Search Documents

By Vladimir J. Ovdenko, Sergey N. Lazarenko, Evgeny L. Schastlivtsev

The term sustainable development shortly appeared in international use in the run of UNO Environmental Protection Conference held in Rio de Janeim in 1992, bears the concept, that evolution process ( associated either with a region or society) should provide mutual interests concern in three aspects - environmental, economic and social ones, admitting as well the fact , that environmental protection should become an integral part of each process and cannot be regarded separately. In this connection it seems quite reasonable to employ the principal of sustainable development, while creating the model of industrial mastering the territory constraint lands viewing their future, for a example, in rich mineral deposit's development. In Kuznetsk coal basin in Russia ( west Siberia) the development of a big coal deposit Erunakovsky is now being underway. The major part of coal will be mined with open cast method, resulting the number of local environmental adverse impacts' occurrence. In such a situation mastering of sustainable development model is given the top priority, demonstrating the essential result of its application in a form of possibility to define the best correlation between economically expedient and environmentally tolerable loads during coal mining. The given paper is an attempt to work out the main principals to approach the creation of sustainable development model in regard of the definite region case. To simplify the presentation of employed approach's main idea, the given below accepts as its component only one of many actually existing environmental profiles, that define mining pro- cess impact on environmental situation -the size of infringed (de- graded) lithosphere surface. Basically, the actually existed correlation between spheres. grounding the model of sustainable development - economic, environmental and social- is given in a form of nomogram under Fig.1. On the presented graph the prognostic values of broken lands relative extent within Erunakovsky boundary ( to be compared with total region area taken as 100%) are arranged on axis x". Axis "y- demonstrates the main values associated with complete economic background, making the intrinsic part of the principal for sustainable development, i.e. annual coal output for the given area. It's top, used in assessment value relates to actual showings implied for the study region project and quantifies 60 min. tons. While regarding the social facet of principal for sustainable development, the applied mode to separate social issues into two types - one profiling a welfare standard of living and the other. pertaining the way environmental situation contributes to the degree of &ologica-1 comfort in people's living was defined as the most expedient and the most adequate in the matter of reflecting the actual terms of vital activity of the society. In accordance with the cited above the economic level showings are plotted on the graph on "-x” axis, while "-y" one obtains the values associated with the point of ecological comfort presented in conventional units. For the maximum possible conventional point of living (corresponding in nomogram to 1.0 value) a welfare level secured with industrial region economic activity revenue was assumed in case the top planned annual amount of coal is delivered ( 60 min. tons). While studying the index of public economic welfare, environmental conditions corresponding level in its original state i.e. prior to any industrial work carried in the region, was adopted as its top possible conventional rating (1.0). The given nomogram displays the integrity of x-f(y) type interrelations plotted on the basis of four separate dependencies, precisely a) the size (area) of infringed earth's surface as a function of coal output in the region; b) subordination of ecological comfort of people's living to the degree of lands infringement; c) interconnection of economic and environmental components of people living; d) interconnection of economic level of people's living and the amount of coal mined in the region. Linked to a single unit, these separate plain dependencies create, as seen from the graph, a more intricate system of bonds, that affords during defining or changing this or that system parameter to secure undergoing changes, that occur to other parameters of the system. The mentioned reason made it possible to consider and drive to mutual conformity ( as far as it is possible) all parameters values for the defined system's relations, that in their turn indicate the concern of other aspects of a sustainable development principal. According to the graph, in order to get not more than 50% of infringed land left after the deposit's mining, it is advisable to ex- tract not more than 40 min. tons of coal annually, gaining in this case the level of ecological comfort for living in the amount of not

