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By G. A. Rowell, T. M. Barczak, J. C. Zelanko

The US Bureau of Mines conducted a multifaceted geomechanics research project at a bump-prone longwall coal mine in eastern Kentucky. This mine presented a unique opportunity to study coal mine bumps because the problem appeared to be confined to areas beneath a sandstone channel of limited areal extent. The Bureau's research plan sought to exploit the uniqueness of the site by monitoring strata behavior and longwall support reactions as the longwall face approached and passed beneath a sandstone channel. The study incorporated static pillar stress and strata deformation measurements, shield support leg pressure measurements and microseismic monitoring, as well as geologic mapping and the collection of observational data. Data collection was underway when a major coal bump occurred on the longwall face at the mine. Pillar behavior (load and deformation), shield loading and microseismic activity were monitored prior to, during and after the bump. Observations in the tailgate indicated that these pillars were failing prior to and following the face bump. Shield load data, evaluated in an effort to identify trends that could be used to forecast coal bumps, appear to provide a measure of the effectiveness of destressing operations conducted on the face (e.g. shot firing). Analyses of microseismic data revealed anomalous values associated with microseismic "energy" prior to the face bump. In addition, microseismic source-location data and underground observations provided some insight as to the sequence of pillar and face bumps.

Jan 1, 1992

By N. Ohmura, H. Saiki, T. Nagaoka

A microbiological flotation system that facilitates the removal of pyritic sulfurfrom coal is reported. The chemoautotrophic bacterium Thiobacillus ferrooxidans is utilized as a pyrite-selective biological surfactant that augments the flotation-mediated rejection of pyritic sulfur. The bacteria were produced by continuous cultivation in 50-L tanks, where it was possible to produce 3.8 x 1012 cells/day when the dilution rate was optimal and the pH was 1.6, a condition that prevented ferric precipitation. Prior to flotation, the bacteria were mixed with coal, which allowed cells to selectively adhere to the pyrite. The adherence of bacteria made the pyrite surfaces more hydrophilic, which facilitated separation by increasing the tendency of pyrite to sink, while the hydrophobic coal floated. When a synthetic coal/pyrite mixture (10.1 % sulfur) was subjected to microbial flotation, the level of sulfur rejection was stable during the entire period of operation (i.e., one hour). In the absence of bacteria, flotation reduced the sulfur content of a coal/pyrite mixture to within the range of 6.7% to 7.4%, with the combustible recoveries ranging from 93.9% to 99.0%. In the presence of bacteria, the sulfur content of cleaned coal was in the range of 1.3% to 1.4%, with the combustible recoveries ranging from 80.8% to 86.3%. Bacterial activity accounted for 57% of the total sulfur rejected through flotation. The desulfurization capacity of the system was further confirmed by testing it with run-of-mine coal. The pyritic sulfur content was reduced from 2.9% in the feed coal to around 1.2% in the froth product, while 80% of the combustible recovery was retained. Microbial flotation removed 60% of the pyritic sulfur from the feed. These findings demonstrated that micro¬bial flotation could be utilized for the removal of pyritic sulfur from coal in a semicontinuous system.

Jan 1, 2000

By Edward C. Heistermann

The Falkirk Mining Company, a wholly owned subsidiary of The North American Coal Corporation, is located 80 kilometers (50 miles) north of Bismarck, North Dakota. Falkirk is under contract to provide 5.08 mil- lion tonnes (5.6 million short tons) annually to the Coal Creek Generating Station, an 1100 mega- watt electric plant owned by the Cooperative and United Power Associations of Minnesota. Coal Creek is located 2.4 kilometers (1 ½ miles) south of the Mine. Electricity from the plant is transmitted to rural electric customers in Minnesota via a 400-kilovolt D.C. power line.

