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By P. E. J. Pitfield, S. D. G. Campbell

The regional shear zone runs along or near the southern margin of the Chinhoyi greenstone belt. It hosts four of the seven mines in the southern part of the belt with recorded production above 100 kg Au, including the top two current producers. Displacement occurred in the late Archaean, producing a shear zone with a trend ranging from northeast at Golden Kopje to east at Muriel. The orebodies at the two mines formed synchronously with the shearing and show strong but contrasting structural controls on mineralisation. The Golden Kopje orebodies plunge to the southwest, parallel to the intersection of curved foliation with discrete shear planes and sub-perpendicular to local stretching lineations. The Muriel orebodies plunge steeply west but parallel to local stretching lineations. The greatest fluid flow during mineralisation of the shear zone was along either the intersection of fabrics (Golden Kopje) or the extension direction (Muriel). There is no unique relation between the orebodies and the extension or shear direction. Where earlier mining has not established the directions of the oreshoots it is essential to understand the possible structurally controlled geometries of orebodies before planning a drilling programme to evaluate their potential

Apr 1, 1996

By Rixon K. L

Hargraves (1983) has reviewed the phenomenon of instantaneous outbursts of coal and gas with special attention to Australian occurrences. The purpose of this discussion is to highlight some problems and omissions in his review that reflect particularly on the authors' structural geological investigations of outburst-prone coal seams.It is surprising to find that much of the authors' reported work has either been glossed over or the results distorted by Hargraves (1983) and it is, therefore, important to clarify the contribution that structural analysis of coal seams can make.For many years it has been known that faults provided preferred loci for gas outbursts (for example, Briggs 1920-21) but the roie that faults played in outburstprone seams was unclear. The present authors reviewed the association of faults and outbursts drawing upon local investigations and the world-wide literature (Shepherd et aI., 1981), and discussed some wideranging aspects that result from the high coincidence of faults and outburst sites. Without repeating large segments of this article the authors will discuss four aspects of coal seam faulting and fracturing.

Jan 1, 1983

Mr Kapp's paper helps considerably to clarify the subsidence/mine geometry relationships in New South Wales. It is in the best tradition of the presentation of field investigations such as has built up internationalknowledge over the years.The Author has plotted in Fig. 10, the maximum subsidence developed over each of a number of panels of different width/depth ratios and concludes that the maximum possible subsidence over a super critical extractionarea is around 0.65 of the mining height and that this is reached with the critical width/depth ratio of extraction of around 1.5.Although, as Mr. Kapp says, the maximum possible subsidence in the UK can reach 90% of the seam thickness extracted the more usual figure for the first seam is up to 80%. It has also been found in the UK that with shallow workings such as those at Kemiro Coliery the critical width of extraction is 2.0 or more times the depth.

Jan 1, 1981

The recent paper by Bain (1981) contains an unfortunate, and potentially misleading typographic error. On page 9, second paragraph, Bain states that, "some samples of drill core from the (Candlow) Formation ... contain up to 90 ppm gold and 2 000 ppm arsenic". The gold value should read 90 ppb. In fact, in the original version of the paper (Bain, 1980), the gold value quoted was actually 60 ppb.Bain also refers briefly to some geochemical data which is attributed to me, and which I feel require some clarification.In addition, the paper contains some factual errors relating to the occurrence of carbonate and potential host rocks to Carlin-type gold deposits in the Candlow Formation.

