless-steel plate. This last process involves the generation of other sub-products due to the electro-chemical dissociation of the electrolyte’s components, or to the addition of chemical additives, or due to specific physical conditions or lack of controls during the operation.

Jan 1, 2019

By Sam Ash

The Murray Mine was one of the first mechanized cut-and-fill underground gold operations in northern Nevada. At the end of 2002, mine production began to decline. So the mine entered a phase of remnant mining and began preparations to close down. The remaining mining took place along the ore waste boundary in previously condemned areas and in challenging ground conditions. To remain economically viable, new techniques were implemented in operations, geology, engineering and rock mechanics. These improvements pushed the edge of known and accepted mining practices and have allowed the Murray Mine to remain in operation.

Jan 1, 2006

By Manqiu Xu, Peter Wells, Gord Bacon

"A methodology is proposed for milling flowsheet design using flotation rate, selectivity and residence time distribution data. The flotation rate data and the difference in flotation rates between the main minerals to be separated are determined from laboratory batch tests. The residence time distribution (RTD) data is calculated using the tanks-in-series model assuming each flotation cell is a perfect reactor. The rate, selectivity and RTD data can be used to calculate the optimum flotation cell number in a bank for rougher/scavenger in a commercial plant.This methodology is demonstrated in a case study of a nickel sulfide ore. It is shown that miniplant rougher/scavenger data can be closely predicted from a batch test using the proposed methodology.INTRODUCTIONIt has been the trend in the past several decades that large flotation cells are increasingly used. This is largely driven by the economics as ore grades are getting lower and the rate of ore processing must be increased in order to keep mining operations economically viable. The development and evaluation of 250 to 500 m3 WEMCO SmartCell and Outokumpu TankCell flotation machines are well in progress. There are many advantages in using large cells: fewer cells, less floor space, lower power cost (although this is now being questioned for fine particle flotation), simpler circuit, simpler control, and lower auxiliary equipment cost (Arbiter, 1999). As flotation cells are getting larger, the importance of cell hydrodynamics, namely, air/slurry dispersion, power intensity, froth transfer, residence time distribution (RTD) and short-circuiting are being fully recognized (Jonaitis, 1999; Weber, et al., 1999; Gomez and Finch, 2002). It is not the intention of this paper to discuss flotation cell scale-up. The intention is to answer what is the optimum number and size of cells in a flotation bank for a given throughput. Figure 1 illustrates this problem facing design engineers. The same mean residence time required for a given throughput can be achieved in a single large cell, or many smaller cells in a row. A few cells of large size may be more cost-effective but they may not be as effective as more cells of smaller size in achieving the optimum metallurgy."

