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Drainage and erosion control often become problems at mine reclamation projects. An example of this was an abandoned surface uranium mine on the south Texas plains. Engineers with the Texas Railroad Commission solved these and other problems with a three-dimensional geomatrix erosion-control matting at the 69-hm2 (170-acre) site, about 70 km (45 miles) southeast of San Antonio. Enkamat, manufactured by Akzo Nobel Geosynthetics, was installed to reinforce the grass waterways that handle runoff from the reclaimed mine. Reclamation was completed in the spring of 1992. Bill Walker is regional manager for Erosion Control Systems. He worked with the contractor to identify an erosion control product that complied with the project's design and installation parameters. "Use of erosion control matting enabled us to reduce erosion control costs," said Mark Rhodes, with the Texas Railroad Commission. "Also, the erosion control material was more aesthetically compatible with the reclaimed landscape." The commission's design called for a terrace ditch to be constructed midway up the reclaimed spoil slope to catch and collect water. Commission engineers selected spoil slopes, ranging from 5% to 8%, to build reinforced grassed waterways to drain the collected water from the terrace ditches. This was done instead of dropping the water at steeper slope location in rip rap- or concrete-lined channels. Enkamat's three-dimensional geomatrix of heavy nylon monofilaments is 95% open space that aids root growth. Its root-reinforcement system resists the hydraulic lift-and-shear forces created by high-volume discharges of water. In addition, fine eddy turbulences are prevented from isolating and uprooting individual roots and clumps of vegetation. Enkamat is flexible and lightweight. This makes it easy to install. And, because grass grows into it over time, it "disappears" after installation. Between 12,540 and 14,200 m2 (15,000 and 17,000 sq yds) were installed in the waterway using 150-mm (6-in.) staples to fasten the matting, said contractor Dale Blackstock. "It was easy to use; no problem at all to put down." Blackstock selected Enkamat from a list of eight different channel lining materials described as acceptable in the commission's design specifications. The design specifications were based on the results of channel-lining performance tests conducted by the Texas Department of Transportation. Rhodes said that the commission's Abandoned Mine Land Reclamation program had installed Enkamat on other abandoned uranium mines and on an abandoned surface lignite mine near Austin. "Pretty much all of our installations of Enkamat have worked out well," he said. "There have been no failures. The reinforced grass waterways have performed as designed."

Jan 1, 1998

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By Paul C. Schmall, Arthur B. Corwin, Hugh S. Lacy, Derek Maishman

For reasons of hydraulic design and constructability, water supply tunnel systems are often located relatively deep, in competent rock. Where the rock lies below a thick cover of water bearing soils, artificial ground freezing offers a low risk, cost-effective method of constructing access shafts. Several examples, in Kansas City, MO and in Brooklyn and Queens, NY, illustrate the use of freezing to provide structural support and ground water exclusion. In the case histories presented, the shafts range from 6m to 12m (20 to 40 ft) in diameter and soils vary from clays to boulder beds with depths in excess of 73m (240 ft) to competent bedrock. Where tunnels for infrastructure are designed to be driven in rock, vertical access may encounter a considerable thickness of overburden soils with groundwater. In such cases, shaft sinking problems are analogous to those confronted by mining engineers over 100 years ago. At that time, newly developed refrigeration machines were set to freeze the ground water and stabilize running sands. Although modern technology has refined our efforts, the basic features of the ground freezing to be described here are the same as those used in the last century. Vertical holes drilled on a circle surrounding the shaft site are equipped with leak-free steelpipes through which a refrigerated brine is circulated. In this manner, a frozen cylinder, seated in rock, is formed when the earth freezes between adjacent pipes. Shaft excavations take place inside a self-supporting, ground-water excluding structure. For mines, no better method of sinking production shafts through deepwater-bearing ground has yet been established. Indeed, major deposits of coal, potash, and salt would have remained inaccessible to this day were it not for the continued use of artificial ground freezing. If mine shaft freezing became relatively commonplace, early applications to civil engineering were few. Mussche and Waddington (1946) mention only five projects between 1892 and 1935; three of these relate to subway or vehicular tunnels in Paris, Antwerp, and (the most extensive) at five locations of the Moscow Metropolitan Railway development. In the United States, the method was used to spectacular effect in 1936, freezing about 3,000 cubic yards of silt to stop a slide

