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By Steve Cole

"The economic and strategic importance of PGE bearing deposits is fuelling renewed exploration and development efforts in the precious metals mining sector.This paper presents a review of the various techniques used in the processing of PGE bearing ores, from primary benificiation to processing routes for the different flowsheets that have been developed amongst current producers. Recent innovations such as preconcentration of ore using gravity separation, recent trends in flotation circuit design and the development of hydrometallurgical routes for the direct treatment of flotation concentrates without the need for smelting will be discussed. Data from recent projects will be presented to illustrate common characteristics of PGE ores and concentrates.INTRODUCTIONThe six platinum-group elements (PGE): ruthenium, rhodium, palladium, osmium, iridium and platinum, together with gold and silver, are considered to be ""precious"" metals. The recent spike in platinum and palladium prices coupled with the projected increase in demand for the metals has fuelled a spate of exploration and new projects in the PGE sector. Historically primary producers of PGE metals have controlled the flow of information relating to the mining and processing of these ores into the public domain. Very little published information relating to individual operations and processes is available in comparison to the abundant literature relating to other metals.The purpose of this paper is to provide a general overview of the issues relating to the processing of PGE ores. Each topic discussed could easily be the topic of a paper or book and so in many cases only the surface of the topic will be scratched. An attempt has therefore been made to provide as many references as possible.This paper discusses the classification and location of PGE ores, the mineralogical factors that affect recovery of PGE's, methods currently used for recovering PGE's and process trends in PGE metallurgy."

Jan 1, 2002

With the employment of a system of ore pillars to support the hanging wall during the mining operations at the North Mine, it was found that, when the stopes were worked out and the pillars remained intact, the main footwall drives developed an undulating longitudinal section due to the shrinkage of the filling in the worked out stopes.These depressions collected water, which made the footwall drives very wet under foot, and everitually worked its way into the pillar workings and the levels below.Water was also finding its way into the shafts because of the silting-up of the level drainage systems and, as both the shafts are down-cast, this tended to increase the humidity of the air entering the mine.To overcome these troubles it was decided to try a system of siphons as a method of subsidiary drainage.SYSTEM AT PRESENT IN USE.All drainage water is collected as near to its source as possible, run into a sump in an adjacent low lying part of the level, and continuously removed from these sumps by a siphon system.Each level has its own system. Water is collected from the sumps by a 1-in. pipe and delivered to a pipe running the length of the level along the back of the footwall drive. At the shaft the pipe is increased to...

Jan 1, 1938

By Norman F. Schulz

The technical excellence of separation achieved in a mineral concentration process, or any other process where two constituents of any kind are physically separated from each other, is expressed uniquely and quantitatively by the separation efficiency (E,): E. = (R - R,), where R is the percentage of the valuable constituent and R, the percentage of the waste appearing in the concentrate. The separation efficiency meets all the usual requirements of a true efficiency and is independent of the nature of the process used. It can be visualized as the percentage of the feed material that, in effect, actually undergoes complete separation while the rest of the feed is distributed randomly and unchanged into the two separation products. The problem of quantitatively assessing the results of a mineral concentration process recurs whenever one tries to decide which of two similar test results is more desirable. Although economic factors may greatly influence the eventual choice of process, they usually have very little to do with the technical excellence of a separation and are, therefore, not a consideration at this point. Concentrate grade and recovery, used simultaneously are the most widely accepted measures of performance, and if both the grade and the recovery are greater for one case than for another, the choice is simple. However, when the results of closely competitive operations are compared, it often occurs that one set of test results shows a higher grade and a lower recovery than another and the choice is no longer obvious. Laboratory investigations of a mineral concentration process applied to a given feed material can often be designed to yield successive paired values of concentrate grade and recovery over the practical range of the course of a separation process.' These results may be plotted as some combination of the cumulative values of grade, weight recovery, or unit recovery in the concentrate. The resulting correlations for two slightly different sets of conditions can then be compared qualitatively, or even quantitatively in particular cases.2 Such comparisons are not applicable, however, to individual concentration test results. The reduction of a separation test result to a single efficiency figure, preferably as a percentage of ideality, would afford the simplest and most precise means for comparing the results of competitive tests. The effective proportion of feed material that undergoes complete separation in any real concentration process, hereafter referred to as "separation efficiency," is such a figure. It is essily calculated from the usual feed and product data and can even be evaluated, in many cases, from data in publications describing various concentration processes. Mineral Concentration Processes and Their Evaluation Mineral Concentration Processes: The overall concentration process may be subdivided into a liberation process which severs the bonds between unlike mineral species, and a separation process, which sorts the resulting liberated particles into two products—a concentrate rich in mineral A and a tailing lean in mineral A. A perfect mineral concentration process is represented by the idealized scheme shown in Fig. 1 in which liberation and separation are not differentiated. The feed mixture consists of two mineral types which are completely liberated early in the process and then sorted by type. All the mineral particles of type A which respond uniquely to the selection forces acting in the ideal separator, report to the concentrate, while all the mineral particles of type non-A report to the tail-

