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By A.M. Wala

A new radioactive tracer method to measure airflow in inaccessible airways has been developed for situations when studies of the mine ventilation system are not possible using conventional flowmeters (anemometers). The new method has been tested in the ventilation laboratory at the University of Kentucky and in the Experimental Underground Coal Mine at the U.S. Bureau of Mines. This new tracer method has been utilized in leakage, dilution and flow velocity measurements in laboratory and field (mine) experiments.

Jan 1, 1996

By Y. F. Vasyuchkov

In general a term "unconventional mining or "non-traditional mining technology" getting more and more familiar to the mining engineering media. There are several alternating terminological combinations such as "in situ mining", "borehole mining", "shaftless mining", "mineless mining", "geotechnology" etc.; specific terms denote particular method of unconventional mining technology: "underground coal gasification in situ", "underground coal combustion in situ", "borehole hydraulic coal mining" and others. Principal scheme of unconventional coal mining technology is shown in Figure I:

Jan 1, 1997

By A. G. Naguib

In essence, any magnetic separator operates on the basis of imparting a preferential magnetic force on particles of higher magnetic susceptibility which are to be separated from particles of lower magnetic susceptibility. A certain amount of directional deviation to the particles takes place thus allowing for collecting each category of particles separately. The forces in action on particles in a wet magnetic separator of the established designs can be analyzed in a simple manner as shown in Figure 1.

Jan 1, 1972

By Paul H. Johnson

Chemical mining may be defined as the in-place extraction of metal values from ores located within the confines of a mine (that is, broken or fractured ore, stope fill, caved material, or ores in permeable zones) or ores located in dumps, prepared ore heaps, or tailing ponds. The term solution mining has, in general, been used to mean the in-place dissolution of water-soluble mineral salts with leach solutions. Chemical mining is presently in its embryonic stage. Basic research needs to be done to understand the influence of physical and chemical factors in chemical mining, and applied research needs to be done on both a small and a large scale to develop practical techniques of field practice. The areas of research include the following: (1) Leaching chemistry, the influence of physical, chemical, and biological factors in in-place leaching (2) Hydrodynamics of chemical mining systems (3) The development of new, as well as old, leaching agents for use under chemical mining leach conditions (4) The development of new means of contacting fine and coarse materials (with solutions) under static leach conditions (5) The development of new processes for generating and regenerating chemical mining leach solutions (6) The development of metal recovery processes

Jan 1, 1969

By Charles T. Holland

The tonnage produced on mechanized automated retreating longwall faces in the United States is growing rapidly. This time last year some 14 faces were operating, and in addition to that, three were known to be down looking towards revision or repairs before being placed back in service. These, it is believed, are now in service. It is estimated at the present time, therefore, that somewhere between 20 and 25 faces are being operated in the country. A survey made last year by the School of Mines, West Virginia University, indicated that longwall mining in general has been beset with a great many problems, some more serious than others. The problems that we propose to discuss here are those pertaining to roof and floor trouble in the tailgates. Longwalls are operating in this country at the present time under an overburden ranging from 200 feet to nearly 3000 feet. Main and tailgate entries are operating in panels of up to five each, with the width of the pillars between these entries varying from 20 to 50 feet. Face lengths are running from 270 feet to 670; actually a face 800 feet long should be possible. The most economical face is often spoken of as being around 450 feet.

Jan 1, 1969

By R. J. Sawyer

*An earlier version of this paper was presented at the IMM meeting, "Use of Cultivated Carbon in Extraction of Gold and Other Metals," held in London in January, 1983, and was subsequently published in Trans. IMM (Sect. C: Mineral Process. Extr. Metall.) in December, 1983. Reprinted with permission of The Institution of Mining and Metallurgy, London. INTRODUCTION This paper describes the development of the processing route for the Alligator Ridge Mine which commenced operation in 1980. It also reviews the modifications which have been made to improve operating efficiency during the three years that the mine has been in production. Alligator Ridge is located 70 miles Northwest of Ely, Nevada. The mine is a joint venture between Amselco and Nerco with Amselco acting as operator. The deposit consists of micron sized disseminated gold imbedded in siltstone. Although definitive conclusions have yet to be drawn as to the age and origin of the deposit, it is probable that it is young by geological standards, perhaps only 5 million years old. It is likely that the gold was brought to the surface by hot springs. As the superheated solutions rose through the porous siltstone, they redeposited minerals in the rock, forming the present ore bodies.