Jan 1, 1982

By W. H. Langer

Natural aggregate is the number one non-energy mineral resource in terms of value and volume and is a necessary commodity for sustainable communities. While aggregate is a non-renewable resource, supplies are nearly inexhaustible on a global scale. Natural aggregates have characteristics that differentiate them from most other mineral commodities: a high number of potential extraction sites, a high volume to value ratio, significantly different set of potential environmental impacts, and regional importance combined with a narrow economic transportation radius. Because of these and other differences, resource management policies for aggregates should differ from ?general? mineral resource policies, even though the end goal is the same for both ? ensuring sustainable resource management. To achieve the goals of sustainable aggregate management each stakeholder ? government, industry, the public, and non-governmental organizations ? will need to accept certain responsibilities, one of the most important of which is to become informed about natural resource issues. Sustainability is about making trade offs among competing objectives, and people disagree about the appropriate balance among the goals of economic prosperity, social equity, and environmental health. One of the primary purposes of mineral indicators is to provide information to decision makers and the public so as to ensure that the public debate about policy choices is grounded in fact. We identify the unique aspects of aggregate resource management and introduce the concepts of sustainable mineral resource management. We then present mineral indicator sets for aggregate resources from Italy, the United Kingdom, and Slovenia to demonstrate that effective mineral indicators should be specific to the resource in question and the public values and objectives.

Jan 1, 2003

By Edward S. Plotkin, Joseph D. Guertin

In a continuing effort to upgrade tunnel design and construction practice, the New York City Transit Authority sponsored a geotechnical monitoring program on a recently completed section of the E. 63rd Street line in mid-town Manhattan. The purpose of the program was to document rock behavior as well as temporary and final rock support systems to provide a possible basis for either confirmation or modificification to design methods and construction specifications. This paper describes those portions of the study associated with rock behavior and temporary support systems. The entire study included monitoring of rock movements above the tunnel as well as adjacent to a 36.6m (120 ft) shaft, monitoring of stresses in steel ribs, monitoring of loads in rock bolts in two test sections and monitoring of stresses in the final concrete lining. PROJECT DESCRIPTION Route 131-A, Section 4 is located in mid-town Manhattan extending from under the Central-Park Zoo on Fifth Avenue eastward under 63rd Street to the end of contract between Madison and Park Avenue. Figure 1 is a Project Locus Map. The contract included approximately 355m (1165 l.f.) of 4 track tunnel(13.4m (44 ft) wide by 13.7m (45 ft) high excavated dimensions) plus a 22.9m (75 ft) by 19.8m (65 ft) by 36.6m (120 ft) deep ventilation shaft. The entire tunnel and most of the excavation shaft was excavated in metamorphic rock, known locally as the Manhattan Schist. The depth to tunnel invert varied from approximately 27.4m (90 ft) at the west end of the contract to 35.Om (115 ft) at the east end. Rock cover above the excavated crown varied from approximately 6.7m (22 ft) at the approximate mid-point of the contract to 18.9m (62 ft) at the western end. A plan view and interpreted geologic profile are illus-

Jan 1, 1979

By James Veri, Andrew H. Stern, Gerald N. Torbert

When mining coal, the occurrence of high methane liberation rates can often overwhelm the existing ventilation capabilities of the working place and result in elevated levels of methane gas. These “gas-off/stand-by condition" events often limit the production of the section and increase the potential opportunity for the occurrence of an ignition. This paper investigates the potential effectiveness of using a continuous miner equipped with an integral dust scrubber which was modified to provide high flow rates to reduce these events during active coal mining. Work was undertaken to quantify various adaptations of the mining system, by simulating site parameters in the controlled environment of an instrumented "full-scale" test facility. Monitoring (in real-time) of velocities and gas concentrations at many fixed points was also undertaken in an attempt to learn the effective limits of the system at high release volumes of gas.

Jan 1, 1997

By Daniel A. Servigna, Pedro V. Porcayo, Humberto F. Preciado

This case study presents successful conversions of two conventional slurry tailings storage facilities (TSFs) to filtered tailings storage facilities in Durango, Mexico, for two mining operations with production rates ranging from 1,200 to 4,500 t/d (1,320 to 4,960 stpd). In both cases, the decision to make the conversion was driven by: • Ability to increase the storage capacity via stacking of filtered tailings. • Reduction of the tailings’ impacted footprint. • Mitigation of long-term geotechnical stability risks. • Alleviation of environmental risks by reducing tailings water as a source of pollutants, and a net reduction in process make-up water demand. The case studies can be divided in two classes: class 1 facilities, where slurry tailings are loaded almost immediately with filtered tailings because of storage area limitations; and class 2 facilities, where the conversion is immediate but the available area allows gradual or eventual loading of the previously stored slurry tailings. The condition where slurry tailings must be loaded almost immediately represents the greatest engineering challenge because of the geotechnical risks of instability and liquefaction (under both static and dynamic loads). Regardless of whether these facilities are converted immediately or eventually, they all require monitoring of pore water pressure conditions within the slurry tailings being loaded as the depth of placement of the filtered tailings above the slurry tailings increases. Additionally, where loading rate exceeds the pore pressure dissipation rate, tailings dewatering may be required to increase the rate of consolidation and strength and mitigate the risks of foundation failure.

