Search Documents

The Sydney Coalfield, Nova Scotia contains extensive undersea longwall mining operations which are expected to continue well into the 21st Century. A long term research program has been initiated by the Cape Breton Coal Research Laboratory to investigate the full spectrum of problems associated with undersea mining and develop specific guidelines for use by the mine operator. At present, research is centred on the detailed evaluation of existing undersea sub- sidence guidelines, the identification and development of instrumentation capable of moni- toring seafloor subsidence profiles and the effective sealing of offshore exploration boreholes. In addition, work is also planned to examine in situ strata permeability changes occurring between the mining horizon and seabed and the feasibility of using artificial islands.

Jan 1, 1986

By Frederick B. Brien

DEVELOPMENT of uranium extraction techniques over the past ten years has been spectacular. Rapid progress has been made by applying chemical, physical, and engineering fundamentals for the purpose of devising simpler processes, developing continuous procedures, reducing reagent consumption, improving final products, simplifying analytical methods, and treating more refractory ores and ores of lower grade. Since the source of many valuable metals in future will be of necessity from lower grade ores of a complex nature, and since hydrometallurgical processes are often the only means of treating such materials, these recently developed techniques will be of increasing importance. Leaching: A great deal has been published about the important role of ion exchange resins in purifying and concentrating leach solutions. Recent applications indicate that the resins will be very useful in simplifying leaching procedures of ores containing two or more valuable metals.1 The valuable metals may be differentially adsorbed by careful control of the oxidation-reduction potential of the pregnant liquors. For example, treatment of vanadium- uranium ores at present may involve salt roasting, water leaching, acid leaching, carbonate leaching, and intermediate fusions. Use of a single acid leach may now be possible, followed by a two-stage ion exchange technique, comprising initial adsorption of hexavalent uranium and rejection of tetravalent vanadium. The solution is then oxidized, producing pentavalent vanadium, which is later readily adsorbed and stripped into a separate eluate.

Jan 9, 1957

By J. E. Lawver, J. D. Raulerson, Charles C. Cook

The agriculture industry has made great strides during the past decade to increase agriculture yields through increased use of fertilizers. Increased use of fertilizers may prevent, or at least delay, mass starvation due to the alarming increase in world population. Phosphate was added to soil as a plant nutrient in the form of calcined bones at least 2000 years ago (Anon., 1964), and man has used phosphate minerals as a source of fertilization in one form or another for at least 100 years. During 1977 the world produced about 116 Mt of phosphate rock, with about 86% used for fertilizers and another 4% for animal feed supplements. More than three-fourths of the total production comes from the United States, Morocco, and the Soviet Union. From a mineral beneficiation point of view, the major sources of phosphate rock and the methods of beneficiation can be classified as follows: marine deposits not containing appreciable carbonate minerals, marine deposits requiring a francolite carbonate mineral separation, igneous deposits not containing appreciable carbonate minerals, and igneous deposits requiring apatite carbonate mineral separation. [ ] Guano, mostly from Chile and Peru, accounts for 0.1% of the total world production, and the calcium phosphates from Ocean, Nauru, and Christmas Islands and the aluminum and iron phosphates from Brazil and Aruba account for less than 4% of the world production and are thus not considered in this classification (Lawver, et al.). At present, marine phosphorite deposits account for about 75% of the world's production; the igneous deposits account for 20%. The igneous deposits low in carbonate minerals are easily concentrated by crushing, grinding, and apatite flotation. The most important igneous deposits are those of the Kola Peninsula, USSR (Woodrooffe, 1972). The igneous deposits high in carbonate materials are of corn appreciably more difficult to beneficiate, but they have been concentrated by froth flotation for a number of years. An interesting but rather complicated flowsheet of this type is at Phalabonva, in the Republic of South Africa (Lovell, 1976). The Phalaborwa deposit is an igneous complex of pyroxenite with a central core of carbonatite surrounded by a serpentine- magnetite-apatite rock called phoscorite. The phoscorite containing about 10% P2O5, 35% magnetite, and 35% calcium magnesium carbonate is currently being processed. The process involves comminuting the material for fiberation and subjecting it to a copper float using a potassium amyl xanthate as collector and triethoxybutane as a frother followed by a magnetic separation of the tailings to produce a feed for phosphate flotation. This process produces a phosphate concentrate containing greater than 36% P2O5 at a P2O5 recovery ranging from 75 to 80%. Considerable success has been claimed for recovering apatite from carbonate-bearing ores at the Jacupiranga Mine of Serrana S/A (Silva and Andery, 1972). The carbonatite currently being mined contains an average of only 5% P205 and is concentrated using a unique flotation process (Andery, 1968) to yield 96% P205 concentrates. The ore contains about 12% apatite, 5% magnetite, 80% calcite plus dolomite, and minor amounts of phlogopite, olivine, zircon, ilmenite, and pyrochlore. Feed preparation consists of crushing to -31.75 mm (-1 M in.), rod milling in closed circuit with hydrocyclones to about 92% (-50 mesh), and two-stage cyclone desliming of the -50 mesh sands at 20 m. Weight recovery in the deslimed feed is normally 85 to 88% and the corresponding P2O5 recovery is usually about 90%. The deslimed feed is conditioned at 60 to 70% solids for 15 min at pH = 8-10 with 0.6 kg/t of causticized starch for iron oxide and calcite-dolomite depression. The conditioned slurry is diluted to 20 to 30% solids, about 0.2 kg/t of fatty acid or soap collector is added to the conditioner discharge, and the reagentized ore is subjected to rougher-scavenger flotation with additional fatty acid added to the scavenger float. The scavenger concentrate is returned to rougher circuit distributor, and the rougher concentrate froth is subjected to two stages of cleaner flotation to yield a final apatite concentrate analyzing 36 to 38% P205. Flotation recovery of P205 is, in general, above 90% when treating fresh carbonatite. The high-carbonate flotation tails normally analyze 1 % P2O5 or less and are suitable for portland cement production. The marine deposits. Types 1 and 2 of central Florida are representative of enormous reserves of phosphate rock that will undoubtedly account for much of the world's production in the near future. Until very recently the sedimentary deposits high in carbonate minerals (Type 2) have not been considered reserves due to the difficulty in making a francolite-carbonate separation. Although no commercial plant has yet been built to beneficiate Type 2 ore, laboratory and pilot plant data indicate the process is viable. If so, the reserves of Florida and similar deposits throughout the world will be substantially increased. A discussion of the beneficiation of these two types of sedimentary deposits and the relation of the resulting concentrates to the fertilizer industry of the United States is the subject of this paper.

