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By Constance F. Acton

This paper presents an overview of -- the carbon-in-pulp process for gold and silver recovery as seen in the early 1980's. Traced back to its origins in the mid-nineteenth century the process development through the present state-of-the art is described. The process technology and equipment design are described in detail for the carbon-in-pulp process. A typical process flowsheet is given with a description of appropriate design criteria. Technical advantages and disadvantages as compared to the traditional countercurrent decantation process are discussed including some illustrative comparisons of capital and operating costs. A summary of currently operating carbon-in- pulp plants throughout the world is given with an insight into future developments in this industry.

Jan 1, 1982

By J. M. Mutmansky

Procedures for the determination of the operating point that results when a fan is connected to a mine and placed in operation have been a source of confusion for decades. A variety of techniques has been proposed, often to the detriment of the practicing ventilation engineer trying to apply the techniques to the multitudinous mine/fan matching problems encountered in the engineering of mine ventilation systems. This paper proposes some different assumptions and definitions that simplify the solution of problems in determining the operating point. As a result, a set of procedures is provided for determining the operating point of a mine/fan system when the outlet areas of the mine and fan are known. To establish a major simplification in the development of procedures for the determination of the operating point for a mine/fan combination, the authors suggest that the mine pressures be redefined and then used in such a manner that the conservation of energy is ensured. This provides the basis for the recommendation that the fan operating point be determined using the fan total pressure and the system total pressure characteristics. This procedure can be applied to fans in the blower, booster, and exhaust positions equally well and to both single-fan and multiple-fan networks. Another advantage is that the method is applicable to fans with or without an evase.

Jan 1, 1984

By J. E. Jennings

The problems of slope stability in open-cast mines are examined. A criterion, the instantaneous stripping ratio, is suggested for use in the design of pit slopes and as an index of control at all stages of the exploitation of the ore body, This criterion may also be related to safety and to the maximizing of profits. Two approaches to safe design of slopes, one from the point of view of soil mechanics and the other from the point of view of rock mechanics, are described and critically compared. The effect on slope stability of the regional tectonic stress field is discussed. Explanation is given of the application of the Terzaghi effective stress concept to problems in both soil and rock environments, including possible effects when borehole deformation techniques are used to measure stresses in rock in situ.

Jan 1, 1963

By M. Kalin, H. D. Sharma

INTRODUCTION The inactive uranium mill tailings investigated in this study are located in two mining districts, Elliot Lake and Bancroft, Ontario, Canada. The sites exhibit a mixture of surface features consisting of dry areas, with or without vegetation and areas covered with water. On the edges of the water bodies, indigenous vegetation has invaded the tailings beaches [ Typha] spp. (Tourn.) L. a dominant plant on these tailings beaches, has been studied for the uptake of radium-226 and lead-210 (Kalin and Sharma, 1981b) from the tailings. It was found that most of the radium-226 remains in the roots of the plants, and that the solubility of radium-226 in control soil differs from that in tailings. The uptake of radium226 by vegetation and other biota is related to the solubility of the element in water. Factors controlling the solubility of radium-226 in uranium mill tailings are of interest in assessing the environments effects of these wastes. Rusanova (1962) found that the soluble or extractable amount of radium from the soil is inversely related to the total concentration of calcium and magnesium in the soil. Experimental work has clarified some aspects of the leachability of radium from uranium mill tailings (Levins, et al., 1978; Wiles, 1978, and others). Halvik, et al., (1967) studied the effects of pH and chemical composition of surface water on the liberation of radium from uranium mill tailings and uranium ore. He found that an increase in pH, up to a value of pH 9, decreases the amount of radium released from the tailings and the ore. A positive effect of calcium chloride was noted on the leachability of radium. Benes (1981) reviewed the physicochemical forms of radium and its migration in water. Based on experimental work, he identifies primary factors which determine leaching of radium from uranium mill tailings. The ratio of the volume of the leaching solution to the weight of the leached sample; the composition of the leached solids and the leaching solution, and finally the pH of the leaching mixture are of importance. He emphasized that a scarcity of field data exists, which would relate experimental work to the actual situation in the tailings ponds. Inactive tailings ponds in Ontario are 16 to 23 years old, and processes which are of importance in evaluating the long term effects of uranium mill tailings in the environment can be studied. The objective of this work was to investigate the leachability of radium-226 from the tailings under field conditions. MATERIALS AND METHODS Sample Collection A description of inactive tailings sites in Ontario, where the tailings and the water for this study were collected, can be found in a report by Kalin (1981). [Typha] spp. specimens were excavated from 16 different locations along with the tailings attached to the roots of the plants. A tailings sample from the surface area, around the plant, was secured before the excavation. Tailings samples were also collected from dry areas free of vegetation in the vicinity of the wetland stand. Surface samples (depth 0-5 cm) and samples at a depth of 20-25 cm were secured. Water was collected from the shallow tailings beaches around the vegetation stands. Sample Preparation The roots of the plants were washed free of tailings with distilled water. The resulting thick slurry of tailings was allowed to settle in the wash basin for five minutes and the supernatant water was decanted. The remaining saturated tailings were brought to 400 ml volume with distilled water in a beaker. The slurry was mixed with a magnetic stirrer for two 24-hour periods. After the first 24-hours, the slurry was allowed to settle for 20 minutes and the supernatant water was removed. The remaining sediments were suspended again in distilled water (400 ml total mixture) and leached for a second 24hour period. The tailings slurries from the different locations had a solid to liquid ratio, which ranged from 0.5 to 1.4 grams of dry tailings material per millilitre. The variations in the ratios are the result of differing fractions of coarse and fine tailings on the sites. In the final leachates, the solid to liquid ratios of the samples were 0.17 (±0.1) g/ml. The pH of the surface water and the leachate was determined with an I.L. Portomatic pH meter. All the tailings samples were homogenized in a mortar and brought to dryness at 75 to 85°C. Approximately 0.5 g of tailings were

