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By Dirk A. Benham, William A. Warfield

INTRODUCTION Whenever a mine operator is faced with a decision between various mining methods or even equipment selection, he must be assured that his decision will have a positive effect on his economic operating goals. For instance, if a mine operator has a choice of using an LHD mining system or a slusher mining system, then his decision to use either must not only fit within the operational limitations of the mining plan (such as production requirements, physical features of the ore body, mine plant facilities, manpower requirements, etc.), but also assure that the ore is being extracted at the least possible cost. To aid in his decision making process, the operator will commonly compare acquisition costs, operating costs, maintenance costs, and the operational advantages and disadvantages of each system in addition to considering his personal observations and feelings. A decision to select a mining system using only those criteria aforementioned will not necessarily lead the operator down the path to maximizing profits. In addition, the operator should also compare the economic effects of each system upon the overall mining operation. A relatively simple method for determining the economic effects is to calculate and compare the break-even grades for each system under consideration. By implementing such an analysis, the operator can now select a mining system which can operate within the limits of the mining plan as well as extract the ores at the least possible cost, thus yielding maximum profits. This paper will present the geologic and physical conditions of a mine being operated by the Hypothetical Mining Company in the Grants Mineral District. An LHD and Blusher mining system are under consideration and are compared by examining operating costs, equipment availability, and the conventional operational advantages and disadvantages of each. Finally, the economic effects of each system upon the overall mine's operation are compared by utilizing break-even grade analysis. DESCRIPTION AND MINING METHODS The Hypothetical Mining Company has located an ore body in the Grants Mineral District. After extensive surface exploration drilling and cost considerations are evaluated, it is decided to open a mine. The following tasks must now be performed: 1. outline ore body and it's basic characteristics, 2. look at mining methods currently in use in the area, and 3. determine mining techniques and methods for exploitation of the ore body. For the sake of argument, the newly discovered ore body is located towards the western edge of the Grants Mineral Belt. The host rock for uranium mineralization is the Westwater Member of the Morrison Formation. The Westwater Canyon sands are overlain by the Brushy Basin Member of the Morrison Formation and underlain by the Recapture Member, all of Jurassic Age. Westwater Canyon sandstone is a fine to very course grained subarkosic sand with interbedded green and red mudstone. The Brushy Basin Member consists predominantly of greenish-gray mudstones and contains thin lenses of sandstone and a few thin beds of limestone. The Recapture Member consists of alternating thin beds of reddish brown, grayish red, and light gray sandstone and siltstone. The Recapture interfingers with the Cowsprings Sandstone in the area of the mine. The mineralization is based on the groundwater flow within the Morrison Formation. Primary ore distribution is based on three basic parameters: ore migration in a geochemical cell, lenticular mudstone interbedding, and enrichment along fault zones. In areas where two fault systems intersect, ore thicknesses of up to 30 m (100 ft.) are encountered. In those areas where the ore is well defined and thicknesses facilitate full face mining, the mining methods are determined more by ground or operating conditions, equipment availability, and access to working areas than they are in the fringe areas of the ore body where grade control is more important and split-shooting and break-even grade determines the mining technique. If a modified room-and-pillar mining method using systematic development is selected, better control of the ventilation, equipment utilization, and mining plan can be obtained. Figure I shows a typical mining area where the ore is thick,

Jan 1, 1982

By Lawrence W. Scott

The performance of RETIMET metal foam as a flame arrester on explosion-proof enclosures was investigated by the U.S. Bureau of Mines both in laboratory tests and at the U.S. Mine Safety and Health Administration's (MSHA's) Approval and Certification Center, Triadelphia, WV. The objective of this research was to develop a permissible pressure vent for use on lightweight, vented, explosion-proof enclosures. In laboratory tests, four grades of RETIMET, a stainless steel foam material, were evaluated: (1) 45 NC-7, (2) 45 NC-13, (3) 80 NC-7, and (4) 80 NC-13. Explosive gas mixtures were prepared by a dynamic flow system. Ignition was by a low-voltage arc. Each grade of RETIMET metal foam successfully arrested the flame front in all methane-air tests. To evaluate RETIMET metal foam on large, commercial size enclosures,a multi compartmented enclosure was designed and tested in MSHA's explosion gallery in Triadelphia, WV. Extensive explosion testing revealed that a minimum vent-area-to-enclosure-volume ratio of 11.33 in2/ft3 is required to keep internal pressure rises below 3 psig. The RETIMET metal foam functioned satisfactorily in all tests as evidenced by the absence of external ignitions.