Jan 1, 1981

It is well known that the electrochemical precipitation of metals from aqueous solutions is conducted by the electrochemical standard potential scale. This scale indicates the order of nobility of different metals. Consequently, the standard potential scale shows the nocivity of metals when they are present as impurities in aqueous solutions containing metals. In the hydrometallurgy of cobalt, nickel and zinc are co-deposited during cobalt electrodeposition. Nickel has almost the same chemical, physical and electro-chemical properties as cobalt. Furthermore, the electrochemical potential of nickel is relatively higher than that of cobalt. For this reason nickel is preferentially reduced at the cathode in the presence of cobalt. The presence of small amounts of zinc in solution leads to contamination of cobalt electrowon. In spite of its lower electrochemical potential in comparison to cobalt, zinc is reduced on the cathode because of its higher hydrogen-overvoltage. But, in the hydrometallurgy of zinc, small amounts of cobalt present solution will contaminate the zinc reduced at the cathode. This is easily explained by the fact that cobalt is more electropositive than zinc. The co-deposition mechanism of sulphur at the cathode is not yet well understood. Great attention has to be paid to the mechanism of co-deposition of sulphur during cobalt electrolysis. This work aimed to investigate the influence of some parameters on the co-deposition of nickel, zinc and sulphur during cobalt electrowinning. The following parameters have been investigated: Co2+concentration in the electrolyte, current density and ammonium sulphate (NH4)2SO4additions. It has been established that important nickel, zinc and sulphur co-deposition takes place at relatively low Co2+ concentration in solution. The behaviour of these elements with regard to the cathodic phenomena has been established.

Jan 1, 2005

"Quang Ninh province is located in the north east of Vietnam and the coal mining in the province generates two thirds of the province’s GDP. The coal is extracted from both open pit and underground mines for the country’s complete coal export and by a share of more than 90% for domestic coal usage. Overburden generated through open pit mines is piled up nearby in the form of large waste rock dumps. The side-hill fill dumping near open pit mines is the source of environmental degradation in the region. Among the current problems at solid waste rock dump sites are dust emissions during and after dumping as well as Acid Rock Drainage (ARD). One of the major existing problems at waste rock dumps is dump stability (comprising erosion, sliding and subsidence issues). The RAME R&D project (Research Association Mining and Environment in Vietnam) is - among other activities - actively involved in the research for stabilisation and rehabilitation of mine waste rock dumps in Vietnam in the form of several sub-projects. The Institute of Mining Engineering I of RWTH Aachen University is leading the sub-project of mine waste dump stabilisation in co-operation with Chair of Geotechnical Engineering and Brenk Systemplanung GmbH. Mine waste dumping in the form of layers as an alternative to the side-hill fill dumping is expected to bring up environmental and economical benefits by minimising the problems of dump sites with the contribution to a sustainable development. Reduction of the post construction settlements, long term stable dump profile with minimum erosion, improvement of material parameters for geotechnical slope stability through compaction from hauling trucks as well as economical evaluation of new dumping concepts are the areas of the current research. The project has been financed by German Federal Ministry for Research and Education (BMBF). The ultimate aim of the sub-project is the compilation of the project results in the form of a "Best Practice Guide" and software based "Decision Support System" for planning and management of solid mine waste dumps."

Jan 1, 2011

Operating well designed and maintained surface mine haul roads is the key to minimising truck haulage costs - in itself a significant component of the total cost per tonne mined. Existing and proposed mining operations are subject to scrutiny, both from the economic and environmental perspectives. In the long run, this scrutiny returns improved efficiencies and leaner, 'greener' operations. The focus of this evolving evaluation process should and will certainly fall on haulage operations - simply by virtue of their contribution to overall cost of operations and sustainability impacts. Whilst the end result - improved efficiency and contributions to resource sustainability - is not in itself problematic, it is the route, or process followed to achieve these improvements that needs to be carefully managed. We can be guided on this journey by our understanding of how a road is designed, and, critically, the technological solutions that can contribute to a company's sustainability initiatives. This paper examines the current state of mine road design and operation, together with the economic and environmental issues associated with under-performance. The paper serves as a basis for evaluating the economic and environmental contributions that technology in road design and management can deliver to sustainable operations. Utilising innovation is the key to long-term sustainability and as a means of positioning the industry for the imminent structural changes to the economy and environment, changes which are now within the operating life of many mines, haul roads and haul trucks.

