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By P. G. Claridge, G. P. Holmes, M. A. Redfearn

"IntroductionThe purpose of this chapter is to describe current practices in operation and maintenance of flotation circuits in Canadian mineral processing plants. The chapter consists of five sections covering description of equipment, typical flow diagrams, operating practice, instrumentation and maintenance. The approach to this subject matter is intended to be practical in nature, hence no deeply technical treatment of flotation theory will be included.Flotation is an extremely common method of beneficiation of ores. Reasons are the relatively low cost of flotation compared to many other processes, the wide range of minerals to which flotation may be applied and the high efficiency of separation which can usually be achieved. The flotation process as we know it today has been developed within this century and it has had a major impact on the mining industry. While flotation is commonly considered to be important because it provides a relatively simple method of separating minerals in complex ores, the availability of flotation has also made a multitude of mineral deposits in Canada into economically viable mining operations because it allowed production of a concentrate of valuable minerals from the ore. These concentrates could then be transported from geographically remote locations to centraHzed plants for further processing, although the existence of the remote and perhaps small ore body alone would not have justified installation of metallurgical processing plants on site and the grade of ore would not have justified direct shipment of ore.Some of the more common minerals concentrated by flotation are sulphides of copper, nickel, cobalt, iron, zinc, molybdenum, lead and arsenic, as well as non-sulphides such as sodium chloride, potassium chloride, phosphates, silicates, chromite, fluorite, limestone and coal. Precious metals are also recovered by flotation, generally in association with another sulphide mineral."

Jan 1, 1989

By Evan Just

DIRECT subsidies for mine production in this country began as an outgrowth of wartime 'price regulation. The price-fixing authorities realized that the volume of production to be required from domestic copper, lead, and zinc mines would invoke costs of production considerably higher than the established ceiling prices. Accordingly, beginning Feb. 1, 1942 the Premium Price Plan was set up, which undertook to pay premiums to mines for production in excess of quota, of 5¢ per lb of copper and 2Y4¢ per lb of zinc or lead. Quotas, except for small mines, were initially established at the average rate of production in 1941, but were later reduced arbitrarily on an individual basis. Subsequently, additional "B" premiums were set for lead and zinc of 2Y40 per lb and a further "C" zinc premium of 2%¢. Additional copper premiums were negotiated on an individual basis, mainly with small mines. Quotas under each of these successive categories were the production obtained at the next lower premium level. In other words, under ceiling prices of 12¢ for copper, 6,14¢ for lead, and 83¢ for zinc, it was possible for high-cost operations to receive prices plus premiums of 17¢ or more for copper, 12¢ for lead, and 16% 0 for zinc on a portion of their production. On the other hand, some of the larger operators received only slight premiums and in at least one instance none at all. The average price plus premium for copper was 12.25¢ in 1942 and rose to 13.65¢ in 1945. That for lead was 6.99¢ in 1942 and 8.73¢ in 1945, and for zinc it was 9.06¢ in 1942 and 11.200 in 1945.

Jan 1, 1948

By X. Nie

A new one-step process was developed to completely eliminate iron and directly obtain pure Ni, Co, and Cu chloride catholyte from nickel rich matte by low temperature (400 0c) chlorination roasting of Ni, Co, and Cu with HCl and leaching with water (PH 7). The effects of roasting temperature, time, oxygen potential and HCI addition and leaching temperature, pH, H2S04 addition and residue sulphur on the recovery of metals and purity of leaching solution were investigated. More than 99 % of Ni and Co and 95 % of Cu in the matte was directly extracted into the solution while 99.995 % of Fe remained in the residue. The iron concentration in the leaching solution was less than 3.6 X 10-5 mol/L.

