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By H. F. Fuls

Skorpion Zinc, an integrated zinc mine and refinery located in the south of Namibia, has a production capacity of 150 000 t/a of special high grade (SHG) zinc. The Skorpion zinc oxide resource will be depleted by 2017. Extensive exploration drilling for additional zinc oxides was conducted without success. With the abundant availability of zinc sulphide concentrates regionally (Black Mountain and Gamsberg in the Northern Cape or Rosh Pinah Zinc and new deposits in the Rosh Pinah region) the life of the operation can be extended by processing of zinc sulphide concentrates. The leach kinetics of zinc sulphides is vastly slower than for oxides, preventing the processing of zinc sulphides in the existing process under current conditions, and Skorpion Zinc faces the challenge to identify a suitable zinc sulphide treatment process that can be integrated with the existing plant. Various zinc sulphide processing options were identified through an extensive literature search and non-viable options were eliminated. Mass and energy balances were established for potential options and an economic evaluation (including capital and operating cost estimation) was conducted to identify a suitable process. It is concluded that ferric leaching of sphalerite in a sulphate medium under atmospheric or pressure leach conditions are suitable processes for Skorpion Zinc. These processes present advantages over all other zinc processing options when elemental sulphur is preferred as the final deportment for sulphide sulphur. Both these processes can be integrated into the existing refinery and will be able to process concentrates with a wide composition range.

Jan 1, 2013

By J. Githiria, C. Musingwini, J. Muriuki

"The maximization of net present value (NPV) is a primary objective in open pit mine planning processes. In an attempt to meet this objective, cut-off grades are considered in all the stages of mining. There are three stages involved in resource extraction, namely mining, processing, and refining/marketing, which are all defined in Lane’s approach. Using Lane’s approach, the economics involved in each stage are identified independently and interact to provide an optimum cut-off grade.A hypothetical block model is used to illustrate how a computer application (Cut-off Grade Optimiser) that was developed in this study is used in the optimization of cut-off grades using Lane’s algorithm. This paper analyses the application of Lane’s approach for a single element in cut-off grade optimization as applied in the calculation of the optimum cut-off grade through a computer-aided application.Although Lane’s approach is not complex it is not widely applied to maximize the NPV of mining operations as the iterative calculations can be lengthy. The computer application developed in this study shows how Lane’s algorithm combined with a linear programming function can optimize cut-off grade with regard to the complex situations faced by mining operations. The application saves users from performing otherwise lengthy, iterative calculations.Background and previous researchAn open pit mining complex constitutes multiple pits, ore processing streams, and metal refining options. An open pit mine produces ore (valuable) and waste material. Ore is transported to an appropriate processing stream for producing concentrate. A refinery processes the concentrate to produce marketable metal product (Lane, 1964, 1988).Cut-off grade optimization is based on Lane’s theory, where a mining system consists of three major stages such as mining, milling, and refining. In this approach the cut-off grade selection is done between six cut-off grades that involve all stages of a mining operation. Selection of the optimum cut-off grade influences the net present value (NPV) by maximizing the net cash flows from the operation."

Jan 1, 2016

[Consider a square block W in an ore body S2, to be valued from one centre sample w, and one corner sample w2 (Fig. 9.1). The sample set is ws = Iw,;w21. We can estimate the block average µw by the average of the sample values x, and x2: µw = 'z (x, + xz). (9.1) The error made is measured by the estimation variance: o' (ws to W) = 2 - (ws; W) - - (ws;ws) - - (W; W). (9.2) However, we know that the centre sample w, gives more information about the block value than the corner sample w2. In the example studied in § 8.4, where w, and w2 were considered to be points, we found that of (w, to W) was nearly one third smaller than o2E (w2 to W), therefore when estimating W one should give more weight to w, than to w2. In general, when evaluating a block W from a set of n samples ws, where WS = l w;w2; ... wn1, a better estimator of µw can be obtained if we give different weights to the sample values, depending on the size and the positions of the samples. Some desirable properties of such an estimator µw are: ? It should be a linear function of the sample values xi: n µw = bi xi (9.3) =t where bi = weight given to sample x . ? It should be unbiased, i.e. its expected value should be equal to the true block value: E [µw - µw] = 0. (9.4) ? The mean squared error of estimation of µw should be a minimum E [(µw - µw)2] = a minimum. (9.5) The estimator which satisfies these properties is a minimum variance, linear and unbiased estimator, and is known as the kriging estimator. The corresponding error of estimation is the kriging error. This estimator is also known as the BLUE (for Best Linear Unbiased Estimator) estimator. This was named by Matheron in acknowledgement of the early work of Krige's in the optimization of the evaluation of the South African gold mines. The following notations will be used: µK kriging estimator. o2k = o2K (w,, to W) kriging error.]

