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By K. A. Natarajan, Abha Kumari

"The effect of applied DC potentials on the bioleaching of a chalcopyrite concentrate in the presence of Acidithiobacillus ferrooxidans is discussed. Copper dissolution was the highest at an applied potential of +600mV (SCE), while all the dissolved copper got cathodically deposited at a negative potential of -600mV (SCE). Electrobioleaching at an applied potential of +600mV (SCE) was established at different pulp densities as a function of time. The effect of applied potentials and electrolytic currents on the activity and growth of bacterial cells was assessed. Preadaptation of bacterial cells to the concentrate slurry and electrolytic growth conditions significantly enhanced copper dissolution. Electrochemical and biochemical mechanisms involved in electrobioleaching are illustrated with respect to oxidative dissolution and biocatalysis of anodic oxidation. IntroductionIn recent years, the bioleaching of low-grade chalcopyrite ores as well as their flotation concentrates has been receiving increased attention. Use of mesophilic bacteria such as Acidithiobacillus ferrooxidans in the bioleaching of chalcopyrite does not yield desirable copper extraction rates due to its easy passivation.Bioleaching of chalcopyrite using A. ferrooxidans has been reported to be very inefficient, and use of thermophilic bacteria at higher temperatures is recommended to improve its bioleaching rates. Bioleaching under imposed applied DC potentials or currents has been shown to enhance not only bioleaching kinetics but also to confer selectivity in mineral dissolution (Abhakumari, 1996; Chirpa Selvi and Natarajan, 1998; Natarajan, 1992a, b,c). Electrochemical principles governing electrobioleaching have been earlier illustrated (Jyothi et.al, 1989). The role of applied potentials on the activity and growth of A. ferrooxidans has also been studied (Natarajan, 1992b). Also, the electrochemical aspects of bioleaching of several sulfide minerals when present alone or in binary and ternary combinations have been reported (Jyothi et al., 1989). The electroleaching of copper-nickel concentrate (Natarajan et al., 1985) and electrobioleaching of sphalerite concentrates have already been studied (Selvi et al.,1998). However, it is possible to enhance the bioleaching rate of chalcopyrite in the presence of A. ferrooxidans under galvanic conditions in the presence of nobler minerals such as pyrite (Jyothi et al.,1989). Also, dissolution of chalcopyrite can be significantly enhanced under applied potential conditions. Ali Ahmadi and his coworkers recently studied the electrochemical bioleaching of chalcopyrite concentrates using moderate thermophiles (Ahmadi et al., 2010, 2011)."

Jan 1, 2015

By David Atkin

This paper presents methods and results of environmental planning conducted for a proposed coal and iron ore transshipment terminal which would be located on reclaimed land west of Lisbon, Portugal. Oceanographic, navigational, marine, structural, material handling, and financial studies of the terminal were conducted along with the environmental analyses. However, because of the proximity of the terminal to heavily used recreational resources, residential areas, and biologically valuable habitat, anticipating and reducing the environmental effects of the terminal to minimum levels was a very important part of the overall project. The results of the study indicate that even the impacts associated with worst-case conditions would be acceptable because of appropriate pollution control measures and planning.

Jan 1, 1983

By N. Edmonson

The demand for sulfur divides into two major categories: agricultural and industrial. Most of the agricultural demand is related to the production of phosphate fertilizer products. Industrial demand is derived from the demands for a large number of products and is related to a variety of industrial processes, no one of which characterizes it in a simple way.

