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By D. Gerhart, J. Hung, J. Hackwood, A. Pix

A computer-based, real-time mine management system has been installed in a number of underground operations. The system’s innovative software application is coupled with wireless data acquisition and database management to provide complete mine monitoring and control execution in closed-loop control fashion. The system is well suited to automatically assign equipment, acquire status, and reassign resources based on status; and to automate information handling. The system has provided mines with improved utilization of the equipment fleet and mining crews, and reduced workplace cycle time. This paper describes the deployment of systems at a number of underground mining operations and discusses future development of the system. Keywords: Computer-based: Real time: Utilization: Mining cycle control: Monitoring: Productivity

Sep 1, 2001

By B. K. Cross

This paper, which adopts a philosophical approach, attempts to analyse the contribution that the computer is capable of making in geological analysis, financial evaluation, and mine project control. Also examined are some reasons for non-utilization of these capabilities. As with most computer associated problems, the cause is primarily human and not mechanical. The paper gives one numerical example of the analysis of geological data and one of the evaluation of mine-design data. The values used in the examples are intended only to emphasize certain points and are not necessarily reproducible.

Jan 1, 1980

By M. R. Khalesi, Y. Kianinia, A. Seyedhakimi, F. Soltani

"The feed ore to Agh-Darreh gold processing plant (Takab City, Iran) contains 2 ppm Au, 7 ppm Ag, and 120 ppm Hg. After cyanide leaching of the feed at pH 10.4, 80% of the total mercury (in the form of cinnabar and metacinnabar) goes to the tailings dam. In this research, separation of mercury from the tailings by flotation was investigated. The effects of collector type and dosage, pH, and the number of cleaner stages were studied. The results showed that after two stages of cleaning a 40–62% recovery of Hg at a grade of 14.3% Hg is attainable. Recycle water from a test in which 50 g/t amyl xanthate and 30 g/t pine oil was used in the flotation stage showed no adverse effect on the leaching and adsorption of gold onto activated carbon.IntroductionMercury, most commonly found in the nature as the mineral cinnabar (HgS), is a hazardous substance because of its mobility and toxicity (Kyle et al., 2012). Roasting is the wellestablished treatment method for mercury ores, in which the sulphur is burned and volatilized and mercury recovered subsequently by cooling and condensation. However, for low-grade ores, flotation is mainly applied for primary concentration due to its comparatively low costs, low environmental hazards, and flexibility regarding feed variations (Bulatovic, 2007). Particularly, cinnabar can be easily recovered by flotation using pine oil and a collector such as xanthate (Bulatovic, 2007; Crozier, 1991). Floatability of cinnabar is reduced due to surface oxidation when it is exposed to the atmosphere after crushing. This problem can be eliminated by adding copper sulphate and lead nitrate or acetate. The best pH for cinnabar flotation is in the range of 6.5 to 8.5 according to Erspamer and Wells, (1954), who reported that cinnabar and stibnite can be floated with long-chain xanthate collectors and an alcohol frother, with cumulative recoveries of 97.8% and 94.7%, respectively. Lead acetate was used as an activator."

