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By J. Partyka

Respirable dust produced during blasting operations in underground mines exhibits both diffusion and convection properties of the airflow. However, the application of the mathematical model presently used to calculate the change in time in the concentration of blasting pollutants is mostly limited to a single drift and cannot be easily utilized for blasts taking place in stopes. This paper discusses an analytical solution based on considerations that enable us to reliably identify the initial condition. Such an approach may be used for any blasting area regardless of its position, its size, or the ventilation system in use during blasting. In addition, a test procedure using field data is established to determined those parameters not measured directly. As well, equations are derived to find the diffusion coefficient, and the position and size of a dust source from the field data. Mathematical considerations The convection-diffusion problem, for a one-dimensional case, is described by the following partial differential equation (PDE): ct. = Dcxx - Vcx + source - sink(1) where C = the concentration of a pollutant T = time V = the mean air velocity in x-direction D = the diffusion coefficient X = distance away from the pollutant source source = any additional pollutant source specified by an initial value problem sink = decay of a pollutant inside the region specified by the boundary conditions. Respirable dust produced during blasting is best considered as an initial value problem (IVP). To derive a mathematical model, the following assumptions are made: •At time zero (t = 0), the dust source is at point x = 0 of an infinite drift. •The problem is considered a one-dimensional case. •The mean velocity of air and the diffusion coefficient are constant. •There is no source or sink of dust other than specified at point x = 0 and time t = 0. We have to solve the following problem [PDE ct=Dcxx-Vcx-m<x...m;0<t<m(2)] ICc(x,0)=f(x)t=0, the initial condition Where c = dust concentration in mg/m3 V= the average mean velocity in a drift in m/s t= time in s D= the diffusion coefficient in m2 /S x = distance coordinate in m. We observe that in problem (2), we have moved the boundaries to[ -m and +m] in order to have an initial-value problem. We introduce new coordinates, s = t and z =x - Vt. To rewrite the PDE in terms of (z, s), we use the chain rule ct = czzt + csst = -Vcz + cs cx = czzx = cz cxx = (cz)x = czz. We now substitute our computed ct, cx. and cxx into the PDE to get new IVP in terms of z and s :

Jan 1, 1991

By N. Arbiter

During the first four decades of flotation practice, close to 200 machine designs were patented, although only about a score were ever used commercially. By the end of that period, the rigors of use had resulted in obsolescence of most pneumatic cells, with a few mechanical cells dominating the field. After World War II, newer designs began to appear but most were modifications in varying degrees of older designs with special features added. Of the dozen or so most recent developments, almost all are pneumatic, relying on compressed air instead of impellers for motive power, although in several cases pumps supplement compressors. Table 1 lists a few of the more fully documented new cells, with their significant characteristics, along with similar characteristics of a widely used mechanical cell. New designs in general are emphasizing: bubble size control; in several cases, methods for delivering feed directly to the froth; methods for introducing and dispersing air through nozzles, eductors, packed beds, or permeable walls; and variable cell volumes, relative to capacities, ranging from very small, with less than a second of retention time, to the largest, with retention times of over 20 min. An important question is the extent to which any of the newer designs are based on locating weaknesses in current mechanical cell designs and then using these as a starting point for innovations. In the present study, the key subprocesses which must take place in a cell are outlined; mechanical cells are analyzed to establish where weaknesses may exist; and then alternative subprocesses are considered which could effect improvements. COMPONENTS FOR CELL DESIGN It is assumed that mineral surface chemistry has been modified as necessary to provide optimum potential floatability. While the cell process may affect this to a degree, cell design is primarily concerned with actualizing the existing potential floatability. The cell design problem is to provide more effectively for the following: 1) Dispersion of air to obtain an optimum bubble size distribution at an optimum ratio of air volume to feed volume. 2) A hydraulic environment which permits most effective contact of potentially floatable particles with bubbles: [Table I - Cell Comparisons* Cell Type Mechanical Column Deister Miller ASH No. of units Multiple Single Single Single Retention time min.10/20 17/27 3.8 ca. is Capacities mt/m3s 0.022/0.066 0.066/0.20 0.11 10.4 st/ft3d 1/3 0.3//0.9 m5.0 469 Air hold-up % 15/20 20 n.a. 50/80 Air escape velocity m/s 0.015/0.0200.020/0.025 n.a. n.a. ft/m 3/4 4/5 n.a. n.a. * References: Mechanical (Harris, 1976,1985), (Bassarear, 1985); Column (Travis, et al. 1987); Deister (Zipperian and Christophersen, 1985); ASH (Air-Sparged Hydrocyclone) (Miller 1989).]

