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By J. F. Myers, S. D. Michaelson, F. C. Bond

This work was undertaken with the object of collecting plant data on rod milling, making laboratory tests on representative samples of the various ores, and arriving at a basis for comparing the relative efficiencies of different mills operating on widely different materials and under varying conditions. These efficiencies consider only power consumption and do not include the other elements of operating costs. The over-all economic efficiency may be quite different from the consumption efficiencies with which this paper is concerned. It is hoped that these collected data will be of value in indicating the most favorable conditions of rod-mill operation. Operating-plant data on 14 different installations from II plants throughout the United States and Canada were collected, and laboratory tests were made on samples of these ores. The plant rod mills were all operated wet, and either in open circuit or with circulating loads so small as to be negligible. Data were obtained on one roll-crushing operation at the Utah Copper Co., for comparison. The rolls were operated wet in closed circuit with 10-mesh vibrating screens, and with a high circulating load averaging about 600 pct. This particular roll operation can be considered as a model of efficiency. The collected plant data are necessarily incomplete in some instances. Each ore was assigned a code number, and the name and location corresponding to the code number are listed in Table I when permitted. The power data in every instance refer to motor input. Net power was reported by the operator for samples 1-30 and 1-45, and this was increased by 20 pct to compensate for motor losses and power factor. The plant rod-mill data are listed in Tables 2 and 3. In Table 2 the speeds are calculated on the basis of partly worn rod-mill liners 2 in. thick. Where D

Jan 1, 1949

By Kiyoshi Azuma, Hiroshi Kametani

Dissolution of a ferric oxide in acid solution is divided into two different types In the accelerated type dissolution proceeds in three stages 1) an inittal reaction during which the dissolved amount of oxide is nearly proportional to the cube root of time ; 2) an accelerated reaction: and 3) a final stage during which dissolution approaches completion. This course of dissolution gives an S-shaped curve in the above sequence OH a log weight percent of dissolved oxide against log time graph. In the parabolic type dissolution proceeds in four stages: first and second stages similar to those of the accelerated type; 3) parabolic-rate dissolution during which the dissolved amount of oxide is nearly proportional to the square root of time: and 4) a final stage. The effect of the kind of acid, the temperature of solution, and the concentration of acid was investigated. The results are discussed and a rate equation is proposed. AN attempt has been made to obtain some information on dissolution of a ferric oxide in various acid solutions. In view of hydrometallurgical processes the dissolution of ferric oxide is of less importance than that of zinc oxide or copper oxide in general. One of a few processes which intend to dissolve ferric oxide is an elimination of iron impurity from acid-insoluble materials. However, most leaching processes of oxide ores and cinders are accompanied with more or less dissolution of iron. According to Pryor and Evans1 the rate of dissolution of a ferric oxide in various acids decreased in the order of hydrofluoric acid > hydrochloric acid > sulfuric acid > perchloric acid. This order was in accord with the order of decrease of the complexing affinity of the anion for ferric ion,

Jan 1, 1964

"""Hazard"" is the general expression used by safety engineers to denote risk, or the danger to property or human beings from the contingencies peculiar to different industries. Where a business, family, or individual is subjected to possible financial loss from any particular hazard, it is customary and is considered good practice to insure for reimbursement of possible loss. Accordingly, we take out insurance for earthquakes, for fire, for accidents, etc., and are charged a different premium for each. The premium rate is based on our relative exposure to the hazard in comparison to the total loss to our particular class of business in our geographical location over a given period.Insurance, then, must be considered as the experience of a group rather than of an individual, and any individual may be fortunate or unfortunate over stated periods. Through the law of averages, the probability of loss because of exposure to hazard is becoming more accurately measured as experience is accumulated and data are analyzed.Lion human life came to have a financial value in the eyes of the law, and employers were held liable for losses through accidents to employees while on duty, insurance companies began to underwrite the accident risks of the different employers and divided them into classes based on the amount of risk involved. Premiums for compensation insurance are based on relative exposure, which is measured by pay roll. The premium is rated in proportion to the severity of losses established by statistical studies.' It soon became evident that the accident experience of employers in the some industry varied widely; that is, there were good and bad risked in the same line of business. This resulted in ""experience rating"" of premiums, meaning thereby that an employer with experience in accident cost was premium rated in proportion to his losses over a stated period (usually three or more years)."

