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By Greg Blerie

The requirements placed on belt conveyors grow increasingly complicated No longer is it sufficient for the conveyors to merely move material from "Point A" to "Point B. " Now these systems must be able to produce a specified flow, guaranteeing a plant can operate as scheduled without peaks and valleys in material flow and production rate. In addition, the conveying system must provide effective control, to minimize the problems and costs associated with the escape of fugitive material these problems include unscheduled shutdowns, premature equipment failures, regulatory headaches, poor neighbor relations, and health and safety issues. The key to accomplishing these goals is the loading and discharge of the conveyor belts. Fortunately, new designs for the chutes that accomplish conveyor loading and discharge can move material efficiently, without blockages and while minimizing the amount of dust generated This innovation is the development of engineered flow chutes

Jan 1, 2004

By Katsunori Okaya, Toshio Inoue

Application of the discrete element method incorporating the simple kinetic model of the ball-to-ball and ball-to-wall interactions enables us to visualize the details of the grinding actions prevailing in the ball mills. We have developed a computer program to represent a generic mill model with which we can simulate tumbling mills, vibration mills, centrifugal mills and planetary mills. While the program can be used for various purposes including design of new equipment, one obvious application of it is the analysis of the conventional tumbling ball mills for better understanding and improvement. The purpose of this paper is to present a model and to observe its behavior in particular attention to the spatial distribution of the energy dissipation as well as the energy spectrum and their dependence on the major parameters such as mill speed, ball filling, liner configuration, slurry viscosity and the hardness of the ore.

Jan 1, 1995

By J. A. Pagliero, R. T. Quiroz, S. H. Castro

Selected voltammograms in tetraborate buffer solutions in the presence of ethyl xanthate or isopropyl xanthate at pH 9.3, for chalcocite, covellite, chalcopyrite, and enargite mineral electrodes, are presented. The concept of critical Eh of dixanthogen condensation (Ehc) is introduced, and the line of Ehc is thermodynamically defined in plots of log [X i] vs Eh, for methyl, ethyl, propyl, and butyl xanthates. A reaction scheme of xanthate oxidation in four steps, involving the dixanthogen condensation step, is proposed. A first attempt to interpret mineral flotation from a more general point of view is made. The electrochemical mechanism of dixanthogen formation and condensation is considered as a particular case of a more general and relevant phenomenon, which is the decomposition of the collector giving hydrophobic colloidal forms, whose associative behavior in bulk solution and at the mineral/solution interface is governed by colloid chemistry.

Jan 1, 1989

By Peter T. Luckie, Mark S. Klima

INTRODUCTION A physical separation partitions a stream of particles into two streams - one containing a concentration of a desirable constituent; the other containing a defficiency - on the basis of a difference in physical properties. Separating ores into concentrates of increased grade and gangue minerals has been practiced for thousands of years. Physical properties of materials which have been used to achieve separations include size, relative density, shape, magnetism, surface charge, color, characteristic emission, resilience and strength. The study of physical separations has also been termed attribute tracking. The term attribute is used to describe the content of the material of interest, or, sometimes, properties not used as a basis of the separation. The content is usually given by its assay; e.g., %Cu, %Fe, etc. Attribute tracking involves material balancing. For a separation using two physical properties, say x and y, the balance is

Jan 1, 1986

By O. A. Bascur

The problem of setting the proper control objectives and/or having several objectives to be met in the implementation of grinding control strategies can be resolved. The design and implementation of a distributed control system for safe operation of a grinding circuit while maximizing profit is presented. An optimal recovery through- put target for a desired grind is accomplished by advanced control strategies at the DCS level. A mill load constraint control technique provides stable operation with a substantial increase in throughput, while the desired grind index is maintained by an inferential grind control technique without using a particle size analyzer. The integration of crushing/grinding and flotation in the overall design was the key to the success of the project. This synergistic effect was obtained by implementing a computer plant management system in the concentrator. The overall/philosophy was to keep the implementation simple while using new technology creatively.

