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By W. C. McClain

The amount of vertical subsidence occurring over a longwall operation is a function of the thickness of material removed, the quantity and quality of any fill material, the width of the extraction, and the depth. The relationship between these variables is empirically determined, and subsidence can be predicted using only a few easily established constants which are subject to local conditions, such as the angle of draw. Associated with the vertical subsidence are both tensional and compressional horizontal strains. The component of strain in the direction of face advance is temporary, while the component perpendicular to the direction of working is permanently retained by the ground. It is these horizontal strains and the vertical strain (tilt) which result in damage to surface structures. It is possible to ameliorate the consequences of mining subsidence by utilizing the well-proved techniques described in this paper in the underground operations and on the surface. INTRODUCTION In the past, American mining practice in stratified deposits has been mostly restricted to production by the room-and-pillar system where a sizable percentage of the salable product is left in place as pillars. These pillars are usually regularly spaced and provide a permanent support for the roof of the mine and the surface above. The longwall system employs an advancing face which may be as much as several hundred yards long and some sort of breaking and loading machine which moves back and forth along the length of the face. The roof is supported behind the face only to a distance required to provide working space for the production machine. The supports are then withdrawn and the roof is allowed to cave. The longwall system therefore possesses the various advantages of complete extraction, thus effectively increasing reserves, while at the same time eliminating the necessity and cost of installing artificial supports such as roof bolts and heavy steel girders over large areas. Longwall mining has long been practiced in Europe, where it has undergone an involved development, culminating in the recent perfection of powered-loading machines and automatic self-propelled supports, offering high productivity. With ever-in creasing

Jan 1, 1967

By H. R. Hanley

The characteristics of alloyed anodes and their influence on the products of electrolysis and power consumption have been noted previosly ously in the literature.' This paper presents data in continuation of the subject given in an earlier paper by the authors.2 As noted in the authors1 previous paper, the objectives of this work are the reduction in the cell voltage and the purity of the cathode product resulting from the use of alloyed lead anodes. The fundamental facts developed in the previous paper were: (I) that the presence of a small amount of calcium lowers the anode potential approximately 50 per cent. below that of pure lead; and (2) that thallium when present to the extent of 4 per cent. rendered the anode practically stable, thereby lowering the contamination of the cathode by lead to a very insignificant amount. The specific features determined in this recent work are as follows: 1. The lowest limit of calcium in the lead which will still give maximum lowering of the anode potential. 2. The characteristics of the thallium-calcium-lead anode. 3. The characteristics of the silver-calcium-lead anode. 4. The characteristics of the thallium-silver-calciurn-lead anode. It was determined that 0.1 per cent. calcium in the anode was sufficient to lower the anode potential approximately 50 per cent. Table 1 shows the characteristics of alloyed anodes tested. The anode composed of 0.1 per cent. calcium, 4 per cent. thallium, and 95.9 per cent. * Professor of Metsllurgy, Missouri School of Mines. t Professor of Metallurgy and Ore Dressing, Hea.d of Department, Missouri School of Mines. Assistant Professor of Metallurgy, Missouri School of Mines. 1 Among others, see C. G. Fink and Li Chi Pan: Insoluble .Anodes for the Electrol-ysis of Brine [Trans. Amer. Electrochem. Soc., (1926) 49, 83-1341 and other papers by these authors. U. CJ. Taintori, 8. G. Taylor arid H. P. Ehrlinger: Lead Alloys for Anodes in Electrolytic Production of Zinc of High Purity. Truns. A. I. M. E. (1929) 192-200. 2 H. R. Hanley, C. Y. Clayton, D. F. Walsh: Investigation of Anodes for Production of Electrolytic Zinc. Trans. A. I. M. E. (1930) 275

Jan 1, 1931

By M. J. Ferrante

The preparation and subsequent reduction of vanadium trichloride was investigated as a process for the production of high-purity vanadium metal. Chlorine was reacted with vanadium oxide in the presence of carbon to vanadium tetrachloride. The contaminant vanadium oxytrichloride was also formed and later converted to tetrachloride by reaction with chlorine in the presence of activated charcoal. Vanadium tetrachloride was stirred continuously during thermal dissociation to vanadium trichloride powder from which volatile contaminants were removed by distillation. The vanadium trichloride powder was reduced to vanadium sponge by controlled feeding of the trichloride into molten magnesium. Excess reductant and byproduct salt were removed from the vanadium sponge by high-vacuum distillation. Batches as large as 6.5 kilograms of sponge metal were prepared at an average reaction efficiency of 98 percent. Vanadium sponge after consolidation by arc-melting contained 820 to 11330 ppm of interstitial impurities and had a hardness between Rockwell B 40 and 70, This metal exhibited excellent workability at room temperature. The purification of impure vanadium metal by electron-beam melting resulted in lowering interstitial impurities from 1,890 to 504 ppm and hardness from Rockwell B 87 to 28.

