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By Simon J. D. Cox, Nicholas J. Archibald, Steve Nichols, Peter B. Ketelaar, Donelle C. Davies, William L. Power

In the mining industry, there is a critical need to communicate complex, three dimensional information about the geology, ore reserves and development plans for mining operations. This information has to be communicated to a diverse audience, the members of which may not be trained in all of the disciplines required to create, utilise, or interpret the information. Software developed to support World Wide Web (WWW) activities has the potential to improve and change the way exploration and mining data and in- formation is organised, presented, and processed. In this paper, we discuss the creation of interactive displays of complex three dimensional data for the mining industry.

Jan 1, 1996

By L. G. Austin

Introduction My first contact with industrial milling was during the time I worked in the electricity generating industry in the United Kingdom. In visits to power stations to investigate either deposits in the boiler furnaces or polluting deposits settling around the stacks. I had to check the performance of the vertical coal pulverizers, since poor pulverization aggravated both problems. Naturally, then, when I came to the USA in 1957 to take a PhD in fuel technology at Penn State, I was put to work to review the science of coal pulverization. After this reviewing, I was completely confused. On one hand, there was a well-developed understanding of stress-strain equations, and a rap- idly developing knowledge of how stressed, brittle solids fractured, based on the Griffith crack theory. On the other hand, reading in the grinding literature gave me: • Kick's Law, which was clearly not correct in the light of modern fracture theory; • Rittinger's Law, which was also clearly not correct; • Bond's Third Law of Comminution, which was claimed to have something to do with the Griffith crack theory, but where the connection between the two was made by intuitive pseudo- scientific reasoning I could not accept; • the choice of mill motor power for the most common type of coal mill, the Raymond pulverizer, was calculated from the fan power required to move air through the mill. Although I could accept the empirical connection between the two, it made no sense from the point of view of fracture energy. Even today, most books or review chapters on size reduction start from these laws. incorrect statements abound in the literature, such as “the Hardgrove Index is based on Rittinger's Law," which it is not, "The Bond theory states that work input is proportional to new crack tip length produced in particle breakage," which is not true, etc. My own test work showed that these "laws" did not fit the data for grinding of coal. At about this time, Epstein (1 948) and Broadbent and Callcott (1 956), following the original work by R.L. Brown (1941) at the British Coal Utilization Research Association, proposed describing breakage as a series of fracture stages. I took their concepts and developed the basic differential equation for a batch grinding process continuous in time, analogous to a batch chemical reactor. Robin Gardner then joined the project and did his PhD on treating batch grinding in the same way as a batch chemical reactor. He found that the basic equation had already been partially derived by Sedlatschek and Bass (1 953) in Germany. We confirmed experimentally the validity of the equations for describing batch grinding (1 962) and formulated the equation describing steady-state continuous grinding in a fully-mixed mill. At about the time this work was published, Gaudin and Meloy (1962) and Filippov (1961) independently published essentially the same equations, but without experimental proof of the validity of the concepts. I will give a brief overview of what these beginnings had led to in the design of mills for size and power, and show some of the results of this more detailed understanding of grinding processes. Concepts of Fracture Mills such as tumbling ball, rod, pebble and autogenous mills and vertical mills such as the Raymond, and E-type apply compressive stress to lumps or particles relatively slowly. Compressive stress applied to a particle of an elastic brittle solid imparts overall strain energy to the solid and produces local regions of tensile stress, (Fig. 1) (Berenbaum and Brodie, 1959). Irwin (1949) showed from solution of the stress-strain solutions that a small hole in a region of tensile stress reduces stress concentration at the hole, that is, the tensile stress at the tip of a crack or flaw in a solid is much higher than the general tensile stress in the region. The longer the crack, the higher the stress concentration. Griffith (1920) hypothesized that when the regional tensile stress is large enough, then the chemical bonds at a preexisting crack tip are stretched to breaking point, as illustrated in Fig. 2. When the bonds break, the crack becomes longer, the tensile stress concentration increases, the situation is unstable and a crack opens up (propagates) a surface of tensile stress, creating its own tensile stress at the leading edge. Stored strain energy is converted to the kinetic energy of the moving stress field, which is analogous to sound propagation through the solid, so the crack tip accelerates to velocities approaching those of sound. The moving crack will pass through regions that were previously under regional compressive stress. The equations for "ideal" and "Griffith" strengths are where a is the intermolecular distance, y is Young's modulus, g is the energy

