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By F. Habashi

Chile is at present the largest copper and rhenium producer and was before World War I the largest nitrate producer. She is also an important producer of lithium carbonate and molybdenum oxide. An important event in the history of the country is the Pacific War of 1879-1883 in which she acquired the northern provinces initially belonged to Bolivia and Peru. In 1888 Chile annexed Easter Island. The Chilean mineral and cultural history is briefly reviewed together with the people who contributed to her development. The present report is the result of a number of academic and industrial visits to Chile during the period 1992 to 2009.

Jan 1, 2011

By W. K. Tolley, D. G. Kotlyar

Silver is an economically important component of many complex, massive sulfide ores. The beneficiation of this type of ore is now almost entirely accomplished by flotation. Silver recovery is often low, and little is known about the flotation behavior of the silver-bearing minerals. Floatability studies were conducted using silver-bearing tetrahedrite and lead-zinc ore. The effects of xanthate-collector concentration, conditioning time, potentiostatic control of the mineral potential and control of the pH on mineral floatability were evaluated. The floatability of the silver-bearing minerals was shown to depend on collector concentration and mineral potential. The effect of controlling the mineral potential during the flotation process is discussed.

Jan 1, 1996

By J. B. Gwiazda

In designing mine workings and adequate roof supports, it is indispensable to determine properties of the strata surrounding the workings. The proper selection of roof supports is based, before everything else, on a thorough knowledge of the roof rock strength, especially tensile strength. This parameter is a vital factor in the delamination of roof rocks and eventual cave-in. It is also important to know the properties of the floor on which the roof supports are set. While driving galleries, it is essential to know beforehand the strength of rocks that will be encountered in the near future during the work. The devices described will help to obtain this information.

Jan 1, 1987

By M. R. Paterson, M. K. Suddaby

Operational and economic factors tend to favor the application of continuous processes in preference to batch processes in the majority of mineral processing applications. In this paper, the requirement for a continuous elution process is established and the development of a large-scale pilot plant is described. It is based on a continuous elution apparatus that has a number of special features. These are the subject of various patent applications. The apparatus is suitable for adaptation to the AARL, Zadra, and other methods of elution. The performance of the atmospheric continuous elution system (ACE) is concluded to have significant advantages over the batch process.

Jan 1, 1988

By René E. Fuenzalida

The copper market has become highly competitive around the world; therefore, operations efficiency is a must for copper producers. The economic outcome in a day-to-day milling is not evident at the concentrator stage, when traditional procedures are used in operations management, based on the supervision of process variables and metallurgical indices. Therefore, in such a complex business, there is a strong need to devise an approach to furnish concentrator's operators with procedures and tools to state and solve the problems related to operational decision making, leading to maximize the economic return of investment; this is what is aimed to be presented in this paper.

Jan 1, 1998

By T. H. Dubaniewicz, J. E. Chilton

The U. S. Bureau of Mines is investigating the application of fiber optic technology to methane and carbon monoxide monitoring. A methane monitoring system is described that is based on differential absorption of infrared light. It can detect concentrations as low as 0.2% and as far away as 2 km via fiber optic cable. The upper range is 100% methane. The system requires no electrical power within the mine so it should operate safely under mine disaster conditions. A hybrid carbon monoxide monitoring system is described that combines a low-powered electrochemical cell with fiber optic telemetry. Testing has shown the system can operate maintenance- free for several months.

Jan 1, 1992

By P. J. Ryan

December 1988, Phelps Dodge Corp. brought an underground gold mine into production in the republic of Botswana. The mineralization is associated with steeply dipping shear zones in Archean greenstones of the Tati schist belt. The gold is intimately associated with sulfides and is refractory in nature. The deposit is being exploited by underhand open stoping methods. Vertical and inclined shafts are used to hoist the ore to the surface where it is treated by conventional crushing, milling, and flotation, followed by fluid bed roasting with cip elution and carbon regeneration. Adjacent deposits are being explored with a view to increasing production. Legal and taxation aspects of operating in Botswana are discussed.