Jan 1, 1981

By R. W. Baker

"The development of uppers drilling at Brunswick Mining is reviewed. This paper describes design criteria for drilling patterns and the research required for selection of a suitable drill jumbo. The success of the uppers drilling program based on operating procedures is highlighted. IntroductionBrunswick Mining and Smelting Corporation Ltd., operates a 10 500 tonnes per day lead /zinc/copper I silver underground mine some seventeen miles southwest of Bathurst, New Brunswick .The No. 12 orebody, first discovered in 1952 by a group headed by M.J. Boylen, had extensive development, including two shallow shafts (500 ft), and test work carried out on it by St. Joe Lead between 1953-1957. However due to complex metallurgy and depressed metal prices, the operation was suspended.In 1961, a group comprised of M.J . Boylen, K.C . Irving and Patino acquired the rights of St. Joe Lead to the No. 12 property. A concerted development plan was initiated along with construction of the necessary sur face facilities and by July 1964, the property was in production at 4500 tons per day.In 1967 Noranda Mines Ltd. acquired operating control of Brunswick Mining and Smelting Corporation Ltd."

Jan 1, 1988

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module (Fig. 1). The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information. Key for Fig. 1 A-Helium-3 neutron detector B-Wax moderator for neutrons C-Irradiation chamber D-Funnel for receiving ore E-Gate F-Transducers G-Biological shield for neutron source H-Transfer tube for neutron source I-Neutron source J- Microprocessor K-Accumulator for hydraulics L-Control panel M-Electro-hydraulic servo valve to control movement of the source N-Electro-hydraulic proportional valve to control movement of the gate O-Hydraulic power unit Irradiation and Counting Equipment Irradiation of the ore is achieved using a 1 X 10 7 neutron*s-1 californium-252 neutron source. The source is contained in an iron capsule and pneumatically transferred between the irradiation/counting and storage chambers. The irradiation/counting chamber possesses 12, 600-mm (24-in.) long helium-3 thermal neutron detectors embedded in an annulus of wax (Fig. 1). Two sets of six detectors are each interfaced to a preamplifier, amplifier and single-channel pulseheight analyzer system, and the two systems are connected to a scaler. A special timing unit controls the transfer of the source and the on and off switching of the detector assembly, in sequence with the movement of the source. Results and Discussion In the experimental demonstration, using the 107 neutron.s-1 source, a 10% uranium ore was measured to ± 10% (2o total error) in 10 cycles and ±6% (2o) in 100 cycles. Since a larger source was not available, it can be statistically predicted that a 109 neutron-s-1 source could be used to measure a 0.1% ore to ± 10% (20) in 10 cycles. The uranium can occur as discrete nonhomogeneous particles in the ore, and the error caused by such a distribution of particles is ± 4% (2o), as determined by