Jan 1, 1978

By Y. -L. Chiang

Paper written on project work carried out in partial fulfilment of BSc (Eng) (Chemical) The generation and control of manganese sulphide particle size distribution using various mixing configurations were investigated. Specifically, this paper discusses how varying mixing intensities on a macro and micro scale affect the resultant manganese sulphide particle size distribution (PSD) at a constant concentration. The mixing variations were achieved using various Y- and T-mixers as premixing devices into a jacketed, agitated vessel. The reagents used to generate the seeds were aqueous manganese sulphate and sodium sulphide solutions. The results showed that the absence of micromixing resulted in very poor control of the particle size distribution, the particle size and the number of particles produced. This was due to the fact that the macromixing time is much greater than the corresponding precipitation time, resulting in different mixing regions existing within the reactor. From the micromixing experiments, the T?mixer was found to be the most effective mixing device at lower concentrations due to the T-mixer providing a faster mixing time than the other mixing configurations, thereby effectively localizing the supersaturation and hence controlling the precipitation. The results obtained from the direct addition without the extension pipe (i.e. short mixing time) indicated a general increase in mean particle size and decrease in total particle number at reagent concentration of 0.007 mol.dm-3as compared to the corresponding micromixing experiments. The phenomenon may be due to incomplete crystallization at the immediate exit of the mixing device. Further experiments need to be conducted before any conclusions can be drawn about the precipitation mechanisms.

Jan 1, 2007

By J. C. van Dyk, F. B. Waanders, J. Brand, C. A. Strydom

"Underground coal gasification (UCG) is a fast-emerging, in situ mining technology that provides access to low-cost energy through the utilization of coal reserves that are currently not technically or economically exploitable by conventional mining methods.Groundwater monitoring for conventional coal mining in South Africa is well established, with SANS, ASTM, and ISO standards for the specific environment, location, and purposes. South Africa’s groundwater is a critical resource that provides environmental benefits and contributes to the well-being of the citizens and economic growth. Groundwater supplies the drinking water needs of large portions of the population, and in some rural areas it represents the only source of water for domestic use. Implementation of, and adherence to, groundwater monitoring standards are thus non-negotiable.The groundwater quality management mission, according to the Department of Water and Sanitation in South Africa, is set in the context of the water resources mission and reads as follows:‘To manage groundwater quality in an integrated and sustainable manner within the context of the NationalWater Resource Strategy and thereby to provide anadequate level of protection to groundwater resources and secure the supply of water of acceptable quality.’ (SABS, 2016).In this paper we propose fit-for-purpose groundwater monitoring standards for a commercial UCG operation. It is important to proactively prevent or minimize potential impacts on groundwater through long-term protection and monitoring plans. Scope of study / standard proposal The purpose of this paper is to set out a National Standard Proposal to emphasize groundwater monitoring as one of the five strategies of the ‘Policy and Strategy for Groundwater Quality Management in South Africa’ (Department of Water Affairs and Forestry, 2000) and to align it with the National Water Act, Act no. 36 of 1998 (Republic of South Africa, 1998).The focus of this standard is directed solely towards groundwater monitoring during underground coal gasification (UCG). Groundwater in this context refers to water as sampled from dedicated monitoring wells around the targeted site, which will include the shallow aquifer referred to as groundwater, and water flow at the level of the underground gasifier, referred to as ‘coal water’."

Oct 1, 2018

By H. Bhikha

The minerals industry is committed to the principles of sustain-ability. Reducing water consumption is a priority area, especially for regions of water scarcity. This paper presents a systemic optimization of the water balance of the Skorpion Zinc refinery with the aim of reducing water consumption. An Aspen Plus simulation of the process is used. The validity of the simulation is tested by measuring key output variables and comparing results to plant data. A number of water minimization scenarios are investigated, Including unit operation and circuit configuration changes. The scenarios leading to the largest reduction in water consumption are through the full recycle of treated effluent water, which results in water savings of up to 19%. Reducing process water and/or recycling of untreated water is prohibited by the build-up of trace elements, which affect product purity. The Skorpion process already features a highly optimized water balance, with unit operational changes merely resulting in a shift in the water balance. Consequently, the largest area for improvement is through the re-use of effluent water. Keywords Skorpion zinc water usage, process optimization, sustainability.

