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By E. Bonates

Truck/shovel systems are a major form of materials handling in open-pit mines in spite of the high associated capital and operating costs. Mine management is usually faced with a decision to purchase new equipment in order to support the steadily increasing on production. These decisions involve a high degree of uncertainty with respect to future demands and technical developments. This paper describes a computer simulation model for evaluating the optimum equipment fleet in an open-pit mining operation. The system is based on a stochastic simulation program which accounts for the various mining and economic parameters. It calculates the type of equipment and the optimal number of trucks needed to meet the production targets. Two types of selection criteria: maximum rate of production or minimum cost can be applied.

Jan 1, 1989

By D. J. Dunn

Grinding balls, simple in shape, are actually sophisticated machine components, and essential elements of mineral grinding mills. Their physical properties and quality must be adequate to achieve design life and grinding mission with few failures. Consistant properties and freedom from breakage are essential. Size, mass, hardness profile, chemistry, and microstructure are important variables which should be designed, specified, controlled, and verified to achieve optimal mill performance. Failure to specify ball properties in detail or failure to obtain required ball properties and quality can cause milling productivity losses and excessive costs that could endanger a milling operations survival in the competive mineral industry.

Jan 1, 1989

By Betty L. Gibbs

The ultimate objective of mine design is to determine the development and untilization of mining resources. Computer Aided Design (CAD) and automated mapping are becoming very popular in engineering and in mining. Any computer program which produces a design or model with the interaction of an engineer can be considered computer-aided design (CAD), but' the term CAD is often associated with interactive graphic design and mapping capabilities. Graphics can be effectively used to describe and model two-or-three dimensional mine workings, orebodies, and geologic structures. Interactive computer graphics is applied to mine design with a combination of deposit modeling techniques and the use of AutoCAD for interactive graphics and mapping. The combined system can perform a large part of the calculations and drafting work for geologic and mining design. Using computer calculation and graphics techniques, comparisons of alternatives, can be made quickly.

Jan 1, 1989

By Bruce H. Van Brunt, Danny L. Taylor

Estimation of local recovered tonnage and its grade above a cutoff is of paramount importance during the exploration and development stages of any mineral deposit. The spatial dis- tribution of grades as well as the cumulative distribution must be defined. In this paper, Probability Kriging, an estimator based upon rank order statistics and the estimation of the distribution of point grades within a panel, is applied to the Manhattan gold deposit. Mapping the conditional probabilities that estimated panel grades exceed a cutoff is useful to the understanding of the distribution of mineralization in a structurally controlled gold deposit.

Jan 1, 1989

By C. J. Gilbert

The mining method at the Cannon Mine uses heading and bench stopes, which are subsequently backfilled with cemented high-strength fill prior to pillar removal. The dimensions of the pillars, 24 feet wide by G5 to 95 feet high, re-quire careful blasting in order to not damage the rock, compromising pillar strength and complicating later removal. Additionally, the proximity of dwellings makes vibration control a must for good-neighbor relations. A blast monitoring study was conducted in order to optimize bench blasting practices to maximize fragmentation, reduce costs, minimize well damage, and minimize ground vibration.

Jan 1, 1989

By Gerald L. Finfinger, Michael A. Trevits

Since 1964, the US Bureau of Mines has been conducting a comprehensive program of research on the origin and control of coal bed gas. Several extensive databases have been developed during the work representing information from every major coal producing region in the United States. Gas content measurements have been made on 885 coal samples from 15 states representing 221 coal beds. Records indicate that the range of gas content varies from zero to over 640 ft3 per ton (20 cm3 per gm). During the conduct of the research program, several drilling and completion techniques have been developed or refined to divert and collect coal bed gas before it enters the mine environment. Such techniques include horizontal, cross-measure and vertical boreholes. Samples of gas collected from boreholes drilled under the program indicate that the primary gas constituent is methane which comprises 97 pct, the remaining hydrocarbons make up approximately 0.2 pct. Nitrogen and carbon dioxide account for the majority of the remaining gases at average concentrations of 1 pct each. Surveys have been made on the total daily gas emissions from underground coal mines. In fact, the most recent survey indicates that 180 mines had emissions in excess of 100 Mft3 per day (2.8 Mm3 per day). Furthermore, total mine emissions were in excess of 300 MMft3 per day (8493 Mm3 per day). An analysis of the data showed that there appears to be no relationship between in-place gas content and coal bed gas composition. Furthermore, no correlation was observed between in- place gas content and daily mine emissions. In a general sense, the higher the gas content the higher the relative gas emissions. Perhaps a more direct relationship would be observed if more complete emissions information (including gob gas borehole production) was available.