Jan 1, 1996

By D. Robert Weir, Verner B. Sefton

The reduction roast-ammonium carbonate leach-hydrogen reduction process for laterites has been under development by Sherritt Gordon for the past 20 years. The process can treat high iron or low iron laterites, with some process modification, and gives extractions of 88-93% Ni or 82-88% Ni, respectively. Five campaigns on a 1 tonne per day Pilot Plant and four campaigns on a 35 t/d Demonstration Plant have been carried out on laterite from three different deposits, and many other laterites have been investigated in the laboratory. The Marinduque Surigao Nickel Refinery in the Philippines was designed on the basis of this work. The process comprises roasting to selectively reduce the nickel and cobalt in the laterite, leaching the calcine in ammonium carbonate to dissolve the metals, recovery of cobalt by precipitation of a mixed metal sulphide, concentration of nickel from the carbonate solution by basic nickel carbonate precipitation and recovery of high purity nickel by hydrogen reduction of nickel from ammonium sulphate solution. Ion exchange has been developed as an alternative to BNC and for recovery of trace nickel from BNC barren liquor. Details of the development of some of the process steps through the laboratory to commercial plant are discussed. The process metallurgy has been confirmed in the commercial plant.

Jan 1, 1979

By J. F. Power

Returning the original soil material to the surface of smoothed mine spoils is n practical means of restoring agricultural productivity. Research has established that high-sodium spoils in North Dakota (a) must be covered with about 70 cm of soil material to achieve maximum production; (b) topsoil must be segregated and spread separately from subsoil; (c) potential production depends on the quality of the soil material returned; and (d) erosion and upward salt movement must be controlled for sustained production. Erosion is best controlled by reducing final grades as much as possible during the smoothing process followed by prompt establishment of vegetation. Almost half the nation's cool reserves are in the Northern Great Plains underlying several million hectares of agricultural land. Social pressures to return mined land to agriculture or other productive uses have resulted in recent enactment of state and federal legislation, which is intended to restore productivity to disturbed areas. All such legislation relating to the Northern Great Plains has stipulated that useable soil material be saved and returned to mined land after spoils are smoothed. This requirement immediately raises questions as to how much soil material is needed, what criteria are important in defining useable soil material, how is the soil material to be handled in the earth-moving operations involved, and how can replaced soil material be maintained and revegetation. Scientists at the USDA Northern Great Plains Research Center at Mandan, North Dakota have conducted research on the reclamation of disturbed lands since 1970. Initially this research was designed to identify and characterize those physical and chemical properties of spoils that are most restrictive to plant growth (1). Basically, they found that any properties of spoils that reduced the efficiency of storage and utilization of the limited precipitation received adversely affected plant growth. Thus, in the Northern Plains, revegetation of mined land was found to depend upon the management of factors affecting the availability of water to plants - the same as on unmined land. After the major problems in revegetation were identified, technologies were developed to circumvent the effects of the growth-limiting factors (2). Basically two approaches were identified by which plant growth-restricting properties of spoils could be corrected: (a) various amendments (gypsum, sulfur, Fertilizers, mulches, calcium chloride, etc.) could be applied to spoils to improve their ability to conserve precipitation and provide water for plant growth; or (b) undesirable spoils could be covered with suitable soil materials to reduce or alleviate adverse effects. The latter approach appeared to be more feasible in most instances And is the approach specified in current legislation. The remainder of this paper discusses research on the use of roil materials returned to spoils and summarizes the information obtained.