Jan 1, 1994

By W. A. Dunkley, W. W. Odell

About two-thirds of the manufactured gas supplied to the public by the gas plants in the Illinois district is cnrbureted water gas. The leading generator fuel is coke, ma.de in by-product c.oke ovens or in retorts, chiefly from eastern gas coals. During the fall and winter of 1917-18, when the demand for coke by war industries and the restrictions of the Fuel Administration in connection with the transportation of coke from eastern States made it almost impossible for many gas companies to obtain their coke requireml'nts, several water-gas plants had to use bituminous coal as generator fud. Many operating difficulties were encountered, some of which were solved more or less satisfactorily in different plants. A number of plants attacked the problem rather blindly because facilities for making tests were lacking and because technical men were too busy with labor and supply problems to study thoroughly the conditions involved or to ascertain what other plants had done with this fuel. It was therefore with the hope of assisting the gas industry during this critical period, and of promoting the use of Central District coals for gas making, that the Illinois Mining Investigations, a joint agency of the U. S. Bureau of Mines, the University of Illinois, and the Illinois Geological Survey Division undertook the, tests discussed in this paper. • This report was compiled under the auspices of a technical committee appointed by Gov. Frank 0. Lowden to promote the use of Central District coals in gas manufacture as a war measure. This committee included the following representatives of the various agencies concerned: U. S. Bureau of Mines, E. A. Holbrook and W. W. Odell; Illinois Gas Association, C. W. Bradley and R. B. Harper; University of Illinois, Dean C. R. Richards, Prof. S. W. Parr, and Prof. H. H. Stoek; State Geological Survey Division, F. W. De Wolf, G. H. Cady, a.nd W. A. Dunkley. F. W. De Wolf was chairxnan of the committee and E. A. Holbrook was secretary.

Jan 1, 1924

By J M. Wallis

Heavy mineral sand exploration drilling prior to the 1970s generally utilised open hole auger and air-blast type drilling. However, mining reconciliations were poor and limitations on accuracy and precision of drilling results were severe.Working with the exploration companies, Wallis Drilling principals designed various systems, including what was termed the ‘Eneabba System’ and then the ‘Capel Twin-Tube System’. Eventually the company with its improvements developed and patented the Aircore Drilling System.Aircore drilling has subsequently become recognised as an excellent drilling system for exploration and mine-sample drilling that gives good precision and accuracy in sampling both above and below water. Its use has also been extended to many other commodities and is used in a number of overseas countries.This paper documents anecdotally the development of the system, identifying problems encountered during the times it was evolving and of the means to overcome the problems as being the documentation of the story by the inventors and developers in Wallis Drilling Company. Aircore drilling system is a Western Australian technological development owing its existence to the designing and incorporation of a number of discrete improvements along the way and recognises the tenacity and inventiveness of all who contributed to the system, in both Wallis and also in a number of companies that gave the impetus to succeed.

Oct 5, 2011

"The Australasian Joint Ore Reserves Committee (JORC), which is sponsored by the Australian mining industry and its professional organisations, developed the JORC Code to provide minimum standards for public reporting. JORC was first published in 1989 and later incorporated into the Listing Rules of the Australian and New Zealand Stock Exchanges making compliance mandatory for listing public companies in Australia and New Zealand."

Jan 1, 2003

By Steve Hubbard, Brad Marchant, Rolando Cuadra

"Triton Mining Corporation acquired the rights to Mina El Limon in Nicaragua in 1993 by international bid. The bid included the gold operations at Limon and an exploration concession of approximately 4000 km2. Triton assumed control of operations and exploration in May 1994.The Limon mine is located in west-central Nicaragua, approximately 100 kilometers northwest of Managua. The climate is tropical, characterized by a wet season from May through October, and a dry season November through April. Gold at Limon is associated with both banded quartz and silicified breccia veins in Tertiary volcanic rocks, belonging to the lower Coyol Group. Mining is from two open pits and one mechanized underground operation. A second underground mine is currently under development for production during 1997.This paper describes the expansion of the milling facility at El Limon from a nominal capacity of 400 tonnes/day, producing approximately 20,000 ounces of gold annually, to the existing 900 tonne per day mill operation. Engineering of the new mill facility was initiated in September 1994 and construction was completed a year later."

