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By D. W. Hendrickson

Cleveland-Cliffs views the mining process as ex- tending from mine planning to cargo loading including auxiliary systems such as materials handling and plant utilities. Process automation is recognized as an area that can provide significant improvement opportunities with less capital requirement than major flowsheet modifications. This article discusses the development and results of a strategy to allow and promote implementation of process-automation improvements at individual mines based on their economic priorities, taking advantage of common systems and concepts for use at multiple mines where practical. Specific process-control upgrade projects at Hibbing Taconite Co. and the Empire Mine are reviewed.

Jan 1, 1998

By Lei Zhang, P. Somasundara

The beneficiation of phosphate is complicated due to its sparingly soluble nature. The surface proper- ties of phosphate and associated minerals are determined, in addition to the ionic composition of the solution, by the dissolution of the lattice ions of the minerals and subsequent complexation, hydrolysis, adsorption, and precipitation of the dissolved species. These, in turn, are influenced by the pretreatment of the mineral. These complex interactions determine the interfacial properties of phosphate and associated minerals and thereby the efficiency of beneficiation processes such as flotation and flocculation. Under- standing of the chemistry involved in such systems is helpful to develop the optimum beneficiation schemes. It is shown that desired selectivity can be achieved by controlling the chemistry of the systems. More important, the need to conduct development tests under plant conditions is clearly demonstrated.

Jan 1, 1999

By S. Fairis Samuelson

Boling Dome. a piercement salt dome in the Texas Gulf Coast basin is (was) the single largest deposit of native sulfur in the Western world. Total sulfur production has been over 81.000,000 long tons and the mine is still in production. Data from over 60 years of operation and 8.000+ wells reveal a salt dome rich in history as well as minerals. Sulfur mineralization and production correlate well with caprock structure and morphology. The occurrence of Mississippi Valley-type deposits in fracture zones is notable in some areas and may warrant additional study. DISCOVERY AND PRODUCTION HISTORY Boling Dome is located in the south- east corner of Wharton County, Texas, 60 miles south-west of Houston. It is a piercement type dome in the Houston diapir province. It was discovered in 1922 when the county surveyor noticed sour gas seeps along the San Bernard River (Haynes. 1942). Several companies including Gulf Production Company. Humble, Texas Company, Sun and others quickly began drilling for oil and gas. Traces of native sulfur were reported in cuttings from many wells penetrating the caprock, but the first commercial quantities of sulfur were not found until Gulf Production's E.M. Chase 3 and 4 wells were drilled in 1927. Texas Gulf Sulphur, then mining sulfur at Gulf Dome in Matagorda County, secured the sulfur rights from Gulf for $3 million in cash and half of the profits after expenses for development and construction of plant and townsite were met (Haynes, 1942). The drilling results from 200 or so Wells revealed a massive deposit destined to become the largest single source of native sulfur in North America. Original reserves were estimated at just over 46 million long tons and were increased in later years. Production by the Frasch method was begun by Union Sulphur in 1928, and continued by Texas Gulf Sulphur Co. in 1929, and Duval Sulphur Co. and the Baker Williams Co. in 1935. Union produced 8,009 long tons, Duval produced 571,123 long tons, and Baker Williams produced 1,453 long tons. Through its deal with Gulf, Texasgulf acquired the largest and best mineralized portion of the dome and has produced 88,544,466 long tons through 1990, and production is planned for several more years. That amount of sulfur is equivalent to a cube measuring 1,152 feet on each side, or a sulfur highway 1 foot thick, 20 feet wide and 14,479 miles long. Peak production averaged 8,376 ltpd in 1948. A price war between North American sulfur producers and Mexican sulfur producers in 1956, followed by the recession of 1957 and 1958 (Haynes, 1959) coupled with a rapid increase in mining depth to below 700 feet began a gradual production decline that was never reversed (Figures 1 and 2).

