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By G. A. Calboreanu

The deterioration rate of any metal or alloy is increased in a erosion-corrosion application because the passivated surface layer is removed by the solid particles contained in the pumping fluid and the material surface becomes corrosion active again. Numerous attempts were made to engineer a stainless steel with higher hardness and unaffected corrosion performance. Alloys of special interest in such an area are the duplex stainless steels particularly the CD4MCu material which has the capability of being hardened due to its ferrite-austenite microstructure. Unfortunately, the CD4MCu with higher hardness comprises different intermetallic compounds which will increase the corrosion rate. The paper is analyzing the erosion-corrosion performance of a new type CD4MCu with 32% higher hardness and superior wear performance. The corrosion performance of this new alloy is better than the corrosion performance of the common CD4MCu material which the industry could supply. The corrosion effectiveness was evaluated with the Huey test, the potentio-dynamic anodic polarization test and the weight loss test whereas the erosion resistance was established with a low stress ball-on-disk type apparatus. The Wilfley hard CD4MCu material had 39% lower weight loss and 32% smaller volume loss than the common CD4MCu alloy The hardness range of the centrifugal pump parts poured and processed with the Wilfley hard CD4MCu is 320-350 HBN versus 250-255 HBN hardness of the common CD4MCu. After approximately 18 months of experimentation and field testing it was clearly established that the life time of the Wilfley hard CD4MCu is higher than the common CD4MCu and consequently centrifugal pumps were supplied to different customers in North and South America. India and more recently Australia. Sincere thanks to all the A.R. Wlifley & Sons Inc. and Western Foundries Inc. employees who supported and help to implement this work. special thanks to George Wilfley. Dave Cobb. Bill Becker Sr. and Wayne Taylor

Jan 1, 1997

By R. Mensah-Biney

The kinetics of loading of gold from bromide solution onto an anion exchange resin, PAZ-4, were investigated in a batch stirred reactor. The initial loading rate was described mathematically by a first-order reaction rate equation. The effect of agitation speed, temperature, resin particle size, resin concentration, initial concentration of gold and bromine in solution on the rate was investigated. The model described fairly accurately the loading process in terms of the variables investigated. The rate constant for the initial loading ranged from 0.5 to 1.3 x10-3 sec-1 and the half-time was 9.3 to 56 minutes. The initial rate constant increased with increasing temperature with an activation energy of 15.7 kJ/mole (3.8 kcal/mole). The rate constant was directly proportional to the concentration of the resin in solution. The subsequent reduction reaction was also modeled by the first-order rate equation and the apparent rate constant was estimated to be 3.0 x 10-6 sec-1 with a half-time of about 44 hours.

Jan 1, 1994

By A. S. Ibrado, D. W. Fuerstenau

"Gold, silver and copper form various complexes with cyanide, of which only the singly charged dicyano form adsorbs appreciably on activated carbon. The adsorption of the dicyano complexes appears to be controlled predominantly by chemical reaction at the surface of a shrinking unreacted core. Metal uptake over time is non-linear, which eliminates control by boundary layer diffusion. The rate of adsorption is not enhanced by a reaction interruption, which suggests the absence of concentration gradients in the pores. Finally, the rate of adsorption increases with decreasing activated carbon size in a manner consistent with the shrinking core model. The results of this investigation accentuate the importance of efficient acid washing of carbon and removing carbon fines from the barren carbon stream. INTRODUCTION With the dwindling reserves of high-grade ores, most of the world’s gold is produced from low-grade ores by cyanide leaching. In the past, leaching was followed by the Merrill-Crowe process, in which gold is precipitated from clarified pregnant solutions by zinc cementation. This process is still being used for high-silver ores, where the cost of carbon stripping becomes prohibitive. For deposits with silver-to-gold ratios of about 4 or less, adsorption on activated carbon would be the usual solution purification step taken. The process simplifies solid-liquid separation and offers a viable solution to preg-robbing organic carbon in some ores. The modern industrial application of carbon adsorption started with the carbon-in-pulp (CIP) process, where carbon adsorption was used mainly as a solution purification step. In 1973, Homestake Mining Company commissioned the first large-scale CIP plant in the U.S.A.1 Later, carbon was added earlier in the circuit to control preg-robbing by adsorbing gold cyanide onto activated carbon as soon as it is dissolved. The scheme is successful because activated carbon is a stronger adsorber than the organic carbon found in preg-robbing ores. This process came to be known as the carbon-in-leach (CIL) process. Numerous research investigations have been performed on the chemistry involved in the CIP process, which at the time of its introduction was not clearly understood. Of these, research on the kinetics of adsorption probably bears the most immediate impact on industrial practice,2,3 where adsorption systems rarely reach equilibrium, if at all, and slight slowdowns in the adsorption rate can translate into significant decreases in revenue. To illustrate this, organic compounds, like flotation reagents, have been known to impair CIP circuits. However, laboratory tests showed that the organic compounds do not significantly affect the adsorption capacity of activated carbon. Rather, the harmful effect may be attributed to a lowered adsorption rate.4,5 Also well known is the faster rate at which fine activated carbon powders reach equilibrium compared to coarse activated carbon, even though the final (equilibrium) amount of material adsorbed per unit weight is the same."