Jan 1, 2010

By Daniel Mapa Clemente

Processing of tin slags in reduction furnaces was studied in order to optimize tin recovery and reduce recirculation rates of slags containing thorium and uranium oxides. A theoretical approach considering a mass balance of the process was developed and adjusted by experimental results in order to predict the final slag composition. A metallic phase (FeSnSi) was discovered trapped in the slag due to high apparent viscosity, raising the FeO and SnO2 contents of the slag sample. This FeSnSi alloy is rich in tin and its recovery would help to reduce the slags recirculation. High contents of ZrO2 were found in the slag, restricting it from reaching complete fluid phase. Through an analysis of the slag in the CaO-FeOx-SiO2 and ZrO2-CaO-SiO2 phase systems it can be concluded that final slags should be modeled aiming fluid phase at working conditions and additions should be planned in order to control ZrO2 levels.

Oct 30, 2017

Mining companies around the world devote considerable effort to production reconciliation and metallurgical accounting. However, although sampling, weighing and analysis methods used to obtain the raw data for metal balancing continue to improve, they are still areas that are sadly neglected. Using the latest metallurgical accounting package to improve metal balancing is fundamentally flawed if the data used are unreliable in the first place. The basic rule for correct sampling is that all parts of the material being sampled must have an equal probability of being collected and becoming part of the final sample for analysis. If this rule is not respected, sampling bias can easily be introduced. Unlike precision, which can be improved by replication, bias cannot be reduced once it is present. Sampling and sample processing procedures must therefore be carefully designed both to minimise bias and to obtain acceptable precision. The most suitable sampling location is the discharge point of a process stream, where increments can be obtained by taking a complete cross-section of the stream with a sample cutter. Sampling devices that take only part of the stream may introduce serious bias and must be avoided. The moisture content and wet mass of the material being sampled are of similar importance to chemical analyses for determining metal content. The accurate determination of moisture content is particularly difficult, because the moisture content of samples can change after the sample is taken, eg due to moisture evaporation. Hence, moisture samples need to be taken quickly and placed in sealed bags or containers that are impervious to moisture. The moisture content should then be determined without delay. To determine wet mass, a range of weighing methods are used in industry, including weighbins, weighbridges, conveyor weightometers and platform scales. The precision and accuracy of these methods vary considerably, so it is important to select a method with the appropriate precision and ensure that the weighscale is well maintained and regularly calibrated.

Jan 1, 2002

By Amir Asadi

Various damage modes that evolve with time in polymer composite laminates, such as transverse cracking, vertical cracking, delamination, and fiber fracture, as well as the interaction among them, influence the time-dependent degradation in modulus (creep) and strength (creep rupture) of polymer composite laminates. This study is focused on modeling the time-dependent evolution of transverse cracking in [90] plies of a multidirectional polymer composite laminate [0m/190n]. This model uses a two-dimensional variational analysis to determine the stress state at any given creep time. The elastic stored energy, determined using this stress state, is compared with a critical energy for fracture to determine the crack evolution with creep time. Model predictions are compared with experimental results to validate the model.