Jan 1, 1984

By Steven I. Majtenyi

There is a growing interest among transportation planners to make more use of the subsurface to preserve the landscape and make transportation more efficient and economical. At the present time, utilization of the underground for transportation facilities is severely restrained by the adverse environmental impact and high construction costs. However, recent studies performed by the United States Department of Transportation indicate that, in spite of such problems, the volume of underground construction has significantly increased during recent years and projections point to continuation of these trends(1). It is estimated that the total volume of transportation tunnel construction will exceed $4 billion in this decade. This level of Investment justifies a significant research program. Due to the common interest in transportation tunnels and to the similarities of problems encountered during their construction, the tunneling research activities of the Federal Highway Administration (FHWA) are closely coordinated with the programs other agencies of the U. S. Department of Transportation DOT); specifically, the Federal Railroad Administration (PRA) and the Urban mass Transportation Administration (UMTA). The FHWA is responsive to the particular needs of these agencies but will benefit from the research results obtained also. In 1971, the Office of Secretary, DOT, initiated plans for an overall coordination of the DOT, FICA, UMTA, and FHWA tunneling research activities. The objectives of this effort have been to reduce tunneling costs by 30% and to increase the rate of excavation by 200% within 10 years. This coordinated program was developed and promoted during 1972 and 1973 by Dr. Eugene L. Foster, at that

Jan 1, 1974

By A. V. Wilke

The metallurgical balance is a tool usually applied in the mining industry. The final result is used for different objectives. All the existing software for the calculation of the metallurgical balance fits the numbers in such way that a logical correspondence between the input and output materials exist. Nevertheless, the accuracy of the metallurgical balance strongly depends on the quality of the input data. Division El Teniente of Codelco Chile has been optimizing the metallurgical balance system the last 14 years. As well, it has participated in the development of corporate norms to optimize the operations of sampling and weighting systems. As a consequence, the precision and the accuracy of the metallurgical balance have been improved in a significant way. Also the understanding of process variability was enhanced. The improvements have allowed Codelco to discover and minimize hidden losses and to better support top management decisions. The present paper describes the main aspect of the continuous improvement process.

Jan 1, 2009

By H. R. Stahi

THE ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classifiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9 1/2 -in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3 ½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 5 2 ¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1 3/4in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges are renovated occasionally, broken pieces being removed and the charge brought up

Jan 1, 1945

By R. C. Bryan, D. B. Jamison

This paper describes a procedure for minimizing the operational costs associated with blending sodium in coal at a surface mine in the Powder River Basin. Mine site blending is performed by adjusting daily production rates for high- and low-sodium pits. These adjustments are designed to produce a consistent sodium value called the blending target. The mine's contractual limit for sodium is sufficiently close to the field average value that any decrease in the sodium blending target may result in an opportunity cost, if the number of potentially minable tons is reduced. Conversely, any increase in the blending target may result in a higher sampling cost if more information is required to keep the blend within contractual limits. Blending parameters were defined using a geostatistical block model of sodium content that was based on analyses of hundreds of in-pit sample holes. Variograms for these data were used to compute the marginal costs of opportunity and sampling for a series of hypothetical in-pit drilling patterns. These marginal costs provided a basis for selecting optimum values for key parameters such as the blending target, production variance and in-pit sampling density. The benefit of implementing these optimized parameter values is that operational costs of sodium blending will be minimized over the long-term. At the mine where this procedure was developed, each dollar spent to increase the sampling density to the optimum value was shown to generate a $10 savings in opportunity cost.