Jan 1, 1999

By Gabriel de Almeida

INTRODUCTION The widespread use of desktop computers by an ever increasing number of people, both for personal and for business and government uses, has brought public attention to surprising stories about computer abuse. Before the advent of microcomputers in the information processing environment, hundreds of amazing misdeeds involving large computers had already been reported, most of them perpetrated by highly skilled professionals. The undreamed enlarging of microcomputer potential has steadily posed a major threat to complex mainframe computing systems, as joyful teenagers become familiar with the habit of hacking around someone else's large processing system, by means of makeshift phone communication networks, based on their intended homework for personal computers -- for the sheer fun of it. Now, the never ending capacity and speed improvement of commercially available micros has given rise to a brand new concern for society: the hazards of processing sensitive information in the powerful, but still unsafe microcomputer itself, whether or not linked to other micros or mainframe networks. This paper deals with some aspects of the problems and presents some practical hints as to enhance protection and minimize risks to sensitive and secretive information when processed on microcomputers. The physical considerations of computer security and pertinent safety precautions will not be discussed, because while essentially relevant to information protection, recommendations are well defined by processing resources suppliers and their importance is more widely recognized by users in general. Definitions and Classifications. The following definitions and classifications will be considered in this report: Information Assets: Any knowledge or information processing, storing or conveying device displayed in one's environment. It consists of two main categories: INFORMATION RESOURCES: End user's application data wherever recorded. PROCESSING RESOURCES: EDP skilled user's hardware, software, peripherals, magnetic media, communication devices, auxiliary equipment, supplies and related operational procedures. Information Protection: In a broad sense this is the set of necessary procedures to assure to the legitimate user the effective control of the operation and allow modifications of information assets, so that their physical or cognitive integrity be preserved or any unexpected loss minimized inspite of accidental occurrences or intentional actions. Classification Processes: Commonly two classification processes are used to define key terms in information protection: SENSITIVE CLASSIFICATION: Pre-defined standards for assessing the vulnerability to losses due to lack of proper processing continuity if information is destroyed, disclosed, secluded, or misused. SECRECY CLASSIFICATION: A spectrum of degrees as to the expectancy of holding exclusiveness of unique information whose destruction, disclosure, seclusion or misuse will cause a negative impact on the legitimate holder environment. Sensitive Information: Any knowledge whose destruction. disclosure. seclusion or misuse - - - implies losses to its legitimate holder. It may be classified as: VITAL INFORMATION: An essential knowledge whose lack of proper availability implies total loss or termination of p