Jan 1, 1998

By J. M. Smith, T. R. Horton, R. L. Blanchard, T. W. Fowler

INTRODUCTION Uranium mining operations release radioactive materials into both air and water and generate large quantities of solid wastes containing low levels of radioactive materials. Solid wastes produced by mining operations remain on the surface at many inactive mining sites in the Western United States. These mining effluents may present a potential health and environmental hazard. Therefore, Congress, in Section 114(c) of the Uranium Mill Tailings Radiation Control Act of 1978, instructed the Administrator of EPA to prepare a report identifying the location and potential health, safety, and environmental hazards of uranium mine wastes and to recommend a program to eliminate these hazards. Several facts and limitations helped shape the method and approach of the EPA study. Little information on uranium mines was available; measurement information that was available on uranium mine wastes was frequently influenced (biased) by nearby uranium mills; time and resources did not permit comprehensive field studies to provide additional data; and there are inherent variations between uranium mines and sites that complicate generic assessments of mine wastes. To accommodate these facts, the EPA developed conceptual models of uranium mines and made health and environmental projections from them. The models were based upon available data from the literature, supplemented with information from discussions with persons inside and outside the EPA, and by doing several short-term field studies in Texas, New Mexico, and Wyoming. When necessary, conservative (maximizing) assumptions were employed. This paper presents a brief account of a part of the EPA study dealing with the potential health and environmental effects caused by active surface and underground uranium mines. Airborne contaminants are emphasized, although solid and liquid effluents are also included. Due to the limited space, only the methods and parameters used and the results of the assessments will be presented here. Anyone interested in the source of the data used and the development of the parameters should refer to the EPA report (Blanchard et al., 1981). The occurrence and emissions of stable elements were included in the EPA report, however, due to space limitations and their apparent small impact, except for possibly at some specific mines, only radioactive sources will be included in this presentation. MODEL URANIUM MINES The model surface mine was located in the South Powder River Basin of Wyoming and the model underground mine was located in the Ambrosia Lake area of New Mexico. These are the prevalent type mines in those areas. The model mines were based on the average production parameters of the 63 open pit mines and the 256 underground mines that were operating in the United States in 1978 (Department of Energy, 1979) and on a report of an extensive study of open pit mines in Wyoming (Nielson et al., 1979). Information contained in environmental impact statements and in reports from federal and state agencies was also used. Parameters for the model mines are listed in Table 1. The surface mining scenario is that 7 pits are opened in the 17-year mine life with overburden from each successively mined pit used to backfill a previously completed pit, resulting in an equivalent of one pit of overburden (2.4 year production) stored on the surface. No backfilling is assumed at the underground mine. Overburden or waste rock, ore, and sub-ore are separated into separate piles that are either rectangular in shape with length twice the width or in the shape of a frustum of a regular cone. Both shapes have 45 degree sloping sides. To account for bulking, the volume of the material comprising the piles was considered to be 25% greater than the volume of material removed from the ground. It was assumed that dewatering was required at both mine sites. Wastewater discharge rates at the surface and underground mines were assumed to be 3.0 and 2.0 cu m per min, respectively. SOURCE TERMS The following radioactive contaminants at active uranium mines were assessed in the EPA report: 1. Radioactive particulates in a) wind suspended dust from waste rock (overburden) pile, sub-ore pile, ore stockpile, b) suspended dust from mining activities (rock breakage, loading and unloading ore and wastes), and c) vehicular dust, 2. Rn-222 emanation from waste rock (overburden) pile, sub-ore pile, ore stockpile, and mining activities, 3. Rn-222 emanation from mine surface areas, and 4. Radionuclides in wastewater discharged to land surface. Estimated average annual dust emissions (item 1 above) from the model mines are listed in Table 2. Emission factors and the assumptions used to estimate these dust emissions are described in detail in the EPA report. Radioactive source terms were computed for each of the sources; dust emissions were multiplied by the concentrations listed in Table 1

Jan 1, 1981

By D. Butler

Subsurface density changes can be detected by very precise surface gravity measurements. A feasibility study, combined with field experience using modern instruments, will indicate the probability of success. In addition to projected anomaly parameters, surface conditions that effect the survey precision must be evaluated. Examples with excavated zones at 9 m (30 ft) of depth and [+] 22 microgal error are compared to the same model at 30 m (98 ft) of depth. The figures illustrate anomaly detectability. The survey error is a function of surface topography and lateral density variation in the very near surface.

Jan 1, 1990

By David Wilson, T. V. Rao

INTRODUCTION The Alchem trona mine, owned and operated by Allied Chemical Corp., is located approximately 32 km (20 miles) west of the town of Green River, and 8 km (5 miles) north of Interstate 80 in southwest Wyoming. Trona, a naturally occuring sodium sesquicarbonate (NaCO3 • NaHCO3 • 2H2O), is used for the production of sodium carbonate (Na2CO3), commonly known as soda ash. Present annual US consumption of soda ash is approximately 9.07 Mt (10 million st), out of which approximately 7.3 Mt (8 million st) is produced from the Wyoming trona mines. About 44% of soda ash is consumed by the glass industry; 3 % by exports; and the remaining portion by paper, chemical, soap, textile, and nonferrous industries. Allied Chemical operates an underground mine at an average depth of about 488 m (1600 ft). There are three other trona mines in the area operated by FMC Corp., Stauffer Chemical Corp., and Texasgulf Inc. (see Fig. 1). Allied Chemical became interested in southwest Wyoming trona deposits in the early 1940s. Explora¬tion work was begun in 1959 with the acquisition of federal leases. Exploitation of the reserves began in 1968. Two shafts, one 6.1 m (20 ft) and another 3.7 m (12 ft), were completed prior to 1968. The 6.1-m (20-ft) diam service shaft is divided in half by a cur¬tain wall, with one half used for men and material and the other for ventilation as a return air shaft. During the period from 1968 to 1974, the 3.7-m (12-ft) shaft was used as production shaft with two 7.3-t (8-st) skips. Due to the increased demand for soda ash, plant and mine expansions were completed during 1974. Under this expansion, the material handling system capacity was increased by constructing a new 2721-t (3000-st) capacity ore bin, sinking a 6.1-m (20-ft) diam produc¬tion shaft, and installing a friction hoist with two 22.7-t (25-st) skips. The Alchem mine manpower is 484 employees and about 2.3 Mt (2.5 million st) of trona was produced during 1976. Three different mining systems are used at the Alchem mine: conventional, continuous, and longwall. Allied Chemical introduced its first longwall unit in 1973, which was also the first longwall system introduced in the Wyoming trona district. At present, four conventional, seven continuous, and two longwall units are in operation. This equipment is operated in ten production areas. GEOLOGY The bedded trona and trona-halite in the Green River basin of southwestern Wyoming is a major re¬source for the nation. There are 25 beds that exceed 0.9 m (3 ft) in thickness and underlie areas of about 2590 km2 (1000 sq miles). There are other minor beds that range in thickness from 0.3 to 1.2 m (1 to 4 ft). Allied Chemical's Alchem mine is located in bed 17. The trona bed numbering system starts at the lowest elevation and increases upwards. The trona beds are in the Wilkens Peak member of the Green River formation. The Green River formation was deposited in a lake that began in early Eocene as a large body of fresh water, shrank in size and became saline, and then became fresh water again. The sedi¬ment during the saline phase of the lake, which includes the trona beds, is called the Wilkens Peak member. The underlying freshwater lake deposit is called the Tipton member and includes a 25.9-m (85-ft) sand unit which is a high pressure, low permeability aquifer. The over¬lying sediment is called the Laney shale. The Green River formation is underlaid by the Wasatch formation and overlaid by the Bridger formation (see Fig. 2). The structure of the beds in the area of the Allied Chemical mine is mostly postdepositional asymmetrical folds of small amplitude. These folds range in size of 3 to 6.1 m (10 to 20 ft) to as much as 45.7 m (150 ft). The trend of the folds is northwesterly with the north¬east flank of the anticlines having a steep dip and the southwest having gentle dips. Trona bed 17 varies in thickness from 2.4 to 3.4 m (8 to 11 ft) over the 93.2 km2 (36 sq miles) holdings of Allied Chemical and is flat lying except for the rela¬tively minor folding and a negligible regional dip of about 0.017 rad (1°). Trona occurs naturally in several physical forms in the Green River basin, being both massive and crystal¬line. The trona ore in the Alchem mine is a fine-grained massive ore with a compressive strength of approxi-