Jan 1, 1989

By Y. J. Wang

The ventilation network problem, having both fixed and nonfixed operating points as control devices, is examined in this paper. Emphasizing the concept of solving a system of n equations in n unknowns, this paper presents the generalized problem formulation, some conditions under which the problem does not have a unique solution, and a solution procedure based on the Newton (or Newton-Raphson) method. A numerical example is included.

Jan 1, 1989

By Robert A. Jankowski, J. Drew Potts

The Bureau has identified several ventilation techniques and a water spray application for reducing auger mining dust levels. The dust control methods all work by improving air quantity or flow patterns at the face. Section airflow is maximized by reducing air leakage between entries. Sealing stoppings with brushed- on mortar, containing fiberglass additives, reduces leakage. Heading airflow is maximized by properly installing the brattice curtain and keeping the area between the curtain and rib free from obstructions. The brattice curtains should be hung at least 1.7 m (5.5 ft) from the ribs in 7.9 m (26 ft) wide entries and kept less than 4.6 m (15 ft) from the face. A tight curtain seal is created by using the mine air pressure to force the curtain against the timbers, roof and floor of the mine. Using both a blowing curtain and an exhausting curtain will isolate some face workers from dusty air. The Bureau designed, auger clearer spray system, can be used to force heading air across the active face. Laboratory tests indicate that the auger clearer spray system can increase face velocities by more than 0.5 m/sec (100 ft/min), while using less than 0.13 l/sec (2.0 gal/min) of water.

Jan 1, 1989

By Dennis A. Swyt

In current post-industrial America, U.S. manufacturing is undergoing major changes both in the technology of its production and in the size and composition of its workforce. At the same time that computers and robotics move the shop floor toward becoming the unmanned "factory of the future" [1], they increase the need within the overall production enterprise for new types of technicians and engineering professionals [2]. Based on previous works which examined, first, the evolution of the technology of manufacturing toward the flexibly-automated factory of the future and, second, the technical- professionalization of U.S. manufacturing sectors [1,3], this paper looks briefly at the workforce of the U.S. metals industry.

Jan 1, 1989

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 Russell B. Dix

Laboratory column leach tests are used routinely to evaluate gold and silver ores for heap leaching. The tests are fairly straightforward. The ore to be tested is placed into an appropriately sized leach column. Then an alkaline cyanide solution is applied to the ore and the effluent solutions are collected. The effluent solutions are then put through a column of activated carbon to recover the gold and silver in solution. Zinc precipitation may also be used for silver ores or if the ore contains a significant amount of soluble silver in addition to gold. The length of the leach time in the tests is mainly a function of how quickly the precious metals are recovered from the ore. At the completion of the tests, the tailings are assayed and the recoveries determined from the calculated head grades. The amount of ore to be used for a test is partly a function of its crushed size. The coarser the size, the larger the sample size for the column. A typical small-scale laboratory column test on ore crushed to -12 mm (-0.5 in.) would usually be conducted in a 100- to 150-mm-diam (4- to 6-in.-diam) column 2 m (6.6 ft) in height, containing about 20 to 45 kg (44 to 99 lb) of ore. A comparable test on the same ore crushed to 50 mm (2 in.) would be conducted in a 200- to 300-mm-diam (8- to 12-in.-diam) col¬umn 2 m (6.6 ft) in height, containing about 80 to 200 kg (176 to 441 lb) of ore. Larger-scale tests can be run by increasing the column heights and diameters. Conducting larger-scale tests results in increased costs for sample procurement, preparation, and running of the tests. The additional data obtained from

