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By K Firmin

The Mt Tom Price orebody of Hamersley Iron Pty Limited consists of a folded hematite - shale sequence capped by a layer of goethite. Consequently, variations in iron content occur within the ore body, and as the overall grade is below that required to make specification products, selective mining must be carried out. A necessary preliminary to quality control of products based on selective mining is the production of advance data on the tonnages and grades of material to be expected in future mining areas. Bench plans showing zones of different grades and types of material anticipated are prepared from structural geological information. The grades of materials on a bench are essentially governed by the positions and attitudes of the shale bands. Quality control is exercised by blending different grades and types of material in primary crusher feed. Iron grade is the prime grade consideration although sometimes the level of phosphorus content must be considered. Pre-production grades are determined from assay of samples taken from blast hole cuttings. As shales are more friable than hematite, product fines are always lower in iron grade than the lump product. The realisation of contract specification grade is more difficult with fines than with lump, so maintaining the correct fines grade becomes a prime factor in quality control. Control of fines grade is presently effected by aiming for an appropriate feed grade to the primary stockpile. Studies to estimate fines grade directly by qualitatively taking into account material types are in progress. Quality control measures are checked by monitoring the crushing and screening plant output, and taking appropriate compensatory actions in the mine when required. Minor fluctuations in quality of fines can be blended out on stockpiles. The quality control measures described ensure constant quality products, and a good recovery of high grade material from the deposit.

Jan 1, 1971

The mIning operations of Bougainville Copper Pty. Limited are located at Panguna on the Island of Bougainville which is part of the Territory of Papua and New Guinea.C.R.A. Exploration Pty. Limited first established mineralization in the Panguna area in 1964. Exploration work since then has established ore reserves estimated to be 900,000,000 long tons of ore assaying 0·48 per cent copper and 0·36 dwt/ton gold. An economic feasibility study completed in July, 1969 enabled a decision to be made to proceed with a major mining operation. The ore will be mined by open cut methods and will be treated in a conventional concentrator with a nominal capacity of 90,000 short tons/day. The concentrate will be pnmped as a slurry through a 16 mile long pipeline to a port established on the east coast where it will be filtered and dried ready for export. Other supporting facilities for the mine include two towns, a 135 MW steam power station, and access roads.

Jan 1, 1971

By W. E. Roscoe

The probability of making an exploration discovery in Canada can be calculated from Derry's (1970) data on exploration expenditures and number of discoveries over the last nineteen years. Considering an average exploration "try" to be $30,000, the probability of a discovery with one try has decreased tenfold from about 0.01 in 1951 to 0.001 in 1969. A plot of these probabilities can be extrapolated to the future, giving a 1969-71 three-year average probability of 0.00096 with 90 per cent confidence limits of 0.00042 and 0.00153. Using this probability, various calculations can be made, such as (1) the probability of making at least one discovery for any given exploration expenditure, (2) the probability of ultimate ruin or survival of the exploration investment process and (3) the expected value of an exploration investment. There is a critical minimum amount per discovery that must be re-invested in exploration, no matter how large the original investment is, for exploration to survive at all in the long run. Currently, this critical minimum is $31,200,000 with 90 per cent confidence limits of $19,600,000 and $71,400,000.

Jan 1, 1971

By N. R. Paterson

"More than a dozen AEM systems are currently being used in the routine exploration for massive sulphides. The systems differ from one another in minor respects, but fall into four main categories: rigid-coupled in-phase/quadrature systems; towed-bird, single-component systems; differential systems; and passive systems.Although field tests are important in demonstrating the performance characteristics of the AEM/airframe assembly (noise-levels, coupling effects, etc.), more meaningful evaluations can be made through physical or computer modelling. The latter approach has been used to compare the response of twelve AEM systems to both vertical and horizontal sheets, simulating a large orebody and a con-ductive overburden respectively. The effects of height changes are also studied.The results demonstrate that each AEM system has its own particular advantages and disadvantages. The relative importance of these is a matter for the geophysicist to judge, on the basis of specific geological and environ-mental information. Data currently available on AEM systems are adequate to permit an objective evaluation of the systems for a particular exploration program."

