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Semiautogenous (SAG) ball mills are being increasingly used today for the coarse grinding of a range of ores, including gold, copper and iron. Larger mines are demanding mills with larger throughputs. This has led to some remarkable developments in mill diameters and installed powers. Conventional design principles - mills with trunnion bearings - have reached their limitations. This article describes one SAG mill with a slide-shoe arrangement and Krupp Polysius' Combiflex twin drive. In August 1997, the world's largest SAG mill with a slide-shoe arrangement was commissioned in Queensland, Australia, operating as a wet mill. The inside diameter is 10.4 x 5.1 m (34 x 17 ft) effective grinding length. The 1.36-ktlh (1,500-stph) SAG mill is driven by two Combiflex drives (twin drives with 2 x 5,500-MW or 7,375-hp installed motor power) in conjunction with a girth gear (outside diameter 11,890 mm or 468 in.). Commissioning of the SAG mill was completed with no notable technical difficulties. The plant runs continuously except for planned maintenance stoppages and process-specific reasons. As this article is written, the machine has been in operation for about 8,000 hours without major problems.

Jan 1, 1998

By Qi Jinduo, Gong Min

On the basis of actual parameter of wall control blast in Nanfeng mine, China, the process of dynamic stress changing between two column charges was studied recently for a presplit-cushion blast by holographic photoelasticity. In the experimental model, two holes were simultaneously detonated from both ends (four ignition points). The bottom of the holes were loaded more than other parts (heavy charge). Under these conditions, four consecutive isochromatic fringe patterns at different times were acquired in one loading run, and the characteristics of the stress distributions were compared at different places and times in the holes. It was found that interaction types of stress waves vary along the holes. The theory is posed that the stress wave which is formed by the interaction of earlier producing P wave and later S wave along a distributed line charge is the most important part in all stress waves. Research indicates that a heavy charge in the bottom of the holes is crucial to producing a presplit-cushion effect at the bottom. "End Effect" doesn't affect the magnitude of stress along linking lines between the bottoms. The chief fields of the "End Effect" are also explained in the paper.

Jan 1, 1998

A numerical model of the spontaneous combustion of coal has been developed and is being used to investigate the nature of self-heating in underground coal mines. Results to date confirm the complexity of the self-heating reaction in coal and illustrate the impact of the factors such as the coal reactivity, mass and airflow through the coal pile. Subtle differences in the chemical, physical and environmental conditions and make a significant difference in the development of a high temperature self-heating or a low temperature benign condition, the subtlety escaping normal underground observations. Additional work is planned to investigate the thresholds for such factors in the development of spontaneous combustion and it the implications for detection by gaseous Products.

Jan 1, 1998

By R. S. Austin

One third of the world's and most US kaolin is produced from Late Cretaceous and Early Tertiary sedimentary strata in Georgia and South Carolina. The kaolin is part of a belt extending to Arkansas of aluminous lateritic deposits, including bauxite. Original sediments were profoundly altered by chemical weathering during Early Tertiary time. Mining exploits those uncommon deposits where extreme leaching and chemical recombination have produced a clay of nearly pure mineral kaolinite. . Its whiteness, brightness, fine particle size, platelet shape, surface chemistry, and chemical composition make kaolin an important pigment, filler, extender and chemical compound in the paper, paint, plastics, ceramics and catalyst industries.

Jan 1, 1998

By Lu Fujun, Chen Xuewei

This paper will analyse the cause-effect relationship of the main factors influencing the production of the fully-mechanised coalface by mainly using the basic system dynamics theory. On the basics of the analysis, the simulated model of the sy

