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By M. J. Latva-Kokko, CA Snyders, J. R. Peltola, M. K. Kostiainen

It is well known that carbon management is essential for minimising soluble gold losses in carbonin- leach (CIL) and carbon-in-pulp (CIP) circuits. Even so, carbon measurements remain predominantly a manual operation with a low amount of control and automation around CIL and CIP reactors. Carbon inventory, i.e. knowing the carbon concentration in each reactor, is crucial for good carbon management. CarbonSense, a novel in-situ measurement system, was therefore developed for continuous monitoring of the carbon concentration in each reactor and paves the way for more advanced control in gold operations. The carbon management and optimisation system is based on Outotec® HSC Sim dynamic process simulations where each process component is modelled in detail. Generated models are validated with real data from a particular operation. The simulator will allow for the carbon control sequence to be realistically tested and the efficiency optimised by controlling the feed and discharge pumps in a hazard- and risk-free process environment. Furthermore, by combining CarbonSense and Outotec® gold courier (online soluble gold measurements) continuous feed-forward and feedback information can be obtained. The optimiser will then define the best running parameters based on the cost function of gold production versus intrinsic factors such as utility costs and adsorption time. Eventually, the application can be installed parallel with a real process island and it can be continuously calibrated with real-time process data. The dynamic model will then create future predictions for the entire process island. The example case presents a CIL circuit, which is simulated using Outotec® HSC Chemistry Sim and operated through Outotec® ACT (Advanced Control Tool) application. Keywords: Carbon-in-leach, CIL, CarbonSense, carbon management, automated control, simulation

Jan 1, 2020

By D. G. Krige

Using a very large number of chip-sample values from a mined-out area of the Hartebeestfontein mine and accepting these as equivalent to borehole values (also areas of 1 km by 1 km as equivalent to 'mines'), the authors apply a series of borehole-valuation procedures to these 'mines', Large undeveloped sections of existing mines are also regarded as the equivalent of such 'mines'. These procedures start with the orthodox arithmetic mean and then introduce additional information in the form of the Iog normal frequency distribution model for Sichel's 't' estimator. Knowledge of the average log-variance within mines leads to the tH estimator. The mean grade of the 'family' of mines to which a new mine belongs is then introduced via regression or elementary kriging. Next, further information on the macro spatial structure of the grade is used, together with the grades of adjacent existing mines in ordinary and simple macro-kriging procedures; finally, via the distribution model, the grade is seen as a function of the logs of the grades and their variances. These are introduced as spatial variables, and both are estimated by macro-Iog normal ordinary and simple kriging procedures with grades and variances from adjacent mines'. Comparisons are also made between macro-kriging on lognormal transformed values versus untransformed values. Conclusions are drawn on the relative importance of the various items of additional information used, on the procedures followed, and on the significance of the advantages gained

Jan 1, 1992

By J. Githiria, D. Wachira

The conventional truck and shovel haulage system has become expensive to operate in most mines due to current strict environmental conditions, increase in depth and size of open-pit mines, increase in labour costs, and fuel costs. These factors have motivated the search for alternative material haulage options in open-pit mines. In-pit crushing and conveying (IPCC) system has been regarded as a plausible alternative in material haulage systems. However, usage of IPCC has been limited to a few mines due to limited knowledge of IPCC performance, and other technical factors of the system. This paper investigates, and evaluates the productivity performance of four types of IPCC systems (fixed, semi-fixed, semi-mobile, and fully mobile IPCCs) using mine productivity index (MPi). Using a model developed by MATLAB, the MPi for the four systems is computed using input variables such as availability, utilisation, performance and number of trucks and shovels in a limestone quarry. The mine productivity index for fixed, semi-fixed, semi-mobile, and fully mobile IPCC is obtained as 67.13%, 69.02%, 70.52%, and 71.52% respectively. In conclusion, the fully mobile in-pit crushing and conveying system is the most appropriate IPCC in the quarry, due to its better productivity performance compared to the other three IPCC systems and the truck and shovel system currently in use.