Jan 1, 1994

By Yuting Xue, Brijes Mishra

The relaxation test on sandstone specimens showed typical behavior in pre-failure region and stepwise behavior in the postfailure region. Overall, more significant stress relaxation occurred within the failed specimens than the intact ones. Numerical simulations were conducted with pre-defined failure plane and with Voronoi tessellation to visualize the relaxation behavior. The model with pre-defined failure plane showed the key role of failure plane and asperity in simulating the step-wise post-failure relaxation behavior. The inhomogeneous stress distribution within failed specimen and the viscous deformation of intact rocks induced high stress concentration at the asperities. The observed step-wise relaxation initiated from the failure of asperity. Furthermore, the simulations with Voronoi tessellation showed time-dependent fracture development during relaxation in post-failure region. The presence of fractures completely changed the stress distribution. Stress concentration occurred at the front area of fractures and at the interacting areas between fractures. During relaxation, fractures still developed with time and the sudden significant increase in the “damage” coincided with the acceleration of stress relaxation leading to step-wise relaxation. Finally, the results showed the possibility of using residual strength as the long-term strength of failed rock.

Jun 25, 2020

By Immo H. Redeker

Introduction In October of 1961 Texas Gulf Sulphur Company contacted the Asheville Minerals Research Laboratory, of the North Carolina State University, and requested assistance in the development of a phosphate ore body in Beaufort County. The main: objectives of the work at Asheville were defined as follows. The Laboratory should try to determine on the basis of drill core samples, submitted by Texas Gulf Sulphur Company, the highest grade of phosphate concentrate that could be obtained with high recoveries from North Carolina ore, the thickness of the ore body that could be treated economically in one section, and the quantity of concentrate that could be obtained per acre. The results should be compared with data obtained from the treatment of Florida ores. It was important to use mineral preparation methods during the initial laboratory work that could be verified later by continuous pilot plant operation and could be scaled-up for a commercial plant. The writer was in charge of the project at the Asheville Minerals Research Laboratory and participated in the development of one of North Carolina's greatest mineral resources from the first drill core examination to the production of the first poundage and tonnage quantities of flotation concentrates. The Lee Creek phosphate mine and mill was put in operation in early 1966 and can produce three million tons of phosphate concentrate a year. Sulfuric- acid, and phosphoric acid plant, diamonium phosphate plant and triple super phosphate plants came on stream in 1967; at that time they were the world's largest. Now that a phosphate industry has been established in the State, it]

Jan 1, 1971

By Ramon B. Santos

The Atlas Concentrator, reputedly the largest in the Far East today, is located on the island of Cebu in the Republic of the Philippines. It started operation on a 4000 TPD capacity in 1955. An expansion program, coupled with significant improvements both in the Crushing Plant and the Mil, made the Atlas Concentrator what it is today, treating about 14,000 tons of ore per day at a comparatively low cost of about $0.50 per ton treated. The Atlas Copper Mine is at present mainly an open pit operation, with ore supplied from two pits (Lutopan and Biga) north of the Concen¬trator. Development of an Underground Mine, situated on the main ore body at Lutopan, which will be a block caving operation, has been underway since 1963, and this underground mining operation is expected to b e in full production by the latter part of 1966.

Jan 1, 1966

By R. S. C. Rogers

The present study examines the use of two vs. one stage of hydrocyclones in a pilot scale ball mill circuit. The approach used for doing this is based on computer simulation with mathematical models that explicitly account for the grinding mill/classification system interaction. Parameters for the ball mill model are obtained using a short-lived radioactive tracer method in open circuit tests of the study mill. The hydrocyclone model, based on the size selectivity concept, is developed and validated in standalone testing of the study hydrocyclones. Simulation procedures are validated through actual tests of the closed circuit grinding configurations of interest. Results of the evaluation indicate that two stages of hydrocyclones in the study circuit would improve classification, increase productivity and lead to more closely sized circuit products when compared to the conventional one stage hydrocyclone circuit. A computer simulation approach for comparing these circuit configurations in larger scale grinding circuits is also suggested.

