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By F. Rosenblum, K. E. Water, J. A. Finch, S. Moon, Y. H. Tan

Previous work has shown that sulphide self-heating occurs in three distinct stages referred to as Stage A, Stage B and Stage C. In this paper the focus is on the transition from Stage A to Stage B, which occurs at ca. 100OC. Background literature hints that the transition corresponds to the transformation of the rhombic form of elemental sulphur to the more reactive monoclinic form that occurs at 96OC. A test apparatus is modified for adiabatic heating to track the transition. The results support this transformation of sulphur as being key to the transition, and the transition temperature is thus modified to 96OC. Variations in a sample’s response under adiabatic conditions are observed and possible reasons are discussed. Testing in adiabatic mode provides new insights into the sulphide self-heating process that complement the test designed to identify propensity to self-heat. Keywords: Self-heating; Stage A/Stage B; transition; rhombic/monoclinic elemental sulphur; adiabatic heating

Jan 1, 2020

By C. Chingwaru, W. Chingwaru, J. Vidmar

"Zimbabwe is endowed with rich deposits of minerals such as diamonds, platinum, coal, uranium, lithium, gold, antimony, iron, and chrome. Bioleaching has been implemented as an efficient and low-cost method to extract metals such as copper, cobalt, and gold from sulphide and/or ironcontaining ores and mineral concentrates in a number of countries around the globe. Zimbabwe, despite being a world leader in mineral wealth, has gone through years of economic stagnation which have brought with them energy shortages. Bioleaching is an innovative way to recover minerals from ores using relatively low-capital-cost and non-polluting technology. Principally, iron- and sulphur-oxidizing bacteria can be used to oxidize iron and sulphide to ferric iron and sulphuric acid, respectively, and the ferric iron oxidizes the insoluble metal sulphides to soluble metal sulphates that can be readily recovered from solution. Although some minerals such as gold are inert to biological reactions, they can be liberated using bacteria that act on certain types of ores and other minerals that co-occur with these minerals. The geology of the mineralized areas in Zimbabwe, rich in chalcopyrite / pyrite, allows a number of microorganisms to be used for the extraction of minerals by bio-oxidation. This paper reviews the potential of bioleaching in the country. IntroductionZimbabwe has rich deposits of minerals including diamonds, platinum, coal, uranium, lithium, gold, antimony, iron, and chrome. Bioleaching is defined as the use of microorganisms to facilitate the extraction of metals from sulphide or iron-containing ores or concentrates (Rawlings, 2004, 1997). Bioleaching has been shown to be an efficient (Blowes, Ptacek, and Jurjovec, 2003; Nordstrom and Southam, 1997; Schippers, 2004) and low-cost (Schippers et al., 2010) method to extract minerals in many countries around the globe. For example, aerobic Fe(II)- and sulphur-oxidizing bacteria have been shown to be up to two orders of magnitude more efficient in the oxidation of pyrite than the chemical processes (Blowes, Ptacek, and Jurjovec, 2003; Nordstrom and Southam, 1997; Schippers, 2004). Additionally, bioleaching can be used to extract minerals from low-grade ores, sulphidic waste rock dumps, and tailings efficiently (Schippers et al., 2010)."

Jan 1, 2017

By L. Petzold

The demand for stainless steel is ever-increasing. As an essential element in stainless steel-making process, so is the demand of chromite concentrate. At the same time, the industry proven technologies for beneficiation of chromite are limited. Chromite is only crushed, screened and jigged to produce a concentrate. However, for ultrafine and low-grade material, no satisfactory beneficiation methods for the enrichment of chromite fines exist in industrial operation. This poses a clear challenge for the industry, which Primetals Technologies has addressed with the development of a new beneficiation process. The aim of this process was to enrich low grade chromite tailings to a quality suitable for the smelting process. Required concentrate grades are different over the world, concentrate grades of Cr2O3=40% and lower are found at the Bushveld Complex, South Africa. The target for the present study was a Cr2O3-grade in the area of 49%. Three different chromite tailings with Cr2O3 grades between 16 % and 29 % were part of a study, executed by Primetals Technologies. All three tailing materials were characterised by Mineral Liberation Analysis (MLA). Based on the physical, chemical and mineralogical findings, process flowsheets were developed. The process flowsheets included grinding to ensure a full liberation of the chromite. The liberated material was classified at a cut size of 63 μm to a coarse and fine fraction. Gravity separation and flotation processes were tested with the coarse particle fraction. Due to the limitations of gravity separation with fine particles, only the flotation process is suitable and was tested with the fine fraction < 63 μm. A Design of Experiment (DoE) with six factors, each three levels, was conducted to identify optimal parameters for the fine chromite flotation. Two different collectors were tested: a phosphoric acid ester and a fatty acid type collector. In small scale test work a chromite concentration suitable for the smelting process was achieved. On this basis Primetals Technologies will develop a pilot-scale test program to research recirculation effects and upscale potential of the process. Keywords: Mineral processing, tailings beneficiation, flotation, density separation, process development, chromite

Jan 1, 2020

By I. A. Mwamba

Titanium aluminide of composition Ti-47.5 at.% Al was prepared by melting titanium and aluminium of commercial purity with additions of precious metals of 0, 0.2, 1, 1.5 and 2 at.% content. The microstructure, hardness and oxidation properties were assessed. It was found that at 47.5 at.% Al, the microstructure of plain TiAl alloys was duplex, consisting of duplex lamellar grains and ? grains. The lamellar grains consisted of alternating lamellae of ? (TiAl) and a2 (Ti3Al), with overall aluminium content in the range of 43?47 at.%. The ? grains had a aluminium content above 50 at.%. The addition of precious metals to the TiAl alloy resulted in the formation of a new phase, with a high precious metal content, occurring mostly in the ? phase, without any major change to the microstructure. Increasing the amount of precious metals resulted in an increase of the new phase amount, at the expense of the ? phase. The hardness of the different alloys was around 300HV10. Overall, the research has shown that addition of precious metals to TiAl resulted in a slight increase in hardness and a significant improvement of the oxidation resistance. This shows that alloying TiAl with precious metals, at the addition levels investigated, does not alter the mechanical properties very much, but improves oxidation resistance.