Jan 1, 1989

By R. Timothy McCrum

It is not often that the Interior Department&apos;s Bureau of Land Management, Office of the Solicitor, and Interior Board of Land Appeals (IBLA) jointly act to do the mining industry a good turn. That happened, however, in the recent decision, Schlosser v. Pierce, 92 IBLA 109 (1986). It was ruled that the theorized "independent mine requirement" was not a part of the discovery standard for mining claim validity under the 1872 Mining Law. The phrase "independent mine requirement" would require the mineral reserves within the boundaries of each individual mining claim in a claim group to support an independent economic mining operation. This, without regard to the fact that reserves on adjacent claims could be mined in a group operation, resulting in greatly reduced mining and marketing costs per ton of ore. Such a rule would render invalid under the 1872 Mining Law mining claims for almost all low-grade, high-tonnage, commercially valuable mineral deposits. Until the Schlosser case, no judicial or administrative decision had addressed the issue of the independent mine requirement. Nonetheless, a series of Interior decisions in the 1970s implied such a rule. Even the Rocky Mountain Mineral Law Foundation reluctantly concluded in 1984 that, "each claim to be valid must be capable of being operated at a profit independently of any other claim," the Foundation did criticize the rule as "excessively stringent" and urged the adoption of a "more realistic rule." Fortunately, the IBLA did just that in Schlosser when it ruled that the "application of a requirement that each claim be independently capable of being mined and marketed at a profit is rejected," 92 IBLA at 134. Now, a year after the 1986 decision, there remains some unwarranted confusion over the issue resolved in, and the meaning of, the Schlosser decision. There is no need for this confusion. The Interior Department&apos;s decision is a clear and legally sound victory for the mining industry. Grazing rights vs. mining claims The Schlosser case was an unusual mining claim contest. A private contest was brought before the Interior Department. Schlosser was a federal grazing lessee and surface owner. He filed against Pierce, the locator of a group of mining claims for bentonite over the same lands. At a hearing before interior&apos;s administrative law judge (alj), Schlosser&apos;s counsel began laying the foundation for the independent mine requirement. Counsel asked the mining claimant&apos;s expert whether he would go through the permitting procedure, road building, and reclamation just to mine the amount of bentonite on any one claim in the group. Schlosser&apos;s counsel proceeded with this line of questioning for each of the 16 claims in the group. Pierce&apos;s mining expert testified that the tonnages of mineral on any individual claim could not be mined, processed, and marketed at a profit if only one individual claim was considered. However, if the claims in the group were operated jointly, all the mineral could be mined, processed, and marketed at a profit.

Jan 8, 1987

By S. K. Mishra

Introduction The history of the Australian coal industry is almost as old as the history of Australia after new settlement. The first cargo of coal was exported from Newcastle to India in 1799. Coal has been mined in Australia for almost two centuries. But there are indications that the first coal was processed in 1877 in the Newcastle area. This is nearly 50 years after the first coal processing in Europe. Processing on any significant scale began in 1944 at the Newcastle Steel Works. In Queensland, however, coal processing did not begin until the second quarter of this century. The dramatic growth of mining and processing of Australian coal since the mid-60s is primarily due to the expansion of the export market in Japan, Asia, and Europe. Australian coals Black coal occurs in most of the major sedimentary basins of Australia. The better known deposits of commercial importance are in the Sydney, Bowen, Oakland, and Gailiee Basins. Apart from the Collie Basin in western Australia, other coal deposits of the Arckinga Basin in South Australia, the Bonaparte Basin in Northern Territory, and the Tasmania Basin are basically unexploited. The Sydney Basin of New South Wales and Bowen Basin of Queensland remain the major producers of black coal in Australia. Currently exploited or potentially commercial deposits of black coal in Australia occur in strata ranging in age from Permian to Tertiary. Estimated black coal reserves are summarized in Table 1. Coals from different ages have significant differences in their petrographic characteristics. Many of the Permian coals are usually rich in inertinite. Jurassic and Tertiary coals are typically rich in vitrinite content. Typical petrographic and chemical characteristics of some important Australian coking coals are summarized in Tables 2 and 3. The reserves of coking coals are restricted in the main to the Permian coals of the Sydney and Bowen Basins. Coal deposits in the latter are at shallow depth. A high percentage of these are being strip-mined. On the other hand, in the Sydney Basin, the majority of the reserves of good quality coking coals are at significant depth. The mining cost of these deposits is substantially higher than those in the Bowen Basin. It is usually possible to make strong coke with Australian Permian coals of vitrinite content much lower than the other international coals. But there are many major Permian deposits in Australia where low vitrinite content presents problems. As they are classified as marginal coking coals, optimization of their pro¬cessing plays an important role in meeting the market specifications of coking coal. The major reason for processing Australian coal is to reduce the ash content. Compared with British and some American coals, Australian coals are of relatively high ash content. Unlike American coals, more than 99% of the Australian coal deposits have sulfur content less than 1%. Consequently, sulfur control has never been a major problem. The coking quality of coal is improved simply through the ash reduction in the marketable product. However, with the increase in intensive mechanization and multi-stage handling of coal be-