Jan 1, 1991

By S. E. Phillipson, P. L. Tyrna

Lineament analysis, a remote sensing application, for geological structures, has been criticized, for its reliance on subjective interpretations. The Mine Safety & Health Administration's Roof Control Division performs lineament analyses with a technique that removes much subjectivity. Study areas are viewed on satellite images with software that allows examination under artificial sun azimuths and elevations. Extraneous lineaments are removed through successive steps of superposition, resulting in selection of the most robust linear features. Field investigations revealed a strong coincidence of joint and. fault zones that were parallel with projected lineaments. The Roof Control Division's successful identification o f lineaments associated with roof falls and high wall failures has prompted some mine managers to incorporate lineament analysis as a mine-planning tool. The identification of potentially hazardous ground conditions would allow mine operators to obviate safety concerns that might otherwise adversely affect the health and safety of miners.

Jan 1, 2003

By L. E. Krys, K. D. Harrison, S. M. Nordwick, C. G. Anderson

Since 1984, nitrogen-species catalyzed pressure leaching has been utilized at the Sunshine Mining and Refining Company for the recovery of precious metals from a refractory sulfide concentrate, The process operates without the use of cyanide. The plant is now the only industrial-scale commercial application of this technology. Since the inception of the process, changes have been implemented to improve efficiencies and throughputs. Two of these changes, the replacement of nitric acid by sodium nitrite and a concentrate regrind, have recently been implemented. This paper will discuss the outcomes of these recently patented plant-process changes. In particular, a comparison between the previously used method and the new leach method is made. Also discussed is the potential applicability of the process to other materials. A companion publication (Anderson, Harrison and Keys, 1996) indicates possible mechanisms responsible for the success of these process changes.

Jan 1, 1997

By David J. Kruse

Although it may be perfectly acceptable to neglect the belt's elastic behavior for many conveyor designs, this can lead to disastrous consequences for others. So how can a designer know in advance if a dynamic analysis is required? Unfortunately, in many cases it is impossible to foresee or predict dynamic problems prior to performing the actual analysis. However, a basic understanding of the principles of conveyor dynamics, as well as knowledge regarding common problem areas, can significantly help the designer make more educated decisions during the initial design stage. This paper will demonstrate how static calculations alone can incorrectly represent the true behavior of the conveyor system. It will also discuss the importance of correctly sizing and locating the take-up system to provide the most economical conveyor design. The use of capstans will also be discussed followed by two examples demonstrating their proper operation.

Jan 1, 2002

By H. M. von Seebach, M. P. Trubeljal, L. G. Austin

Six US coals and a crystalline quartz were tested in an instrumented laboratory high-pressure grinding rolls machine equipped with 100-mm-diam. rolls. The machine was choke-fed to maximize the capacity. As expected, the specific capacity decreased with increased grinding pressure for each material tested. Among the six coals, there were significant differences in the critical angle of nip, the effective feed porosity and the variation of strain at the gap relative to the grinding pressure. Although a precise correlation was not obtained, the overall effect of these factors was that, at a given grinding pressure, the specific capacity increased as the Hardgrove grindability index of the coal increased. The relationship between the specific capacity and the grinding pressure for a given coal was independent of feed size down to 150 µm. However, smaller feed sizes produced anomalous results, and the flow within the feed hopper was restricted.

Jan 1, 1996

By A. I. Kalugin, A. A. Abramov, D. E. Kozlov, D. V. Magazanik

Based on theoretical and experimental researches of phosphate and carbonate minerals flotation in the presence of soluable glass at various pH values and sodium silicate concentrations, it has been established that the presence of soluble glass in solution leads to formation of silicate compounds on the surface of phosphate and carbonate minerals causing change in their floatability for the worse. The harmful influence of soluable salts is done by tying of collector as metallic soaps, forming precipitates on the surface of minerals to be floated and causing their depression of flotation, and breaching of the froth etc. Depressing action of slimes and products of interaction with collector depends on the ratio of values and signs of potentials of the mineral surfaces. The possibility of phosphate and gangue minerals separation can be evaluated by comparing the minimum-necessary concentrations of collector required for their flotation.

