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By V. Mathebula

Paper written on project work carried out in partial fulfilment of B. Tech (Mining Engineering) degree The purpose of this study is to establish whether to mine the Bottom of Pit (BOP)/TOB 4 reserve before the commencement of the Block Cave or whether this should be disregarded and the reserve be extracted at the end of the life of the Cave. The primary goal of this report is to investigate and propose a practical mining method to safely extract the reserve given the set constraints. This report quantifies the additional revenue the TOB 4 will contribute to Finsch?s overall NPV thus enhancing Finsch?s business case and strategically enhancing the SBP. Several underground mining options have been investigated. The proposed mining method includes the installation of a series of extraction tunnels below the TOB 4 reserve from 53 level, with a slot access tunnel system. Ground will be loaded by LHDs from the draw points and tipped into 50-ton dump trucks at a transfer point, from there the ground will be transported to 53 level No. 1groundpass where it will join the existing ground-handling system. Primary development is scheduled to begin October 2002 and last for a period of approximately 7 months until early April 2003.The production is broken into 4 phases. Phase 1 is scheduled to begin mid-April 2003. The final production phase will be completed by November 2005. The production scheduling has been designed in order to remain ahead of the Block 4 undercut in order to ensure stability and safety. Due cognisance was taken of the health and safety of all employees concerned during the design of the mining and ground-handling processes for the Bottom of Pit. As far as is reasonably practicable, the BOP was designed and will be constructed and equipped to provide conditions for the safe operation of machinery and a healthy working environment for people underground. Relevant requirements of the Mine Health and Safety Act regulations pertaining to mining operations have been identified and adhered to during all phases of the design process.

Jan 1, 2004

By Clive Chadbourne

Investigations of world wide coal face strata control practice, the study of theoretical calculations and detailed engineering analysis have been combined to design and select a powered roof support suitable for use in conjunction with the extraction of 4.5-m (14.7- ft) of the 7.5-m (24.6-ft) thick Warwickshire Thick Coal Seam. The interface between geotechnical criteria and engineering design has been closely considered along with the following 3 important factors: Firstly, the need to further improve strata control and operational characteristics whilst producing a wall integrated overall face design. Secondly, to enable the supports to be rapidly transported from face to face in roadways driven at the minimum size for operational requirements. Finally, to provide hydraulic power, automatic control and monitoring systems which will improve operations reduce manpower requirements and improve management information. Clear design requirements were determined in order to allow selection on a commercial basis. A method of tendering is described with the subsequent analysis and selection process. In the event oversupply of coal in the UK resulted in the order not being placed. Brief consideration has been given to supports suitable for reserves to be mined In a few year's time where there is potential for extraction well in excess of 4.5-m (1 4.74.

Jan 1, 1994

By Michael R. King

The application and use of steel fibre reinforced concrete in precast tunnel lining design is a growing trend due to its advantages in performance over plain concrete, and ease of manufacture compared to traditionally reinforced concrete. However, the design process and performance characteristics of the material are poorly served by current codes and standards. This paper provides some basic design guidance for segmental linings, and compares the material properties for the different material options. A typical example is considered to compare the possible results of lining arrangement and thicknesses for a plain, reinforced, and fibre reinforced design solution.

Jan 1, 2005

In order to deliver a quality shotcrete and concrete product underground for the extensions to the Cullinan Diamond Mine operations it was decided to develop a new, simpler concrete product to meet underground support requirements. The aim was to develop a concrete product that not only reduced costs but also met the performance criteria of the current range of concrete products used on Cullinan Diamond Mine, a 25, 45 and 60 MPa concrete and a fibre reinforced wet shotcrete. The concrete product batching and delivery system was to be from surface as opposed to the current underground batching systems. The operational principle was to include an automated surface batching plant with delivery of the concrete product to a borehole containing suspended pipelines delivering to an underground storage facility from where the product is delivered to the working area by further pumping or agicars. Keywords: concrete, shotcrete, full flow, frictional loss, pipeline transport, positive displacment pump