Jan 1, 1971

By Joseph Yancik

El presente texto señala el método de evaluación que determine el sistema óptimo de los explosivos de una voladura. Los componentes básicos para una voladura óptima son: características de la roca, planificación de la mina, preparación, manipuleo, equipo de procesado y sistema de explosivos. Es preciso anotar que el punto óptimo de una voladura no está constituído por un grupo estático de condiciones. Si consideramos que los costos de perforación y disparo representan una parte significativa del costo total de la operación de minado, no es posible ignorar avances tecnológicos que ofrecen substanciales reducciones en el costo. Em este sentido, el sistema de potencia de campo, se plantea como el sistema adecuado para la predicción rendimiento de un explosivo a través de la determinación de su potencia, "in situ" bajo condiciones indicadas.

Jan 1, 1971

By John V. Beall

On January 1, 1948, we boarded the cage at 7 a.m. and were dropped to the 900 level of the Brunswick shaft. The previous evening, we had attended the New Year's party at the Miners' Guild Hall and as this soireée had not disbanded until the small hours of the morning, we had only had about two to three hours sleep. At the 900 level station, we fell in beside our mining partner, a Cousin Jack, who was known by all as Porky. Porky didn't say anything as we plodded down the drift, we didn't say anything either-in fact there was little conversation by anybody.

Jan 1, 1971

By Judith A. Eisele

The possibility of improving the volatilization of gold through the formation of a metal-gold chloride vapor complex was investigated. Gold-silica mixtures, containing ferrous and nonferrous minerals and salts as additives, were treated in a laboratory tube furnace with a flow of chlorine gas at various temperatures. The addition of pyrite (FeS2) to a charge of gold-silica increased the volatility of the gold from 3, 20, and 62 percent of the total gold to 62, 93, and 100 percent at 200°, 300°, and 400° C, respectively. Chemical analyses of the condensate indicated the volatile species to be FeAuCl6.

Jan 1, 1971

By John V. Beall

As Mother Nature arranged it, the hottest potential mineral target in the country today, the Stillwater complex had to occur in a spectacular natural setting-the north margin of the Bear Tooth Mountains of Montana. What more appropriate backdrop for the August 17-18 field hearings of Senator Frank E. Moss (D-Utah). His subcommittee, on Minerals. Materials and Fuels, heard 37 witnesses ponder the value of minerals vs esthetics or other uses of federal lands. On August 17, four Army Hueys swept the senatorial party, the press, Forest Service Officers and such distinguished witnesses as Montana Governor Forrest H. Anderson over the Custer-Gallatin National Forests, besting the thin air of 10,000 ft p1us, they set down on Iron Plateau.

Jan 1, 1971

By R. Mulligan

"The Cordilleran tin belt extends from the Seward Peninsula in westernmost Alaska to Mexico. Throughout the Canadian section, tin occurrences are mainly confined to a belt along the northeastern fringe of the Western Cordilleran sub-region; the boundary of this sub-region with the Eastern Cordilleran sub-region is most simply defined as the practical eastern limit of granitic and metamorphic rocks. The stratified rocks underlying this restricted belt are mainly late Precambrian and early Paleozoic miogeosynclinal sedimentary rocks, which are more or less metamorphosed but otherwise are generally equivalent to the less severely deformed rocks underlying the Eastern Cordilleran sub-region. They contrast with Late Paleozoic - Mesozoic eugeosynclinal volcanic-sedimentary assemblages farther west, but the granitic intrusions throughout are mainly of Late Mesozoic age. Thus, narrowly defined, the tin belt largely coincides with a lead-zinc-silver metallogenic province, whereas copper deposits characterize the eugeosynclinal regions to the west. In addition to the tin occurrences, which include the Sullivan lead-zinc-silver mine, Canada's only significant producer, the belt contains all the beryllium, most of the tungsten and many Iithophile-type molybdenum occurrences. Several types of occurrences of each metal are represented. Lithium occurrences in the Canadian Cordillera are notably rare, and these few are insignificant."

Jan 1, 1971

By P. C. Kapur

An approximate but simple and practical solution to the discretized batch grinding equation has been obtained. The solution is based on a new grind function, a combination of selection and breakage parameters and particle-size distribution of the feed. The grind function, which is readily computed from limited experimental data, has significant implications for analysis of the particle-size reduction processes. Using the practical solution, three ball-mill comminution systems—dry dolomite, wet quartz, and limestone—have been simulated with acceptable precision. The application of the discretized batch-grinding equation1 to routine comminution analysis is frustrated by two principal difficulties. First, the selection and breakage parameters of the grinding equation must be computed separately. This is usually a difficult and uncertain task although, recently, a number of experimentally and computationally efficient strategies have been proposed for determination of these parameters?" Second, for actual computation, the solution to the batch-grinding equation becomes progressively tedious and unwieldly for finer particle-size range.1,2 The purpose of this part of the report is to present an approximate solution to the batch-grinding equation, which is practical in that it seems to overcome the aforementioned obstacles. The solution involves the concepts of a new grind function, a lumped parameter of the selection-breakage functions, and particle-size distribution of the feed, which is readily computed from a limited amount of experimental data.

