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I found the article most interesting and timely. I consider it imperative that mine case studies like this one are undertaken and published so that the practicing mining engineer and geologist can obtain a better idea of what is the preferred estimation and grade control technique to use, given the site-specific geology, grade distribution, etc. Mining companies, with operating gold mines, need to be encouraged to either initiate such a study in-house or support a research project by financing a graduate student at a local university. The latter method is preferable since it not only provides research support but also an excellent training opportunity for young mining engineers and geologists. Ideally, what is needed is exploration drill-hole, blasthole, and mine production data that can then be used to assess the efficiency of different estimation and grade control methods by comparing the results obtained using the drill/blasthole data with the actual production data. It is hoped that more mining companies will realize that the rewards from this kind of research greatly outweigh the sometimes perceived problem with releasing data to a university. I noticed that the authors used the blasthole data to represent the actual values, the validity of which relies heavily on the two key assumptions with regards to blasthole assays and in-pit ore-waste selection procedures. Could the authors comment in more detail on these assumptions and what criteria they used to check the validity of the assumptions? Since the article was published, have the authors had the opportunity to compare their estimates with the actual production data? It would be interesting to compare other estimation methods with OK and IK. For example, how well did the polygonal method used at the mine compare with the results given in the paper? I would like to mention that B.L. Kwa completed a similar study (Kwa and Mousset-Jones, 1986) using exploration drill-hole, blasthole, and production data from the Alligator Ridge mine in Nevada. The study compared a variety of estimation methods, including IK, and the final results are shown in Fig. 1. The IK method exhibits the smallest difference between estimated and production results and is closely followed by the indicator moving window (Ind MW) method. Reply by Y.C. Kim We do concur with Professor Mousset-Jones with respect to the merit of such case studies performed at a local university. In answer to his request for comment on the validity of the two key assumptions made in our paper, we can only say that these are reasonable assumptions that are very difficult to validate. The assumption of blasthole assays being accurate can be partially validated by a careful analysis of the nugget component of blasthole variograms. The assumption that the in-pit ore-waste selection procedure in use was an effective one can only be validated through comparison between the mine estimates of ounces of gold versus the actual recovered ounces of gold at the mill. Neither one of these two assumptions has been validated to date due to several reasons. For example, the mine went through a change in management since the study. Similarly, no compelling reasons have been presented to perform detailed variogram analysis of the blasthole data. We do, however, believe that the obtained results (i.e., indicator kriging performs better than OK in a highly skewed deposit) will remain the same in relative terms, regardless of how valid these two assumptions may be. The most recent com¬parative study by D. S. Lindsey of the Montana College of Mineral Science and Technology also shows a similar ranking between ordinary kriging versus probability kriging, which is an extended version of indicator kriging.