Jan 1, 1998

By A. Probst, N. Moores, J. A. Finch

"In 1994 McGill and INCO (Sudbury operation) initiated transfer of technology originally developed to detect level in flotation columns to thickeners. A conductivity probe was installed in a concentrate thickener to monitor the mud-line and solids load. The probe consisted in a vertical tube with 32 equally-spaced electrode rings connected in pairs to give 31 conductivity cells. Conductivity was measured as a function of depth (so-called conductivity profile) from which the mud line and % solids could be inferred. The probe was installed in October 1994 and has been successfully detecting profiles since then. This communication includes a description of incidents and experiences during this technology transfer process including: probe installation and start-up; software modifications to accommodate operating conditions and operators' needs; calibration procedures to adjust to changes in performance as a consequence of long term use; and preliminary observations on the last stage, still in progress, signal interpretation and use to facilitate operation, particularly by providing early reaction to upstream disturbances.IntroductionThickening is one of the basic unit operations in mineral processing and hydrometallurgy used to separate suspended solid particles from a solid-liquid dispersion by gravity settling. As the dispersion enters the thickener, the solids settle and clarified water overflows the top. A thickener, which may be continuous or batch, is basically a large vertical cylindrical tank provided with devices to feed the dispersion and to remove the settled solids. The tank provides residence time for the particles to settle. The dispersion enters the thickener through the feedwell whose function is to dissipate the momentum of the feed stream and to create relatively quiescent conditions around the entering point. Rotating rake arms move the settled solids towards the discharge point located at the bottom centre of the tank.There are normally three distinct zones in a thickener: a top layer of clear water, a middle layer called a sedimenting zone in which particles are settling, and a bottom layer called a compression zone in which particles are pressed together as a consequence of the weight of accumulated particles. Proper operation of a thickener requires two simultaneous conditions: no solids in the water overflowing the top and the target solids content in the underflow pulp. To achieve these conditions is difficult because sensors for detecting the sediment line (interface between clear water and sedimenting zone) and the solids content in the compression zone are not commercially available."

Jan 1, 1998

Abstract-This paper discusses the fundamental question, "what is an adequate sample size?" in the light of recent developments regarding parameters of Gy’s well-known formula for predicting sampling variance. New understanding of the granulometric factor, g, and the liberation factor, l, had significant impact on the results of calculations involving Gy’s equation. For the granulometric factor, justifications are given as to why nominal comminution size should always be reported as P95 (i.e., the screen size rejecting 95% of the mass). A small but important change in the model to estimate the liberation factor makes the difference between a predictive formula that works or does not. Other parameters and considerations regarding their practical implementation are commented upon in light of the author’s experience.

Jan 1, 1998

In contrast with mining in areas close to the main centres, mining in remote areas is characterised by comparatively high initial investment and relatively high running cost for the life of the mine. This is because of:the real cost of establishing and running the operation at a location distant from the necessary infrastructure; and the perceived risks priced into their costs by the multitude of suppliers involved in the project. Optimal co-operation between the mine and the supplier in developing the right rela tionship and the correct technical solution permit both cost elements to be reduced. This re duction may well mean the difference between success and failure. This paper examines the possibilities for optimising the contribution of the equipment supplier to the overall success of the process. Issues covered include the necessity to provide more than merely equipment fit for the purpose , but also the need to maximise the benefits of equipment ownership over the life of the mine. A simple method of determining the optimum timing for equipment replacement or rebuilding is proposed . Furthermore the maintenance contract (or performance agreement) is suggested as a means of leveraging the supplierÆs commitment to the mining operation. The specific examples are based on experience with mobile equipment but the methodology has relevance to other equipment and supplies.

Jan 1, 1998

By P Dight

Increasing the strength of backfill has been the subject of much research. In the former Soviet Union most methods concentrated on the fill mixture including chemical additives, binder re-grinding, water electromagnetic activation and the use of vibration and turbulent activation. For the most part, this research has been confined to the laboratory where the results were not representative of field conditions. These methods also were directed at increasing the mass strength and were not specifically aimed at strengthening the fill where it may subsequently be exposed for instance on the side or underneath as it required for underhand cut and fill methods of mining. In the Republic of Armenia, a method of fill strengthening known as electro-activation has been adopted. The deposit it an epithermal gold deposit which to-date has been mined using underhand cut and fill with timber support. The backfill comprises cement and pumice placed in 2.5 m lifts. Central to the method of electro-activation is the placement of the electrodes in the stope in which the polarity and the electrical potential are changing during filling. The process promotes the binding of the fill layers, the bond to the side walls and to the adjacent fill. There is a significant increase in the compressive and tensile strength. The process has been used with beneficial effects on reinforced concrete and could be used in any mine utilising hardening fills.

Jan 1, 1998

By J. J. Kollé

A compressed-water hydraulic pulse generator (HPG) for nonexplosive blasting of rock and concrete is described. Energy is stored in the HPG by compressing water to ultrahigh pressure. A large, fast-opening valve discharges the compressed water, resulting in a high-energy impulse. The HPG can be used for small-scale bench blasting by discharging the water through a nozzle that is closely fitted into a borehole. A bench blasting system requires an HPG, a drill and an indexing mechanism to allow insertion of the HPG nozzle into a drilled borehole. Systems with discharge energies ranging from 40 to 250 kJ have been used to excavate hard rock, including andesite, granite and quartzite. The largest system can excavate 15 t/h (16.5 stph) of massive quartzite based on a 2-min cycle time for hole drilling. Blasting with an ultrahigh -pressure hydraulic pulse generator was demonstrated as an effective means of excavating hard, massive rock under the confined conditions of a tunnel opening or underground sloping panel. Hydraulic blasting provides a continuous mechanized mining capability that is well suited for deep-level mining and urban excavation sites where explosives are prohibited.