Jan 1, 1981

By P. R. ZETTWOOG

HISTORICAL BACKGROUND A program in France to develop personal [a] dosimeters has been initiated 1974. The patent on which is based the present device was obtained in 1972 * . From 1972 to 1974, the possibilities of applying certain ionograph track detectors to the spectrodosimetry of radon daughters was explored. The first prototype were produced in 1974. It took four years (from 1974 to 1978) to produce an autonomous dosimeter whose components has a sufficient life span, especially for the turbine motor unit. Qualification in the laboratory was obtained in 1977. In 1978 it was obtained in the mine for technology (autonomy of 12 hours and a life span of more than one year) and in 1980 for monitoring. 300 dosimeters have been tested in underground mines all together. Indispensable peripheral equipment were also developed from 1976 to 1980 : calibration devices, equipment to prepare and develop the films, read out systems. The concept of an "Integrated System of Individual Dosimetry" (ISID) based on a personal [a] dosimeter measuring exposure to radon daughters, thoron daughters, ore dust and external irradiation doses was proposed at the end of 1980. Since January 1st 1981, ISID is used on a routine basis in some french mines, situated in remote area, and appears to be very competitive with the ambiant dosimetry. The latest version of the dosimeter is produced in mass series since June 1981 and should equip all french mines in 1982. DESCRIPTION OF THE INSTRUMENTATION DEVELOPMENT OF THE DOSIMETER MEASURING HEAD The measuring head is based on the use of ionographic film to detect a tracks. In fact, the measuring head is a spectrodosimeter which measures separately over the period of exposure: - the potential [a] energy inhaled due to the decay of Po 218, Po 214, and Po 212 ; - the number of Rn 222 atoms inhaled ; - the inhaled total [a] activity of the five long-lived emitters present in the ore dust. The contribution to the total inhalable potential [a] energy of these various radionuclidesin a typical underground mine is studied in Appendix I. The measuring head described in detail in Appendix II, is able to satisfy all the implications made in the ICRP recommendations. Appendix III deals with the use of this measuring head in the cases where the equilibrium factor is lower than 0.1. This situation occur in open-pit mines where account must be taken of the Rn 222 contribution, which is no longer negligible in relation to that of its daughters. CURRENT DOSIMETER PERFORMANCES Table I shows the characteristics of the latest dosimeter. Appendix IV should be consulted concerning qualification of the dosimeter in the laboratory and in mines, technological development which finally produced the [a] dosimeter and its peripheral equipment, and technical presentation of the ISID (Integrated System of Individual Dosimetry) based on the concept of a multirisk personal dosimeter. Data on the installation and operating costs of such a dosimeter, which would seem to be competitive, are also given in this Appendix. ADVANTAGES OF PERSONAL DOSIMETRY AS COMPARED TO AREA MONITORING The results of the first eight months of experiments carried out under real conditions in an underground mine site are given in detail in reference 8. Area monitoring : The monthly exposure per worker to inhaled Rn 222 was determined from the knowledge of time spent in various areas of the mine and for the different mining operations, as well as from numerous and systematic sampling of the Rn 222 concentration in all work places. Personal dosimetry : The exposure to potential energy from radon daughters was measured by an [a] dosimeter developed by the CEA and worn by each of the 160 miners during eight months. In this way 160 x 8 pairs of monthly individual exposure values have been obtained which can be statistically studied. This test was decisive for us because it proved that the [a] dosimeter was technically sound (very few defects over one year for 160 dosimeters) and especially that personal monitoring devices were superior to area monitoring devices. The following conclusions can be drawn 1. The exposure distribution obtained by personal dosimetry is log-normal. This is true for the results on the whole as well as for groups of results relating to certain explanatory variables. See fig. la, 1b, 1c. 2. The exposure distribution obtained by area monitoring does not correspond to any type of distribution. If the results of personal monitoring are taken as a reference, area monitoring tends to underestimate the high exposures and overestimate the low exposures.See fig. 2. 3. [a]-energy exposures are underestimated when calculated from radon exposures and the equilibrium factor found in the considered mines. This is due to episodes or to zones of high radon concentrations not registered