Jun 1, 2010

By Martial P. Corriveau

The preparation of coal is essentially the same for export as it is for the domestic market. The big difference, I would say, is that, while coal for certain domestic uses (utilities, for instance) need not always be of the highest quality, as long as it is uniform, the buyer of coal for export invariably demands a well-prepared, top-quality product. And he is willing to pay for it, as long as he knows what he is buying. This is where quality control enters the picture. The day is long past since the export market was the dumping ground for coals which could not be sold on the domestic market. Today, the export buyer wants American coals not so much for a price but for the necessity of blending the American coals with foreign coals to achieve the desired chemical and/or physical characteristics, To blend intelligently, he must know in detail what are the qualities of the coals to be blended. Sampling and chemical analysis facilities at the piers provide the information on the coals exported, These facilities are maintained for the export buyer by independent laboratories; however, in one instance, the buyer operates his own laboratory. With very few exceptions, and Clinchfield?s Moss No. 3 mid-volatile coal is one, cargoes shipped overseas are loaded as blends of two or more coals. Up to now, foreign requirements have not been too clearly defined; therefore, shippers must be prepared to blend mixtures of the desired combination. Successful blending requires an accurate knowledge of the coals shipped. And this makes it necessary for the coal producer to maintain his own control laboratory.

Jan 1, 1959

By S Jackson, I Arbuthnot

The use of High Rate Thickeners has gained wide acceptance in the mineral treatment industries. The design of these units is characterized by an efficient deaeration and flocculation chamber, and introduction of the flocculated feed solids into a preformed floc bed zone. By these means the settling process is accelerated and consequently High Rate Thickeners achieve a significantly higher throughput per unit area than conventional thickeners.   The object of a control system on a High Rate Thickener (HRT), as on a conventional thickener, is to obtain the required underflow density and overflow clarity at all feed conditions. The feed conditions which may vary include volumetric and mass feed rate, pulp density, particle size and particle type. A conventional thickener has a large volume and cross sectional area, and changes in performance occur slowly û the operator has time to observe and react to these changes to maintain control. On a HRT, there is a much reduced residence time and therefore control parameters need to be reset more frequently. Although manual control is possible, the use of automatic control systems is therefore desirable to avoid excessive operator attention, while optimizing control of density and clarity. Automatic control of a conventional thickener is usually achieved by an underflow density loop which regulates the rate of withdrawal of underflow. A turbidity meter is sometimes used to measure overflow clarity and regulate the addition of a coagulant or polymer flocculant. These measures are not adequate on a HRT, because they would not give optimum control and could lead to excessive flocculant use. To achieve optimum operation in a HRT, the floc bed level should be maintained above the feed inlet point in the thickener. This bed level has become the key variable around which most control systems have been designed. However, a number of problems have occurred which has resulted in poor control, or only partial control of thickener performance. We will describe some of these problems, and a solution which is simple and effective.

Jan 1, 1990

By W. C. Hellyer

In an autogenous grinding circuit in which feed at approximately nine inches top size is reduced to a size suitable for subsequent processing, the build- up of a "critical size" fraction in the media causes problems. A "critical size" fraction has been defined as, "media too small to effect reduction by impact grinding of ore coarser than a quarter of an inch and too large to be broken by the largest size of media in the charge. 9, (1) The buildup of a "critical size" fraction reduces the capacity of the mill, increases the grinding power requirements per ton of finished product, and generally produces a finer grind than is desired. It is general practice to overcome this problem by the addition of large diameter steel balls to the grinding charge. This has certain disadvantages, such as; (i) an increase in mill liner wear; (ii) wear on the steel balls, and; (iii) some loss in flexibility in grinding circuit operations resulting from difficulties in removing the steel balls by means other than by grinding out.