Jan 1, 2011

By Q. Li, B. Ricks

Over the last several decades mineral concentrate long distance slurry pipelines have been installed and successfully operated. Nearly all were installed via the buried construction method. Their record of depend ability and economy has been well proven. However, these systems operate in regions where seasonal soil frost penetration is shallow or non-existent. In recent years interest in long distance mineral concentrate pipelines in northern cold regions has been increasing. For cold regions one must consider seasonal temperatures and soil frost penetration in order to ensure reliable operations through the year. Herein a case study is developed where heat balance and temperatures for small and large diameter long distance slurry pipelines are compared. The study development is as follows: ?Rheological properties are analysed over range of temperatures from near freezing (4ºC; 39ºF) to relatively very warm (40ºC; 104ºF). ?Then assuming similar climatic conditions, soil properties, and pipeline length, a study is carried out for two pipeline diameters, namely: a small diameter (nominal8 inch; 219.1 mm) and a large diameter (22 inch; 558.8 mm). ?The study focus is on maintaining slurry temperature above freezing throughout the pipeline length by illustrating a decision based on burial depth versus coating the pipeline with insulation.

Jan 1, 2007

By G. M. Evans, S. Mitra, G. J. Jameson, M. M. Hoque, E. Doroodchi, R. Moreno-Atanasio

In this study, the detachment behaviour of 314 μm diameter silica particles from a single 3 mm diameter bubble rotating in a confined cavity connected to a cross-flowing water channel was investigated. The liquid cross-flow generated a rotating flow inside the confined cavity, which was quantified using 2D particle image velocimetry (PIV) measurements. The fluctuating (x-y) velocity components inside the cavity for each cross-flow velocity, U, were then used to estimate the corresponding time/volume average turbulence intensity values, Ti. Turbulence intensity ranged from 9.3-16.2%. In the absence of a bubble, the rotating vortex within the cavity created a pressure field which decreased towards the centre of the vortex; wherein the lowest pressure zone was located at the centre of the cavity. The vorticity into the cavity was found to increase with increasing the turbulence intensity. A bubble, both without and with particles attached, was introduced into the cavity under different flow conditions and its motion was tracked using the shadowgraphy technique. Experimentally, it was found that the bubble-particle aggregate was trapped at the centre of the cavity (low-pressure region) following a spiral trajectory. An interface resolved 3D computational fluid dynamics (CFD) model was developed to predict the particle detachment conditions within the confined cavity. CFDpredicted vorticity in the absence of a bubble agreed well with experimental observation. The CFD analysis indicated that the vorticity inside the cavity was not significantly changed when a single bubble was introduced. On this basis, the computed vorticity was utilised in describing the force balance applied to the attached particle. Computationally, it was found that the critical turbulence intensity, beyond which the particle was detached from the bubble, was 14%. This value was consistent with the experimental observation of detachment at Ti at 15.7%, and above, but not at lower values. Keywords: Bubble, particle, turbulence intensity, vorticity, particle detachment, CFD, PIV

Jan 1, 2020

By S. Krishnannair

Singular Spectrum Analysis (SSA) is an effective data-adaptive multimodal tool for monitoring chemical processes. However, monitoring multiscale signals using principal component analysis obtained from the multilevel decomposition of process signals using basic SSA assume the upper control limits of monitoring charts (" ) to have a Gaussian distribution. This assumption deteriorates the performance of basic SSA when applied to the monitoring of nonlinear processes. To address the issue in this paper, the kernel density estimation (KDE) was used to estimate upper control limits of monitoring charts for basic SSA based nonlinear process monitoring techniques. The monitoring performance of the resulting basic SSA-KDE approach was then compared with a basic SSA approach of which upper control limits of the monitoring charts were determined based on a Gaussian distribution. Application on nonlinear dynamic simulated process and base metal flotation plant chemical processes showed that the SSA-KDE approach provided better overall performance than the SSA approach with Gaussian assumption-based upper control limits for its monitoring charts. Keywords: Fault detection, singular value decomposition, singular spectrum analysis, kernel density estimation

