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By M. Pearson, L. Zunti

Pressure vessels in hydrometallurgy and many chemical process industries throughout the world require the use of specific lining systems to protect the parent vessel materials from corrosion and/or abrasion. Typically, these lining systems are required for processes that are corrosive and operate at elevated temperatures. Two basic lining system types arc: 1) refractory linings in conjunction with an impervious corrosion resistant membrane, and 2) metal-clad vessels consisting of titanium or other corrosion resistant metal bonded to a carbon steel shell. Operating conditions play a key role in selecting the type of lining system to be utilized for a specific application. Lining systems in turn have an impact on the process vessel size, which ultimately affects the capital cost of the vessel. The capital costs for each type of lining system must be considered for each application, as well as maintenance concerns, quality control, and delivery. It is clear that operating conditions play a key role in the final selection of lining systems utilized for specific projects based on initial capital cost, life-cycle maintenance costs, and overall service life. This paper will examine the fundamentals of each lining system, their inherent technical strengths and weaknesses, and present relative cost comparisons for each lining system as they apply to specific operating conditions and vessel sizes.

Jan 1, 2004

By C. H. Lee, K. Adham, D. Lemieux, J. P. Mai, A. M. S. Hussein, S. Francey

Copper concentrates contain an increasing amount of arsenic, as the older sources of cleaner ores are depleted. At the same time, the market-determined price penalties on arsenic rich concentrates are rising due to limited arsenic treatment capacity at smelters and increasingly stringent environmental emissions limits, necessitating arsenic removal from the copper concentrates. Heating the concentrate to about 700°C can remove significant amounts of arsenic (as elemental and/or sulfide); however, it is important to achieve that removal economically and with due consideration to post-roasting arsenic (toxic waste) treatment. A comparison is made among different roasting options (kilns, multi-hearths and fluid beds) to qualify their potential application to arsenic roasting of copper concentrates. Both direct and indirect methods of heating as well as inert, air and oxygen atmosphere options are considered. Technology comparisons are made with respect to the roaster’s operation, its off-gas treatment and arsenic stabilization. INTRODUCTION In the last twenty years, the arsenic content in globally produced copper concentrates have risen steadily due to the depletion of cleaner ores and the increase in complex concentrate production. Smelter contracts typically limit arsenic to less than 0.5wt%, with a penalty paid per unit of additional arsenic and other contaminants such as antimony. While penalties can vary, they can be on the order of $70/dry metric tonne of concentrate per 1% of arsenic. Strategies used by miners to avoid the price penalties have been to blend clean and complex concentrates, and to remove arsenic from complex concentrates to achieve smelter limits. Arsenic can be removed from copper concentrates by the heat-induced decomposition of the predominant enargite mineral (Cu3AsS4) at ~700°C. Roasting at such elevated temperature is a commercially proven option, but still suffers from process, cost and environmental challenges. After roasting, the released arsenic sulfide fumes can be separated from most of the entrained concentrate dust in cyclones. Then, the As2S3 laden off-gas can be oxidized to generate a still volatile oxide (As2O3), which can be further separated from the residual concentrate dust in a hot electrostatic separator. Next, the off-gas can be cooled with air or water, to either deposit the arsenic oxide in a dry or dissolved state. Dissolved arsenic acid can be oxidized to As(V) and reacted with iron to yield a stabilized ferrihydrite precipitate or scorodite that can be safely disposed. In the case of dry arsenic trioxide, it can be mixed with reactants (e.g. iron) and fluxes before thermal stabilization in a glass matrix, thus eliminating the need for costly oxidation.

Jan 1, 2019

By Joseph P. Brennan, Robert L. Vines

A historical comparison of the frequency of fatal coal mining accidents occurring underground in mines of various size ranges in the United States and in member countries of the Commission of the European Communities is presented. Frequencies are normalized on hours-worked and tons-produced bases. The comparisons rely on published and confirmed data and on theory for selective treatment of data in an effort to overcome reporting differences and basic differences in the mining conditions and system employed. Trends are identified ad projections made of the effect on accident records of various industry growth patterns. Conclusions are drawn regarding the reasons for differences between the records of U.S. underground coal mines and the records of those mines in European countries.