Jan 1, 2017

By S. Annavarapu

"Fragment size distributions play an important part in the design and planning of block cave mining operations. Though several methods have been proposed for estimating fragment size distributions in advance of mining, the calibration of these estimating procedures to field measurements has been a challenge. The Block Size Estimator (BSE) program was developed for providing an assessment of primary fragmentation expected in a block cave mine using drill-core piece length data and joint characteristics from the exploration and geotechnical evaluation programmes. In order to validate the results from the BSE program, the drill-core piece length data and the joint characteristics in different rock types from the DOZ block cave mine in Indonesia were used to help generate fragment size distributions in different areas of the mine. The predicted fragment size distributions were compared with the fragmentation observed at the drawpoints during operation of the mine.This paper presents the details of the validation of the fragment size estimates by the BSE program using fragment size distribution data from the drawpoints at the extraction level at the DOZ block cave mine, and the challenges encountered in developing reasonable correlations between the estimated and measured fragment size distributions. IntroductionThe assessment of fragmentation is an important aspect of the design and planning of the block cave mining method, and forms an integral part of the design of the excavations at the extraction level and the selection of material handling systems for transporting the ore to the processing stations. Secondary blasting requirements can also be estimated based on the fragment size distributions developed for the block cave (Laubscher, 1994).Several methods of estimating fragment size distributions in block cave mines have been developed based on joint set parameters estimated from structural data from oriented core drilling campaigns and mapping of available excavations and outcrops. The BCF program (Esterhuizen, 1999) estimates the fragmentation using a combination of empirical, analytical, and rational methods to model the behaviour of materials during the primary and secondary fragmentation processes. The program JKFrag uses advanced tessellation of joint traces to create rock blocks and then generate fragment size distributions (Brown, 2002). Nickson, Coulson and Hussey (2000) presented the details of processing of the geotechnical data from the Mont Porphyre project, including an assessment of block sizes from the data collected on the core logging sheet. The distance between natural fractures was used to create distributions of block lengths, and a distribution of potential block sizes was estimated assuming that the blocks had equal dimensions in all directions. The results agreed closely with the results of the BCF program. Hadjigeorgiou, Grennon, and Nickson (2002) demonstrated the use of oriented borehole data to provide characteristic block size distributions for a mining project using a software package called Stereoblock, which generates a three-dimensional joint network for a given volume and calculates the volumes of blocks created by the intersections of these joints, simulated as circular planes. The distributions of joint orientation and joint spacing generated from statistical analysis of the oriented core data are used for the creation of the Stereoblock model. The diameter of the circular planes representing the joints is determined based on the mean and standard deviation of the joint trace length obtained from scan-line mapping."

May 1, 2019

By E. J. Magri

Consideration is given to the problem of the estimation, from limited data, of the likely grade and tonnage for a new mining property that is to be exploited by the selective mining of two metals that have substantial contributions to make to the total revenue. In particular, the case of a new gold and uranium mine in South Africa is analysed as follows. (I) The applicability of the underlying lognormal bivariate model for different support (i.e. ore unit) sizes is examined. (2) The necessary parameters for the bivariate lognormal models for different block sizes are estimated from abundant chipsampling data from a section of the Hartebeestfontein Gold Mine using alternative approaches, and the results are compared. (3) A method is given for obtaining the necessary parameters for a tonnage-grade relationship relative to a joint pay limit from the very limited information likely to be available at the end of the exploration stage of a gold mine, and the results are compared with those obtained from the large volume of data available from a minedout area.

Jan 1, 1982

By S. Northey, N. Haque, S. Stoddart, W. Bruckard

Development of new technologies and procedures for managing mineral residues and wastes may benefit from a clear understanding of existing tailing management practices. There is a considerable variability in tailing management techniques used in the Australian copper and gold operations, with associated differences and flow-on effects for the mine site water balance. A database was created, capturing a total of 41 copper and gold processing operations in Australia. The database provides information on tailing generation rates, tailing properties (solids contents and particle size distributions) and tailing storage facility (TSF) types and configurations (e.g. total TSF surface area and wetted surface area). Significant variability was found for the above-mentioned parameters. The solid tailing generation rates ranged from 0.5-37 Mt per annum in 2015. Furthermore, both the solids contents and particle size ranges of the tailing were highly site-specific, and largely dependent on the processing flowsheets. Solids contents of tailings ranged from 45-74% by weight whilst P80 sizes of the concentrator (flotation) feed varied from 75-185 μm. The estimated pumping distance was up to 15 km. The presented data can potentially be used to better tailing management strategies by estimating the potential recoverable water and recycling it, thereby thus reducing tailing footprints. Keywords: Tailing management, tailing storage facility, water balance, mine site, flowsheeting

Jan 1, 2020

By H. Ruthert

An understanding of wear in systems conveying solids is extremely important from the standpoints of design and economics. The lifetime of any bulk solids-handling system depends on the type of solids present and the operating parameters of the transportation system. The relationship between wear in backfill pipelines on South African gold mines and the transportation system's operating parameters is well understood. However, the effect of the transported solids on the pipe wear is less well understood, because of an inability to characterize particle sharpness. The sharpness of the particles, together with their size and hardness, is responsible for the aggressivity of particles, or their ability to cause damage and wear. This paper presents a novel method for the characterization of particle sharpness. The results presented rank particles according to the sharpness of their surfaces.