Jan 1, 2006

A single aggregate mix design was developed based on a slag/Portland binder blend and crusher sand. The primary design criterion, after performance, was the fluidity of the shotcrete and concrete product to ensure compatibility with existing manufacturing and shotcrete equipment as well as the pipeline transport system Cullinan Diamond Mine was investigating. Initial work had determined an optimum slag/Portland to aggregate ratio while retaining a water to slag/Portland binder ratio of approximately 0.4 to maintain the flowability of the product. A super-plasticizer was used to further increase the flowability of the product as opposed to increasing the water content. Finally a polypropylene reinforcement fibre was included in the wet shotcrete product to improve the energy absorption characteristics of the shotcrete. Previous work on the strength development of the slag/Portland based shotcretes and concretes suggests that strength class based on the EN 206 specification is C40/50 while the toughness classification based on the EFNARC plate test is c, or greater than 1000 joules1. Typically, Cullinan Diamond Mine wanted a shotcrete with a uniaxial compressive cube strength of 45 MPa and an energy absorption, based on the EFNARC plate test, in excess of 700 joules. This is an EN 206 strength class of C36/45 requiring a uniaxial compressive core strength of 30.5 MPa and a toughness classification of b1. In addition to the shotcrete development, Cullinan Diamond Mine required concrete products with a compressive strength range of between 25 and 60 Mpa for differing construction operations underground to be manufactured from the same mix design and raw materials as for the shotcrete. Finally, Cullinan Diamond Mine carries out extensive dry shotcrete operations and an investigation into the possibility of using the single aggregate mix design for dry shotcreting was also entered into. This paper presents data that indicates that, under the terms of reference, a wet and dry shotcrete and concrete product appears to be possible from a single aggregate less than 6.7 mm in size, and that a slag/Portland binder is suitable for use as the hydraulic binder in all the products, and that the necessary performance parameters of strength and energy absorption can be achieved.

Jan 1, 2006

Coking coals are very much in demand on a world wide basis, particularly the higher rank coking coals, which form an important part of many coking blends and which represent the smallest portion of the world'scoking coal reserves. Conventional coke oven plants, using these coals, will predominate at least until the next century.Blending schemes for conventional coking require definition of the basic properties of coals so that the factors influencing the carbonization behaviour may be assessed. Several concepts, which are relatively simple, ate available to help promote a better understanding of the bask factors controlling -coke strength and are also useful in suggesting possible lines to follow in the blending of coals for coking.The interpretation of carbonization behaviour and coke structure in terms of petrographically defined coal rank and type provide a basis for planning and exploitation of coal Teserves. The optimum parameters for coking are suggested as follows.1. Type- Vitrinite content (% m.m.f. by volume) 55 ± 5%2. Rank- Mean maximum reflectance of vitrinite 1.30 ± O.10% Carbon content of vitrinite (% d.m.m.f.) 87·5 ± 1·0%The value of a coal should be assessed in relation to all its possible uses in blends. This is a more realistic and productive approach, as compared with considering coals only in relation to the results of blending with"good" coking coals. Blending schemes have been outlined to assist in the assessment of the blend potential of coals of all ranks and types.Reserves of higher rank coking coals in N.S.W. are very limited and should be conserved by blending with the low rank coking coals, of which there are large reserves.Examination of blending theories in relation to N.S.W. coals suggest that blends in the broad categories of minimal downgrading of the higher rank N.S.W. cok...

