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Joseph E. Thropp, Jr., Indiana Harbor, Ind.:—To what do you attribute the fact that in some localities the by-product coke sells at a premium over the ordinary bee-hive coke for foundry use ? If the cell-structure is preferably soft and rather thick in coke for foundry use, as I understand it, do you mean that by-product coke has soft and thick cell-walls ? William H. Blauvelt, Syracuse, N. Y.:—As I say in one of the paragraphs of my paper, the tendency seems to be towards the manufacture of foundry- and furnace-coke under somewhat different conditions. The development of the best cell-structure in coke, of course, has not reached perfection, but we are making progress, and we are inclined to believe that under ordinary conditions the best furnace-coke, that is, coke with the hard cell-wall and the well-developed cell, is produced at fast coking-time and high heats, whereas the best foundry-coke seems to be produced at somewhat lower heats. In one of our plants we are making furnace-coke in 16-hr. coking-time, and foundry-coke in 20-hr. time. Foundry-men want the strong, tough coke. I think the furnace-men present will agree that the best furnace-practice requires a hard coke, with a well-developed cell-structure. Replying to the question as to the reason for higher prices being realized for by-product foundry-coke than for bee-hive coke, of course, the immediate reason is that it suits the foundry-men better. Perhaps one cause may be that we are able, by the selection of coals, which is not usually practicable at bee-hive plants, to get a quality and analysis best suited to foundry-work, with high carbon-content, and with cell-structure and physical strength especially suited to the foundry-men's ideas.

Jan 1, 1913

By N. Arbiter

Fracture under low velocity free-fall and double impact, and under slow compression have been-investigated. The pattern of breakage and the size distribution of resulting fragments for sand-cement and glass spheres have been determined. Photoelasticity methods were used to simulate the stress distributions in free-fall impact in order to explain the observed patterns of breakage. Oblique fracture planes, occurring only in free-fall impact, develop along lines which coincide with the trajectories of principal compression as determined by mathematical analysis Breakage efficiencies for different modes of fracture were compared for both types of spheres. For the same specimen and loading system, static loading and low velocity dynamic loading induce geometrically similar stress fields resulting in reasonably similar fracture patterns and shapes of fragments.

Jan 1, 1968

By M Carter

The Great South Basin is located beneath the northwest part of the Campbell Plateau, a+broad, nearly level, triangular shaped submarine platform, southeast of the South Island of New Zealand. It encompasses an area of 77 700 square kilometres and contains a sequence of Cretaceous to Tertiary sedimentary rocks which are up to 6000 metres thick near the centre of the basin (Shirley, 1983). The western boundary of the basin is adjacent to the steep Aucklands Slope. In the+ north it merges with the Canterbury Basin and to the south onlaps against the Pukaki Rise, a basement high. The eastern margin of the basin is bounded by the northeast - southwest trending steep linear Antipodes Scarp. In the extreme south the Antipodes Scarp curves westward and joins the north trending Aucklands Slope. The Campbell Plateau is dominated by tensional tectonics. The Antipodes Scarp represents a major region of transcurrent faulting initiated by Cretaceous rotation within the Campbell Plateau. Later movements are associated with the Alpine Fault of the South Island (Sandford, 1980; Anderton, Holloway, Engstrom, Ahmad and Chong, 1982).

Jan 1, 1983

By R. Salter, C. K. Chase

Although Asarco's Silver Bell Unit, 40 miles west of Tucson, Ariz., is known primarily for copper production, molybdenite is also produced as a byproduct in the 8000 tpd flotation mill. The Silver Bell ore comes from two pits approximately two miles apart. Copper mineralization is essentially chalcocite resulting from two to three-fold secondary enrichment in a highly altered zone. A typical porphyry copper deposit, Silver Bell ore is of medium hardness. The ore consists of about one part molybdenum to 65 parts of copper. In the copper concentrate which is the starting point for separate molybdenum recovery, the ratio is also one part molybdenum to 65 of copper. Molybdenum mineralization is primary. Enrichment is not evident but it is notable that the molybdenite is more common in the more siliceous, harder rocks. The actual molybdenite occurs as diversely oriented flakes of fairly constant grain size, mostly in fractures associated with quartz veins. Most of the mineral appears unoxidized; secondary coatings on the molybdenite flakes are fortunately rare. In May 1962 a modified, improved version of the molybdenite circuit was put into operation.