Jan 1, 1987

By Sidney A. Hiscock

INTRODUCTION Lead and zinc are today rightly regarded as sister metals. Historically, however, they differ markedly with lead being known and in general use as the metal since 3000 BC in several countries which at that time might have been described as mature economies. Zinc was not isolated and recognised as an industrial metal with distinct and valuable properties until about 1700-1800 AD (although it had, of course, been utilised as a constituent of brass for some 2000 years previously). c were well established as industrial metals by the beginning of the 20th century. The present paper briefly reviews the overall growth in world consumption* and changes in the three main areas which today represent the mature economies or industrialised areas, i. e. ,Europe, the United States and Japan. Trends in overall consumption are determined by the demands for the individual uses for lead and zinc. The changes in patterns of use in the industrialised areas - individually and collectively - and reasons for the changes are also considered. LONG TERM TRENDS - 1900-1984 The World Picture Since the beginning of the 20th century the world consumption of lead (3.9 million tonnes in 1984) has grown almost fourfold and that of zinc (4.7 million tonnes in 1984) is almost eleven times greater than it was in 1900. The long-term trend has been one of continuing expansion for both. metals although the pattern has been severely disturbed at certain periods, for example in the years immediately following the two world wars and at times of marked recessions in industrialised areas. Apart from such major setbacks, new levels of consumption have been established regularly every few years. Recently, however, consumption has not really grown and with only a slow and partial recovery in world industrial activity at present, there is no indication that the peak consumptions of 4.2 million tonnes of lead, set in 1979, and 4.8 million tonnes of zinc (1973) will be exceeded in the near future. The growth in world lead and zinc consumption since 1900 is shown in Figure 1 (which includes changes in copper and aluminium for comparison). Overall growth over the period 1900-1984 has been about two percent a year for lead and about three percent for zinc (annual growth for copper has also been about three percent). Increases in tonnage consumption and growth rates by decade for lead and zinc are summarised in Table I. However, the growth rates by decade sometimes conceal very large individual annual increases and decreases. For example, in 1975 lead consumption fell by 13 percent and zinc consumption by 22 percent compared to 1974. In 1976 consumption of lead and zinc rose by 12 percent and 18 percent respectively. The years when consumption 'landmarks' (ie one, two, three and four million tonnes) were first reached are shown on Table 2. Clearly, the pattern of shorter periods being required to attain each extra million tonnes of consumption was broken in the 1970's. The consumption of zinc overtook that of lead for the first time in 1940, and since 1946 has always been at a higher level. Trends in industrialised areas The two key industrialised areas at the beginning of the 20th century were Europe and the United States which between them accounted for some 95 percent of world lead consumption and 98 percent of zinc consumption. Figures 2 and 3 show the growth in lead and zinc consumption, and Figures 4 and 5 the percentage shares for each area since 1900. Europe has usually been the major user of lead except for some years in the 1920fs, most of the 40's and early 50's when consumption in the United States was larger. In 1900, Europe accounted for about 64 percent of world lead consumption and the United States just over 30 percent. Currently, Europe takes about 40 percent of world consumption and the

Jan 1, 1986

By E. W. Hawk

he power generation industry may be required to reduce carbon dioxide (C02) emissions if regulations related to global climate change are enacted. Coal-fired generation emits more than 80% of the power sector CO2 and would be a likely target for CO2 reductions. Compliance with the Kyoto Protocol will require a 33% reduction from the projected year 2012 emission level, even with moderate load growth. This article describes an analysis of power industry CO2 reduction options and their costs. It assesses how a generator would make compliance decisions under a mandatory CO2 emissions reduction program. Carbon sequestration, fuel switching and new plant construction are considered. The Global Climate Change Treaty signed at the Rio de Janiero Earth Summit in June 1992 set the long-term goal of stabilizing CO2 concentration in the atmosphere. The December 1997 Kyoto Protocol established short-term mandatory targets for the United States and 37 other developed countries. The United States' target is a 7% decrease by 2012 from the 1990 "greenhouse gas" level. Developing countries are exempted from reductions, according to the Kyoto Protocol, despite projections that indicate a significant increase in their emission levels. The Kyoto Protocol must be ratified by the US Senate. US power industry carbon emissions were 490 Mt (540 million st) in 1990 and are projected to be 680 Mt (749 million st) in 2012, according to the US Energy Information Administration. Coal-fired generation emits more than 80% of the power sector CO2. Assuming that the power industry is only required to reduce emissions by a proportional share, a 33% emission reduction, or 225 Mt (248 million st) of carbon, would be required to comply with