Jan 1, 1978

By R. R. Klimpel

The industrial processing of ores containing precious metals such as gold is generally very mining site specific, e. g. McQuiston and Shoemaker (1985), vanZyl et. a1. (1988), Fuerstenau and Hendrix (1990), and Hausen et. a1. (1990). The variety of various mineral processing techniques employed is extensive and includes size reduction, size classification, gravity separation, roasting, cyanidation, chemical and/or bacterial leaching, ion exchange and/or solvent extraction, and froth flotation. Gold occurs in the earth's crust on the average in very low concentration estimated by various sources to be between two and five parts per billion. For a specific ore deposit to be considered as an economic source of gold, generally a concentration of near one part-per million (one gram / Mton) or more is required. From a mineralogical viewpoint, gold often occurs in its native or free form. However, it can also occur as gold compounds of telluride, antimony, and selenium contained within either sulfide minerals such as pyrite, arsenopyrite. and pyrrhotite or, to a less extent, within silicate, oxide, and carbonate minerals. In many locations, alloying of gold with silver is found. The optimal processing method for recovering gold is determined by many factors such as the mineralogy of the gold bearing minerals as well as the associated mineralogy of the gangue minerals, the type of liberation profile of the gold bearing minerals upon' size reduction, and the particle size of the gold itself. Free or native gold particles can vary in size from the extreme of large nuggets to very finely disseminated gold associated in a complex sulfide mineral matrix. Typically free gold particles of several hundred microns or larger are efficiently recovered by one or more of a number of gravity based techniques due to the large specific gravity differential of gold (19.3 to 15) versus typical gangue minerals (5 to 2.7). When gold is associated with a sulfide (usually pyrite) matrix, historically most industrial plants involving size reduction and subsequent liberation have operated in a size range of 50 to 80 percent less than 75 microns. Grinding finer than this has been normally not economical unless the ore has a high gold content in the feed ore.

Jan 1, 1998

By Kirsten Corin, Belinda McFadzean, Cyril O’Connor, Jenny Wiese

The Bushveld Complex of South Africa contains almost 90% of the world’s reserves of platinum group minerals (PGMs). In the flotation of PGMs, there are significant challenges arising from the need to treat ever-decreasing grades of the relevant ore deposits. The major challenge in the flotation of these ore bodies is the control and management of the gangue minerals, particularly silicates such as orthopyroxene, plagioclase feldspar, and pyroxene which are often rimmed with talc which makes them naturally floatable. It has been shown that various polysaccharide depressants such as CMC and guar have different properties in terms of depressing the gangue minerals. Since the PGMs are often associated with sulphides, copper sulphate is widely used as an activator in PGM flotation but can inadvertently activate the gangue minerals as well as reduce the recovery of PtTe2 which accounts for up to 40% of the Pt in the Platreef ore body. Depressants also reduce the mass of solids reporting to the froth and can thus destabilise the froth. This effect on the froth can be mitigated by using higher frother dosages or water of higher ionic strength. Reducing chromite recovery is of critical importance since high levels may negatively affect the downstream smelting process. Chromite recoveries can however be reduced through the application of gravity separation or reducing entrainment through reduced water recovery. In summary, due care needs to be taken to carry out site test work to develop an optimum ratio of collector, frother, activator, and depressant to ensure that the highest grades and recoveries of the PGEs are obtained while reducing depressant dosage as much as possible.

By J. Kennedy

Rare Earths: Understanding The United States complete failure of status and relevance in Rare Earth Oxides, Elements and Alloys. Although rare earth oxides are strategic and critical in their own right, elements and alloys are even more important for the development of "Green Technologies" and advanced weapons systems. America does not have an active domestic source of rare earth oxides. America does not have any Heavy Rare Earth Elemental or Alloy capabilities. The magnitude of this failure continues to become more apparent as the manufacture of REO based electronics and magnetic components moves eastward towards China. China has succeeded in outmaneuvering The United States and the world in its quest to control the Rare Earth Oxides, Elements and Alloys, and the many industries dependent upon REO. What can be done?

Jan 1, 2010

By John E. Tilton

Mining is one of the oldest human activities. It goes back at least to the Bronze Age and possibly even the Stone Age. The Romans mined copper in Spain and tin in England some 2,000 years ago. The writings of Agricola in the 16th century show that mining was important in the Middle Ages. Today, more mineral commodities are mined in greater tonnages than ever before. This long and fascinating history closely tied with the economic development of the human race is one of the charms of the mining industry. It also, unfortunately, gives rise to some widespread misconceptions. Old industries are generally considered stodgy with mature and stagnant technologies. Costs tend not to decline as in newer industries, and may even increase with real wages. Costs are particularly expected to rise in the case of mining as the best mineral deposits are depleted. Thus, mining is often thought of as a sunset industry, destined to decline over time as the resource base is depleted. These perceptions have several important implications. They argue against economic policies that rely too heavily on mining and mineral processing to foster growth and development. The best and the brightest young people, along with society's other valuable resources, should be encouraged to enter newer industries. There the technological opportunities and the prospects for growth are more promising. These perceptions also suggest that the ability to compete in the mining industry depends largely on the quality of a company's or country's mineral deposits. Mining will take place first in countries with the lowest-cost deposits. Once these are exhausted, mining will shift to countries with the next best deposits. From time to time, new discoveries may also produce a shift in the location of mining, [ ] and may even for a time keep the costs and prices of mineral products from rising. Over the longer term, however, rising costs and prices are inevitable, as the world is forced to exploit poorer mineral deposits.