Jan 1, 2002

By J. D. Miller, Jon J. Kellar

Researchers in flotation chemistry have long sought to describe collector adsorption phenomena in greater detail including kinetics, orientation and interactions between adsorbed species. Now a quantitative description can be made in-situ using FTIR internal reflection spectroscopy. Details of the adsorption process at the surface of mineral crystals, machined into internal reflection elements, can be described from first principles, provided the refractive indices, incident angle, and wavelength of light are known. Collector orientation can be established based on spectral characteristics of plane polarized and perpendicular polarized light. Examples of the use of in-situ FTIR internal reflection spectroscopy are given for the sapphire-sodium dodecyl sulfate system and for the fluorite-sodium oleate system. It is expected that quantitative in-situ analysis of collector adsorption reactions by FTIR internal reflection spectro-scopy will continue to grow in importance and will provide an important new dimension in our understanding of flotation chemistry.

Jan 1, 1989

By F. N. Kissell

The Bureau of Mines used sulfur hexafluoride (SF6) tracer gas to evaluate the effectiveness of gob ventilation and/or sealing practices at three coal mines, each having different problems associated with their mined-out areas. The purpose of these ventilation studies was to better understand whether current techniques employed for ventilation or sealing are successful at minimizing the potential for gob fires and explosions. The work performed at each mine is discussed: One is a longwall operation that uses ventilation to carry off hazardous gases, and two are room-and-pillar operations, each with far different concerns, that seal off the gobs to isolate them from the main mine ventilation. In all cases, employment of the Bureau's SF6 tracer gas technique resulted in answers to questions raised by the mine operators concerning the effectiveness of their gob ventilation or sealing practices in preventing mine fires.

Jan 1, 1987

By Raymond E. Krauss

The McLaughlin Mine is a world class epithermal gold deposit located in the coastal thrust fault of Northern California, U.S.A. Homestake adopted a proactive strategy in the permitting of the Mine. Extensive environmental baseline studies were undertaken. Engineering and process alternatives were evaluated for environmental impact along with other factors. A unique pressure oxidation process was selected. Screening studies were conducted as a basis for the selection of waste rock and tailings disposal sites. Project engineering and permitting were fast tracked. Construction required eighteen months, and the first gold bar was poured in March of 1985, six years after the initial discovery. Project facilities are designed and operated so as to minimize environmental impacts. A comprehensive monitoring program demonstrates permit compliance. Reclamation of the site is ongoing throughout the Mine's twenty year life. Nonmined lands are managed to maximize wildlife habitat value.

Jan 1, 1990

By Andy Ingriselli, Andrew Famiglietti, John Westland, Greg Landry

Atlas Dewatering was contracted to install and maintain a groundwater control system for the new Bessarion Subway Station and the Tail Track Structure on the Toronto Transit Commission’s new Sheppard Subway Line. Atlas met the challenges presented by the project by closely reviewing the contract documentation provided by the Program Geotechnical Consultant, Golder Associates and by working with the general contractors, Bondfield Construction Company and PCL Constructors Eastern, inc., throughout the duration of both projects. In addition to concerns about soil loss, excessive settlement and seepage, the consultant was concerned about spreading hydrocarbon-contaminated groundwater between the two aquifers present at the Bessarion Station site. By installing a system of multi-level vacuum eductor wells around the perimeter of the excavation, Atlas was not only able to provide a dry excavation for construction, but, due to a unique well design was also able to prevent cross-contamination of the aquifers.

Jan 1, 2001

By L. E. Allen, J. F. Abel, D. R. Morgan

The Waste Isolation Pilot Plant (WIPP) is a geologic repository for disposal of transuranic waste and is operated by the U.S. Department of Energy (DOE). The repository is located at a depth of 655 m in the Permian age salt beds of the Salado Formation, 40 km east of Carlsbad, NM. A key component in the design of the WIPP repository is the installation of concrete structures as panel seals in the intake and exhaust drifts after a panel has been filled with waste containers. The DOE has proposed several panel closure design modifications to the EPA. EPA approval of the changes is required since the details of the designs are specified in EPA’s final rule as a condition for WIPP certification. The Environmental Evaluation Group, in its role of providing technical oversight on the WIPP project, has evaluated the proposed modifications. An analysis of this evaluation is discussed.