Jan 1, 1981

By Roy H. Paddon

This paper describes the 6, 000 tpd, three stage haulage system at the Ozark Lead Company's mine in southeastern Missouri. It details the methods and problems pertinent to the loading, hauling, sorting, crushing, storing and hoisting of waste rock and ores obtained from underground development headings and production areas. The Ozark Lead Company is a wholly owned subsidiary of the Kennecott Copper Corporation. Its mine is located nine miles south of Bunker in Reynolds County at the south end of the "New Lead Belt" mining district. Development started in May, 1964, by drilling a seven foot diameter (six foot cased), 1, 250 foot development-ventilation shaft. Drifting began in August, 1965, to obtain a bulk ore sample, to test ore continuity, and to drive two mine levels while the pre-grouted production shaft was being sunk. Over 7, 900 feet of 12 by 24 foot drifting was done from this small shaft. In March, 1966, a 20 foot diameter, three compartment, concrete lined production shaft was collared. It intersected existing development headings at a depth of 1, 195 feet in April, 1967, and was completed along with a concentrator and other facilities in June, 1968. Mine production began that month, with 5, 000 tpd reached in 1970.\

Jan 1, 1971

By Louis P. Gignac

The Lac Shortt gold deposit is located in Northwestern Quebec, at some 115 km west of the Opemiska Division and 380 km northeast of the Lac Dufault Division, as shown on Figure 1. Falconbridge's exploration crews drilled the first six holes in the gold-bearing ore zone in 1979. After further diamond drilling from surface in 1980, the decision was taken to carry out an underground exploration program. A three-compartment shaft was sunk to a depth of 340 metres in 1981 and 1982; it was followed by an extensive exploration program, which included 1073 metres of cross-cutting and footwall drifting, 645 metres of ore drifting and 11 200 metres of diamond drilling. The underground work was completed by mid-winter 1983. It was followed by a complete feasibility study of the Lac Shortt Project. After some $Can 14.2 millions in exploration expenditures, the Board of Directors of Corp. Falconbridge Copper approved in July 1983 additional expenditures of $Can 45 millions to bring the Lac Shortt Mine in production at the capacity of 261 000 tonnes per year. Gold price was $US 425 per ounce at the time.

Jan 1, 1985

By Stephen M. Voynick

Two hours driving time west of Denver, CO and 21 km (13 miles) north of Leadville, Bartlett Mountain rises 4,065 m (13,337 ft) high in central Colorado's Tenmile Range. Bartlett Mountain is one of Colorado's most recognizable peaks, especially when the late afternoon sun casts a surreal, reddish light across its most prominent feature - the Glory Hole. This is a gaping, half-mile-wide, 245-m- (800-ft-) deep crater. Created by subsidence from 60 years of massive underground mining, the Glory Hole is a monument to the Climax Mine, one of the great achievements in American mining and metallurgy. For most of this century, the name Climax has been synonymous with brilliant mining engineering, innovative metallurgical research and entrepreneurial promotion of a once-obscure metal - molybdenum. Climax overcame an array of economic, technological and social problems to become a model for underground development of massive deposits of low-grade ores. It is one of the few mines in the world where cumulative production is measured in the billions of dollars. The Climax mineralization was emplaced 30 million years ago. Magmatic intrusions formed a large, mineralized stockwork within the granite country rock of Bartlett Mountain. Associated hydrothermal events forced superheated metal- and quartz-rich solutions through fracture zones, silicifying the country rock and depositing huge amounts of molybdenite, or molybdenum disulfide. The mineralized rock, a light gray, finely crystalline, silicified granite, is laced throughout with blue-gray quartz veinlets containing molybdenite. Although the richest ores graded only about 1% molybdenite, the Bartlett Mountain deposit contained more than 2.7 Mt (3 million st) of elemental mo- lybdenum.