Jan 1, 1989

By R. B. Coleman, T. P. McNulty

HISTORICAL PERSPECTIVE Environmental protection is not a newcomer to the minerals industry, there having been numerous examples dating from decades ago. Sometimes the awareness of a need for preservation or restoration derived from the ethics and principles of the engineers and management responsible for development and operation of a property. On the other hand, pressures to clean up industrial messes frequently were applied by special interests such as farming and ranching, recreational and conservational groups, and local governing bodies. Mineral industry environmental controls in this century have therefore been a patchwork of good intentions and bad, clean operations and dirty ones. The Hoboken siphon converter for converting copper matte while effectively capturing sulfur dioxide at a relatively high concentration was developed in Olen, Belgium in 1928, a full sixty years ago. The alternative was closure of the Hoboken smelter by Belgian government health officials who had become intolerant of sulfur dioxide emissions which were causing serious respiratory ailments in the river valley in which Olen is situated. As early as 1958, the Anaconda Company had a full-time botanist on its metallurgical research staff with the responsibility of learning how to vegetate the extensive tailings deposits at Anaconda, Montana. By that time, considerable progress had already been made in restoration of Silver Bow Creek, which at one time was “dead” for as far as 40 miles down- stream from the point of entry of acid mine drainage in Butte.

Jan 1, 1989

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By G. G. Marino

Little comprehensive information has been reported on the behavior of room-and-pillar mines. The objective of this paper is to present case data on mine failures in the Illinois basin for use in practice. Presented are results of an ongoing study and details on the site characteristics of cases where sags have developed on the surface. Site data are reported to show the geologic, mining, and sag conditions that existed. Sags mainly develop from pillar, floor, or pillar-floor failure. The character of the sags depends upon the type of mine failure as well as the overburden response.

Jan 1, 1989

By T. O. Kuhl

The Royal-Mountain King Project is 7.2 kilometers (4.5 miles) northwest of Copperopolis, California. Reserves of 8,250,000 t (9,100,000 st) at 2.32 g/t (0.068 oz/st) gold will be processed at 2500 t/d (2750 st/d) from three pits. Production is scheduled to begin the first quarter of 1989. District geology consists of northwest trending sequences of Jurassic meta-sedimentary and meta-volcanic rocks. Three ore bodies are located along the Hodson fault which trends northwest and dips 40 degrees northeast. The Gold Knoll and North Pit ore bodies are hosted in carbonate and silica altered meta-volcanic rocks whereas the Skyrocket ore body is hosted in quartz stockwork zones in meta-sedimentary rocks. Gold is present as inclusions and micro-veinlets in pyrite.

Jan 1, 1989

By H. R. G. Steeghs

Recently developed organic binder systems have the advantage of eliminating bentonite contamination. This results in better chemically reactive taconite pellets. Both acid and fluxed pellets show increased R40 reducibilities, whereas mechanical and physical properties of the taconite pellets are not adversely affected. Blast furnace performance of these pellets is excellent, and substantial economic benefits are achieved.

Jan 1, 1989

By R. E. Brewer

Pilot scale flotation tests were conducted with a 1-ft-diam column at the Barrick Resources (USA) Mercur mine, Tooele County, UT, to scavenge unleached (0.01 to 0.018 oz/st) gold values from the tailings of a carbon-in-leach (CIL) circuit. The optimal recovery was 35 pct of the unamenable-to-leaching gold, at a grade of 0.353 ounces of gold per short ton, using a mixture of dithiophosphate, xanthate, and diesel fuel as a collector. This was achieved by floating approximately 1.6 wt pct of the CIL tailings. The gold values are believed to be contained in arsenic-bearing minerals, attrited carbon, and carbonaceous material in the ore. Redox potentials were measured with a gold electrode during the conditioning stage prior to flotation. The correlation between flotation recovery and redox potential values could be used to control reagent additions. Treatment of a previously leached and weathered tailings material, averaging about 0.055 ounces of gold per short ton, was also investigated using the column. The tailings were first water leached at approximately 600 C, and the water-leached tailings, which contained approximately 0.03 ounces of gold per short ton, were then fed to the column flotation unit. The effects of pH and sodium sulfide addition on gold flotation recovery and redox potentials were examined. maximum gold recovery, 36 pct, was achieved at pH 8.5 with a sodium sulfide addition of 0.18 lb/st.