Jan 1, 1971

By R. W. Wein, L C. Bliss

Northern ecosystems are dominated by a harsh physical environment which dictates a need for biological adaptation. As yet, it is unknown just how these adapted plants and animals will tolerate industrial development. Our work, sponsored by university, government and industry, is part of an effort to evaluate this developmental impact. Our present studies and the results to date will be discussed. Other problems to which the environmental student may address himself in the future to ensure against environ-mental degradation will be suggested.

Jan 1, 1971

By George J. Simchuk

The selection of a stripping method and subsequently, the stripping equipment, are two of the most important decisions confronting a project engineering staff during the initial feasibility study of a new coal stripping operation. The selected stripping units average from 60 to 70 percent of the total capital outlay for mine equipment and their selection will affect the type, size, and quantity of the attendant equipment, such as loading shovels, drills, haulage trucks, and bulldozers. This paper is intended to report on an in depth study by Utah's Technical Services Staff regarding the selection of stripping equipment for the Goonyella Coal Mine in Queensland, Australia. We will touch briefly, on the objectives considered in this selection; the various alternatives which were explored; the general assumptions that were followed throughout the engineering and design of an appropriate stripping system; and computations necessary to determine the amount and size of stripping equipment. INITIAL EXPLORATION Utah Development Company (presently 90% owned by Utah Construction & Mining Co. and 10% by Australian investors) commenced systematic geological studies of the Bowen Basin coal fields of Central Queensland in 1961. The Blackwater Mine as shown in Figure 1 has since been established in this area and is now producing 3.0 million tons per annum of washed coking coal. Geological reconnaissance was then extended to the northern part of the Bowen Basin and in 1963 Utah's coal prospecting area covered 2444 square miles of Central Queensland. The authority to prospect north of Blackwater has since been trimmed to 810 square miles by relinquishments to the State.

Jan 1, 1971

By G. E. Rouse

Anomalies of sulfur dioxide and other gases have been defined, by direct measurement, around various sulfide bodies. These readings were made (1) in soil gas and (2) in the atmosphere, from both ground vehicles and, to a limited extent, from helicopters. Test areas included porphyry copper and molybdenum deposits, sulfide bearing vein systems, pyrite zones associated with sandstone uranium accumulations, and pyrite-bearing mine dumps. Some of the deposits are covered by as much as several hundred feet of overburden or barren rock, while others outcrop, or have been exposed by mining operations. Depending on local conditions, background values of sulfur dioxide in soil gas and the atmosphere range from 1 to 15 parts per billion. Using present techniques, anomalous readings are recognized at approximately two times background. Apparent "depth penetration? on known deposits has ranged to several hundred feet. Readings have been made over known deposits from a helicopter and ground vehicle moving at up to 40 mph. The rate at which the survey may be carried out depends largely on the size and extent of the anomaly. The existence and magnitude of sulfur dioxide anomalies around sulfide-bearing deposits is a function of geology, target size, soil moisture, and climatic conditions such as temperature and temperature changes, barometric pressure, wind velocity, and precipitation. Industrial S02 pollution can effectively mask the presence of the anomalies sought.

Jan 1, 1971

By D. M. Hendrick, D. K. Fountain

Since the discovery by conventional prospecting of the Horne deposit in 1920, the Noranda area has produced 82 million tons of copper-zinc-gold-silver ore from sixteen massive sulphide bodies. Recent exploration in the area has consisted of systematic studies employing geology, geophysics and geochemistry. Due to the depth of the deposits being sought, surface geo-physical surveys are limited in their effectiveness as DIRECT indicators. However, because of the geological environment, surface geophysical methods, and the Induced Polarization method in particular, can be success-fully employed as INDIRECT indicators of ore deposits. Examples of survey results from various properties in the Noranda area illustrate the advantages and limitations of the Induced Polarization method as an exploration tool in this area.