Sep 1, 2021

By J. H. Viljoen, A. Heyns, J. F. Morrison, C. H. Wyndham, N. B. Strydom

Discussion M. J. Martinson* (Associate Member): This is another interesting paper in the series which the Human Sciences Laboratory is publishing on the physiological cost of the more strenuous manual tasks performed underground. As a mining engineer-albeit presently an academic one -this contributor is more concerned with the practical significance of the study in the field of stope cleaning than with physiological niceties, but before discussing some of the practical issues raised by the paper perhaps the authors would care to reply to lay comment on three physiological matters. Firstly, in two places the authors make statements to the effect that the maximum oxygen intake of 'average' (sic) Bantu mine workers is 3.0 l/min. What do the authors mean by 'average'? Maximum oxygen intake in the individual is probably some function of inter alia age, physical fitness, state of health, ambient partial pressure of oxygen, diurnal and/or seasonal changes in metabolism, and possibly cumulative exposure to dust and other occupational contaminants. Since maximum oxygen intake is apparently such an important quality in manual labourers we ought to have a clear picture of its distribution in the entire labour force, and also know precisely how the individual's quantum is affected by factors such as those mentioned above. Secondly, arising out of the same statements and ignoring the reservation implicit in the previous paragraph, is it correct to say that the 'average' Bantu mine worker has a maximum intake of 3.0 l/min? In the August 1966 issue of the South African Journal of Science four of the present co-authors reported the mean maximum oxygen intake of a random sample of 338 Bantu recruits of different tribes to be 2.63 l/min with a standard deviation of 0.542 l/min, and these figures were repeated in a paper published in the November 1968 issue of this Journal. Admittedly the mean maximum oxygen intake of the six subjects used in the present study was 3.05 l/min, but since they were selected on the grounds of previous shovelling experience they can hardly be regarded as being 'average' recruits. Thirdly, it seems to me that the authors are perhaps a little uncritical in their acceptance of the '50 per cent of maximum oxygen intake' level of activity for endurance work. I am not unmindful of the evidence upon which this level is based, but I suggest that the evidence is at best tenuous and that the whole topic of fatigue and 'optimum' levels of productivity under conditions of high muscular activity and considerable psychological stress is a lot more complicated than would appear from the authors' ready acceptance of the 50 per cent level. Under present working conditions the question is largely rhetorical because it would need an army of supervisors to ensure that all members of a typical gang work at 'optimum' levels throughout the shift, but if individual performance should ever become more readily determinable the issue might become crucial to management and malayisha alike. So much for physiological matters; now a few random thoughts on some wider, practical issues raised by the paper. In the paragraph headed Material the statement is made that 'The mixture (of fine, medium-sized and bigger rocks) was considered to be representative of an average underground rock sample'. Elsewhere in the report the authors specifically draw attention to the effect of particle size on the oxygen consumption of men shovelling graded material, and intuitively one imagines that the size distribution of the material being shovelled would be highly significant in a study such as this. Presumably the authors were forced to assess size distribution subjectively because of the lack of objective data on the subject, and possibly the necessity for this assumption highlights how little we know quantitatively about any of the variables in stope cleaning systems. In this connection it may be remarked that although the present study was largely designed to show how dip and stoping width effect energy expenditure in shovelling, no data are given-nor, probably, are they available-to show how the industry's stoping tonnage is distributed in terms of these two factors. It seems to me that if the mining industry seriously hopes to improve stope productivity it must start by measuring the basic variables in existing systems. Elaborating somewhat on the previous paragraph, there also seems to be a tremendous gap between the present study-and its predecessors-and the prospect of gainfully using the information so carefully garnered by the authors. It would be reassuring to hear from the authors that their work on the physiological aspects is being integrated in some master plan for increasing productivity of manual labour. This I fear is not the case, and if productivity has changed at all in recent years I would hazard a guess that the change has been brought about by random factors rather than by systematic engineering design. The Human Sciences Laboratory has worked extensively in stopes in the past, and possibly the time has come when it should widen the ambit of its operations to include engineering-type analyses of cleaning systems, with particular emphasis on the economics of rapid cleaning by manual labour. Perhaps experiments could be conducted with stakhanovite-type gangs consisting of men selected for their high maximum oxygen intake and superior motivation; these men might be offered special incentives in the shape of extra pay