Jan 1, 1984

By R. J. McLaughlin

Acid mine drainage (AMD) can occur as a natural phenomenon through oxidation of sulfide minerals. Oxidation by air and water can be exacerbated by mining. Because AMD can exhibit low pH, it is capable of extracting heavy metals from mineral deposits. The result is highly acidic drainage with elevated heavy metal concentrations -the combination of which can cause extensive environmental damage. Because of the importance of these factors, this paper is focused on sulfate based AMD. Due to widespread mining, the occurrence of AMD is ubiquitous. As a result, several approaches to treatment have been developed. Chemical treatment using alkali is most commonly employed to treat AMD. There are alternative methods of treatment, as well as alternative ways of handling treatment sludges or concentrates. Dewatering of sludge or concentrates is beyond the scope of this paper.

Jan 1, 1996

By A. W. Bryson, J. D. le Roux, B. D. Young

Several rate models for the adsorption of gold cyanide from aqueous solution onto activated carbon are given in the literature. The purpose of this study is to critically examine their ability to simulate the adsorption process over a wide range of conditions. An account is given of a comprehensive set of experimental work on the kinetics of gold adsorption onto a coconut shell- based activated carbon. Equilibrium isotherms were also determined. It is concluded that the homogeneous surface-diffusion model fits the experimental data reasonably well and provides a good balance between simplicity and descriptive accuracy. The model contains two parameters: the extra-particle mass-transfer coefficient and an effective intraparticle surface-diffusion coefficient. Both can be evaluated readily from simple adsorption experiments.

Jan 1, 1991

By T. Sun, J. Kou, C. Sun, D. Gong

"HPGRs (high pressure grinding rollers) have demonstrated higher energy efficiency than conventional comminution equipment due to their application of particle-bed breakage mechanism. In the past decades, a large number of studies have been conducted to understand the characteristics of particle-bed breakage, however, few of them concerned the difference between single-particle and particle-bed breakage in compression comminution. This comparison is able to demonstrate possible advantages of particle-bed against single-particle in a breakage event due to the isolation of the effect of force loading mechanism. The present study investigated progeny fragment size and energy efficiency of these two breakage methods at variable energy inputs and particle sizes. The results show that at low energy inputs of around 0.25 kWh/t, the numbers of fines resulting from the two breakage methods are very similar. However, at higher energy inputs of around 1.0 or 2.0 kWh/t, particle-bed breakage generated significantly more fines. From the perspective of newly generated surface energy, particle-bed breakage is obviously more energy efficient than single-particle breakage at energy inputs between 1.0 and 2.0 kWh/t. It was also found that HPGRs demonstrate high energy utilization rate in generating new surface energy within their typical range of specific energy input. INTRODUCTION Rock comminution is an important and energy-intensive process in mining operations. Nowadays, decreasing head grade and mineral liberation size of ore deposits greatly increase the energy cost of comminution process. HPGRs have been successfully applied in the mining industry to improve the energy performance of comminution circuits. Typically, 10 to 30% of energy saving can be achieved due to the application of HPGR (Wang et al., 2013; Casteel, 2005; Günter et al., 1996)."

Jan 1, 2018

By A E. A Northcote

The drive to maximise performance, reduce maintenance times, improve safety of operation and maintenance on mining equipment while keeping both operational and capital costs down has provided good nourishment to the creative thinking of designers of skip loading facilities for hoisting systems. Early systems were as simple as manual loading which evolved into a surge bin feeding a skip through a chute finally maturing into a measuring flask discharging into a chute to feed the skip and more recently high capacity skip feed conveyors. This paper will compare the concepts of the skip loading flask and the skip loading conveyor systems over a range of winding distances and hoisted tonnages. The comparison will draw on actual installations and review operational, maintenance and safety. An order of magnitude capital cost estimate will be used to complete the financial comparison. This will enable the reader to understand the concepts and costs and assist in selecting the appropriate skip loading system.