Jan 1, 1971

By G. W. Smith

"The adsorption of dehydroabietylamine acetate on quartz, hematite, rutile and baddeleyite has been deter-mined for concentrations ranging from 0.3 to 5,600 µ. moles/I and for pH values ranging from 2.5 to 10.5.The surface tension, UV and IR spectra, equivalent conductance and ion concentrations for dehydroabiety lamine acetate solutions have been determined.Contact-angle and flotability studies have established that good collection may be obtained with this amine but that selectivity depends on the suitable use of depressants for some of the minerals."

Jan 1, 1971

By S. W. Nabbs, W. D. Bachman, A. W. Last

A study has been made of wet, autogenous grinding of disseminated copper ores, including testing of a large number of samples from. Kennecott Copper Corp.'s Chino mine. The grindability of the various samples was found to vary over an extremely wide range, even among samples taken from diflerent locations within the same ore body. These variations in grindability can be qualitatively correlated with the composition of the rock and the fracturing and alteration that had occurred. Because of the large-scale nonhomogeneity of disseminated copper ore bodies, extensive and careful testing of autogenous grinding is needed. An accurate correlation of the grinding data from various samples with the geology of the ore body and with mining plans will indicate if variations in grinding mill throughput can be avoided in practice by mixing or blending of mill feed and if, therefore, autogenous grinding can be applied successfully. In the past 15 years, comminution by primary autogenous grinding has been investigated and applied economically in the minerals industry, particularly in iron-ore processing. Successful application of primary grinding is described by A. A. Dor,l.' P. B. Dettmer,334 R. R. Turner, and W. F. McDermott." other authors describing the advantages of primary grinding include H. Hardinge,' R. T. Hukki,' D. S. Coyle" and F. C. Bond,"," to mention only a few. For the past several years, Kennecott Copper Corp.'s Metal Mining Div. Research Center has been investigating primary grinding of disseminated copper ores. Both fully autogenous and semi-autogenous grinding has been explored. Ores from three of Kennecott's western operating properties and from several development properties have been tested. Host rocks include diabase, schist, silicated and argillaceous limestones, shale, granite porphyry, quartzite, and granodiorite. The object of this paper is to supplement the general knowledge in the field of primary grinding by presenting information gained from the studies with the various disseminated copper ores. Data are presented which show that carefully considered test programs are required to determine the practicality of applying autogenous grinding to a specific mining-beneficiation operation. Particularly this is so if more than one rock type occurs within an ore body or if there are wide variations in the degree of rock alteration within a deposit. Extensive sampling and subsequent testing of the ores within a mine are required if serious pitfalls are to be avoided and good engineering data acquired. We hope that this presentation will stimulate studies to increase the understanding of the fracturing process applicable to large pieces of rock, a factor which is important in autogenous grinding. Test Facilities and Procedures Both pilot-scale and full-scale test facilities were used to investigate primary grinding characteristics of the disseminated copper ores. A 6-ft Hardinge Cascade mill, a 24-ft Hardinge Cascade mill, and a 5-ft Aerofall mill were used in the studies. The Cascade mills were operated wet, while the Aerofall mill was operated both wet and dry. The Cascade mills were operated both for fully autogenous and for semi-autogenous grinding; the Aerofall mill was operated only semi-autogenously. Fully autogenous grinding, as the term is used in this presentation, is defined as the direct grinding of coarsely crushed rock by the action of the rock mass. Semi-autogenous grinding is defined as the grinding of coarsely crushed rock with the rock mass supplemented by a steel ball charge. Most of the test work was done using the 6-ft Cascade mill, operated under fully autogenous grinding conditions. Only limited testing was conducted with the 5-ft Aerofall mill and only one ore was tested using the 24-ft Cascade mill. This presentation is concerned, therefore, primarily with fully autogenous, pilot-scale grinding in the Cascade mill. Fig. 1 presents a diagram of the test facilities used with the 6-ft Cascade mill. Principal test equipment consists of a 6-ft Cascade mill, an 18-in. spiral classifier, an elevator, a 2 x 9-ft belt feeder, a conventional integrating kilowatt-hour meter, and a recording wattmeter. In testing, feed charges, adjusted as to weight, were deposited as evenly as possible on the belt feeder every 15 min. The weight of each charge was adjusted to maintain a mill charge of 27 to 30% of mill volume. The classifier was adjusted to obtain the desired size-consist of the final product, generally at 20% + 100 mesh. After equilibrium was established in the circuit, the feed, mill discharge, classifier sand, and classifier overflow were sampled and their size distributions determined. During the sampling periods, the power drawn by the mill was measured to determine the energy consumed in the reduction of the size consist of the ore solids. Fully Autogenous Pilot Grinding Tests on Various Kennecott Ores Typical test results obtained in the 6-ft Cascade mill with samples from Kennecott's Ray Mines Div., Chino Mines Div., Utah Copper Div., and a development property designated as Mine A are presented in Table 1. As shown, the ores that were ground most easily in the pilot-scale mill were composite mine samples obtained