Jan 1, 1987

By S. B. Bealko

Mining haul trucks comprise the majority of the equipment used in underground limestone mining operations and are known to emit high levels of noise. A previous study conducted by the National Institute for Occupational Safety and Health (NIOSH) indicates that 70-90 percent of all miners have a noise-induced hearing loss (NIHL) great enough to be classified as a hearing disability by retirement age. These results demonstrate the public health need to protect the hearing of workers in the mining industry, including haul truck drivers. ng of workers in the mining industry, including haul truck drivers. Cab enclosures present an opportunity to isolate the haul truck operator from both truck and other noise in the mining environment. A total of twenty-five haul truck cabs were studied which were divided into three style (treatment) categories determined by soundproofing features and technology for noise reduction: old-, new-, and retrofitted-style. This study examines the contribution of cab acoustics, operator performance, and maintenance to noise reduction for each cab style. Dosimeters were used to measure 8-hr time weighted average sound pressure levels (TWA8 SPLs) inside and outside the cabs to determine if different acoustical treatments affect cab attenuation. Adherence to the Mine Safety and Health Administration (MSHA) Permissible Exposure Limit (PEL) of 90 dB TWA8 (with a 90 dB threshold) was used as the main indicator of overall noise reduction achieved. Dosimetery results indicate only 2% of the samples exceeded the PEL, but samples could still be reduced much further. Descriptive and comparative statistics indicate that noise levels inside the new-style cabs are significantly lower than the other two cab styles. Also, data suggest that there is no difference in noise exposures when comparing the old-style to retrofitted cab styles. Operator influence (opening doors and windows) was a significant factor for increasing noise exposure. This paper demonstrates that properly designed cabs can achieve major noise reductions, but noise levels could still be reduced much further below the MSHA PEL. New-style cabs, equipped with modern noise-reduction treatments, exhibit much lower noise exposures than the other two cabs styles, and the effectiveness of the current noise-reduction treatments for retrofitted cabs is questionable. Haul truck driver observations indicate that improved noise exposure reduction training is needed. Finally, specific targets for future noise reduction research are suggested that will further contribute to the prevention of hearing loss for haul truck operators.

Jan 1, 2008

By S. Ash

Murray mine was one of the first Mechanized cut and fill underground gold operations in Northern Nevada. At the end of 2002 mine production began to decline and Murray entered a phase of remnant mining and began preparations to close down. The remaining mining took place along the ore waste boundary, in previously condemned areas, and in highly challenging ground conditions. To remain economically viable new techniques were implemented in operations, geology, engineering, and rock mechanics. These improvements pushed the edge of known and accepted mining practices and have allowed Murray to remain in operation today. History The Murray mine is located on the Jerritt Canyon property, which is located 45 miles north of Elko, NV in the Independence mountain range. Mining operations started at Jerritt Canyon in 1980 out of several open pits. In 1993 underground mining began in conjunction with a decline of surface mining operations. The transition from surface mining to underground mining moved the property into a new era. At the time there were limited examples of successful underground operations in the area from which to draw experience. Several mining methods were attempted with varying degrees of success and underhand cut and fill proved to be the most effective method. To date Murray has mined 1.2 million ounces at an average grade of .335 OPT utilizing this mining method.

Jan 1, 2006

By John W. Rozelle, David S. Bolin

INTRODUCTION During the normal course of a mining project, a feasibility study will be prepared that addresses all the aspects of the project including geology and mineral inventories, mining and geo- technical, metallurgical processing, infrastructure, environmental considerations and economic analyses. These studies are prepared for various reasons. Occasionally, a feasibility study will be prepared solely for use within a company as justification for development expenditures, however, these documents are frequently submitted outside the company either to raise funds from financial institutions or from public financing on one of the various securities exchanges. Mineral industry consultants are often called upon to participate in the preparation of feasibility studies or to conduct an independent audit and review of an internally prepared study. As the recoverable mineral resources form the foundation upon which the balance of a feasibility study is based, including the economic analysis of the project, the resource aspect of the study is examined very carefully by financial and securities organizations. The balance of this discussion is directed toward the various geological and resource estimation aspects of feasibility studies that experience has shown to be of particular importance. Much of this discussion will be directed toward open pit gold operations, although the principles are applicable to any type of deposit. 2.0 GEOLOGIC RESPONSIBILITIES IN FEASIBILITY STUDIES The principal geological effort in the preparation of a feasibility study is generally dedicated to the following areas: - Preparation of Geologic Reserve/Resource Estimates - Condemnation of Sites for Facilities, Access Roads, Leach Pads, Waste Dumps and Tailings Disposal - Collection of Geotechnical Data - Provision of Mineralogical and Petrological Data Pertinent to Metallurgical Processing - Selection of Appropriate Bulk Sampling Sites for Collection of Metallurgical Samples Each of these items will be addressed in detail in the following sections. 3.0 RESERVE/RESOURCE ESTIMATION An ore reserve estimate consists of the determination of a volume, or tonnage, and a grade for the deposit. There are numerous cases in the mining industry of projects that have experienced severe difficulties when operational ore grade, tonnage or metallurgical amenability have proven to be deficient compared to predicted estimates. An error in estimation of tonnage generally leads to more long term problems affecting mine life and production schedules. However, a significant error in the estimation of grade, particularly if mined grade is lower than anticipated, can have an immediate, and often disastrous, effect upon a project. The geologic preparation leading to an ore reserve estimate generally follows a logical sequence familiar to