Jan 1, 1992

By Dragomir D. Kocic

At its lowest hierarchical level, the mine ventilation control system consists of local systems controlling the concentration of noxious gases in the air currents of various branches of the ventilation network, and of local systems controlling the air flow rates through those branches in which it is not necessary to regulate the concentration of any noxious gases. With respect to the noxious gas concentrations, the manipulated variables are air flow rates, while with respect to the air flow rates - the manipulated variables are aerodynamic resistances of the corresponding branches. Since there is an interdependence between the air flow rates in the ventilation network branches controlled through changes in aerodynamic resistance, the total system of mine ventilation control falls in the class of multivariable systems. The treatment of individual branches of a controlled ventilation network as separate plants presupposes an absolute autonomy of local systems, which is an untenable concept. The present paper is an attempt to break up the multivariable control system of the noxious gas concentrations and the air flow rates into autonomous (mutually unrelated) local control systems. The plant discussed is a ventilation network of arbitrary configuration with a single main axial fan. Because of the complexity of the problems involved, the multivariable control system of noxious gas concentrations and air flow rates has been described as linear, with constant and lumped parameters.

Jan 1, 1980

By R. Larry Davis

Over the past decade there has been a continual decrease in domestic oil reserves accompanied by a sharp increase in the demand for energy. Even with increased imports of foreign oil, the energy demand has not been met. This has forced many users to turn to substitute fuels. Of the many substitute fuels available most users are turning to and creating an increased demand for low sulfur coal. Because of the increased demand for low sulfur coal as well as the inherent problems in producing eastern coal, many producers have turned their attention to the large deposits of low sulfur coal in the western United States. Western coal mining is, however, not without its problems such as high moisture content, high ash content, highly variable sulfur content, multiple seams, and/or pitching seams. It is to aid in the solution of these problems that this chapter is directed.

Jan 1, 1979

By J. R. Mattern

The engineering geology of the Rio Tinto Borax Boron Mine creates challenges in providing practical slope design recommendations to mine planning engineers. Low strength sedimentary units interbedded with very low strength lakebed clays surround the orebody. Faulting has offset and juxtaposed these units into variably adverse orientations, while groundwater is compartmentalized both vertically and horizontally within fault blocks. A collective program of geologic data collection, geomechanical strength testing, geologic modeling, slope depressurization, confirmation of underground stope locations, and documentation of slope behavior provides vital information for design investigations and sensitivities. Analysis results lead to risk assessments and operational procedures that add value to the business by optimizing slope configurations and minimizing risk.

Jan 1, 2004

By L. A. Puchkov

The paper presents the outline of the results in the algorithms and software of the microprocessor-based system for the powerful drugline bucket automatic motion, as well as the technological process control and functional diagnostics of the electrical equipment. The system is, aimed at the increasing both the draglines' working efficiency and the professional, competence of the operators. The excavator-dragline, equipped by the automatized operation, diagnostic and information processing system, possesses considerable technological: flexibility, which can widen therefore its field of application. The tasks of the dragline bucket motion optimal operation, programming of the excavation and overburden dumping processes, as well as the tasks of the control and the analysis of the technological parameters and the diagnostics of electrical equipment conditions, are defined in the paper. In our paper we present the results of these tasks being solved, providing the algorithms of the main mechanisms for the automatic operation for the swinging bucket controlled motion, the algorithms of the statistical analysis of the working efficiency parameters the algorithms of the diagnostics for electrical driving systems. The algorithms of the motion operation are constructed using the predicting dynamic models. The diagnostic algorithms at the normal working conditions form a number of algebraic invariants always equal to Ø with the accuracy up to when the electrical equipment is parametrically adequate. The algorithms of evaluating and analysis of the excavation efficiency make use of the direct measurement and analytical identification. The working condition and quality of the operation and control algorithms are investigated by simulation and natural testing. The results of the mathematical simulation of the bucket motion (with the 20 m3 capacity) of the excavator-dragline with the boom of 90 m in length are given in the paper.