Jul 1, 2025

By Richard C. Harris

Substantial economic benefits exist today for fine coal preparation and improved technologies coupled with more stringent air quality standards should increase those benefits in the future. The costs of cleaning can be determined with reasonable certainty with proper consideration given to mining methods, plant feed characteristics, refuse disposal and marketing impact. The benefits cannot be as accurately determined but include impact on generating capacity, plant maintenance, emission standards, slagging, fouling, transportation, and handling. Site specific cleaning offers particular benefits in some situations to tailor the preparation plant facility to boiler characteristics. Improved communications and methodologies need to develop to allow the coal producer and the coal user to clean coal to the degree that results in the lowest overall cost for electrical generation.

Jan 1, 1988

By Anthony D. Hammond

Geographic Information Systems (GIS) are powerful analytical and data display tools for mining industry management, capable of linking spatial features with tabular data and integrating communications across departmental boundaries. This GIS application concentrates in four technical areas of underground mining: ownership and mineral claims; exploration management; production; and mine safety. GIS provides land and exploration managers with the ability to consolidate and archive claim data and to georeference property maps defined in local coordinates into world coordinate systems. It facilitates optimal siting and querying of service installations and refuge chambers relative to production centers. GIS's proximity analysis identifies areas of potential safety concern and points to and maps the shortest routes to emergency exits, facilitating the prompt evacuation of mine personnel.

Jan 1, 2002

By Wilson Blake

A mathematical model based on the finite element method of stress analysis has been used to describe the behavior of the western wall of the Kimbley Pit as its slope was steepened from 45 degrees to 57 degrees. Theoretical stress changes because of slope modification are slight except at the toe of the modified slope. Magnitudes of stress created are small when compared with the strength of a competent rock. Resulting displacement patterns point out the need for determining the in situ state of stress at a site prior to excavation or modification. A comparison of model results with field measurements shows that, in general, the model has described the measured behavior of the west wall of the Kimbley Pit remarkably well.

Jan 1, 1968

By Tore M. Larsen

The Rødsand Gruber mining company is a subsidiary of the Norwegian industrial corporation Elkem AS, and its mine is situated in the northwestern part of southern Norway. The ore from this mine is a relatively low grade, magnetite ore with a small amount of vanadium. The ore bodies are very irregular in shape, resulting in a high amount of waste rock intrusion in the stopes. The annual underground production in the mines is approximately 1 million t. Because of varying dimensions of the ore bodies, both transverse and longitudinal stopes are employed. The inclination of an ore body is usually 1.04 to 1.59 rad (60 to 90°), and the horizontal width varies up to 100 m. Sublevel stoping has been the traditional mining methods in this mine. Until a few years ago, the ore was primarily blasted using horizontal blasthole fans drilled from Alimak raises. Recently, however there has been an increasing demand for a les expensive mining method. Alternate systems were considered, and finally a method based upon large blasthole stoping was found to be the most suitable. GENERAL LAYOUT The general layout of a stope is shown in Fig. 1 As with all the figures in this chapter, only the prin¬ciples of the method are emphasized and the details may not be valid for any particular site in the mine. In transverse stopes, the horizontal width of the ore determines the length of the stopes. The average length is approximately 50 m, the width of the stopes is 20 m, and there is a 10-m wide vertical pillar between. The distance between the main levels is 60 m. Usually a horizontal (sill) pillar 10 m thick is left below the mucking level. Each stope is first undercut along the entire stope length. This undercut is made with 50.8-mm (2-in.) diam drill holes in a fan pattern. At the moment this is being done by a trackless Atlas Copco Promec jumbo equipped with two Cop 130 pneumatic rock drills. The drill drift is 4 X 4 m, and the crater is 20 m high. This leaves a 36-m stope height to mine with large blastholes. To achieve a fair portion of vertical drilling withlarge blastholes, there are two parallel drill drifts at the top of each stope. These drifts are 4 m high and 9 m wide, separated by a 4-m wide pillar in the middle of the stope. To ease the drilling of perimeter holes, the drifts extend 1 m into the vertical pillar on each side of the stope. The pillar will thus be only 8 m wide on the drill drift level, but 10 m wide down the rest of the stope.