Jan 7, 1964

The assistance that can be gained by the use of aerial photographs in the geological examination of a mineral field is clearly illustrated by the results of a rapid reconnaissance survey of the newly-discovered Mount Fitton talc, deposits, situated in the north-eastern extension of the Flinders Ranges in South Australia.The talc deposits were found to occur about the margins of a large lenticular mass of dolomitic marble, interbedded with hornfelsic tillites and fluvio-glacial sediments of Upper Pre-Cambrian age. Aerial photographs show clearly the extent of the dolomitic marble on which the growth of a species of eucalypt-Eucalyptus Gillei-is prominently developed, and which assists in defining the limits of this relatively light-coloured formation. (See Fig. 1).Upon field investigation it was found that the talc, deposits are related closely to the margin of the dolomitic marble where the latter has been subject to minor folding and faulting. The deposits themselves have been formed by metasomatic replacement of schistose dolomite. The contact of the dolomite with the underlying and overlying glacial beds is also noteworthy for the presence of minor occurrences of copper, silver, lead and zinc minerals.An examination of photographs covering a wide area reveals an outcrop of a massive granite to the north-east of the dolomitic marble, Fig. 2, which was almost certainly responsible for the regional metamorphism of the dolomitic and glacigene sediments, and the source of the mineralising solutions.All these features, as well as the broader structural features of the area, which are extremely important in the understanding of, individual commercial depQsits, were quickly determined with the assistance of the photographs.It is useful, briefly to relate the procedure adopted in examining this recently developed field as an example which may be well worth following on other fields, even if exposures are not so clearly defined. The method to be outlined, of course, will be of maximum advantage. Where rock exposures are good, where the ore bodies are likely to be widely spread, and where such bodies are related to some major geologi-cal control. Typical of such deposits are pegmatitic, hydrothermal and related ore bodies associated with the periphery of an igneous intrusion, or ore bodies related to particular sedimentary horizons in a regionally metamorphosed province.Procedure.The first requirement is the preparation of a detailed reconnaissance geological plan of as large segment as possible of the potentially mineralised province. The use of aerial photographs is par...

Jan 1, 1950

By Fred W. Brackebusch

"The use of paste technology for surface tailings disposal may offer economic, technical, safety, and environmental advantages for both underground and surface mining operations. Paste is a non-Newtonian fluid having some shear strength during transport and placement. The strength is augmented after placement by consolidation, water loss, and by addition of chemicals. Economical dewatering systems are necessary to the success of surface paste tailings disposal. In an underground mine, the paste backfill plant can produce paste for surface disposal about 50% of the time because only about 1/2 of the tailings are needed for backfill. The distribution of paste to the surface stack involves expensive pumping equipment and thus a nearby location of the stack is desirable. The viscosity of paste can often be manipulated by modifying pH. Placement of the paste in a designed stack configuration depends on topographic and climatic factors as well as paste properties.INTRODUCTIONThe use of paste technology for mine backfill is becoming accepted for underground mines mainly because of ground control and economic reasons. Paste technology is just starting to be applied for tailings disposal with only a few hardrock metal mines worldwide using the technology; however, it has been used extensively for red mud disposal in the aluminum industry. The disposal of dry tailings is very attractive from environmental and safety points of view because of limited water which might potentially escape and pollute river systems. Dry stacking systems, however, are expensive because of filtration and truck transportation costs. Paste can be produced from most tailings streams by special types of thickeners. Furthermore, it can be transported in pipelines, therefore paste technology may be more economic than dry stacking but still have some of the major advantages. Understanding of paste technology has improved through experience with operating underground backfill systems in recent years and because of testwork and research programs by mining companies, consultants and equipment vendors. Thus we may be on the verge of taking a new look at tailings management using paste technology."