Jan 1, 1999

By David Muir, Slobodanka Vukcevic

Gold is commonly associated with copper minerals, particu­larly in the more geologically recent rocks of Northern Australia, Chile and the Pacific Rim. Traditionally, oxidised copper-gold ores are selectively mined and processed for gold by cyanidation, ensuring that levels of reactive copper in the ore are kept low to minimise cyanide consumption; whilst sulphide copper-gold ores are floated and gold is ultimately recovered as the by-product from copper smelting and refining (La Brooy, 1992). Very few process­ing options exist for treating copper ores which cannot be selec­tively mined or floated, and for treating floation tailings which often contain significant gold (Muir et.al.,1989). The ammonia-cyanide leach system stands out as a unique method of selectively extracting gold without copper using economic levels of cyanide (1-2 kg/t NaCN) and ammonia (1-5 kg/t). Although first patented in 1904, it proved difficult to control and predict per formance with different ores. However, recent research and development in Western Australia by Ruane (1991), Hayes and Corrans (1992), La Brooy et.al.(1991), Muir et.al.(1993), Muir and La Brooy (1993), has led to an under­standing of the mechanism, and the commercialisation of the process at the Paris Dump (W.A.), Akjoujt (Mauritania), and in the Eastern Transvaal (Laxen,1994). Efforts are now underway to pilot and develop a tailings treatment plant at Telfer following successful laboratory trials. To date, process optimisation has relied on trial and error and large numbers of tests covering a range of conditions. Thus the results are specific to the ore. At Murdoch University, process optimisation is now focused on measuring and control­ ling Eh, pH, NH3 [Cu(II)], and Cu-CN species at the optimum levels identified from the recent mechanistic studies. In this way, the optimum composition required to process any ore should be readily established and maintained.

Jan 1, 1995

By Arnolfo A. Valdivia

Prior to 1971, there were several clinical trials to evaluate programs for early detection of lung cancer. Among these, the Philadelphia Pulmonary Neoplasm Research Project,(3) the Veterans Administration Study published by Lilienfeld,(7) and the controlled trial of the Kaiser Foundation Health Plan showed an overall five year survival rate of 8% for newly detected cases (the same as the national statistic for unscreened patients). In 1971, the National Cancer Institute initiated three randomized, controlled mortality studies using lung cancer screening of persons at high risk (male smokers over 45 years old). The studies are being conducted at the Johns Hopkins University Hospital, the Mayo Clinic, and the Memorial Sloan-Kettering Cancer Center. The studies have slightly different designs in the combination of sputum cytology and chest x-rays. At the Mayo Clinic the study group is offered screening with sputum cytology and chest x-rays every four months, whereas the control group is advised to have an x-ray and cytology every year. No reminders are sent, and it is believed that only about 20% of the control group is screened. At Johns Hopkins and Memorial, both experimental and control groups are offered annual chest x-rays. The experimental group is additionally offered sputum cytology every four months.(5) At present all of the programs show that screening can detect cancers that are undetectible by other means. However, at this time mortality rates in the control and experimental groups are not significantly different in any of the three studies. OUR PROGRAM Our clinic is located in Grants, New Mexico and we provide most of the pre-employment physical examinations for the mines operating in the Grants area (Kerr McGee Nuclear, Homestake Mining, United Nuclear, Western Nuclear, and Ranchers). In the examinations, we obtain the previous mining history of the worker, a chest x-ray, a sample of sputum for cytological examination, and a blood sample. We also provide routine annual physical examinations of the workers, with special interest in the detection of bronchogenic carcinoma. In the early seventies, we did not have a definite surveillance program. We did not know whether we should have a program like the one started at Memorial or like the one started at the Mayo Clinic. After long consideration, we decided to have a program that does not demand a sputum cytology and chest x-ray every four months, but that allows as many chest x-rays and sputum cytologies as needed to diagnose lung cancer as early as possible. We believe that, if a screening method for cancer is to be optimally effective, it must detect the process at stages early enough for curative therapy. We order a test depending on the age of the miner, the race, the mining history, the smoking history, the radiation exposure levels, and the results of the previous chest x-ray and sputum cytology. With the help of the computer, we have a list of all the miners who should be watched closely because of age, race, mining history, smoking history, radiation exposure, etc. Examination of the miners is performed at our clinic, where all the records are kept. The sputum is collected there but examined in Grand Junction, Colorado, by Dr. Geno Saccomanno. There are two ways to collect sputum. The best way is to collect three consecutive morning samples. For this, we need the cooperation of the miners. They have to follow these instructions and mail the bottle containing the sample to Grand Junction. "Instructions for obtaining a good cough specimen" The enclosed plastic bottle contains a preservative solution, so do not empty out the liquid in it. When you go to bed, place the plastic bottle at your bedside where it will be handy in the morning. When you first get up in the morning (before breakfast) try to cough up some "phlegm" from deep in your chest, and spit it into the liquid in the bottle. Try coughing several times. If you have difficulty coughing, try inhaling deeply the steam from a teakettle (or home-type inhalator). Keep the amount of saliva (ordinary spit) that you put into the bottle along with the cough specimen as small as possible. Do not collect the "phlegm" or mucous that comes from the back of your nose. Put the cap back on the bottle, and shake it vigorously for two minutes. If the amount of material you have coughed up is quite small, then keep the bottle at your bedside for three or four days, and each morning try to add another cough specimen. After obtaining your cough specimen, repack the bottle in the mailing container, and attach the enclosed mailing label. It does not require any postage stamps. Unfortunately, some miners "forget" to mail the sample and end up with an incomplete physical examination. To avoid this some companies, like Homestake, request that we obtain the sample in our clinic by forcing cough and expectorant with a nebulizer machine. This method does not give as good a sputum sample as the previous one, but we do get a sputum sample for every miner. The policies of different companies, in regard to annual physical examinations are different. All