Jan 10, 1969

By J. Bal, V. D. Dao, W. Li

"SynopsisThis study proposes a methodology for evaluating the risk of coal bumps in multiple-seam longwall mining. Both the stress field and the total energy release (TER) during retreat were evaluated in the specified case involving multiple-seam mining using the LaModel program. The results of numerical simulations indicated that both the peak vertical stresses on the panel edges and the peak TERs in the outby longwalls increased significantly as the horizontal offsets were reduced from 60 m to zero. With the comprehensive consideration of the stress field and TERs, a conservative offset of 60 m was ultimately adopted when developing the gateroads of the lower panel in the field. The field measurements indicated that coal bumps were avoided completely by employing the proposed design, and the maximum roof-to-floor and rib-to-rib convergences of the tailgate during retreat were only 360 mm and 576 mm, respectively.IntroductionCoal bumps – sudden, violent bursts of coal – pose a serious threat to the safe excavation of coal. Their occurrence depends on the properties of the surrounding rocks, stress field, and dynamic disturbance. The complexity of the mechanism of coal bumps is further enhanced in multiple-seam mining due to seam interaction. Multiple-seam mining is most frequently found in the coalfields in eastern China, such as Xuzhou, Xinwen, Zaozhuang, and Huainan. In the past, due to issues of availability, economics, and ignorance of the potential risk in multipleseam mining, coal seams in such conditions were mined without proper planning to account for seam interaction. Fortunately, studies published over the past three decades have raised awareness of the problems associated with multiple-seam mining, and indicated that the ground control issues induced by multiple-seam mining can be avoided and/or minimized with proper planning.A number of technical studies pertaining to the design of multiple-seam mines have been published. These studies can be divided into two groups based on their methodologies of design: empirical (Chekan and Listak, 1993; Haycocks and Zhou, 1990, Matetic et al.; 1987a, 1987b) and numerical approaches (Li et al., 2000; Maleki, 2007; Mark, 2007; Peng, 2007; Zipf, 2005). The majority of the extant studies on the interaction mechanisms in multiple-seam mining have focused on the load transfer during operation, resulting in the proposal of the pressure bulb theory (Forrest et al., 1987; Haycocks and Karmis, 1983; Peng, 2008; Peng and Chandra, 1980) and pressure arch theory (Holland, 1973; Oram and Ponder, 1997) to guide multiple-seam mine design. However, neither, of these theories can be used to evaluate the potential for coal bumps."

Jan 1, 2015

Contribution by G. Narsimhan It was with much pleasure that I went through D.F. Malan's excellent paper. The following explanatory and critical note has been the result of my deep study of the paper. Although by qualifications and experience I have been a chemical engineer, involved over four decades with problems in the modeling and simulation of transport-limited (heat and mass) chemical reactions in multi-phase systems, I have of and on interested myself in challenging problems in extractive metallurgy. This is the first time I have veered of to continuum mechanics, at the age of 71! Author's reply. I thank Professor Narsimhan for his extensive contribution to the paper. It provided me with the stimulus to derive an extended model. However, there are some inaccuracies in his arguments that need further discussion. Professor Narsimhan's main concern is that there is not a good enough fit between the derived viscoelastic Kelvin model and the experimental data supplied by Güler1 (see Figure l1 on page 218 of my paper). I will therefore explain the choice of a viscoelastic model in more detail and also make specific comments on the arguments given in his contribution.

Jan 1, 1996

By F. K. Crundwell

The rate of dissolution of sulphide minerals is frequently interpreted in terms of the passivating behaviour of a layer of material on the surface. This interpretation has driven research to identify and possibly remove this layer from the surface, and thus yield higher rates of dissolution. The identification of such layers has proved difficult, with different researchers claiming different compositions for possible layers. An alternative framework has been pursued by the author and co-workers, who have argued that the rate of dissolution is governed by the electronic structure of the minerals which give rise to their semiconducting properties. Researchers sceptical of this approach have argued that semiconducting properties are only important in highly pure substances. In this paper, we review the electronic and semiconducting properties and discuss how they impact dissolution behaviour. These properties affect their colour and dissolution properties even though the minerals are not high purity substances. Keywords: Dissolution, leaching, passivation, semiconductor, chalcopyrite