Feb 1, 1926

By Pope Yeatman

THE Committee on the Elimination of Waste in Industry, of the Federated American Engineering Societies, in its report says that "Waste in -industry is attributable to four causes: 1. Low production, caused by faulty management, of materials, plant, equipment, and men; 2. Interrupted production caused by idle materials, idle plants, and idle equipment; 3. Restricted production intentionally caused by ,owners, managers, or labor; 4. Lost production caused by ill-health, physical defects and industrial accidents. The committee sent out elaborate questionnaires, and in them divided the responsibility into Manage-ment, Labor, and Outside Contracts. These question-naires showed that in the different industries examined the responsibility of Management was from 50 to 80 per cent. and of Labor from 9 to 28 per cent. The percentage of responsibility that Management was held accountable for was much larger than any of us supposed. However, when this question is studied, we realize how necessary it is to select, in the first instance, suitable personnel, in the form of directors, managers, heads of departments, experts, etc. We would like to -believe that the percentage of responsibility of Management for waste in the mining business was less than that quoted above, but there are so many mining properties started up and operated by men of no experience in such lines, and there have been so many improvements in late years in mining and metallurgical methods not taken advantage of, that we can not believe that. the percentage will vary much from that already quoted. Quoting from a report approved by the Committee on Management of the American Society of Mechanical Engineers: Management is the art and science of preparing, organizing, and directing human effort applied to control the forces and util-ize the materials of nature for the benefit of man.

Jan 3, 1922

By S. J. Green, R. D. Perkins

Little data exists on the high strain rate behavior of geologic materials. Uniaxial stress tests by Kumar1 and by Serdengecti and Boozer2 present some results to strain rates in the range 10 to 103 per sec. The two references cited are the only ones the authors found for high strain rates. Techniques for experiments at high strain rates are not well developed and, in general, leave questions with respect to interpretation of data when the material behaves in a brittle manner. The most reliable data at very high strain rates (104 to 106 per sec) are found in some recent flat-plate work3-5 where experimenters have impacted one plate into another, thereby creating a uniaxial strain loading. The standard uniaxial stress-strain curves cannot be obtained directly from these tests; however, yield (or fracture) may be studied by measuring the "elastic" wave magnitude. These data then may be compared to uniaxial stress data by resorting to some preassumed yield/fracture criteria. Even with the most liberal interpretation of existing data, little can be said of the effects of high strain rate on stiffness, yield/fracture, and post-fracture behavior of geologic materials. It is the purpose of this work to begin to add to the understanding. The data presented here are intended to provide material information for use in calculating the response of geologic materials to high rate deformations. These data can be applied directly to determine parameters such as the elastic modulus and Poisson ratio as a function of stress. Furthermore, these data can be used to derive an understanding of yield/fracture and the effect of rate on the yield/fracture surface. The work presented is a first step, and is

Jan 1, 1972

This paper summarises the results of a study carried out for the Sustainable Minerals Institute, University of Queensland, aimed at identifying the major water-related issues facing the minerals industry internationally and the research required to support the development of practicable solutions. An overview is given of the sources and uses of the worldÆs water resources and of the regional and global sustainability of water supply and management. Agriculture is the worldÆs major user of water; it wastes more water than the minerals industry uses. Water reform and its potential impact on the minerals industry are discussed with particular reference to Australia and South Africa. The sources, uses and management of water on mine and processing sites are outlined. Particular attention is paid to the importance of groundwater as a source of often saline water that can be used in mineral processing and for other purposes. Aquifer storage and recovery has the potential to help conserve water resources. Generally, the major uses of water on mining and processing sites are for mineral processing and refining and for the associated tailings disposal. Water consumption for these purposes may be reduced by recycling process water, using saline water, paying close attention to water management, thickening tailings and using techniques such as dry stacking and paste. A change in the minerals processing paradigm in respect of water use could come through research into new, improved and less costly dry processing techniques. The treatment of mine and processing water will be required increasingly before water is reused, stored, discharged or sold, and before sites are decommissioned. Research and development are required to reduce costs and develop new treatment methods. Mine closure and decommissioning raise a number of long-term water issues requiring continuing generic and site-specific research and development.