Jan 1, 1998

By Seedsman R

The fundamentals concepts of roof support are similar between coal mining, metal mining and tunnelling and yet there are different design approaches, different hardware, and different levels of site investigation. This should not be the case and this paper discusses the key geotechnical issues that should underlie all roof support design - the need to reinforce defects. A design method to build a beam in laminated rock under high horizontal stresses is presented. The method allows for changing rock strengths, defect properties and varying horizontal stresses. At one level the design can be implemented with design charts and tables. Lessons learnt during the implementation of the designs at 5 different mine sites are Discussed.

Jan 1, 1998

By J. Ling, J. Xiao, A. R. Laplante

"The most challenging gold gravity separations occur when gangue has a high s.g., gold is fine, or a combination thereof This paper examines such separations, and evaluates the performance of a 3-in Knelson Concentrator at various rotation speeds, a 4-in SuperBowl, and a laboratory Mozley Multi-Gravity Separator (MGS). Synthetic feeds, primary cyclone underflows and flash flotation concentrates are tested.It is concluded that higher rotation speed are mostly useful for very fine (< 3 7 µm) gold recovery, with potential loss of recovery in the 37-212 µm range. This problem is traced to the collapse of the flowing film at higher Gs, which hinders the percolation of particles of intermediate size. Increasing fluidization water flow can alleviate this problem, but the recovery of very fine particle then drops. The MGS does not operate with the same principles, and can achieve good to very good recoveries of very fine to fine gold particles, but the unit has a high cost to capacity ratio, and there is no evidence yet that it can yield smeltable concentrate grade or a concentrate that can be upgraded with a shaking table.IntroductionThis presentation is a sequel to similar CMP presentations of the past eight years; all focused on practical aspects of gold recovery. The first hard rock application of the Knelson Concentrator (Laplante et al, 1990), studies of the grinding and classification behaviour of gold and other malleable metals (Banisi et al, 1991; Laplante et al, 1995), the development of a 3-in Knelson based methodology to study gold gravity circuits (Laplante and Shu, 1992), a comparison of the Knelson and Falcon (Laplante, Laplante et al, 1993), the presentation of a test to determine the amount of gravity recoverable gold, or GRG (Laplante et al, 1996), and the use of synthetic materials, in particular tungsten, to mimic gold (Laplante et al, 1995) were presented. A summary of many practical implication of the McGill research was presented in 1994 (Laplante). Whereas gold gravity recovery was less popular and largely jig-based (in North America) at the beginning of the eighties, many gold plants now incorporate a gravity circuit, and the Knelson Concentrator is the most commonly used unit."

Jan 1, 1998

By Patrick Clark

As part of the Mine Waste Technology Program (MWTP), a large-scale column test using bacteria and a sulfide ore was conducted at McClelland Laboratories, Inc., in Sparks, Nevada. Four technology providers participated in this demonstration. A mining company in Nevada provided the sulfide ore, process effluent and make-up water. A process effluent column, a chemical oxidation column, and the four biological treatment columns were set side by side on a test pad outside the laboratory building. The process effluent and chemical oxidation columns were used for experimental control. The goal of this project was to evaluate the viability and feasibility of using these biodegradation technologies to reduce the cyanide to the regulatory level of 0.2 parts per million (ppm) within an acceptable timeframe and with low operational costs. The column testing began on December 3, 1998 and operated until May 17, 1999. One technology provider reached the regulatory level within the 158 days. The remaining three biotechnology columns were approaching the regulatory limit at completion of the demonstration. The MWTP is funded by the U.S. Environmental Protection Agency (EPA) and is jointly administered through an interagency agreement with the U.S. Department of Energy (DOE).