Jan 1, 2005

By A. E. Back, S. F. Ravitz

When manganese ores are leached with sulphur dioxide, a large part of any zinc in the ore usually is extracted with the manganese.1 In the dithionate process,2,3 in which the manganese ore is leached with dilute sulphur dioxide gas and the manganese is subsequently precipitated by adding lime to the pregnant solution, the zinc is precipitated with the manganese. These facts suggested the possibility of applying the dithionate process to the recovery of zinc from oxidized ores or calcined sulphide concentrates. Accordingly, an investigation was begun, and preliminary tests were made on several ores and concentrates. The results were encouraging, but the need to devote available funds and personnel to other projects made it necessary to drop the work before the various factors involved in the process could be studied in detail. A brief mention of the process,4 however, elicited a number of inquiries, so that it appeared desirable to report the results of the preliminary work. The dithionate process as applied to zinc may be summarized briefly as follows: The ore or calcined concentrate is leached with dilute SO2 gas to dissolve the zinc, largely as zinc sulphate, in accordance with the following equation in which [ZnO] represents the "oxide zinc" in the ore or calcine: [ZnO] + SO2(g) + 1/2 O2 = Zn++SO4— [1] If, in the leaching step, the ore or calcine is suspended in calcium dithionate solution, the sulphate ion, as it is formed, is precipitated immediately as calcium sulphate, which is subsequently filtered off with the insoluble ore residue: Zn++SO4— + Ca++S2O6— = CaSO4(s) + Zn++S2O6- [2] An alternative procedure is to suspend the ore in water, filter off the insoluble ore residue after leaching, add calcium dithionate to the resulting zinc sulphate solution, and then filter off the calcium sulphate precipitate. In either event the zinc is obtained in the form of a pregnant solution of zinc dithionate, from which it can be precipitated as relatively pure zinc hydroxide by adding milk of lime: Zn++S2O6— + Ca(OH), = Zn(OH)2(s) + Ca++S2O6- [3] The zinc hydroxide precipitate is calcined to form the final zinc oxide product: Zn(OH)2 = ZnO(s) + H2O(g) [4] Theoretically the process is completely cyclic with respect to dithionate ion, the calcium dithionate required for reaction 2 being replaced by that formed in reaction 3. In practice, of course, small amounts of dithionate would be lost in the various filter cakes, depending on the economic washing limits. In the leaching step, however, some of the SO2 is oxidized to dithionale instead of sulphate, accord- ing to the following overall reaction: ZnO + 2SO2(g) + 1/2 O2(g) = Zn++S2O6-, [5] so that the dithionate losses are made up, at least in part. Ores and Concentrates Tested The analyses of the ores and concentrates tested are summarized in Table 1, which also includes the analyses of the calcines produced from the concentrates. The concentrates were roasted in a laboratory muffle furnace at the temperatures indicated until evolution of SO2 could no longer be detected; shallow clay dishes ordinarily were used; but calcines 2a and 2c were prepared in iron dishes, with the result that these two calcines were badly contaminated with iron scale. The "complex" concentrate, sample 1, consisted essentially of sphalerite grains containing inclusions of pyrite and other sulphide minerals too minute to be liberated by grinding within economic limits. In the " marmatitic" concentrate, sample 2, the iron was present largely as marmatite (that is, as iron sulphide in solid solution in zinc sulphide) rather than as pyrite inclusions. The zinc was present largely as the carbonate in the two "oxidized" ores, samples 3 and 4, and as the silicate in the "willemite" ore, sample 5. Leaching Tests The leaching cell was a 3-liter glass beaker placed in an electrically heated water bath manually controlled to maintain a temperature of 35°C in the leach slurry. The slurry was agitated by means of a flat paddle-type stirrer

Jan 1, 1950

By Turgut Yalcin

"Recent years have seen a resurgence of interest in flotation cell design, and one of the areas where attention is being focused is flotation in centrifugal fields which aims at increasing the flotation rates by forced bubble-particle contact.A new cycloflotation device was developed at Laurentian University and tested on the nickel ore of Falconbridge Limited in Sudbury. In comparison with conventional flotation, cycloflotation yielded at least seven times faster flotation rates in terms of the mass of solids floated per unit time per unit flotation volume, while producing concentrate grades similar to those of conventional flotation. These results, which were obtained with a bench-scale ""cyclofloat cell,"" suggest that cycloflotation has the potential of setting a new direction in flotation technology.IntroductionFroth flotation is no doubt the dominant separation technique in the mineral processing industry. Since its inception early in this century, much research effort has been and is still being expended on this process. Until recent years, emphasis seems to have been placed on the chemical aspects of flotation, but there is now an increased awareness for the need to improve flotation machine performance and more and more attention is being directed towards this area. In a 1982 article on ""Flotation Machines"" (Young 1982) it is stated that "" ... by now the design of the machines used would have reached a high level of perfection, ... In fact, this is not the case ... "".The process of flotation in a flotation device may be considered to take place in several distinct stages as follows:"

Jan 1, 1994

By L. Avramovic, B. Han, R. Jonovic, Z. Stevanovic, K. Haga, S. Dimitrijevic, A. Shibayama, L. Godirilwe, Y. Takasaki