Jan 1, 2003

By Lennart Malmqvist

Decision-making in exploration deals with experience, knowledge, and sometimes faith, but very often financing and profit, mineral resource aspects, employment, and general political ambitions are parts of the scene. At one end of this spectrum, there is the prospector and his decision on the actual project. The exploration business now expands from its pioneering phase for the benefit of a single company to a wider spectrum which includes the benefit of the large mining consortium and a national development program. Its objectives are also to maintain employment and to favor long-term stability in a mining district or a mining company. The prospector then has new aspects and questions to consider such as, what to do in mining districts where there is a shortage of ore resources but there are still useful investments in mines, processing plants, etc. To analyze the exploration process, therefore, seems to be of great importance in a time of rapidly changing conditions. As in other areas of complex decision-making, for example, in the fields of investments, of production, and of the armed forces, computers most certainly have an important role to play in future exploration decision-making systems. However, we must first analyze our own business as closely as possible to find out what kind of information we need, i.e., why we recommend an exploration program as a good or a poor one. An important part of the debate about exploration is the case study. Case studies on exploration are few and most of those written are only about successful exploration programs. Where are the case studies of the failures and the not so successful programs? Every prospector knows that the majority of exploration projects fail. Consequently, the most important experience seems to be hidden in the failures.

Jan 1, 1979

By Frank W. Osterwald

Violent, spontaneous destruction of coal faces and ribs during, what are commonly called, bumps endangers and at times destroys life and property in mines of the Book Cliffs coalfield, Carbon County, Utah. Because these mines supply much of the coking coal for the West Coast steel industries, mining is necessary, although costs are high and operating conditions are very difficult at times. As operations continue, mining will go to greater depths, and the difficulties will increase. Whether or not mining can be continued successfully depends in part on whether or not incidence of bumps can be controlled successfully. For this reason, the U.S. Geological Survey, at the request of and in cooperation with the U.S. Bureau of Mines, undertook a study of the Sunnyside No. 1 mine to determine what geologic factors were related to the incidence of bumps. This report presents the preliminary results of that study and indicates that some observable geologic features in the mine may be related directly or indirectly to bumps, and that other features may be used to infer stress distribution around workings that will aid in design of roof support measures. Coal is mined in the Sunnyside No. 1 mine from depths as great as 2500 ft by room-and-pillar methods. The surface topography is rugged with many steep cliffs and deep canyons, hence the load on the coal varies greatly within short distances. Much of the Sunnyside coal is under high stress, and while being mined, it shatters continuously' and falls from the face. Continuous shattering (bumping) makes mining easy, but if the shattering is interrupted, greater stress is built up and violent bumps may result when the coal is mined.

Jan 4, 1962

By H. M. Chance

RECENT research work has shown that coal can be produced, at reasonable cost, from almost all coal-mining districts containing not more than 3 to 8 per cent. of ash. From coal so produced, an abundant supply of coke, suitable for blast-furnace operation, can be made containing from 12 down to 6 per cent. or less of ash. These statements apply broadly to the Appalachian coal belt, including most of the mining districts in Pennsylvania, Ohio, West Virginia, Kentucky, Tennessee, and Alabama. The exceptions will generally be found among those coals carrying a prohibitive percentage of sulfur. Any gravimetric heavy fluid method of separation will effect a reduction in the percentage of sulfur, but this reduction may not be sufficient to produce, from coal unfit for metallurgic use, a product that can be so used. When a large part of the sulfur is in the form known as "organic sulfur, "or is present as very .finely disseminated and homogeneously distributed pyrite, possibly it cannot be eliminated from the low-ash coal produced. When the sulfur exists as pyrite in distinct layers, bands, or nodules, or when it is finely disseminated but is irregularly distributed, often a low-ash coal low enough in sulfur to make good metallurgic coke can be produced from relatively impure coals. It has generally been assumed that the quality of coke on which blast furnaces must be operated depends on the geographic location off the furnace, and that the grade of coke on which a blast furnace must operate is that produced from the best coal obtainable by careful mining; or if the coal is prepared by washing, hand picking, or other means by which the foreign impurities are removed, that the grade of coke on which the furnace must operate depends on what has been termed the "inherent ash" of the particular coal.