Jan 1, 2020

By A. Ndibalema

Blasting is one of the primary functions in any mining operation. Despite the fact that it generally constitutes between 30% to 40% of the mining cost, blasting also affects the cost-effectiveness of down steam activities such as load and haul, safety, tyre life for haul trucks, crushing and milling. Over the past three decades, significant progress has been made in the development of new technology in an attempt to reduce costs and increase efficiencies and productivities of blasting activities. Mine to mill project concepts have paved the way for such developments. Other advancements have included sophisticated computer modelling technologies for blast design and performance analysis. There are many factors that affect the cost of blasting. These can be divided into two categories; controllable and uncontrollable. The latter includes costs related to depreciation, exchange rates and government legislation with the former category including contractor versus owner blasting costs, efficiency and productivity issues and product quality. Controllable factors have one common denominator which can determine their outcome; direct site management influence. Excavator digging rates and final floor conditions give a good measure of blasting performance, and alternatively cost effectiveness. This paper is aimed at benchmarking against a blast optimisation database, understanding the relationship between costs and identifying opportunities for improvement. Understanding the cost drivers allowed a typical cost model to be developed for the current mining operation at Geita and this was used to drive continuous improvement with emphasis on closing the information loop.

Jan 1, 2008

By H. E. Bartlett

Previously, the grade of ore at the shaft head was taken to be equivalent to the grade measured in the faces by means of chip sampling. The tonnage mined from the stopes is determined from survey volumetric measurements. Estimates of shaft head grade and tons include the grades and tons of waste, development ore, sweepings from old areas, and other sources, not all of which are sampled. The tonnage of ore delivered at the shaft is often estimated using skip factors. These factors are subject to manipulation and not all the skips are filled to the same level. Consequently, both the grade and tonnage of ore delivered at shaft head are subject to a measure of uncertainty. For many years, these uncertainties have led to disputes between shaft managers and metallurgists over the reconciliation between shaft tons and grade and plant tons and grade. These problems are compounded when several shafts feed a central processing plant. Shaft bonuses and a shaft?s profitability are affected by poor gold allocation methods. Without measurement of the grade and tonnage from all shafts, a poorly performing shaft could benefit at the expense of the other shafts. In order to improve the measurement of grades and tonnage at the shaft head, which is the custody transfer point between the mine and surface operations, South African mining companies have developed a methodology to measure the mass of ore from each shaft using rail weighing equipment, and the run-of-mine grades delivered from each shaft using cross-stream (go-belt) sampling. Harmony Gold Mining Company has implemented the rail weighing and go-belt sampling methods for their shafts. This paper deals with the statistical tests on the procedures and control measures that are in place to establish confidence in the gold allocations. The metal accounting system for examining the allocations follows the first principle of the AMIRA P754 Metal Accounting Code of Practice namely ?The metal accounting system must be based on accurate measurements of mass and metal content. It must be based on a full check-in check-out system?. The check-in is mass and grade of run-of-mine ore and the check-outs are the plant input grades and the recovery and residue from the plant. The balances between check-in and check-out are the subject of the statistical analyses, which have been conducted using the classical statistical tests and procedures that are generally used in all fields of endeavour. It is concluded that the gold allocation procedures that rely on go-belt sampling and rail weighing of ore from shafts has been successfully implemented, and that the system, which has been in operation at Harmony since 1993, has won general acceptance from stakeholders, with the gold allocations being perceived as being fair to all. Apart from the on-mine issues of bonuses and fair allocations, accurate metal accounting assists in the broader issues of generating accurate production figures for management. An added benefit of weighing and sampling ore at the shaft head has been that shafts now have daily moving average grades, which are used for grade control. The go-belt sample at the shaft head is the first accurate sample of production in the value chain from face sampling to gold bullion. Keywords metal accounting, go-belt sampling, run-of-mine sampling, codes of practice, AMIRA P754, train weighing, statistical analysis, gold split, allocation of gold, metal balance, grade control.

Jan 1, 2014

By L. P. Fatti

An important feature affecting the classification of a remotely sensed scene is that the data are, in general, spatially correlated. Various approaches are discussed towards incorporating the correlation between neighbouring pixels when they are classified on the basis of their spectral reflectance data.

Jan 1, 1987

By K. M. Banks

The hot ductility of in situ melted tensile specimens of micro-alloyed steels having C contents in the range 0.12?0.17% (mass %) have been examined over the 700?1000°C temperature range. An improved testing method for simulating the continuous casting process was used, which takes into account both primary and secondary cooling conditions. Increasing the N content to electric arc furnace levels (0.01% N) was found to cause a serious deterioration in ductility. V-N steel gave better ductility than Nb-containing steels due to less precipitation. From a cracking perspective, low- N steels are generally recommended but, when not feasible, a combination of Nb and V gives even better ductility. However, to be sure of avoiding transverse cracking in higher N steels a small addition of Ti is required. This resulted in a decrease in the fraction of fine particles and in accord with this better ductility. Transverse cracking of industrial slabs was then avoided.