Jan 1, 1981

By Priscilla P. Nelson, Francis L. C. Fong

Introduction Fracture mechanics is the study of the behavior of cracks and their modes of propagation. During cracking, energy is consumed in creating new surface area. The more energy a material absorbs, the higher is its fracture resistance. This fracture energy is provided either by work done externally, by the release of elastic strain energy within the cracked body, or by a combination of these two effects. Fracture resistance is measured in terms of either the stress intensity factor, K, or the energy release rate (also referred to as crack driving force), G. For linearly elastic materials, these two quantities are related by the following equation: [ ] where [u] is Poisson's ratio and E is the modulus of elasticity. Both K and G are direct functions of the applied force and the crack length. The process of rock cutting with full-face tunnel boring machines (TBMs) involves the indentation of a rock surface by rolling disc tools. Excavation occurs as fractures, extending from the locus of chip indentation, propagate between concentric disc grooves and liberate rock chips from the intact rock face. Since disc cutting is achieved by a fracture process, it is logical to anticipate that differences in cuttability of a variety of rocks should be related to the variation of fracture material properties among the rocks. Such a relation has been demonstrated between the performance of full-face TBMs and the critical crack driving force for excavation in relatively high strength, brittle, low porosity, isotropic rock (Nelson et al., 1985). The objective of this paper is to report on the results of a continued investigation of the relationship between fracture material properties and the effectiveness of disc cutting. Additional reliable data on full-face TBM performance is difficult to obtain, so that the results of disc linear cutter tests performed by the Transport and Road Research Laboratory (TRRL) (U.K.) are utilized in this study. Fracture Material Property Evaluation If a crack tip is subjected to a perfectly plain strain condition, many materials will behave in a brittle manner, and the material resistance to fracture is constant during crack growth. This resistance is referred to as the plain strain fracture toughness, believed to be a material property. The standard testing procedure for this parameter is given in the American Society for Testing and Materials (ASTM) test designation E399. To ensure plain strain conditions, acceptable specimen geometries are specified. Unless these geometric specifications are adhered to, the measured fracture resistance is regarded as the apparent fracture toughness, and a subscript Q is used, as [ ]. The methodology adopted by ASTM E399 is within the framework of Linear Elastic Fracture Mechanics (LEFM). This approach assumes that material resistance against cracking is developed instantaneously, without initial stable or subcritical crack extension. Prior to crack advancement in rock, however, much of the crack driving energy is absorbed in a localized volume in front of the crack tip, referred to as the plasticity or microcracking zone. This capability for energy absorption substantially increases material resistance, and a crack may be arrested unless an increased crack driving force is applied. The LEFM or plain strain assumptions would underestimate the actual fracture resistance of such a material. Such a conservative approximation may be desirable in the design of structures against fatigue or brittle failure. In mining or tunneling, however, an underestimation of crack resistance can lead to the selection of underrated excavation equipment and a decrease in production efficiency. The development of fracture resistance in rock is best described by the material resistance curve, or R-curve, which is a plot of fracture resistance as a function of crack extension. Typically, peak resistance is only reached after a finite amount of subcritical crack growth. The