Jan 1, 1984

By Hassan El-Shall, Rosana M. Macedo Borges, Ponisseril Somasundaran, Regis Stana, José Renato B. Lima, Patrick Zhang

The phosphate fertilizer industry is a large activity in Brazil because the soil has a phosphorous deficiency. The phosphoric acid that is produced is extensively used in the production of fertilizer. It is produced by reacting phosphate carbonaceous rock with sulfuric acid, generating phosphogypsum as a by-product. In 1991, there were about 29 million tons of phosphogypsum stored in Brazil. This has increased to an estimated 280 million tons in 2001. While the phosphogypsum is used in the cement industry and also in agriculture, the quantity used is much less than the production rate, and the stockpiles are still increasing. The radioactivity of this waste may not allow its continued use as a construction material and some other uses. However, Brazilian phosphogypsum contains a smaller quantity of radioactive materials than is reported for phosphate rocks from other countries. One important reason is the geological origin of the Brazilian apatite. All the large deposits being currently mined are igneous, not sedimentary, as is the majority of the phosphate rock produced around the world. An extensive study is being conducted to characterize the phosphogypsum generated in two production facilities in Sao Paulo state to develop safe uses for this waste and determine the environmental impact caused by the present method of phosphogypsum disposal.

Jan 1, 2002

By V. C. Kelessidis

Demand for multifunctional drilling fluids continuously increases. In the quest for more hydrocarbons, drilling has extended to deep water areas (now in excess of 2.000 m) and to very deep wells (now in excess of 6.000 m). In these extreme conditions, low temperatures, as low as 1-5°C on the sea bed, high temperatures, in excess of 200°C, and high pressures, more than 1.000 bars, in the bottom of the well, are frequently encountered. In these harsh environments, synthetic oil and oil based drilling fluids are used, but the former are expensive while strict environmental requirements have limited the use of the latter. Thus, drilling industry has turned back to water based drilling fluids which use bentonite as primary viscosifier and as a fluid loss control additive. However, in conditions of high temperatures, bentonite suspensions gel and loose their exceptional low temperatures properties, thus, additives are needed to overcome such problems. In this article we review these challenging drilling environments and establish the requirements for the multifunctional drilling fluids. Our research indicates that demand for continuing use of bentonite by drilling fluid industry is strong. Furthermore, we bring forward the opportunities for new additive development which can expand the operational ranges of bentonite suspensions.

Jan 1, 2007

By Mark S. Dunsbier, Steven J. Schatzel

Large roof outbursts began occurring in November 1986 at Domtar's Sifto Salt Mine in Goderich, Ontario, Canada, working the bedded Salina A-2 Salt unit. The outbursts, measuring about 13-20 m in diameter, released high concentrations of methane from cavities that extended into carbonate roof units. Spontaneous ignitions of methane gas during an outburst seriously burned a mine technician. The Canadian Ministry of Labour contacted the U.S. Bureau of Mines for technical assistance to aid in mine managements1 efforts in alleviating the problem. Initial gas content determinations of salt samples from the outbursts by acoustic test methodology revealed no methane enrichment. In- mine core drilling and subsequent methane content testing indicates the main source of methane in overlying roof units. Drilling the roof units demonstrated the potential of draining methane from areas showing signs of impending outbursting and alleviating the danger. Three mechanisms are considered the most likely means by which methane was introduced to the mine environment. These are (1) hydrocarbon gases released from brines, (2) gas transport along horizontal bedding planes or gas accumulations in bedding separations, and (3) mining-induced fracturing. The abundance of higher hydrocarbons in gas yielded by non-evaporite rock samples from the Goderich mine may be indicative of a former association with liquid petroleum.