Jan 1, 2002

By B. K. Shibler, I. S. Solet, H. W. St. Clair

"This study was undertaken to provide a more adequate knowledge of the chemistry of the sulfide volatilization process for recovery of tin from low-grade Bolivian ores. The investigation was partly supported by the Department of State under its program for scientific cooperation with South American republics.The results of this investigation demonstrate the conditions under which cassiterite (Sn02) may be converted to stannous sulfide by reaction with pyrite or sulfur vapor. Quantitative data are given on the rate of sulfidization and on the vapor pressure of stannous sulfide.INTRODUCTIONAlthough the United States is the world's largest consumer of tin, virtually no tin ores occur within our national borders. Most of the world's tin supply comes from alluvial deposits in the Far East and in Africa, The major sources of tin ores are Malaya, Indochina, Bolivia, Belgian Congo, Nigeria, Thailand, Burma, China, and the United Kingdom. Bolivia is the only important source of tin in the western hemisphere. It was the chief source of tin f or the United States during World War II. The Government-owned tin smelter at Texas City, Tex., was built primarily for smelting Bolivian ores and depends largely on Bolivian ores for its operation. Bolivia is the only major tin-producing country in which the tin ores occur as vein or lode deposits requiring underground mining. Most of the tin from the Far East and Africa is mined by dredges from alluvial or placer depos its. It is recovered as a high-grade concentrate of cassiterite relatively free from impurities. The tin minerals in the lode deposits are more intimately associated with other minerals and more difficult to separate into a clean concentrate. Furthermore, a large part of the tin in the original ore is lost during concentration. The ores as mined contain only a few percent of tin, so they must be milled to separate the gangue minerals and produce a concentrate containing 30 to 40 percent tin before they are shipped to the smelter. The losses of tin that occur during milling vary from 18 to 65 percent. The average recovery of tin from Bolivian ores is probably 50 to 60 percent."

Dec 1, 1954

By John W. Straton

Following the completion of the paper presented in February, 1972, data were received from many other mines and some original respondents submitted more complete questionnaires. The computer program was changed to permit the results to reflect more clearly all the data reported by the mines. The original paper was revised to reflect the new information. There were significant changes in some numerical values, particularly where the size of the sample changed appreciably. Some limited samples considered originally in arriving at conclusions were not relied on in this revision. The revised tables and text are of the same design and format as the original paper.