Jan 1, 1968

By C. G. Anderson

Today, with a stringent economic and environmental climate prevailing in the copper business, there is increased interest in evaluating new processing alternatives for production. Hydrometallurgical pressure oxidation of copper concentrates is one of the more viable approaches and several technological candidates have emerged. Of these, an overlooked but, ironically the first industrially proven methodology, utilized nitrogen species catalyzation in the oxidizing pressure leach system to produce copper via SX/EW. Given its many inherent process advantages, this may prove to be a feasible process alternative for the future. In this paper, the history of the system and its application to copper concentrates and ores will be outlined. In particular, a novel methodology for effective recovery of precious metals from chalcopyrite concentrates will be discussed. Finally, the perceived economics of chalcopyrite concentrate treatment with this unique industrially proven process will be delineated.

Jan 1, 2001

By Michael S. Wieland

"Toxic fumes cause fatal andnonfatal incidents in underground mining, where the working environment tends to trap the fumes, hindering the restoration of non-harmful conditions. Workers can underestimate the residual fume toxicity and return to the work site too quickly. The toxic fumes from charge explosions depend upon a number things: formulation ingredients, mixture uniformity, water resistance, hole contaminates, rock hardness, and dust interactions. Transitory hole-to-hole waves can cause partial desensitization, ruin the reaction kinetics, and worsen the fume toxicity.Traditionally the toxic fume hazard for a candidate explosive is resolved from a requisite tally of fume components. The resolution starts with the fume spectrum, a tabulation of specie concentrations, transformed to standard reference conditions and readjusted to remove the air dilution. The resultant influence, which is referred to as relative fume toxicity RFT, is deduced by a formula rule with some chosen constants and the requisite concentrations. The RFT result is normally compared to an RFT criterion that represents the worst-case (maximum) tolerable fume toxicity, as stipulated by regulations or otherwise."

Jan 1, 1997

By J. T. Hollow

The Fort Knox Mine is located in Alaska?s interior, where the average ambient air temperatures range from -24°C (-11°F) in January to 16°C (61°F) in July. The mill processes a free-milling gold ore utilizing both a gravity recovery circuit and conventional cyanide leach/carbon-in-pulp circuits. Mathematical models have been developed to accurately predict the impact of leach circuit slurry temperature on gold leaching, carbon adsorption and cyanide destruction kinetics. Additionally, an energy-balance approach has been used to model the seasonal variations in slurry temperatures throughout the Fort Knox mill. The energy-balance model, combined with the kinetics models previously developed to predict plant performance, was used to justify a tailing thickener installation at the Fort Knox Mine. This paper describes the development of the energy-balance model and its prediction of plant performance assuming a tailings wash thickener to reduce heat loss from the mill. Also presented is an analysis of the post-project plant performance that validates the model and shows that the mine produced an additional 596 kg (19,155 oz) of gold while reducing mill reagent costs by $6,302,000 during the first 38 months of operation.

Jan 1, 2007

When the Mining Bill came before the Victorian Legislature not so many months ago, the question arose whether mining managers required education, and to what extent. Some members of the House expressed themselves in a way which appeared to leave the impression that mining managers would be all the better if they were not brought up in a scientific school.To try and prove what we owe to science I will give a short history of colonial mining from the earlier period to the present time and compare results.I believe I am right in saying that South Australia is the pioneer colony in mining and to her belongs the credit of not dnly being the first mining colony, but also that her beginning turned out so successful, as the history of the Burra Burra copper mine will testify. At about the same time a little galena mining was carried on nearer to Adelaide, but as this did not prove a success and was only worked by tributing parties, little notice was taken of it.