Jan 1, 1971

By Robert B. Bittner

Large diameter piles (2 to 3 m diameter, both driven and drilled) are being used more frequently by designers of bridge foundations. The use of these large piles allows the pile cap to be positioned off the bottom but usually still below water. Conventional cofferdams as shown in Figure 1 below are not suited to this type of pile cap. However, float-in cofferdams offer a viable alternative. This paper describes the float-in cofferdam system, its initial development for use on the Bath-Woolwich Bridge in Maine and its most recent application on the new Carquinez Bridge in California.

Jan 1, 2013

By Seal. Thom

Large heap leach operations around the world contain significant under-leached metal values that present recovery challenges to operators. The first challenge is to find the location of the highest inventory. High resolution resistivity ? has proven sensitive enough to detect such zones in the interior of heap leach pads to target treatment or re-leaching. Newmont Mining Corp. has adapted the high pressure fracturing technology, used in oil wells to stimulate select zones in heap leach pads to add reagents in 3-D for re-leaching, changing zonal chemistry for improved recovery and heap closure. High Resolution Resistivity ? is able to subsequently monitor injected solution flow and determine directional permeability during Hydro-Jex stimulation. The combination of these technologies has shown strong economic success, for selecting targets in the heaps and monitoring solution application. These technologies are now available to the mining industry for adoption.

Jan 1, 2008

By H. P. Jacobi

"INTRODUCTIONPine Point Mines, operated by Cominco Ltd., is located approximately 65 miles east of Hay River, N.W.T. on the south shore of Great Slave Lake.The milling operation started in the fall of 1965 with a capacity of 5,000 tpd. An addition was put into operation late in 1968 to bring the nominal capacity up to 8,000 tpd. The present milling rate is 11,000 tpd.With a feed grade of approximately 2.9% Pb and 6.0% Zn the annual concentrate production amounts to roughly 130,000 SDT of lead concentrate and 370,000 SDT of zinc concentrate for a total of 1/2 million tons of product.This tonnage is moved by railroad. Pine Point is served by the Great Slave Lake Division (G.S.L.) of the Canadian National Railways which connects to the Northern Alberta Railway System (N.A.R.) at Roma Junction, at Peace River, Alberta. The system in turn connects to the CNR and the CPR in Edmonton, Alberta. From here the concentrates move inland chiefly to Trail, B.C., with some shipment to other Canadian and U.S. smelters, and to the port facilities of Pacific Coast Terminals at New Westminster, B.C. for offshore export.This paper will focus on the problems associated with the movement of lead and zinc concentrates from Pine Point to port facilities on the west coast."

Jan 1, 1975

The placer mining industry of the West Coast region of the South Island is based largely around the production of alluvial gold, although other valuable minerals are also present. Since the first major discoveries in 1864-65, reported gold production has totalled around 264 t (8.5 Moz). During the past 10 years, annual production has totalled around 1460 kg (47,000 oz). It is anticipated that gold production over the next three years will continue at a level of 1400-1800 kg.   Annual gold production peaked during 1866 at approximately 17,170 kg (550,000 oz) and has not exceeded 100,000 oz since 1885. This reflects the ease with which gold was won from shallow, rich deposits that were present at many localities. Subsequently, there has been a general decline in the availability of this type of deposit, a trend that will continue in the future. This paper outlines the present state of placer mining in the West Coast region, gives some historical background to the industry and briefly predicts future trends.