Jan 1, 1997

By R. R. Klimpel

Over the last six years, several new commercial families of sulfide mineral collectors have been developed. This paper summarizes some plant data and use experiences gained with each of these product families. The first is a series of products (F-series) based on a chelation mechanism of attachment. The second of these families is a series of products (S-series) based on an electrochemically controlled mechanism of attachment. The major advantage of the F-series collectors is the very high metal recovery capability in complex ores involving Au, Pt, Zn, Ni, and mixed Cu minerals. The major advantages of the S-series collectors are the very high pyrite rejection capability along with reduced lime requirements in Cu/Mo plants.

Jan 1, 1995

By William R. Oakley

In order to best-discuss the met spar industry let me take the liberty to review our involvement, that is, my company's. Our organization has been involved in the fluorspar industry for approximately thirty-five years. We started our activity in Western Kentucky as a mining company and a beneficiator of spar, that is, grinding and grading. After a few short years, the mine became contaminated and rather than invest in a flotation plant, the property was sold and our spar activity was moved, primarily to Mexico. It was decided not to establish ourselves as a Mexican corporation, so, it was necessary to take the position of backer-financier of numerous mining operations. We served in this capacity for quite some time. We also built processing plants on the border in Texas, that is, sizing, grading, etc. in order to market into the steel industry. In 1957, generally speaking, fluorspar for the steel industry was purchased as gravel fluorspar with a maximum of 10% -1/16". The price of this material was $23.50 to $24.00 per ton f.o.b. railcar at the Rio Grande Crossings and about $26.00 or $26.25 f.o.b. barge, Brownsville, Texas. Usually the material was purchased on an effective basis. As you know, 2-1/2 times the silica total, subtracted from the total CaF2 gives the effective calcium fluoride. This was normally the extent of the chemical specifications. Consequently, material was shipped with no regard whatever, to anything other than effective units. The variations in material was absolutely unbelievable. In addition to this, quite frequently, cars were shipped, as they say in the industry, topped. This being, good material placed in the top of cars or barges over bad or low grade. The entire shipment may very well have met the effective unit spec, but the complete imbalance of the sub grade vs. the high grade more or less guaranteed that the consumer was probably using a great deal of low grade and therefore, his usage had to be way higher than it should have been. He came to the conclusion, that because of these conditions and the new technology developing in the steel industry, it was just a matter of time that the entire raw material industry would have to improve. Rather than spend our efforts directed to a substitute flux, we put together a sizing and grading plant in order to ship a controlled crude product, with independent laboratory analysis confirming size and quality. The age of control was started. We began shipping 70 to 72-1/2, 80 and 85 effective with sizing 1/2" by 2", 1" by 3", and 2" by 5" with absolute control on the amount of fines. A material with no fines was a premium product. Incidentally it sold for $24.50 -60 effective up to $28.50, $29.50 for 80 effective.