Jan 1, 2010

By J. M. Eldridge, P. Balk

Phosphosilicate glass films were formed, by reacting gaseous P2O5 with SiO2, over a large range of temperature (800° to 1200°C) and gas phase composition (nearly two orders of magnitude of effective P2Ospressure). The film compositions generally corresponded with the liquidus curve, delineating the maximum solubility of the tridymite Phase of SiO 2 in phosphosilicate liquid solution at the temperature of film formation. It is shown that the P2O5 concentration of the phosphosilicate liquid film tends to decrease by reaction with the underlying SiO 2 layer until the liquidus curve is reached. The validity of the thermodynamic argument used in this explanation is supported by the results of a determination of the composition of borosili-cute films, prepared by reacting gaseous B2O3 with SiO2 at different temperatures. The kinetics of phosphosilicate film formation were described by a model predicated on a steady-state diffusion of P2O5 through the film. UNDERSTANDING of the processes leading to formation of phosphosilicate and borosilicate glasses is of great importance for producing passivating layers on FET devices. Passivating films with optimum characteristics are preferably formed in a separate step, independent of the doping of the semiconductor.' The results of an investigation carried out to gain improved insight into the mechanism of glass formation are presented in this paper. It would be expected that application of the known Pz05-Si02 and B 2 O 3-SiO2 phase diagrams should be useful in extending understanding of the glass-forming processes. However, the question of the propriety of treating thermally grown SiO2 in these binary oxide systems by the methods of equilibrium thermodynamics must be considered when this application is attempted. Although Sah et a1.' and Allen et al. 3 investigated the kinetics of formation of phosphosilicate glass (PSG), they failed to adequately relate their diffusion models to the occurrence of experimentally observed phases in the PSG/SiO 2/Si system. Horuichi and yamaguchi4 investigated the diffusion of boron through an oxide layer and described their results in terms of a model similar to that of Sah and coworkers. More recently, Kooi 5 and Snow and Deal6 reported the compositions of PSG films formed by depositing P2 O 5 onto SiO2. These compositions apparently coincide with those at the liquidus curve delineating the maximum solubility of crystalline SiO2 in phosphosilicate liquid solutions. These authors did not discuss the thermodynamic implications of their results on the structure of thermally grown SiO2 films. The structure of thermally grown Sio2 films and that of vitreous silica are generally thought to be quite similar. Since the solubility of a substance depends on its structure, it is relevant that the solubility of vitreous silica in water7 is highly reproducible, like the solubility of thermally grown SiOz in phosphosilicate liquid. Furthermore, the vitreous silica-water system appears to be in true thermodynamic equilibrium (viz., the same solubility value can be approached from both supersaturated and under-saturated solutions). Sosman7 suggested that a type of two-dimensional lattice may form at the silica/solution interface, resulting in the observed solubility behavior that is characteristic of a crystalline solid. An alternative explanation may be that vitreous silica has a microcrystalline grain structure. Other investigators have suggested that vitreous silica has essentially the structure of B cristobalite,' or is composed of microcrystals of p tridymite or cristobalite, or a mixture of both. Presumably the grain size would be sufficiently large to minimize any appreciable contribution of the grain boundaries to the solubility of the crystalline matrix. The present investigation was carried out to clarify the significance of the boundaries in the Pa,-SiO, and B2O3-SiO2 Systems in determining PSG and BSG (borosilicate) film compositions. Furthermore, the kinetic data for PSG film formation were extended, using a wider range of formation parameters than were previously reported. One model describing the kinetics of film formation will be presented that is compatible with the thermodynamics of the Pa5-Si02 system. EXPERIMENTAL PROCEDURE Glass Film Preparation. SiO2 films (1000 to 8000A thick) were obtained by oxidation of silicon substrates in dry O2 at 1100°C. PSG and BSG films were prepared by exposing these layers to gaseous oxides obtained by reacting high-purity POC13 and BBr3, respectively, with O2. A double-columned saturator was used to ensure complete saturation of the N 2 carrier

Jan 1, 1969

By V. Lessels, D. J. Buckwalter, D. H. Rose

At White Pine Copper Co.'s operation in the Upper Peninsula of Michigan, the company has been losing more than four lbs of copper in each ton of sand tailings. With an average rate of 12,000 tpd of sand going to the tailing pond daily, the company has been losing more than 14 million lbs of the red metal annually. With an eye aimed toward lessening this loss, White Pine engineers in 1959 initiated a study on techniques for recovering this "lost" copper from the sands. Inasmuch as the flotation operation was considered to be optimized, a logical approach seemed to be leaching. For various reasons ammonia, ferric chloride, acid, and biological leaching were ruled out as impracticable. This led to a consideration of cyanide leaching as practiced in the precious metals industry.