Jan 1, 2020

By C. Birch, J. Perold

"Due to the challenging geological environment of the Postmasburg Manganese Field (PMF), a study was conducted to determine if any benefits would derive from the inclusion of resistivity data during threedimensional (3D) modelling of the manganese resource. This was achieved by estimating manganese resources from 2011/2012 drilling data and comparing them with manganese resources estimated from the same drilling data and resistivity data collected during 2013 and 2017.Both models were adjusted to limit their extent to the same 3D modelling space. Significant volume and tonnage differences were observed for all lithological units. The greatest differences were noted in the manganiferous zones of alteration – 7.200 Mt for the geological model versus 3.700 Mt for the geoelectric model.This study showed that the inclusion of resistivity data can reduce exploration costs significantly, as a direct consequence of the resistivity data allowing more accurate siting of boreholes. This decreases the number of boreholes, samples, and analyses required due to the 3D electrical delineation of mineralized areas prior to drilling.An additional benefit is the ability to more correctly forecast the net present value of an operation due to more accurate estimation of manganese resources and stripping ratios. This is clearly demonstrated by the estimated gross profit estimation of R409 million for the geological resource model versus R264 million for the geoelectric resource model.The addition of resistivity data can, therefore, reduce exploration costs and can increase confidence in geological and financial modelling. It would be reasonable to conclude that this approach could also be used for karsthosted massive sulphide deposits. IntroductionAccording to Corathers (2014) and Gajigo, Mutambatsere. and Adjei (2011), South Africa holds between 75–80% of global manganese reserves, followed by Ukraine (10%), Australia (3%), India (3%), and Gabon (2%). The known, land-based, manganese resources located elsewhere are estimated at 2%.The majority of economically important sedimentary manganese ore deposits in South Africa are situated in the Northern Cape Province (Gutzmer, 1996). Manganese ores of the Postmasburg Manganese Field (PMF) were discovered in 1922, and the Kalahari Manganese Field (KMF) in 1940. According to Astrup and Tsikos (1998) and Gutzmer (1996), mining of manganese ore from the PMF ceased during 1989 in favour of the superior quality of manganese ore from the KMF."

Mar 1, 2019

By R. Nyirenda, B. Besa, N. Dzimunya

The accurate estimation of peak particle velocity (PPV) is crucial during the design of bench blasting operations in open pit mines, since the vibrations caused by blasting can significantly affect the integrity of nearby buildings and other structures. Conventional models used to predict blast-induced vibrations are not capable of capturing nonlinear relationships between the different blasting-related parameters. Soft computing techniques, i.e., techniques that are founded on the principles of artificial intelligence, effectively model these complexities. In this paper, we use the random forest (RF) algorithm to develop a model to predict blast-induced ground vibrations from bench blasting using 48 data records. The model was trained and tested using WEKA datamining software. To build this model, a feature selection process using several combinations of Attribute Evaluators and Search Methods under the WEKA Select Attributes tab was performed. The correlation coefficient of the actual data and RF model-predicted data was 0.95, and the weighted average of the relative absolute error (RAE) was 10.9%. The RF model performance was also compared to the equivalent-path-based (EPB) equation on the testing data-set, and it was seen that the RF model can effectively be used to predict PPV. The study also demonstrates that the EPB equation is a suitable empirical method for predicting PPV.