Jan 1, 1998

Many studies have been conducted over the past 3 decades on the leaching of gold in solutions containing thiosulfate, ammonia and copper. However, most of these studies have been carried out using ore samples, with very few fundamental studies of this system being attempted. The majority of the published studies have failed to acknowledge the undesirable side reaction between copper(H) and thiosulfate, let alone account for its effect. The fact that the solution chemistry is constantly changing makes it very difficult to perform a fundamental study of gold leaching. With the advent of the rotating electrochemical quartz crystal microbalance (REQCM), such fundamental studies are possible as this instrument measures mass in real time. This has allowed leach kinetics to be measured in a matter of minutes and hence at essentially constant solution conditions; The present paper highlights that a sufficient concentration of copper(II) and ammonia are required in order to achieve appreciable gold leach rates. However, the use of air (oxygen) to maintain the copper(Il) concentration catalyses the undesirable copper(II) reaction with thiosulfate. The presence of ammonium ions is shown to have an effect on this undesirable reaction as well as on the gold leaching ?reaction. Preliminary electrochemical studies into .gold electrowinning using the REQCM show that gold(I) thiosulfate cannot be readily reduced to metallic gold in the potential region at which cementation usually occurs. However, it is well known that gold cementation onto copper occurs rapidly. The current study addresses this discrepancy by studying the effect that copper and silver have on the reduction of gold thiosulfate. It has been found that the presence of both silver and copper enhance gold deposition, particularly at low over potentials. While this is beneficial in cementation, the presence of copper and silver in solution has been found to cause passivation of the cathode surface during electrowinning at &apos;higher overpotentials.

Jan 1, 1998

A very wide range of technologies are utilised in the mining and processing of ores and a great deal of effort is expended in the development of ænew technologiesÆ to improve mining performance and economics. Traditionally, technology and people have been treated as separate aspects of an operation, when in fact they are inextricably linked. The successful implementation of new technologies is entirely dependent upon the actions, enthusiasm and skills of people, both as individuals and working in groups. This paper discusses the role of technology in mining and how new technologies are implemented by the industry. The paper argues that existing technologies have not been fully or adequately utilised and that effort should be re-focussed on the implementation of technologies that have been taken for granted for many years. The need for greater emphasis to be placed on the management of the human-technology interface is discussed.

Jan 1, 1998

By R. V. Ramani

"In this paper, the current status of respirable dust control in coal mine atmospheres is presented. The advancements in the engineering control of dust generation and entrainment, and of reducing the exposures are summarized. The results of the Health Surveillance Program are briefly reviewed. The results of research conducted in the Generic Mineral Technology Center for Respirable Dust are presented. The 1995 NIOSH criteria document on respirable coal mine dust is briefly reviewed. The fundamental scientific and engineering issues in need of research are discussed. A detailed bibliography is attached.IntroductionFor coal mines in the United States, the current permissible exposure limit (PEL) of 2mg/m3MRE-equivalent concentration (an 8-hour TWA concentration measured gravimetrically) was established by Public Law 91-173: the Federal Coal Mine Health and Safety of 1969 (the 1969 Coal Act). This standard is reduced when the respirable quartz content in the dust sample exceeds 5%; the reduced standard in mglm3 is obtained by dividing 10 by the percent quartz. The OSHA and ACGIB PEL for respirable coal mine dust is also 2 mg/m3 , and OSHA PEL for respirable quartz is 100 µg/m3 . The 1969 Coal Act also prescribed two activities in the Coal Mine Respirable Dust Sampling Program and the National Coal Workers' Health Surveillance Program n to control respirable dust in mines, and the coal worker's pneumoconiosis (CWP) among the coal miners.There are three principal methods by which personal exposures to respirable dust are controlled. The first is engineering control where the aim is to reduce the concentration in the mine atmosphere to or below the mandated level. The second is the provision of personal protective devices which allow the miner to work in dusty atmospheres without increasing personal exposure. The third is the use of administrative controls where the miner's exposure is controlled through work schedules and procedures to keep miner exposure at or under the mandated levels. The 1969 Coal Act specifically prohibits the use of respirators in lieu of environmental control. The compliance sampling program essentially does not give any credit to the use of administrative controls."