Jan 1, 1987

By Roderick K. Davey

The objective is to develop a short-range mining plan which satisfies management and operating criteria and also "dovetails" with the long-range mining plan. In the past years, this objective was accomplished entirely by manual methods. Today, the computer is used to assist the engineer with this task. The results provided by these programs usually are incomplete and require further manual manipulation. The advantage of using the computer programs is that the program can be used to evaluate alternative plans on a first-cut basis minimizing the manual work. Subsequently, the most promising alternatives can be analyzed in more detail, commencing with the first-cut results. In this specific case history, one technique for using computer programs to assist with short-range mine planning will be discussed. Some of the basic operating and management criteria considered directly and indirectly are as follows: (1) operating criteria which include minimum and maximum shovel cut size, equipment performance, haulage roads, requirements for blending, and plant capacities; and (2) management criteria which consist of production requirements, cash flow requirements, return on investment, market demands, and availability of capital.

Jan 1, 1979

By A. C. T. Bigg

Heavy media separation is one of a number of ore-concentration techniques for coarse? feed which, when the ore is amenable, offers economic advantages in the mining stage and/or the processing stage. In mining, simpler and cheaper methods may be used if a greater proportion of waste material can be included in the mining blocks that can later be re-moved efficiently at a coarse size. In the mill, pre-concentration can mean considerable savings in capital and operating costs by down-scaling of equipment sizes particularly in the grinding section. Conversely, throughput may be increased using the equipment already in-stalled in the concentrator.

Jan 1, 1988

On my arrival in New York after the Russian examinations John D. Ryan, president of Anaconda, asked me what my plans were. I told him I wanted to be useful somehow in the World War that we had entered while I was away, but I did not know how. Ryan and Thayer told me that the new American Air Corps was gathering a group of executives, most of them young, to go to France as Army officers to help with the industrial side of our aviation activities. This would include construction of military airports, maintenance, and supply for our rapidly growing Air Corps. They thought I might be accepted in this group of about 30. It seemed to be just what I wanted, and I put in my application at once. My application was accepted about three weeks later and I was told to come to Washington to take physical and other tests and to meet others of the group. I liked them. They were a cross section of competent American manufacturers, bankers, automobile executives, and merchants. I was the only miner. I had many discussions in the days in Washington with members of our organization and others I knew who were going into war service. Attitudes were very different from those 55 years later. There was no discussion as to whether war was right or wrong. We were in it and must do all we could to help. At the same time there was a holdover from the days when, at least for the elite, war was high adventure. Even the most high-minded did not then realize that war is a horrible form of wholesale murder and, in 1917, it would have been treason to suggest this. I had two weeks before our industrial group was to start work which gave me time for a hurried trip to Berkeley where Dorothy and I announced our engagement.