Jan 8, 1967

By C. W. Wark, Fred Wise

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By Fred Wise, C. W. Wark

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By W. H. Corbould

THE township of Cloncurry is situated in the northwestern part of Queensland, about latitude S. 20° 42' 53" and longitude E: 140° 30' 25". Townsville is the port through which all the trade passes and is about 470 miles east-northeast of Cloncurry township. The Cloncurry copper district covers an area about 200 miles long by 40 miles wide. Unfortunately the mines are very scattered over this area, there being no large groups of mines. Mount Elliott mine is situated about 70 miles south from the township of Cloncurry, and Mount Oxide mine about 150 miles north of Cloncurry. The district has been known for the last 35 years, and in the early stages a large amount of capital was expended with little to show for it; but within the last four years the railway has been extended from Townsville to Mount Elliott and about 40 miles north of Cloncurry, which has made possible the treatment of the ores of the mines on this extension. The Mount Elliott company within the last few years has smelted and converted on its mine 169,688 tons. of ore for 20,554 tons of blister copper, containing 20,377 tons of copper, 35,642 oz. of gold and 28,841 oz. of silver, at a cost of about £33 per ton of blister copper, f.o.b. Townsville. This cost included all ' development, wages, salaries, general management, office, agency, legal expenses, freight, etc. The gold and silver contents would practically pay for freight to Europe, refining, realization charges, etc. The costs of supplies are as follows: Firewood 15s., coke, 75s., mine timber, 54s., and coal, 60s. per long ton; explosives: dynamite and gelignite, 50s. and blasting gelatine, 65s. 6d. per case.

Jan 1, 1915

By M. K. Barnett

[ ] ENGINEERING research consists, broadly speaking, in the investigation of the effect of the variations in a number of factors on some property of a product or characteristic of a process. The unambiguous description of the product or process requires that the property or characteristic be expressible in quantitative terms. For the purpose of practical control, it is likewise desirable that the independently variable factors governing the product or process be quantities, although this ideal goal is not always attained. Maximum efficiency-i.e., the attainment of the maximum amount of reliable, useful information relating to the influence of the factors on the property of the product or characteristic of the process, for the least expense-is an important requirement for the satisfactory operation of a research organization. Obviously, a necessary condition for the attainment of maximum efficiency is an adequate knowledge of the specialized field in which the experimental research lies. While necessary, this condition is not sufficient. Proper attention must also be given to the general principles of experimental design together with the modem methods for the analysis and evaluation of experimental data. The present discussion is confined to a single design, the so-called "factorial" design. This design is of special interest in that it attains its goal, maximum efficiency, by directly departing from the

Jan 1, 1947

By M Dusci, A Trueman, S Poos, R Baumgartner, M Brittan

The Canahuire epithermal Au-Cu-Ag deposit in southern Peru is a complex and challenging deposit for metallurgical studies. It has issues of gold locking in sulfides and non-sulfides, deleterious elements and acid rock drainage (ARD). From scoping study to feasibility study, there has been considerable improvement in the understanding of the important geometallurgical characteristics that need to be measured and modelled. This paper presents the current understanding and the requirements to continue to provide optimal models for mining studies.Significant effort has been put into understanding the genesis of the deposit. Improved spatial and temporal understanding of lithology, alteration, structure and mineralogy was used to build a geological and alteration model of the deposit.Early metallurgical programs quantified the effect of head grade on process plant recoveries, but the geometallurgical and statistical analyses did not identify other physical characteristics that influence these recoveries. The studies improved the understanding of the nature of locked gold which can occur as minute inclusions (< 5 ¦m) in sulfides (primarily pyrite with lesser arsenopyrite, pyrite being 100 times more abundant than arsenopyrite) and gangue minerals (quartz and carbonates) as well as solid solution in arsenopyrite and pyrite.The presence of pyrite and the limited buffering capacity of waste rocks increases the potential ARD, which will require mitigation costs before, during and after operations. Waste rock ARD potential has been classified and incorporated in the geometallurgical model. The current classification of waste rock at Canahuire, using international classification standards, is limited due to the complexity of the deposit.Complex deposit metallurgy necessitates a complex process flow sheet, including flotation and cyanidation. Economic recovery will be constrained by the relationships between economic variable grade, other elements and mineralogy. These constraints and inter-relationships are difficult to establish throughout the deposit with a single test. A follow-up geometallurgical program is under investigation and will attempt, with the help of additional test work, to map areas of higher and lower recoveries. It is hoped that metallurgical response can be linked to measurable parameters other than head grade. If this is possible, these factors can be incorporated into future geometallurgical models.CITATION:Baumgartner, R, Dusci, M, Trueman, A, Brittan, M and Poos, S, 2013. Building a geometallurgical model for the Canahuire Epithermal Au-Cu-Ag Deposit, Southern Peru - past, present and future, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 51-58 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By Everton Pizzio