Jan 1, 2011

By J. A. Maritz

Since the dawn of mining, pillars have been used as primary support to ensure stable workings. Early designs were based on trial and error, after which more scientific means developed over time. A vast amount of progress has been made, especially in soft-rock room-and-pillar design methodologies, from which hard-rock design theories developed with minor changes to constant parameter values. The commonly used Hedley and Grant method for hard rock and Salamon and Munro methodology for soft rock draw on the tributary area associated with the pillar, the width-toheight ratio of the pillar, and a back-analysed strength reduction factor. In these methods, only the vertical stress, or stress normal to the pillar influences the load applied to the pillar. This investigation considers the possible influence of shear stresses on pillars in a room-and-pillar layout in single reef planes and multi-reef environments, based on elastic numerical modelling methods. The possible shear stress poses a safety and financial risk to the design process, whereby an undersized pillar would lead to unstable working conditions, whereas oversized pillars could lead to an under-utilized ore resource.

Jan 1, 2014

The variety and sources of sampling errors have been studied since the late 1800s, but the pioneering work of Pierre Gy in the 1950s provided insights into sampling as an error generating process. Total sampling variance is a measure of the total sampling error, but disaggregating the variance into its component parts is problematical. Ten sources of error and error generating processes have been identified, some of which can be eliminated through rigorous application of the principle of correct sampling. Others can only be minimized but never eliminated. Three principal questions relating to the appropriate mass of a sample, the fragment size of the sample and the size of the associated error are addressed. These questions relate principally to the Fundamental Error (FE), the only error that can be estimated before the sampling event. Aspects of constitutional and distributional heterogeneity are examined, as are the essential components of Gy?s formula including the shape factor (?), the granulometric factor (g), the mineralogical composition factor (c), and the liberation factor, l. Applications of Gy?s equation to the problems of sampling in the minerals industry are explained using real-life examples. The approach to sampling taken by Pierre Gy provides an excellent inferential relationship between the mass of material to be sampled, the size of the fragments being sampled, and the relative variance of the sampling error. It also provides solutions for three of the main problems associated with the Theory of Sampling.

Jan 1, 2007

By N. Sinclair

"In open pit dozer bulk push operations, the material left resting against the highwall after blasting poses an issue. This is because dozers cannot physically get in behind the waste to begin the bulk pushing process. Instead, an excavator can be used to side-cast this waste for the dozers to access. An alternative technique is to have the dozers ‘side-cut’ the material, which involves the dozers running parallel to the wall, digging out the material, and eventually rotating their blades to begin bulk pushing. Industry practices suggest the latter technique is less productive and economical than using an excavator. However, such excavator sidecasting ‘economic viability’ assumes that smaller/ancillary type excavators are used (i.e. less than 23 m3 bucket capacity). So, what happens if some mine sites are restricted to using relatively larger excavators (> 23 m3 bucket) that are preferred for higher priority digging tasks? When can they avoid using their excavator? This paper explores multiple scenarios that have been designed and developed to simulate data results, which identify conditions in which the dozer side-cutting technique becomes more economically viable than excavator side-casting in terms of decreased unit cost per loose cubic metre. IntroductionOptimizing the cost of waste removal operations is important for the profitability of opencast coal mines. In the past decade, dozer bulk pushing has emerged and taken over from traditional waste removal operations such as dragline and truck-and-shovel methods at various sites around the world (Beatty, 1994). In a typical bulk push operation, once the bench has been blasted, waste is left resting against the highwall, which can either be cleared using a loading unit or by having the dozers ‘side-cut’ the highwall.Dozers perform most effectively when slot dozing side-by-side, perpendicular to the highwall. However, when dozers are required to remove material off the highwall, they must push parallel to the highwall, rather than perpendicular, hence ‘side-cutting’ the material to ensure a clean face. While dozer sidecutting is possible, it has been found that the majority of mine sites use a loading unit in this situation to avoid dozer inefficiency. Other perceived issues with dozer side-cutting include higher maintenance requirements and safety risks to dozer operators working in close proximity to the highwall."