Jan 9, 1985

By Rebecca S. Stanevich

BACKGROUND Diesel exhaust is an animal carcinogen, although evidence of the carcinogenicity in humans is "limited". (World Health Org. , 1989) Of the eight cohort studies in the 19808, five found an increased risk for lung cancer among workers who may have been exposed to diesel exhaust (RR from 1.3-1.5, SMR from 101-132). (Howe, et al., 1983) Only two studies accounted for smoking indirectly. Of the ten case-control studies in the 1980&apos;s, eight reported increased risk of lung cancer among those populations purportedly exposed to diesel exhaust. (Benhamou, et al. , 1988) All but one of these adjusted for smoking and other occupational exposures. The odds ratios ranged from 1.2 to 2.1. In two of these studies, miners experienced a two to three fold risk. (Benhamou, et al., 1988) Bladder cancer has also been linked to diesel exhaust in cohort and case-control studies. Excesses of approximately 50% were observed among truck drivers in studies by NCI. (Hoar, et al., 1985) In the National Bladder Cancer Study, risks for truck drivers rose with duration of employment. A study by NIOSH in Ohio also found an excess of bladder cancer among truck drivers, particularly among those with 20 or more years of employment. (Steenland, et al, 1987) Despite the large number of studies in which the relation between diesel exhaust exposure and lung cancer has been examined, exposure to diesel exhaust has been quantitatively estimated only in the studies of railroad workers and truck drivers. (Woskie, et al., 1988) The excess lung cancer risk experienced by workers in these two occupational groups was modest, ranging from 40 to 90% for those with long-term employment. However, diesel exhaust exposure among railroad workers and truck drivers is lower than that seen among miners. Several markers of diesel exhaust have been utilized in these studies. Diesel exhaust is a complex mixture of particulates, vapors and gases. One group of these gases, the nitrogen oxides, react with polynuclear aromatic hydrocarbons, forming mutagenic and possible carcinogenic nitrated derivatives. (World Health Organization, 1989) For this reason, and because it is easy to measure and analyze, NO, was frequently used as a marker of diesel exhaust exposure. As part of the Garshick (Garshick, et al., 1988) study of railroad workers, Woskie measured NO, exposures of 477 railroad workers. (Woskie, et al., 1988) Geometric mean exposures of 13 job groups ranged from 12 to 80 ppb. In a New Mexico potash mine, (Attfield, et al. , 1982) exposures of from 100 to 3300 ppb nitrogen dioxide

Jan 1, 1992

By C. J. Hall

The simplest natural ventilation cycle is similar to a Joule or Brayton cycle. This comprises two reversible adiabatic processes in the shafts and constant pressure processes underground and on surface. The thermal efficiency of such a cycle is given by the temperature increase in the downcast shaft divided by the absolute temperature of the air at the bottom of the downcast shaft, and is independent of the temperature increase underground. This efficiency is comparable to the Carnot efficiency, which is the greatest possible value between constant temperature reservoirs, source and sink. Two variations of the Joule cycle are investigated, using numerous work examples. In the first, dry air is used as the working substance, a fan is provided at the top of the upcast shaft, and various mass rates of flow are used with frictionless shafts and a constant resistance value in the horizontal mine workings. The lost work less the fan work gives the natural ventilation work, and this divided by the enthalpy increase in the workings gives the thermal efficiency. The thermal efficiency increases dramatically with mass rate of flow, although the enthalpy increase remains constant. This appears to violate the second law of thermodynamics, since the frictionless case has a Carnot efficiency. When the lost work is added to the enthalpy increase, however, the efficiency decreases slightly from the Carnot value, to conform with the second law. This increase in efficiency with mass rate of flow is often referred to as the reheat effect, although the air is not heated. The second investigation is similar to the first, except that the air entering the downcast shaft contains a small amount of water vapor, and there is evaporation in the mine workings only, until the air entering the upcast shaft is saturated. There is, of course, condensation in the upcast shaft. In this case, it is not possible to calculate the Carnot efficiency, but the thermal efficiency is 20 to 40% of the similar dry air values. This decrease is much more than can be accounted for by the larger value of specific heat for water vapor. In addition, the natural ventilation work is almost constant with increased mass rate of flow, and the thermal efficiency based on heat addition tends to decrease. The thermal efficiency based on heat and lost work decreases in conformity with the second law. The reasons for the differences in behavior between dry and moist air are not readily apparent.

Jan 1, 1991

By Graham Taylor, Mandy Agnew

Hardpans developing in tailings storage facilities (TSF) in Australia have been investigated in relation to Acid Rock Drainage (ARD) generation. This research had three main aims. Firstly it was to obtain an understanding of the physical and chemical characteristics of tailings residues which support the in-situ formation of cemented layers and hardpans. Secondly it was to examine how effective hardpans are at reducing ARD generation, and thirdly it was to investigate the promotion of hardpan formation to inhibit ARD and pollutant generation. This paper presents the data collected during investigations into the effectiveness of hardpans to reduce ARD generation. Three basic types of hardpans have been identified and examined: laterally extensive surface hardpans, laterally extensive cemented layers; and laterally discontinuous hardpans developing in locations of seepage. Their mineralogy, morphology, lateral extent, depth and rate of formation are dictated by sulfide and gangue mineralogy and content, milling and depositional techniques of the tailings and the climate. As the sulfides oxidise, acid and heavy metals are generated, while gangue mineral degradation is enhanced, adding to elemental loads. These liberated ions are then available for precipitation of secondary mineral cements through concentration during evaporation. In TSFs where hardpan development is restricted by mineralogy, grain size or climate, hardpan formation can be enhanced in seepage zones (eg. seepage from waste rock dumps) where tailings are exposed to seepage waters laden with heavy metals and sulfate. The effectiveness of hardpans in the inhibition of ARD generation is dependent on reducing oxygen diffusion and water infiltration into the underlying tailings. A reduction in water infiltration will diminish the transport of contaminants developed during sulfide oxidation and subsequent reactions. Water and gas movement through the hardpans at the CSA (Cu, Zn, Pb) and Elura (Pb, Zn, Ag) mines has been investigated. Permeability tests indicate the hardpans have reduced the infiltration rates by 10 to 100 times that of the fresh tailings. Oxygen diffusion rates were calculated and were shown to be up to 1000 times lower in hardpans than in fresh tailings. Such decreases are explained by persistent elevated &apos;degrees of saturation&apos; within the hardpans. These conditions were verified using a conceptual model of sulfide oxidation reactions, which allowed comparisons of diffusion coefficients with modelled and observed depths of oxidation within the tailings impoundments.