Jan 1, 1982

By Rush H. Spedden

Antoine M. Gaudin was a vigorously creative man and throughout his career an internationally respected leader of his chosen profession of mineral engineering. To his professional colleagues and clients and to his former students he was both "Professor Gaudin," their wise, inspiring, and enthusiastic teacher, and "Tony," their warm and beloved personal friend. Thus, his teachings and his unique inquisitive-creative approach to problems live on and grow in the accomplishments of his students. Antoine Gaudin was born at the turn of the century in Smyrna, Turkey, where his father was general manager of a French-owned railroad. It was there that he spent his earliest years, literally in the shadow of classic marbles from diggings at the site of Aphrodesias, which his father had rediscovered. From this, Tony developed an appreciation of the arts. He eventually surrounded himself with art treasures; he reveled in the beauties of nature, and he cherished his choice collection of rhododendron and azaleas. His artistic traits led him to take up painting when his daughter, Elinor, sought his help in blending her tints for a water color. In typical fashion he was not satisfied until he had become adept with pastels in capturing the muted tones of distant landscapes and the technical properties of light and color. Unfortunately not many paintings were produced. His paternal grandfather was Marc-Antoine Gaudin, for many years secretary of the Academie des Sciences in Paris, and a well known chemist and mineralogist whose primary interests included the synthesis of precious stones. In an autobiographical sketch written in 1939, Dr. Gaudin recalled: "As a child I saw many of the original synthetic sapphires and rubies which my father had inherited. This, coupled with my father's avocation of mining ventures and archaeology, early decided that I would become connected with the mineral industry." One can speculate that the archeological activities of his father, the mineral systhesis experiments of his grandfather, and his own early studies of mineralogy and geology formed a base from which his special talents as an experimentalist were to develop. Laboratory research under Professor Gaudin's personal guidance was often an intensive, fast- moving, and exciting experience--especially to a new graduate student. First of all, Professor Gaudin was a highly skilled observer with seem-

Jan 1, 1976

By Ted D. Johnson

INTRODUCTION The report of the workshop on Alternate Energy Strategies estimated that the world's reserve of uranium oxide is about 700,000 tonnes. In order to extract these reserves, it is estimated that between 100 and 200 million tonnes of solid waste and of the order of 1 billion cubic meters of liquid waste will be generated. In addition, radium within tailing will continue to generate radon gas as it decays for many years after closure. Hence, the handling, storage and possible use of uranium mine waste is a matter which deserves serious consideration and planning. Since the properties of the waste will largely dictate the method of handling and storage or use of the waste, the properties of the waste should be determined, to the degree possible, prior to the design and subsequent operation of the mine. Uranium mine waste has both solid and fluid constituents. The solids include soil and rock removed during site preparation, similar material removed during development of the mine and uranium tailing generated by the mill. Soils can be stockpiled for reuse during reclamation or used as earth materials for construction of fills, roads, dams and soil linings. Rock can be stockpiled, used as mine backfill or can be used for earth construction in fills, roads and dams. The tailing is usually placed in a tailing storage area. The fluid constituents are both liquid and gaseous. The liquid waste is generally composed of the mill water; however, additional waste water can be generated by dewatering processes and diversion of surface runoff. The gaseous constituent of the uranium mine waste is radon generated by the radioactive decay of radium in the ore tailing. DEFINITION OF PROPERTIES The following presents a brief definition of some of the more important properties discussed later. The definitions are not intended to be comprehensive, but are useful as a basis for understanding.

Jan 1, 1980

By T. S. Golosinski

Most of the 154 t (170 st) truck fleets that were predominant in large surface mines in the 1970s and 1980s are approaching the end of their economic life. One replacement option considered by mine operators is a change to the 2 18 t (240 st) trucks. A comparison of load capacities between the 1.54 t (170 st) and the 2 18 t (240 st) trucks indicates a large productivity advantage in favor of the 218 t (240 st) trucks. This advantage, however, is often reduced by a number of site-specific factors. Therefore, a careful evaluation is required to define the actual benefit of changing to the large trucks. This article discusses some important factors that have an impact on truck size selection, based on several different studies. 218 t (240 st) trucks The 2 18 t (240 st) rear dump trucks were first introduced for surface mine service almost 10 years ago. However, it is only during the last few years that this equipment has gained acceptance in the surface mining industry. About 250 units were employed in surface mines worldwide at the beginning of 1992, and further purchase commitments are being made. Encouraged by the mining industry’s interest, all major off-highway mining truck manufacturers now offer 218 t (240 st) models (Table 1). At present, the existing 218 t (240 st) truck fleets are dominated by electric drive trucks. However, recent introduction of a 2 18 t (240 st) mechanical drive truck rounds out the development of this truck size.