Jan 1, 1997

By W. G. Jewitt

THE Box mine, situated in Saskatchewan on the north shore of lake Athabasca, is unique in two respects. It is the first and, so far, the only gold mine in Saskatchewan, and it is the only gold mine in Canada to employ a large-scale forced caving system as the main mining method. Discovered in 1934 by Tom Box, an Edmonton prospector, its subsequent history, while breaking no records for speed, has been one of steady and more or less normal development. Camp and plant are situated close to the shore of Beaverlodge bay, sorne 50 miles east of the Saskatchewan-Alberta boundary and 35 miles south of the north boundary of the Province. The railhead at Waterways, Alberta, is about 220 miles to the southwest. Prince Albert lies some 450 miles to the south. The surrounding country presents a rather rugged variation of the usual Precambrian topography, with hills rising five to six hundred feet above lake Athabasca. The numerous lakes and islands, together with the wide expanse of lake Athabasca, compose an attractive scene, at least in the not over-long 'open water' season. The winters are long but not extremely cold, judged by prairie standards. The small lakes break-up around May 20th and the 'freeze-up' stops navigation early in October. Though apparently remote, the Box mine is fortunate in having reason-ably cheap transportation from Edmonton. The 300-mile railway from Edmonton to Waterways is the most expensive section, the rate for 1st class freight being $1.70 per hundred pounds and for 5th class freight $0.77 per hundred. Freight from Waterways to the mine, approximately 140 miles down the Athabasca river and 90 miles along lake Athabasca, is delivered by barges and tugs for 50 cents per hundred pounds. All freight is delivered during the 'open water' shipping season, which lasts from about .Tune 7th to October 7th. In normal years, barges are loaded to 3 to 3 1/2feet draught up to September 1st. After that date, and earlier in dry years, shallow water at Waterways reduces the barge loads. Barges of 100 to 200 tons capacity handle all mining freight, including pieces up to 15 tons in weight, without undue difficulty.

Jan 1, 1940

By Thomas Gales

Longwall mining has a high potential for producing bulk production from a single face. Production is interrupted for a considerable time when it is necessary to move the equipment to a new block of coal. The frequency of such moves can be reduced by increasing the dimension of the block to be moved. Merely doubling the tonnage in a block to be mined may increase production by more than 21,000 tons per year. The time taken to move the equipment can be reduced by careful preplanning, industrial engineering analysis, and specially designed equipment. Reducing its move time from 30 days to five days can increase production by more than 25,000 tons per year. Increasing the panel dimensions and reducing the move time can give an increase in pretax profits of more than $600,000. A comprehensive bibliography is included.

Jan 1, 1977

By Robert T. Chapman

Many of us are privileged in that we have witnessed the transition from mining methods used in the first quarter of the twentieth' century to those practiced today. In the late Twenties, for example, many mines still used hand mucking methods. In the middle Thirties, more than one large metallurgical limestone quarry still used the 4 cubic yard railway steam shovel formerly in use at the Panama Canal. Mules could occasionally still be found underground, where they provided the-main-source of motive power. Within the last decade, and even within the past five years or less, great improvements have been made in blast-hole drilling, ore loading, and ore hauling equipment. Industrial demand for minerals can be expressed in exponential terms as compared to that at the turn of the century. Even marine mineral resources are receiving accelerated attention. One overall result is that larger and lower grade orebodies can be exploited at a profit. This profit must be made despite, for example, a static monetary price for gold, and despite the continual cost escalation involving labor, materials, supplies, and equipment. Accordingly, the application of the necessary degree of knowledge and judgment required to meet production, cost, and profit goals is paramount.