Jan 1, 1991

By Michael J. Hannan, Thomas F. Torries

The nature of industry structure is examined and the structure of the aluminum, copper, lead and zinc industries are described by the use of factor analysis. Capacity is defined and the relationship between industry structure and capacity utilization is examined. It is demonstrated that excess capacity is possible in industries with competitive or monopolistic characteristics, but the effects of the overcapacity are vastly different in the two cases. In a competitive industry, cycles of overcapacity with attendant fluctuating prices are expected. A change in industry structure from monopolistic to competitive characteristics over time causes profits and economic rents to be transferred from the producer to consumers of the mineral.

Jan 1, 1989

By R. R. Pierce, C. E. Semler, K. G. Thomas, R. A. Williams, H. J. . H. Pieterse

This paper highlights the experience and knowledge gained at Barrick Goldstrike Mines Inc. with reference to the pressure oxidation circuit. Particular attention is given to the different linings employed on the autoclaves, flash and condenser vessels. Experience gained during the three years of operation is given. A test program and results are discussed. Quality assurance is highlighted. Barrick Goldstrike is 100% owned by American Barrick Resources Corporation, Toronto, Ontario, Canada.

Jan 1, 1993

By Paul E. Senseny

Data from triaxial compression tests per- formed at confining pressures between 0 and 100 MPa are used to evaluate parameters in a contitutive model that is based on microcracking. The tensorial formulation is reduced to equations that are specific to the simple Loading imposed in triaxial compression tests. Procedures for evaluating the model parameters are described, dependence of parameter values on confining pressure is discussed, and prediction of failure is assessed.

Jan 1, 1986

By K. Hansson, A. Jedborn

Background In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore, from a system's point of view, slurry transporta¬tion would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alterna¬tive examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasi¬biliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. l . The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then tempo¬rarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situ¬ations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe¬line, and personnel are needed. In a first phase (Fig. 2), a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1989

By Anders Sellgren, K. Hansson, A. Jedborn

In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore. from a system's point of view, slurry transportation would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alternative examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasibiliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. 1. The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then temporarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe- line, and personnel are needed. In a first phase (Fig. 2). a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1990

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Sublevel long-hole stoping is a large tonnage stoping method that may be used when there is a fairly com¬petent straight vein and the walls will stand with little support. It is a method that lends itself to a high degree of mechanization. Two variations of the method are illustrated in Figs. 1 and 2. Fig. 1 illustrates development typical for a 3-m (10-ft) wide vein. Initial development is by a lower haulageway, which may be either track or trackless. A two-compartment raise is driven in the footwall of the vein. At about 12 m (40 ft) above the main head¬ing, a footwall scram is driven and pockets through the vein are cut on 7.6-m (25-ft) centers. Each of the sub¬level blocks selected in this example is 45.7 m (150 ft) long by 12 m (40 ft) high. Stoping starts at the slot raises at stope endlines and retreats to the timbered raise. This retreat sequence starts on the lowest sublevel and the opening cut is blasted into the slot raise. This retreating system is used in order to provide a good safe back over the miners working in any given sublevel. A crawler-mounted down-the-hole or conventional pneumatic drill puts the holes down from one sublevel to the next on any desired drill pattern. Gravity flow brings the broken ore down to the scram level. The illustration shows a large slusher mounted at the timbered raise and slushing from the pockets back to the raise chute. Development of the ore block in Fig. 2 is typical for a 5.5-m (18-ft) wide vein. Initial development is by a trackless heading driven in the footwall of the vein. Pockets are then put into and through the vein on 7.6-m (25-ft) centers. For ease in using trackless rubber tired mucking equipment, these pockets are crosscut to the vein at approximately 1.04 rad (60°) rather than per¬pendicular. A two-compartment timbered raise is driven down from one level to the next. Endline slot raises are again put through from level to level. Each of the sub¬level blocks selected in this example is 40-m (130 ft) long by 33.5 m (110 ft) high. Stoping starts at the slot raises and retreats back to the timbered raise. Long holes are drilled from one sub¬level down to the next by starting between the lowest pair of sublevels and then working up the ore block. Mucking out of the pocket drawpoints is accomplished by rubber-tired load-haul-dump (LHD) units which then dump into a centrally located orepass.