Jan 1, 1993

By C. K. Chabedi

"This paper discusses the difficulties associated with the potential exploitation of the deep multi-seam resources east of the Daarby fault in the Waterberg coalfield. The resources occur at a depth greater than 250 m and the thickness of the coal is roughly 110 m, but the top 50 m comprises coal intercalated with shale and the bottom 60 m contains five seams with sandstone and shale partings. Various factors affecting multiple seam mining at these great depths are discussed with reference to lessons learned from local and international experience on multi-seam mining. Field geological and geotechnical data was utilized to assess the stability of the roof of the seams. There is no specific rock mass rating for the Waterberg area, therefore approximate coal mine roof rating (CMRR) values were used to propose appropriate support strategies. Analysis of Multiple Seam Stability (AMSS) was used to analyse the strength of the parting or interburden between the various seams, the mining sequence, and the interaction between the various seams.The research indicated that it is possible to mine seams with a low CMRR at high mining rates using longwall mining, although support for gateroads is expected to be expensive, time-consuming and onerous to install, and will impact gateroad development rates. It will not be possible to simultaneously mine zones in close proximity and failure of the interburden is predicted, thus dangerous mining conditions are anticipated. However, it will be possible to mine just two of the eleven zones using longwall mining. IntroductionThis paper discusses the difficulties associated with the potential exploitation of the deep multi-seam resources east of the Daarby fault in the Waterberg coalfield using bord and pillar and longwall mining. Figure 1 illustrates the location of the Waterberg coalfield relative to other coalfields in South Africa, while Figure 2 shows the areas containing deep and shallow resources. The Waterberg resources are expected to contribute to South Africa’s future energy requirements, and are currently exploited at the Grootegeluk open pit mine. Grootegeluk produces coal mainly for a power station, with the higher quality product supplied as metallurgical coal. The deposit is technically unique and challenging, being a multi-seam coal deposit with a total of 12 seams over a thickness of 110 m as shown in Figure 3, including Zone 5. There is currently limited knowledge on the multi-seam mining of the deep Waterberg resources. Multi-seam mining utilizing the bord and pillar method has been practised in South Africa before, but at a depth of less than 100 m in the Witbank coalfields. Multi-seam mining in thin seams has also been performed in the Natal coalfields at a depth of less than 160 m, but mostly using bord and pillar mining and secondary mining (partial pillar extraction)."

Jan 1, 2016

By F. P. Greyling

The aspect of sustainability of ferroalloy processes has long been a serious matter for consideration in many of the world?s industrialized countries. A number of countries can be named where ferroalloy plants that were once flourishing, have all but disappeared. One of the most important factors playing a role in this type of occurrence is the cost and availability of quality raw materials. Given that all natural resources will be depleted with time, users are compelled to develop new processes that are sustainable in the changing environment. The DC arc smelting furnace is one such process that has proved itself, and has gained increasing popularity over the past two decades. This process, amongst other benefits, directly addresses the use of more readily available and more affordable raw materials in ferroalloy production. Since early in the new millennium, GLPS has been involved in various studies for different clients of furnaces for the production of ferroalloys. Two of these recent studies have culminated in the design, construction and commissioning of DC arc furnaces for the production of ferrochrome. One of these furnaces is a 10 MW unit at Mogale Alloys, and the other a 60 MW unit at the Middelburg plant of Samancor Chrome. This paper discusses the approach taken in the design and construction of these two furnaces, which, although varying significantly in size and capacity, are very similar in many other aspects of their design. Keywords Furnace, DC arc, and ferrochrome.

Jan 1, 2010

By E. Bakhtavar

There are many near-surface deposits with considerable vertical extent that have the potential to be mined by a combined method of open-pit and underground methods. In this regard, there is often a point called ?transition depth? where a decision has to be made whether to continue deepening the pit or change to underground methods. Recently, optimization of the transition from open-pit to underground operation has become an important challenge in mining engineering. Optimally, to model the transition problem on the basis of maximization of the profit from open-pit and underground mining, (0-1) integer programming has been used. In this method, block economic values of open-pit and underground methods should be imported to the transition model. A hypothetical example is presented in order to assess the model in detail.

Jan 1, 2012

By H. L. Monro

CONTRIBUTION BY J. R. F. HANDLEY Mr Monro is to be congratulated on his concise treatment of a very complex subject. As a graduate with only Mathematics I in my curriculum, I am unable to comment on the mathematical treatment of the factors affecting his supply/demand model, but the logic does not escape me. The main conclusion is that any excess of supply to the spot gold market will depress the price in a normal market, and particularly in a falling market.