Jan 1, 1997

By Rao P. K, Sahoo P. K

The main constituents of ferromanganese slags are silicates of Mn, Ca and AI. Factors affecting the recovery of manganese from a low manganese content ferromanganese slag (about 16 per cent MnO) during acid leaching were studied. These included time, temperature, agitation, sulphuric acid addition, solid:liquid ratio and particle size. Agitation rates greater than 300 rev/min. leaching periods greater than 30 min and temperatures above ambient were found to have no effect on Mn recovery. Dissolution of Mn was found to be highly dependent on the amount of acid added.About 90 per cent of the Mn was recovered when 5·5 times the stoichiometric amount of acid was used. About 28 per cent of the Mn was recovered when the stoichiometric quantity of acid (with respect to Mn alone) was used. However, the recovery increased to about 75 per cent when the total stoichiometric requirement of acid with respect to Ca, Al, Mg and Fe as well as Mn was used for leaching. The extent of Mn dissolution also depended on the particle size of the slag. The percentage of Mn recovery (using 5·5 times the stoichiometric amount of acid) increased from about 60 per cent to about 92 per cent when particle size decreased from -36 +52 to -150 +200 mesh (B.S.S.). When the results of this study were compared with those in a previous study involv,ing the leaching of a high manganese content slag (containing about 40 per cent MnO) results were similar except that there was a gr-eater consumption of acid in the present case (i.e. low manganese content slag).

Jan 1, 1979

By Thomas Novak, Sampurna Arya

Controlling dust generation and keeping it below permissible limits to meet federal dust standards at the working face of a roomand-pillar coal mine is a challenge for a mine operator. With the recent changes in federal dust regulations requiring lower worker exposure, maintaining compliance has become increasingly difficult. The current most effective practice of dust control at a continuous miner face in an underground mine is the use of a flooded-bed scrubber. A study carried out by the National Institute for Occupational Safety and Health (NIOSH) indicated that a flooded-bed scrubber could achieve cleaning efficiencies between 58 and 90%. But, the operation of such a system is maintenance intensive. The flooded-bed scrubber screen becomes clogged with dust particles and requires frequent cleaning to maintain performance. However, the dust control issue is not solely a mining industry problem. Other industries face similar issues. The University of Kentucky collaborated with Toyota Motor Manufacturing on the development of a novel wet scrubber, called the vortecone scrubber, for capturing oversprayed paint particles in automotive paint booths. The vortecone scrubber achieved a cleaning efficiency of 99% and required minimal maintenance. This study aims to assess the ability of this vortecone scrubber to capture respirable dust in an underground coal mine. The paper presents the results of computational fluid dynamics (CFD) modeling of this vortecone scrubber. It discusses the effects of air quantity and dust particle size on the performance of the scrubber.

By Rüdiger Wunsch

The new high speed Vienna - St.Pölten line and the Lainzer tunnel are to be connected to the existing Western railway line in Vienna. This involves building an underground cut and cover branching tunnel which will be connected to the Western railway line by ramps. A key structure is the Herzmansky Street tunnel, which has a length of some 820m. It is to be constructed as a cut and cover tunnel using the top down method. The walls of the tunnel are designed as a water resistant structure consisting of double walls, an outer wall of bored piles and an inner reinforced concrete wall. The bored pile wall will be used as a temporary retaining wall during the construction phase and will also be a structural part of the per­manent tunnel wall. The retaining walls of the open pit, which will have a depth of up to 14m and a width of 12m, were investigated as a non-contiguous bored pile wall and overlapped bored pile wall. Preventing ground water from penetrating into the pit is of paramount importance. The high ground wa­ter level is the determining factor for the dimensioning of the bored pile wall and results in the require­ment for a water resistant concrete structure. It has been necessary to adopt adequate measures, therefore, in order to reduce the hydrostatic water pressure during the construction phase, assuming economical pile dimensions.