Mar 21, 2011

By D. Wines

"Slope stability analyses have traditionally been undertaken in two dimensions assuming plane strain conditions. Although three-dimensional analysis techniques are widely available and used routinely for open pits, two-dimensional analyses are still more common due to the relative ease of model construction and the relatively rapid simulation times.Two-dimensional analyses will often produce different results to three-dimensional analyses for the same slope. It is generally thought that two-dimensional analyses will produce more conservative results. The main reason for the differences is the ability of three-dimensional analyses to account for the three-dimensional nature of the various model inputs, including the slope geometry, the distribution of soil and rock mass domains, the orientation of geological structures with respect to the excavation face, the orientation of the in situ stresses, and the distribution of pore pressure. In some cases, for a long, straight slope in basic geological conditions, two-dimensional analyses can provide a reasonable representation of the problem. However, in many cases, the inability of a two-dimensional analysis to represent the true three-dimensional nature of the problem will lead to unrealistic analysis results.This paper discusses the reasons, as detailed above, for the differences in two-dimensional and three-dimensional analysis results. Work by others is summarized, and additional numerical analyses are performed to provide an improved understanding of the effects of slope geometry, structural orientations, and in situ stresses on predicted stability. Case studies are presented for both stable and unstable slopes, and the behaviour of these slopes is related to the three-dimensional nature of the slope geometry and geology.The previous work, the new analyses performed here, and the case studies show that it is often important to provide a realistic representation of the slope in three dimensions in order to obtain reasonable stability analysis results. This is particularly true for hard rock environments where structurally controlled failure mechanisms are most likely. The paper also highlights the fact that back-analysed properties obtained from one analysis technique are not necessarily applicable to forward analyses using another technique."

Jan 1, 2016

By James W. Reeves

DuPont was the titanium metal pioneer who put the greatest effort into both R&D and commercial sponge and powder metallurgy production. This was concentrated in the period 1945 through 1963. The commercial results were development of the chloride process for TiCl4 using ilmenite, development and improvement of the Kroll process, development of a modified Hunter process to make titanium powder and commercial development of a direct powder process. The powder metallurgical process failed because of the inability to meet powder chloride specifications. All this work remained a trade secret until published in NMAB -392 by Mahla in 1983. There were other processes explored by DuPont, both improved sponge and powder processes. Much of this work has been continued by others but none, as of yet, resulting in new improved titanium metal capacity.

Jan 1, 2001

By D. Lindeman, R. M. Nendick

"From a technologist's point of view, there have been few fundamental 'breakthroughs' in the process of gold extraction since the turn of the century. However, the industry has made a great deal of progress in its response to the problems that have arisen: a changing demand for products, lower grade and more refractory ores, inflated labour and materials cost and the increasing control of governmental legislation. The trend in North America and South Africa to alleviate these problems has been to adopt a common policy to process larger tonnages of lowgrade ore through more efficient milling circuits. Here, however, is where the similarity ends.The purpose of this paper is to emphasise the different principles of practice employed by extraction plants in North America and South Africa to achieve the same common goal, namely gold bullion.2.0 INTRODUCTIONTwo very important aspects of extraction metallurgy for which the metallurgical engineer is responsible are plant design and plant commissioning.In North America and South Africa the importance of these aspects is due to the fact that both are related to the financial success of the project of which they form part. Good design is imperative if the capital invested in both the plant and the whole mining undertaking is to reap maximum profit. Furthermore, good design leads to a carefully planned and well executed start-up procedure resulting in bringing the plant into production on time, at designed efficiency and within budget.Due to the competent nature of the quartz pebble conglomerate and quartzite ore on the South African Witwatersrand, the characteristics of the ore type are not much different from one mine to the other. Thus a significant operating knowledge and experience has been built up over the years and there is considerably less risk in selecting a gold extraction process route."

Jan 1, 1992

By Michael D. Justason

This paper details the static and Statnamic load testing on a 600 mm square pre-stressed concrete pile. The pile was located at Pier 15 on the 20 million-dollar Bayou Chico Bridge Project in Pensacola, Florida. The pile was 10.5 meters long and had a design load of 1.3 MN. Three static load test cycles were performed in November, 1996. The Statnamic load test was performed in January, 1997. The Davisson Failure load of the first cycle of the static test was 3.7 MN, and the Davisson Failure load of the Statnamic test was 3.2 MN. The load-displacement curves for the two types of tests were very similar, indicating that the Statnamic load test performed well in the sandy soils. Strain gauges embedded in the pile showed 35% skin friction and 65% end bearing for both test methods. Strain gauges were located at equal spacing at five elevations in the pile. A toe accelerometer verified the rigid body assumption of the Unloading Point Method currently used in the analysis of Statnamic load tests. The data collected from the pile instrumentation was of excellent quality and represents some of the best research to date comparing static and Statnamic load tests.