Parafusos utilizados para fixação de partes submetidos a carregamentos cíclicos estão sujeitos a falhas por fadiga. No estudo do conjunto carcaça/corrediça de um compressor fixado por parafusos prisioneiros - ISO M24x3, que fixam o flange da corrediça falharam prematuramente por fadiga. O objetivo deste trabalho é desenvolver uma metodologia para otimizar o desempenho em fadiga do conjunto parafusado. Para tanto, foram realizados ensaios de fadiga visando o melhor processo de fabricação no desempenho em fadiga. Observou-se que o processo de laminação da rosca seguido de tratamento térmico é o mais adequado para o caso em estudo. A conclusão principal foi que a configuração atual de união do flange com a carcaça, resulta num baixo índice de aperto que é verificado através das cargas máximas que são transferidas diretamente ao parafuso e que aumentando o comprimento do parafuso através da inserção de uma bucha extensora, aumenta a vida em fadiga

Aug 17, 2017

By Michael Murphy

A multi-stage triaxial procedure has been developed for the enhanced servo controls and feedback systems at the National Institute for Occupational Safety and Health (NIOSH) rock mechanics laboratory. Due to the difficulty of preparing multiple rock specimens for single-stage triaxial tests at different confinements, multi-stage testing provides a huge advantage if only a few samples of coal measure rocks can be prepared from the drill core. A multi-stage test can provide mechanical properties, post-failure properties, and a complete failure envelope for a single specimen. For the multi-stage test procedure, a parameter used to define the slope of the stress-radial strain curve was developed and monitored in real time in order to successfully predict peak strengths at different confining stages. For this paper, multi-stage tests have been conducted on sandstones of varying strengths, including Berea sandstone which has well-documented properties, and fine-grained and coarse sandstone from a western US coal mine. For the tests conducted on the fine-grained sandstone, it was found that the ramp rate was an important factor in the success of obtaining peak strengths during the multi-stage tests. For each rock type, a table including Young?s modulus, Poisson?s ratio, unconfined compressive strength, friction angle, cohesion, residual friction angle, and residual cohesion are reported. The values from these tables have been used as input parameters for projects at NIOSH requiring numerical model input parameters.

Jan 1, 2014

By S Taylor, B Lumsden, G Long

Statistical metallurgical plant trials are becoming more widespread, as a valuable tool for increasing plant profitability. Generally, a single parameter like metal recovery varies under the trial condition while all other variables such as average metal feed grade or average metal concentrate grade are equal for the two test conditions during the period of the test. Metal recovery is usually the measured variable because concentrate grade is fixed by smelter terms, and plants operate to maximise recovery at the required concentrate grade. However, in all flotation plants there is a relationship between concentrate grade and metal recovery, and under the same conditions increasing one results in a decrease in the other. At MMGÆs Rosebery Mine the zinc concentrate produced is generally above smelter specifications, so zinc concentrate grade can vary during a test. During trialling of magnetic conditioning in the Rosebery zinc circuit, the difference in both zinc recovery and zinc concentrate grade were analysed. The analysis of changes in zinc grade and recovery during the trial raised some interesting questions about the relationship between concentrate grade and recovery. A paired statistical method was developed and tested to combine in each pair the change in concentrate grade and recovery in a single result that could be analysed by a paired t test. This method was found to be very sensitive to combined, varying changes in concentrate grade and/or metal recovery. It was also evaluated on plant test results from another flotation plant to verify its applicability in other plants. CITATION:Long, G, Taylor, S and Lumsden, B, 2012. Statistical plant trials - Efficiently combining concentrate grade and metal recovery results, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 355-362 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