Jan 1, 2000

By G. I. McPhail

The authors have conducted several investigations into the prediction of the beach profile of high-density tailings and have found that there are a number of important considerations that need to be taken into account over and above the reliability of the beach profile prediction methodology. These considerations relate to the chemistry of the slurry and specifically the influence of pH and flocculation, both of which have a significant influence on the rheology of the slurry at the discharge point. A number of examples are discussed and a potential testing protocol that will assist in assessing the influence of chemistry ahead of the beach profile predictions is presented. A methodology for assimilating the results of the testing into the likely field situation is discussed, and suggestions are made on application of the results in beach profile prediction. The suspected influence of residual flocculation on the stratification

Jan 1, 2012

By N. Sano

The chromium oxide in stainless steel slags usually exists as Cr2O3. However, there is a possibility that it is oxidized to hazardous CrO3 after disposal or during use. Therefore, some technical developments to reduce chromium oxide by using a smelting reduction technique have been reported recently. This paper describes the fundamentals of the technique which have been applied to ferrochromium production and reviews the recent results in literature.

Jan 1, 2004

By deur R. R. Mears

The expected curtailment of activities within the gold-mining industry by the closure of the more marginal mines, and the shortening of the life of the industry as a whole as a result of rising costs and/or an uncertain gold price, are factors that are bound to harm the South African economy in a number of ways. The industry has no option but to regard the gold price, which is determined on the world market, as a given quantity. The remaining ore reserves al tend to be of a lower grade and/or to be found at greater depth. Therefore tighter control of working costs would seem to be almost the only option open to the industry. Working costs can be divided into two broad categories: costs emanating from the environmental characteristics of the orebody, and costs emanating from the general inflationary environment in which the industry has to purchase its inputs. Working costs stemming from the physical environment also fall into two categories: those relating to the grade of the orebody, and those resulting from working at increasing depth and temperature, together with the need to pump increasing quantities of water. The industry is both capital- and labour-intensive in absolute terms, since it requires large inputs of labour as well as of capital. It is limited in its scope to adjust its production function by becoming more labour- or more capital intensive, and is dependent on a large labour force, which cannot easily be replaced by capital inputs. On the other hand, large amounts of capital are required for investment in shafts, development, machinery, and equipment. Since the early seventies, the working costs of the South African gold-mining industry have increased at an annual rate that was significantly higher than the general rate of inflation. This rapid rise in working costs has caused some mines to find themselves in a difficult profit position. At times, the gold-mining industry has even been compelled to restructure its cost/revenue relationship rather fundamentally by revising the grade of ore worked sharply upwards or downwards, as the case may be. The rise in the gold price since 1972 has made it possible for the industry to introduce significant changes in its Black wage policy. The main aim was to increase the permanency of the Black labour force and to encourage local Black workers to join the industry. At the same time, significant increases were also experienced in the cost of machinery and stores, as well as of capital, mainly as a result of inflation. In the study on which this paper is based, the author attempted to measure the influence of environmental factors on working costs, using the physical consumption of electricity as a criterion for measurement. He found that the usefulness of this criterion in relation to factors such as depth, temperature, and water quantities pumped was invariably more than neutralized by the effects of greater productivity and economies of scale. The effects of these environmental factors on the working costs of the industry could therefore not be determined to any degree of accuracy. It seems that good progress is being made with the development of capital-intensive forms of technology aimed at coping with rapidly rising working costs and other effects of increasing depth. Techniques are also being developed to improve rock-cutting procedures and to eliminate the use of explosives in order to introduce a continuous mining process. On the whole, the gold-mining industry seems to be more sensitive to general inflationary pressures than to the cost effects that are associated with depth, temperature, and water quantities pumped. Greater control is possible over physical environmental factors such as the grade of ore mined, temperature, rate and nature of development of a mine, and adaptation of the daily working cycle to a continuous mining process. Wolfe and others hold the view that the industry will lose control over labour costs as a result of the 'sharp increase in wages and the growth of Black trade unions. However, the use of more capital-intensive technology will allow the industry to reduce its labour force, and a smaller, better-trained, and more permanent labour force will enable the industry to cope with the increasingly difficult physical environment and higher wages.