Jan 1, 1992

By D. X. Tang, G. Wilson, Y. Luo, S. S. Peng

Knowing the location of voids/fractures in roof rock will be greatly helpful for roof support design. The drilling parameters obtained during normal roof bolt installation cycle provide a fast and economic way for determining void location in roof rock. In this research, feed pressure and rotation torque are found to be good indicators for identifying the voids, based on the results of a series of laboratory experiments. A set of computational rules has been developed to locate voids in mine roof using the measured roof drilling parameters and implemented in MRGIS, a software package developed to allow mine engineers to make use of the large amount of roof drilling parameters for predicting roof geology properties automatically. Anew feature, 2D map+3D figure display, is also added to MRGIS.

Jan 1, 2004

By John W. Wilson

Every operation connected with the working of a deep-level mine has, to a greater or lesser extent, a direct bearing on the costs of production, and hence, profitability of the mine. If adequate mine planning is not done in the early stages in the life of a mine, considerable expenditure can result from the cost of maintaining mine service tunnels, production delays in stopes caused by poor strata control, distortion of fittings in hoisting shafts, rockbursts and so on. In fact, the optimum profitability of a deep-level, hard-rock mine can only be achieved if the layout of the shaft systems, ore passes, tunnels, stoping areas and other underground excavations are designed to withstand or accommodate the rock stress conditions that will be created by mining operations. In South Africa where the deepest mines in the world are located, access to the underground workings of most deep-level gold mines is via vertical shaft systems and a network of tunnels connecting the shaft systems to the stoping areas. In current practice, there are two basic methods of extracting the gold-bearing reef,* namely, the scattered stoping method and the longwall stoping method. The choice of the stoping method is governed largely by the anticipated gold distribution and payability of the reef in the area, the geological conditions, the degree and complexity of faulting and dyke intrusions, the depth at which the gold-bearing reef is to be mined and the inclination and width of the reef. The longwall system was first introduced at the East 717 Rand Proprietary Mines Limited in 1942 in an effort to combat the rockburst hazard which appeared to increase with increase in depth. This mining method became standard practice on that mine for stoping operations 700 ft. below the surface. Since that date, several deep mines on the Central Rand, East Rand and West Wits gold mining areas have adopted this method.

Jan 1, 1977

[Costs may be classified according to their source and behavior: Scale of aggregation - Total costs - Unit costs Time period for which the costs are charged - Product costs - Period costs Controllability and relevance for decision-making and control - Controllable costs - Noncontrollable costs - Sunkcosts - Opportunity costs Their basic source - Production costs (also termed manufacturing costs) - Nonproduction costs How the costs change with changes in activity - Fixed costs - Variable costs - Semivariable costs Traceability to individual units of production - Direct costs - Indirect costs Use in planning (budgeting) and subsequent control - Standard costs - Normal costs ]

Jan 1, 1999

By J. W. Shuster

Test of trolley-diesel prime power is another step in the search to exploit the unused potential of electric wheel driven haulage trucks. The output of the present engine - a 700 horsepower diesel - is not adequate to exploit fully the 1200 horsepower potential of the two electric wheel drive motors. Power was supplied to a modified 100-ton haulage truck through a double wire trolley system. The vehicle was tested with various payloads on a 7% adverse grade haul road. More significant conclusions developed are: Power - Horsepower to propel the vehicle can he more than doubled to improve truck cycle time. Productivity - Trolley power up adverse grades can improve [producivity] rates from 18 to 36%, dependent upon haul road profile.

Jan 1, 1968

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By S. R. Bryan, J. D. Stephens, D. R. Rothbard

The distributions of gold and associated elements in mineral grains from a pyritic gold ore were studied by optical microscopy, and by SEM and SIMS imagery. Products studied were a sulfide flotation concentrate, a calcined sulfide concentrate, a cyanide leached calcine, and a cyanide leached pressure oxidation residue. Individual pyrite grains are compositionally zoned and solid solution gold occurs in some, but not all zones. Most of the gold in calcined products is present as metallic coalesced grains that are smaller than 100 angstroms. Gold lost into cyanide leached residues from the calcines tends to be in “semi-dense” grains that cannot be distinguished from gold-free grains except by SIMS analysis. The gold in cyanide leached pressure oxidation residues occurs as solid solution gold in unoxidized pyrite remnants.