Mar 30, 2023

By R. Simon, P. Pagé, L. Li, P. Yang

The Mathews-Potvin stability method is widely used in the Canadian mining industry as a starting point to determine the maximum dimensions of mine stopes. However, it cannot be applied to inclined (more frequently encountered) mine stopes without conducting numerical modelling to obtain the stress factor A, defined as a function of the ratio of unconfined compressive strength of intact rock to the induced principal stress on the exposed stope walls. The need to conduct numerical modelling significantly limits the application of the Mathews-Potvin method. In addition, given its empirical nature and main application for preliminary design, it is deemed undesirable to conduct numerical modelling, especially elaborate modelling. Alternatively, theoretical methods can provide a much simpler and quicker way to estimate stresses around stopes and the corresponding stress factors. Over the years, a large number of studies have been conducted to estimate stresses around openings excavated with various crosssections. However, theoretical or graphical solutions remain unavailable for mine stopes that typically consist of horizontal floor and roof, and two parallel inclined walls (hangingwall and footwall). To remedy this situation, a series of numerical simulations is first performed for openings with vertical and inclined walls, including typical stopes commonly encountered in underground mines. A group of empirical solutions is then formulated to estimate the induced principal stresses at the roof centre and mid-height of the stope walls. The validity and predictability of the proposed solution have been verified using additional numerical simulations. The proposed solution can thus be used to calculate stresses and the resultant stress factors A around typical mine stopes with any inclination angle and height to width ratio, under any in-situ stress state, without conducting numerical modelling.

Aug 1, 2021

By C. Musingwini, M. Mohutsiwa

"SynopsisCapital cost estimates are important in decisions on whether a project will be approved, mothballed, or abandoned. In South Africa, junior coal miners do not have extensive databases of historical projects from which to estimate capital costs. The purpose of this paper is to establish formulae that can be used for estimating capital costs of developing coal mines in a coal-producing country, using South Africa as a case study. The costs are estimated to an error of magnitude level of -30% to +50%, which is suitable for a concept study level, using a parametric estimation technique. The study uses data from completed coal mining projects from selected coal-producing countries. Three formulae are developed and presented for estimating capital costs of underground bord and pillar, surface shovel and truck, and dragline operations.IntroductionCapital cost estimates play a critical role in deciding whether projects will proceed, be delayed, or abandoned. It is therefore important that capital cost estimates are carried out accurately as determined by the estimation guidelines based on the level of estimation conducted (Shafiee and Topal, 2012). The lack of literature on estimation of capital costs in many coal mining countries is adversely affecting junior coal miners, since they do not have an extensive database of historical projects on which to base estimations. In South Africa, the difficulty in estimating capital costs is attributable to lack of publicly available data on completed South African mining projects (Hall, 2013)."

Jan 1, 2015

By A. C. Lawrence, K. F. Mauer

Human factors in stope productivity -a field experiment by K. F. MAUER, M.A. (Visitor), and A. C. LAWRENCE, B.Econ., D.Com. (Member) An experiment in which three human variables were manipulated in eight stopes was conducted in a gold mine over a period of roughly six months. The three variables, which each had two levels, concerned the degree of movement from gangs, the quality of the Bantu production supervisors, and the extent to which gangs were composed of a single ethnic group (Malawians). Time was included as a variable with three levels in the analysis of the findings. On the average, stable gangs produced 12 per cent more than gangs with a high percentage of transfers, gangs with better production supervisors were more successful by nearly 9 per cent than those with poorer production supervisors, and homogeneous gangs were more productive than heterogeneous ones by as much as 15 per cent. In addition, it was found that a number of the variables exercised a joint influence on stope productivity and that the effects of the variables were accentuated by the len gth of time they had been in operation. V

Jan 10, 1973

By J. H. van Dalen, P. J. Brandt, F. W. Wessels

A maximum of ten process streams (Merrill barren solutions) can be monitored sequentially with this on-line analyser, and gold concentrations as Iow as 0,005 mg/I can be determined. The design is a further development of the Tell-tale analyser constructed by the Corner House Laboratories during the period 1965-69. The barren solution is preconcentrated automatically by solvent extraction, the solvent being a solution of I per cent Aliquat 336 in di-isobutyl ketone. An aqueous-to-solvent ratio of 40: I is maintained. The extracted gold in the organic phase is determined by flame atomic absorption, each analysis taking approximately II minutes. The long-term reproducibility of the absorbance readings over a 24-hour period of repetitive sampling of four simulated process streams varied from 6,9 to 4,1 per cent relative standard deviation for gold concentrations of 0,0 I to 0,05 mg/I.