Jan 1, 1972

By F Gervilla, J Laukkanen

The Kevitsa Ni-Cu-PGE deposit in northern Finland contains disseminated Ni-Cu-Fe sulfides, and platinum-group minerals (PGM). The 2.05 Ga-old ultramafic-mafic, funnel-shaped intrusion is intruded into a volcanosedimentary sequence. The major ore minerals are pentlandite, hexagonal pyrrhotite and troilite, chalcopyrite, cubanite and magnetite. Pyrrhotite and pentlandite are replaced by pyrite, millerite and heazelwoodite near the surface. Nickeline, maucherite and gersdorffite are locally abundant. A pilot feed sample of 575 t of the main orebody at Kevitsa had average grades 0.45 per cent Cu, 0.36 per cent Ni, 1.85 per cent S, 0.33 g/t Pt, 0.34 g/t Pd and 0.19 g/t Au, with 0.06 per cent Ni contained in silicates. Higher PGE contents (2 - 27 ppm) occur in a vertical N-S bearing fault zone. The PGM (n = 6022) mapped in specimens of four drill holes, consist of sperrylite 34.7 vol per cent, braggite and cooperite 18.5 vol per cent, geversite 17.2 vol per cent, moncheite 14.7 vol per cent, merenskyite 6.7 vol per cent, stibiopalladinite 4.3 vol per cent, sobolevskite 1.9 vol per cent, kotulskite 0.7 vol per cent, michenerite 0.6 vol per cent, keithconnite 0.3 vol per cent, rustenburgite 0.1 vol per cent, melonite, froodite, laurite together 0.1 vol per cent and only a few grains of sudburyite, irarsite, vysotskite, erlichmanite, hongishiite, Pt-Cu-alloys, three undefined Pd and Pt bismuthinides and antimonides and a Pd-Cu-bearing Pt-Ni-Fe sulfide. Most of the PGM, 62 per cent, are enclosed in silicates, 29 per cent are intergrown with sulfides, nine per cent are within sulfides. A portion of the PGE occurs in solid solution in sulfides and arsenides. Ni flotation yielded smelter-grade concentrates averaging 12.2 per cent nickel. The average sulfide Ni recovery is 78.8 per cent and total Ni recovery is 65.3 per cent. The average Pt is 10 g/t at 55 per cent recovery, Pd 7.7 g/t at 38 per cent recovery and Au 1.6 g/t at ten per cent recovery. Cu concentrates averaged 28.6 per cent Cu with an average recovery of 78 per cent and further 11 per cent in the nickel concentrate. The copper concentrate contains 4 g/t Au at 24 per cent recovery. The average Fe to MgO ratio 17.9 is well above the targeted minimum of 4.5. The pilot results show that the previously used TETA (triethylene tetramine) can be avoided and good quality, smelter-grade concentrates can be produced from the Kevitsa ore at economical levels of recovery by conventional flotation methods.

Jan 1, 2008

By B B. Beamish

Long-hole directional drilling for in-seam exploration was introduced to New Zealand at Huntly West Mine in 1989. Following the success of this initial drilling program the Coal Corporation of New Zealand (CoaICorp) purchased its own drilling equipment and proceeded to adopt the technique with in-housetraining. From July 1991 to September 1992 a total of 15 455 metres was drilled at Huntly West Mine. During this period several modifications were made to the drilling technology to suit the ground conditions. The resulting interpretations of the drill information enabled reliable seam definition to be incorporated in mine planning. Since January 1993, the technique has been transferred to the Huntly East Mine to assist with seam definition for the planned development westwards under the Waikato River. The benefits of this drilling technique to coal mine planning are now well documented in Australia and New Zealand as well as other overseas countries. The possibilities of adapting the long-hole drilling technique to hard rock applications has been illustrated recently with successful cross-measure drilling at Huntly East Mine.

Jan 1, 1993

By R. V. Higgins

Many contributions to methods of forecasting the performance of waterfloods of a five-spot pattern have been made by investigators. In most of these forecasting methods, a previously selected sweep efficiency is used, and other simplifying assumptions are employed to obtain the objective. In this paper the calculated oil recoveries from several five-spot single- and multiple-layer stratified reservoirs were made with a minimum of assumptions, and because of the techniques used, no previously selected sweep efficiency was required. As a result, the accuracy of the forecast is increased. Permeability curves and continuously changing saturation gradients are used in every section of the pattern. By means of a potentiometric model, the flood pattern is divided into channels, the centers of which are the streamlines. The channels are divided into cells. As flood water enters, the permeability and saturation of each cell change with time. In this study the production from multiple-sand layers of different absolute permeability but with the same relative permeability-saturation relations are combined. The results are shown for a single- and 10-layer reservoir with a permeability variation of 0.60. The oil rate, water-oil ratio, and recovery are shown as functions of time for a 2- and 20-centipoise oil starting in primary phase and after fill-up.