Jan 1, 1987

By Arthur M. Langer

Since the promulgation of the first Federal Asbestos Standard by the Occupational Safety and Health Administration in 1972, other federal agencies have modified the standard to better carry on their own unique missions. The instruments used to identify and measure asbestos, the sampling protocol, and the criteria used to define asbestos, have been modified to some degree. The Mine Safety and Health Administration regulates and controls asbestos dust in the mining and mineral commodity industries. However, crushed stone and processed ores contain mineral fragments that are frequently difficult to distinguish from asbestos. Mineral nomenclature, instruments for particle analysis, and sampling strategy must be accommodated to some degree to make asbestos control workable and meaningful. Precedent in other agencies has made consideration of these changes possible. Newly identified amphibole asbestos minerals have further complicated the agency?s regulatory charge. Changes in its Asbestos Standard are now being considered. Crushed taconite ore in the Eastern Mesabi highlights many of these issues. ©2008 Published by Elsevier Inc. Keywords: Amosite; Grunerite asbestos; Crocidolite; Riebeckite asbestos; Tremolite asbestos; Actinolite asbestos; Winchite asbestos; Richterite asbestos; Arfvedsonite asbestos; Fluor-edenite asbestos; Asbestos defects; Cleavage fragment; Mineral habit; Mesabi amphiboles

Jan 1, 2008

By Woodcock J. T

When an aqueous suspension of pyrite and a manganese ore reacts with oxygen, manganese is dissolved by the ferrous sulphate and sulphuric acid produced in situ by oxidation of pyrite. In the absence of added acid, the optimum conditions are: a molecularratio of Mn02/FeS2 of about 1·0, a temperature of about 1400 C., and an oxygen pressure between 25 and 90 lb./sq. in. Under these conditions, complete extraction of manganese is obtained in six hours leaching. An addition of sulphuric acid enables an Mn02/FeS2 ratio of 2·0 to be used successfully.Manganous sulphate decreases in solubility with increase in temperature above 1000 C., but even when the solubility limit is reached during leaching, reaction to form MnS04 still proceeds, and on cooling the MnS04 redissolves.INTRODUCTIONMany manganese deposits contain secondary manganese oxide minerals which are difficult to concentrate. Considerable attention has therefore been given to their hydrometallurgical treatment. In most manganese oxide minerals, the manganese is quadrivalent and must be reduced to the divalent state for dissolution. Reduction and dissolution can be accomplished in a number of ways, one method being to use aqueous ferrous sulphate solutions.A possible source of ferrous sulphate is pyrite as it is known that ferrous ion is an intermediate product in the aqueous oxidation of pyrite. It was therefore thought possible, that by heating an aqueous suspension of manganese oxide and pyrite with oxygen, the ferrous sulphate and sulphuric acid formed from pyrite would provide the necessary leaching reagents for manganese in situ. There are certain advantages in producing ferrous ion in situ...

Jan 1, 1958

By Gerald T. Homce, James C. Cawley

Electrical arcing injuries are the most common form of electrical injury in mining. MSHA data show that 381 "non-contact electric arc burn" injuries occurred in the mining industry from 1996 - 2005. A more detailed study of data from 1990-2001 showed that five types of apparatus - circuit breakers, conductors, nonpowered hand tools, electrical meters and test leads, and connectors and plugs were involved in two-thirds of the lost workday injuries. Many of the protections from electrical arcing injury available to workers in other industries stem from the application of National Fire Protection Association Standard 70E - Standard for Electrical Safety in the Workplace (NFPA 70E). NFPA 70E is a generally accepted industry guideline document, but excludes underground mining from its scope. Many of its practices and procedures, however, can be applied in the mining workplace to supplement the existing electrical safety requirements of Title 30 Code of Federal Regulations. This paper describes how NFPA 70E can be applied to mining workplaces after an analysis of the mine power system electrical arcing hazards has been completed (and Hazard/Risk Categories determined) using the tables from NFPA 70E or the calculation method as described in IEEE 1584. NFPA 70E can then be used to determine the levels and types of personal protective equipment (PPE) needed to protect miners against burn injuries. Appropriate lockout and tagout (LOTO) procedures for use in the mining industry are also described. Suggestions for better selection of electrically-rated hand tools are included. Recommendations for electrical meter selection and use based on transient withstand capability are also discussed.