Jan 1, 1978

By Shangyong Lin, Wei Sun, Weiping Liu, Muhammad Azam, Junfeng Chen, Yujie Zhao, David Asubonteng, Michael Ngoie

Copper concentrates are widely used in smelting operations worldwide and typically consist of chalcopyrite ( CuFeS2) as the primary mineral, with chalcocite ( Cu2S) and bornite ( Cu5FeS4) as secondary copper-bearing minerals. However, these concentrates often contain hazardous impurities, such as the fluorine ion, which can result in various complications during the smelting process. These issues include equipment corrosion, gas treatment requirements, environmental hazards, and strained community relations. Furthermore, high fluorine content in copper concentrates can lead to penalties and limitations during sales. To address these challenges, this comprehensive review assesses current techniques for removing fluorine from copper concentrate. Drawing upon recent research and the findings in the field, the review analyzes the advantages and disadvantages of existing fluorine removal methods and identifies critical issues requiring attention in the future. This review provides valuable insights into the current state of fluorine removal from copper concentrates and serves as a practical guide for researchers and industry professionals in this field.

Aug 2, 2023

By Robert B. Crookston

This paper presents a resume and status report: of mining technology appropriate to commercial oil shale development in the United States today. The nature and extent of oil shale in Colorado and Utah are discussed. A look at most of the prominent activities now underway emphasizes the questions "where are we now?" and "where are we going?" Various alternate mining methods are reviewed and a bibliography of oil shale mining is included.

Jan 1, 1980

By M. Sadegh Safarzadeh, Erich U. Petersen, Jan D. Miller

"An acid bake-leach process has been identified and applied to an enargite concentrate containing ~67% enargite, ~27% pyrite (FeS2), ~3% chalcopyrite (CuFeS2), ~6% sphalerite (ZnS) and <1% galena (PbS). Sulfuric acid baking of this concentrate resulted in the transformation of the concentrate to S8, As2O3, HFe(SO4)2•4H2O, PbSO4 and Cu0.47Fe0.53SO4•H2O, all of which (except elemental sulfur and PbSO4) dissolved during the water leaching step. More than 87% Cu and 64% As were leached when the concentrate was baked at 200°C over 7 hours with a sulfuric acid-toconcentrate weight ratio of 1.6. Less than 1% As was lost to the gas phase. Upon water leaching, copper and arsenic were released to the solution at a concentration ratio of about 3:1. While enargite is very reactive during the sulfuric acid baking reaction, other minerals present in the concentrate such as pyrite, galena and sphalerite may compete to consume sulfuric acid and be converted into their corresponding sulfates at different concentrations. The purpose of this paper is to present our examination of the effects of sulfuric acid dosage on the extractions of Cu, Fe, and As from an enargite concentrate. In this regard, the concentrate, baked concentrate, and the leach residues are characterized and a conceptual process flowsheet is described for the acid bake-leach process for the treatment of enargite concentrates.IntroductionThe occurrence of enargite (Cu3AsS4) in copper ores and concentrates and the associated high arsenic values in such resources have resulted in a considerable decrease in the demand for such concentrates because they cannot be economically processed using conventional technologies such as smelting and atmospheric leaching (Mihajlovic et al., 2011). The exploitation of such problematic deposits through competitive treatment strategies will be required in the near future in order to satisfy increasing pressure from the copper market. It is expected that the average arsenic grade of copper concentrates will increase from 0.17% in 2008 to 0.22% by 2020 (Mayhew et al., 2011).The acid bake-leach process was identified and selected for further investigation at the University of Utah as a possible process alternative for the treatment of enargite concentrates. In this case, adequate amounts of concentrated sulfuric acid are mixed with the concentrate and heated to the temperature range 150-300°C. Subsequently, the baked material is leached with water/weak acid and the acid bake-leach process has some advantages over conventional sulfuric acid leaching (Prater et al., 1970). During water leaching, metal sulfates are released in solution, after which they can be extracted through well-known hydrometallurgical techniques. While the majority of the work has been done on oxide minerals, there is a limited record of sulfuric acid baking of sulfides."

Jan 1, 2015

By C. J. B. Dreyer

In terms of resources of amphibole asbestos only South Africa can be considered. There are only two asbestos producing fields, one in the Northern Cape Province producing only crocidolite and one in the Eastern Transvaal producing only amosite. The banded ironstone sequence in which amphibole asbestos occurs is discussed, and the genesis of the fibre is briefly reviewed. The properties and characteristics of amphibole asbestos are discussed and compared to those of chrysotile. Approximately 300,000 tonnes of amosite and crocidolite are produced annually. Crocidolite has a potential for a fivefold increase of production, which could be sustained beyond the year 2000. An estimate of the resources of the various producing fields in South Africa is made. Mining methods are reviewed and evaluation techniques discussed.