Jan 1, 1993

The vast majority of the roughly 100 million rock anchors installed in mines in the USA each year use resin cartridges1. About 4.5 million of these bolts are installed using a mechanical shell in addition to the resin to create an active (pre-tensioned) bolt. Over 1 million of the bolts are passive cable bolts and typically have an effective grout length of 1.2 m, regardless of the cable length, which could be as long as 6 m. The successful performance of the resin grouted bolts depends on several parameters, including the annular gap between the bolt and hole wall, which should be relatively small, ideally from 3 mm to 5 mm. This requirement, combined with the high viscosity of the resin, produces a high back-pressure that can cause the bolt being installed to buckle or not be installed to standard. It is this back-pressure that limits the effective grout length with passive cable bolts and causes the mechanically anchored bolt failures (typically called ?spinners? where the mechanical shell does not anchor). This creates potentially unsafe conditions and wastes time and money. A purpose, built rig was used to mimic underground installations and record the back-pressures during full scale applications in the laboratory. This information was used, and is still being used, to reduce the failures and sub-standard installations by producing improved designs. In addition, a flow model was calibrated that can act qualitatively to estimate the back-pressures and can be used as a crude screening process before full scale prototypes are built and tested. To date, the results obtained have been used to stop the use of a mechanical shell due to the proven higher rate of failures. A new, improved mechanical shell is being field tested and another system is underdevelopment. The use of the rig is therefore ongoing to develop improved mechanical anchor systems. It is too early for real data; however, anecdotal evidence seems to indicate that significant improvements can and will be made.

Jan 1, 2011

By R. M. Mashishi, T. Malehase, O. J. Okonkwo

The concentrations of selected trace elements (As, Cd, Hg, and Pb in run-of-mine (ROM) coal, discard, and coal product were investigated to assess the efficiency of beneficiation in reducing trace element concentrations and ascertain any implications regarding environmental compliance to regulatory frameworks. Samples were collected from a colliery in Mpumalanga Province twice a month for a period of 24 months. The samples were ashed to approximately 0.21 mm, then digested in a mixture of 70% HNO3 and 40% HF and analysed using inductively coupled plasma-mass spectrometry (ICP-MS). Except for Pb, the mean concentration of all elements decreased from the ROM stage to the coal product stage. The order in which trace elements occurred from the highest to lowest throughout the production chain was Pb>As>Hg>Cd. With the exception of Cd, the mean trace element concentrations in the discard were above the total concentration threshold (TCT) set for landfill disposal. The As, Hg, and Cd concentrations in the product were well below the thresholds for all land uses. However, Pb concentrations in the product coal were above legal limits, which is of concern with regard to environmental compliance and the performance and marketability of the product.

Dec 13, 2023

The Namakwa Sands heavy mineral deposit is located along the West Coast of South Africa and the mine is a world class producer of high quality zircon, ilmenite and rutile concentrates from essentially unconsolidated marine and aeolian sands of Cainozoic age. The objective of this study was to characterize rutile with respect to distribution, grain size, textures, colour and mineral chemistry within the Namakwa Sands orebody with the aim to explain its overall poor recovery. A representative suite of heavy mineral concentrates from various sections of the orebody and stages in the recovery circuit has been investigated microscopically and by means of SEM-EDS and LA-ICP-MS. The rutile grain size distribution displays a wide range and is directly related to its variable chemistry and consequently density. In placer deposits such as Namakwa Sands, mineral sorting is a function of hydraulic equivalence, and high density grains are smaller than the lighter grains when deposited under similar conditions. Grain size is also a function of sediment maturity, and the highly mature red aeolians and (RAS) component of the deposit has a coarser grain size than the less mature orange feldspathic marine sands (OFS). The above primary characteristics of rutile are responsible for the loss of the coarse-grained fraction during screening and, given the empirically determined relationship that increased substitution elements reduce conductivity, also during electrostatic separation. Provenance studies using geothermometry have shown that the heavy mineral suite has been sourced mainly by the proximal medium-to high-grade Namaqualand Metamorphic Complex. The positive relationship between substitution elements and temperature of formation explains their high concentration in rutile of this deposit. The heterogeneity of rutile, produced by the combination of a high temperature primary source and the typical marine-aeolian placer genesis is deposit specific and unfortunately not conducive to high recovery levels in the current Namakwa Sands beneficiation circuit.