Jan 1, 1972

By Goldney L. H

The authors review the literature and discuss the application of methods to the design of thickeners for ilocculated metallurgical pulps. The use and evaluation of flocculents is described. Experimental data show that design based upon laboratory settlingdata is acceptable. The economic aspect of design is also discussed.INTRODUCTIONFor many years there has been little change in the methods for the design of thickener installations. While modifications have been made in the method of applying the data all methods have been based on the observee settling rate of the solids. The development of a theoretical approach based on physical data would involve so many variables which affect the settling rate that the more practical approach of direct observation seems unlikely to be replaced as the basis of design.The advent of synthetic fiocculents has introduced a modification in operation which affects design and should be considered. The authors review the subject and attempt to bring it up to date with present day procedure. Until recently the original work of Coe and Clevenger (1916) remained the basis of thickener design. They described their observation of the different zones encountered in settling, namely clarification, feed, transition, and compression or thickening- zones. The solids pass through the transition zone in settling from feed density to some higher density where settling ceases and ·some compression or compaction effect controls. the density obtained. The point at which the settling zone ceases to exist is known as the critical point.

Jan 1, 1959

Draft tube circulators are now widely used instead of Pachucas for suspension of slurries in large tall tanks.Precipitators in the alumina industry are an example.Draft tube/tank diameter ratio is critical. Two independent criteria are given for optimum ratio. The hydraulic and energy relationships are defined. These require an impeller of very high specific speed. For a stable and efficient system the impeller should have aerofoil blades with a pitch angle which decreases from hub to tip.The draft tube must be matched to the impeller with inlet baffles above the impeller and flow control vanes below it. Tip recirculation must be suppressed. The tank itself should have a filleted flat bottom.Such a system exhibits the behaviour of axial flow devices with the characteristic undesirable behaviour at low flow.Installed and operating power requirements depend critically on hydraulic design and to a lesser extent on speed reducer efficiency.The problems of re-suspension have been overcome by a patented slot system which also improves behaviour at low flow to permit a more economic design. A number of other practical operating considerations are also discussed.

Jan 1, 1982

By H. D. Stolovitch, M. L. Lavallée

"An engineering system used to analyze and improve the performance of industrial grinding circuits has been adapted into a training program. The systematic approach to the design of training was applied in order to:Determine if training was an appropriate and effective tool for the transfer of this new technology;Define the technical content of the program;Establish program specifications based on learner and context characteristics;Design and test prototypes of the instructional modules; and,Establish a strategy for plant implementation and maintenance of the training program.This paper thus describes how a engineering system is transferred to the workplace through the use of designed instruction.INTRODUCTIONTraining can change human performance in a permanent way. It can improve productivity for the overall benefit of an operation. Organizations can therefore use training as a mean to achieve their specific goals. Training is not the solution to every performance problem. However, in many cases, it can be a very effective way to attain desired performance.The benefits of training are often difficult to assess. However, to remain competitive, organizations must introduce new products and innovative technologies. Often, they can do both most effectively through suitable training interventions.In order for the effort and investment of training to provide optimal results, two points of view should be adopted.• An overall view of the system in which training is to be applied.• A systematic approach to the design of training (Stolovitch and Keeps, 1989)."

Jan 1, 1992

By John C. Bowling

Many recent mine developments in North America are located in remote northern regions with extreme winter temperatures in the range of -20oF to -60°F. In addition, many locations are subject to very heavy snowfall. This paper discusses some of the key considerations in the planning and design of mine plant buildings, particularly concentrators, located in these regions. SUMMARY & CONCLUSIONS The subject is considered under three main headings: I. Building Out line and Structural Frame 2. Wall and Roof Materials 3. Foundations The plan outline of a mine plant building will naturally be governed by the scope and layout of the process equipment it contains. The selection of roof profile, on the other hand, will generally depend upon a number of factors extraneous to the process e.g. snow conditions. When these factors are considered for concentrators in cold locations, a flat or stepped roof will usually be more economical and suitable than a sloping roof. The materials normally used for the structural framework of concentrators are the same as used for other industrial buildings, namely steel, concrete and wood. Structural steel, however, is the most commonly used material for concentrators in