Jan 1, 1989

By S. P. Mehrotra, P. C. Kapur

A phase-species classification is suggested for the mathematical models of flotation kinetics. Important models are discussed and the structure of their rate constants and parameters is analyzed. Split or enhancement factors are formulated for different kinetic models. The steady state dynamics of the flotation circuits can always be described in terms of these factors. For simulation of circuits, feed is characterized in a formal manner starting from a multivariate distribution of particles in flotation rate, size and mineral content. Synthesis of optimum circuits which simultaneously searches for the best process variables and structural parameters from a generalized master network is described. Trends in the areas of search algorithms, role of middlings and feed identification etc. are indicated.

Jan 1, 1989

By H. Haas

European coal mines are facing a very difficult situation. Despite a steadily increasing world energy consumption, the sales of European coal mines have been going down and production has to be adapted. The reason is the collapse of energy prices aided by the decreasing rate of U.S. $ to Deutschmark in the last few years. The problems are not due to the development of European mines themselves. Assured energy supplies are an important condition for the health of a nation's economy. Contribution of domestic energy production ensures some degree of independence from world market fluctuations as to supply volumes and price, and maintaining jobs and should be achieved within the framework of economic constellations. Governmental energy programs are planned towards meeting this end.

Jan 1, 1989

By Syd S. Peng

Coal mine ground control is the science of the application of rock mechanics principles to coal mining operations. More acceptance of ground control techniques by the US coal mining industry has been seen in the 1980s. It has been demonstrated many times that a proper ground control design is key to a successful coal mining operation. However, numerous in-mine applications of various ground control techniques by the author indicate that cur¬rent techniques need improvement. The biggest problem in coal mine ground control is that miners are dealing with materials of rapidly varying properties and occurrence. And only the exposed surface can be identified positively. Thus, in attempting to solve the ground control problems, many unknown factors are assumed. Some of these problems deriving from such assumptions are discussed here. Coal pillars There are two types of coal pillars - barrier and chain pillars - depending on the purpose. Barrier pillars are those located between panels or sections or mines. Their objective is to isolate the mined-out effects of one from the other. Chain pillars are characteristic of room-and-pillar mining. They are designed to support the overburden load and protect the surrounding entries and crosscuts. Pillar design is the most popular and, frequently, the only ground control technique employed by coal operators for mine design. There are many pillar design formulae available (Bieniawski, 1983; Peng, 1986). Nearly all consider both size and width/height effects. Specifically, pillar strength decreases with increases in pillar size and increases with increase in the ratio of width to height. However, the manner of those changes varies with investigators. Consequently, the pillar size determined varies considerably from investigator to investigator. And the variation grows much more pronounced as the seam depth increases (Fig. 1). The question, then, is which one should be used? One may say each formula is developed for a particular coalfield. A particular formula may have been used to design a pillar, which has not failed. Does that mean it was over designed? Not necessarily, because no one can confirm its stated "safety factor." Recent case studies (Peng, 1986; Tang and Peng, 1988) show that all pillar formulae are too conservative for chain pillar design in room-and-pillar mines due to the following two problems: • Uncertainty about in situ pillar strength. All formulae employ uniaxial compressive strength of selected samples tested in the laboratory and scaled to that for underground pillars. The scaling factors are not well proven. In situ testings by Bieniawski (1969) suggested that 1.5 m (5 ft) pillars are representative of in situ pillars that, after being tested in various US coal mines, are not universally applicable. • Uncertainty about the actual loading on the pillar. The most frequently used tributary loading theory defines the loading as total overburden weight tributary to each pillar (from roofline to surface). This is too conservative when it applies to coal measure rocks. The overburden consists of layered strata of different stiffness. Many stiffer strata will self-support over small roof spans. Depending on the location of these strata above the coal seam, it will reduce pillar loading in various proportions. It follows that total structural analysis using, for instance, finite element method is more appropriate. It can take into account the interaction of roof and floor strata with coal. This has been demonstrated by Peng et al. (1986,1988) on several case studies. However, the validity of the finite element structural analysis depends on the accuracy of in

Jan 1, 1989