Jan 1, 1970

By André Pomerleau, Michel Grégoire, René del Villar

"The operation of industrial flotation columns requires the control of at least two variables, the interface position and the bias rate, by manipulation of some appropriate operating variables. Problems arise due to the reliability of existing methods of measuring the bias (i.e bias = tailings water - feed water), a situation which has often forced the industry to disregard this control loop. Moreover, when using such a measuring approach, the identification of the process dynamics is impossible. A second problem arises from the possible interaction between both control loops that might call for the use of a more complex multivariable control strategy.Recent work done at Lava1 University has demonstrated the feasibility of an independent sensor for bias, which models the relation between the conductivity profile across the interface and the bias value using a neural network algorithm. A 250 cm height, 5.25 cm diameter Plexiglas laboratory column was equipped with a series of conductivity electrodes in its uppermost part (across the interface) to measure both interface position and bias rate. Using such equipment, the flotation column dynamics were identified. The results thus obtained permitted the design and implementation of a distributed PI control strategy, where bias was associated to wash water rate and froth depth to tails rate. Both PI controllers were tuned using a frequency-response tuning method. Results of both identification and process control are presented and discussed.INTRODUCTIONFlotation columns are used for the separation of valuable minerals from a carrier gangue. Typically, this device is a tall (9 to 14 m height) cylinder (1 to 3 m in diameter) although other cross sectional shapes are also used. Finely ground ore is fed as a pulp at about 2 m from the top of the column, and pressurized air is introduced through a sparger at the bottom. If the pulp has been properly conditioned with adequate chemical reagents (collectors and modifiers), the valuable mineral particles falling by gravity will attach themselves to the rising air bubbles, and the air-particle aggregates will rise to the top of the column where they overflow as a stream called the concentrate. The gangue particles,, that do not attach to the air bubbles, fall down the column being withdrawn at its bottom in a stream called the tails.A particular characteristic of flotation columns is the addition of a fine spray of water at the top. As the water flows down through the froth zone, it washes out the entrained gangue particles from the mineralized bubbles, thus improving the quality of the concentrate recovered at the top. As a result of its cleaning action, this water stream is called the wash water and the upper zone, the cleaning zone. In order to ensure a downward flow of wash water, a proper water balance must be kept in the froth zone, otherwise the water added at the top might shortcircuit to the concentrate without performing any cleaning action. This net downward flow of water inside the froth zone is called the bias and represents an important operating variable of a flotation column."

Jan 1, 1999

By C. J. Brown

Controlling the level of impurities in copper refinery electrolytes is essential to achieving acceptable cathode purity. Purge of electrolyte is not an entirely satisfactory solution since the cost of neutralization, waste disposal, and replacement acid can be appreciable. The APU® process utilizes ion exchange resin to sorb sulphuric acid while rejecting metal salt impurities. Extensively employed for recovery of waste pickling and anodizing acids since 1977, the system is now being used at several refineries for removal of contaminants such as antimony, bismuth, iron, and nickel. The purified acid, which is recovered from the resin by water elution, is recycled back to the refinery tank-house, while the metals are typically recovered from the de-acidified by-product solution by precipitation.

Jan 1, 2005

In This Chapter [Finding the dust source Ventilation and dust collector malfunctions Upgrading the dust controls Design stage ventilation planning ] This chapter explains how to reduce respirable dust76 in hard-rock tunnels during excavation by using tunnel boring machines (TBMs). The first steps in combating a dust problem are to take dust samples to pinpoint the source, check the ventilation system, and check the dust collector. If the ventilation system and dust collector are operating properly, then other dust controls such as water sprays and conveyor belt scrapers must be upgraded. For tunnels in the design stage, recommended air quantities are provided. FINDING THE DUST SOURCE AND LOOKING FOR VENTILATION MALFUNCTIONS [The first steps in fighting a dust problem are to take dust samples to pinpoint the source, check the ventilation system, and check the dust collector. Without knowing the exact source, efforts to reduce dust are hit-and-miss (mostly miss).] Taking samples to pinpoint the dust source. In tunnels with high levels of airborne dust, the first task is to pinpoint where the dust enters the airstream. Most dust originates from rock breakage at the tunnel face, but the location where this dust enters the airstream can vary. Dust can leak from behind the TBM face shield, from gaps in the ventilation duct, or from a mal-functioning dust collector. It can be entrained into the air from the muck on a moving conveyor belt. It can even be shaken loose from the underside of the belt as it passes over the idlers. As a start, to locate the dust source, dust samples and air quantity measurements should be taken at the following locations:

Jan 6, 2003

By Hideki Shimando

Periodic weighting and windblast which are caused by the dynamic effects of overburden in a longwall panel significantly affect longwall mining This study describes past experience with wind blast at Miike Colliery, proposes a new loosening system using hydraulic fracturing and moistening, and also discusses strength deterioration in hard-to-collapse roof rocks due to water in order to solve the problems caused by the effects of dynamic roof rocks behavior in a longwall face

Jan 1, 1998

By M. H. Pettibone

During the past decade, the size of grin- ding mills has steadily increased. Mills 11 m in diameter with 9 mW motors have been installed with even larger units on the drawing board. The rising costs of energy and labor have spurred interest in improving efficiency while reducing capital investment and operating labor. Autogenous and semiautogenous grinding have experienced renewed interest as efforts to reduce operating and capital costs have increased.

Jan 1, 1981

By Fred N. Kissell

In This Chapter [Measuring the gas emission from the coal Methane at the top of the silo Methane at the load-out area and Actions taken after a silo explosion] Methane accumulations in coal silos have resulted in the occasional silo explosion. These can be quite violent and dangerous because coal dust adds to the strength of the blast. However, with the appropriate precautionary measures, methane accumulations in silos can be greatly reduced. Mine Safety and Health Administration (MSHA) regulations at 30 CFR2 77.201 require that the methane content in the air of any coal silo be maintained below 1.0 vol %.3 Also, MSHA requires that methane tests be conducted before any electrical equipment is energized, unless a continuous monitor capable of deenergizing the electrical equipment is used.4 Measuring the gas emission from the coal. The first necessary step in dealing with silo methane issues is to measure the gas emission from the coal going into the silo. Such measure-ments allow one to estimate the silo ventilation needs and permit a comparison with the methane controls used at other mines that have similar gas levels. The gassiness of the coal can be measured by taking conveyor belt grab samples.5 Matta et al. [1978] measured the gas emission from conveyor belt grab samples using a simple desorption test. To conduct the test, they collected several grab samples of coal, weighing a few pounds each, from the conveyor belt entering the silo.6 They then sealed the coal into an airtight sample container that was equipped with a valve and short hose along with a pressure gauge. Every few hours they opened the valve and bled the emitted gas into a water-filled graduated cylinder that had been inverted and placed in a pan of water (Figure 11–1). The results are shown in Figure 11–2.

Jan 6, 2006

By D Gelling

Mineral sands dry plants consist of a large number of repetitive unit operations with significant transport delays, in the form of stream pipes, conveyors and elevators. As part of the P255 mineral sands sponsorship and using the JKMRC developed conductor/non-conductor (CNC) gauges, control loops have been implemented at the multi-duty dry plant of Cable Sands Limited at Bunbury. This paper describes the loops for kV control on a rougher high tension roll machines (HTR). The options available are product grade control and ratio control using the JKMRC CNC gauges, which measure the percent conductors in a stream. The purpose of control is to stabilise the circuit operation against disturbances such as atmospheric and feed disturbances.

Jan 1, 2001

One of the problems associated with the maintenance of roadways in underground coal mines is that of dealing with roadway closure. Sidewall weakening technique is based on the transference of the high stress zone further away from the vicinity of roadway. This transferral is carried out by weakening parts of sidewalls of roadway by cut- ting slot, drilling a series of parallel holes or blasting during development stage. In the paper, the effect of sidewall weakening on the stability of roadway is studied by finite element analysis and field experiment. The finite element analysis shows that sidewall weakening can control the development of yield zone around a roadway, improving the stability of roadway. It is also shown that if in- situ horizontal stress is lower or higher than the in-situ vertical stress, it is difficult to obtain roadway stability by sidewall weakening. The field experiment shows that roadway closure can be controlled by sidewall weakening with blasting.