Jan 1, 1976

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 Guillermo P. Salas

The Consejo de Recursos Naturales No Renovables of the Mexican government entered into an agreement with the United Nations in 1964 for the mineral exploration of a large area in Estado de Sonora, Mexico. The agreement called for a systematic exploration for copper in northern Sonora based on photogeologic mapping, aerial reconnaissance, and geochemical stream sediment sampling. The UN-CRNNR investigations resulted in the detection of anomalous copper values in sediment from the Arroyo Cruz de Cafiada. The anomaly was followed upstream to its source, which ultimately proved to be the La Caridad ore body (see Fig. 1). Soil copper and molybdenum values were found over a large area, and further detailed geologic and geophysical investigations were instituted of the anomalous area. David Coolbaugh was largely responsible for the follow-up work. An induced polarization (IP) survey found a roughly circular region of "highs" that coincided with a high soil copper anomaly. The circular anomalies ringed the topographic feature that ultimately was found to contain the Caridad ore body. Soil geochemical copper response over the ore body was subdued, because the soils were derived from a leached capping, whereas leaching in fringe areas that contained less veining was less intense. The ore body was also ringed by the typical high pyrite zone, and this was the anomaly detected by the IP survey. The UN began drilling in the fall of 1966. Eleven holes tested the IP-Cu geochemical high, but none were of commercial interest. The project was returned to the Consejo and was reviewed in the field by G.P. Salas and V.F. Hollister. It was found that the leached capping had not been tested by the past activities, and the Consejo elected to explore that area in 1967. The capping was evaluated by Salas and Hollister and drilling by the Consejo began in 1967. Commercial grade ore intercepts were found in the first hole. The Consejo drilled nine holes, of which eight were commercial. At least 90.7 Mt (100 million tons) of 0.8% copper were indicated by the Consejo in eight holes. The lesson learned in this discovery is that evaluation of leached capping is an important guide to exploration. The general-

Jan 1, 1991

By Stephen G. Hardcastle

CANMET has developed a simple yet very effective tracer gas method to test the efficiency of stench gas emergency warning systems which can easily be performed by mine personnel. This test method uses sulphur hexafluoride (SF,) as a surrogate for such stench gases as ethyl mercaptan. The test has many advantages: it is non-disruptive to production because SF, is odourless, colourless; it is safe to use as it is inert and only a small volume, insufficient to reduce oxygen levels, is necessary to dope the air with a measurable concentration; and most importantly it provides definitive results to assess the adequacy of the warning system. Specific findings of these tests have included, over doping sections of the mine and subsequent production loss due to sickness, the warning being ignored, the injection period being too short, excessive dispersion, and inadequate notification times. The simple test protocol involves releasing a known volume of the SF, under similar circumstances to those of the stench gas, and then sampling for its presence downstream. These samples are collected in air-tight syringes which are then analyzed in a laboratory with a gas chromatograph. By direct proportion, between the size of the stench and tracer gas releases, it is possible to convert the SF, concentration into equivalent stench gas concentrations profiles. From these, it is possible to define, the initial arrival time of the gas, the time it takes to reach a "target" notification concentration, whether the gas peak will exceed TWAEV, STEV and CEV exposure limits, and the duration of the peak above these levels. Since 1993, this test protocol has been employed a total of eight times on four different Canadian mine properties and the results are being used to optimize the stench gas release system with respect to volume of gas used, duration of injection, and number of injection locations.

Jan 1, 1997

By J. A. Herbst, J. E. Bohrer, Y. C. Lo

A systematic approach has been developed to analyze a phosphate grinding circuit with a view toward determining promising alternatives for increasing mill capacity. This analysis was carried out and capacity increases were predicted in the context of detailed population balance models developed for the plant ball mills and classifier systems. Model development and verification was performed using data obtained in a series of batch tests and two plant test campaigns. The effects of feed rate, percent solids, ball size, and mill material filling, etc., on grinding were investigated in the laboratory and plant. Residence time distribution tests were also performed to determine the transport behavior in the mill. The population balance models were found to provide accurate predictions of full-scale grinding circuit performance. Based on the findings of this study, additional capacity in excess of 25% was predicted for these ball milling circuits by a proper combination of ball size change, mill dilution, and adopting screen classification in place of hydro-cyclones. These predictions have been confirmed by plant tests in which the recommended operating changes have been implemented. Understanding the effects of operating variables on breakage kinetics and the interplay between kinetics, transport, and classification has contributed significantly to increasing the mill capacity.