Jan 1, 1989

By G. Konstantopoulou

Environmental degradation due to abandoned quarries is an international problem, more intense when these sites are found in urban or suburban zones, in environmentally sensitive areas or are visible from sites of special interest. Rehabilitation solutions applied worldwide, range from simple relief restoration and planting to more sophisticated site exploitation in order to serve specific social needs. The elaboration of rational rehabilitation plans for every site presupposes deep knowledge and comparative evaluation of all the parameters of the problem. In this paper, a methodological approach of rehabilitation solutions for abandoned quarry sites is elaborated, based on domestic and international experience and practice. Solutions are grouped in eight main categories. In every category the framework of prerequisites and confinements (environmental, technical, social, economic, legal) for their application in Greek quarry site is investigated. A comparative assessment of alternative restoration solutions is introduced, using tools supplied by environmental economics, cost-benefit analysis, G.I.S. and multi-criteria analysis and follows in two stages: 1) Establishment of a set of criteria aiming at a priority list for a number of quarry sites located in a geographical or administrative district. 2) Assessment of alternative solutions for every single quarry from the previous priority list for the selection of the most adequate using technical, economical, social and environmental criteria.

Jan 1, 2008

By W. F. Riggs

INTRODUCTION Flotation equipment is becoming more definitive in design as well as greater in dimension. The last ten years have witnessed the implementation of large flotation cells ranging in size from 28 to 85 cubic meters. However, small plants still operate with cells as small as 3 cubic meters. These changes have increased the operational sensitivity of the flotation system to the chemical reagents used. No scientific study has been undertaken to develop the relationship between the size of equipment and the response of reagents in the flotation system. Hence, this article is an attempt to present several observations from past experience for the purpose of stimulating future research in an area previously ignored and perhaps subsequently not used to advantage. CONVENTIONAL FLOTATION Recent advances in conventional flotation cell design have been made in two areas: a. Flotation cells are becoming larger, and b. Basic design of flotation cells has changed. They can be classified into two types: deep cell and shallow draft cell. Size and design changes of flotation cells have a greater effect on frother performance than on the performance of other chemical reagents. Hence, the discussion in the following sections is restricted to the impact of frothers in the flotation system. Flotation Cell Size The advent of increasing flotation cell size has resulted in the increase of frother consumption. There appears to be a correlation between reduced horsepower (kilowatt) consumption and increased frother consumption per ton of ore processed in the circuit. This is more prevalent when the frother chemistry remains unchanged. Another reason for increased frother consumption is the increased distance between the rotating central shaft and the lip of the cell for larger flotation machines. Increasing this distance results in increased surface area corresponding to the larger cell size. Reduced power input per ton combined with the need for longer, stronger load support of the froth bubble as a result of increasing cell size has resulted in the development of a new series of stronger and more persistent frothers. By the same token, these newer, stronger frothers are difficult to control when applied to the smaller cells. This is true even at lower dosages because the break point between too much frother and not enough frother for optimum metallurgy is normally very narrow. This problem is often exaggerated when operators have established an optimum setting for pulp level and air for the particular frother in use. Generally, the tendency is to operate the flotation cells simply by adjusting frother dosage. The increase in frother strength to accommodate the large cells often results in the realization, after a retrofit, that the existing pump box or sump which is still handling the same volume of concentrate is no longer adequate in size. Flotation Cell Design The two most common flotation cell designs installed during the past ten years are the deep cell and the shallow draft cell. The deep cell provides for complete mixing of air and pulp at the bottom of the cell. The shallow draft cell does all the mixing/pumping in the bottom haft of the cell whereas the mixing of the air is performed further up the shaft and is brought into contact with the pulp midway up the vortex. There are many preferences for one or the other. It is the author's contention that both designs have very valid concepts and must simply be suited to the particular situation as required by the operator. The following comments are subjective observations which should be explored in greater detail. The deep cell mixes both air and pulp together at the bottom of the cell requiring a full depth ascending of the homogenized air to the surface to form the froth layer. This design, in combination with charged air, provides for good concentrate grade control under adverse conditions of selectivity, especially when faced with a fine grind and a slime-laden gangue. A selective frother can be applied without a competitive loss in recovery. One of the conditions for good selectivity in this environment is operating at medium pulp solids concentration in the range of twenty-six to thirty-four per cent (26-34%). When a higher solids concentration is used in combination with a coarse grind, it becomes difficult to maintain the same recovery with a selective frother due to higher pulp viscosity. This condition sometimes results in an unacceptable coalescence of the air bubbles as they ascend the full height of the deep cell. The increased pulp viscosity at higher solids concentration can also retard the flotation kinetics. The combination of decreased surface area of dispersed air and slower kinetics can result in reduced recovery causing coarse particles to continue to drop