Jan 1, 1981

By John M. Marek

INTRODUCTION This paper is for the mining manager, generally meaning mine superintendent, chief engineer, and others who have control over the mine operation responsible for feeding the insatiable beast called the mill or processing plant. The author is not from mine management but from the field of technical mine engineering consulting. This fact illustrates the authors gross inadequacy to comment on this subject. On the other hand, the consulting profession allows visits to many mine operations, worldwide, supplying many different commodities. The common complaints and misunderstandings that have been heard worldwide will be shared within the following few paragraphs. In order to discuss optimization, one must first define it. There certainly will be many papers in the following sessions that claim to optimize something, or other, based on a technical definition of cost, efficiency, throughput or some other quantifiable basis. DEFINE OPTIMIZATION Actual optimization of an operation is not as easily achieved as most of us in the technical world would like. Borrowing terminology from the technical world, the boundary conditions and the optimization function are usually not well defined. The problems in many cases do not lend themselves to quantification. Typical management procedures are to establish goals or strategies that set the approach or general boundary conditions to the overall plant. Three simple cases are listed below as possible corporate strategies to optimization of a mine-process plant complex. 1) Maximized discounted cash flow 2) Minimized long term cost of product 3) Maximized resource utilization This list seems simple when presented in the above context, uncluttered by constraints such as market position, capital availability, or environmental limitations. Sometimes the best approach is a prioritized combination of the above items. Typically, the Board or CEO has established these strategic priorities. They may

Jan 1, 1991

The dissolution kinetics of cobalt in iodine/iodide solutions was investigated using a rotating disc technique. Variables studied here included temperature, rotating speed, concentration of lixiviants, and pH of the solution. The overall dissolution reaction was found to be limited by mass transfer of tri-iodide through the diffusion boundary layer with Arrhenius activation energy of less than 16.2 kJ/mol (4 kcal/mol) over the temperature range, 2-42°C. The pH of the solution had little effect as long as the pH of the solution was less than 8. However, when the pH of the solution became greater than 8, the dissolution rate decreased dramatically due to passivation. The rate of dissolution was in good agreement with the dissolution model developed by the authors earlier.

Jan 1, 1994

By B. DANIEL MARJAVAARA, I. A. SOFIA LARSSON, ANNA-LENA LJUNG

The flow field and coal combustion process in a pilot-scale iron ore pelletizing kiln is simulated using a computational fluid dynamics (CFD) model. The objective of the work is to investigate how the thermal effects from the flame affect the flow field. As expected, the combustion process with the resulting temperature rise and volume expansion leads to an increase of the velocity in the kiln. Apart from that, the overall flow field looks similar regardless of whether combustion is present or not. The flow field though affects the combustion process by controlling the mixing rates of fuel and air, governing the flame propagation. This shows the importance of correctly predicting the flow field in this type of kiln, with a large amount of process gas circulating, in order to optimize the combustion process. The results also justify the use of down-scaled, geometrically similar, water models to investigate kiln aerodynamics in general and mixing properties in particular. Even if the heat release from the flame is neglected, valuable conclusions regarding the flow field can still be drawn.