Jan 1, 1978

By K. Terhaar

The objective was to determine operating protocols for a lead and silver column flotation cleaner, being considered for installation at Sunshine Mining & Refining Co.&apos;s Kellogg mill in Northern Idaho. In this column, tetrahedrite and galena were floated, and pyrite and other gangue minerals, including siderite and silica, were depressed. The procedure was to develop a dynamic simulation using a linear transfer function approach. The transfer function related cause and effect variables such as the influence of collector and depressant dosages on recovery and grades. Bench-scale flotation tests followed by pilot column testing provided the data used to develop a linear transfer function. The transfer function was derived using a spreadsheet. This model was applied to simulate how the flotation column could be made to respond to disturbances in feed grade. The results were that xanthate, zinc sulfate, and sodium cyanide were found to be suitable controls for grade and recovery. Simulations showed the best sequence of manipulations of these reagents, and these sequences were useful for future operator training. The conclusion was that linear transfer functions were able to simulate lead and silver recovery, develop scenarios for operator training, and provide a foundation for future on-line control.

Jan 1, 1999

By Sahendra Ram, Ashok Kumar, Rakesh Kumar, AMIT KUMAR SINGH, Dheeraj Kumar, Arun Kumar Singh, Rajendra Singh

This article offers an examination of a retreat mining method conducted at the Pinoura Mine in India from 2010 to 2017. The method is atypical for coal mines of India’s coalfields. The method executes depillaring in a single pass and does not require the formation of galleries and support installation in splits, thus, reducing the cycle time. A variety of observations are discussed throughout this work and emphasis is placed on field measurements and depillaring under weak roof (RMR = 40–45). This paper discusses design techniques, specifically the estimation of the snook (remnant pillar) size, and important practical observations after the fact. The execution of this method was eventually conducted in eleven panels.

Aug 4, 2020

By D. J. Barratt

The application of autogenous and semi-autogenous grinding circuits in recent years has contributed toward substantial savings in capital and operating costs compared to conventional circuits, particularly for large scale copper and molybdenum operations. Such savings have been realized from the deletion of operating, maintenance, equipment purchase and substantial civil/structural costs associated with secondary-/tertiary crushing and screening, fine ore storage and conveying. In spite of the downturn in the copper- moly industry in North America, there is still an interest in autogenous and semi-autogenous grincting for copper-zinc, gold and silver ores and in the overseas market. Both single-stage and two-stage circuits are in operation. Primary circuits operate autogenously or semi-autogenously in either open or closed-circuit. Critical sizes are extracted and crushed for recycle in some cases (e.g. ABC circuit). Secondary grinding is conducted using ball mills or pebble mills.