Jan 1, 1971

By D. F. Malan

This paper gives an overview of the difficulties associated with determining the strength of hard-rock pillars. Although a number of pillar design tools are available, pillar collapses still occur. Recent examples of large-scale pillar collapses in South Africa suggest that these were caused by weak partings that traversed the pillars. Currently two different methods are used to determine the strength of pillars, namely, empirical equations derived from back analyses of failed and stable cases, and numerical modelling tools using appropriate failure criteria. The paper illustrates that both techniques have their limitations and additional work is required to obtain a better understanding of pillar strength. Empirical methods based on observations of pillar behaviour in a given geotechnical setting are popular and easy to use, but care should be exercised that the results are not inappropriately extrapolated beyond the environment in which they are established. An example is the Hedley and Grant formula (derived for the Canadian uranium mines), which has been used for many years in the South African platinum and chrome mines (albeit with some adaptation of the K-value). Very few collapses have been reported in South Africa for layouts designed using this formula, suggesting that in some cases it might yield estimates of pillar strength that are too conservative. As an alternative, some engineers strongly advocate the use of numerical techniques to determine pillar strength. A close examination unfortunately reveals that these techniques also rely on many assumptions. An area where numerical modelling is invaluable, however, is in determining pillar stresses accurately and for studying specific pillar failure mechanisms, such as the influence of weak partings on pillar strength. In conclusion, it appears that neither empirical techniques nor numerical modelling can be used solely to provide a solid basis for conducting pillar design. It is therefore recommended that both these techniques should be utilized to obtain the best possible insight into a given design problem. Owing to the uncertainties regarding pillar strength and loading stiffness, monitoring in trial mining sections and in established mining areas is also an essential tool to test the stability of pillar layouts in particular geotechnical areas.

Jan 1, 2011

By A. H. Swart

Unplanned stope panel collapses occur on most near-surface and shallow mines in South Africa. Although these incidents often occur during blasting, they pose a major threat to the safety of underground workers and the economic extraction of orebodies. Most stope panel collapses occur due to excessive panel lengths (resulting in excessive stope spans) that are not designed according to a systematic design procedure or methodology. Instead, panel lengths are often dictated by the equipment in use and by previous experience under similar conditions. Hence, a rock engineering design methodology for the design of stable stope panels between pillars is of vital importance for optimum safety and production in shallow mining operations. Using the proposed design methodology described in this paper, rock mechanics practitioners and mine planners should be able to identify and quantify the critical factors influencing the stability of stope panels. The critical factors should then be used as input to the design of stable stope panels that will provide the necessary safe environment for underground personnel working in stopes. The proposed design methodology is an iterative process, which includes aspects such as rock mass characterization, estimation of rock mass properties, identification of potential failure modes, consideration of appropriate stability analyses and other elements of the rock engineering design process. This process continues after implementation. During monitoring, new data should be included in the database and the design process repeated on a regular basis. Design approaches included in the proposed design methodology include empirical methods such as different rock mass classification methods, analytical methods such as elastic and Voussoir beam analyses, kinematic analyses, probabilistic analyses, and numerical analyses. Keywords: stability, stope panels, design methodology, hardrock mines, shallow mining operations

Jan 1, 2005

By P. J. Bethell

Background Increasingly complex geological conditions, with the presence of high ash binders in the mining horizon were encountered at the A.C. I. Dugout Mine near Price, Utah. This led to the production of high ash coal from the Dugout Mine during 2003 and 2004. This material had poor marketability and posed operating issues for the utilities burning the coal. Ash levels increased from typical levels in the 9-12% range to between 15 and 25%. etween 15 and 25%. Reviews of the mine plan and expected geological conditions indicated that further periods of high ash production could be expected. Action Plan Detailed sampling of the mine high ash production was carried out. The majority of the coal from the Dugout Mine is trucked to the Savage Coal Terminal (SCT) at Wellington, Utah where it is stored in multiple piles by quality. Coal at the mine is segregated by a CoalScan 2600, on-line analyzer into high ash, low ash and rock chute (waste material). From these piles, the material is crushed prior to shipment to SCT. The samples obtained from the various high ash piles were subjected to detailed sizing and washability analysis. Once this information was obtained, alternative circuit configurations were evaluated to process this material into a salable product. Water supplies in the arid terrain where the plant is located are scarce. This factor played a part in the final circuit design.