Jan 1, 1997

By Howard N. Egan

EASTERN COAL CORPORATION is located in the heart of the Appalachian Mountain area of extreme Eastern Kentucky, The mountains are of sedimentary origin, and have been deeply eroded by streams as the area has been thrust up during geologic time. There are many seams of coal interspersed between beds of shale and sandstone. While some of the seams are thin, there are a number that can be economically mined. We are mining the Pond Creek, or No. 2 Gas Seam, at a rate in excess of 6,000 tons per day. Our central preparation plant utilizes tromp baths for the plus 3/8" size and Rheolaveur for the minus 3/8" portion. No provision was made to recover the fine coal, which was included with the plant effluent pumped to a 12 acre refuse pond. The pond is impounded behind the coarse refuse dump. This pond rose 15 feet last year. Obviously, no pond could stand many years at this rate. Limited pond space in this rugged area of steep hills and narrow valleys dictated a revised disposal method. Our first thought was that it would be necessary to dry the fine refuse sufficiently to permit inclusion with the coarse refuse, which is hauled uphill by monitor and then trucked to final disposal.

Jan 1, 1967

By H. D. Dahl

In January, 1969, the Central Research Division of Continental Oil Company initiated a research program on roof control. The purpose of the research was to improve roof stability. Toward this end, it was necessary to determine which of the many geological and geometrical variables have a significant effect on roof stability. The isolation of these significant variables would allow research effort to be concentrated where it would be most effective in improving roof control. The research effort was directed toward two major objectives. One was to de-fine the geological and/or geometrical parameters that affect the severity of roof conditions in any particular area of the mine. This study was concerned with roof fall intensity. The second was to define why roof falls in Humphrey No. mine (and in northern West Virginia-south-western Pennsylvania coal mining areas) are oriented so that they occur primarily in north-south rooms or entries. This study was concerned with roof fall orientation. To provide a reference framework for further discussion, a geological description of the area is presented in Part II. In Parts III and IV, the procedures, methods, and 'results of the research applied toward the afore-mentioned major objectives are presented. Part V contains the conclusions and recommendations for-warded by this study.

Jan 1, 1971

By J. H. Wilhelm

Continuous thickeners can be sized either from previous experience, continuous pilot-scale test results, or from batch settling tests. Although it is a relatively simple process to size full-scale thickeners when previous experience or pilot-scale operating data are available, batch testing is often the only realistic method for sizing thickeners due to time restrictions, budget limitations and availability of sample material. Many authors have, therefore, suggested means of running batch tests and analyzing the resulting data. Fundamental thickening theory had its beginning with the work o£ Coe and Clevenger(1). In order to determine the size of the thickener settling area, they proposed that a series of batch settling tests be run at various initial concentrations. The solids handling capacity at each concentration can then be determined and the lowest capacity (maximum area requirements) used to size the thickener. Kynch(2) developed a mathematical analysis in which the relationship between solids concentration and settling rate can be obtained from a single batch settling curve. Talmage and Fitch(3), Yoshioka(4)(5) and Dick(6), have enlarged upon these works to give a basis for methods of sizing thickeners.

Jan 1, 1979

By A. E. Koski

This paper describes the implementation of specialized software in mine planning at the Cleveland-Cliffs' managed Empire Mine on Michigan's Marquette Iron Range. The application puts together years of mining experience on the eastern end of the Marquette Range beginning in the early 1960's. The Empire Mine, in its thirty-fifth year of production, uses the software to model a complex and highly variable low grade magnetite ore body, capturing the economics of mining and processing to the final product. With special attention given to geostatistical methods, the software is used for all geologic modeling, ore reserve estimation, ore control, pit design and short and long range planning.

Jan 1, 1999

By James J. Fallon

The industrial minerals industry is defined loosely and a specific list is presented. Criteria by which to judge prospects and examples are provided. General trends of the industry are observed.

Jan 1, 1984

By W. L. Josey

The Stibnite Project is an open pit/cyanide heap leach gold property located in the Yellow Pine mining district of central Idaho. The gold ore is in an oxidized zone of an epithermal deposit which is associated with a fault system related to Tertiary igneous activity of the region. The faults transect a roof pendant of metamorphosed sediments within crystalline rocks of the Idaho Batholith. Seasonal mining is at a rate of 3900 mt/d (4300 st/d) of ore with a 2:l waste to ore ratio. Gold and silver are recovered through a process of cyanidation, carbon adsorption and desorption, and electrowinning. Disposal of the leached ore requires chlorination of the residual cyanide.