Jan 1, 1982

By V. A. Leach, Lokan. K. H., S. B. Solomon, R. S. O’Brien, L. J. Martin, K. N. Wise

INTRODUCTION The development during 1979 of a relatively small, but high grade (10,000 tonnes uranium at an average grade of 2 per cent), uranium ore body at Nabarlek in the Northern Territory, Australia offered an excellent opportunity to obtain detailed radiation data for an open cut mine operating during the dry season. The ore body (Queensland Mines Limited-1979), which was completely extracted in a period of four and a half months, consisted of a vein type deposit dipping at 30 to 45 degrees and contained a central core of pitchblende in massive and irregular pods, surrounded by lower grade fine grained disseminated pitchblende. Mineralisation extended from the surface to a depth of 72 metres over a length of 230 metres with an average but variable thickness of 1D metres. Ore near the surface had been heavily weathered and complex secondary minerals were formed which had dispersed from the main vein. Mining was carried out with large earth moving equipment. Overburden and weathered surface ore were removed initially with scrapers. At greater depths bulldozers were used to rip and assemble ore and rock at each level, and these were removed by large trucks to the ore and waste rock stockpiles. Where necessary, blasting took place during shift changes each evening. Mining was essentially continuous with two ten hour alternating shifts working for thirteen days out of fourteen. At the completion of mining a relatively small excavation (335m x 185m x 70m) remained, and this will serve as a tailings repository during the milling phase. FIELD MEASUREMENTS The inhalation of radon daughters, arising from the radioactive decay of radon gas is well established (Archer et. al. 1973) as a potential hazard in the uranium mining industry. Control over radon and its daughters to ensure that recommended exposure limits are not exceeded is achieved by providing adequate ventilation, and under normal circumstances natural ventilation from an open pit should be sufficient. However, during the dry season it is not uncommon for stable atmospheric conditions, with little horizontal air movement, to develop - particularly at night - and significant radon daughter concentrations may accumulate. Throughout the entire mining period measurements were therefore made of radon and radon daughter levels at representative locations within the pit and on the ore stockpile as it developed. Initially these measurements were carried out manually, using the Rolle method for radon daughters, (Rolle 1972) and .scintillation cells (Lucas 1964) or a two filter tube for the determination of radon (Thomas 1970). For the latter half of the period however, a continuous recording instrument, developed within the Laboratory was used to provide a detailed record of radon daughter levels within the pit. At the same time, continuous readings of wind speed and direction, and vertical temperature gradient between 10 and 3D metres were recorded on a 30 metre meteorological tower, situated 800 metres from the pit. Radon Emanation Rates It is evident that radon and radon daughter concentrations depend on the grade, or more particularly, on the surface radon emanation rate of the ore which is exposed. Accordingly, as the mine progressed, detailed measurements were made of both of these quantities. The surface emanation rate of radon was determined for each ore bench as it was exposed by placing an extended array of canisters, filled with freshly degassed activated charcoal, face down on the ore for a known time. These canisters, which had previously been calibrated in the Laboratory, adsorb radon with high efficiency, and the total radon adsorbed is measured after retrieval by detecting the gamma rays from the trapped radon daughters (Countess 1977). At the same time, as each canister was placed in position, a measurement of the local ore grade was made for each location. This was achieved with a calibrated sodium iodide scintillation detector, adjusted to detect the 609 keV gamma ray from the isotope 2148i, a decay product of radium. Finally, measurements were made of the radiation field 1 metre above the surface, with a gamma ray survey meter, which was calibrated in the Laboratory. The relationship between the scintillator count rate and ore grade was determined by comparing the scintillator output with the gamma monitor, and relating the latter measurements to ore grade (Thomson and Wilson 1980). It was observed that while emanation rates and ore grades varied widely, the ratio of emanation rate to ore grade was in general fairly stable. A plot of this ratio is presented as a function of depth below the original surface in Figure 1. For most observations the ratio is constant at a value of 80 Bq m-2 s -1 per unit ore grade, where ore grade is expressed as percentage of U308. At the surface however, where the ore was weathered, the ratio was about a factor of three higher, and at two particular depths, where high grade pitchblende was being removed, it was very much lower. This was not unexpected as earlier Laboratory studies of drill core samples from Nabarlek had indicated that the emanation coefficient (the fraction of radon produced within the ore which escapes from the mineral particles) decreases with increasing ore grade.