Jan 1, 1992

By William M. Schafer

Mining projects which contain sulfide mineralization face increasing regulatory hurdles due to environmental concerns over acid rock drainage (ARD) and the ability of long-term reclamation to protect water quality. Among regulatory agencies, much of the perception about sulfide projects has been based on conditions at historic mining sites at which little or no effort was made to control ARD. With proper design measures, ARD production can be greatly reduced or eliminated. The science of ARD prediction and control must be expanded to enable mine planners to anticipate the effect of various facility design options on ARD potential. Some common-sense examples of geochemical prediction techniques and mitigation measures for waste rock dumps, spent ore heaps, and mine pits are described.

Jan 1, 1992

By M. L. Harris, I. Zlochower, M. J. Sapko, I. E. Perera

Experiments conducted with limestone dusts and dolomitic marble dusts have indicated that when rock dust is wetted and subsequently dried, it becomes a solid, non-dispersible cake. However, in order to be effective, rock dust must be able to disperse as individual particles to inert. To counteract this, rock dust manufacturers created treated rock dusts that will resist caking after moisture exposure. National Institute for Occupational Safety and Health (NIOSH) researchers conducted a series of laboratory-scale experiments on four base rock dusts and their treated counterparts to assess the effectiveness of various anti-caking additives after being exposed to moisture and then dried. The dusts were exposed to moisture using humidity cabinets having a high relative humidity (99% RH) and by also exposing the rock dust bed to water through bottom wicking. The dusts were then evaluated for dispersibility after drying using the NIOSH-designed dust dispersion chamber. The anticaking additives were different concentrations of stearic acid, oleic acid, and xylene-based surfactants. All results were compared to a reference rock dust used to conduct large-scale experiments in the Lake Lynn Laboratory (LLL), Fairchance, PA. When the untreated dusts were dried after exposure to moisture for one day, no dispersion was measured. However, rock dusts treated with anti-caking agents were readily dispersible even after exposure to moisture for six months. This report details the analysis and characterization of anti-caking additives using the NIOSH-designed dispersion chamber and the 20-L explosion test chamber. INTRODUCTION In order for a coal dust explosion to propagate, five elements are required. In addition to the fire triangle in which three elements are necessary to sustain a fire—fuel (coal dust), heat (ignition source), and an oxidizer (oxygen in air)—a combustible dust explosion requires dispersion of a dust cloud (pressure wave) and confinement of a dust cloud (underground mine entry) (1). Coal dust explosions can originate anywhere in a mine. For example, when a small volume of a flammable methane and air mixture is ignited in a closed end, the high-temperature gases rapidly expand to create a pressure wave, sometimes referred to as a “pioneer wave,” that may steepen into a shock wave as it propagates away from the ignition source. The shock wave produces a wind that disperses the dust preferentially from any elevated and exposed surfaces (roof, ribs, belt structure, cribbing, etc.). Factors that are known to affect the intensity of an explosion are the dust particle size, the location of the dust within the entry, the dust dispersibility, and the volatility of the coal dust (1). Particles less than 75 μm (< 200 mesh) in size are most reactive (2). Dusts that are located on the roof, ribs, and elevated surfaces are more readily dispersed than dusts on the floor. Therefore, application of quality rock dust in sufficient quantities is essential to inert coal dust and to prevent continued flame propagation. 30 CFR 75.402 and 75.403 require 80% total incombustible content be maintained within 40 ft of the face, and 30 CFR 75.2 defines rock dust acceptable for use in underground coal mines. It also defined that “when rock dust is wetted and dried, will not form a cake”.

Jan 1, 2019

By Robert H. Scott

Developments during the past few years have brought coal-based composite fuels to commercial readiness. During the next few years, composite fuels will become available from a number of suppliers. Potential users must evaluate the combustion properties of these fuels to determine equipment and operating changes required for their use. Most work to date has been done with boiler applications. However, many potential non-boiler applications exist in the metals and minerals industry. These applications require specific evaluations appropriate for the processes involved.