Jan 1, 1982

By Robert D. Coyle, David L. Emmons

Introduction The Cove gold-silver deposit is located in Lander County, NV about 50 km (30 miles) southwest of Battle Mountain, NV (Fig. 1). The deposit is in the north-central part of the Fish Creek Mountains within the McCoy mining district. Gold was first discovered in the McCoy district in 1914 by Joseph H. McCoy (Vanderburg, 1939). From 1928 to the early 1930s, minor production occurred from shallow underground workings. Total gold production was less than 311 kg (10,000 oz; Stager, 1977). The district was inactive until the mid-1960s when it was explored by Bear Creek Mining, Summa Corp., Houston Oil and Minerals Corp., Gold Fields, and Tenneco Minerals. Tenneco began a detailed reevaluation of the McCoy deposit in the fall of 1984. Tenneco's decision to place the deposit into production was based on a geological reserve of 8.7 t (280,000 oz) of gold (Kuyper, 1987). Subsequent exploration and development drilling of the McCoy skarn deposit has more than doubled the mineable skarn reserves to 19 t (615,000 oz) of gold. In September 1985, Tenneco began development work on the McCoy gold-skarn deposit. The project is an open-pit, heap leach mine. The first gold was poured in April 1986. Echo Bay Mines acquired Tenneco's precious metals properties in October 1986. The mining operation and district exploration started by Tenneco is continuing under the Echo Bay's ownership. The Cove deposit was discovered in January 1987 and production began in February 1988. The deposit will begin as an open-pit, heap leach operation. It will be converted into an open-pit and underground mine with a 9-kt (10,000-stpd) mill. Geologic setting The Cove gold-silver deposit is situated on the east flank of the Fish Creek Range. Surface elevations range from 1460 to 1680 m (4800 to 5500 ft). The deposit is 1.6 km (1 mile) northeast of the McCoy gold-skarn deposit. The Cove deposit is hosted by the Augusta Sequence of Triassic age. Three formations of the sequence are mineralized at the Cove. In ascending order they are the Favret, Panther Canyon, and Augusta Mountain formations. The Favret formation is a gray, locally silty limestone that is commonly fossiliferous. The base of the formation is not exposed. Exposed and drilled intervals are more than 60 m (200 ft) thick. Conformably overlying the Favret formation is the Panther Canyon formation. This formation is 215 m (700 ft) thick and consists of a siliclastic and carbonate sequence of conglomerate, quartzite (sandstone), mudstone, limestone, and dolomite. It is a distinctly coarsening-upward sequence of rocks. The top of the unit is conglomerate with chert and quartzite clasts up to 150 mm-diam (6 in.-diam). The Augusta Mountain forma-