Jan 1, 1990

By Claude Bazin, François Robichaud

"The control of the operation of semi autogeneous grinding (SAG) mills is mainly achieved through adjustments of feed rate, water addition and speed of rotation. With the introduction of image based sensors it is possible to have on-line measurements of the size distribution of the ore feeding a SAG mill. This information could be used for the development of a control strategy to regulate a SAG mill feed size distribution to a set-point. The control of the size distribution is carried out by adjustment of the speed of the ore extractors underneath the ore stockpile. The potential and difficulty associated with the implementation of such control are assessed using a multivariable statistical analysis of data collected from the operation of the grinding circuit of the Agnico-Eagle Goldex mine in Val d’Or, QC. A control strategy for the ore feed size distribution is assessed using a simulator of the ore stockpile operation.INTRODUCTIONThe popularity of autogeneous (FAG) and semi autogeneous (SAG) grinding mills has been increasing over the last 20 years (Napier Munn et al., 1996). These mills use the ore or a combination of ore and steel balls as the grinding media. Since the ore is used as a media for the size reduction the mill ore feed size distribution impacts the performance of the mill. Several papers were recently published on this subject (Dance et al.,2011; Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011) showing the advantages of preparing the feed size distribution to maximize the circuit throughput or mill efficiency (Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011). With the availability of particle size sensor based on image analysis of the ore on a conveyor (Partanen et al., 2001; Bobo et al. 2004; Edwards et al., 2002; Wei & Craig, 2009) it is possible to control the size distribution of the ore feeding autogeneous or semi autogeneous mills (Partanen et al., 2001).This paper describes a method to control a size distribution index to a set point via the manipulation of the speed of the extractors that withdraw the ore from the stockpile feeding the grinding circuit. The paper consists of 4 sections. The first section describes the considered grinding circuit. A statistical analysis of the collected data on the grinding circuit is presented in the second section. This analysis demonstrates the necessity of using real-time adaptation mechanisms to develop a control strategy for the size distribution of a grinding circuit. The third section describes the method used to simulate the ore stockpile operation and the last section describes the preliminary development and assessment of a control strategy of size distribution of the ore feeding a SAG mill."

Jan 1, 2013

By A. McGarr

The shear stress t that can be sustained by the rock mass in the environs of a mining-induced earthquake controls the near-fault peak ground velocity v of that event according to v=0.25(ß/G)t, where ß is the shear wave speed and G is the modulus of rigidity. To estimate t at mining depths, I review the results of four studies involving Witwatersrand tremors that relate to the bulk shear strength. The first and most general analysis uses the common assumptions that the seismogenic crust is pervasively faulted, has hydrostatic pore pressure before mining, and an extensional stress state that is close to failure. Mining operations reduce the pore pressure to zero within the mine and redistribute the stresses such that, in localized regions, the state of stress is again at the point of failure. Laboratory friction experiments can be used to estimate t in the zero-pore-pressure regime. Second, model calculations of states of stress in the vicinity of mining at about 3 km depth indicated the shear stress available to cause faulting near the centre of a distribution of induced earthquakes. Third, laboratory experiments combined with microscopic analyses of fault gouge from the rupture zone of a mining-induced event provided an estimate of the average shear stress acting on the fault to cause this earthquake at a depth of 2 km. Fourth, moment tensors determined for mining-induced earthquakes usually show substantial implosive components, from which it is straightforward to estimate t. These four different analyses yield estimates of t that fall in the range 30to 61 MPa which implies that near-fault particle velocities could be as high as about 1.5 m/s. To the extent that the causative fault ruptures previously intact rock, both t and v, in localized regions, could be several times higher than 61 MPa and 1.5 m/s.

Jan 1, 2002