Jan 1, 2019

To determine the characteristics of a mineral deposit, the usual practice is to take samples, analyse the properties of those samples and infer the characteristics of the deposit from these properties. This analysis can be done using statistical methods. In the present work we are concerned with two types of statistical approaches, classical statistics and spatial statistics. If one uses classical statistics to represent the properties of the sample values, the assumption is made that the values are realizations of a random variable. The relative positions of the samples are ignored, and it is assumed that all sample values in the mineral deposit have an equal probability of being selected. The likely presence of trends, zones of enrichment, or pay shoots in the mineralization, is ignored. The fact that two samples taken close to each other are more likely to have similar values than if taken far apart is also not taken into consideration. In contrast, spatial statistics will be used if one chooses to consider that the sample values are realizations of random functions. On this hypothesis, the value of a sample is a function of its position in the mineralization of the deposit, and the relative position of the samples is taken into consideration. The similarity between sample values is quantified as a function of the distance between samples and this relationship represents the foundation of spatial statistics. There are few situations where classical statistics can be used. The assumption that all sample values in the mineral deposit have an equal likelihood of being represented will be satisfied only if the sample values are randomly distributed, or if the sample positions are random. Sample values are in fact never randomly distributed within a mineral deposit. Further- more, geologists usually avoid taking samples at random (random sampling), as it is correctly accepted that samples located on a regular grid, or approximately on a regular grid, give more information than randomly located samples. In practice, classical statistics should be used only in the early stages of exploration, when the number of samples available is relatively small and the distances between samples are large. In these circumstances, and whenever the information available is not sufficient to permit the use of spatial statistics, application of the methods described in this chapter is justified.

Jan 1, 1978

By P. S. Piper

In the 1960s it was suggested that the volume of stope convergence, and hence the rockburst hazard, could be reduced by the emplacement of backfill. Recent advances have facilitated the placement of large quantities of backfill underground, and a reassessment of its potential benefits as part of a regional support strategy is now timely. In this paper the principal criteria of effectiveness of backfill are taken to be its impact on face energy release rate (ERR) and average pillar stress (APS) levels, and maximization of acceptable ore extraction ratios. The possible effects of backfill on excess shear stresses (thought to modulate large seismic events) are also briefly reviewed. Analytic solutions describing backfill properties and their impact on ERR and APS in certain idealized geometries are also presented, and compared with the results of an extensive suite of computer simulations, in which parameter sensitivities are explored. Results indicate that a regional support strategy is required at depths exceeding 2 km in order to maintain acceptable levels of ERR, and backfill ribs can on their own fulfil this role at depths less than 3,3 km. At greater depths use of stabilizing pillars becomes imperative, but here too supplementary backfill ribs can reduce ERR and APS levels and permit significant increases in ore extraction ratios.

Jan 1, 1988

By S. G. Mgenge, J. D. Steenkamp

Tronox Namakwa Sands operates two ilmenite smelters for the production of high-titania slag and low-manganese pig iron. Slag or metal is tapped from one of two product-specific, bi-level tap-holes. In this paper the refractory design specific to the tap-hole refractory layout is described, as well as practices for slag and metal tapping and refractory maintenance. Further developments planned are briefly described.

Jan 1, 2014

Anglo Platinum is in the process of introducing a number of initiatives, varied in nature, on a scale that will impact operations in the short, medium and long-term. The advent of mechanization on a large scale and the introduction of new mining technologies therefore require a different approach to the design of ventilation and cooling systems. In addition, the ventilation and cooling systems serving increasingly deeper operations will place a greater demand on electrical power consumption?this in an environment where power generation is limited and increasingly expensive. Ventilation and cooling systems currently being employed in platinum operations have evolved from practices first pioneered in deep level gold mines. However, in modern and future platinum mines, deepening of operations will take place in a somewhat different environment?particularly in terms of geothermal settings and electrical power utilization. This provides the opportunities to devise novel and energy-efficient ventilation and cooling strategies tailor-made to meet these challenges. The need for refrigeration and cooling at depth will increase the capital and operational cost component of ventilation systems and the medium-term limitations on the availability of electrical power will require the provision of innovative systems. The increased use of diesel-powered vehicles underground, the viability of energy recovery and feasibility of thermal storage opportunities are discussed in this paper. Some of the challenges and the solutions currently proposed in terms of ventilation system design, refrigeration plant selection and chilled water distribution methodologies are described. Changes in strategies core to this discipline that affect the health and safety of employees and the profitability of modern operations are deemed necessary to support the transformations that are envisaged.

Jan 1, 2008