Jan 1, 1998

Kroemer(1-3) and Kroemer et al.(4) define the isoinertial technique of strength assessment as one in which mass properties of an object are held constant, as in lifting a given weight over a predetermined distance. Several strength assessment procedures possess this characteristic. Most commonly associated with the term is a specific test developed to provide a relatively quick assessment of a subject&apos;s maximal lifting capacity using a modified weight-lifting device.(1,5) The classic psychophysical methodology of assessing maximum acceptable weights of lift is also considered an isoinertial technique under this definition.(6) While the definition provided by Kroemer(1) and Kroemer et al.(4) has been most widely accepted in the literature, some have applied the term "isoinertial" to techniques that differ somewhat from the preceding definition, such as in a description of the Isotechnologies B-200 strength testing device.(7) Rather than lifting a constant mass the B-200 applies a constant force against which the subject performs an exertion. The isoinertial tests described here apply to situations in which the mass to be moved by a musculoskeletal effort is set to a constant. Is Isoinertial Testing Psychophysical or Is Psychophysical Testing Isoinertial? As various types of strength tests have evolved over the past few decades, some unfortunate developments in terminology have arisen to describe and classify different strength assessment procedures. This is particularly evident in sorting out various tests that have been labeled "isoinertial." One example was cited earlier. Another problem that has evolved is that the term "isoinertial strength" has two connotations. The first is the conceptual definition: isoinertial strength tests include any strength test in which a constant mass is handled. In practice, however, the term is often used to denote a specific strength test in which subjects&apos; maximal lifting capacity is determined using a machine and a constant mass is lifted.(1,5) Partially as a result of this dual connotation, the literature contains references to both "isoinertial strength test" as a psychophysical variant(8) and the psychophysical method as an "isoinertial strength test" (4,9) To lay the framework for the next two chapters, the authors will briefly discuss some operational definitions of tests of isoinertial and psychophysical strength. In stating that the isoinertial strength test is a variant of the psychophysical method, Ayoub and Mital(8) refer to the specific strength test developed by Kroemer(1) and McDaniel et al.(5) Clearly, this isoinertial protocol has many similarities to the psychophysical method: both are dynamic; weight is adjusted in both; and both measure the load a subject is willing to endure under specified circumstances. However, while both deal with lifting and adjusting loads, there are significant differences between the psychophysical (isoinertial) technique and the Kroemer-McDaniel (isoinertial) protocol in their procedures and the use of the data collected in these tests. For our purposes, we designate the

Jan 1, 1998

By P G. Hicks

The influence of feed size and pulp density on the specific rate ofbreakage of quartz was investigated. Plots of log fraction of feedremaining within the feed size interval against grinding time allowed the application of the first order grinding model to be examined. Substantial non-linearities were observed for all feed sizes due to environmental and material factors. For feed sizes above 1.3 mm, the specific rate of quartz breakage was higher with rod milling than ball milling, but below this size the reverse was true. Material that remained in the feed size interval after the initial grinding broke more slowly in subsequent grinding than the original feed material. Grinding media and time as well as feed size influenced the effects of pulp density on grinding. More dilute pulps (eg 50 per cent solids density by weight) were favoured for higher rates of breakage in rod milling but for ball milling, the optimum pulp density appeared to be 60 to 70 per cent solids by weight.

Jan 1, 1998

By M. Cai

Transformation of the measured borehole deformations induced by stress relief relies on the traditional assumption that the rock is continuous and homogeneous. Systematic and quantitative tests were carried out to check the performance of devices commonly used for measuring borehole deformations within different types of discontinuous rock conditions. Test results provide a reliable basis for considering the influence of discontinuities on in situ stress measurement and for choosing the best measuring device for a given rock type and condition.

Jan 1, 1998

The total cost of horizontal development can be minimised by optimising the cost of each of the component operations (eg drilling, blasting, etc). Component costs should not be minimised. To optimise blasting results (ie the mean advance per round, muckpile characteristics and overbreak), operators need to optimise the blast design and drilling and blasting procedures. The costs of pulling extremely long (say >6 m) rounds, producing unnecessarily fine fragmentation, and minimising overbreak are excessively high. Operators should continually consider how they can shorten the cycle period by saving time at the face. To enable it to be matched to the pertinent rock properties, the blast design needs to be based on as much relevant experience as possible. To develop the most valid correlation between blasting results and blast design, the latter must be implemented carefully and accurately, but without sacrificing speed excessively. The total cost can often be reduced by using: 1. computer-controlled jumbos (to reduce the period and cost of drilling and to increase drilling accuracy); 2. larger diameter (51 mm to 64 mm) and longer (4.5 m to 6 m) blastholes; 3. at least three large diameter ( ¦89 mm) relief holes that are 0.2 m to 0.3 m longer than the blastholes; 4. explosive products that are selected on the basis of proven reliability and overall cost efficiency rather than price; and5. stemming (to increase the amount of effective work done by each kilogram of explosive and, hence, to reduce the powder factor and possibly the number of blastholes and the costs of drilling and initiation). Non-perimeter blastholes should be charged with ANFO or a bulk emulsion of established performance. Upper perimeter blastholes should be charged with a bulk explosive with a much lower effective density. But where geology varies considerably, greatest control over overbreak could possibly be achieved with the versatile pre-assembled charge shown in Figure 1. This paper presents numerous potential ways in which the total cost of development can be minimised. It concentrates on recommending blast design features and the safe, rapid, cost-effective implementation of designs.

Jan 1, 1998