Jan 1, 2000

By R. Svensson

Mill linings play an important role in all comminution processes with tumbling mills. The selection of the lining profile and the lining material has a great impact on overall grinding cost and mill throughput. The cost of the lining itself is minor in comparison with media cost, energy cost and the value of the material produced in the mill. The total scope must be considered when mill linings are selected. This paper presents case stories and explains the technology behind: •Different lining profiles in rubber and steel. •Orebed magnetic lining. •Steel-cap rubber lining. Introduction Grinding mill liner life has been a subject of extensive studies. Performance of mill liners in terms of wear life and how it relates to lining material and profile is abundant in the literature. The most economical liner material and design has been selected based on "cents per ton milled". The evaluation of the effect of liner design on power and grinding media consumption which is a far more significant part of grinding costs, has been less comprehensive (Fig. 1). Lining materials Steel. The most commonly used lining material is steel, calculated on a volume basis. Steel linings have a long tradition and a number of alloys is being developed. The main families are: •High chromium white iron. •Cr and Ni white iron (NiHard). •High chromium martensitic steel. •High carbon Cr-Mo perlitic steel. •High carbon Cr-Mor martensitic steel. •Manganese austenitic steel. Steel linings can be successfully used in most applications. Rubber. Rubber linings were introduced 25-30 years ago and are today commonly accepted to be the most economical linings for secondary and regrind applications. Rubber linings offer long service life in these applications and generally [ ] outlast steel in wet applications, where corrosion is a factor on the steel. The cost relation between rubber and steel has changed dramatically over the years. In the past rubber was more expensive than steel, but today, the cost of rubber is close to half the cost of steel, calculated on a volume basis (Fig. 2). The present cost situation has made rubber linings an economical alternative in several primary grinding applications as well. The wear rate of rubber in primary applications is higher than the wear rate of steel, thus giving a shorter wear life. This can be compensated for by making the rubber lining thicker. The cost advantage of the rubber allows for this, but care must be taken not to jeopardize the grinding performance. Rubber, as a lining material, offers some distinct advantages over steel lining, such as: •Ease of installation and removal. •Safer to handle, less risk of injuries. •Faster replacement. •Reduces noise.

Jan 1, 1996

By James D. Reilly

The time-worn adage "an ounce of prevention is worth a pound of cure" was never more applicable than when used in reference to coal mine safety. Many individuals have dedicated many hours to developing effective accident prevention programs, and much has been accomplished. Each production escalation, however, creates new hazards and safety problems which need special study and research. It is a never ending cyle. To study the progress of coal-mine safety is to raise memories of many individuals who have been dedicated to removing the hazards that have developed with increased productivity and deeper penetration into the earth for coal. Some vivid examples of this dedication are the works of the men who pioneered the use of the safety lamp and the use of rock dust in mining.

Jan 1, 1968

By John A. Kaufman, Terry D. Vandell

Recent mining regulations have increased the need for detailed engineering and environmental information pertaining to the hydrogeologic environment in mining areas. The cost of obtaining such hydrogeologic information has also increased. After mineralization of potentially economic grade ore and tonnage has been identified, and after completing preliminary mining feasibility studies (suggesting that development is merited), a second exploration phase can incorporate a simple, cost-effective baseline hydrogeologic data reconnaissance program that is limited but useful. Using exploration drill holes only, this program can provide a direct indication of where ground water is encountered. Also, reasonable estimates can be made on the relative production of water from various water-bearing strata, a composite potentiometric surface map can be prepared, and the general chemistry of the ground water can be characterized. This article outlines what typical baseline hydrogeologic data are needed and describes how it can be easily obtained during exploration drilling. Introduction Major engineering and environmental ground water parameters can be identified early in a project by using, drill hole data obtained from a mineral exploration program, designed by the mine geologist and hydrogeologist. Most importantly, data obtained in the exploration drilling program can be used to design a logical, defensible, and cost-effective baseline ground water data reconnaissance and monitor well program, which would be required for mine design and ultimately for mine permit approval. Drilling Where possible, exploration drill holes should be drilled with conventional rotary drilling equipment using air or, if necessary, water as the drilling fluid. Rotary air or water drilling generally allow good delineation of the first major saturated or water-bearing zone. If mud must be used during drilling, distinct hydrogeologic zones can still be defined, although the task may be considerably more difficult. Water Levels Static water levels can be obtained from selected exploration drill holes at specific depths while additional boreholes are drilled. Such measurements should not, however, be conducted until the hole has been developed with air (no water) and has stabilized at each of the depths to be measured. Such water-level information can be particularly useful when more than one aquifer exists. Where mine development is eminent, the hydrogeologist should install 50-100 mm (2-4 in.) diameter PVC casing in select drill holes. The PVC should be slotted, gravel packed, and properly sealed so additional hydrogeologic data can be collected for specific hydrogeologic zones at a later date. Such data could include water level information, permeability estimates based on injection and /or pumping tests, and water quality data. Where casing is not installed in drill holes, or when the drill holes are abandoned, the holes should be properly grouted and back-filled to the ground surface to prevent potential future ground water contamination. Monitoring Air Lift Water Production Estimates of the relative production from various water-bearing strata can be obtained from selected exploration drill holes. If ground water is encountered, its air lift (ground water flow) production can be measured at distinct intervals using a surface water flow measuring device, such as a portable Parshall