Jan 6, 1924

By William H. R. Gibney

Pine Point Mines is located in the Canadian Northwest Territories about 1,200 kilometers north of Edmonton, Alberta, close to the south shore of Great Slave Lake. Originally staked in 1898 by prospectors en route to the Dondike gold rush, it was not until the mid-1940's that large scale exploration was undertaken and Pine Point Mines Limited was formed. Work was suspended in 1954 while studies were made to resolve one of the main problems, transportation. In 1961, under the Canadian Government "Roads to Resources" program, agreement was reached for the construction of the Great Slave Lake Railroad and plans made to put the property into production. A townsite was established in 1963 and plant construction commenced. The first production was in the form of high grade ore which was shipped to Kimberley, B.C. for concentration starting in late 1964. The Pine Point concentrator started up in late 1965 when hydroelectric power became available. Initially designed for 4,540 metric ton per day, facilities were expanded to a designed 7,250 metric ton per day in 1969. Problems associated with mining operations at Pine Point are not unique but are augmented by the isolation, climatic conditions, topography, high water table and spread out nature of the operation. Maintenance of a reasonably stable work force is the number one problem but this is gradually being overcome as more local people are trained and oriented from a nomadic type life to an industrial society. The purpose of this paper is to give a general description of the mining operations of Pine Point Mines Limited and to review some of the problems encountered in the operation along with problems associated with northern operations.

Jan 1, 1970

By W. F. Carew, F. A. Crossley

The stress for rupture in 500 hr at 1000° F has been reported to be about 13,000 psi higher for Widmanstitten than for equiaxed microstructures for the Ti-7A1-3Mo alloy.1,2 Also, limited data indicated Widmanstatten and equiaxed micro-structures to have approximately the same creep resistance at 1000°F.' Recently, data have been obtained which show Widmanstatten structures to be significantly more creep resistant than equiaxed structures at 1000°F, see Fig. 1. These data represent several ingots and several heat treatments of the solution treat and age type. The thermal history and tensile properties corresponding to the creep data are given in Table I. The creep data were taken from creep-rupture tests. The associ- ated rupture data are given in Table 11. It may be noted that total creep deformations are considerably greater for equiaxed structures. Previous evaluations of stability indicated that all of the structures tested are stable with respect to exposure to stress at elevated temperature.1,2 For a minimum creep rate of 10-4 in. per in. per hr the difference in stress levels is indicated to be about 17,000 psi. Holden,et found an irreversible contraction upon heating three a-ß titanium alloys in the equiaxed condition above their ß-transus temperatures. The alloys were: Ti-6Al-4V, Ti-155 A (5Al-1.3Cr-1.5Fe-1.2Mo), and C-130 AM (44-4Mn). The net contractions in length between the original equiaxed and the final Widmanstatten structures measured at room temperature were: 0.4, 0.4, and 0.7 pct, respectively. The greater creep and rupture resistance of the WidmanstStten structure of the Ti-7Al-3Mo alloy may be associated with the phenomenon

Jan 1, 1959

By Gordon C. Hicks

Prior to 1950 most fertilizer manufactured in the United States was produced in a nongranular form. In such form, the material caked when stored and was extremely dusty when applied in the field. About this time a wide interest developed in the fertilizer industry for the manufacture of a granular, high-analysis fertilizer of improved physical properties. Thus, TVA began work on fertilizer granulation about 1950 using both a pugmill (double-shaft paddle mixer) for a thick slurry of nitric phosphate and a continuous ammoniator-granulator drum for superphosphate-base mixed fertilizer. Granulation in these units was mainly by agglomeration. The liquid phase generated by the moisture and by the heat of reaction of ammonia with acidic materials, such as superphosphate and sulfuric or phosphoric acids, resulted in agglomeration of the solid ingredients. About 1956, TVA began studies of the use of a pan or disk granulator for production of ammonium nitrate or other fertilizers containing a large proportion of ammonium nitrate, such as 30-10-0 ammonium phosphate nitrate. This is the layering method of granulation. More recently, TVA has demonstrated pan granulation of urea from a highly concentrated solution and granulation of ammonium polyphosphate melts in a pugmill. The bulk of TVA studies throughout the years has related to use of the drum granulator. About 200 granulation plants in the United States have been licensed and built to use the TVA continuous ammoniator-granulator; it is estimated that more than half of the 20 million tons of mixed fertilizer used in the United States in 1974/1975 was produced in these plants. Some other granulation processes developed for use in the drum granulator include diammonium phosphate, monoammonium phosphate, urea- ammonium phosphate, ammoniated superphosphate, and most recently, NPK fertilizers by melt granulation using a pipe reactor.