Jan 1, 1986

By J. M. Applegate, C. L. Morgan

Introduction In the same way that a blueprint communicates informa¬tion about a structure, a systems architecture should provide information that guides the construction and implementation of a computing system. Taking the blueprint analogy a step further, a plot plan can also show where, in the "bigger picture," a particular design fits (if in fact it does) and what a builder needs to know about the environment in which he or she will operate. In the context of business systems, a meaningful architecture should convey the following: • the primary purpose of computer-based systems rela¬tive to particular tasks or operations, • the relationship of the systems to each other and an indication of their interaction and dependencies, • the relationship of the systems relative to the organiza¬tion and its objectives, • indications of where IS/IT resources can best be allo¬cated to support the organization and • the overall context in which new systems and technolo¬gies can be assessed for both their impacts on current systems and on the organization itself. In theory, a well-designed systems architecture should allow an organization to focus its resources on those aspects of its business that matter most, thus, leveraging investments in information technology to a competitive advantage. In addition, the way that a computing systems architecture is represented should be un¬derstandable to non-IT staff, and it should be eas¬ily recognized by the people it supports. This pa¬per attempts to show that a meaningful systems model exists and that its applica¬tion to mining can achieve the desired results. The following discus¬sion is based on the au¬thors interpretation of a concept referred to as the computer integrated manu¬acturing (CIM) model, as developed by Allen-Bradley. An inherent strength of the CIM approach is that the model combines aspects of business-reporting structures with the manufacturing process itself. What results is a systems frame¬work that is aligned with relevant business processes and is, therefore, easier for non-IS/IT personnel to interpret. The successful application of the CIM model can be seen in a large number of both discrete and process-manufacturing facilities throughout the United States. This paper provides a brief overview of the CIM model and then describes how the model can be applied at a conceptual level to the mining industry. Finally, this paper proposes an extension to the CIM model that the authors believe will make it easier to understand and apply. CIM model As shown on Fig. 1, the CIM model begins with the most detailed levels of process equipment and works its way up to the highest level of consolidated financial reporting. Systems at each level are characterized by common attributes, users, time frames, etc., providing a clear picture of their interac¬tion. Further analysis of the relationships between systems and between levels provides insight into the opportunities for integration. Regardless of the industry or processes involved, the CIM model provides a framework in which the characteristics of systems and the data they manage can be predicted based on the nature of the business function they support. In general, as data moves upward from the lower rungs of the CIM model, it changes from de¬scribing the physical na¬ture of a process to describ¬ing its financial aspects (e.g., from productivity to cost). The time frame to which data applies also transitions from a very short duration, perhaps only milliseconds, to financial or regulatory reporting time frames that may span years. In a simi¬