Jan 1, 1989

By J. M. Smith, T. R. Horton, R. L. Blanchard, T. W. Fowler

INTRODUCTION Uranium mining operations release radioactive materials into both air and water and generate large quantities of solid wastes containing low levels of radioactive materials. Solid wastes produced by mining operations remain on the surface at many inactive mining sites in the Western United States. These mining effluents may present a potential health and environmental hazard. Therefore, Congress, in Section 114(c) of the Uranium Mill Tailings Radiation Control Act of 1978, instructed the Administrator of EPA to prepare a report identifying the location and potential health, safety, and environmental hazards of uranium mine wastes and to recommend a program to eliminate these hazards. Several facts and limitations helped shape the method and approach of the EPA study. Little information on uranium mines was available; measurement information that was available on uranium mine wastes was frequently influenced (biased) by nearby uranium mills; time and resources did not permit comprehensive field studies to provide additional data; and there are inherent variations between uranium mines and sites that complicate generic assessments of mine wastes. To accommodate these facts, the EPA developed conceptual models of uranium mines and made health and environmental projections from them. The models were based upon available data from the literature, supplemented with information from discussions with persons inside and outside the EPA, and by doing several short-term field studies in Texas, New Mexico, and Wyoming. When necessary, conservative (maximizing) assumptions were employed. This paper presents a brief account of a part of the EPA study dealing with the potential health and environmental effects caused by active surface and underground uranium mines. Airborne contaminants are emphasized, although solid and liquid effluents are also included. Due to the limited space, only the methods and parameters used and the results of the assessments will be presented here. Anyone interested in the source of the data used and the development of the parameters should refer to the EPA report (Blanchard et al., 1981). The occurrence and emissions of stable elements were included in the EPA report, however, due to space limitations and their apparent small impact, except for possibly at some specific mines, only radioactive sources will be included in this presentation. MODEL URANIUM MINES The model surface mine was located in the South Powder River Basin of Wyoming and the model underground mine was located in the Ambrosia Lake area of New Mexico. These are the prevalent type mines in those areas. The model mines were based on the average production parameters of the 63 open pit mines and the 256 underground mines that were operating in the United States in 1978 (Department of Energy, 1979) and on a report of an extensive study of open pit mines in Wyoming (Nielson et al., 1979). Information contained in environmental impact statements and in reports from federal and state agencies was also used. Parameters for the model mines are listed in Table 1. The surface mining scenario is that 7 pits are opened in the 17-year mine life with overburden from each successively mined pit used to backfill a previously completed pit, resulting in an equivalent of one pit of overburden (2.4 year production) stored on the surface. No backfilling is assumed at the underground mine. Overburden or waste rock, ore, and sub-ore are separated into separate piles that are either rectangular in shape with length twice the width or in the shape of a frustum of a regular cone. Both shapes have 45 degree sloping sides. To account for bulking, the volume of the material comprising the piles was considered to be 25% greater than the volume of material removed from the ground. It was assumed that dewatering was required at both mine sites. Wastewater discharge rates at the surface and underground mines were assumed to be 3.0 and 2.0 cu m per min, respectively. SOURCE TERMS The following radioactive contaminants at active uranium mines were assessed in the EPA report: 1. Radioactive particulates in a) wind suspended dust from waste rock (overburden) pile, sub-ore pile, ore stockpile, b) suspended dust from mining activities (rock breakage, loading and unloading ore and wastes), and c) vehicular dust, 2. Rn-222 emanation from waste rock (overburden) pile, sub-ore pile, ore stockpile, and mining activities, 3. Rn-222 emanation from mine surface areas, and 4. Radionuclides in wastewater discharged to land surface. Estimated average annual dust emissions (item 1 above) from the model mines are listed in Table 2. Emission factors and the assumptions used to estimate these dust emissions are described in detail in the EPA report. Radioactive source terms were computed for each of the sources; dust emissions were multiplied by the concentrations listed in Table 1