Jan 1, 2008

By Jim Dodd, Jim Paynter

"We would like to acknowledge the efforts of the following key people who helped to make this project a success.Ron Jung, instrumentation supervisor for the Golden Giant Mine, spent many hours reviewing process and instrument drawings, selecting instruments and providing hands on support during commissioning.Andy Keough, underground production supervisor for the Golden Giant Mine, was responsible for readying the underground distribution network for paste backfill.H.A. Simons Ltd. personnel who did an excellent job providing sound conceptual and detailed engineering.Fred Brackebusch, of Mine System Design who designed the paste backfill delivery systemPierre Lacombe, mill superintendent of the Louvicourt Mine, generously shared information and operating experience of their paste backfill plant.AbstractThroughout the 11 year history of the Golden Giant Mine, the primary backfilling method made use of consolidated rock fill. New backfill technologies have developed over recent years which now make it possible to use large quantities of fine mill tailings as backfill.In January of 1996, construction began on a new paste backfill plant at the Golden Giant Mine. Six months following, the plant was commissioned. This paper describes the design, commissioning and operational experience to date of the paste backfill plant."

Jan 1, 1997

The ore treatment expansion programme at Mount Isa Mines required the erection of a new lead-zinc concentrator to be known as No. 2 Concentrator. This concentrator was designed by Western-Knapp Engineering Company in conjunction with staff of Mount Isa Mines Limited. Metallurgical design followed conventional wet milling practice, but the size of many plant units is larger than previously used at Mount Isa. The. total capacity of the new plant is 150,000 tons per four weekly period, and the plant is designed to operate in two equal parallel sections.Erection was contracted on a turnkey basis, and the plant was commissioned as a production unit on 23rd June 1966.The existing concentrator, No. I, was converted to treat all copper ore, and the total capacity of the two plants is now 390,000 tons per four weekly period.

Jan 1, 1968

Why A Friction Hoist? B EFORE discussing the equipment, it would be instructive to review the factors that influenced the management of Opemiska Cop-per Mines to purchase a friction hoist rather than a conventional double drum hoist. The mine required that 250 tons per hour be hoisted from a depth of 2,100 feet. They also specified that the top speed should not exceed 1800 feet per minute. Eight ton capacity skips in balance and a maxi-mum hoisting speed of 1725 feet per minute would meet the required <Output on a duty cycle as follows: Loading time Creep away from loading pocket Accelerate (at 2.5 ft/sec2) 10 seconds 5 seconds 11.5 seconds Full speed run Decelerate (at 2.5 ft/sec2) Creep into dump Cycle time Trips per hour Tons per hour 72 seconds 11.5 seconds 5 seconds 115 seconds 31 31x8 = 248

Jan 1, 1962

This is a sampling of the papers presented at the conference, the proceedings can be bought from the SAIMM The technique reported here involves measuring the air pressures generated within sealed boreholes located behind blasts. Results from twelve free face blasts in competent rock environments a represented. There were no instances of high pressure gas penetration, even at distances corresponding to less than one blast burden. Rapidly generated underpressures were measured in most holes, with the magnitudes of the peak pressures decreasing with distance behind the blasts. It has been proposed that the under pressures are the result of new volume created due to crack formation and overall rock mass dilation. The experimental evidence is consistent with this, including a correspondence of the onset times of the peak underpressures with the nominal firing times of the nearest blastholes. New post-blast cracks were observed when using borehole video cameras in the monitoring holes. A calculation of new volume produced within the monitoring holes is proposed as an indicator of damage. While the data exhibit considerable scatter, there is a power function dependence on scaled distance, which resembles that of typical vibration data. The data also lie within the normally quoted vibration regimes for rock damage and are consistent with measured extents of damage. It is concluded that this technique may provide an inexpensive and practical comparative measure of near-field rock damage due to blasting. It is further concluded that gas penetration is not a common phenomenon. Damage appears to be mainly attributable to the nearest blastholes and the control of charge mass in perimeter blastholes is emphasized for damage minimization.