Jan 1, 1966

By Lynn Huntsman

INTRODUCTION The Sunnyside mines of Kaiser Steel Corp. are lo¬cated on the west flank of the Book Cliff Range, a tongue of the Wasatch mountain range in Utah. The mines are 193.12 air km (120 air miles) southeast of Salt Lake City and 40.23 km (25 miles) east of the town of Price. The area has a semiarid climate with an an¬nual rainfall of 304.8 to 381 mm (12 to 15 in.) and a temperature range of -23.3 to 32.2°C (-10 to 90°F). The base of the Book Cliff Range is approximately 1981 m (6500 ft) in elevation with some mountain peaks in excess of 3048 m (10,000 ft) (Fig. 1). The Sunnyside mines were opened in the late 1890s and have been operated continuously since that time. Kaiser Steel leased the No. 2 mine in 1942 and pur¬chased the entire property in 1950. There are three mines adjacent to one another, each with its separate access, ventilation, and haulage. The three mines en¬compass an area extending 12.87 km (8 miles) along the outcrop and dipping northeasterly 2.41 to 3.22 km (11/z to 2 miles) into the Book Cliffs. During 1975, only the No. I and No. 3 mines were operating to sup¬ply 908 184.7 t/a (1,000,000+ stpy) of clean coal to the Kaiser Steel mill at Fontana, CA. Sunnyside coal is blended with coal from the Kaiser Steel mine near Raton, NM, and with purchased coal from noncompany sources to form the mix for the coke ovens. At that time, Sunnyside operated three longwall units and five to seven continuous miner units. Typical production for that time came from one miner unit pillaring an area not conducive to longwall mining; one or two miner units on construction or modernization projects, and the remaining units used to develop future longwall panels. Longwall mining has been a part of the Sunnyside scene since 1961 when the first panel was set up in the lower Sunnyside seam of the No. 3 mine. Through 1975, 27 longwall panels were mined (Fig. 2). Of the more than 47 173 604 t (52,000,000 st) of clean coal that were taken from the property through 1975, 6 350 292.9 t (7,000,000 st) were mined by longwall methods. Coal was brought to the surface in 4.5-t (5-st) mine cars and crushed to -152.4 mm (-6 in.) before entering the preparation plant. Coarse coal is cleaned by Baum jigs and fine coal (-28 mesh) is cleaned in flotation cells. The coal is reduced to 50.8 x 0 mm (2 x 0 in.) and belted to an aboveground storage pile with a ca¬pacity of approximately 113 398.1 t (125,000 st). The coal is gravity fed through two 1.52-m (5-ft) diam culverts directly into railroad cars, which pass beneath the storage pile through an arched excavation. Eighty-four 90.7-t (100-st) cars make up a unit train which makes the round trip to the steel mill in four days. Additional trains are scheduled in as demands dictate. Loading rates in excess of 13 607.8 t/h (15,000 stph) have been attained. Geology The coal seams of this area are of Cretaceous age and are classed as a medium volatile bituminous coal. coal, with a sulfur content less than 1 % and 6 to (20-mile) zone of the Book Cliffs, and this area is owned or under lease by either Kaiser Steel or US Steel. Coal in adjacent areas is noncoking but an excellent steam coal, with a sulfur content of less than 1 % and 6 to 7% ash. Sunnyside coal contains 40% volatile, 53% carbon, less than 1 % sulfur, and 51/2 to 61/2 % ash after washing. There are two minable seams throughout most of the Sunnyside property known locally as the upper and lower Sunnyside. Separation between the seams varies from less than 300 mm (1 ft) near the southern boundary adjacent to the US Steel Columbia mine, to over 12.19 m (40 ft) in areas of the No. 3 mine in the center of the Sunnyside property. The upper seam varies in height from 1.2 to 1.98 m (4 to 61/2 ft) while the lower seam varies in height from 1.52 to 4.28 m (5 to 14 ft) in areas of No. I mine where the two seams apparently come together. The majority of early work was con¬centrated in the upper seam which has a stronger roof and a lower sulfur content. When Kaiser Steel began mining the No. 3 mine in the early 1950s, it mined in the lower seam beneath the old workings with conven¬tional and continuous mining equipment. Numerous bounces and pressure points created by unextracted pillars and barriers in the upper seam made this a low¬production, high-cost mine and brought about the ad¬vent of the first longwall system. The first unit, while small by today's standards, successfully mined panels that had been all but given up and saved the mine from closure. There were areas that could not even be mined by longwall because of the erratic extraction in the upper seam. The system was successful enough to allow operations to gain access to and develop virgin areas of the upper seam. Development of longwall mining in the No. 3 mine has been confined to the upper seam since late 1964. With the almost complete extraction afforded by the longwall system, it is felt the lower seam can be successfully worked in future years (Fig. 3). In most of No. 1 mine, only the lower seam (or lower seam and upper seam combined) is minable, except in the northwest portion where a rock split has again separated the seam into two 1.83 to 2.13-m (6 to 7-ft) thick beds. In this area, the upper seam or upper split is being mined. The separation between seams at the present workings in this area varies from 0.609 to 6.1 m (2 to 20 ft). It is hoped that most of the lower split can be mined using shield supports and/or what¬ever new technology is developed in the coming years. The coal seams in this area dip to the northeast at an average grade of 10%. The dip near the outcrop reaches 22% and gradually decreases downdip. Over¬burden ranges from a few meters at the outcrop to 762 m (2500 ft) over some of the deeper workings. In the next several years, some overburden will be in excess of 914 m (3000 ft) as the lower mine workings extend ever deeper under the Book Cliffs. To cope with the