Jan 1, 2015

By K. J. Shelswell

This work describes the design, analysis, and comparison of discrete simulation and static spreadsheet models developed to compare the oreflow efficiency of two conveyor-based material handling systems; one approach with a conveyor stream in series and one with parallel conveyor streams. Analysis and comparison indicated that daily rates were significantly affected by random conveyor failures in both designs. The compound effect of random failures limited the effectiveness of long belt assemblies in series, while the parallel conveyor stream design partially alleviated this constraint. The static spreadsheet model predicted a lower average daily oreflow rate for the series conveyor stream design and a higher average daily oreflow rate for the design with parallel conveyor streams when compared to the discrete simulation results. These discrepancies were the product of model sensitivity to operational factors incorporated in the system and provided the opportunity to further use the discrete simulation models for trade-off comparisons and sensitivity breakdowns of mine operating practices and material handling system efficiencies in both the series and parallel belt stream designs

Feb 23, 2014

By Jey K. Jeyapalan

A characteristic common to most tailings dam failures caused by earthquakes or excessive rain storm is that the tailings tend to liquefy and flow for substantial distances as a viscous fluid, with the potential for extensive damage to property and life. Failures of El Cobre, Chile (1965), Mochikoshi, Japan (1978), Aberfan, Wales (19661, Buffalo Creek, West Virginia (1972) are examples of such catastrophic incidents. In these four incidents more than 400 lives were lost, and the loss of property was approximately $200 million. In order to be able to assess the potential for damage in case of such a failure, it is necessary to predict the possible extent of flow slide movement. This paper discusses simple procedures for performing such analyses. Also, in order to illustrate the use of these analysis procedures in tailings dam design, two case histories are analyzed.

Jan 1, 1982

By Ronald J. Roman

As in every business venture, the primary objective of a mining operation is to make money for those who supply the venture capital. Therefore, all decisions concerning the operation of the mining operation should be made in light of the resulting effect on profit. The ultimate control is on profit, but the process is controlled, for example, by aiming for a recovery or concentrate grade knowing its effect on profit. The normal procedure used to determi.ne the profitability of a venture is to "present worth" all costs and incomes back to some arbitrary year zero. Every company has formed a management decision as to what it will require as a minimum return before it will undertake an investment. Once the economic analysis is completed, investment will be made only if the venture shows a positive present worth. If two alternatives are available, the alternative with the larger positive present worth is the one chosen. The same principles can be applied to the selection of a production schedule.