Jan 2, 1983

By Ramon S. Pizarro, W. Joseph Schlitt

Introduction Trying to synthesize innovations occurring in gold processing is not simple. At present, this technology is exploding, brought about by the high price of gold. This price has compelled the mining industry, as a matter of survival, to produce gold. The frantic search for gold deposits continues, from glacial areas to tropical jungles. The search has been fruitful. New exploration techniques have helped to locate deposits. Some are amenable to conventional processing techniques. Others are in unusual settings that challenge the ingenuity of extractive metallurgists. Ore deposits containing traces of invisible gold, refractory mineralization, and impurities and deleterious materials that previously defied treatment are now being profitably processed. Metallurgists have developed various processing methods and equipment for gold ore treatment that reflect differences in ore characteristics, grade, mineralogy, reserves, topography, climate, and location. These processes can be categorized as: • For nonrefractory or readily leachable gold ores - vat leaching, heap leaching, and agitation leaching. • For ores containing free gold - gravity concentration methods, followed by mercury amalgamation, and • For refractory gold ores - flotation, roasting, pressure oxidation, followed by cyanidation leaching. In actual practice, a combination of these processes may be required. This is because gold mineralization, although usually simple, may be associated with other mineral species or with clays that make the ore difficult to treat. Furthermore, after the gold has been solubilized to form a cyanide complex, one of two important processes must be applied to recover the gold. These ore zinc cementation (Merrill-Crow process) and Carbon adsorption, desorption, and electrowinning. New techniques for leaching and recovery are also advancing. These will become the bases for the next generation of gold processing plants. Nonrefractory Gold Ores Nonrefractory ores with readily leachable gold are among the .easiest to treat. Even so, non-refractory ores from epithermal deposits are likely to contain too much clay. When these ores are damp, they may clog the screens and cone crushers. Depending on the type of process, scrubbing the crushed ore is very effective in overcoming clay problems. After dewatering, the slimes enter the cyanidation process and the thickener overflow is stripped of gold values and recycled. Semiautogenous (SAG) mills are also gaining popularity due to lower capital costs and ability to treat wet, sticky ores. If properly sized, the secondary crushing and screening plants, together with a secondary ball mill, can be eliminated. When the ore is of such a low grade that it can not support fine grinding, vat or heap &apos;teaching may be applicable. These systems are being actively applied in the US and elsewhere. Rapid construction and low capital costs are important features of these plants. Vat leaching is ideally suited for fast leaching ores and small tonnage deposits located in areas with harsh winters or wet tropical climates. For effective vat leaching, the ore is crushed to less than 25 mm (1 in.), agglomerated with lime or cement to control fines, and leached with sodium cyanide solution (upflow or downflow). The degree of size reduction is a tradeoff between increased costs of comminution and improved gold recovery from smaller particles. However, the size of material

Jan 11, 1984

By Stuart R. Cotterill, Jack D. Nelmark

Introduction Opencast surface mining requires blasthole drilling to achieve mine planning. No single method of mining is right for all conditions. Occasionally, mine planners introduce a project with little regard to drilling and blasting. These costs are usually a significant percentage of total mine overhead. One drilling and blasting method that can be used to reduce overall mine expense is inclined or angle hole drilling. But inclined-hole drilling will not cut costs unless a set of required circumstances is met. Several conditions of mine design indicate the use of angle drilling. Geotechnical problems must first be analyzed. Primary considerations are safety, fragmentation, production, and maintenance. Secondary considerations are, in some instances, reduction in amount of explosives, which lowers blasting costs; reduction of practical fragmentation beyond blasting pattern limits; and reductions in overall costs due to reduced volume of excavation. Future trends in blasthole drilling may include incorporation of drill equipment designed to allow rapid change from vertical to inclined drilling. This will allow the operator to adapt to changing hole requirements with minimal production loss. Another improvement may be the ability to chamber holes at any elevation throughout the hole depth. This will allow more charge to be placed at strategic locations to improve fragmentation. Merits of Drilling Inclined Holes Drilling is performed to create voids for explosives. It is done either to recover minerals or to excavate a volume of earth. The manner of mining or excavation can greatly affect the economics of the overall project. Many innovations and methods have been tried, but the most practical and economical is still drilling holes and using explosives to fragment the material. To break up the material, the practice has been to place charges within the mass. This is done by drilling holes of a specific diameter and depth into the material. The holes are of such magnitude as to develop sufficient energy to break or loosen this material, providing for easier removal. In general, more open or exposed faces must exist to aid material breakage when blasting. Inclined drilling can be advantageous in open-cut mining with highwall depths of at least 24 m (80 ft). A minimum of two faces are open in open-cut mining and the material can be thrown outward to cover pre-mined volume. This cuts cost by eliminating the need to remove or relocate a large portion of material. Sometimes the material is of such a mass that more energy per unit is required to loosen it than can easily be done by blasting vertically. Inclined holes allow a more uniform loading of drilled holes. The unit of energy is raised, so little or no secondary blasting is required. This usually comes about only if the main plan calls for drilling holes deeper than 24 m (80 ft). For shallow holes, spacing and burden can be made closer to allow sufficient energy to blast the toe. If the open face is in only one