Jan 1, 1988

By H. Keith Henson

The purpose of this paper is to describe the possible application of two relatively new technologies: flat load cells, and improvement seeking computer programs (referred to in the literature as heuristic programs); to a very old problem, inventors control of ball mills. The combination of these two technologies seems most applicable to grate discharge mills operated closed circuit. It is possible that other applications such as controlling the inventory of autogeneous mills might also be found. Direct control of ball mill inventory has been the goal of a number of investigators over many years. Many patents and articles on the subject have claimed success, but operating installations are hard to locate. An excellent review of the advantages and problems of direct inventory control is contained in an article by Oswald and Ziegler (1). Figure 1 is a conventional representation of throughput vs. inventory with other factors such as slurry consistency and rock twee held constant. There are common sense, physical arguments? and experimental evidence supporting the existence of a unimodel curve of this twee? though direct measurement of the shape is difficult and how it moves as rock type changes is riot well known. Figure 2 represents the problem faced by all mass measurements of inventory, a tare weight of about 25 time the inventory mass. Worse yet, the tare is variable due to ball and liner wear and the Periodic addition of balls. The problem then breaks down into two parts: 1) obtaining enough mass

Jan 1, 1977

By Priscilla P. Nelson, Francis L. C. Fong

Introduction Fracture mechanics is the study of the behavior of cracks and their modes of propagation. During cracking, energy is consumed in creating new surface area. The more energy a material absorbs, the higher is its fracture resistance. This fracture energy is provided either by work done externally, by the release of elastic strain energy within the cracked body, or by a combination of these two effects. Fracture resistance is measured in terms of either the stress intensity factor, K, or the energy release rate (also referred to as crack driving force), G. For linearly elastic materials, these two quantities are related by the following equation: [ ] where [u] is Poisson's ratio and E is the modulus of elasticity. Both K and G are direct functions of the applied force and the crack length. The process of rock cutting with full-face tunnel boring machines (TBMs) involves the indentation of a rock surface by rolling disc tools. Excavation occurs as fractures, extending from the locus of chip indentation, propagate between concentric disc grooves and liberate rock chips from the intact rock face. Since disc cutting is achieved by a fracture process, it is logical to anticipate that differences in cuttability of a variety of rocks should be related to the variation of fracture material properties among the rocks. Such a relation has been demonstrated between the performance of full-face TBMs and the critical crack driving force for excavation in relatively high strength, brittle, low porosity, isotropic rock (Nelson et al., 1985). The objective of this paper is to report on the results of a continued investigation of the relationship between fracture material properties and the effectiveness of disc cutting. Additional reliable data on full-face TBM performance is difficult to obtain, so that the results of disc linear cutter tests performed by the Transport and Road Research Laboratory (TRRL) (U.K.) are utilized in this study. Fracture Material Property Evaluation If a crack tip is subjected to a perfectly plain strain condition, many materials will behave in a brittle manner, and the material resistance to fracture is constant during crack growth. This resistance is referred to as the plain strain fracture toughness, believed to be a material property. The standard testing procedure for this parameter is given in the American Society for Testing and Materials (ASTM) test designation E399. To ensure plain strain conditions, acceptable specimen geometries are specified. Unless these geometric specifications are adhered to, the measured fracture resistance is regarded as the apparent fracture toughness, and a subscript Q is used, as [ ]. The methodology adopted by ASTM E399 is within the framework of Linear Elastic Fracture Mechanics (LEFM). This approach assumes that material resistance against cracking is developed instantaneously, without initial stable or subcritical crack extension. Prior to crack advancement in rock, however, much of the crack driving energy is absorbed in a localized volume in front of the crack tip, referred to as the plasticity or microcracking zone. This capability for energy absorption substantially increases material resistance, and a crack may be arrested unless an increased crack driving force is applied. The LEFM or plain strain assumptions would underestimate the actual fracture resistance of such a material. Such a conservative approximation may be desirable in the design of structures against fatigue or brittle failure. In mining or tunneling, however, an underestimation of crack resistance can lead to the selection of underrated excavation equipment and a decrease in production efficiency. The development of fracture resistance in rock is best described by the material resistance curve, or R-curve, which is a plot of fracture resistance as a function of crack extension. Typically, peak resistance is only reached after a finite amount of subcritical crack growth. The