Jan 1, 1971

By R. E. Le Messurier

GENERAL DESCRIPTION The mines are located at Broken Hill in the western part of New South Wales, Australia, 930 km west¬northwest of Sydney. Mining Methods Where pillars must be left between stopes, extraction of a block of ore between two or more levels is carried out in stages: 1) Stopes are mined overhand by horizontal cut¬and-fill stoping, or by long-hole stoping. The various methods of horizontal cut-and-fill stoping used are desig¬nated as: -Loader cut-and-fill stoping with horizontal blast¬holes and diesel load-haul-dump (LHD) units. -Open stoping with horizontal blastholes and com¬pressed air or electric scrapers, or with diesel or electric LHD units. Square-set timber stoping, where ground condition precludes open stoping. The backs are sup¬ported with timber square sets. Pattern rockbolting of the backs may be used as an intermediate stage between open stoping and square-set stoping. 2) Vertical pillars are mined by undercut-and-fill stoping or by vertical crater retreat stoping. 3) Where a horizontal pillar has been formed below the level above the primary stope, this level pillar, usually 6.1 to 8.2 m in depth, may be mined by over¬hand square-set stoping or by undercut-and-fill. Long-hole stoping is used at the Zinc Corp. (ZC) and New Broken Hill Consolidated (NBHC) to mine parts of the zinc-rich A and B lodes. Drilling is carried out from drill drives and, in some cases, sills. Hole sizes used are 165 mm (61/2 in.), 110 mm (4'h in.), and 65 mm (21/2 in.) nominal diam with hole lengths being limited to 50, 30, and 20 m respectively. Ammonium nitrate-fuel oil (ANFO) and water gel explosives are used. Broken ore is loaded from drawpoints at the bottom of the stope using rubber-tired diesel-driven LHD units which deliver ore to an orepass. Annual production at Zinc Corp. and NBHC accord¬ing to mining method is given in Table 1. At Zinc Corp. mining is being carried out at depths varying between 200 and 832 m. At NBHC it ranges from 528 to 915 m below the surface. The average travel time to the stope is 30 to 45 min. GENERAL ORE BODY REQUIREMENTS Size, Shape, and Dip There are six separate lodes which carry minable ore in the mines of Zinc Corp. and New Broken Hill Consolidated. The lowest of these and the main ore bodies presently being mined at Zinc Corp. are the lead lodes, consisting of a siliceous fluoritic ore body (No. 3 lens) and overlying it a calcite rhodonitic ore body (No. 2 lens). Each possesses characteristic ratios of lead, silver, and zinc, although the grade within each lens is variable, in places being as high as 30% Pb. There are also low¬grade portions which consist chiefly of poorly mineral¬ized rhodonitic material. No. 3 lens terminates within the Zinc Corp. leases, while No. 2 lens continues into the NBHC leases where it continues to provide a significant proportion of pro¬duction from the mine. Overlying the No. 2 lens are the various zinc-rich ore bodies. In ascending sequence these are lower No. 1 lens, upper No. 1 lens, A lode, and B lode. A typical assay for the small No. 1 lens group ore bodies is 8% lead, 50 g/t silver, and 20% zinc. The `A' lode is a large low-grade ore body in which the higher grade sections assay 4% lead, 30 g/t silver, and 10% zinc; larger quantities of lower grade ma¬terial also occur. B lode, the main zinc lode, typically assays 5% lead, 30 g/t silver, and 17% zinc. B lode increases in size as it extends south, and in NBHC is a major ore body, contributing about 52% of the production. All the lodes pitch south, the average pitch being 0.52 rad (30°) in the Zinc Corp. leases and rather flatter than this in the NBHC leases (see Figs. 1-3). The ground is all competent rock. Production requirements at Zinc Corp. are 900 000 t and 1 200 000 at NBHC.

Jan 1, 1982

By S. C. Suboleski

In a high risk venture such as mining where capital is committed and contracts are signed on the minimum of information, decisions are based primarily on production forecasts derived through consideration of such factors as management-labor relations, mining practice, variations in physical conditions and the equipment applied. Among these factors, the physical conditions offer the least opportunity for control and, due to the wide range of mining conditions that may be experienced, little effort has been made to develop objective methods for measuring their effects on mine out- put. The aim of this paper is to demonstrate a method for assessing these relationships that can be applied under a wide range of conditions both in new mines and in working sections where there is a need to determine the impact of proposed operational modifications. psing empirical data from active underground mines, a mathematical model was developed to derive a set of regression equations for predicting production as a function of mining conditions