Jan 1, 1962

By R. F. Favreau, Patrice Favreau

"Ever since we have started to assist those responsible for blasting with blast simulation studies, we havebecome aware of an interesting aspect, namely that in 90 % of cases we are asked ‘‘Can you seek ablasting method which gives better fragmentation than the method which we use now’’. But after a visit to the site, and the evaluation of the fragmentation obtained with their present blasting method, it very often becomes evident that it is the ease of mucking and not the fragmentation which requires improvement. Hence we then proceed to do a simulation study the results of which are improved mucking rather then more fragmentation, and those responsible for the blasting always become very satisfied."

Jan 1, 2017

By Henry Siegmann

HISTORY OF PRODUCTION AND USE In 1899 the consumption of fluorspar in the United States was reported as 16,000 tons. The invention of the open-hearth method of steel manufacture, plus the beginnings of a fluorine chemical industry, pushed consumption of fluorspar to 250,000 tons by 1918. In contrast to these consumption statistics of the past, the most recent statistics available (1972) indicate a total fluorspar usage in the United States of 1.352 million st (Table 15.4D.1). Fluorspar remains the world's cheapest source of fluorine, and its unique properties make the mineral the most desirable fluxing agent for steel refining. The linking of fluorine with carbon in the late 1930s resulted in the production of fluorocarbon, and constant growth in production over the years has formed the backbone of a strong fluorine chemical industry. For example, in 1972 it is estimated that total fluorocarbon production amounted to 2.113 million tons, and that, in this process, over 1.351 million tons of acid-grade fluorspar was consumed. The producers of primary aluminum long have been concerned about the fast dwindling reserves of natural cryolite. They found the solution to their problem by utilizing fluorspar for the manufacture of synthetic cryolite and aluminum fluoride. Nuclear weapons, as well as the growing number of peaceful uses for atomic reactors, constitute a market for uranium hexafluoride. This chemical and chlorine pentafluoride, used as an oxidizer with rocket fuels, are products that are as much a novelty as fluorocarbons were 25 years ago. Judging from past success in the field of fluorine chemistry research, it is safe to assume that the market for these chemicals will grow in the future. Three major industries account for the bulk of the fluorspar consumption: steel refining, fluorine chemical, and primary aluminum production. These industries have had a pattern of fabulous growth in the past, and their future plans point to a continuation of this pattern. It seems logical that the fluorspar usage also should continue its upward trend. Considering the growth of the fluorspar industry since 1899, and the obvious dependence on imports of fluorspar as a raw material, one of the most puzzling aspects of the past has been the apparent lack of coordination between fluorspar production and fluorspar consumption. A few authoritative voices have spoken out concerning the need to expand fluorspar production in order to meet consumption of the future. During World War II, the U.S. Government took positive steps in this direction by authorizing purchases at attractive prices for its "stockpile program." As a consequence, new fluorspar producing facilities were built, not only in