Jan 8, 1988

By Harold Stocker

INTRODUCTION This presentation is concerned with the administrative and technical capability of the Atomic Energy Control Board (AECB) to assure compliance with the individual exposure limit for radon daughters (currently 4 WLM per year). It is not concerned with the epidemiological bases for setting the exposure limit. Moreover, the intent is to show how sophisticated methodologies and advanced technologies, applied to radon daughter concentration measurements in uranium mines, convey the spirit of compliance by providing better estimates than do the historical methods. These better estimates mean that more accurate and more precise estimates of each worker's exposure are determined using these more modern methods and devices. The estimates so derived should provide more convincing evidence to an individual worker that his assigned exposure is a valid indicator of his true exposure. In addition, a perspective on the exposure estimate in relation to the exposure limit is given as further evidence that an exposure limit is not the dividing line between "safe" and "unsafe" exposures. A brief description is given of the compliance aspects of the Atomic Energy Control Regulations and of the limitations of purely statistical non-compliance procedures. Most of the emphasis of the paper will be placed on the uncertainties associated with conventional radon daughter exposure determination and the means being employed (and anticipated) to reduce these uncertainties. NON-COMPLIANCE Under current Atomic Energy Control Regulations (1978), the annual individual exposure limit for radon daughters is given without reference to the possible methods of sampling and calculation of radon daughter exposure and without any reference to possible uncertainties or their magnitudes. This is common in such statutes, the details of sampling, calculation, error analysis, and so on, being left for licence conditions or provided as a specific guideline to the licensee on the matter of compliance with the Regulation. Since the exposure limit is contained in the Regulations, compliance with it is absolute, as with any other law. In Canada, a state of non-compliance with the radon daughter exposure limit exists when an exposure (attributed to an employee) is reported by the licensee to exceed the limit. No uncertainty in the measurements or in the overall determination of exposure is reported nor is any requested. Removal of the worker and the loss of his services are the immediate and direct penalty suffered by the licensee for failure to maintain the exposure at, or below, the limit. A worker may be re-instated to employment for the balance of the reporting period only if the licensee can assure the AECB that further significant exposure to the worker will not ensue. In other jurisdictions, such as the United States, non-compliance is defined on a statistical basis. For example, NIOSH, the National Institute for Occupational Safety and Health presents procedures for calculating the 95% Lower Confidence Limit (LCL) in order to "compare the results of occupational environmental sampling to an occupational health standard and make a decision with a known chance of making an incorrect decision that a state of non-compliance exists" (Leidel and Busch, 1975). (In the nomenclature of this presentation, exposure limit would be used in place of "standard", in the NIOSH sense). Furthermore, it is emphasized in the NIOSH document that such numerical comparisons "are necessary only if the sample mean is greater than the standard". The NIOSH document points out, quite correctly, that the "statistical procedures presented below will not detect and do not allow for analysis of highly inaccurate results, i.e., systematic (non-random) errors or mistakes ... If a systematic error is known to exist in an instrument or analytical procedure then correct the sample mean of the data before analyzing for non-compliance". It is certainly not the purpose of this paper to criticize the sophisticated statistical approach to non-compliance as given in the NIOSH document or in similar approaches used in other jurisdictions. Rather, the purpose is to approach, with some introspection, the question of the determination of exposure by the employer for his employee and especially the employee's understanding of, and confidence in, the accuracy of the exposure determination and its relation to the exposure limit. DETERMINATION OF EXPOSURE Historically, in uranium mines, exposure to radon daughters for an individual miner

Jan 1, 1981

By Richard L. Bullock

At the beginning of the previous chapter, variation in the types of room-and-pillar stopes were briefly identified so that the reader could begin to understand the extensive application of this mining system. These same variations will be reviewed and elaborated upon in this chapter while discussing the methods by which the material is extracted. FULL-FACE SLICING BY DRILLING AND BLASTING When the entire mineralized thickness is taken in one pass of mining, it is known as full-face slicing. There is no mineral of economic value intentionally left either in the floor or the back. Theoretically, there is no limit as to how high the face could be in a single pass. But there are practical limitations to equipment size, pillar stability, and control of local loose slabs on the face, roof, and pillars that dictate current practices. The range of face height of the 15 typical room- and-pillar mines covered in Table 8 (Sec. 1, Subsec. 7) from the Dravo report, 1974 was 1.7 m (5.5 ft) to 9.8 m (32.1 ft), the average being 5.2 m (17.1 ft). In the Gaspe copper mine, Murduchville, Que., the drill jumbos were constructed with an extendable tower for full-face mining up to 15.2 m (49.9 ft) if necessary. However, a more common practice at that mine is to take the ore in multiple slices (Hall, 1959). In many metal mines the normal practice when starting stoping is to drive a single development drift into the ore zone a distance that will allow at least four or five rooms to be opened on each side of the initial drift. Since this opening will probably serve as the main haulage drift, it should be kept as straight and as level as possible. As to the other criteria for the stope developments, it might be well at this point for the mine planner to review the remarks presented earlier (Sec. 1, Subsec. 7) on drifts, entries, and crosscuts for production. After the initial drift is driven, it should be slabbed to the full room width if it is not that wide already. Next, the future pillars should be marked on the ribs and the rooms driven between them. If the ore extends beyond the length of the initial drift and is to be mined in the normal sequence of mining, care should be taken in a random room-and-pillar system to see that the extension of the initial roadway remains straight. If pillar "spotting" is left to the shift foreman or to the miners, they probably will not realize the importance of maintaining a uniform pillar line next to the roadway, and of keeping the line straight. Invariably, a pillar will end up in the middle of the future drift extension resulting in a "dogleg" in a possible main haul- age road. While this error is not catastrophic, too much weaving in and out of pillars will certainly slow down future production haulage and needlessly increase costs. Face or Breast Drilling and Blasting Practices To drill and blast the initial advances into the rock, usually some form of cut pattern must be used. In room-and-pillar mining, when a pattern is drilled in a face and that face is the only exposed surface to which all of the rock must be broken, the pattern is known as a "face round" or "swing." The face to which the rock must be broken is referred to as a "free face." Thus the most common way to advance a room into virgin rock with only one free face is by drilling swings. If, after breaking around a pillar, taking down back, or taking up bottom, there is a second (or third) face exposed, a group of holes drilled nearly parallel to a free face and breaking to it is known as a "slab round" or "slabbing." In some mining districts it is also known as "slashing." Obviously, because two faces exposed offer less resistance to fragmentation than a single face, slabbing requires less drilling and fewer explosives than required to break the same tonnage with the same degree of fragmentation with swings. Therefore, it behooves the driller (or supervisor) to plan the combination of