Jan 1, 1986

By C. Ptacek, D. Hiller, J. L. Janbor, H. Mccreadie

Between 1991 and 1994, metallurgical residues from a pressure-oxidation autoclave at the Campbell Au-quartz mine at Red Lake, northern Ontario, were neutralized with dolomite-rich flotation tailings and were deposited in a polyethylene-lined disposal area. Mineralogical analysis indicates that most of the As occurs as a crystalline ferric sulfarsenate, and some is associated with Fe oxides and with K-H3O jarosite. Field pore-water geochemical data from 1994 and 1995 showed increases in Fe(II) and As(III) concentrations, suggesting that the sulfarsenate and jarosite are dissolving; none of the discrete As phases included in a thermodynamic database limit dissolved-As concentrations, and these concentrations may continue to increase. These results demonstrate the potential for Fe and As loading through groundwater systems.

Jan 1, 1998

By Henry Brehaut

The International Network for Acid Prevention (INAP) was formed in 1998 by 17 international mining companies to promote significant improvements in the management of sulphidic mine materials and the reduction of liability associated with acid drainage through knowledge sharing and research and development of technology. Since its inception two panels have been formed to identify and move forward with research projects and eight networks have been constituted to initiate interaction and knowledge sharing within the membership in specific areas of technical interest. An INAP web site has been developed to facilitate the sharing of information and to encourage the exchange and development of new ideas.

Jan 1, 2000

By Robert C. Peddicord

Application of Quality Management philosophies and techniques commonly referred to as Managing for Continuous Improvement, Team Management , or Managing for Consistency is required in today's mining, manufacturing, and service industries in order to meet quality demands, beat competition, lower costs, optimize operations, effectively use personnel, improve morale, and enhance profitability. Most of these phiosophies and techniques are known to today's managers but either have fallen into disuse or were never allowed to surface in the corporate environment of the organization. Managers should review and adopt the philosophies and techniques of Quality Management and use them to revitalize their organizations and bring about a "paradigm" shift in the way we manage, thereby achieving the goals listed above.

Jan 1, 1991

By J. F. McDonald

On December 8th of last year, the new St. Clair River tunnel was holed-through to Port Huron, Michigan. The tunnel, commissioned by the Canadian National Railway, was needed to replace an existing tunnel that had been built between 1889 and 1890, and had remained in continuous operation until the new tunnel was opened to traffic in April of this year. The original tunnel, after more than 100 years of service, was still structurally sound, but was too small to accommodate the double stacked container cars and other advances in the railroad's equipment that had developed over the years. Prior to the construction of the first tunnel, all freight was barged across the river. It was a slow, expensive process, which in winter was sometimes impossible. The nineteenth century tunnel was built in twenty months, utilizing two shields that mined to the middle from each end. The lining, which is still in excellent condition, consisted of bolted, cast iron segments, with an O.D. of 6.4 meters. It was the first sub-aqueous tunnel built in North American, and it's use of compressed air under the river and hydraulic power on the shields may also have been firsts. The digging was done by hand, and the muck hauled out by mule drawn wagons. Boulders and methane gas were reported to be a constant problem during the drive, and compressed air sickness was common, but when the two shields came together in August of 1890, it was reported that the alignment was within 6 mm. The new St. Clair River Tunnel was constructed for basically the same reason as the original. The larger equipment which the railroad is using today would not fit through the old tunnel, and once more they were forced to ferry the cars across the river.

Jan 1, 1995