Jan 1, 1981

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Victor E. Archer

INTRODUCTION Man is exposed to many agents which induce mutations in germ cells and/or cancer at work, at play, and at home. In this total mix of mutagenic and carcinogenic agents, how important are radon and its daughters? Before man moved into caves and other permanent dwellings, the principal mutagenic and carcinogenic agent to which he was exposed was natural background radiation--cosmic rays, radium and potassium-40 in his food, plus gamma rays and radon from the soil and rocks. When man moved into caves, captured fire, and began to preserve and store foods, his exposure to carcinogens and mutagens took a quantum leap. Carcinogens and mutagens appear to act in the same way, that is, by altering the DNA or nuclear proteins of cells. Most mutagens are carcinogens, and vice versa, so when I say mutagens from here on, I will be referring to both. The relationship of the two is emphasized by the fact that administration of a carcinogen to a group of animals not only increases cancer rates among the exposed animals, but also among their progeny (Tomatis 1979). Environmental Mutagens Smoke from man's fires, overheated foods, and foods preserved by smoking, resulted in ingestion and inhalation of many polycyclic aromatic hydrocarbons--many of which are mutagens. Caves and houses with tight windows and doors tend to collect the radon which is constantly emanating out of soil, rocks and concrete, so man's exposure to the radon daughter component of background radiation increased several fold. Preserving food by salting or pickling with material that contained nitrites and nitrates led to increased ingestion of nitrosamines, which are potent mutagens. When his grains and other foods were stored in slightly damp rooms, fungi or mold would grow on them. Several of these fungi are now known to produce very potent mutagens. The best known of these is aflatoxin B (Ramachandra 1979). It may seem strange that a living organism would produce a mutagen. One might think that it would scramble its own genetic heritage. The reason it does not is that it produces the mutagen in an inactive form. It can be activated only by an animal's enzyme systems after being eaten. When man moved into cities, the collective smoke from wood and coal fires further increased his exposure. That particular smoke has now mostly disappeared, but has been replaced by smoke from automobiles and industry. When man moved into the age of technology, his exposure to mutagens again increased dramatically. Many mutagenic chemicals, from benzene and beta naphthylamine to a long array of pesticides and tobacco products have been added to our environment. Excess deaths from cancer are now being observed among chemists in most industrialized nations. Mutagens are even found in much of our wine, beer, and whiskey (Keller 1980). Some of the chemical mutagens were widely used in food or in other commercial products before their potential was discovered. Striking examples of this is the original butter coloring agent and the polychlorinated biphenyls that have been widely used in brake fluids and electrical transformers. Large quantities of them have been discarded or disposed of in a careless manner--in such a way that many of them have contaminated our food, our ground water and air (Landrigan 1981). In this nation, with the help of several recent laws, we were just beginning to get control of the industrial chemical mutagens. With the relaxing of these laws that is currently going on, it appears that it will be many more years before we really bring chemical mutagens under control. Many nations have yet to come to grips with this problem. On top of this massive array of chemical mutagens we have now added radiation from many artificial sources. For most of us this means medical X-ray and fallout from nuclear weapons testing. Ionizing radiation is one of the most potent mutagens, so it has caught the public eye, and its contribution cannot be ignored. Fortunately, by the time we started using radioactive materials in quantity with the Manhattan Project, we had experience with radium and X-ray (some of it bad); we knew enough radiobiology and enough about methods of radiation protection so that most nuclear laboratories have had a phenomenal record of radiation safety. Radiation is one new technology with great potential for harm that has not exhibited that potential except for a few isolated situations like that of radium dial painters, uranium miners and atomic bomb victims. Uranium miners slipped into this list almost by accident. We could have protected our uranium miners just as well as we did the workers in nuclear laboratories; but we failed to do so. Why didn't we? The reason is simple. The Atomic Energy Commission was charged with protecting the health of their workers. They did not wait for a pile of bodies before they introduced controls. Congress appropriated the money, and taxpayers were willing to pay for the protection against radiation. Miners unfortunately did not work for the Atomic Energy Commission. Although mine operators were ignorant about radiation, the key item was that in the 1950s nobody was willing to pay the extra costs of adequate ventilation to control the high levels of radon and radon daughters in uranium mines. Control was not achieved until new laws and regulations were passed which made it compulsory. BIRTH DEFECTS AND CANCER

Jan 1, 1981

By Robert Snow, Yingxue Yu, Patrick Zhang

The FIPR/SAPR process is FIPR's third approach to develop a viable alternative to the Crago "Double Float" process for phosphate flotation. SAPR stands for Single-collector, All-anionic Phosphate Recovery. The FIPR/SAPR process - offers a universal flowsheet for any anionic reagent system and flotation feed of varying sizes. For an unsized or fine flotation feed, the FIPR/SAPR process consists of the following steps: 1) high-solids conditioning with an anionic collector; 2) anionic rougher flotation, with the rougher concentrate sized at 48 (or 65) mesh and the +48 mesh recovered as a final product; and 3) cleaning flotation of the -48 mesh fraction from step 2. In a variation of step 2, the rougher concentrate from the first two cells may be collected as a final product, and the rougher concentrate from the last two cells sized at 48 or 65 mesh. The new process was tested with a blend of anionic collectors, achieving single-digit Insol at 85 + % recoveries.