Apr 20, 2021

By Irving Hauser

The determination of new ways to cut costs in then times of economic uncertainty is essential in running a successful business. Arco Petroleum Products Company a Division of Atlantic Richfield Company recognized an opportunity for a substantial cast savings in transporting calcined petroleum coke to market by simply building an economical port facility at a location that would enable them to drastically cut their normal rail haulage distance.

Jan 1, 1982

By K. N. Han

Coal is becoming increasingly important as a major energy source in the face of the world's depleting energy mainly based on oil reserves. Coal is by far the most abundant fossil fuel both in the United States and abroad. Currently, because of increased mine mechanization, the fineness of preparation plant feed is increasing. Further, in the future as the grade of coal diminishes and the sulfur content increases, upgrading preparation of coal is very important. Because the density of coal is less than that of rock materials that occur as impurities, coal is separated from such waste largely by gravity-controlled processes for particle size between 5 inches and 28 mesh. If the size of coal particles is below 28 mesh, froth flotation is a good method for upgrading fine coal. Previously, numerous investigators have tried to upgrade coal values by froth flotation(1-16) and have achieved moderate success. However, the recovery and grade of coals in the concentrates are not always satisfactory, especially when the particle size is small. In general, when the size of coal is below 100 mesh, the above methods suffer from poor grade and low recovery. The reason for this is that as size decreases, the effects of gravity diminish, and other forces take effect.

Jan 1, 1985

By D. Malhotra

The economic viability of a mining operation depends on continuous process improvements. Diagnostic testing and plant auditing are essential useful tools to improve performance and reduce operating costs. Case studies are discussed to illustrate benefits of these techniques to improve plant efficiency.

Jan 1, 2001

By M. R. Yalamanchili

Recently, noequilbrium electrokinetic measurements using laser-Doppler electrophoresis were reported by Miller et al. (1992) for alkali halides and the sign of their surface charge thus determined. These results are generally as expected from simplified lattice ion hydration theory with KCI, under normal circumstances, being the most significant exception. In this regard, lattice defects (F centers and oxygen imperfections) have been found to influence the surface charge of KCI and its electrophoretic mobility. In particular, the presence of oxygen defect states in the crystal lattice has been found to be responsible for KCI's unexpected behavior. The concentration of oxygen defect states in KCI, as characterized by UV/Vis absorption, was varied in order to study their influence on the electrokinetic behavior and subsequently on the flotation response of KCl. At low oxygen contents « 60 ppm) KCI is positively charged as expected from lattice ion hydration theory. However, typically, KCI contains substantial oxygen (>300 ppm) sufficient to impart a negative surface charge which accounts for the selective flotation of KCI from NaCl by hetero-coagulation with oppositely charged alkyl amine hydrochloride collector colloids.

Jan 1, 1994

By William J. Cavanaugh

In 1967, Lone Star Industries, Incorporated, established a program to carry out the necessary Research and Development work to bring the Snyder Process to commercial application. The purpose of this paper, is to present the detailed metallurgical results gathered in comparing the Snyder Process of Comminution to conventional grinding techniques for the milling and beneficiation of certain ores and other materials. The ore primarily dealt with in this paper is a fine-grained magnetite ore, containing approximately 23% acid-soluble iron, from the Lake St. Joseph ore body of Steep Rock Iron Mines, Ltd., (Canada). This particular example was selected for presentation in this paper because extensive comparative results with conventional methods of grinding were available. Also briefly described in the paper are the comminution of other ores and materials such as porphyry copper ore and alumina.

Jan 1, 1973

By Ehsan Moosavi, Kamyar Tolouei, Mehran Gholinejad

The literature defines rockbursts in slightly different ways based on their generation mechanisms, field phenomena, and degree of damage, but in a way that is almost consistent. Generally, a rockburst refers to seismic events that mainly happen in underground excavations, for purposes such as underground mining, road and railway tunnels, nuclear power, etc., and that are followed by the collapse and outburst of rocks. Long-term rockburst prediction can be abstracted as a classification problem mathematically, in which we can introduce deep learning to solve it. Further, using deep learning in rockburst prediction can overcome drawbacks such as subjectivity and inconsistency brought from previous traditional approaches. A rockburst data-set, was employed to evaluate the current method for predicting rockburst grade, and the good results of overall success rate were obtained. The results indicated that the proposed algorithms can speed up parameter optimization search, the proposed method is robust model and might hold a high potential to become a useful tool in rockburst prediction research.