Jan 1, 1980

By Barry Hunsaker

I cannot help but be reminded, as I stand here today, of my boyhood years and heritage, both of which are inextricably tied to this very valley and are, in a: bay, apropos of the theme of this Keynote Session. I remember the scarcity of people. For example, 48, 000 in Phoenix and 436, 000 in the State in 1930 and 65,000 in Phoenix and 499, 000 in the State in 1940. Today Phoenix boasts a population of over 500, 000 and the State a population of 1, 500,000. I remember the' lack of facilities, goods and services, including gas and electric power. I recall stories of my grandfather who had the freighting or teaming contract to haul all the supplies and generating equipment, by wagon mind you, up the old Apache Trail to Roosevelt Dam. . .a three day trip from the nearest railhead which, by the way, was not Phoenix, for it did not merit a railhead on those days, but the village of Maricopa approximately 25 miles south of Phoenix. I remember chopping firewood, heating wash water on an open fire in a No. 3 wash tub at the rear of the house and doing my school homework by the light of a kerosene lamp just 25 to 35 years ago within what is now a large metropolitan area

Jan 1, 1965

By F. E. Senftle

Preliminary geophysical surveys for nickel deposits are usually carried out by airborne magnetometer and surface electrical methods. Once an anomaly is discovered by these techniques, one must resort to chemistry to identify nickel and determine its concentration. Consequently, a method to determine nickel quickly and quantitatively at the exploration site could help guide detailed sampling and provide analytical data that might otherwise be delayed at distant laboratories. In addition, borehole probes capable of indicating nickel in the rock adjacent to the hole, should prove attractive to the exploration geologist. With these needs in mind, preliminary investigations have been made to demonstrate an in situ analytical technique which appears feasible for nickel exploration. The method depends on the very high energy gamma rays emitted when nickel absorbs a thermal or low energy neutron. These neutron capture gamma rays should be distinguished from the lower energy gamma rays emitted during decay of radioactive nuclides as in conventional activation analysis.

Jan 1, 1971

By D. Robert Weir

The complex mineralogy of the Porgera resource has a pronounced effect on process selection for the recovery of refractory gold and silver. Intimate association of gold with pyrite necessitates a high degree of sulphur preoxidation in order to achieve acceptable levels of gold recovery. Acid pressure oxidation has been demonstrated to be the only commercially proven process capable of achieving the required level of pyrite oxidation. Gangue mineralization dictates preconcentration of the sulphides to limit the acid consuming capacity and carbon dioxide generation during oxidation. However, the relationship between concentrate sulphur grade and gold recovery to concentrate ultimately governs the optimum degree of sulphide preconcentration.