The concentration of Mt. Morgan sulphide ore has always been attended by serious loss of gold. Until recently, with a simple selective flotation flowsheet, even the lowest grade of ore milled-about 2 dwt. per long ton-thislloss was 0·9 dwt., whilst for average to high-grade ore, 3·5 to 5 dwt., the figure was 1·5 to 2·0 dwt. Early in 1937 it was decided to equip the research laboratory for an investigation into the condition and location of the escaping gold and the physical and mineralogical barriers preventing its recovery.With the completion of: this equipment in July 1937, trial and error testing, which had been explored to exhaustion, was abandoned and a five months' research programme mapped out and commenced. Based on the knowledge so gained, a flowsheet combining primary desliming, bulk flotation of the natural slime, selective flotation of the deslimed fraction, and cyanidation of the selective flotation tailing was drafted and tried with encouraging results in a 70-lb. per hour pilot plant. This flowsheet, modified to meet current circumstances, is being dove-tailed into the existing 1,500-ton sulphide ore concentrator, though the cyanidation treatment is actually carried out in the oxidized ore mill, to which the flotation tails are pumped.RESEARCH EQUIPMENT(1) Five tube elutriator (Fig. 1) t built in the company's workshop.

Jan 1, 1939

By Mills R, Shannon W. T, Finch J

As an extension to investigations into methods of upgrading the Iitania content of ilmenite concentrate from beach sands using 100 g/day, a small pilot plant treating 23 kg/ day of ilmenite was used.In the pilot plant the iron oxide in the ilmenite was reduced to iron at 1200°C using carbon (coal/char) without melting the material, using a tunnel kiln heating cycle, the charges contained in saggars, The iron was then separated from the titania fraction by ball milling, wet magnetic separation, and screening. The resulting magnetic product was high in iron (about 80 per cent Fe) and the non-magnetic product high intitania (about 70 per cent TiO2).From this trial sufficient data were obtained to enable the preparation of a flow sheet for a plant producing 62,300 tonnes/year total products, and an estimate of costs to be made. This has shown that for an estimatedcapital expenditure of nearly $6,000,000 and annual production costs of about $4,000,000, the cost per tonne of high titania product (76 per cent TiO2) would be nearly $70.

Jan 1, 1977

By S. Parada, P. T. Jones, M. Guo, S. Smets

The corrosion behaviour of a high-quality rebonded magnesia-chromite refractory in contact with a vacuum oxygen decarburization (VOD) slag with high Al2O3content (15?20 mass per cent) is investigated under VOD process conditions by rotating finger tests in a vacuum induction furnace. The refractory degradation mechanisms are studied by the characterization of corroded microstructures using scanning electron microscopy, equipped with energy dispersive spectroscopy. The influence on the refractory wear of the process temperature and the Al2O3content in the slag is discussed. Keywords: corrosion, magnesia-chromite refractory, VOD slags, vacuum induction furnace

Jan 1, 2004

By Pamplinm J. W.

"INTRODUCTION For the past six years the United States Bureau of Mines in cooperation with the University of Alabama, Tuscaloosa, Ala., the Missouri School of Mines and Metallurgy, Rolla, Mo., and certain operating companies of the Florida land-pebble phosphate district have conducted an investigation to determine the amenability of phosphatic sands to concentration. 4 This paper discusses briefly an investigation of the concentration of fine phosphatic sanads of the land-pebble district of Florida. The process used - namely, the tabling of material previously given a selective oiling treatment - is new in its application to the phosphate industry, although old in principle.The feed to the standard phospidate washery is a hydraulically wined mixture of phosphate pebbles, quartz sand, and clay in which the quartz sand grains are segregated in the fine sizes. Me standard method of concentrating (washing) consists of wet-screening. at about 20 mesh to recover the plus 20-mesh material. 5 Lately some of the washeries have introduced hydraulic classifiers for sizing of the screen undersize to recover the coarser spigot products, but even this additional treatment allows much of the fine phosphate to be wasted."

Dec 1, 1932