Jan 1, 2002

By Dave Southworth

After Abe Lee discovered gold (in April 1860) near the site of present-day Leadville, Colorado, he and is group staked nearly all of California Gulch with "speculative claims." They did so in an effort to grab any property that might be potentially profitable. But the plan did not work. New prospectors demanded a miners' meeting, from which arose the official "Bylaws of California Mining District, California Gulch, Arkansas River." These bylaws, adopted on May 12, 1860, established provisions as to the number, size and type of claims that could be filed. Fair laws invited more newcomers. By mid-summer, 5,000 people with tents and wagons were scattered from one end of the gulch to the other.Among the earliest arrivals to the region that spring were Horace Austin Warner Tabor and his wife, Augusta. H.A.W. Tabor, who was to become a pillar in the development of Leadville, arrived at California Gulch during a food shortage. He sacrificed his oxen so that hungry miners could eat. By doing so, he immediately became the friend of many. Augusta established a little store where she sold her baked goods and provided meals. Like most of the men, Horace prospected the surrounding hills. He only did so, however, for a short time. California Gulch yielded more than $5 million in gold during the next five years. By 1866, however, the placer gold was gone and most of the miners had packed up and left. Charles J. Mullen and Cooper Smith, who were grubstaked by a Philadelphian named J. Marshall Paul, found a new, rich lode at the Printer Boy Mine in 1868. It was the first underground gold mine in the Leadville area and marked the transition from surface mining to underground hard rock mining. Its discovery created a new but modest flurry of activity. Heavy black sand covered the region and made gold mining difficult. In 1875 William Stevens and Alvinus Wood decided to have the sand assayed. They discovered that it contained 15 ounces per ton of silver and was rich in carbonate. Stevens and Wood kept their find a secret for nearly two years. When the word spread in late 1877, the silver boom was on. In 1878 prospectors, merchants and others trickled into town from all directions. Dr. David H. Dougan closed his office. in Alma, crossed Mosquito Pass and set up shop in Leadville. Nobody seemed to need a doctor, however, and Dougan sat in his office for 28 days without a customer. On the 29th day he received word that there had been a mining accident and he was needed immediately. -That was the begin¬ning of his illustrious career in Leadville. Dougan was a successful physician, became mayor in 1881 and was later president of the Carbonate National Bank. Leadville was a simple community of log structures in 1878. By June of the following year, however, it had blossomed considerably. According to Cass Carpenter, as of May 1,1879, Leadville had "...19 hotels, 41 lodging houses, 82 drinking saloons, 38 restaurants, 13 wholesale liquor houses, 10 lumber yards, 7 smelting and reduction works, 2 sampling works for testing ores, 12 blacksmith shops, 6 livery stables, 6 jewelry stores, 3 undertakers and 21 gambling houses where all sorts of games are played as openly as the Sunday School sermon is conducted." Additionally, there were 36 brothels. When George E. King and other architects arrived, Leadville took on an air of elegance and sophistication. In sharp contrast to some of the Victorian-style, gingerbread-laced homes in Leadville, its suburbs were comprised of many tenement shacks and plain cabins mixed with a few commercial buildings, many mines and smelters. Mines sprinkled the area east of Leadville like pepper from a pepper shaker. Wherever there was a cluster of mines, a camp usually sprang up. Most were very close together. South of Leadville the giant Arkansas Valley lead smelter sprawled out between Stringtown and Bucktown in an area known as Smelter Valley. Founded in 1879 as the Billing and Eilers Smelter, the Arkansas Valley plant was operated by the American Smelting and Refining Company. Several other smelters operated in the vicin¬ity. In April 1878, Horace Tabor grubstaked prospectors August Rische and George Hook. For $64 worth of supplies and tools, he was guaranteed a one-third interest in their findings. Rische and Hook discovered the Little Pittsburg (often spelled Pittsburgh) Mine and made Tabor an instant millionaire. Tabor once purchased a "salted" mine, the Chrysolite; but instead of considering it a bad deal, he further developed the property and struck rich silver ore. Later, Tabor purchased the Matchless Mine for $117,000, possibly the only investment he made without partners. During its peak years, the Matchless yielded $1 million worth

Jan 1, 1998

By James Sheridan

Risk can generally be defined as the possibility or chance of a loss corresponding to uncertainty. Analysis is the process for examining the components and relationships of a complex system in order to determine the impacts of its variables. Risk analysis is therefore a course of action taken to analyze multiple variables of a system that is subject to speculative outcome. When risk analysis is applied to mine development, it becomes the means for assessing and measuring those parameters that contribute to economic viability. Mining can be a risky business. It is composed of an assortment of factors that contribute to its financial outcome. Of these factors, the most consequential output is the net present value (NPV). Applying risk analysis in mine development can identify those factors having the most significant impacts and applying limits to those impacts. Mathematical computer modeling using Monte Carlo techniques is available and can be customized to suit specific mine development ventures. Essentially, those parameters that are not fixed values can have input variability. Rationally deduced engineering estimates should be the basis for determining the initial values applied as inputs in the models. Risk analysis therefore provides the user a means to identify, rationalize and assess the risks in mine development in an informed, calculated, and unbiased format for making economic decisions.