Jan 1, 1981

By W. S. Hannan

History Leaching of sulfide ore was first attempted by the Phelps Dodge Corporation about 1900 at Bisbee, Arizona, when Dr. James Douglas conducted experiments in an attempt to apply the leaching techniques then in use at Rio Tinto, Spain. Although this initital attempt was unsuccessful, research work was continued which culminated in successful leaching operations at Tyrone, New Mexico, in 1917 and at Bisbee beginning in 1920. Leaching has continued intermittently since that date, with material from the Lavender Pit being leached today. The Lavender Pit Leach Dump Stripping for the Lavender Pit was started in 1951, with mining operations on stream in 1954 and continuing to the present. All material removed from the pit has been classified as waste, leach material, milling ore, or direct smelting ore, depending upon grade and degree of oxidation. Lavender Pit leach material has been stockpiled in a single dump, called No. 7 Dump. This dump covers an area of approximately 135 acres and average about 300 feet in height. The dump has been deposited over a period of 20 years by truck haulage and contains several layers, each of which has a permeable boulder layer at the bottom and a compacted layer at the top. Leaching of Lavender Pit material was started in 1956 and has proceeded simultaneously with dump construction. Former leach pond areas and haulage surfaces have been ripped with D-9 tractors before deposition of succeeding layers of ore.

Jan 1, 1972

By Ernest Y. Baafi

The technique of linear kriging as a grade estimation method of mining blocks is reasonably established in the mining industry. Because of the nonbias condition imposed in kriging which requires that the kriging weights (?i) add up to one, negative weights are sometimes used in kriging. The use of negative weights is not justifiable from the practical standpoint. Also, negative block grades can be obtained in some extreme situations which makes no practical sense. To guarantee non-negative weights an additional constraint is required in the kriging system of equations. Unfortunately, the nonnegative weight constraint results in a set of non- linear system of equations. In this paper, a derivation of the resulting set of nonlinear equations is presented and a numerical algorithm is developed for the solution of the equations. Because the algorithm developed does not remove negative weights arbitrarily, optimal nonnegative weights are always guaranteed. Results obtained using the developed algorithm were compared with those using the standard linear kriging, in terms of computer time and the estimation results. Both the computer time and the kriging variance were found to be higher with nonnegative constraint, as expected. However, some systematic pattern was observed in the kriged values (i.e., over or under estimation) with or without the nonnegative constraint. The study also demonstrated when one should generally expect negative kriging weights in practice, and what practical impact the negative weights can have on the kriging estimates.