Among the various methods of thick-seam mining already well-tried in practice longwall caving with extraction of a single slice, as thick as possible, abandoning the top and/or the bottom coal, seems to be the most suitable one for the Australian coal mining industry. The production units of the future should be equipped with heavy double-drum shearer-loaders, wide armoured conveyors, (with one or two central chains) and shield support. Material of this type has been developed for seams up to 6 m thick. Output figures of more than 6,000 tROM/d from a single face can be looked upon as a quite realistic expectation. The problems of co-ordinating such large production units with the whole layout of the mine are discussed. The factors to be taken into account in the planning work and for preven- tion of mine fires caused by spon- taneous combustion are dealt with. For economic reasons the highest possible utilization factor of the very expensive equipment must be aimed at. Furthermore, the seams or parts of seams to be extracted should be carefully selected.

Jan 1, 1976

Ore petrology will fundamentally determine the types of metallurgical processes employed for ore treatment, their operational efficiency and the nature of processing problems encountered. It is therefore paramount that the petrological charact

Jan 1, 1998

By Liu Liu

In the recent 10 years the rapid development of Chinese steel industry has attracted the attention of the world. The technological progress of converter steelmaking has played an important role in this process. First of all, slag splashing has been extensively adopted by Chinese steel companies in various sizes of converters from largest to smallest ones, and even including those using special hot metal containing V and Ti. Remarkable metallurgical results have been achieved. Up to now the average converter campaign life has reached nearly 10,000 heats, and the longest exceeds 30,000 heats in three melt shops. Based on this process, the long-life combined-blowing converter steelmaking technology has been developed. Slag splashing has been used to form a gas permeable "slag mushroom" on the top of bottom nozzles to protect the nozzles. In order to keep the effective bottom blowing, furnace bottom thickness control technique has been developed to prevent the furnace bottom growth. It makes the service life of bottom nozzle the same as the converter campaign life (average 10,000 heats and longest > 30,000 heats). The complete combined blowing was ensured through out the furnace campaign. Concentration product [C][0] is in the range of 0.0023 ~ 0.0030%, the T.Fe in the slag is lowered by 4 ~ 6%, and the [0] in steel is decreased by 100 ~ 250 ppm.