Jan 1, 1997

Recent focus by terrestrial mineral explorers (and investors) on cobalt has increasingly highlighted the future critical need for this metal, with new-age batteries being the primary utilisation. We present an overview of the world’s current terrestrial cobalt resources and put this in perspective of currently known seabed resources, particularly polymetallic nodule and crust deposits. We focus on comparison of the key exploration, development and exploitation risks for such deposits, and provide a framework for financial evaluation and appraisal.

Jan 1, 2018

By Elizma Ford, Vincent Sithole

"Two shortened test procedures for estimating the Bond ball work index are compared for suitability of use on Zambian/DRC-type copper–cobalt ores. The first procedure is based on a batch milling test while the second procedure is based on a limited locked-cycle test. Both procedures were conducted using a standard Bond test mill. INTRODUCTION The Bond ball work index is widely used in the field of minerals processing as an indicator of ore grindability. The work index is determined by means of a locked-cycle test procedure that typically requires seven to eight cycles to stabilise at a re-circulating load of 250%. A typical test requires between three to four kilogram of ore sample and around eight hours of operator time. For geometallurgical test work applications, where a large number of variability samples are to be tested, it would be highly beneficial to use a test procedure that requires less sample mass and operator time. Certain test work laboratories, for example SGS and JK Tech, are already offering various types of shortened test procedures for estimating the Bond ball work index. This paper discusses a comparison of two shortened test procedures that were evaluated for suitability of use on Zambian/DRC-type copper–cobalt ores. EXPERIMENTAL Test Procedure 1 – Batch Milling The first test procedure is derived from a batch milling approach proposed by the University of Hacettepe (Aksani, 2000). Some modifications were, however, made to the data analysis method from that originally proposed. During this test procedure, a 700 mL volume of sample is milled in a standard Bond test mill for a pre-determined number of revolutions. The feed and product particle size distributions are measured for each test. These results are then used to back-calculate a cumulative breakage rate parameter (k) for each particle size, according to the following formula:"

Jan 1, 2015

"With the plethora of cable bolts on the market, geotechnical engineers can be baffled by the differences in performance claims from the manufacturers but perhaps more importantly they are sometimes at a loss in making an optimal selection matching behaviour with a particular site condition. To this end a new testing methodology has been developed that can cater for the varying types of cable bolt behaviours. The methodology is based on the laboratory short encapsulation pull test (LSEPT) as recommended in the British Standard and overcomes several deficiencies such as cable rotation. Of most significance however, is its inability to measure the post-peak residual performance of high load transfer capacity cables and consequently, there is an inference that these cables may be susceptible to sudden anchorage failure with relatively little yield. This paper details the results of anchorage performance tests of a range of cable bolts and the differences in performance with rock strength and drill hole diameter.INTRODUCTIONA research project was undertaken to develop a new laboratory-scale test apparatus to study the axial performance of fully grouted cable bolts that are used in the Australian underground coal mining industry. Work by Thomas (2012) and others in recent years found there were a number of deficiencies in the testing methodologies making it difficult to determine and compare the performance of the dozen or more different types of cable bolts on the market. This has meant that geotechnical practioners have been hamstrung in optimizing the design of ground support systems in underground environments that are increasingly more arduous.A cable bolt is a flexible tendon consisting of a quantity of wound wires that are grouted in boreholes at defined distances between holes in order to provide ground reinforcement of excavations (Hutchinson & Diederichs, 1996). They were initially introduced into the underground hard rock mining industry in the 1960s (Thorne & Muller, 1964) and since the early 1970s have been brought to coal mining operations."