Jan 1, 2002

By Richard O. Y. Breese

Geologically and commercially, the term diatomite is applied to the nearly pure sedimentarv accumulation of diatom frustules- the microscopic skeletons of unicellular aquatic algae belonging to the class of golden brown algae, bacillariophyceae. The sediments are fine-grained, siliceous, and consist primarily of amorphous opaline silica with only minor amounts of organic residue, secondary minerals, and co-deposited non-diatomaceous or crystalline clastic debris. In the geological sense, the name diatomite implies sedimentary accumulations that have reached appreciable thickness and when thick enough, such accumulations consequently may have possible commercial potential. Although the term diatomite is popularly and inappropriately applied to any sediment in which there is an abundance of diatom frustules, alternative terminology is more correctly employed to describe the less pure diatomaceous sediments, for example: clay-bearing diatomite or diatom-bearing clay. Synonyms in current usage include diatomaceous earth and kieselghur. More antiquated and obsolete terminology includes tripoli-powder, tripolite, and infusorial earth. Worldwide, diatomites occur within Tertiary to Recent lacus- trine and marine sedimentary facies. Although diatomite is wide- spread throughout the world, deposits that contain high purity, commercially versatile ore are uncommon. Physical properties of the diatom and of processed diatomite that provide unique commercial value in a broad spectrum of market end-uses include ornate fine structure, low bulk density, and high porosity and surface area. Properties of equal importance are mild abrasiveness, high absorptive capacity, insulating ability, relative inertness, and high brightness. End-use markets are diverse and range from insulating brick and absorbents through quality sensitive filter aids and premium quality functional fillers. Notwithstanding both the economic attractiveness of the specialty markets and the commercial versatility of the high purity deposits, deposits of lesser purity are mined in nearly all parts of the world for less demanding uses. Mining costs are minimized through open pit quarrying, but in Europe, Asia, Africa, and South America underground mining methods are also employed. Blasting is not required because diatomite is soft and easily broken loose with mechanized equipment. Following gentle crushing, the ore is dried, milled, and processed into one of three broad categories of products: naturally milled, straight calcined, and flux calcined grades. Each of these principal categories is further subdivided into additional grades through particle size adjustment of the powders. Agglomeration of particles, alteration of fine structure, and color change are achieved through calcination. Flux calcination further accentuates these changes. Production throughout the world is dominated by the United States, followed by Romania, the former Soviet republics, and France. Other major producers include Spain, Mexico, Iceland, Korea, Japan, and Germany. Denmark is a major producer of Moler earth products, which consist of a diatomite-clay mixture. HISTORICAL BACKGROUND The first industrial use of diatomite can be traced back some 2000 years to the Greeks and to the use of diatomaceous earth as a component in lightweight building brick and in ceramic pottery. It was not until the mid-1800s, however, that the unique properties of diatomite were first recognized and the market end-uses explored and developed. One of the most important of the early uses followed the development of dynamite by Alfred Nobel in the mid-1860s, whereupon diatomite was used as a component of the explosive to improve stability and safety. Other early uses included low temperature insulating and refractory bricks and as a component in insulating and fireproofing construction panels. During the 1920s, processing technology underwent a very rapid evolution with the development of calcination, flux calcination, and air classification technologies. Through these technologies, innumerable different size and grade classifications could be made for market applications and for rapidly diversified end- uses. Today, the application of processed diatomite for filter aid is the largest of the quality sensitive end-uses. Specific filtration uses include the clarification of beer, wine and liquor, vegetable oil, syrup and sugar, pharmaceuticals, and swimming pool water, to name a few. As a functional filler and extender, processed diatomite is ideally suited for application in paint, rubber, and plastic formulations. Other filler applications include use as an anti-blocking agent in plastic film, anti-caking agent for fertilizer, thermal insulating material, catalyst carrier, polish, abrasive, pesticide and fertilizer carrier, and chromatographic supports. Throughout much of the world diatomite is still used as a component of insulating brick and as an absorbent.

Jan 1, 1994

By F. M. Doyle, J. P. Dwyer

Introduction The mining industry has a poor public image. It is often perceived as a despoiler of the landscape and a polluter of the environment, and there is little recognition of societal needs for the raw materials and fuels that this industry produces. There are many reasons for this poor image, but an important one is that a substantial number of mining operations have caused significant damage to the environment. Many of these operations, particularly those that now are abandoned, continue to have an adverse impact, principally in the form of water pollution.1 In some cases, the damage caused by this pollution extends well beyond the immediate region around the mine. Chemical contamination of ground and surface waters can, and does, occur in nature when these waters flow through or over undisturbed, mineralized rock masses. However, the problem is greatly exacerbated by mining because the mining activities increase, by several orders of magnitude, the permeability and the surface area of rock that can be contacted by percolating water. Water flowing through either the mine workings or the waste rock dumps becomes contaminated. Pollution can also arise if leachate from tailings ponds enters either groundwater or surface waters. This leachate may contain chemicals used in processing the ore. The magnitude of the threat posed by mining wastes depends on the type of minerals in the deposit, and on the chemicals used to process the ore. The magnitude of the threat also depends on two other, site-specific factors. One is the volume of water flowing into the mine and through the dumps, which depends on rainfall patterns, topography, and whether the mine workings are surface or underground. The other is the proximity of the pollution source to groundwater or surface waters, which depends on both the climate and geology. These factors influence the concentration and the quantity of polluted water forming and reaching the receiving water. Mining activities present a potential problem of enormous scale. Vast amounts of rock are mined each year2 and, as noted above, substantial problems have arisen at some mine sites. These problems are usually exceedingly costly to address, and it is imperative that these mistakes not be repeated. There are indications that we can accomplish this goal; we now better understand the processes that cause pollution and have methods to prevent or contain it. In addition, in the past few decades society has started to recognize the need for all industries to operate in a manner that maximizes the health and safety of workers and the public and minimizes adverse environmental impacts. A variety of recently enacted federal and state laws regulating public health and environmental hazards reflect the increasing concern over these issues. Many statutes, such as the federal Clean Air Act and the Clean Water Act, encompass activities undertaken by the mining industry. Other laws, such as the Surface Mining Control and Reclamation Act, have been written to regulate specific hazards posed by this industry. At the present time, the federal government. through the Environmental Protection Agency (EPA), and several state governments, are preparing new regulations specifically tailored to control the disposal of mining wastes. Given the magnitude of the potential environmental and health risks posed by disposal of mining wastes and the regulatory costs to the mining industry (both of which are due to the tremendous volume of waste generated), it is especially important that regulators have a good understanding of the different types of mining wastes and the risks they are likely to pose in various geographic and climatic circumstances. Only then can a regulatory program be designed with suitable procedures and substantive criteria to protect the environment and ensure compliance, while avoiding unnecessary or wasteful expenditures. California has long enjoyed a large, active mining industry. Today it produces a wide range of mineral commodities. The revenue from production of non-fuel minerals in California ($2.85 billion in 1988) greatly exceeds that of any other state. About a third of this revenue is from the production of inert sand, gravel and crushed stone. However, the remainder is from the production of a wide range of materials. These include boron minerals, rare earths, asbestos, talc and pyrophyllite, and gold. In earlier decades, an even broader range of minerals was produced, including mercury, copper, lead, zinc, and barite. There are acute environmental problems at some of these early mine sites. Because in 1987 the state of California was engaged in a revision of its regulations for mining waste disposal, the legislature commissioned a study by researchers at the University of California at Berkeley to investigate and review: the magnitude of the problem posed by these wastes, • the best feasible control measures available for waste management and disposal, • the problems of abatement and cleanup at so-called "problem" mine sites, • the current regulations imposed on mining waste disposal by various federal, state, and local agencies and the effectiveness of these regulations. This paper summarizes the principal findings from this study (Mining Waste Study, 1988). The lessons from this investigation are likely to be of general interest because of the diversity of mineral products produced, and the wide range of geographic and climatic conditions encountered. The acute problems seen at some of the older mines are typical of the western United States, and teach useful lessons on strategies that should be implemented at new mines to prevent recurrences of these problems.