Jan 1, 1976

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

By Michel R. Julien, Michel Lemieux, Jean Cayouette, Daniel Talbot

"The Louvicourt Mine located near Val d’Or, Quebec has been in operation since 1994. It produces copper and zinc concentrates. The tailings generated from the ore processing operations have a strong net acid generating potential. Louvicourt Mine, a grass root project, was designed for closure with the best available technology at the time of design. In order to inhibit short and long time acid potential generation, sub-aqueous disposal was selected at the design stage. Given the fact that disposal in a natural lake was ruled out up front for obvious reasons related to loss of natural habitat and risks of permitting delays, a man made facility built with dams was planned. This mine includes therefore the first fully man-made sub-aqueous tailings disposal facility built in Canada. The requirement of using sub-aqueous disposal had serious implications on the placement of tailings. This tailings facility located about 9 km from the mine site has been selected based on the available natural confinement, the favorable foundation and hydrogeological conditions. This paper describes the site and the design criteria that were used. A presentation is also made on the placement technique to maximize storage and its impact on the operation of the basin. The performance of the basin, the dams and the decant structures is reviewed.Résumé : La mine Louvicourt, située près de Val d’Or, Québec est opérationnelle depuis 1994. Elle produit des concentrés de cuivre et de zinc. Les résidus générés par les opérations de traitement du minerai possèdent un fort potentiel net de génération d’acide. Le parc à résidus de la mine Louvicourt, implanté dans un milieu naturel intact, fut doté dès sa conception initiale de la meilleure technologie de fermeture disponible à l’époque, soit la déposition sous couverture aqueuse et ce, dans le but de bloquer le potentiel de génération d’acide à court et à long terme des résidus. Etant donné que la déposition dans un lac naturel fut mise de côté dès le départ pour des raisons évidentes reliées à la perte d’habitats naturels et au risque d’introduire des délais dans l’émission des permis, un étang artificiel délimité par des digues a été conçu. Cette mine fut donc la première au Canada à être dotée d’une installation anthropique de disposition des résidus sous une pleine couverture aqueuse. L’exigence d’une déposition sous couverture aqueuse a eu des implications sur le mode de mise en place des résidus. L’emplacement de l’installation de gestion des résidus a été choisi en un lieu situé à environ 9 km du concentrateur sur la base du confinement naturel disponible et des conditions hydrogéologiques et géotechniques favorables. Cet article décrit le site et les critères de conception utilisés. Une présentation est faite sur la technique de mise en place utilisée pour maximiser l’entreposage et son impact sur l’opération du bassin. La performance du bassin, des digues et des structures de décantation est également revue."

Jan 1, 2004

By Joseph E. Worthington

Increases in the price of gold starting about five years ago, and rapid fluctuations since then, have caused a substantial popular interest in the noble metal. Gold prospecting had been at low ebb since the end of World War II in most of the United states except for a flurry of activity in Nevada occasioned by the Carlin and Cortez discoveries. The last five years, however, have seen a great awakening of interest in gold exploration as well as an opportunity for education for many explorationists in the geology and other vagueries of gold deposits. The present writers have participated over the last few years in a search for gold deposits throughout the United States. This search was greatly influenced by the geologic studies of Dr. Wm. Spence at the Haile Gold Mine in South Carolina as reported in a companion paper. Spence's studies resulted in definition of a geologic model that has proven useful in a number of other geologic environments in the united states. These programs were all carried out under the auspices of Cyprus Mines Corporation and permission by Cyprus to present these results is gratefully acknowledged. The writers, however, reserve for themselves all responsibility for the results and conclusions. The exploration program to be described began in the southeastern United States. Spence's studies provided a reasonable basis for interpreting and understanding the geology of the Haile deposit and also provided a means for predicting additional occurrences. This concept was then expanded to other gold occurrences in the Piedmont and became an effective tool for evaluating other prospects. Finally it became evident that the concept could be applied in some western gold districts. The following discussion will then include descriptions of other Piedmont examples and a few western districts that also appear to fit the pattern.

Jan 1, 1978

By M. K. Larson, J. K. Whyatt

Proper employment of numerical stress analysis design tools is based on the demonstrated ability of a model to capture key elements of the geologic site model and accurately simulate how these elements interact with a mine design. While these tools have progressed markedly, they are, at heart, a gross simplification of the abundant complexity of a natural setting and its response to mining. A generic deep longwall site model was developed that includes aspects of the geology of deep coal mines in the Wasatch Plateau and Book Cliffs coal fields of Utah. The site model contains a set of common features and observations of how these features typically respond to mining. This site model was the basis for evaluating use of empirical, boundary element and volume element stress analysis tools to analyze the distribution of stress around a deep longwall panel. More specifically, this evaluation examined shifting of stress to panel abutments and gob, distribution of stress in the abutment and deformation and failure of bridging strata. Measurements of abutment stress changes at two sites in the Wasatch Plateau region were used to illustrate model calibration. Overall, these comparisons highlight the considerable differences between methods. Volume element tools can incorporate considerable detail and have fewer underlying assumptions, but this detail carries a considerable computational cost. Boundary element tools are much more efficient. However, this efficiency comes at the cost of added assumptions. These assumptions were challenged by the presence of a strong sandstone unit in the overburden, leading to boundary element results that depart significantly from volume element results. Empirical rules are the simplest, but are even more burdened by assumptions, many of which are implicit in underlying cases. Insight into the nature and impact of underlying assumptions in each method is essential to the proper use of results in mine design.