Apr 26, 2022

Ten years ago, only 3% of Consol's underground coal production resulted from using the longwall mining method. Today, that figure has grown to 45% and has helped the Pittsburgh-based concern fuel an upswing in productivity. The company reported that productivity at its underground mines, excluding labor strike periods, increased 24% between 1978 and 1981. Last year, an additional 8% increase was achieved. Efforts such as improved safety, training, engineering, and maintenance programs as well as the closing of unproductive mines has contributed to these gains, according to Joseph M. Richards, Consol's executive vice president of operations. But, Richards adds, a great part of the credit goes to longwall mining. What is Longwall Mining? To fully appreciate the advantages of longwall, one needs an understanding of how mining is done by other methods. For instance, in both conventional and continuous mining, rooms or tunnels are dug in the coal seam. Pillars of coal are left standing to support the roof as mining moves forward in the seam. In conventional mining, a slot is cut in the base of the seam and holes for blasting are drilled. After blasting, the coal is loaded. A more efficient method - continuous mining - uses a machine that combines cutting and loading. Longwall mining, by contrast, works backwards from the end of a panel. Two parallel sets of tunnels (development panels) are cut into a seam with continuous mining machines, then connected with a set of crosscuts at the far end. Min¬ing then proceeds back toward the entrance. Longwall mining employs some very productive machines. Large drum shearers, for example, run on a track along the length of the coal face for some 150-180 m (500-600 ft). The shearers cut the coal which falls onto a conveyor system or pan that also runs the entire length of the face. The longwall crew works under a roof of heavy steel supports called shields. As the coal is removed, shearers, pans, and shields are moved forward by hydraulic rams. The roof is allowed to collapse behind the shields. In addition to increasing production, longwall mining is considered to be safer because the miners work under strong steel shields. New designs have made them more rugged and maintenance free, as well as more stable on soft mine floors. Early roof supports, called chocks, had gaps in the roof canopies. The new roof shields now cover the entire working area and are better able to prevent roof material from falling into the working area, thereby improving safety and productivity. Longwall mining also results in a cleaner product in many mines. If the first layer of rock over the coal seam is thin, it might have to be removed with the coal in a room-and-pillar operation so a

Jan 11, 1983

By R. O. Burt

The paper examines the role of gravity concentration in both an historical contest and within the modern mineral processing industry. It discusses the principles of the process on both a theoretical and practical level. The importance of ore mineralogy and its effect on the choice of the wide range of equipment available is discussed. The paper concludes with a discussion on the relative merits of gravity concentration to other forms of processing and seeks to answer the classical question - "Quo Vadis?"

Jan 1, 1982

By R. J. Roman

Several research groups have spent considerable time investigating coarse ore leaching. Coarse ore leaching includes both heap and dump leaching. The term "heap leaching" will be used in this paper to mean any type of coarse ore leaching. The results of their work are generally in agreement: the progress of leaching can be described by the shrinking core model and its associated mathematics. The shrinking core model is schematically represented in Figure 1. The particle is contacted by the leach solution on its outside surface and the solution migrates inward by filling the pores on the particle. As the solution filling the pores comes in contact with grains of ore mineral the reagent chemically reacts with the ore mineral to solubilize the contained metal. The solubilized metal diffuses outward to the particle surface and is washed away by the leach solution flowing past the particle. The leaching reagent in the particle interior is replenished by the diffusion of reagent from the surface of the particle to the interior of the particle. If the chemical reaction between the leaching reagent and the ore mineral is relatively fast compared to the diffusion rate of the reagent in the solution in the pores of the particle then leaching proceeds from the outside of the particle inward resulting in a "shrinking core" of unreacted ore surrounded by a "leached shell". The rate controlling step in the reaction can initially be a chemical reaction between the reagent and the ore mineral grain, however, once the reaction has proceeded a short period of time diffusion of the leaching reagent through the leached shell will be the slow step and hence the rate controlling step in the production of metal from an ore particle.