Jan 1, 1992

By Everett M. (Original by) White

Numerous articles have been written about the coal mine ; he has been compared with brave men who have gone out to conquer some unknown hazard throughout the ages. Now, however, modern mining is no longer something to be avoided, but a vocation that anyone can be proud of, with remuneration equal to or above that of almost any industry. A concerted effort is being made each day by both the operators of most mines as well as the state and federal inspectors to see that coal mines are as safe or safer than work places in any other industry. How- ever, it is true that mine fires and explosions have occurred in the mining industry since its inception. One of the earliest accurate references to such an occurrence is found in the transactions of the Royal Society of London. In the Colliery Managers' Handbook, 1898, Caleb Pamely writes of an explosion that occurred at Fatfield, England in August 1708, resulting in the deaths of 69 persons. He states: "Three of them, viz, two men and a woman, were blown quite up from the bottom of the shaft, fifty-seven fathoms deep, into the air, to a considerable distance from the mouth of the pit." It is interesting to note that women were employed as miners in England during that period. In the United States, the first record of coal mined was in 1702 by the Jamestown colonists. In 1786, about 20 years after Pittsburgh, PA was first settled, a charter was granted by William Penn to mine coal on the banks of the Monongahela River. Then in 1803, the first important coal company, the Lehigh Coal Co., was established. From this beginning, the mining industry continued to grow in importance and the first state mining law was formulated in 1870. The Pennsylvania state mining law was followed in 1873 by Ohio's, and in 1876 by Maryland's. A mine explosion at Avondale, PA in 1869 stimulated state legislatures to formulate these early laws. In 1907 near the town of Fairmont, WV, there were, within three days, two explosions resulting in a total of 597 deaths. Because of this, the federal government took action and by 1910, the US Bureau of Mines had been established. This was perhaps the most significant change in health and safety for miners during the first half of the 20th century as the USBM concentrated its entire efforts in the study, research, and testing of methods totally directed towards improvement of the health and safety of the miner. After a mine disaster in Illinois, a federal law was passed in 1940 calling for federal coal mine inspectors. Then in 1969, Congress enacted the Federal Coal Mine Health and Safety Act following a mine explosion near Farmington, WV. In 1977, the law was revised and included noncoal mining operations.