Jan 1, 1984

By W. W. Harvey

The work to be described was part of an exploratory research program directed toward identifying potentially useful but little known reactions in extractive metallurgy. In particular, the major copper sulfide minerals, as a group, have long been considered insoluble in nonoxidizing acids (e.g. Forward and Barren, 1J60; Parsons and Ingraham, 1970; Peters et al. 1972). This generalization was considered to include nonoxidizing conditions in warm aqueous MC1 solutions (Ingraham et al. 1972) and, indeed, was consistent with our own chemical and mineralogical experience. On the other hand, under hydrothermal conditions, copper and sulfide species are carried in solution and deposited as sulfides in mineralized veins. The question that poses itself is: what are the mildest conditions of temperature and fluid composition under which commercially important copper sulfide minerals can be solubilized in the absence of an oxidizing agent? A reference to H2S formation during oxygen-sulfuric acid pressure leaching (Warren, 1958) of chalcopyrite, when the oxygen partial pressure was allowed to fall, provided a clue. In the present study, experimentation with chalcopyrite concentrate in dilute hydrochloric acid, under re flux conditions (in Pyrex flasks) and with an inert purge gas, revealed substantial amounts of H2S production and iron solubilization, cite negligible copper solubilization, after 2 ½ -4 days. Under comparable conditions the extent of reaction with dilute sulfuric, acetic and phosphoric acids was markedly less.

Jan 1, 1976

By Paul Hosford, B. Thrall

There has been a surge of exploration activity and interest in mineral properties located in cold Northern climates, with gold and copper properties as the two most attractive targets. There are already a number of mining operations located in the Arctic, for example, Red Dog, Polaris, and Lupin, all located in the Canadian Arctic. The logistics of sustaining operations, the engineering design and construction methods for such inhospitable regions are now relatively well understood. However, at least three of the proposed new mines in the Yukon Territory, Canada are designed to be heap leach operations. Rescan Engineering has recently completed Feasibility studies for two heap leach operations in the Yukon, one of which, the Brewery Creek Mine, is now in operation. The purpose of this paper is to explore some of the engineering and operating concepts that have been developed to make heap leaching an appropriate process method for gold recovery in a cold climate. Many of the readers will be familiar with the general and specific details of heap leaching, and it is not the intention of the authors to deal with any of the established "do's" and "don'ts" that have been addressed in numerous previous papers, but only with those pertaining to cold weather operations. This paper also examines these design concepts in specific relation to the results of the first winter's operation at Brewery Creek Mine, the first "far" North heap leach operation.

Jan 1, 1997

By Z. T. Bieniawski

Results of in situ investigations are discussed providing comparisons of rock mass deformability measurements by three methods: 'petite sismique' technique, the Goodman borehole jack and flat jacks. The tests were performed in exploratory tunnels constructed for site characterization purposes of an underground hydro-electric scheme. Due to the large number of data (41 petite sismique measurements, 45 Goodman jack tests and 19 flat jack measurements), it was possible to evaluate thoroughly the testing techniques employed. The purpose of this paper is to indicate the problems encountered in the interpretation of the results and to offer suggestions for selecting in situ tests. INTRODUCTION In situ tests are expensive and time consuming but they are an important aid in tunneling applications since they can provide deformability data of rock masses in contrast to rock material characteristics obtainable from small laboratory specimens. Some in situ tests are relatively easy to perform but their reliability may be in question unless a comparison of the data obtained by several different test methods is possible. Although many in situ techniques are available for determination of rock mass deformability (1), those more convenient to use are the Goodman jack, flat jacks and 'petite sismique. These methods are considered convenient because the Goodman jack is a borehole device; flat jacks can provide not only deformability data but also information about the stress field around excavations; while 'petite sismique' is based on seismic refraction surveys. Accordingly, these three methods have been selected for comparison.