Jan 1, 1980

By M. Arlandi-Rodríguez, and A. García-Berrocal, A. Alonso-Jiménez, C. Lopez-Jimeno

Spain is a major producer of dimension stone, which is one of the most important industrial sectors in the northwest of Spain. This justifies the maintenance of these industries, although the technical and environmental requirements for their continuity are increasingly demanding. The main target of this paper is to present a new methodology in order to obtain k-ratio stress values on the basis of flat jack measurements and finite element numerical back- analysis with Rocscience© RS2 software. Continuous iterations were performed to try to find a relationship between the k₀ and the measured stress in the tunnel roof and sidewalls. It focuses on technical aspects of the method for designing mine chambers in the slate industry that were previously mined by open pit methods. The size of the chambers impacts the operating cost, and the dimensioning of a correct support ensures better efficiency and safety. Correct knowledge of the k₀ therefore is critical for the design of the support. The proposed new method entails a simplification of the classical procedure for estimating the k-ratio.

Sep 1, 2022

By B. Genc, P. Malepfane

This study proposes a new approach to reduce the environmental impact during surface mining by improving environmental factors through the incorporation of carbon footprint into strategic mine planning. Two key variables were analysed by contrasting and comparing the baseline and optimised models generated. These variables include net present value, which helps determine the financial feasibility of the mining models and their carbon emissions. The variables were analysed by using linear regression, Spearman’s correlation, a P-value hypothesis, and an exponential decay test. Results show that at Mine A the projected net present value was 62% higher than when running a baseline model. Furthermore, the net present value from the baseline model portrayed a high rate of decay at 0.667 versus 0.118 from the optimised model. This indicates the baseline model’s sensitivity to long-term project timelines and fluctuating markets. The optimised method presents an opportunity to manage emissions during the high operational costs phases. It is represented by a 31% improvement in carbon emissions during the periods of high operational costs, whilst maintaining a positive net present value when compared with the baseline model. It was shown that the optimised model presents a better chance of achieving financial feasibility and balancing environmental impact, thereby ensuring that Mine A can align with the environmental aspect of economic, social, and governance principles through lowering the carbon footprint resulting from its hauling operations.

Jan 30, 2026

By N. Hadebe, A. Mampuru, F. Jantjies

Mintek conducted test work to study the effect of media wear when using Mintek’s Milling-in-raffinate technology. The grinding media consisted of high-chrome and ceramic balls. The tests were conducted using a laboratory 5L 316 stainless-steel rubber-lined mill. Magotteaux supplied the high-chrome and ceramic balls. The raffinate was synthetically prepared using a typical Cu/Co (2.2 g/L Cu, 3 g/L Co) composition. The milling rate, media wear, mill, leach extraction efficiencies, gangue acid consumption, and residence time were studied. The tests indicated that the high-chrome balls had a higher corrosion rate than the ceramic balls. The ceramic corrosion rate was lower than 0.1 mm/y. The highchrome balls performed better during the leach stages with faster milling time, higher leaching efficiencies for both Cu and Co, lower gangue-acid consumption, and lower residence time

Jan 1, 2023

By P. Ramsaywok, S Ndanduleni, G. Legodi

The objective of this paper is to share the optimisation journey of a pilot Reflux classifier (RC) circuit in the Exxaro GG1 plant, at Grootegeluk beneficiation complex in Lephalale, South Africa. The RC circuit was commissioned in 2016 to produce semi-soft coking coal at its fines circuit; however the stability of operation as well as the quality control of the coal product was not fully achieved. Various test campaigns and optimisation studies were conducted to understand and optimise the RC circuit at GG1. The studies revealed circuit design inefficiencies; provided an improved technical understanding of the current pilot circuit and the parameters required for the circuit to reach a stable operation. The operation was stabilised with its current inefficiencies to produce an average 12.5% ash semi-soft coking coal product for sustained periods, above the initial target of 10.3% ash. Keywords: Reflux classifier, coal