Jan 1, 1986

By Alan Boughey

The world's largest grinding mill operates at the Cadia Hill Mine in Australia. Svedala Industries supplied the 12.2-m- (40-ft-) diam, 20-MW Cadia gearless, semiautogenous grinding mill (SAG). This was a leap of more than 40% above the largest operating SAG mill. This article reviews the significant items involved in the development, design, manufacture, installation and operation of this mill. Description of the Cadia project The Cadia Hill Mine is located 25 km (15 miles) southwest of Orange, in the Central Tablelands of New South Wales, Australia. The mining lease covers about 3,100 hm2 (7,660 acres) of the Cadia Valley. The Cadia deposit is a large, homogeneous, low-grade, porphyry-type, gold-copper resource. Current reserves stand at 202 Mt (222 million st), grading 0.73 g/t (0.02 oz/st) gold and 0.17% copper. They contain 146.2 t (4.7 million oz) of gold and 347 Kt (382,500 st) of copper. The operation consists of openpit mining and ore processing to produce gold bullion and a gold/copper concentrate. The operation processes 17 Mt/a (19 millon stpy) of ore to produce 91 kt/a (100,300 stpy) of concentrate and 9.1 t/a (293,000 oz/year) of gold. The plant operates two 12-hour shifts a day and had a targeted availability of 94%. The plant facilities include a copper concentrator, o water supply system, an openpit mine, tailings pipeline and dam, and a concentrate pipeline and dewatering facility located in Blayney, about 35 km (22 miles) from the concentrator. Significant features of the process plant include o primary crusher and stockpile, a reclaim conveyor system feeding, the Cadia SAG mill (Figs. 1 and 2) and two 6.7-m- (22-ft-) diam ball mills (Fig. 3). A portion of the gold is extracted in a gravity circuit. The balance is extracted with the copper in the flotation section of the plant. Concentrate is thickened and piped about 35 km (22 miles) to the filter plant and rail loadout facility located in Blayney. Tailings are pumped to a tailings dam and process water is reclaimed from the tailings dam. The water system is designed to control and capture available runoff from local creeks, and to store these waters in dams and weirs. In addition, a major portion of the water for the process plant is treated effluent from the city of Orange and a smaller portion from Blayney Mill selection. The low-grade ore posed a dilemma for Cadia. To achieve a satisfactory economic operation, large-scale processing was required. The selection of the mills, equal to the largest designed and operated at the time, gave a plant configuration consisting of two 11.1- x 5.8-m (36- x 19-ft) SAG mills and three 6.1- x 9.1-m (20 x 30-ft) ball mills. This configuration did not provide the most satisfactory economic model for the project. A review of the single line configuration indicated that a significant reduction in the overall capital cost of the plant could be achieved by increasing the size of the mills and reducing the quantity. Increasing the size of the mills,

Jan 1, 2001

By R. A. Jankowski

Although the number of operating longwall mining systems has remained relatively consistent over the past five years, longwall production levels have increased significantly during this period. Longwall production currently accounts for approximately 50% of U.S. underground coal production. While longwalls are highly productive and offer other advantages, operations employing this method of mining continue to experience dust compliance problems. This increased longwall productivity has meant that far more dust is being produced. An improved understanding of the longwall face ventilation system and advancement of face ventilation technologies are necessary to ensure all face personnel are allowed to work in an environment that is free of excessive levels of airborne respirable coal mine dust. The Dust and Toxic Substances Control Branch of the National Institute for Occupational Safety and Health's Pittsburgh Research Laboratory has examined several basic principles of the longwall face ventilation system, and has evaluated the effectiveness of numerous improved face ventilation techniques for longwall mining systems. These include identifying improved techniques for measuring face ventilation parameters on longwall mining sections and investigating the fundamental relationship between face air flow and the entrainment and dilution of respirable coal mine dust. Studies have been completed to determine the impact on face dust levels from using belt entry air to ventilate the longwall. Novel methods have been identified to increase the amount of face air flow and to manage face air flow to most effectively minimize face workers' dust exposure. Unique systems of auxiliary face ventilation have been developed and evaluated at full-scale, simulated longwall test facilities. The theoretical and applied aspects of each of these principles and technologies are discussed. Application of these results, throughout the longwall mining industry, as documented from surveys conducted in the early 1980's and 1990's, have reduced the health hazard associated with excessive exposure to respirable coal mine dust.

Jan 1, 1998

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 Kunzo Amano

Fully-bonded rock bolts have been widely used in construction sites and mines in recent years. The purpose of this investigation is to prove the effectiveness of rock bolt supports in unstable ground through in-situ tests and numerical analysis. 2. MEASUREMENTS AND PROCEDURES OF IN-SITU TESTS The fully-bonded rock bolts as the mine road¬way supports were tested in the Fukasawa mine which is located in the northeast district of Japan (Fig. 1). The Fukasawa mine (one of the Kuroko mines) started the production in 1973, and the present monthly production is slightly less than 20,000 tons. The ore deposit which exsists at about 400 m depth is divided into several small lenticular shaped bodies. They are distributed apporiximately 1500 m in NW-SE direction and 700 m in NE-SW direction. The hanging wall is basaltic lava and the footwall is tuff. The rocks surrounding the ore bodies are soft but relatively intact, so the trackless mining system has been adopted from the beginning of the development.