Jan 1, 2009

By K. Oraee

Mining of copper ore and the related industries play an important role in Iran?s economy. Copper ore production in Iran accounts for 75% of the total production in the Middle East. All activities related to copper, that is, exploration, production, refinement etc are owned and managed by state owned companies. Decisions have been made to privatize all these activities. To assess the success of this privatization process, different components of the industry must be subject to economic analysis. In this paper, the cost estimation of all operations has been carried out. The total potential and attainable revenues have also been estimated using historical data available from the industry. Risk is a major factor in all mining activities and the copper industry of Iran has been subject to risk factors arising from many internal and external variables. These variables have been recognized and their effects evaluated. The effect of risk on the economic feasibility of the operations in the industry has also been analyzed.

Jan 1, 2011

By K. L. Lownie

For research and test data to serve a useful purpose, it is essential that it be derived from some known standard test procedure that allows for comparative analysis as well as independent replication and verification. Fan performance data provided by manufacturers is not absolute, and is influenced by the geometry of the test installation. For these reasons, fan test standards have been prepared in order to establish a clearly defined test procedure that allows for comparison of different fan designs, and also allows for the independent verification of test data. Internationally recognized Standards such as BS848, ISO5801, and AMCA 210 etc. are highly prescriptive, and precisely specify test duct configurations, measurement procedures, measurement planes and instrumentation to be used. Provided the laboratory tests are conducted strictly in accordance with the relevant Standard, the performance data will accurately reflect fan performance within the appropriate tolerances for efficiency, pressure and flow when tested in the defined configuration. Major differences in performance exist between a fan tested in accordance with a Fan Test Standard, and an identical fan installed in a mine ventilation system. The aerodynamic performance of any fan will be influenced by the quality of the fan manufacture, the flow system within which it is operating, the location of measurement planes, and the system stability. The flow distributions present in a standardized Laboratory test are unlikely to be repeated in mine ventilation systems, and this will also disadvantage fan performance. Unless the fan supplier takes all of these influences into account, there will an inevitable reduction in fan aerodynamic performance, and the end user will experience a shortfall in expected flow through the mine. The effect on fan performance can range from minor inconvenience to catastrophic fan failure resulting in major rectification costs and significant loss of mine production.

Jan 1, 2009

By D. G. Klaren

Self-cleaning fluidized bed heat exchangers are conquering the world for reason that they are able to solve any severe fouling problem in indirect heat transfer due to the circulation of cleaning particles through the tubes of vertical shell and tube exchangers. In most cases, these particles consist of chopped metal wire. Moreover, self-cleaning fluidized bed heat exchangers perform with excellent heat transfer at very low liquid velocities, which makes completely new designs of shell and tube configurations possible. Further, self-cleaning fluidized bed heat exchangers show a strong shear thinning effect in non-Newtonian liquids. All the above characteristics make the self-cleaning fluidized bed heat exchanger an ideal tool for the indirect heating of laterite nickel slurry in High-Pressure-Acid-Leach (HPAL) plants. Preliminary tests and evaluations of these heat exchangers have shown that the performance and price of these heat exchangers used for the production of nickel and cobalt from a laterite slurry, cannot be matched by existing conventional heat exchange systems. This article explains the principle, design, lay-out and enormous advantages of the self-cleaning fluidized bed heat exchange technology in a plant suitable for an annual production of approx. 45,000 ton nickel and 4,500 ton cobalt from laterite slurry through high pressure'acid leaching.

Jan 1, 2002