Jan 1, 1977

By Klaus Emicke

The paper describes the lead refining process operated at Norddeutsche Affinerie (N.A.). Incoming materials are different grades of lead with varying percentages of impurities: Cu, Te, As, Sn, Sb, Bi, precious metals. The facilities are designed for charge-wise operation. After drossing, generally Te, As, Sn and Sb are removed substantially by the Harris Refining Process using fused NaOH. A certain selectivity during refining is achieved by controlled addition of NaNO3. Another refining step is decopperizing by means of elemental sulfur. Complete separation of Te, As, Sn and Sb is accomplished by chemical methods in a leaching plant. A Howard Press is used in the Parkes Process for desilverizing. After vacuum dezinking in kettle, Bi is removed, if required, by means of Ca and Mg. In an end refining step Ca, Mg and Zn are removed by having the lead melt treated with NaOH and NaN03. The refined lead as such, or in the form of special alloys, is cast by an ingot casting machine into 50-kg ingots. Zinc which is used as a precipitant in the Parkes Process is recovered by distillation from the high grade Ag-Zn-Pb crystals in Faber du Faur retorts while the Pb/Ag alloy is fed to the cupellation furnace. Ca and Mg are removed from the Ca-Mg-Bi crystals formed in the course of Bi removal, and the Bi-bearing lead is cast into anodes. The electrolytic lead refining facilities at N.A. are operated largely for the purpose of separating Pb and Bi. These facilities are also used particularly for the refining of Bi-bearing lead which had gone through the above process, but from which Bi has not been removed by Ca and Mg.

Jan 1, 1970

By A. M. Gaudin

UNTIL recently little attention has been paid to the effect of particle size on flotation. This has been especially true of material finer than 200 mesh.1 Particles of different sizes must behave differently in a flotation circuit; increased information along this line would be of practical use as well as of theoretical interest. The object of this study has been to determine what effect particle size has on floatability. The problem has been approached in two ways. One method consisted in a study of samples of the final products of several mills. These samples were sized and each size-product was analyzed for its important constituents. Bakelite briquettes2 of some of the products were made and studied under the microscope. This investigation is being continued, with particular reference to quantitative microscopic analysis of intermediate products of flotation plants. M In the other method of study, mixtures of pure minerals were ground to different degrees of fineness and floated. The products were then sized and analyzed. A study of the behavior of mixtures of pure sized minerals would have been illuminating, but after trial it was considered impractical on account of the tediousness of the large amount of elutriation required and the fact that the mineral particles would oxidize while being sized, preparatory to flotation. Broadly speaking, the following conclusions have been reached: 1. Results obtained by the two methods of study are in agreement. 2. The usual notion about the relative ease of flotation of very fine particles does not agree with the facts: 3. There is a time sequence of flotation with regard to particle size, the medium coarse particles floating first. 4. No successful method of floating extremely fine pulps has been found. 5. Recovery is optimum in a certain well-defined size range. 6. Selection is optimum in another well-defined size range.