Jan 1, 1991

By J. L. Weeks

Discussion by J.L. Weeks Discussion I wish to comment on Kissell and Sacks's paper "Inaccuracy of area sampling for measuring the dust exposure of mining machine operators in coal mines." It is well established in industrial hygiene that area sampling is not an accurate measure of personal exposure. To measure personal exposure, one must use a personal sampler. Kissell and Sacks, using data from coal mining, reached the same conclusion. This is the right answer but it is the wrong question. This issue arises because the machine-mounted respirable dust continuous monitor (MMRDCM) is being considered to monitor coal-miners' exposure to respirable dust. This device may be mounted on a continuous mining machine, and it would take a modified area sample. I refer to it as "modified" because a mining machine is not really a fixed location, as is the typical area sample. The concern with whether this machine would provide an accurate estimate of the machine-operator's exposure, however, overlooks important features of the MMRDCM. Specifically, it displays dust concentration as it occurs and at the place where it is needed, not 10 days later, as is the case with the current system. Used in this way, it allows for timely intervention to control exposure to dust and it leads to greater first-hand knowledge of specific conditions and mining practices that result in excess exposure. If one knows the dust level is excessive at the time the dust is being generated, it is possible to use known and effective methods of dust control to reduce exposure. By measuring again, one can determine whether control methods have worked. This is a generic procedure followed in any industrial process: measure, adjust and measure again. It leads to more effective and more efficient dust control and is preferred to taking a sample, waiting 10 days for the results and then trying to apply that information to a situation that no longer exists. The MMRDCM also records the dust level automatically, and it runs continuously, making it harder to tamper with. Dust monitoring in underground coal mines has been corrupted by some operators having submitted fraudulent samples (and being convicted on criminal charges because of it), weakening the credibility of the current system. The MMRDCM is, in part, a response to this corruption, and an effort to focus attention on dust levels and dust controls rather than on the perpetrators of fraud. Use of the MMRDCM is an evolutionary change from the current system. With the current system, one takes a personal sample of exposure of the "designated occupation" (DO, defined at 30 CFR 70.2 (f)), the occupation regularly exposed to the highest level of dust on a section. Typically, on sections using continuous mining machines, this is the mining-machine operator. The rationale behind DO sampling rests on some of the findings that Kissell and Sacks report. They reviewed several studies that showed that dust concentration varies with location on a mining section and that there is a concentration gradient with the highest concentration occurring closest to the face. The logic behind DO sampling strategy is that if the dust concentration at the place where it is highest is below the exposure limit, then concentration elsewhere on the section is also below the limit. In this way, one will have achieved the objective or keeping dust concentration below the exposure limit for all miners on the section by measuring the exposure of only one miner. The

Jan 1, 2003

By B. Fernow

(Read at the Philadelphia Meeting, February, 1878.) THE question of preserving the forests in this country is an important one, not only to trades using wood but to the whole nation, and though agitated for many years has not received that general consideration which its broad bearing demands. Those who are interested in and working for reform in the treatment of American forests are looking to the government for help, suggesting the creation of a commission to study the conditions of forests at home and abroad, and hope that a sound system of managing the woodlands will be thereby inaugurated. This reform will cost time and money, all the more because the government can bring to bear upon private owners hardly any power but that of good example and encouragement. There is, however, another aspect of the question to which I wish to draw your attention, namely; the more economical use of the valuable material which our forests offer, a more careful and exhaustive utilization of its constituents in all cases where wood or its products are applied. The wasteful consumption of wood in the United States is in every way so enormous as to justify the statement that by merely adopting an economical husbandry in this particular the destruction of forests might be delayed for many years, while the neglect of this imperatively needed and immediately practicable reform deprives legislative remedies of much of their value. In this regard I wish to draw the attention of metallurgists, especially those who are interested in the manufacture of charcoal iron, to a more economical and profitable use of wood in iron manufacture. For thousands of years of human history wood was the main, nay, the only material used for fuel. Mineral coal came into use only in the fourteenth century, and it is only recently that it has largely displaced its less carbonaceous predecessor as fuel. The art of concentrating carbon in the form of charcoal for easier transportation and readier use in distant places was, according to Pliny, known to the ancients long before Christ, and was practiced in all countries, especially in connection with the iron- manufacture.

Jan 1, 1878

By Murray W. Grabinsky

The in situ performance of Cemented Paste Backfill (CPB) systems can be monitored to improve understanding of the paste?s behaviour and therefore optimize the geomechanical design of the cemented paste backfill system (i.e., minimize binder content, minimize fill barricade complexity, maximize fill rate-of-rise during backfilling, and minimize cure time before resuming production blasting in proximity). However, undertaking such a monitoring program is inherently complex, because the rate-of-rise and accompanying self-weight induced stresses occur at a time rate that is similar to the rate of hydration and associated increase in the paste?s stiffness and strength. In addition, arching within the stope complicates the overall stress distribution. Therefore, to obtain a complete picture of the CPB?s response, it is necessary to measure several variables (3-D total pressures, water pressures and suctions, hydration rates, and blasting induced vibrations) at a given ?point?, and take all such measurements at many points throughout the fill mass as well as at barricade locations. The overall testing program can therefore be logistically difficult to implement and have an apparent high cost. However, if the program is successful, the benefits arising from improved mine design can significantly outweigh the seemingly high initial capital and labour costs of the monitoring program. A case study involving an instrumented paste fill in a tall and narrow Alimak stope is briefly outlined. Plans for a similar experiment in a large long-hole stope are then considered.