Jan 1, 1982

By G. L. Oldright

The first Report of Investigations of the present series (No. 4562) was on the fundamental properties of lead compounds, and the manner in which they set limits to the richness of lead products treated in the usual sintering-blast furnace units. These properties of the compounds were emphasized, as' it was desired, also, to learn the limitations of smelting the same compounds in present ore hearth practice; the hearth has a competitor in these "usual" units, even with the rich lead feeds best adapted for the hearths. Operating conditions that were followed in older metallurgical processes, when rich products, were treated, were outlined. The second report of the series (R. I. 4563) was on the development of the ore hearths of the types now used in the Tri-State from the viewpoint of the features, functions, and dimensions of the earlier hearths that had been most changed. Experimental data, for both well and poorly desulfurized charges being considered, were given to show the relative changes in the sizes and compositions of the charges after they were added to the ore (Scotch) hearths. A general picture was given of the charge column and of the conditions prevailing there when the operation was satisfactory. In this third paper, the composition of the gases prevailing in the charge column at various elevations will be correlated with what has been learned of the nature of the charge column in order to trace the steps of operation by which lead sulfide is converted into metallic lead. The metallurgical factors that govern operation and the lines along which improvements may be made will be noted.

Jan 1, 1949

By Davis RA, Wasp EJ, Cowper NT

The transportation of solids by pipeline is not new. In 1891 a United States patent broadly covering a method of pumping coal with water was granted to W. C. Andrews. His pilot plant was constructed on what is now the site of the New- York Public Library in Manhattan. His process was applied to pumping coal mine waste back into the worked out portion of coal mines. On October 26, 1967, the world's first long distance iron ore slurry line was put into service by Savage River Mines. The commissioning of this 5.0 million dollar installation climaxed three years of development, testing, and construction and proved the Company's faith in pioneering' this unique transportation method for iron ore. This paper will discuss some of the more pertinent items in connection with that three-year period. Also to be discussed will be the operation of this line during the past eleven months.