Jan 1, 1982

By Wilson Blake

The occurrence of rock bursts in some deep mines throughout the world constitutes a severe operational hazard. In addition to endangering the safety of the miners, the rock-burst problem is a major factor in mining economically in deep mines because of the loss of production and expense owing to cleanup and repairs following a burst. Developing means of controlling or eliminating rock bursts has been the subject of intensive research for over 50 years. While much has been learned about the effects of rock bursts, little is known about their basic causes; hence, control measures have been at best only partially effective--attested to by the number and severity of bursts that continue to occur throughout the world each year. In the Coeur d'Alene mining district of north Idaho, the rock-burst problem continues as the silver and lead veins persist at depth. These veins are presently being mined at greater than 7,000 feet below the surface and will be mined to some 10,000 feet in the future. The majority of rock bursts in this district occur in pillars. While pillar bursting has been essentially eliminated in some deep-mining districts by longwall techniques, modification or application of this technique to the thin, irregular, and near vertically dipping veins of the Coeur d'Alene district has not been found to be practical. The creation of pillars is a consequence of the horizontal cut and fill mining method used. Figure 1 shows a vertical longitudinal projection of the silver vein at the Galena mine, illustrating the mining method. Stoping progresses between raises and is carried upward between levels. A horizontal pillar is created as a stope approaches a mined-out level. Recent broad-band micro seismic monitoring of rock-burst-prone stopes by the Bureau of Mines has shown that potential burst zones could be delineated weeks in advance of a burst and that a definite sequence of events took place prior to the occurrence of a pillar burst (1). Hence, time was available for control measures to be effected when a stope pillar reached a critical burst geometry.

Jan 1, 1971

By H. B. Oblad

Coal/pyrite fines were froth floated in a laboratory cell to examine the rate of pyrite recovery during conventional coal flotation. The effects of collector, particle size and carbon to total sulfur ratio on the kinetics of recovery were investigated. These data coupled with microscopic examination suggest that miniscule amounts of coal that remain attached to the pyrite surface affect the rate of flotation of pyrite.

Jan 1, 1985

The purpose of this paper is to demonstrate how the Computational Fluid Dynamics (CFD) technique can be an attractive approach to evaluate the performance of face ventilation systems. To present the capability of the CFD for such a study and for design purposes, a series of simulations using CFD have been performed at the Department of Mining Engineering, University of Kentucky, for different face ventilation arrangements (e.g. different scrubber geometries, brattice (curtain) and continuous miners locations, airflow quantities, etc.). Results from these simulations are displayed as three dimensional graphs and discussed. Needs for the field validation of these results and refinements of design methodologies are also discussed.

Jan 1, 1998

By R. Raby

During the 1980 decade, the rapid price increase for flake graphite has generated a high degree of interest for the discovery and the development of alternate occidental sources of supply in response to fluctuations in quality and availability from current producers. Alike many industrial minerals, graphite draws its complexity not only from its mineralogy and unique properties but also from the very large spectrum of industrial applications each calling for specific quality criterion. As new high technology uses are being developed, more rigid specifications will apply. Discovered during the latter part of 1988 by Mazarin inc., some 30 Km South of Fermont near the Quebec-Labrador border, the Lac Knife deposit was quickly identified as one of the world major long term alternate source of stable supply for this mineral. This paper also reviews the status of the project development as of late 1991, while updating previous papers published by Industrial Minerals1990 and Mining Magazine 1990.

Jan 1, 1992

By Jorge Amaya, Nelson Morales Varela, Francisco Saavedra, Gonzalo Nelis, René Gómez

Precedence constraints are broadly used in mathematical models and algorithms for mine planning to model, for example, slope angles in open pit operations or connectivity in underground mines. Unfortunately, these constraints do not capture relevant geometrical aspects for the design of the mine like minimum bottom space of pits or smoothness of the economic envelope in block caves. In this work, we present different approaches that utilize precedence constraints generating economic envelops that are more suitable for design purposes. The numerical experiences show that the proposed techniques produce high value envelops with better geometrical properties and that the methods proposed can be computed efficiently.

Jun 25, 2023