May 1, 2008

By J. Andrew Leach, Robert T. Tape, Thomas G. Harper

A practical approach to elope stability design of heap leaches up to 90 m (300 ft) high is presented based on recent experiences with heaps throughout the Western U.S. Shear strength parameters for foundations, ore and liner systems are considered, with particular attention given to frictional contacts between geomembranes and clay or geotextiles. Stability of both ore and containment dikes is discussed including the effects of seismicity, staged dike construction, progressive heap development and in-heap solution levels. Design criteria and methods of analyses are presented along with techniques for improving stability.

Jan 1, 1987

Partnership with MiHR ? AHRC and MIHR jointly developed Mastering Aboriginal Inclusion for Mining ? Customization specific to mining Mining is unique? ? History of Aboriginal peoples and the land ? Unique community needs ? Effects of Impact Benefit Agreements ? Remoteness of mines ? 1200 Aboriginal communities located w/in 200km of active mines or exploration ? Long term effects of the land

May 1, 2009

By A. F. Lucas

The rock-salt deposit of Petite Anse, in Louisiana, has been known for many years. A description of it, with an account of the method pursued in its exploitation, was contributed in 1888 to the Transactions of the Institute* by Mr. Richard A. Pomeroy. But this locality does not by any means comprise the whole of the resources of the State in rock-salt. A few months ago the writer published a general account + - of all the Louisiana localities now known, together with particulars concerning their technical and commercial development. In view of the fact that active enterprises in this line have increased in number, and promise to constitute an important, if not controlling, factor in the supply of salt for the immense demand of the United States (including, as it does, the consumption of this mineral as a raw material in chemical manufactures), the present paper has been prepared. While it necessarily contains much of the substance of the article above mentioned, it is accompanied with illustrations and details not heretofore published. Moreover, the former article is out of print, the publishers having no copies left of the number of their journal in which it appeared, and the writer is therefore warranted in the belief that a paper placing the essential facts on record in a form more permanently accessible to mining engineers will be acceptable to the members of the Institute.

Jan 1, 1900

By P Voulgaris

The Osborne Mine commenced production from an open pit in December 1994 with the first ore being hauled from underground one year later. Currently all ore is hauled up a decline by trucks until the shaft system is commissioned in early-1998. The mine is hosted within a multiply deformed and complex sequence of metamorphic, igneous and metasomatic rocks of mid Proterozoic age that are concealed beneath 20-40 metres of Mesozoic sedimentary cover. Epigenetic and largely hypogene copper - gold mineralisation occurs in a close spatial relationship with quartz - magnetite ironstones and is an important example of a significant mineralisation style in the Mt Isa inlier. Life of mine concentrator production, to year end 1997, totals 3.2 Mt of ore at 3.40 per cent Cu and 1.13 g/t Au. Total metal produced was 97 000 tonnes of copper and 2900 Kg of gold (95 600 oz) from some 348 000 tonnes of dry concentrates averaging 28.01 per cent Cu and 6.99 g/t Au. At the start of 1998 OsborneÆs total measured and indicated resource was stated at 11.38 Mt at 3.24 per cent Cu and 1.21 g/t Au with an ore reserve of 11.42 Mt at 2.90 per cent Cu and 1.09 g/t Au. Exploration efforts in and around the mine have been increased since 1996 and have resulted in an early success, with a 28 per cent increase in ore reserves since mine start-up, and the establishment of a 1.5 Mt inferred resource open to depth. During 1997 the performance of the concentrator has exceeded expectations with throughput having been increased to 1.45 Mt per annum. The Mining Cycle at Osborne has proceeded from discovery through feasibility and is currently seen to be in the latter part of the development stage and/or the early part of an expansion stage. This paper will describe the role of geology and exploration within these stages and will highlight how a number of key interpretations and decisions based on geology have both favourably and unfavourably affected the project.