Jan 1, 1992

By C. Sun

The Leeville underground mine has been in operation since 2006 and currently produces 3,200 tons of ore per day with an average gold grade of 0.4 opt. The mine has a large horizontal extent of 2,500 feet by 500 feet and a vertical extent of 20 to 200 feet. High density of joints occurs within the host rock. Based on Rock Mass Rating and field observation, the Leeville ground conditions are characterized as poor to very poor. As such, heavy ground support is required for the mining operation in Leeville. Ore zones with a height in excess of 10 m (35 feet) are considered for longhole stoping. A typical stope measures 6.0 m (20 feet) wide, 30 to 55 m (100 to 180 feet, depending on the ground conditions) long and 10 to 30 m (35 to 100 feet, depending the thickness of ore body) high, yielding 10,000 to 20,000 tons when completed. Narrow ore zones, less than 10 m (35 feet), are mined using cut and fill. These drifts are 4.6 m (15 feet) wide by 4.6 m (15 feet) high. Based on the afore-mentioned poor ground condition, the establishment of a ground control monitoring system is key to operating both safely and successfully. The ground control monitoring system is composed of comprehensive instrument design and installation. Instruments include Multi-Point Borehole eXtensometers (MPBX) and Stretch Measurement to Assess Reinforcement Tension (SMART) cables. The monitoring system covers the mining zones and infrastructure openings. After about two year?s applications and measurements, useful data has been and continues to be collected to use in future stope and infrastructure structure design. The system provides the information making engineer and operation teams to do their job more confident and efficiency. Keywords: Ground control, Monitoring system, Long hole stope, Instrumentation.

Jan 1, 2009

By G. T. Austin

Every year the United States consumes millions of dollars worth of abrasive materials as blasting media. Entrained in either a gas or liquid stream or propelled by paddles or wheels, they are directed onto the surface of another material to modify that surface. The modifications to impacted surfaces include cleaning, cutting, deburring, etching, finishing, and peening. Blasting media can be either organic or mineral abrasive materials. Information and data presented are for only a portion of the mineral abrasives, they are; fused aluminum oxide, coal slag, copper slag, garnet, metallic shot and grit, nickel slag, olivine, silica sand, silicon carbide, and staurolite.

Jan 1, 1994

By Sebastian Avalos, Julian M. Ortiz, Willy Kracht

Semi-autogenous grinding mills play a critical role in the processing stage of many mining operations. They are also one of the most intensive energy consumers of the entire process. Current forecasting techniques of energy consumption base their inferences on feeding ore mineralogical features, SAG dimensions, and operational variables. Experts recognize their capabilities to provide adequate guidelines but also their lack of accuracy when real-time forecasting is desired. As an alternative, we propose the use of real-time operational variables (feed tonnage, bearing pressure, and spindle speed) to forecast the upcoming energy consumption via machine learning and deep learning techniques. Several predictive methods were studied: polynomial regression, k-nearest neighbor, support vector machine, multilayer perceptron, long short-term memory, and gated recurrent units. A step-by-step workflow on how to deal with real datasets, and how to find optimum models and final model selection is presented. In particular, recurrent neural networks achieved the best forecasting metrics in the energy consumption prediction task. The workflow has the potential of being extended to any other temporal and multivariate mineral processing datasets.

Jun 16, 2020

By H. M. Woodrooffe

In the U.S.S.R. three of the world's major-known phosphate deposits are located. These have an estimated reserve of 2,600 million short tons of contained phosphorus. The best known lies in the "Khibiny Massif', an alkalaic rock formation some 600 square miles in area with mountain peaks rising to an altitude of 3300 ft above the surrounding terrain. Location and Climate The Khibiny Massif lies north of the Arctic Circle in the Kola Penin-sula, 100 miles south southeast of Murmansk, at a latitude of 67° 30' north and a longitude of 35° 30' east. It is serviced by rail from Murmansk and Leningrad (900 rail miles south). Figure 1. The phosphate operations are centered about the two cities of Apatity and Kirovsk. The former has a population of 35,000 and is on the Leningrad¬Murmansk railway. Kirovsk, 15 miles to the east, is now a city of 85,000. The region is at the limit of the tree line with most of the ground cover typical tundra. The mean annual temperature is -4°C and, in a normal year, precipitation occurs on 240 days. Because the temperature does not exceed freezing for 300 days each year, virtually all of the precipitation is snow. The two towns are nestled between mountains and, because of the weather, extensive avalanche control systems incorporating major engineering construction have been introduced to divert snow slides away from the populated areas. General Kirovsk was first settled in 1923 and the apatite occurrences were recognized shortly after. In 1933, the first concentrator for the recovery of the mineral was completed, producing in its first year of operation 170,000 tons of concentrate. With the growth and demand for phosphate in the U.S.S.R., a second concentrator was erected at Apatity. It calve on stream in 1963. Because of the severe climate, in the early days mining was limited to 85 days per year. With the introduction of new techniques and improved mining methods, the season has been extended to 115 days.

Jan 1, 1972