Jan 1, 1931

By B Hall

In feasibility studies the choice of optimum cut-off grade is typically based on an evaluation of tonnage grade curves and manual design of stopes at one or two cut-off grades. In the industry these stope designs are usually based on break-even cut-off grades, but there is little consensus on what cost and price assumptions ought to be included. A new tool for automated stope design has been developed that generates sets of stope shapes for a nominated range of cut-off grades, or depending on the nature of the mineralisation, cut-off values based on net smelter return calculations in the block model. The stope design procedure includes constraints in stope and pillar geometry for the selected mining method and can be used to evaluate sublevel interval and sensitivities to other design parameters in the specific geological and geotechnical environment. These ænested stopesÆ are analogous to ænested pitsÆ in open cut design, and can be used as a key input in the æhill of valueÆ methodology for the evaluation of the optimum cut-off grade in underground design. The results of many consulting studies indicate that a break-even cut-off grade will not be the optimum cut-off grade. A particular benefit of the new stope design tool is that the design rules used to establish stope shapes in the feasibility study can then be used in mine production, in an optimal and repeatable method. This methodology is being advanced in the AMIRA P884 PRIMO project, which is a new research collaboration coordinated by AMIRA International. The project has brought together active research groups in Australia and Canada, with the goal of developing a flexible framework and set of software tools for optimisation in underground mine design.

Jan 1, 2009

By V. K. Grotzky, F. J. Fraikor

A number of investigators have noted the importance of various precipitation reactions on the properties of commercial-purity beryllium.1-5 Carrabine, for example, has demonstrated the interaction of iron and aluminum on the mechanical properties of hot-pressed polycrystalline beryllium.6 He considered the determining factor on elevated-temperature properties to be a reversible AlFeBe, precipitation reaction which could be directly related to the Fe/Al impurity ratio of the beryllium. Any aluminum in excess of the stoichio-metric amount required to form A1FeBe4 was presumed to remain primarily in the grain boundaries. We have observed, by transmission electron microscopy, such a grain boundary effect after various heat treatments of beryllium sheet. The starting material in this investigation was bare-rolled beryllium ingot sheet (0.001 to 0.005 in. thick), in contrast to the hot-pressed powder utilized by Carra-bine. It should be noted from the ingot sheet analysis shown in Table I that the Fe/Al weight ratio was near unity which would provide some free aluminum although the total aluminum and iron content was still small. Fig. l(a) shows some aluminum or aluminum-rich precipitates in the grain boundaries of a beryllium foil which had been quenched from 1200°C into brine at -TPC. Similar grain boundary precipitates were found, particularly at grain boundary intersections, in other samples quenched from 1100° and 1000°C. The length of annealing time at these elevated temperatures did not appear to substantially alter the amount of grain boundary precipitates, thus attesting to their relative insolubility. No matrix precipitates were observed other than the normal BeO and Be2C inclusions. A microprobe analysis of the larger precipitates at

Jan 1, 1968

By Frank R. Barnako

Coal mining is a hazardous occupation, but tremendous progress has been made in reducing accidental injuries and deaths in the mines. Let us take a look at the hazards in coal mining and the accident record. COAL MINE SAFETY TODAY Underground coal mining creates underground open areas, rooms, and haulageways. The roof (top) and ribs (sides) become highly stressed, and because coal and slate are very weak in tension or compression they frequently fail and fall from the roof and ribs unexpectedly. This is the most hazardous condition miners must cope with underground, and it has consistently produced the greatest number of fatal injuries in the coal mining industry. Only very sophisticated equipment, but which is not yet practical for use underground, is able to detect the buildup of stresses that may eventually lead to a roof or rib fall. Today's miner still looks for fresh cracks in the roof or rib, listens for the "drummy" sound, and feels vibrations from loose material in the roof, when tapped with a sounding stick, in order to find material which may fall. Once located, it is barred down, but if this is not possible, it is then secured to prevent it from falling. The next great& hazard is the equipment used to mine or haul coal. It is large, heavy, reacts quickly to its controls and is operated in con- fined areas with low lighting levels. The record of the last few years shows that fatalities attributable to mining and hauling equipment are about equally divided. Thorough training as part of a complete safety program is necessary to teach employees the hazards of loose roof, rib, and machine operation in order that they can work in the safest way possible. Unless there is such a program, the trade is learned by trial and error with a resulting tragic.