Pyrite (FeS2) and chalcopyrite (CuFeS2) which could not be recovered, are present in mine tailings discharged from Bor copper mine. About 0.3% of copper (Cu) is contained in these tailings, therefore it can be considered as a useful copper resource. At the same time, pyrite is known to promote acid mine drainage, therefore it is necessary to detoxify the tailings. An efficient method has been proposed to recover copper and reduce the environmental impacts of tailings by a combined hydrometallurgical process of high pressure leaching (HPL), solvent extraction (SX) and electrowinning (EW) process. HPL process yielded high copper leaching rate (>95%) when water (H2O) was used as a leaching media. On the other hand, pyrite contained in the tailings was converted to iron oxide by high pressure leaching. Stability evaluation of the residue obtained from high pressure leaching, confirmed almost no elution of metal ions. The percentage of copper extracted and stripped from the Cu-loaded organic phase reached 97.4%. Moreover, it was confirmed that a low iron concentration in the electrolyte of copper electrowinning, resulted in the improvement of current efficiency of copper electrodeposition. INTRODUCTION Mine tailings are a by-product obtained from the flotation process and there are billions of tons of already existing tailings all over the world. Moreover, several billion tons of additional copper tailings will inevitably be produced in the coming years (Onuaguluchi & Eren, 2012). With the declining average copper ore grade, alternative sources of copper are highly sought, mine tailings have a comparable copper grade to that of low-grade ores (0.2–0.3%) therefore, it can be said that copper mine tailings present a very important source of secondary raw material for copper production (Antonijevi´c et. al., 2008). Consequently, it is necessary to confirm the possibility of the valuable metals (such as copper) recovery from mine tailing in which copper content is about 0.2–0.3%. Our group investigates the flotation and high pressure leaching processes for recovering of copper from mine tailings. Previous work on related samples include copper and nickel flotation tailings from Botswana, and a low–grade copper ore sample from Chile. Impurities, which are associated metals in mine tailing such as iron (Fe) and aluminum (Al) co-dissolve into the solution during the HPL treatment process. Therefore, separation of copper from impurities is necessary for selective extraction of copper from mine tailings and upgrading the copper concentration in the leachate to enable further purifying of Cu by electrowinning. Thus, it is important to investigate the possibility of applying solvent extraction for concentration of copper from leachate solutions of mine tailings (Onuaguluchi & Eren, 2012; Antonijevi´c et. al., 2008). In this study, LIX-84I was used as an extractant of copper from the leachate of high pressure leaching of mine tailing. As a fundamental solvent extraction investigation, the extraction behavior of copper was studied using a simulated solution of the leachate from the high-pressure oxidative leaching of mine tailing. The effects of; solution pH, extraction time, extractant concentration and phase ratio on copper extraction and sulfuric acid (H2SO4) concentration on stripping were investigated.

Jan 1, 2019

By Loyal Clarke, H. H. Storch

"The discovery of large deposits of polyhalite (K2S04.MgS04.2CaSO4.2H2O) in Texas has led to an investigation of possible industrial proecsses for the extraction of potash from this material. This mineral was first discovored in small quantities in the European potash deposits in or about the year 1800. Chemists fire thought it to be a form of calcium sulphate, but a more accurate analysis by Strohmiyer indicated the composition as given by the above formula.The decomposition of polytalite by aster is very slow and results in the formation of a layer of gypsum at the surface of the polyhalite which retarts the solution of the potaseium and magnesium sulphates. To obtain a substantial yield of K2S04 in solution it is necessary to use large amounts of water at room temperature, for he concentration of K2S04 is limited by the low solubility of syngonite (K2S04.CaS04.H2O) whereas at 100º C. several days boiling are necessary to obtain more than a 6 per cent K2S04 solution even when finely powdered polyhalite is employed.In 1854 H. Rose discovered that calcination, followed by leaching with water, resulted in a larger extraction of the potassium and magnesium sulphates, as compared with leaching without precalcination. During the course of a series of investigations concerning formation of oceanic salt deposits, Van't Hoff and his pupils obtained sufficient data to construct the equilibrium diagram for the system K+Mg+SO4--Ca++.H2O for 25° C. and for 83° C. These data are very well summarized. in the International Critical Tables' and by Boeke and Eitel. From a consideration of the diagrams for 83º C. one would conclude that an extraction of polyhalite with moderate amounts of water at 33º C. would result in a solution whose maximum concentration of K2So4 would be about 3.6 grams per 100 grams of water and of M2SO; about 4.0 grams per 100 grams of water, corresponding to equilibriumwith the solid phases: Polyhalite potassiam-calcium pentasulphate and anhydrite.The fact that it has been found possible to obtain as high as 11.4 grams of K2S04 per 100 grams of water (see table 6) and an equivalent amount of MgSo4 (8.1 grams per 100 grams of water was obtained corresponding to a ratio K2/Mg of 0.98; indicates that the formation of these complex salts is very slow. To be sure, this results was obtained at 100° C. instead of 83º C. However, the work of Anderson and Nestall 9 on the double salts of calcium and potassium sulphates at 100º C. indicates that the diagram at 100º C. C. would not be greatly different from that at 83º C. This slow rate of approach to equilibrium when combined with precalcination of,of polyhzli-,a is the 17,asis of some of the more p:olisinE methods of extracting potash from polyhalite."

Jun 1, 1930