Jan 9, 1984

By D. L. Slyter

Coeur d'Alene Mines Corporation's Thunder Mountain project in Valley County, Idaho, is one of the most remote active mining operations in the US. High altitude, heavy snows, and proximity to wilderness complicate operations in this environmentally sensitive area. Mining is conducted by conventional open pit techniques while processing utilizes on-off heap leaching and Merrill-Crowe zinc precipitation. The project is one of a few similar operations which detoxify the leached ore prior to its removal from the leach pads. Both leaching and neutralizing ( detoxifying) circuits are closed and independent loops. After leaching, alkaline chlorine solutions destroy residual cyanide in the heaps. Calcium hypochlorite solution for the cyanide destruction is prepared by injecting chlorine gas into milk of lime solution.

Jan 1, 1998

By A. L. Holmes

Since the 1981 record of 54 mm tons of US fertilizer consumption, grain surpluses in the US have brought about massive acreage reduction programs, decreased fertilizer consumption, and resulting price pressures caused by excess capacity. The strong US dollar that reduced export sales of both grain and fertilizer, and new offshore manufacturing capacity using US engineering/contracting know-how have combined to put US producers of phosphate fertilizers in uncompetitive postures. While recent events suggest the industry may be growing out of this capacity problem, the cost of US phosphates will continue to be disadvantaged by such external factors as taxes, environ- mental regulations, and high power costs. There is no reason to expect help from the government in the form of reduced environmental regulation and taxation. In order to survive, phosphate fertilizer producers will have to develop technologies and adopt strategies that result in world competitive production costs. Current responses to the need to reduce costs include reduced energy costs through cogeneration, innovative reclamation techniques that create valuable land, and in- creased raw material recovery achieved by operator proficiency. While new technology offers long-term benefits for the producer, present efforts are concentrated on more mundane management efforts to reduce costs. Major attention is being given to the problems of labor utilization and productivity. Raw materials suppliers are being asked to renegotiate contracts which more appropriately reflect conditions in an industry operating at about 50% of capacity. Achievements in these and other areas will allow the producers to continue operating until longer term benefits materialize. `

Jan 1, 1986

By Val Cachic, Harry M. Castrantas, Craig McKenzie

In April 1986, FMC- Paradise Peak Corporation began milling at the Paradise Peak Mine, which is located eight miles southeast of Gabbs, Nevada. The mill processes about 4DOO tons per day of ore through the mill and cyanide leaching circuit. Gold and silver are recovered via the Merrill-Crowe zinc precipitation process. Tailings are routed to a tailings pond located about one half mile southeast of the mill facility. The pond, lined with 60 mil MDPE, was 25 acres at the start of the cyanide detoxification, covers 90 acres and will ultimately cover 180 acres. Solution is reclaimed from the pond and circulated back to the mill.

Jan 1, 1998

By Ronald Allum

Arizona leads all other states in the value of gemstones produced with an annual production which exceeds $3,000,000. Gemstones are an important component of the mineral wealth of the State. Turquoise, peridot, chrysocoll a, petrified wood, fire agate and crystallized mineral specimens are among the gemstones commercially mined. Numerous additional deposits are operated by hobbyists. The State&apos;s reputation as a source of gemstones and specimens for the collector and lapidary adds to the economy of the State by attracting tourists. Production of gemstones further adds to the State&apos;s economy through employment in the value added gemstone processing industries and by employment of Native American craftsmen.