Jan 1, 1973

By H. V. Winchell

An examination of the Pachucal and El Oro districts in July, 1920, although cursory and incomplete, disclosed facts of more than passing interest to the student of ore deposits; and inasmuch as the literature on these districts is limited and some of their features are little understood, a brief account may prove valuable as reconnaissance observations to be amplified by more thorough and competent observers in the future. Where no complete monograph of important mining camps has been published a partial presentation of the data is often useful, as it sometimes leads to fuller and more accurate descriptions, by stimulating discussion. The following notes are presented with the consent of A. F. Main, managing director of the El Oro Mining and Railway Co., whose fund of information and wide acquaintance with both districts alone made it possible to cover so much ground in a short time. I have also to acknowledge gratefully the aid of C. A. Lantz and D. S. Calland of the Santa Gertrudis and Real del Monte companies, respectively, in connection with permission to inspect geological features of interest in their mines. Pachuca, Hidalgo, Mex. This mining camp has been described by Ezequiel Ordonez2 and by A. F. J. Bordeaux.3 Reported to have been discovered in 1522, it has produced silver valued at between $250,000,000 and $375,000,000. Subsequent to the discovery here, by B. de Medina, of the patio process. in 1557, its production was increased, and now, using cyanidation, the output of the camp is nearly one-sixth of the production of the world. Topography and Climate The Pachuca district consists of a series of mountains and valleys, the minor features of which extend in an easterly and westerly direction and reach elevations of 8000 to 10,000 ft. (2438 to 3048 m.) above the

Jan 1, 1922

The price of oil and other minerals may not make the costly business of offshore exploration seem -attractive at present. There is little doubt that the future search for the strategic resources will tend towards deeper waters particularly on the continental slope margins. In addition with the realisation that active tectonic environments (eg the Makassar Straits) cannot be excluded as potential hydrocarbon sources there is an increasing demand by the explorationists for deep water survey capability. Such deep water research has until recently been restricted to a few specialised academic institutions ;however with the anticipated extra, demand from other institutions, companies, and government agencies, Marconi Underwater Systems Limited (MUSL)and the Institute of Oceanographic Sciences in the UK have come together to allow MUSL to provide the deep water mapping system GLORIA on a commercial basis -for the first time. The GLORIA package comprises a deep water(more than 400m) side scan sonar system which, by operating at a frequency around 6.4 kHz can insonify swathes on the seafloor up to 60 km wide. In addition, single channel seismic, 3.5 kHz high resolution profiles, precision echo sounding, magnetic and gravimetric data all combine to provide as complete a regional interpretation as possible. The GLORIA system combines surveying at a reconnaissance level, covering as it does up to 20,000 sq.km per day, with a very high reliability, so that even in- economically unfavourable times, it can-still be considered as an invaluable and cost effective primary exploration tool. Analysis of the GLORIA data enables the identification of areas of particular interest requiring more detailed study by the more costly multi-channel seismics at a later date.

Jan 1, 1987

By V. I. Lakshmanan, I. G. McColl

"Heap leaching of gold ores in permafrost conditions is feasible provided certain criteria are met. These criteria include relatively short heap leach residence time and the availability of waste heat.This paper describes the conceptual design of such a system for a mining operation in the North West Territories of Canada. The design includes the concept of continuous countercurrent leaching in enclosed ore heaps.Indicative capital and operating costs are given.INTRODUCTIONIn April 1984, Acres Davy McKee Engineering Inc. (ADM Eng.) in association with Ontario Research Foundation (ORF), was invited to carry out a preliminary design, cost estimate and financial analysis to assess the viability of gold heap leaching in a permafrost region in North West Territories, Canada.An orebody was being explored which could, by nature of its mineralogy, be upgraded substantially by crushing and screening in the size range of 1/4 in. to 3/4 in. It was proposed to process the screen undersize through an existing gold mill on site and to treat the oversize material in a heap leach facility to be constructed adjacent to the existing mill.Despite the formidable challenges imposed by isolation, permafrost and harsh winter conditions, the heap leach project was considered worthy of examination because of the relatively high grade of the heap leach material (> 0.1 oz/ton), very short leach times (in the range of 1 to 4 days) and the availability of substantial quantities of waste heat from the diesel generating station on site.Design challenges included the necessity to maintain the permafrost in frozen condition, to prevent freezing of ore and leach solution, to limit size of construction materials to that which could be carried in a Hercules freight plane and to contain water usage at an absolute minimum level."

Jan 1, 1987