Jan 1, 1969

By M. E. Rossi

This paper discusses the use of Conditional Simulation (CS) models to provide mill feed grade profiles for different time periods. The conventional approach of classifying block estimates as ore or waste generally does not adequately represent dilution, lost ore and the anticipated grade control data, including variability of mill feed from the pit and possible stockpiles. The detailed simulation of grade control practices can be done, but at the Pre- or Feasibility stage of the project may be impractical, since drawing dig limits on multiple benches is tedious and time consuming. Using the Feasibility Grade Control (FGC) method as developed by the Centre for Computational Geostatistics (CCG, University of Alberta-Edmonton), the CS models can be processed quickly to establish dilution and lost ore, and predict daily or short-term feed to the mill. The simulated nodes can be aggregated into selective mining units with adequate representation of likely geometries to be recovered from the pit, to mimic anticipated selectivity. An example from a large copper deposit is shown. The study provides an estimated stream of tonnages and grades fed to the plant by mining periods of interest and based on a specific mine design and schedule. The resulting cash flows support the economic and financial analyses of the Feasibility Study; also leads to improving mine schedules; and aids in developing blending schemes if necessary. INTRODUCTION The conventional approach of separating mineralized material into ore or waste categories does not always include a realistic amount of dilution and ore loss. Also, often it does not provide a good estimate of the grade that will be fed to the mill. If the grade control estimates are unbiased, they will over the long run represent the long-term average of grades fed to the mill; but the anticipated variability of mill feed from the pit and stockpile sources may be misrepresented. Some references on traditional grade control implementations can be found in Chunnett (1982); Davis et al. (1989); Srivastava et al. (1992); and Nowak (1992). Alternative methods based on economic variables improved significantly on the purely grade-based control methods, see Douglas et al. (1994). Further developments involved the use of geostatistical conditional simulations (CS) and profit models, which in several cases have been successfully implemented (Aguilar and Rossi, 1996; Rossi, 1997; Deutsch et al., 2000; Norrena and Deutsch, 2002; Neufeld, 2005; Jewbali and Dimitrakopoulos, 2009). But they can be impractical, mostly due to the lack of understanding of the methodology involved; even if semi-automatic dig limit drawing is implemented, there is always a need for is tedious and time consuming checking and corrections, even if using a single profit model derived from the simulated grades.

Jan 1, 2019

By John D. Bradley

Stibnite is 76 miles east of Cascade, Idaho, which is 78 miles due north of Boise. The altitude at Stibnite is 6,500 feet, and between Cascade and Stibnite the Forest Service road crosses two summits - Big Creek at 6,800 feet and Warm Lake at 7,300 feet. The ownership of most of the Meadow Creek group of 190 claims, which is optioned by the Bradley Mining Co from the United Mercury Mines Co. may be traced back from their consolidation by the United Mercury Co. through ownership by prospectors and small promotional companies to the discovery of the Thunder Mountain district in 1900 by the Caswell brothers.

Jan 1, 1942

By Chas. F. Jackson

This paper is one of a series dealing with mining methods and costs, prepared and published under the sponsorship of the United States Bureau of Mines and made possible through the cooperation f many mining companies in the United States and foreign countries. The preparation of this paper follows a field study of top-slice mining during which the author visited and discussed with operating officials a large number of iron mines in the Lake Superior region. During this study special note was made of changes in top-slicing practice during the past decade, prior to which the author had acquired several years experience with the method through active connection with a number of mines employing top slicing.

Jan 1, 1931

The principal activities of the Bureau of Mines Explosives Research Center from July 1, 1962, to June 30, 1963, are described briefly in part 1. Part 2 gives short abstracts of publications that appeared during fiscal year 1963; in general, the research reported in these publications antedates the report period.

Jan 1, 1965

By Edgar Carlos García Flores

El presente texto señala al Perú como modelo de sistemas de explotación minera para Bolivia -en tanto desarrolle, prepare y explote las minas con calidad y así incrementar la producción de sus operaciones mineras. Asimismo, describe los procedimientos idóneos para seleccionar las tablas geomecánicas, la ingeniería de explosivos, equipos de acarreo y los elementos de sostenimiento adecuados para las excavaciones subterráneas con el objetivo de tener operaciones de calidad, seguras y óptimas.