Jan 1, 1985

By K. L. Boyd

"Can the ever-increasing tonnages of ore and waste being mined today be accommodated simply by using bigger and better crushers? This paper reviews the latest primary crushing and large cone crusher designs. It highlights how—in addition to using the larger crushers now available—maximizing the utilization percentages in each of the four stages of the rock size reduction circuit will assist in meeting the larger production requirements of existing and new mining operations. Its goal is to help consultants and operators select the optimal equipment and custom-designed items in the material handling flowsheet from the mine to the mill, maximizing operating hours per year. INTRODUCTION Are larger crushers the answer to meeting the demands of existing or new mining projects for increased tonnage? Boyd (2002) predicted that larger crushers for primary and fine crushing would be beneficial and contribute to the further development of mineral processing. Larger crushers for primary and fine crushing applications are now available. The rock size reduction process from the mine to the selected milling process facility has not changed since man first broke a rock to get at the interesting mineralization within. The rock size reduction process circuit from the mine to the mill can be summarized in four stages: • Stage 1: Mining (rock breakage and fragmentation) • Stage 2: Pre-screening (scalping) • Stage 3: Primary crushing • Stage 4: Fine crushing (secondary, tertiary and quaternary) Plant throughput and ore delivery schedule will form the basis for flowsheet design and equipment selection. Once the project design criteria and the mine plan have been established, the circuit’s utilization percentage (percentage of hours when the circuit is being utilized) can be optimized for each of the four stages. The mining and process departments are the key stakeholders in the selection of the final comminution circuit, and both departments should be familiar with and contribute to the design of the circuit. This paper focuses on the four stages of a rock size reduction circuit from the mine to the mill, equipment selection, various ore and waste handling flowsheets, and improvements in crusher station design, all with the purpose of maximizing availability and plant utilization percentages. What is a large tonnage? From 1907 to 1914, the American Panama Canal excavated 300 million tonnes of material in 7 years (History of the Panama Canal, Wikipedia). Today’s large mining operations can move the same material in 10 months, using large crushers or multiple crushers for ore processing and large crushers and a mixture of truck haulage and conveyors for handling the large waste tonnages. Some examples include:"

Jan 1, 2017

By James B. Fletcher

The Miami mine is located near the town of Miami, Gila County, Arizona. The Miami mine started mining in 1910 and finished in July, 1959. The orebody was divided and mined as follows: Tons High Grade 24,400,000 - Top slicing, sub-level caving, etc. Low Grade 105,100,000 - Block caving Mixed ore 9,800,000 - " Low Grade #2 13,100,000 - " Total Mining 152,400,000 GEOLOGY Ore minerals are largely in pre-Cambrian Pinal schist intruded by tertiary Schultz granite porphyry and covered to some extent by Quarternary Gila conglomerate. The structures are highly faulted and shattered. The Miami fault on the east cuts off the ore and the Pinto fault on the southwest caused re-oxidation of enriched sulphides producing mixed ore. MINERALOGY The chief mineral is chalcocite, with chalcopyrite, bornite, covellite, malachite, azurite, chrysocola, cuprite, native copper, and molybdenite as minor minerals. The ore minerals occur in seams, veinlets and disseminated particles. In block caving there is always some dilution, and drawing of ore stops when the grade drops past an economical point. This results in a small amount of copper left in the stopes below the capping. There is some copper in the capping, probably below 0.03% or less than 0.6 lbs. per ton. Over the low grade #2 orebody and portions of the high grade there was an oxide capping which was not mined. The crushed pilars are another source of copper. The leaching at Miami is an attempt to recover this last remaining copper of a worked-out mine. The enrichment at the Miami mine had gone more nearly to completion than most ore bodies in the southwest. There is practically no pyrite in the capping and very little in the ore.

Jan 1, 1971

By J. H. Daniel

The reduction of accidents will play an ever increasing role in the mining of coal and hardrock ores as mining operations become more mechanized, the equipment becomes more costly and sophisticated, and as mineral production exploits deeper and more marginal coal and ore deposits. The most effective means to reduce accidents is through the design of mining systems and equipment. Results of safety research by the U.S. Department of the Interior, Bureau of Mines, to design safer coal extraction and haulage systems and to reduce electrical accidents are presented.

Jan 1, 1991

By Phillip E. Butler

Since early 1973, the Pennsylvania Department of Transportation (PennDOT) has been actively engaged in the utilization of coal mine refuse in the design and construction of highway embankments. The Soils and Geological Engineering Division of the Bureau of Materials, Testing and Research has the task of managing the program which includes field investigations, laboratory testing, engineering evaluation, planning, coordination, implementation, construction control, and review of construction performance. The objectives of the utilization program, in removing colliery refuse banks and utilizing the material in construction, are the environmental and aesthetic rehabilitation of those areas in the Anthracite and Bituminous coal fields which have been plagued for such a long time by air and water pollution, and unsightly landscapes; and the economic rehabilitation of those areas occupied by refuse banks by providing usable land which can now be developed as residential, business, or industrial sites.

Jan 1, 1975