Jan 1, 1998

By Dickson B. L

Groundwater geochemistry is controlled by chemical interactions with aquifer rocks, and offers a powerful geochemical exploration technique for locating concealed mineral- isation. Techniques are applied for identifying constituents of aquifers using the compositions of their groundwaters, collected from the surface through drill holes, water bores or springs. Even slowly moving groundwaters have passed through large volumes of rocks which surround any sample site. Products from these rock-water contacts modify the groundwater. By relating water composition to the minerals from which the dissolved species may have been derived, geochemical exploration information is obtained for a significantly larger zone than is represented by drill hole rock samples. In this paper we describe some solution chemical tech- niques we are applying in mineral exploration, particularly in areas of Australia where ground- waters are highly saline and contain enhanced concentrations of trace metals. EQUILIBRIUM MODELS FOR SALINE GROUNDWATERS Central to our technique is the assumption that in low rainfall regions, aquifer rocks and contained groundwaters are in chemical equili- brium. Solubility calculations then allow us to predict upper limits to the soluble products of groundwater interaction with aquifer rocks and to identify minerals with which the groundwater is saturated. Whilst equilibrium calculations require input of concentrations of all dissolved species, they are actually based on "activity" of dissolved species rather than the analysed concentrations. Unless waters have very low ionic strengths (a measure of total dissolved solids), activity of a solute is not equal to its analytical concentration but equals concentration multiplied by an "activity coefficent". Several methods exist for calculating mineral solubilities. The ion-association model, commonly used in programs such as WATEQ (Truesdell and Jones, 1974) and later deriv- atives such as MINTEQ, assume the formation of complex ions such as MgSO4(0), CaCl+, which are described by formation constants. Activity coefficients are obtained by the extended Debye

Jan 1, 1988

By H. D. Kiddle

IN mining in Malaya has been a courageous ad- venture in private enterprise, attended by all the hazards of prospecting in unexplored areas of tropical jungles-jungles that still cover four-fifths of the country. To assess the contribution of an industry to a country's revenue, it is necessary to consider not only duties levied on the products, but also the revenue from all businesses and individuals directly and indirectly deriving income from the industry. Exact figures are not obtainable, but from 1850 until shortly before the first World War, tin mining in Malaya was a dominant factor in the Federation's development. Even at the turn of the century, tin export duty alone represented 45 pct of total Government revenue, while in the 40-year period up to the end of 1937, tin ore duties collected by the Federated Malay States Government amounted to 383,- 81 0,284 Malayan dollars,* approximately one-fifth of the total revenue for the same period.

Jan 11, 1957

By E. Rupprecht, J. Volkwein, K. Cashdollar, R. Gürtunca

The National Institute for Occupational Safety and Health (NIOSH) mining research program at the Pittsburgh Research Laboratory (PRL), includes the areas of coal mine dust control and monitoring, and explosion prevention. One very important project at PRL is development of a continuous recording person-wearable respirable dust monitor. This Personal Dust Monitor (PDM) is a respirable dust sampler and gravimetric -equivalent analysis instrument that is part of a belt-worn mine cap lamp battery unit. Prototypes were successfully tested in the laboratory and in four underground coal mines. The PDM testing demonstrated that it was convenient to wear and use, robust, and provided accurate and timely data. NOSH is also conducting research in the program area of coal mine explosion prevention, suppression, and mitigation. Aspects of this research program include: full scale explosion research, laboratory scale research, frictional ignition of methane, mine ventilation seal construction, prediction of coal mine methane emissions, mine monitoring systems, flame cutting and welding operations, spontaneous combustion, and emergency preparedness.

By D’Arcy A. Shock

In this year's review, the concept of solution mining has been expanded to include total solution as well as the technology of leaching, or partial solution methods. Leaching methods continue to make strides in the processing of metallic minerals, particularly in copper and uranium production. No quantitative data are available on the increase of leach operations; however, it is known that there were more systems installed in 1969 and that the estimated 15% of U.S. domestic production of copper produced in 1968 was by leach methods.1 While the number of publications appearing during the year were few,2,3 there were special sessions in this field at the AIME Annual Meeting and SME Fall Meeting in 1969 and more scheduled for these meetings in 1970. Significant areas that should be mentioned are: the continuing interest in bacteria to convert sulfide sulfur to acid thus reducing reagent costs in both copper and uranium sulfide ores, the exploration of leaching agents other than common mineral acids or alkaline solutions,4,5 kinetic studies of leach mechanisms and the development of microfractures for increased leach solution penetration by use of chemical or atomic explosives.6,7

Jan 1, 1970