Jan 1, 1984

By F. Ikumapayi

Hydrogen peroxide production was measured during the grinding of a complex sulfide ore, and its oxidizing effect on solid surfaces was investigated using Fourier transform infrared spectroscopy (FTIR) with diffuse reflectance attachment measurement. In turn, an attempt was made to correlate the formation of hydrogen peroxide, surface oxidation and sphalerite flotation. Additionally, in order to predict and minimize detrimental production problems due to the recycling of process water in sulfide ore processing, the effects of major components of calcium and sulfate species present in recycled process water and the effect of temperature on sphalerite flotation were investigated through bench-scale flotation tests using complex sulfide ores. The significance of process water species in flotation was studied using tap water, process water and simulated water containing calcium and sulfate ions. Formation of hydrogen peroxide was revealed during the grinding of the complex sulfide ore, and its formation was counteracted by diethylenetriamine (DETA). The FTIR spectrum of the pulp solid fraction showed varying degrees of oxidized surface species, which are related to the concentration of H2O2 analyzed in pulp liquid. Bench-scale flotation using two different complex sulfide ores showed that sphalerite recovery is better in process water than in tap water. Flotation results also indicated a varied recovery of sphalerite at different temperatures in either tap water or process water.

By R. D. Smith

The Bureau of Mines (BOM) evaluated the technology, capital investment, and annual operating costs of emission control for further reductions of lead in ambient air at the East Helena, Montana, smelter, and the Glover, Herculaneum, and Buick smelter-refineries in Missouri. This study seeks to analyze the technology available to the domestic primary lead industry for attempting to further reduce lead emissions below 1985 levels. Through enclosure of concentrate unloading and handling areas and all lead-processing buildings, followed by venting these enclosures to the atmosphere through baghouses and the application of appropriate local controls, emissions can theoretically be reduced by a total of 2,424 pounds of lead per day from 1985 levels at the three Missouri operations. Using conventional technology, aggregated additional capital costs of $250,014,000 and annual operating costs of $18,507,800 are estimated for the four operations, resulting in a total cost of $0.063 per pound of refined lead. Continuous drossing techniques would assist in meeting OSHA standards in process areas. Continuous drossing would change capital and operating costs to $258,344,800 and $16,104,700, respectively, for a total cost of $0.062 per pound of refined lead. First quarter, 1986 open market price for refined lead was $0.184 per pound.

Jan 1, 1986

By Gunnar Nord, Per-Anders Persson, Duri Prader

The paper gives a short resumé of the findings of the joint US - Swedish "Mechanical Boring (MB) or Drill and Blast Tunneling (DB) ? Workshop" in Stockholm December 1976. New developments and trends in both MB and DB techniques in Europe since 1976 are described, and information is given on some interesting new projects. Differences in contracting principles between the US and Europe lead to different practices for preinvestigations, design, excavation and arbitration and legal affairs. It is shown that this often leads to a higher cost in the US than in Europe for similar excavation jobs. A typical US tunnel job also has more men on the site than a similar job in Europe. This adds to the cost difference and is only partly balanced by a slightly higher rate of excavation in the US. A comparison is made between two quantitative methods of predicting MB penetration rates. These are in reasonable agreement. A probable future development in machine design is an increase in cutter load towards perhaps 25 tons per disc where the tunnel diameter is large enough to house the machine size dictated by the size of the very large electric motors required. Recent European developments in drill and blast technique include improved methods for cautious blasting, precision hydraulic drilling of small diameter holes, and non-electric interval blasting techniques. These developments still keep the drill and blast method ahead of mechanical boring costwise for tunnels in igneous rock whereas for most sedimentary materials mechanical boring has a lead, particularly for long tunnels.

Jan 1, 1979