Jan 1, 2020

By D. R. Young

Drilling exploration over the Wesizwe Platinum Limited?s Pilanesberg Project has identified the Merensky Reef at depths varying from 594 m to 1 552 m depth. Merensky Reef thickness varies from 0.05 m to 3.63 m and contains PGE(4) mineralization above a cut-off grade of 1.0 g/t over intersection widths of 0.95 m to 2.63 m. Four distinct facies types have been identified over the property, viz.1) Normal Merensky Reef, generally 1.2 m thick, bounded by thin top and basal chromitite layers and composed of upper feldspathic pyroxenite and lower feldspathic olivine pegmatoids containing base metal sulphide mineralization overlying poikilitic anorthosites. 2) Single chromitite Merensky Reef, generally0.15 m thick and usually devoid of internal silicate material, overlying variable stratigraphy containing significant base metal sulphide mineralization. 3) Detached Merensky Reef, generally1.0 m thick, bounded by thin top and basal chromitite layers and composed of feldspathic pyroxenite or pyroxenite containing base metal sulphide mineralization overlying Merensky Pyroxenite. 4) Normal footwall Merensky Reef, generally 1.8 m thick, bounded by thin top and basal chromitite layers and composed of feldspathic pyroxenite containing base metal sulphide mineralization overlying norites also containing significant base metal sulphide mineralization. The PGE(4) content?s (g/m2) distribution for the global intersection data appears to be lognormal, and conversion to lognormal space produces a near normal distribution which could be interpreted as a compound lognormal distribution. The mean PGE(4) contents estimated by these three different distributions are the same, and thus do not impact on the contained metal. Data for the Normal Merensky Reef facies type are sufficient to indicate that a normal distribution is followed for this facies type, and therefore does not need to be transformed to log space for estimation of the mean and the 90% confidence limits. A global evaluation based on all of the data is compared with a facies-weighted evaluation. The latter method is noted to provide a better definition of the associated estimation errors and thus potential classification of Mineral Resources. A valuation of $20.95 for Inferred and $41.90 for Indicated PGE(4) ounces is applied to both of the evaluations and compared. The global evaluation is considered to over-value the project by approximately 40%, thus indicating that valuations based on global summary sampling data that have a compound distribution may be materially flawed, and that cognizance of the underlying geology is crucial for unbiased valuation.

Jan 1, 2006

By S. Bradshaw, A. Mabentsela

"Physical and numerical modelling methods were followed to identify mechanism(s) for tundish slag entrainment in a bare tundish and one with a flow control device (FCD).The physical and numerical models made use of water and paraffin to model steel and slag respectively. Observations from the physical model showed that the steel-slag interface remains immobile in both cases. Entrained paraffin formed droplets approximately 1 mm in diameter. Results from both models (numerical and physical) showed that in both cases (bare and FCD case), areas of high entrained slag concentration exist near the inlet region. The entrained slag concentration decreases towards the tundish endwalls.Flow patterns and velocities tangential to the steel-slag interface from the numerical model showed that slag entrainment in both the bare tundish and tundish with a FCD possibly takes place via two mechanisms. First, the slag moves across the steel-slag interface via mass transfer; secondly small velocities tangential to the interface at depths greater than 10 mm below the interface carry the already ‘entrained’ slag into the bulk steel phase. These tangential flow patterns are dominant in the inlet region, hence the high concentration of entrained slag in this region. IntroductionThe tundish serves as the last metallurgical vessel through which steel passes before solidifying in the moulds. It is in the tundish that the last traces of non-metallic inclusions should be removed, otherwise the inclusions carry over to the moulds and cause defects in the steel product. In the tundish, the slag is less dense than the melt and thus resides on the top of the melt. The slag provides a sink into which the non-metallic inclusions float to and dissolve. The slag also protects the melt from air and heat loss. However, when slag is entrained in the melt as a result of increased turbulence or shearing at the steel-slag interface, slag can become a source of nonmetallic inclusions.Numerous studies on flow patterns in the tundish have been carried out since the introduction of the tundish. These studies were aimed at improving the flow characteristics in the tundish (Chattopadhyay, Isac, and Guthri, 2010; Cloete, 2014; Jha, Dash, and Kumar, 2001; Jha, Rao, and Dewan, 2008;"

Jan 1, 2017

By N. du Preez, J. Christie, F. Crundwell

Precipitation is a key process in the recovery of copper and cobalt. While there is a wealth of experience in operating precipitation steps, improving their performance remains challenging. Chemically, the reaction might be controlled by reaction kinetics, be limited by oxygen mass transfer, or be at equilibrium. In this paper, we examine these different operating regimes and demonstrate how they can be identified from plant operating data, process modelling, and pilot plant test work

Jan 1, 2023