Jan 1, 2000

By S. R. La Brooy, P. C. Earley

Gold and Resource Developments NL (GRD) operates the Macraes Gold Project in East Otago in the South Island of New Zealand, treating a carbonaceous, refractory sulphidic gold ore. The ore is concentrated by flotation, then reground, followed by CIL cyanidation, to minimise loss to the preg-robbing natural carbonaceous material. As the mine has progressed there has been some variation in the preg-robbing nature of the ore, with average overall gold recovery generally in the range 70- 75 %. The Macraes Gold Project has been operating since 1990 and still has an estimated 11 year mine life, which has stimulated the continuing drive to enhance gold recovery. Testwork in 1994 showed there was no improvement in CIL gold recovery after conventional acid pressure oxidation of Macraes Gold Project flotation concentrate. Hence roasting was initially considered as the preferred option to improve gold recovery, being the only established technology to deal with both carbonaceous preg-robbing and sulphide encapsulation. However Newmont Gold Company pressure oxidation technology, as used at Twin Creeks in Nevada, has proved to be applicable to the Macraes Gold Project concentrate. Pilot scale testwork has shown that both roasting and pressure oxidation could achieve similar CIL gold recoveries on the concentrate. However, capital cost, simpler arsenic stabilisation, the ability to operate autothermally at a lower concentrate sulphur grade and simpler environmental consenting factors, has tipped the balance in favour of pressure oxidation. Overall average gold recovery is expected to improve to approximately 84% with adoption of pressure oxidation. GRD has appointed Minproc Limited as EPCM contractor to install pressure oxidation at the Macraes Gold Project, together with upgrading the existing flotation and grind circuits at a cost of NZ$ 30M (US$ 16M). Construction is 60% complete and on schedule for 1 October 1999 start up.

Jan 1, 1999

By R. Poulin, R. Pakalnis, D. Peterson

Introduction A comprehensive project has been completed at the University of British Columbia that focused on the development of a commercially competitive composite cable bolt. Laboratory evaluation consisted of pull tests, preliminary shear tests, grout column testing and scanning electron microscope evaluation of the composite cable bolt. Field implementation at five underground mines throughout Canada were completed for technical and overall performance evaluation of the new product. The project began in 1989 under the sponsorship of HDRK Ltd., a consortium of mining companies including Inco, Noranda and Falconbridge Ltd. The objective was to develop a cuttable cable bolt that would serve as support for a continuous miner within a hardrock environment (Mah, 1994). The bolt was required to be cuttable by a continuous roadheader without adverse effects to equipment, personnel and the milling process. From the beginning, the focus was to employ existing cable bolt installation technology (Goris et al., 1994) whereby the cable bolts are grouted within a 4.8-cm-diam (1.08-in.diam) or larger employing a portland cement grout having a water: cement ratio between 0.35:1 to 0.4:1. The prototype, developed from Phase I, was technically comparable to that of conventional steel cable bolts. However, high cost and limited accessibility hindered its commercial acceptance. The second phase of the project was sponsored by the British Columbia Science Council and Pacific Pultrusions Ltd. of British Columbia to locally develop a glass fiber cable bolt. This phase of the project has resulted in the DAPPAM cable bolt (Pakalnis et al., 1994). DAPPAM Cable bolts are the second most used support method in Canadian underground mines after mechanical rock bolts. It is projected that 870,000 m (2.9 million ft) of cable bolts are installed annually in Canada (Poulin et al, 1994). The objective of the research program was to develop a glass fiber cable bolt with similar pullout strengths to the conventional steel cable bolts (267 kN - 60,000 lbf). More than 200 laboratory pull tests were conducted employing standard cable bolt test procedures, as outlined by the US Bureau of Mines (Goris, 1994), to assess optimum rod and surface composition, geometry, number required, pullout and shear characteristics, among others. This resulted in the DAPPAM cable bolt (Fig. 1). It consists of 10 individual 0.6cm-diam (0.25-in.-diam) rods enveloping a 2.2-cm (0.8-in.) outside diameter, high-density polyvinyl grout tube. The composite used in the construction of the rods is comprised of about 65% glass fiber within a 35% polyester resin matrix (by volume). Through pultrusion (a continuous fabrication technique), the fibers and resin are combined into a composite material. The strength of the overall composite is largely governed by the material properties, the alignment [ ] of the fibers relative to the applied load and the density and interaction of the overall components within the composite. A further modification to the composite was the addition of a sand grit to the surface of the individual rod. This increased the overall bond strength between the rod and the cement-based grout, used to solidify the cable bolt to the rock mass. The grout tube has an inside diameter of 1.9 cm (0.75 in.) and is rated at a pressure of 1.72 MPa (250 psi). During installation, the grout is pumped through the annulus of the grout tube thereby filling the hole from toe to collar. Laboratory and field trials concluded that a minimum drill hole diameter of 4.8 cm (1.08 in.) be used to ensure optimum grout coverage. Table 1 compares the design strengths (as tested) for DAPPAM vs. the ultimate strengths for a conventional 1.6cm (0.6-in.) steel cable bolt. The characteristics for the steel cable have been derived directly or indirectly by Goris et al. (1994). A design tensile strength of 289 kN (65,000 lbf) was proposed for the composite developed. However, recent tests have indicated maximum breaking strengths of more than 400 kN (90,000 lbf) (Fig. 2). Figure 2 shows the load vs. displacement pull out curves for steel and DAPPAM cable bolts. The methods of testing were standardized so as to be able to compare the results directly. The shear strength of the fiberglass cable is largely a function of the surface area sheared. Shear strengths of 89 kN (20,045 lbf) were obtained for the configuration shown in Fig. 1 when sheared normal to the axis of the cable. The testing procedure was similar to the shear tests performed on steel by Goris (1993). It is important to reinforce the implication that critical bond strength has on the success of an overall support system.