By Michael Godard, Peter Cleyle

"Suncor Energy’s Project Millennium increased oilsands production capacity by an additional 95,000 barrels per day to its present level of 225,000 barrels per day. This massive venture included an expanded Mine, new Extraction and Upgrading facilities, as well as, steam and utilities infrastructuresIn the Primary Extraction process, removal of the bitumen from the oil sand begins with slurry preparation and is followed by primary , secondary and tertiary separation processes. The final cleaning step of the bitumen froth is the Secondary Extraction process, that removes residual water, mineral and solids to produce a clean diluted bitumen suitable for Upgrader feedstock.This paper will review the flowsheet of the Millennium Primary Extraction facility, from its recent commissioning through the ramp up to design production rates. In addition, new technologies in the Primary Extraction flowsheet will be reviewed.INTRODUCTIONAs conventional supplies of oil and gas are depleted in Canada, the Athabasca oil sands industry of Alberta becomes an increasingly important source of energy. Today, high quality crude from Alberta’s oil sands accounts for roughly 30% of Canada’s daily petroleum production. According to government projections, over the next 7 years, oil sands production is expected to more then double, and by 2010 more than half of Canada’s crude oil is expected to come from the oil sands.Oil sands deposits consist of a layer of sand saturated with bitumen covered by a muskeg overburden and underlain by silt and clay. The oil sand is mined where the overburden is economical to remove, typically less than thirty meters thick. The Athabasca oil sand deposits around Fort McMurrray is ideal for mining. Bitumen can be described as a diverse collection of hydrocarbons, organosulfur, and nitrogen compounds."

Jan 1, 2004

By J. Rogans

This paper discusses the comparative performances of the three activated carbon adsorption circuits at the West Wits operations of Anglogold. These circuits consist of two pump-cell circuits (at Two Plant, formerly Elandsrand and at Three plant) and a conventional counter-current adsorption circuit (at Mponeng plant). The implications of the improved performance of the pump-cell plants are discussed.

Jan 1, 2002

By S. Narasimha Rao

Significant amounts of lateral load may act on pile foundations supporting structures such as transmission line towers, tall structures, bridge abutments, offshore platforms, quay wails, harbour structures and anchorages for water front structures. There were several theoretical and experimental studies conducted on laterally loaded piles in clays to determine ultimate lateral capacity and deflections under working loads. Under a lateral load, a pile deflects in the direction of load application and the soil in front of the pile offers resistance to this movement. The ultimate lateral resistance of a pile can be estimated considering the statics of the pile for a known distribution of soil resistance with depth. In this paper, considering the contributions from both lateral shear drag and frontal passive earth pressure, a method has been proposed to estimate the ultimate lateral capacity of rigid piles in clays. The lateral shear drag has been estimated in terms of adhesion and shape factor. A few load tests were conducted on model piles subjected to lateral loading and these experimental values are used in verifying the proposed method. The reported results obtained from both model tests and field tests were used in support of the predictions made. In all the cases, there is a good agreement between the predicted and measured capacities.

Jan 1, 1996

By William M. Dowall

The recently completed in-pit crushing and conveying project at Phelps Dodge's Morenci Mine is rated among the world's leading materials handling projects. The operation involves the crushing and transport of more than 109 kt/ d (120,000 stpd) of copper ore, using regenerative conveyors capable of providing more than 5.5 MW (7000 hp) of equivalent electrical energy. The Morenci Mine is located 330 km (200 miles) east of Phoenix, AZ. Phelps Dodge Corp. owns an 85% interest in the mine. The remaining 15% is owned by Sumitomo Metal Mining Co., Ltd. and Sumitomo Corp. Copper has been continuously produced from the Morenci district since the late 1800s. Until 1989, run-of-mine ore was transported from active mining faces to the Morenci and Metcalf concentrators by truck and rail haulage. In that year, a system of in-pit, semi-portable crushing plants and overland conveyors was commissioned. This substantially reduced truck haulage and effectively eliminated rail haulage in the pit. The system used two semi-portable, in-pit crushing stations equipped with 60/89 gyratory crushers. The crushed ore from both plants reported to the first of two overland belt conveyors, arranged in series, for transport from the pit to a coarse ore stockpile. The system was "backed-up" by existing locomotive rail haulage.

Jan 1, 1993

By Fennell B

A process for the recovery of cobalt from a speiss produced at Sulphide Corporation is presented. The speiss is leached in oxygenated sulphuric acid to dissolve the metallic iron at a pH of around 4. The resultant iron (11) is subsequently oxidized in the leaching vessel and precipitates as goethite. Acid is regenerated in the precipitation allowing dissolution of further speiss. Subsequent to the conversion of most of the iron in the speiss to goethite, conditions become sufficiently oxidizing to dissolve cobalt. Cobalt is not co-precipitated under the leaching conditions, yielding a pregnant solution containing the cobalt and some iron (11). The principal advantage of such a process is the significantly lower acid consumption over that obtained by complete dissolution followed by selective precipitation. An impure product of cobalt hydroxide can be obtained from the pregnant solution by neutralization.