Jan 1, 2000

WHILE in the employ of the Broken Hill Associated Smelters and the Electrolytic Zinc Oompanies at their South Melbourne Research Station, the writer was detailed to investigate the principles underlying the Parkes Process with a view of determining:1. The Interdependence of the variables:(a) Quantity of zinc required per charge,(b) Temperature of stirring of each zincing,(c) Silver content of the bullion after first zinculg,(d) Silver content of the bullion after second zincing,(e) Silver grade of the crusts produced for retorting, with reference to the method of working the Parkes Process then in operation at Port Pirie.2. The possibility of effecting improvements in the Port Pirie practice of desilverising, especially with respect to:(a) Silver grade of the crusts produced for retorting.(b) Quantity of zinc required per charge.The method of desilverising in practice at Port Pirie was a two-stage zincing.In both zincings the bullion was cooled to its solidification temperature and the silver crusts which separated from the lead were completely removed.The crusts from the first zincing were retorted, and the crusts from the second zincing were returned to the first zincing of the next charge.

Jan 1, 1925

Gold was determined directly in eight different mill cyanide solutions, by atomic absorption spectroscopy in an acetylene-air flame. Some of the mill solutions were diluted with water to bring them into the most sensitive absorption region (5-50 p.p.m. Au) whereas those solutions with very low gold content &#171;2 p.p.m.) were concentrated by evaporation.Calcium was complexed with ethylene diamine tetra-acetic acid (EDTA) to avoid the undesirable effect of a precipitate on atomization.When added to standard gold solutions, EDTA had no effect on absorption, nor did any of the impurities common in mill cyanide solutions.Known amounts of gold, ranging from 2 to 30 p.p.m., were added to mill solution&apos;aliquots and recovered to within 100 &#177; 2&#183;5 per cent. For duplicate determinations under normal worKing conditions, standard deviation was 1&#183;32 p.p.m.Excellent agreement was found between atomic absorption gold values and accurate fire assay values.INTRODUCTIONAtomic absorption spectroscopy has been shown to offer a rapid and accurate method for the determination of trace amounts of a number of metal ions. In the field of primary metallurgy atomic adsorption has been used for routine determinations of lead, copper, silver, zinc, cadmium and iron in a variety of raw materials and intermediate products.The scope of atomic absorption spectroscopy was reviewed recently by Willis (1963).The noble metals silver, gold, platinum, rhodium and palladium were determined by atomic absorption from chloride solutions by Lockyer &#183;and Hames (1959). In their determination of gold down to 1 p.p.m., they found no interference from the noble metals, nor from considerable excess...

Jan 1, 1964

An assessment of methods currently available for the accurate determination of low concentrations of platinum in ores and mineral concentrates indicated that prior treatment of the sample by fire assay procedures is necessary. To enable rapid determinations to be carried out, instrumental methods applied to the metallic bead produced from the fire assay treatment have been evaluated, considering maximum flexibility for concentration variations, minimum chemical operations, and adequate sensitivity.Atomic absorption spectrometry proved suitable, giving solution determination limits of 0&#183;4 p.p.m. Pt, 0&#183;1 p.p.m. Pt, or 0&#183;06 p.p.m. Pt under various extraction conditions using aqueous and organic agents. Thesedetermination limits related to 1 assay ton of sample (32&#183;67 g) produced easily readable and accurate readings within the range of platinum contents (0&#183;03-2&#183;0 p.p.m.) of economic interest in deposit evaluation and prospecting. The technique shows promise of also being suitable for determination of palladium and gold.

Jan 1, 1970