Jan 1, 2004

By S. J. Rothman, L. T. Lloyd

S. J. Rothman and L. T. Lloyd (Argonne National Laboratory)— Hagel's paperz0 discusses the question of the mechanism for self-diffusion in bcc transition metals. Recent work21"23 has led us to agree with him in reabandoning the ring mechanism in favor of a defect mechanism, most likely vacancies. On the other hand, we think that Hagel's attempt to group together all self-diffusion in bcc transition metals is wrong. We prefer the scheme of Peart, Graham, and omlin," who separate them into two groups. One group contains the anomalously behaved metals P titanium,', y uranium, and perhaps zirconim.' The other group contains a iron," this alloy was held at 1800°C for 2 hr before quenching, its structure was all a. B. C. Giessen, I. Rump, and N. J. Grant (Authors' reply)—We very much appreciate the discussion by Deardorff and Kato regarding the a to /3 transformation temperature of hafnium. In the W-Hf program, as well as in a current study of the system Cr-Hf, standard precautions were exercised against contamination, including vacua of less than 10"5 mm Hg and a weight control check of all specimens to insure an impurity rise of less than 100 ppm. Following receipt of the above discussion, a revision was made in the annealing process. Arc-cast specimens were etched with a 10 pct HF-90 pct HNO solution until 1 mm of surface, including a possible oxide film, was removed. Annealing was done in a hafnium container which provided a long diffusion path for gaseous contaminants. Finally, the heat treating time was shortened. Structural changes were noted to occur quite slowly. A 99.5 at. pct Hf alloy showed 70 pct /3 after 30 min at 1700°C and was 100 pct /3 after 10 min at 1715°C. After 24 hr at 1740C, the alloy was found to be 10 pct/3. This change is due to contamination; it was estimated that it corresponds to an increase of 60°C in the transformation temperature over a 24 hr period. The corresponding change in 30 rnin would be slight. Using the same hafnium as in the above publication, and three new alloys at 98, 99, and 99.5 at. pct Hf, the transformation temperature was indicated to be 1715" 15°C. Since this hafnium contained 4.4 at. pct Zr, a correction of 40 5C, based on an assumed approximately linear transformation temperature curve, must be made for pure hafnium. This yields a value of 1755 20°C in good agreement with Deardorff and Kato's value. Because the slope of the boundary is small, this permits a good extrapolation, and we regard the W-Hf alloys as well suited to check the Hf transformation temperature. niobirn,'' and molybdenum,22 whose activation energies correlate with the melting temperature and whose Do values follow Zener's predictions.28 We have no quarrel with Hagel's placing chromium into the second group; we disagree with his interpreting our data to make uranium fit into that group. Hagel considers that the measurements of self-diffusion in uranium below about 950°C represent combined volume and grain boundary diffusion. This contention is based on the curvature of the penetration plot for our 891°C run,4 which also has a nearly straight line plot of log c vs x. Hagel has extrapolated these characteristics to our 803.5C run and fitted his Arrhenius plot for self-diffusion in y uranium to our four highest temperature points only. We think this is wrong for the following reasons.

Jan 1, 1963

By A. A. Lavrinenko

There are discussed the possibilities of using of flotation tailings of platinum-copper-nickel sulfide ores of the Norilsk region for the preparation of backfill mixture instead of a specially produced rock and binder materials to reduce the cost of stowing operations and impact on the environment. It is shown that due to the presence in the tailings of flotation up to 10% of sulfide sulphur, which reduces the strength of the fill materials, and more than 1 g/t PGM it is necessary the preliminary extraction of the sulfides in the product for further processing that give an opportunity to reduce the losses of valuable components. It is shown that by the flotation of the current tailings it is possible to extract 91% of sulfide sulphur, 60% of palladium and 30% platinum in 17% sulfide concentrate containing 2.1 g/t palladium, and to reduce the sulphur content in the tailings up to 0.7%. Due to the easy oxidation of pyrrhotite, there is recommended fractional delivery of the collector with a total consumption 200 g/t. There are discussed results of the flotation study with an additional collector of PGM (sodium diizobutyditiophosfinate), which one show the possibility to increase extraction of platinum on 45%, as well as extraction of sulphur and palladium. For the flotation of sulfides and PGM from the current tailings there is recommended a combination of collectors for primary flotation and only butyl xanthate using in control operations. The possibility of using sludge fraction of the current tailings as a component for the preparation of the combined cementing materials after exposure of sulfides is analyzed. It is shown that the efficiency of the flotation of sulfides from this fraction can be enhanced using selective flocculants. There are discussed the advantages of implementation of completely hydrophobic butadiene-styrene copolymer in comparison with partially hydrophobic polymers. There is discussed an efficiency of the mechanochemical treatment of old tailings in turbo mill and sulphuric acid to remove oxide films from the surface of pyrrhotite before its flotation. Due to the pretreatment operation there is achieved increasing of sulphur extraction and after the introduction to the flotation circuit three control operations there is produced the fill material with sulphur content 2.5%. Keywords: tailings, copper-nickel ore, pyrrhotite, butyl xanthate, sodium diisobutyl dithiophosphinate, selective flocculation, butadiene-styrene copolymer, mechanochemical treatment

Sep 1, 2012