Jan 1, 2010

By K. Lomberg

In this paper the process of sampling and assaying as it relates in particular to the platinum group elements (PGEs) is presented. A discussion of the geology of PGE deposits is included to provide an understanding of the mineralization that is considered necessary prior to sampling, as this has a direct bearing on the mineral resource estimate and also the selection of an appropriate mining cut. The discussion includes the selection of the sampling technique (drilling or face sampling), the size of the sample relating to the requirements of the laboratory, the potential mining cut, as well as a reference to sampling theory. The selection of an assay technique that is appropriate and will yield the optimal result based on sample turnaround, precision, accuracy, and cost is then considered. Examples are drawn from work undertaken on the Bushveld Complex (Merensky Reef, UG2 Chromitite Layer, Platreef) and Great Dyke (Main Sulphide Zone). A discussion of the appropriate QA/QC programme includes the chain of custody, CRM selection and the QA/QC process. Keywords sampling, assaying, QA/QC, platinum group elements.

Jan 1, 2014

By A. J. Rogers

The KX36 kimberlite was discovered using a high-resolution gradient aeromagnetic survey, with follow-up detailed ground geophysics and deflation soil sampling used to assess and prioritize the geophysical anomalies identified from the initial aeromagnetic survey. Prospect drilling intersected kimberlite in the first hole drilled on the KX36 anomaly. Preliminary evaluation has shown that KX36 is diamondiferous, and a more detailed evaluation and sampling programme is currently underway. The area around KX36 has been prospected by numerous companies over the last few decades, with sporadic kimberlite indicator mineral (KIM) recoveries in regional deflation sampling indicating the possible presence of kimberlites, but without sufficient resolution to ?home in? on specific targets. Previous (lower resolution) aeromagnetic surveys and soil sampling campaigns also did not define anomalies with sufficient detail to guide prospect drilling effectively. The discovery of KX36 highlights the prospectivity of areas of Botswana under deep Kalahari cover that have been explored previously. Keywords kimberlite exploration, central Kalahari, overburden, high resolution aeromagnetics, deflation sampling, KIM.

Jul 1, 2013

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 R. T. Muller

Contribution to the above paper by R. T. Muller (Visitor) Demands are being made and pressure is being brought to bear on employers in South Africa to narrow or eliminate the wage gap between their White and Black employees. Few thinking people will dispute the fact that a constant improvement in the wages of Black personnel is essential, but this cannot be done merely by a redistribution of income in South Africa.

Jan 7, 1974

By S. Kahraman

The prediction of rock properties from indirect testing methods is important, particularly for preliminary investigations since indirect tests are easier and cheaper than the direct tests. In this study, we investigate the predictability of the uniaxial compressive strength (UCS) and Brazilian tensile strength (BTS) of rocks from the indentation hardness index (IHI) obtained using point load apparatus. Forty-six different rock types, 14 of which were igneous, 15 were metamorphic, and 17 were sedimentary were tested in the laboratory. The UCS and BTS values were correlated with the corresponding IHI values and the results were statistically analysed. The influence of rock classes on the relationships was also investigated. A strong correlation between UCS and IHI was found for all data. The correlation between BTS and IHI is not as strong as the correlation between UCS and IHI. However, it is in the acceptable limits. When the regression analyses were repeated for igneous, metamorphic, and sedimentary rocks, the correlation coefficients were generally increased. The results show the UCS and BTS can be estimated from IHI. In addition, the effect of rock classes on the relationships between IHI and both UCS and BTS is important.

Jan 1, 2012

By B. Talbot

How this simple plant generates a large revenue stream. The concept behind the Tailings Scavenging Plant (TSP) is simple: it is a non-complicated modular plant which is designed to be more stable than upstream plants. This, plus the use of ?high energy?, enables the plant to recover any metal missed by the upstream operation. Impala Platinum?Mineral process started pilot plant exercises in June 2003. The pilot plant ran for one year before the capital for the full-scale plant was approved. This year was used to motivate, design and cost a full-scale plant. The metal produced during the pilot and feasibility study enabled this part of the project to run at a small profit. The scale-up to full-scale was relatively simple as the pilot plant had a capacity of 180 m3/hrand the full-scale plant was designed for a capacity of 550 m3/h. The design for the full scale plant was completed in May 2003, with ground breaking taking place on 19 May 2003. The construction phase of the project went well with the commissioning of the first two modules taking place on 22 October 2003. The remaining two modules were commissioned two weeks later. The plant was at full capacity by the 20 November 2003. A review of the plants performance for the first 5 production months is as follows: the plant has required only a few minor modifications and additions, it has run well in excess of 95%availability with no major wear apparent. In addition to this, the plant has produced 3 times the metal forecast.

Jan 1, 2004