Jan 1, 1982

By R. N. Torbin

Foster-Miller has recently developed a new concept.for a highly specialized shaft lining system. This innovative shaft lining system uses a continuous slipforming placement technique. lining the shaft from the bottom up such as in a raised bored shaft. Although this part is not unique, the concept is designed to use polyurethane foam instead of concrete for the liner material. Furthermore, the concept is designed to attain rates of advance of 30 m/hr. The polyurethane foam liner is designed to be thinner (100 mm) than a conventional concrete liner (300 mm). It will eliminate the high cost associated with the erection and operation of an on-site concrete batch plant. The system will be designed to be operated remotely, requiring few personnel for its operation. This system (or parts of it) may be appropriate where fast construction speed is required or where a jointless liner is needed.

Jan 1, 1988

By S. Dixon

The objective of any business is to maximize the value derived from the commodity produced. There is a point where additional expenditure does not return a value sufficient to justify the additional expense. The treatment of high acid consuming copper ores is a perfect example of this principle in which the relationship of acid consumed to the percentage of Copper extracted is vital to the economic value of the final product. This paper describes a method of evaluating column leach test data to determine the economic optimum conditions for leaching high acid consuming copper ores. The effect of acid consumption on the net economic value of the Copper extracted is very significant in the case of high acid consuming ores. The focus of this paper is the relationship of acid concentration in the raffinate, and acid cure dosage to the economic value of a leach cycle.

Jan 1, 2004

By S. Veeramasuneni, Qingxia Liu, Susan Dichter

Gypsum (CaSO4-2H2O) is an abundant and naturally occurring mineral. Under elevated temperatures (250 to 300°F) gypsum loses its crystalline water and is converted to hemihydrate (CaSO4 -0.5H2O). The hemihydrate form is also called plaster that is widely used in wall- board, joint compound, and many other industrial applications. At temperatures higher than 450°F (up to 1,200°F), gypsum loses all crystalline water and is converted to anhydrite form (CaSO4). The loss of crystalline water at elevated temperatures is an endothermic reaction. This property makes gypsum an attractive filler and fire retardant in certain thermosetting plastics. The applications of gypsum in plastics, food, pharmaceutical, and paper coatings will be reviewed in this paper. The future of synthetic gypsum, a by product of various industrial processes, will also be discussed.

Jan 1, 2003

By Mark Cadzow, Rodney Lamb

CARBAD is a recently developed Australian process for the recovery of gold and platinum group metals (PGM). The process relies on the selective agglomeration phenomena of the gold and PGM to capture the mineral in a carbon-hydrocarbon media. The precious minerals are then extracted from the loaded media by a solvent gold room process producing a relatively pure concentrate. The media is then reformed and recycled. The process development from batch test work to a 5 t/h commercial scale CARBAD plant is discussed. The operation and results of processing a variety of ore types at several alluvial and hardrock gold mines is examined. The operating and capital costs for the CARBAD process shows that it is significantly less than that the conventional CIP/CIL. The environmental safety of the process is a major feature where precious metal deposits are located within urban environments or within wilderness areas.

Jan 1, 1989