Feb 1, 2015

By A Esfahanipour, M Osanloo

Determination of cut-off grade?strategy is one of the most important stages of open pit mine planning and design. It is the parameter directly influencing the financial, technical, economic, legal, environmental, social and political issues in relation to mining operation. Choosing the optimum cut-off grade strategy that maximises the economic outcome has been a major topic of research workers for nearly one century. Many researchers have contributed in devising methods and algorithms, such as dynamic programming, linear programming, optimal control and so on for various aspects of its determination. In this paper, a non-linear mathematical programming for cut-off grade strategy optimisation is presented, considering the three main stages of mining operation introduced by K F Lane. In this model maximisation of net present value of mining operation, under the three constraints of mining stagesÆ capacities, considered as the optimisation criteria. Due to the discrete representation of the mining resource, the proposed non-linear formulation is approximated by a non-linear signomial geometric programming. According to non-convexity and the complexity of the proposed model, an augmented Lagrangian genetic algorithm was used to find the optimum cut-off grade strategy under varying and fixed price circumstances. To validate the proposed non-linear model efficiency, the results were compared with the results obtained by the K F Lane methodology. It was found that the proposed non-linear model works efficiently in the determination of cut-off grade strategy. According to the simplicity of the structure of non-linear programming modelling in comparison with dynamic programming it is hoped that, further development of this model would certainly provide the ability of considering managerial and technical flexibilities as well as incorporating more real mining conditions in the determination of cut-off grade strategy optimisation.

Sep 26, 2011

By Minghao Hu, Hongbo Zhao, Jun Wang, Wenqing Qin, Rui Liao, Central South University

"In the present work, X-ray diffraction (XRD) analysis, X-ray photoelectron spectroscopy (XPS) analysis and electrochemical analysis were carried out to investigate the intermediate species and electrochemical dissolution process of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms. For bioleaching of chalcopyrite by the sulfur oxidizing microorganism Acidithiobacillus caldus, the redox potential was always lower than the appropriate value and chalcopyrite was mainly directly oxidized to polysulfide (Sn 2-), thus resulting in low copper extraction. As for bioleaching of chalcopyrite by the iron oxidizing microorganism Leptospirillum ferriphilum, chalcopyrite was mainly directly oxidized to polysulfide at the initial stage where redox potential was lower than the appropriate value and resulted in low dissolution rate. At the middle stage where redox potential was appropriate (about 350–480 mV vs. Ag/AgCl), chalcopyrite was mainly transformed to intermediate species of Cu2S rather than polysulfide, thus resulting in high dissolution rate. At the later stage where redox potential was higher than the appropriate value, chalcopyrite was mainly directly oxidized to polysulfide which caused the later passivation of chalcopyrite. Finally, a model of electrochemical dissolution process of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms was provided.INTRODUCTIONBiohydrometallurgy is considered as one promising mineral processing technique in processing low grade ores due to the advantages of simplicity, low cost and eco-friendly. Chalcopyrite (CuFeS2) is the most abundant copper-bearing resource, accounting for appropriately 70% of the known reserves in the world, while is recalcitrant to bio-hydrometallurgy mainly due to the coated passivation layer (Klauber, 2008; Li, Kawashima, Li, Chandra, & Gerson, 2013). However, the detailed dissolution and passivation mechanisms of chalcopyrite bioleaching are still uncertain. The microorganisms assisting chalcopyrite dissolution can be mainly classified into three types, including mesophilic microorganisms (e.g., Acidithiobacillus ferrooxidans, Acidithiobacillus thiooxidans and Leptospirillum ferrooxidans), moderately thermophilic microorganisms (e.g., Acidithiobacillus caldus, Sulfobacillus thermosulfidooxidans and Leptospirillum ferriphilum), and extremely thermophilic microorganisms (e.g., Metallosphaera sedula, Sulfolobus sp., Acidianus sp., and Sulfurcoccus mirabilis). Moderately thermophilic microorganisms have a better prospect in the future industrial application of chalcopyrite bio-hydrometallurgy (Olson, Brierley, & Brierley, 2003; Rohwerder, Gehrke, Kinzler, & Sand, 2003). Hence, investigating the detailed dissolution and passivation mechanisms of chalcopyrite bioleaching by moderately thermophilic microorganisms is of important significance, which will provide essential information for promoting chalcopyrite bioleaching by moderately thermophilic microorganisms in laboratory or industrial plant."

Jan 1, 2016