Jan 1, 1990

By Roe-Hoan Yoon

Froth flotation is widely used for separating particulate materials from one another in aqueous media. The basis of this important separation process is to control the hydrophobicity of particles. However, it is difficult to predict the kinetics of flotation based on contact angle (?), which is a thermodynamic property. A solution to this difficulty would be to develop a relationship between ? and the gradient of the hydrophobic forces across the solid-liquid and bubble-liquid interfaces. Such information can be obtained using the surface force apparatus (SFA), the atomic force microscope (AFM), and the thin film balance (TFB). The present communication reviews the results of the measurements conducted using these methods, and offers a methodology for deriving a rate equation for bubble-particle interactions

Jan 1, 1999

By J. P. Wilson

The Tessenderlo-Kerley Leach process has been shown to be effective in the removal of residual copper from molybdenite concentrates. Concentrates of varying mineralogy, including those containing relatively large amounts of chalcopyrite, have been successfully leached on laboratory and pilot plant scales. The lixiviant system is ammonia and ammonium bisulfite with air / oxygen addition. An abrasive is added which greatly enhances the kinetics, and it is subsequently screened out and reused. The dissolved copper is easily recovered using existing technology. Extraction profiles, reagent consumptions, and mineralogical data are presented.

Jan 1, 1999

By Bruce H. Van Brunt, Danny L. Taylor

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

Jan 1, 1989

By James F. Albus, Vincent C. Grimshaw

High pressure fluid energy grinding, or jet milling, is a versatile method of continuously fracturing solids into fine particles to gain substantial increases in exposed surface area Applications in industrial minerals grinding include calcium carbonate, talc, barytes, mica, diatomaceous earth, bauxite, rutile, coal, kaolin, sepiolite, diamonds, etc The basic principles can be extended to provide continuous drying and delivery of discrete par¬ticles from a wet process slurry The main design and operational features of jet mills and dryers are described, relating energy consumption for both air and steam milling and hot air drying techniques Mean particle sizes in the order of 1 micron are achievable from pre-milled feeds, or maintained from similar slurries in the drying mode Aspects of feed preparation and multi-stage processing are discussed

Jan 1, 1983

By M. M. Hajdasinski

For decades, the internal rate of return (IRR) project-evaluation criterion has been under increasing criticism. Torries (1998a) defends the beleaguered criterion. He claims that, as a project-evaluation and ranking instrument, the IRR is as suitable as the net present value (NPV). Torries also maintains that using the IRR to rank projects does not require the incremental approach and introduces his own alternative ranking approach. This paper disputes Torries' pro-IRR arguments and demonstrates that the IRR is a theoretically flawed and operationally dysfunctional criterion whose true meaning has persistently been misinterpreted. In a forthcoming paper by the author it will be shown that Torries' IRR-based project-ranking approach is flawed as well.

Jan 1, 2000

By G. M. Li

It is shown that seriously preoxidized arsenopyrite can adsorb a large amount of copper ions as unoxidized one, but without, release of corresponding amount of Fe ions; seriously preoxidized pyrite cannot adsorb copper ions, but with release of a large amount of Fe ions. However, if the minerals are rapidly oxidized with potassium permanganate solution, both arsenopyrite and especially pyrite can adsorb a large amount of copper ions, but still without release of corresponding amount, of Fe ions. Cu adsorption (if there is any) on both oxidized arsenopyrite and pyrite is not based on the Cu-Fe exchange mechanism.

Jan 1, 1991

By Felicia F. Peng

Two modes of collector addition techniques are applied in the column flotation for cleaning ultra fine coal. The collector can be: 1) added directly into the coal slurry through the feed system; and 2) gasified collector and transported in the gas phase through the bubbles. Operating parameters of the column flotation were investigated through three phases of work plans for Pittsburgh No.8 seam coal. In present study, the column flotation using two modes of collector addition techniques were tested and compared. Ninety percent or higher combustible recovery can be achieved by using both methods. However, utilizing the gasified collector addition technique in column flotation has more potential than direct liquid addition technique to obtain the higher recovery at same collector dosage.

Jan 1, 1993