Jan 1, 1995

By A. B. Holland-Batt

In the 1960s and 1970s, incremental development rather than innovation was the norm for gravity separation. However, the last 15 years have seen a number of new developments in this area. The article reviews the traditional gravity-separation techniques and then examines the newer developments such as centrifugal separators. Emphasis is placed on novel devices or improvements that offer better control or improved performance.

Jan 1, 1998

By Michael Elliot-Jones

The fundamental competitive structure in world coal markets will be modified by the entry of such new participants as Colombia and the newer mines in South Africa. The US will, in most markets remain the swing producer, although a number of low cost US mines will continue to compete in world markets. Exchange rate swings will persist as a major hazard for coal exporters.

Jan 1, 1986

By R. J. Pruett

Selective flocculation enables the production of high-brightness kaolin products from kaolin ore once considered too poor in quality for beneficiation into paper-coating pigments. Anatase, Ti(1-X)FeXO2, with X = 0 to 0.05, is a common iron-bearing impurity in sedimentary kaolin that can be unresponsive to kaolin beneficiation processes, such as magnetic separation, due to its extremely fine particle size (<0.3 µm), or chemical bleaching, because of its chemical stability. Described here are selective flocculation technologies that separate anatase by flocculating anatase or flocculating kaolinite from dispersed slurries. Applying the advances in the chemical and mechanical dispersion of ultrafine particles and the selective aggregation of kaolinite or anatase particles have resulted in the selective flocculation processes practiced today that have increased process recoveries from those technologies practiced beginning in the 1960s.

Feb 1, 2012

By John Frostiak, Don Welch, David Landriault

The use of paste technology for underground cemented backfill is widely accepted in the mining industry today as a potentially cost effective alternative to cemented rock and hydraulic slurry fill. While paste backfill is not the appropriate backfilling method for all mines, there are presently 23 operating tailings paste backfill plants throughout the world. This new technology has gained increasing acceptance as a backfill method over the last decade and its use is now expanding into the surface disposal of tailings using the stacked, thickened tailings disposal concept. Over the next 18 months there will be ten new paste plants commissioned in the world. Five of these ten plants will be used exclusively for surface tailings disposal or a combination of surface disposal and underground backfill production. The thickened tailings disposal concept, developed by Dr. Eli Robinsky in the early 1970&apos;s, has been in use for decades at several mine sites throughout the world (Kidd Creek Mine site in Northern Ontario), but always utilizing non segregating high density tailings slurry central discharge, stacked tailings placement. With the use of a higher solids content paste, the typical 2-degree beach angle can be increased to ten degrees or more with no runoff water and minimal shrinkage crack development through desiccation. This paper presents the design of the first mine to implement true surface paste disposal, Barrick Gold Corp&apos;s Bulyanhulu Mine in Tanzania. The paste system will be commissioned in the first quarter of 2001 allowing performance data to be collected to define the full potential of paste technology for surface tailings disposal.