Jan 1, 1983

By M. P. Hay

"For a model to be successful it must address the concerns and needs of a target audience. Further, the model must demonstrate the same patterns of behaviour as the real system, and need only be complex enough to meet this objective. If the model is too complex it can become unusable to those who need it most.With this in mind, SUPASIM has been developed for the practical metallurgist requiring only laboratory flotation rate data as input. Some 35+ successful plant simulations have been completed since 1985. The approach of SUPASIM is to characterise an ore, organise information in an understandable way and add structure to the complex flotation situation and hence support the metallurgist to interpret the relationship between variables and optimise the flotation process.The paper describes the basis on which the model was developed and sets out to provide an insight into interpreting laboratory flotation kinetics and their physical meaning in a production plant and the influence floatable gangue has on float performance and plant design. A case study illustrates how the ratio of slow floating mineral to gangue can influence circuit design and improve recovery on a plant scale.INTRODUCTIONThere is an enormous amount of useful data in the results of a laboratory scale rate test. Each variable of fraction and rate, for mineral, metal and gangue has a physical meaning in the plant. The link between float kinetics and the plant in respect of recovery, grade, pulp density, pulp flowrate, residence time, number of stages of cleaning and circuit configuration has been established by the author in the process of developing SUPASIM."

Jan 1, 2005

By Edmund J. Cardoza

WELCOME TO THE DEEP FOUNDATIONS INSTITUTE 26TH ANNUAL CONFERENCE AND WELCOME ESPECIALLY TO THIS SEMINAR ON NON-DESTRUCTIVE TESTING OF DRILLED SHAFTS, PRESENTED BY MEMBERS OF THE DFI DRILLED SHAFT COMMITTEE. MY NAME IS ED CARDOZA; I AM A VICE PRESIDENT WITH THE MILLGARD CORPORATION, A DRILLED SHAFT CONSTRUCTOR FOCUSING PRIMARLY IN THE EASTERN HALF OF THE UNITED STATES.

Jan 1, 2001

By Martin J. Kenwright

Within the famous Harrods department store, a new lift shaft has been constructed, servicing a depth of 22m below ground level as well as the existing 6 storeys of the store. The shaft will connect with a new goods tunnel which is to be excavated beneath the shop, as well as new floor space above ground level which has resulted from the infilling of a previously existing lightwell within the store. The construction of the deep shaft while maintaining uninterrupted use of the store by the Client led to a number of innovative solutions being undertaken. Empirical assessments were made of settlements anticipated from the construction process, which took place within a congested environment of existing foundations. Various elements were designed and installed to mitigate the effects of the inevitable ground movement on the surrounding overlying structure, which supported sensitive high quality finishes and mechanical and electrical installations The reinforced concrete lining of the shaft was contained within a cofferdam of minipiles. The minipile placement accuracy (1 in 150) was vital as it could not interfere with the subsequent tunnelling works. As the scheme progressed this became even more critical since the minipiles were needed as support for parts of the permanent works This paper describes:- the high quality placement and control of the minipiled cofferdam, the methods employed in assessing ground movements and the techniques used to mitigate against the effects of these movements.

Jan 1, 2000

By H. K. Lin

Residual cyanide in cyanidation tailings can be detoxified by its oxidation to cyanate, as in the Inco and Degussa processes. However, cyanate, the product of detoxification and other oxidation processes, may be reduced to cyanide in reducing environments, e.g., deepwater discharge of cyanidation tailings. This study shows that the reduc-tion of cyanate to cyanide occurs and that the reduction rate increases with decreasing solution redox potential. Solutions with lower pH favor the reduction rate. No significant difference was observed between the results of tests prepared with distilled water and seawater, and the presence of cyanidation solid tailings does not show an effect on the reduction. Because cyanate can be reduced to toxic cyanide, causing environmental concerns, the deepwater discharges of cyanidation tailings must be handled carefully.

Jan 1, 2008

By James H. Cobbs

The utilization of large diameter drilled holes for mine shafts seems to have had its beginning in Europe about 1850 (28) where the mining industries of Germany and The Netherlands first used drilled shafts. Recent shafts in softer formations of diameters as large as 25 feet and as deep as 1, 700 feet have been drilled in both The Netherlands and Germany. The first shaft drillers utilized cable tool or churn drilling techniques which limit both hole diameters and drilling speeds. Though some churn drilling of shafts is still performed, rotary drilling has been the dominant technique for many years.