Jan 1, 1988

By J. W. Carson

Introduction Many bulk solids such as coal, limestone, and fine moist ores often flow unpredictably out of bins, hoppers, and boreholes. Vast amounts of time and money are spent-with or without success-on flow promoting devices. In addition, many solids segregate in storage, feed erratically, flood, arch, rathole, and stick to bin walls reducing the live capacity. This article discusses bulk solids flow properties and provides guidelines for designing storage bins, hoppers, and feeders to provide reliable flow. Bulk Solids A bulk solid is a granular material consisting of discrete solid particles-submicron to several inches in size-that is handled in bulk form (for example, in a bin or silo). The gas filling the voids is generally insufficient to fluidize the material and the liquid (usually water) content is assumed to be less than saturation. Types of Flow Problems To design an efficient storage system, an engineer must be aware of problems that can arise during storage and flow of bulk solids: • No flow. A stable arch or pipe (rathole) develops within the solid, stopping flow. • Erratic flow. Momentary arches form within the solid, (ratholes), partly or completely empty out, and then collapse. • Flushing. Powders become fluidized, aerate, and flush uncontrollably through the bin outlet. • Lack of design capacity. A large part of the stored material remains stable around a rathole. Manual prodding, severe vibration, and even explosives may be used to restore flow. • Segregation. Smaller particles tend to sift through the voids between larger particles causing accumulations of fines and coarse in various areas of the bin. • Product degradation. Some materials spoil, cake, or oxidize if they remain in a bin too long. This is usually caused by a first-in-last-out flow sequence. • Level control. Measuring the volume of material left in a bin is difficult if a rathole develops. In one portion of the bin the level control device may indicate the bin is full; in another portion it may indicate it is empty. Eliminating Flow Problems All flow problems can be minimized or prevented by suitable design of the bin/feeder unit. This involves knowing the flow properties of the solid, then choosing the appropriate type of bin and feeder to establish the desired flow pattern. There are three types of flow patterns that may occur in a bin: funnel flow, mass flow, and expanded flow. Funnel flow occurs when some of the material in the bin is stationary while the rest is in motion. Funnel flow bins, shown in Fig. 1, are characterized by: • Low head room. A flat bottom or shallow hopper bin has large storage capacity with minimum height. Unfortunately, all this storage capacity may not be usable ("live"). • First-in-last-out flow sequence. The first material put into a funnel flow bin is the last material withdrawn because the flowing material is concentrated in the region directly above the outlet. • Ratholes. Because of the flow channel that forms, a stable rathole develops if the material has sufficient cohesive strength. • Erratic flow. Once a rathole becomes partially or completely empty, it may collapse creating a stable arch and erratic flow. • Flushing. Collapsing ratholes cause flushing and flooding of fine material. • Segregation. If a bulk solid has a tendency to segregate, this tendency will be accentuated as the material is discharged from a funnel flow bin. Funnel flow bins are suitable for coarse, free flowing, nondegrading solids, when segregation is not important. Mass flow occurs when all the material in a bin is in motion whenever any is withdrawn. Char-

Jan 3, 1983

By J. A. Brierley, S. R. Hutchins, C. L. Brierley

Introduction Many gold and silver ores are refractory to conventional processing technologies. Thus, precious metals recovery is uneconomic. These resources can consist of ores, flotation concentrates, mill tailings, and other reserves. The refractory nature of these materials can be ascribed to such components as iron sulfide minerals, silica minerals, tellurides, organic compounds, and sulfosalts. Precious metals locked within iron sulfide minerals account for a large number of refractory gold ores. These minerals consist principally of pyrite or arsenopyrite. They must be oxidized to liberate the encapsulated precious metal and allow its contact with the leaching agent. Carbonaceous gold ores represent a unique class of refractory precious metal ores in general. Not only does gold encapsulation by iron sulfide minerals occur (Hausen, 1985), but these ores also contain organic matter that interferes with gold recovery by cyanidation. Direct interference can be attributed to either occlusion of gold within organic matter or formation of a stable gold-carbon complex. The magnitude of this problem is debatable. Wells and Mullens (1973) examined two carbonaceous gold ores and found no association between the gold and carbonaceous material. A more common problem is indirect interference, whereby the aurocyanide complex formed during cyanidation is sorbed by the native organic matter and is not available for recovery. This phenomenon is referred to as "preg robbing" (Osseo-Asare et al., 1984; Hausen and Bucknam, 1984). Guay (1981) summarized various methods that can reduce preg robbing during processing of carbonaceous gold ores. However, some are not applicable to carbonaceous ores that have significant amounts of gold locked in iron sulfide minerals. Several methods exist for oxidizing iron sulfide minerals and rendering refractory precious metal ore amenable to conventional recovery processes. These include various embodiments of roasting (Jha and Kramer, 1984; Brown, 1984; Arriagada and Os-seo-Asare, 1984; Smith, 1986), pressure oxidation (Muir et al., 1984; Beattie et al., 1985; Thompson, 1986; Weir et al., 1986; Argall, 1986), and chemical oxidation (Scheiner, 1971; Jackson, 1983; Van Weert et al., 1986). Each process has been successfully used for refractory precious metal ores. In many instances, however, their use is precluded by capital and operating costs relative to the value of the ore. An alternative to these methods is microbial pretreatment - a biological process whereby iron sulfide minerals are degraded and the liberated precious metal values can be recovered by conventional technologies. Bioleaching refractory precious metal ores is a relatively new concept compared to roasting and chemical oxidation. It is rapidly becoming established as a viable pretreatment alternative. Numerous studies have been conducted in the laboratory and on a pilot scale (Yudina et al., 1979; Lawrence and Bruynesteyn, 1983; Pol &apos;kin et al., 1985; Marchant, 1985; Livesey-Goldblatt, 1985; Gibbs et al., 1986). Comparative tests have demonstrated that bioleaching can be equal to or better than roasting or pressure oxidation in terms of product recovery and process economics (Fridman and Savari, 1983; Livesy-Goldblatt et al., 1983; Karavaiko et al., 1985; Gilbert et al., 1986; Hackl et al., 1986). Bioleaching&apos;s main drawback is that pretreatment requires days rather than hours. This can lead to excessive operating costs for difficult-to-treat ores. In all of the tests reported thus far, the leaching microorganism used has been thiobacillus ferro-