Jan 1, 2001

By A. V. Glembotsky, N. A. Greshnova, V. P. Seregin, V. A. Glinkin

Based on a new alternative depressant of zinc sulphide - dimethyldithiocarbamate- modern technology has been developed for the treatment of complex polymetallic ores. In addition to its general function (depression of sphalerite and iron sulphides to some extent), this double-acting reagent is a promoter for galena and silver minerals. Laboratory and plant tests, as well as industrial experience have all shown consistently higher silver recovery to the lead concentrate as compared with sodium cyanide, resulting in less pollution and a safer environment

Jan 1, 1995

By J. R. Caylor

The Mine Safety and Health Administration (MSHA) advanced several initiatives last year to further protect miners&apos; health. Regulatory efforts included the implementation of a Hazard Communication rule for all mines, work on the Single Sample rule for all coal mines, and the Plan Verification rule for underground coalmines. Development and application of new technologies are actively encouraged in preventing miners&apos; hearing loss and respirable dust overexposures. MSHA&apos;s successful x-ray program for screening of black lung among coal miners completed its third year. Compliance Assistance efforts are focused on Best Practices, outreach, and technical support for operators&apos; specific challenges.

Jan 1, 2003

By B. Towson, R. F. Dalton, G. W. Seward

This paper reviews the successful application of oxime reagents for the solvent extraction of copper and considers the role that well focussed research initiatives can play in bringing further radical change to the metallurgical industry. Particular reference is made to the Cuprex Metal Extraction Process (CMEP), for the treatment of low grade or difficult to treat sulfide concentrates, using the reagent Acorga CLX- 50 to recover the copper and the reagent Acorga ZNX-50 for the recovery of zinc from chloride solutions. Other new extractants, currently in development, are also referenced.

Jan 1, 1994

By Brian H. Kaye

Efficient design of grinding circuits for efficient production of ultrafine material requires rapid feedback of information concerning the fineness of material leaving the grinding device Generating this information requires efficient sampling and adequate rapid characterization of the material leaving the grinding machinery Recent developments in electrooptic inspection of dry and wet material leaving a grinding device are critically reviewed New methods for characterizing ultrafines are discussed A rapid operating automated cybernetic (intelligent) permeability device for automatic control of grinding circuits using Blaine Index and other fineness indices is described

Jan 1, 1983

By S. C. Suboleski

Forecasts indicate that the acid rain bill currently before Congress is likely to cause an increase in demand of 30 to 40 million tons per year of compliance quality coal from Central Appalachia. This represents a doubling of this segment of the coal business. This paper will examine how this doubling can be met, what the likely implications for pricing are, the major uncertainties associated with these predictions, and the likely outcomes of the growth in compliance coal. Finally, some implications for fuel planners will be discussed.

Jan 1, 1993

Design models for belt conveyors used in mines predict both the belt power requirements and belt tensions. Conveyors with slightly downhill profiles are more complex to predict due to the way that models such as CEMA interpret friction. Performance of a particular conveyor in relation to other conveyors in the same system requires an analysis of natural stopping times, braked times and take-up location. A review of CEMA-based PC models is presented for design of underground multi-flight belt systems.

Jan 1, 1995

By Carl C. Martin

The treatment of ores in preparation for fatty acid flotation is characterized by a wide divergence in methods, and attempts to define "common practice", even for the same type of mineral, are usually of little value because of the numerous exceptions. Such factors as desliming, pulp density, pulp temperatures, and reagents for conditioning are successfully employed under conditions representing almost extreme opposites. More precise control of variables is necessary than with sulphydric flotation processes It appears likely that variations result from independent development of a process from the multitude of variables that characterize fatty acid flotation and the ores of non-metallic minerals. Consequently, a symposium is a suitable way of presenting a summary of pretreatment methods for fatty acid flotation processes, together with froth control practices, as they exist today.

Jan 1, 1959