Jan 1, 2015

By Franklin Allison

THE influence of cobalt and nickel on the properties of steel might be readily expected to be very similar. The two elements occupy close and somewhat unusual positions in the periodic table, their chemical behavior and compounds are so completely analogous and their various physical properties, including the ferro-magnetic characteristic, show such a remarkable correspondence, that there is no doubt of a strong resemblance existing between these two metals. However, the variety of, properties that are obtained from steel is due, largely to the fact that the element iron possesses allotropic forms. Alloying elements that affect these allotropic forms, or alter the nature of the allotropic changes, exert a profound influence on the resulting steel. In the case of nickel and cobalt, although these elements are very similar chemically and physically, it is found that they differ distinctly in their individual effects on the critical points of iron. Both nickel and cobalt form a series of homogeneous solid solutions with iron at temperatures above 1800° F. The difference in their behavior does not occur until changes begin to take place in the solid state below this temperature. An inspection of the respective thermal equilibrium diagrams at the lower temperatures makes the opposite effects of nickel and cobalt clearly evident.

Jan 3, 1927

By Kuo En Chang

Many rare earth - transition metal (RE-TM) compounds absorb/desorb hydrogen reversibly at room temperature and atmospheric pressure. For LaNi5 which is used as a battery material, the electrochemical nature of this process is well known. In the case of Nd-Fe-B which has found wide application as a powerful permanent magnet, the absorption of hydrogen has both significant advantages and disadvantages. However, a fundamental understanding of the hydrogenation process at ambient temperatures and its effect on subsequent corrosion behavior is currently lacking. This paper reports the results of the hydrogenation of NdFeB two alloys, (NdDy)15Fe78.3A10.7B6 and (NdDy)15Fe67Co5V4A12B7, which are compared with that of LaNi5. Results show that hydrogenation can be accomplished by either electrochemical methods or immersion in dilute sulfuric acid. For the latter method, the change in lattice parameter indicates that the level of hydrogen absorbed by the NdFeB particles corresponds fairly closely to an overall composition of Nd2Fe14BH2.7. Cyclic electrochemical charge/discharge experiments show that the maximum hydrogen discharge capacities of the NdFeB, NdFeB-Co-V, and LaNi5 are 40, 60, and 210 mA?h?g-1, respectively. Anodic potentiodynamic polarization measurements and calculated values of icon reveal that absorbed hydrogen in NdFeB-Co-V and the NdFeB may lead to significant increases in corrosion rates.

Jan 1, 1995

By Steve Wilson, Bern Klein, Jeff Drozdiak, Stefan Nadolski, Andrew Bamber

"In the adaptation of high pressure grinding rolls (HPGRs) to hard-rock mining applications, circuit development has mainly focused on placing HPGRs in the role of tertiary crusher in the preparation of ball mill feed. In order to take advantage of the potential energy benefits associated with HPGR and stirred mill grinding technologies, work was done to investigate the potential of and resultant energy benefits of placing the HPGR in series with a high speed stirred mill in replacement of the conventional cone crusher / ball mill arrangement. The proposed circuit was designed to operate with two stages of HPGR grinding followed by high speed stirred milling.Using a pilot-scale HPGR unit inline with a 20 liter stirred mill, the design criteria, circuit arrangement and consequent comminution energy requirements for a circuit processing copper sulphide ore from Teck Resources’s Mesaba Copper project were determined. These results were then compared to two other circuits, an HPGR / ball mill circuit and a conventional crusher / ball mill circuit processing the same ore.The testwork indicated at a proof-of-concept level that the HPGR / stirred mill circuit could potentially be appropriate for this duty and further work on the concept is merited. While the results indicate that the HPGR component of the circuit was more energy efficient than conventional cone crushing, the energy efficiency of the stirred mill was found to be lower than that of a comparable ball mill undertaking the same duty. However, increased energy efficiency of the stirred mill portion of the circuit is possible with further optimization and further work on the concept is indicated."

Jan 1, 2010

By E. Chen, M. Barati, K. Coley

The reaction between carbon and slags is of significant importance in direct smelting of iron ores, EAF slag foaming and copper slag cleaning processes. This reaction is known to occur via gaseous intermediates and it is important to have fundamental data to understand the relative importance of each reaction step. In this study, the rate of CO2dissociation on different slag surfaces is measured using an isotope exchange technique. It was found that increasing iron oxide content up to 30 wt per cent does not have a significant effect on the rate constant while increasing from this limit, increases the rate constant dramatically. Activation energy of CO2dissociation on slags of unit basicity and different iron oxide contents is in the range of 180±40 kJ/mol. It is found that rate constant dependence on oxygen potential in the slag phase is increasing with iron oxide concentration. In the case of copper making slags, the influence of copper oxide is found to increase the magnitude of apparent rate constant, while decreasing the activation energy considerably. Key words: Gas-slag reaction, CO2dissociation, copper slag, kinetics, isotope exchange