Jan 1, 1981

By T. J. Jr. DeMull, F. G. Miller, J. P. Matoney

For some time a need had existed in the minerals processing field for a relatively efficient separator that would treat high tonnages of particles in the intermediate size range, i.e., those particles too large for froth flotation and too small for conventional gravity-type separa¬tors. Among those devices developed to meet this need are the centrifu¬gal specific gravity separators. These devices employ centrifugal acceleration to assist gravitational acceleration in separating light¬density minerals from heavy-density minerals. In the category of centrifugal specific gravity separators are the heavy-media centrifugal separator and the water-only cyclone. The two major centrifugal heavy-media separators, i.e., the heavy-media cyclone and the DynaWhirlpool, as well as the water-only cyclone, are discussed in terms of: design features, operating variables, operat¬ing data, and flowsheet design criteria. Examples of plant applications are given in the field of coal processing as well as the processing of other minerals such as iron ores, potash, and tin. Finally, the subject of the staging of centrifugal separators and their use in combination with other separators is discussed. PRINCIPLES For coarse sizes of minerals, efficient specific gravity separations have been possible for many years with open-bath vessels using the natural settling velocity or buoyancy of the particles. These bath ves¬sels process ore by utilizing micron-size solid particles suspended in the slurry fed to the separator. The inclusion of these particles in the slurry increases the effective density of the separating fluid to allow particle separations to be made at densities greater than that of water. However, if vessel size is to remain within economical limits, the particles processed in the bath vessel must have high settling rates in a IG gravitational field. Because of this requirement, heavy¬medium bath vessels are usually restricted to processing +V4-in. sizes. To extend efficient specific gravity separation to smaller sizes, the gravitational acceleration of particles is replaced by centrifugal acceleration. The settling of a small particle in a fluid in a centrifugal force field is similar to that found in a static bath except that the acceleration due to gravity, g, is replaced by a centrifugal acceleration where v, is the tangential velocity at radius r: V=kdm(P-P,), V'. (J) µ In more practical terms where the particles settle in a suspension of finer particles comprising the heavy media and with an effective suspension density p" V = kdm' (P P ) . v'. (2) U r To date the most effective use of this principle has been obtained with devices that rotate a liquid or suspension within a stationary enclosure in order to create centrifugal force. Cyclones are the most common devices used for this purpose, because they generate centrifu¬gal forces far greater than the force of gravity and therefore not only have high capacities but can treat finer sizes than bath-type vessels can. The two main types of cyclones used by industry are the heavy-media cyclone and the water-only cyclone. Also quite widely used is the DynaWhirlpool, which, though based on the same princi¬ple, differs in design from the conventional cyclone. HEAVY-MEDIA CENTRIFUGAL SEPARATORS Like the bath vessels, the heavy-media centrifugal separators em¬ploy media composed of micron-size particles suspended in water. However, the centrifugal force generated in these separators accentu¬ates the difference in settling rate between particles of different density and thus makes possible separations of finer size particles than can be treated in bath vessels. The two most common heavy-media centri¬fugal separators are the heavy-media cyclone and the DynaWhirlpool. Heavy-Media Cyclone Although cyclones were originally developed for use as classifiers or thickeners, it was later found that they could also effectively serve as heavy-media separators.63. 64, Design Features Fig. I I is a schematic of a typical cyclone developed to serve for any one of the following purposes: as a classifier, thickener, or specific gravity separator. The cyclone consists of a cylindrical section joined to a conical section, usually having an included angle of between 14° and 25°. Feed enters the cyclone tangentially through an orifice attached to the cylindrical section. The overflow orifice is located in the base plate of the cylindrical section. The vortex finder, a tube attached to the overflow orifice, extends into the cyclone from the base plate of the cylindrical section. The underflow orifice is located at the apex of the conical section. As some medium together with mineral particles is fed through the feed orifice, a vortex with a hollow air core extending from the overflow to the underflow orifice forms in the cyclone while hollow spray discharges form at each of these orifices. Under the influence of the centrifugal force, high specific-gravity particles move through the medium to the wall of the cyclone and descend in a spiral flow pattern to the underflow orifice. Those particles in the feed stream, lower in specific gravity than the feed medium, follow the major portion of the flow to the center of the core where they are caught in the high-velocity upward central current and are carried out through the overflow orifice. Fig. 12 shows a family of curves that illustrates how materials of varying specific gravity are recovered by a cyclone. Since the specific gravity of the medium is 1.40, the particles of 1.40 sp gr actually act as part of the medium and, regardless of the particle size, split between the underflow and overflow of the cyclone in proportion to the volume split of the medium. Particles higher in specific gravity than 1.40 are recovered in the underflow of the cyclone at increasing rates as the difference in specific gravity increases and the particle size increases. Particles lower than 1.40 in specific gravity are dis¬charged through the overflow orifice at increasing rates as the specific¬gravity difference increases and the particle size increases. However, all the curves originate at the fluid-flow ratio point for the finest particles of any gravity. The fluid-flow ratio is defined as the ratio of the rate of fluid flowing from the underflow to the rate of fluid

Jan 1, 1985

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By M. Javier

In business mining, dilution is the result of the operating parameters that occur too often and cause serious consequences to mine performance, questioning TO WHOM BENEFITS DILUTION. Mining persists in changing perception of real meaning of dilution. The lack of interrelation between mineral processes and nature compromises the understanding of dilution; this absence of interrelation occurs due to the misinterpretation of nature, otherwise, Geomimic could be helpful to filling the gap of misinformation. NATURE provides LIMITED metals in low content deposits; it is a challenge to obtain them. Nature never considers dilution. Furthermore, though dilution is a human concept, dilution is accepted and promoted. Reality of dilution is in its errors accumulation that translates into technology level or status applied in nature. It is doubtful that dilution can continue surviving if it is ?meaningless?. It will discourse: Is dilution illegal? If the change occurs in the prediction of mining will prevail a different destiny, it is important to touch the cores of the business; otherwise, it would represent repetition of the same ?negative? effects, as usual. Mining defines many losses during the disturbance of nature, from NPV elaboration to the metal recovery to commerce, demanding BETTER DESIGN. This paper introduces and applies MANAGEMENT BY LOSS, as first steps, to limit risks and uncertainties in mining. As result, the management of dilution will be obtainable under mining practices which promotes its character of SUSTAINABILITY.

Feb 23, 2014

By Alexander May

In its role of providing technology to promote the efficient use of minerals and mineral process waste, the U.S. Bureau of Mines has conducted research at: its Tuscaloosa Research Center to identify and develop high volume uses for phosphogypsum. Mixtures containing a maximum of 50 percent phosphogypsum, 6 to 10 percent lime, and the remainder fly ash had compressive strengths as high as 4,800 psi. The effects of the mounts of phosphogypsum, lime, and fly ash were investigated. Also, the effects of fly ash to lime ratios were studied.

Jan 1, 1983

By Bryan A. Bartley

High intensity, autogenous grinding has been developed in the hard rock, quarry industry and is finding wide application in mineral processing. The autogenous vertical shaft impactor is described and action within the machine is illustrated. This machine is compared with the cascade mill, in the three characteristics of impact, inter-particle rubbing and nipping, of which impact is the most important. There is a seven fold increase in impact level over that of the cascade mill. The autogenous impactor has characteristics of crushing and grinding, preferential crushing, abrasion resistance and drying. These are illustrated by examples of industrial applications, where the machine is now used and range from abrasives, such as Carborundum to phosphate rock.

Jan 1, 1998

By Rick T. Richins

Despite what may be overwhelming evidence that natural degradation of cyanide and cyanide attenuation in hydrogeochemical processes are capable of reducing or eliminating potential water quality problems associated with heap leach facilities, the recent sharp increase in the price of gold and wide interest in the cyanidation process has focused all regulatory eyes on cyanide leaching. Regulations continue to increase, design criteria are more strict, and detoxification requirements approach or exceed the level of detectability. Current governmental regulation of mining and milling operations using cyanide varies considerably from state to state. The lack of regulatory consistency at this level is largely attributable to the current Bevill Amendment exclusion of mine wastes as potentially hazardous wastes, the undefined status of the Environmental Protection Agency's ongoing mine wast study effort, and the mining industry's reluctance to develop a consistent policy regarding state versus federal mine waste regulation.

Jan 1, 1998