Jan 1, 1979

By Byron Belew, Juris R. Harlamovs, D. B. Dreisinger, John L. Watson, G. A. Kordosky, Thomas O. O’Keefe

Unwanted impurities, present in raw materials, are often coextracted along with desired constituents during leaching. Due to its abundance and ease of extraction, the presence of ferric iron in leach solutions, solvent extraction circuits, and subsequently in electrowinning circuits, has always been a problem. In the case of zinc hydrometallurgical processing, iron is frequently an impurity in leach solutions. Purification of such solutions by solvent extraction, using di (2-ethylhexyl) phosphoric acid (D2EHPA) in a kerosene diluent, results in ferric ions being readily extracted along with zinc into the organic phase. Ferric iron has proven difficult to strip from the organic phase, but it is known that ferrous iron may be easily stripped from D2EHPA, and hence a process to effect reduction of iron present in the organic would have great potential. This research investigated a process to directly reduce ferric to ferrous iron in the organic phase, and to strip the ferrous iron using a dilute acid solution. The process utilized metallic zinc as the reductant and a low concentration sulfuric acid stripping solution. The results of the laboratory tests illustrated that iron could be readily removed from the organic phase by simultaneous direct reduction and acid stripping. The reductive stripping process was subsequently utilized in industry at the laboratory and pilot plant levels. The results of the industrial tests were encouraging and demonstrated the potential of this novel concept.

Jan 1, 1993

By Maria S. Adusumilli, A. Bhaskara Rao

ABSTRACT The presence of gold in lateritic crusts over mafic and ultramafic volcanogenic complexes in / and greenstohe belts, has called the attention in several parts of Brazil and also in Australia. Recent studies reveal the variety of structures and mineralogy of such crusts, and their importance in correlation with host(bed) rock types, as exploration guides, specially for greenstone belt metallogeny. The expected nature of growth of plates and/or nuggets in such crusts is under detailed study, and is considered as possibly due to dissolution and reprecipitation phenomenae of disseminated-fine-low grade ore in the host rock, in this investigation, laterite geochemistry, though incipiently discussed todate, is of fundamental importance and significance. Lateritic gold can be classified texturally as: dusty and diffused, fine, thin platy, taDular, and nuggets. Such gold is pure, and is often restricted to nodular (pisoliths, ooliths, concretions) forms or the matrix, that constitute the crusts or lateritised gravel beds which have an essentially goetnitic nature. Further this gold is recovered directly by mining the crusts or the adjoining placers with the metal derived from them. The thickness and extension of laterite crust zones is extremely variable depending oh the paleoclimatic conditions, and is conditioned to the favourable lithology and its structural expression. However, the "growth" of gold in these laterites will be conductive , without doubt, to several small scale mining activities, specially in the developing countries and regions limited to latitudes 309N and 309S.

Jan 1, 1987

By Cengiz Karagüzel, Ömer Canıeren

The mobility of gold and silver metals from refractory ores is very low due to free metal ions in their bodies that bind the transportable and free metal molecules or they are encapsulated in minerals such as sulfides, arsenosulfides, carbonaceous, and clayey materials which make it difficult to extract them. As known from the literature, the enrichment of oxide refractory gold, silver ores, Ag, and Au-rich flotation concentrated from sulfur refractory ores can be leached in the chloride-hypochlorite solutions. In this work, the leaching approach was studied to extract silver ions from refractory ores in the presence of hypochlorite media under high pressure. The mineralogical and scanning electron microscopy analysis of the sample indicated that the silver was encapsulated in barite and quartz minerals in fine sizes. In order to obtain the optimum leaching conditions for the silver ore, the effects of several parameters, namely, temperature, leaching time, NaCl and NaOCl concentration, solid ratio, and particle size on the leaching recovery were investigated in detail. The leaching experiments showed that the highest silver recovery of 62.35% was obtained under the following conditions: temperature of 150 °C, oxygen pressure of 1800 kPa, leaching time of 2 h, concentrations of 2 mol/dm3 and 0.4 mol/dm3 for NaCl and NaOCl, respectively, the solid ratio of 30%, and particle size (d80) 15 μm. The results obtained from this study indicated that the enrichment of refractory silver ores to extract the silver encapsulated in barite and quartz minerals could be possible in the chlorite-hypochlorite under high pressure, a relatively shorter leaching time compared to conventional cyanide leaching.

Jan 28, 2021