Jan 4, 1988

By D. G. Krige

Discussion by D.G. Krige I read the paper "The jackknife approach" with interest. There appears to be a widespread misunderstanding of the kriging variance as a measure of the error variance of the kriged grade estimate of an ore block. Virtually all mineral deposits display skew frequency distributions and are subject to the well-known proportional effect. Thus, it is usual to find in an ore body, when subdivided a really into subpopulations, that the variances of the grades (or the total sills of the variograms) for the subpopulations will be correlated with the corresponding mean grades (usually linearly with the squares of the mean grades). Thus, the variogram will be station. When this feature is evident, the relative variograms for the subpopulations (i.e., with variogram values divided by the relevant mean squared) will be similar, or they could even be accepted at an averaged model for the whole ore body with stationarity. Using this approach, the kriging variance will reflect the relative error variance that is to be multiplied by the kriged block grade estimate squared to yield the usual kriging variance for purposes of confidential limits. The objection that the kriging variance does not take into account the local grades used for a block valuation then falls away. In the example given by the authors, the global mean grade is not recorded. But the global variance, taken at the value of the total sill, is 0.005, whereas the variance of the 11 values used in the example is 0.000152. The two variances differ by a factor of nearly six. If the relative variogram approach is applicable in this case, the need for a jackknife technique will fall away. Also note that, for skew distributions, the confidence limits for ore-block estimates will be skew as well, and symmetrical normal limits will not apply. RepIy by G.S. Adisoma and M.G. Hester There are a few interesting ways to utilize the block-by-block kriging variance that the authors have encountered. One of them is to classify a particular block into a proven-probable reserve or a possible-resource category, using a somewhat subjectively determined kriging variance value. Here, the kriging variance still represents the drilling density, i.e., data configuration more than the data values. Professor Krige mentions another ingenious approach through the use of relative variograms. This will yield a block kriging variance that takes into account local (block) grade, as he correctly points out, for purposes of confidence limits. However, we are more interested in quantifying the uncertainties associated with grade estimation in a much larger scale than a block. e.g., for a mining phase or the minable reserve. The jackknife, through the block kriging shortcut, provides a simple solution to obtain a single estimation variance for a general shape. The authors&apos; main reference discusses different ways to handle skewed distributions typical in mineral resource estimation, e.g., the weighted jackknife approach or data transformation. Also, as with any technique, we can always improve our estimate by subdividing into smaller populations that are geologically and statistically separate domains. As we have mentioned; the jackknife is but a tool in our collection of tools to solve estimation problems. The jackknife&apos;s appeals are its simplicity and versatility. It is handy when no other solution is available, or when the available solution is not practical. It is used with other methods, not to replace them altogether. ?

Jan 1, 1997

By Patrick M. James

Many in mining believe that the industry is slipping behind in public image, in public mineral education and, particularly, in mining&apos;s basic right to do business. It is critical to turn things around - to change the rules of the game soon - before mining loses any chance of recovery. In the industry&apos;s experience, there is a technique of project development and operation that can help gain ground on public opinion. Only recently has this technique had a name, and that is sustainable development. What is sustainable development? Sustainable development is broadly accepted and supported by the international business community. And significant thought and work has been devoted to the subject. Research into the topic was initially confusing because the definitions of sustainable development are written in terms the writer, and his or her constituency understands. To some, it refers to environmental protection. To others, it is about social responsibility. And to the developer, it is about economic benefit. Slowly a theme emerged from the research that seemed to be part of all the definitions. That theme has three parts. First, the development, whatever or wherever it is, must contribute economic value to the developer and the stakeholders of the development. Stakeholders in this context include employees, neighbors, residents of the surrounding area, citizens of the country, suppliers and all others affected by the development. Next, sustainable development must contribute environmental value. This means that at a very minimum it must be neutral and not cause irreparable damage. And last, and what is new for miners and most other businessmen, it must contribute real social value to the stakeholders. How to use the concept? With that three-part theme as a working definition, the mining industry must find ways to use that “the challenge is for an individual company to translate what sustainable development means for them. There is no one model..." Donald W. Gentry is president and chief executive officer of Polymet Mining and former professor emeritus of mining engineering at Colorado School of Mines. He added: "While simple to state, the concept is anything but simple to implement. Present trends indicate ... one wins the sustainable development argument by staking out the high ground early..." In other words, start using sustainable development principles at the beginning of the project. New ways of doing business The mining industry&apos;s challenge is to implement sus¬tainable development universally around the world. Paternalism as a way of doing business is dead and a new model of doing business is emerging. All types of businesses are learning to deal with the stakeholders of new developments. Mining cannot be an exception to these trends. To understand how the international mining business came to sustainable development, it is useful to look back a few years. For many North American mining engineers, there were few opportunities for foreign assignments from the 1960s to the 1980s. American companies had been asked to leave many of the Latin American countries. Africa was going through a continent-wide upheaval. The Soviet Union and China had Eastern Eu-

Jan 1, 2000