Jan 1, 2003

By V. Galla

In the present paper, two large diameter columns 2 m tall and 1.3 m diameter filled with crushed (finer) and run off mine (coarser) heap leach materials respectively were used to perform unsaturated flow studies. Instruments measuring moisture content and matric suction were placed in the columns at different depths as water was applied at calibrated flow rates from the top of the columns. The outflows from the columns were recorded using tipping bucket rain gauge placed at the bottom of the columns. Unsaturated hydraulic parameters were also obtained for the heap leach materials, scalped (passing 4.75 mm sieve size) and whole (passing 37.5 mm sieve size) sample. The measurements obtained from the columns were later compared with unsaturated flow modeling results. The comparison indicated sensitivity towards the sample size selection for unsaturated hydraulic parameters, and initial conditions considered in the model.

Jan 1, 2009

By Larry D. Hartzog

For detailed information you are referred to "Heap Leach Solution Application at Rochester" by S. N. Dixon and A. L. Wilder which will be presented at the SME Meeting in Phoenix, January 1988.

Jan 1, 1998

By Kenneth P. Katen, Robert H. Trent

GENERAL DESCRIPTION Shortwall mining is one of the newest mining meth¬ods to be introduced into the world coal industry. The technique of shortwalling seems to have a number of origins, but it is generally agreed that it is an outgrowth of pillar-and-stall mining where timbering was employed as the primary means of roof support. In early opera¬tions the face was undercut by a shortwall cutter, hand¬bored, and blasted. The coal was loaded by hand or gathering arm loader. Australia, having experienced success with this method, advanced the technique through the use of mechanized roof supports in 1968. Although there were several attempts at shortwalling in the US, it was not until 1973 that Eastern Associated Coal Corp.'s Federal No. 1 mine successfully extracted a shortwall panel. Table 1 is a list of shortwall opera¬tions in the US as of April 1976. Shortwall mining is a method of mining in which a continuous miner cuts and loads from the short end of a rectangular pillar of coal while hydraulically powered self-advancing roof supports provide protective cover. Characteristics of a shortwall system include panel di¬mensions (Fig. 1) of 610 to 1219 m (2000 to 4000 ft) in length, face widths of 30 to 61 m (100 to 200 ft), straight-line ventilation, and production machinery, with the exception of roof supports, commonly used on con¬tinuous miner units. GENERAL REQUIREMENTS AND LIMITATIONS The justification for shortwalling is primarily eco¬nomic, i.e., increased production and recovery concur¬rent with savings in material and labor. The method also has a number of secondary benefits including the much needed progress it has stimulated in deep mine technology, the alternatives it presents to some of the problems resulting from the 1969 Coal Mine Health and Safety Act, and is perhaps a key to the solution of the enigmatic labor problems which plague the industry. The advantages of shortwall mining over continuous miner units include the following: a highly effective method of ventilation and dust control; a greater per¬centage of extraction (approximately 85%) ; fewer equipment moves, hence more efficient operating time; cost savings on materials such as roof bolts, rock dust, ventilation curtains, timbering materials, and the labor needed to supply, load, and transport these materials; a cyclical pattern of mining, each phase of which, i.e., development and retreat, lasts for approximately three months and appears to be psychologically motivating as well as highly productive; and improved safety. In addition to these positive aspects, a shortwall sys¬tem also has advantages over a longwall system, and these are listed as follows: the capital outlay for a shortwall system is one-third to one-fourth that for a longwall (1976 prices); the shortwall system utilizes equipment that is used on a continuous miner section, with the ex¬ception of the roof support system; there is a great deal of flexibility in areas with numerous gas wells and in areas where surface subsidence cannot be tolerated; the shortwall has reduced tailgate problems, since the tail¬gate entry serves as a return airway only; and the cyclical pattern of mining insures that the mining machinery pro¬duces coal even while the roof support system is being moved to another panel. There are several parameters which govern the se¬lection and use of a shortwall system. They are all of equal importance, and hence are not listed in any par¬ticular order. These are deposit size, seam height, seam inclination, roof and floor conditions, depth of over¬burden, subsidence and chain pillar extraction, man¬power, ventilation, and materials handling. Deposit Size Reserves available for shortwalling should be suffi¬cient for at least an eight to ten-year period, since this seems to be the effective life, without rebuilding, of a hydraulic roof support under average conditions. Of

Jan 1, 1982

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By Leonard Harris

It gives me great pleasure to be here today to tell you a little about how Newmont Gold chose to use roasting at its Nevada operations for the treatment of its refractory gold ores.

Jan 1, 1992