Jan 1, 1971

By Basil W. Mercer

Increased emphasis on industrial pollution control has intensified investigation of waste treatment technology for methods of reducing the amount of pollutants discharged by industry. The mining and mineral processing industry can point to a record of considerable accomplishment in the field of wastewater treatment and reclamation because these operations are so frequently an integral part of the entire operation of an ore dressing facility. This is amply demonstrated by metal producers in the arid regions of the Southwest where water shortages require maximum use of available water. For example, the copper industry in Arizona during 1960 employed an estimated 184 billion gallons of water of which 87% was reused water. (1) Only 0.0003% was discharged to streams and the remainder was consumed by evaporation (8%) and seepage (5%). The most severe pollution problem facing the mining and ore processing industry today is that resulting from acid mine drainage. This problem is largely associated with the mining of coal although it occurs to a lesser degree with the mining of other materials. The degradation of streams by acid mine drainage in the vicinity' of the Appalachian coal fields has been widely publicized and a major effort is underway to determine the best means of correcting this situation. Other forms of pollution from mining operations are less well-known, but it is anticipated that improved or advanced waste treatment techniques will be required for segments of the industry to comply with new pollution control regulations. The current emphasis on redefining hazardous levels for heavy metals and various chemicals with regard to the total environment will in all likelihood lead to tighter discharge controls in some cases. It is the purpose of this paper to review some of the waste

Jan 1, 1971

By Lewis M. McNay

The Bureau of Mines located and examined 292 burning coal refuse banks throughout the Nation's coal-producing regions in 1968. These coal waste fires, extending over 3,200 acres, produce poisonous, acrid gases and particulate matter that pose a threat to health and safety in the surrounding areas, damage vegetation, and cause the deterioration of nearby structures and buildings. Methods of extinguishing coal refuse fires are discussed, and an approach to eliminating future occurrences of these fires is presented.

Jan 1, 1971

The ore deposits investigated are composed primarily of hcmatite and can be grouped into three categories depending on the amount of leaching of silica and carbonate which has occurred from the original iron formation, and on the amount of recrystallization of the hematite. Minor constituents are magnetite. goethite, and quartz.Texturally. there are five different types of hematite, formed at different stages during the genesis of the ores.These are1. anhedral hematite grains formed by recrystallization of hematite already present in the original iron formation.2. hematite formed by martitization of euhedral magnetite, 3. massive aggregates of fine- to medium-grain hematite formed by recrystallization of (l) and (2).4. tabular or prismatic hematite grains occurring around the boundaries of the massive aggregates or martite euhedra. and in fractures and5. specular hematite. which is present in fractures in ore types (l) to (4).Minor and trace-element analyses, determined using an electron-probe microanalyser. indicated that only silica (less than 0·01 per cent) and cobalt (less than 0·002 per cent) were present in the magnetite and hematite from unaltered iron formation and in the hematite from the ores. Low concentrations of trace elements are typical of magnetite in ores formed as marine precipitates. On the basis of the analyses alone. it proved impossible to distinguish between the five hematite types. A determination of the variation in rare-earth trace element contents between iron oxides in the iron formation and iron oxides in the ores indicates that the rare earths were depleted during ore formation. A possible explanation is that they were removed in solution with silica as it was leached from the iron formation.

Jan 1, 1971

By William T. Barrett, James E. Lawver, James L. Wright

One of the problems facing both the designer and the user of magnetic separators is the evaluation of the body forces acting on particles in the magnetic field of the separator. Perhaps the most elegant method of evaluating the body force associated with a drum-type magnetic separator is that recently suggested by W. F. Brown.' Brown measures the radial component of the field intensity 1/2 in. from the drum surface and fits a Fourier series in the angle about the drum axis (Fig. 1). Once the coefficients of the series are known, it is then possible to calculate the values of the radial and angular

Jan 1, 1971