Jan 1, 1982

By Nora M. Hutt, Kevin A. Morin

There is currently some pessimism over mixing of net-acid-neutralizing (NAN) and net-acid-generating (NAG) rock to obtain indefinite near-neutral seepage from mined-rock piles. This is due in large part to the lack of well- documented, successful cases of blending and layering, and to the knowledge of unsuccessful attempts. Past work often defined a potentially successful mix in terms of the ratio of net-acid-neutralizing rock to net-acid-generating rock. For example, one case study indicated the bulk pile-wide ratio was approximately 3: 1. Since this exceeded a common ratio used in acid-base accounting, no acidic drainage was predicted. However, acidic drainage appeared within two years. A large part of the ambiguity in mixing arises from attempts to predict success based on a ratio. This paper looks in detail at simple discrete-zone mixing, where at least one truckload of NAG rock is placed in a particular location with no further rehandling or dozing. The lack of acidic drainage in this scenario is dependent on the intervening distance between the discrete NAG zones, along flowpaths occupied by NAN rock. Calculations show that the bulk mixing ratio of a mined-rock pile is not an accurate indication of whether near-neutral drainage can be maintained. For example, a pile with a bulk ratio around 300: 1 could still release some acidic drainage. Calculations also show that the minimum separation distance of the NAG zones (L) can be so great, even with typical NAG and NAN rock, that the mixed pile could safely contain only a few percent of NAG rock, which would not be sufficient for many minesites with NAG rock. The comparison to a more elaborate, computer-based statistical model shows that the simple discrete-zone approach will provide a reliable yes-no indication of whether any acidic drainage will leave a pile. However, predictions of whether combined downstream drainages will be acidic depend on numerous site-specific physical and geochemical factors within and below a pile that cannot be modelled in a general way, like the number of flowpaths and the nature of their intersections.

Jan 1, 2000

By S. J. Stokowski

An extensive drilling and laboratory testing program indicates that diagenetic dolomitization of the Oligocene Sombrerito Fm. limestone, southwestern Dominican Republic, is detrimental to the usefulness and quality of the potential construction aggregate resource. The first part of this program was the collection, photographing, and geologic logging of diamond drill cores to allow lithologic subdivision of the strata into potential quarry layers. The primary lithostratigraphic units are: I (youngest) approximately 10 m of white, dense marine limestone; II, approximately 5 m of white, interbedded dense and chalky marine limestone; III, 2+ m of tan, chalky calcitic dolomite. Unit I has been often altered by lateritic weathering and/or dolomitization; Unit II is often altered by dolomitization. The quarry layers were individually crushed, by core, into coarse aggregate sizes for chemical and physical (specific gravity absorption, MgSO4 soundness, sand equivalent, and LA degradation) analyses. The non-weathered units decrease in coarse-aggregate yield and quality with depth and dolomitization. Dolomitization typically causes about 5% lower yield of coarse aggregate, a sand equivalent of about 20 units below the original value, an increase in absorption of about 3%, an increase in the LA of up to 13%, and a decrease in the average 5-cycle MgSO4 soundness by about 17%.

Jan 1, 1987

By A. L. Boerner

Off-highway tire costs represent a significant portion of the total mining costs in openpit, truck-shovel operations. To extend the lives of tires, the site-specific factors influencing tire life should be investigated Valuable data can be obtained by combining historical operational data captured by a mine dispatch system with detailed tire and maintenance histories. Assuming that failure counts satisfy a Poisson distribution, F-tests can be applied to determine significant deviations from desired or target counts. The methodology developed in this paper can be used to identify the haul routes and loading and dumping zones that have significantly higher incidences of tire failure. Corrective action can then be targeted at these zones by modifying the shovel or truck operator practices; changing the haul route maintenance and clean up practices; or altering the bench, ramp or waste dump design parameters. An example is given of the correlation of tire separation and tire impact failures with haul routes for a fleet of nine Caterpillar 793C trucks in operation at the Chuquicamata Mine in northern Chile. A high incidence of impact failures was observed for the ex pit ore crusher and for tires installed in the left outer rear position on the trucks. The latter was postulated to be due to a preference for drivers to execute right-hand turns prior to dumping. Recommendations were made to revise crusher and dump cleanup practices and investigate the feasibility of super-elevating the surface of dumping zones to reduce ground pressure on the left rear tires.

Jan 1, 2002