Jan 1, 2004

By P. S. B. STEWART

As part of a continuing study of mechanisms of breakage, the grinding at Nairne Pyrites Pty. Ltd. was traced by the addition of barytes and the results compared with parallel small scale laboratory tests. It was confirmed that the same basic mathematical functions described the behaviour of the full scale ball mill as those established in the small scale mill. However, an additional classification function was required to define the behaviour for the rod mill and was evaluated as a function of size from the experimental data. The classification and grinding of the pyrites ore was assessed from steady state size distributions around the circuit.

Jan 1, 1971

By A. H. Leuschner, P. J. Kruger, J. Kruger

INTRODUCTION This paper deals with some practical aspects of the use and interpretation of dosimetric methods for assessing the exposure of workers to natural uranium. Consideration must be given not only to the scientific value of the method of dosimetry, but also to the practicality of the method in the particular working environment. It is generally accepted (IAEA, 1976) that uranium does not present a hazard during the mining and extraction processes up to the final concentration stage, that is precipitation as ADU or calcining to the oxide. Further processing (e.g. conversion to oxides, fluorides, etc.) is generally associated with the nuclear industry. It is customary for uranium processing facilities to be managed according to industrial norms rather than according to those applied for the purpose of radiation protection. This is mainly because of the large quantity of material to be processed and the fact that uranium is considered to be of a low radiotoxicity. Although ventilation is used to some extent, provision is not made for the same level of protection as would be required in a radiochemical facility. It is typical in such a plant to find that the process itself may be enclosed and ventilated, but the ventilation is inadequate to cope with an accidental release of material. The working environment is poorly ventilated, sometimes only by natural ventilation, and one frequently finds areas with high airborne dust levels, designated as mandatory respirator areas. Protection of personnel is dependent on personal protective equipment. Radiation protection is not of prime concern and personnel are not specifically trained in this subject. It is in this type of environment that acceptable monitoring procedures and personnel dosimetry must be established. The effectiveness and practical application of the dosimetry must also be judged against this working environment. DOSIMETRIC TECHNIQUES The methods available for assessing personnel exposure, whether it be in terms of chemical or radiological hazard, include urinalysis, faecal analysis and [in vivo] monitoring. Faecal analysis does not present itself easily as a method for routine use. Techniques for [in vivo] monitoring have been developed recently and as expensive instrumentation is required, it is not generally available. That leaves urinalysis as a dosimetric technique for routine use, and it is probably for this reason that urinalysis is still widely used. Guidelines for the interpretation of urinalysis results, as originally provided by Neuman (Neuman, 1950), are still used in practice, even though it has become clear that this method has severe limitations as regards the assessment of the dose as a result of the intake of class W or class Y compounds (Alexander, 1974). Typically a level of 300 µg [U/C] urine can serve as an indicator of an acute exposure (above which chemical damage to the kidneys may occur) and a level of 100 µg [U/C ] urine can serve as an indicator for an investigation. Such levels are determined and are used in conjunction with several factors, such as mode of intake (ingestion or inhalation), solubility of compound (D, W or Y), means of intake (acute or chronic), environmental monitoring results and frequency of urine sampling. Since the ICRP concept for internal radiation limitation changed from that of the critical organ, i.e. the single organ of greatest significance under the circumstances, to that of effective dose equivalent, i.e. account being taken of the total risk due to the exposure of all tissue, the maximum permissible organ burden (MPOB) was replaced as secondary dose limit by the annual limit on intake (ALI) (ICRP, 1977). The ALI values are calculated from the committed dose equivalents of the various organs, and are used to determine organ burdens, which are then used to interpret the dosimetric results. Using the ICRP model (ICRP, 1979), a comprehensive calculation was made by Johnson (Johnson, 1980), giving organ burdens and excretion data in terms of intake for acute and chronic exposures for different categories of uranium compounds. The question arises as to what extent the direct measurement of intake can be utilised as a method of dosimetry. This will require a method of personal intake measurement for each individual worker, with both ingestion and inhalation being taken into account. A measurement of

Jan 1, 1981