Jan 1, 1973

By Chris Bennett, Glen Kosick

"Process modeling and equipment sizing procedures for SAG circuits has advanced significantly over the past two decades. However, no matter how exact the hardness tests, models and scale-up calculations, if the ore samples being used for design do not accurately represent the orebody then major design errors can be encountered. This paper will describe a hardness testing method for identiQing the power requirement profile of an orebody, to an extent that design errors due to poor sampling can be greatly reduced or avoided altogether. It will provide some insight into the magnitude of hardness variability that can be encountered in an orebody. Finally, the paper will preview how detailed hardness distribution data can be incorporated into mine block models, to optimize SAG circuit design for an orebody, both fi-om a capital investment and throughput point of view.INTRODUCTIONIt has long been known that many SAG/AG' circuits experience wide swings in throughput. These throughput variations are almost always of a magnitude not anticipated in the plant design; consequently, they are detrimental to the expected return on investment for the mining companies. It is clear that the throughput swings are mainly caused by changes in ore hardness. In 1993, in an attempt to address this problem, MinnovEX began investigating the possibility of using a small scale SAG test (2 kg charge) to predict SAG power requirements from drill core samples. The objectives of the small scale test were to enable mining companies to do many more SAG hardness tests than currently undertaken for a particular orebody. To make this possible, the SAG tests had to use a minimal amount of sample, be cost effective, and give acceptable accuracy in the prediction of SAG power. If this small scale SAG test proved accurate, then the high sampling frequency would yield results that approach the true hardness distribution of an orebody. This hardness distribution could then be used to derive the distribution and variability of specific energy required throughout the complete orebody. This specific energy (kWh/t) can be thought of as the power required to grind one tonne of ore per hour (kW/tph) to a specified size. Once the true distribution of power requirement throughout the whole orebody is understood, the grinding circuit design team can make an informed decision regarding the hardness value that is most appropriate for sizing the SAG equipment. Throughput losses due to hardness variation could be deliberately planned for or avoided altogether (a capital cost decision)."

Jan 1, 1999

By V. A. Melezhik, P. K. Skufin, L. P. Nilsson

these factors into account. Theoretical and

Aug 1, 1994

By Zhuo Chen, Jun Zhou, Jiemin Zhou

Samples of the reacting suspension particles were taken from the reaction shaft of a commercial flash furnace at different heights. The processes of matte and slag formation and morphology of the samples in the shaft were observed and analyzed. Variations of matte and slag composition in the shaft were investigated. By analyzing the samples taken from the settler at different locations, settler processes were studied such as slag formation, agglomeration and separation of matte-slag phases, Fe3O4 reduction, copper-in-slag reduction etc. In accordance with the characteristic of matte and slag formation the flash furnace can be divided into 6 functional regions, and a mechanism model called as “Micro bath process” is proposed for matte and slag formation in a commercial flash furnace. INTRODUCTION During flash smelting, concentrate particles scattered in the reaction shaft complete their oxidation reactions in suspension. The droplets produced fall into the settler beneath with a small amount of un-melted solid particles, agglomerate and separate into matte and slag layers which are tapped out from their tap holes separately. With respect to the extent of slagging reaction in the reaction shaft, there isn’t any commonly accepted understanding in literature (Kimura, Ojima, Mori & Ishii, 1986; Kemori & Kondo et al.,1988; Kemori, Denholm & Kurokawa, 1989; Fagerlund &Jalkanen, 2000). Some assume that the slagging reaction in the shaft can be ignored and some propose that all slag is produced in the shaft. With regards to the characteristics of the physical-chemical changes of concentrate particle in the reaction shaft, although there are some research results reported from pilot scale and commercial scale experiments (Kemori, Ojima & Kondo, 1988; Hagni, Vierrether & Sohn, 1988; Jorgensen, Taylor, Dickinson & Stevenson, 1992; Swinbourne, Simak & Yazawa, 2002; Kivinen & Tuominen, 2005), they are generally limited themselves to theoretical analysis and calculations. No studies are published concerned with the processes of matte and slag generation in the shaft, quantitative analysis of the processes, variations of matte and slag composition and morphologies of the melts along the axis of the shaft, etc.

Jan 1, 2019