Jan 1, 2012

By M. P. Brandt, T. P. Knoetze, I. I. M. Kessler

It has generally been accepted that the volatile mater in a coal is the factor that determines whether or not a specific coal dust is an explosion risk. This paper presents the results of an attempt to relate the explosibility index (Kex) of South African coals, as determined in a AD-litre explosion vessel, to the volatile mater content of the coal and the calorific value of the volatiles. A good relationship between the calorific value of the volatiles and the explosibility index was established, which makes it possible to calculate the violence of an explosion, as expressed by Kex, for South African coals. The Kex value is but one of a number of parameters that influence the explosion hazard in a specific colliery and cannot be used as the sole predictor of the propagation of an explosion underground. Other parameters include the ignitability of the coal, temperature, humidity, dispersibility of the dust, methane concentration, amount of incombustibles present, and ventilation.

Jan 1, 1993

By M. R. MacIver, M. Pawlik, S. Finora, H. Alizadeh

The dispersion of mature fine oil sands tailings (MFT) by the addition of sodium lignosulfonate was studied using contact angle measurements, optical scattering, and particle size measurements. Lignosulfonate caused the release of residual bitumen to the slurry surface and settlement of fines. Dynamic contact angle by captive bubble measurements were carried out on bitumen to determine the effect of the dispersant on the wettability of bitumen. Lignosulfonate caused a change in the aggregation state of the MFT and increased sedimentation of particles, as indicated by optical scattering measurements. Zeta potential measurements indicate that this change was caused by an increase in adsorption of lignosulfonate onto particles, which caused improved settling of the colloid. Ultraviolet Visible (UV) spectrophotometry and organic carbon content measurements showed that only a small amount of the reagent remains in the aqueous phase of the tailings, which is critical for decontamination of recycled water. Keywords: Oil sands tailings, bitumen recovery, dewatering, dispersion, surface chemistry

Jan 1, 2020

By C. Birch

"Mineral reserve statements as well as optimizing mine planning require a block model with grades for each mining block. A cut-off grade is determined by considering the mining costs as well as the expected revenue for each block. These grades are estimated using various techniques, but in reality there is uncertainty in the grade values. These uncertainties result in two types of error. A type I error is where material is classified as ore and mined, but the true value is below the cut-off grade and the material is therefore waste. This material constitutes dilution. The type II error is where material is estimated to be below the cut-off grade and is classified as waste, whereas the true grade is actually above the cut-off grade. This material is not mined and the value is lost.This research considers the value of the lost ore and the costs of dilution under various degrees of uncertainty. Simulation using @Risk and mixed integer linear programing (Excel Solver) is used in a financial optimizer model to maximize either profit or net present value. This is applied to four Witwatersrand tabular gold deposits to investigate the impact of block grade uncertainty on cut-off grades.When optimizing for profit, value may be added by adjusting the cutoff grade slightly downwards. When optimizing for NPV, value may be added by lowering the cut-off grade significantly for the lower discounting rates. At higher discount rates, the lowering of the cut-off grades should be reduced, and in some cases an increase in cut-off grade may be required. Each mine reacted differently to the optimization and thus there is no single rule that can be applied across all tabular Witwatersrand gold deposits."

Jan 1, 2017

By Q. Cui, B. Z. Zhou, Z. G. Song, Y. G. Zhu, G. B. Zheng

The complex refractory Cu-Pb-Zn polymetallic ore in Xinjiang, China is a massive sulphide ore with complex mineral components. In the past two years, the ore mineralogy has changed, resulting in a decline in metallurgy, with too much zinc reporting to the lead and copper concentrates. A detailed mineralogical study indicating the presence of secondary copper minerals, which could be activating the sphalerite early in the present circuit, was used to develop a novel reagent and flotation flowsheet termed “Pb-Cu co-flotation, Cu intensified flotation, Zn flotation”. It was tested and found to significantly contribute to improving the flotation performance of the ore. In the Pb-Cu co-flotation circuit, successful depression of pyrite and sphalerite was achieved using a BK612- Na2S based depressant system. Most of the galena and some of the Cu bearing minerals with excellent floatability were collected with the use of the selective collector BK906. The rest of Cu bearing minerals were floated in a copper Z200 circuit using a BK537-lime-zinc sulphate based depressant scheme. Zinc flotation from the copper tailing was performed in a SIBX circuit using lime as a pyrite depressant. Using the proposed flowsheet, overall Pb, Cu, and Zn recovery increased significantly to 88%, 87%, and 90% respectively. Keywords: Complex polymetallic ore, Cu-Pb-Zn, Pb-Cu co-flotation, particle size

Jan 1, 2020

By E. Tabosa, S. Vianna, K. Duffy, W. Valery, B. Andrade, L. Pyle, P. Holtham

Pneumatic flotation cells are high-intensity flotation machines with rapid kinetics, resulting in short residence times. Examples include the Jameson, PneuFlot and ImhoFlot cells. These cells do not use impeller-like mechanical cells. Rather, the air and pulp are mixed in a continuous stream of high fluid velocity through a downcomer or venturi to disperse the air into fine (approximately 0.5 mm) bubbles and maintain particle suspension. Froth wash water is often used to enhance concentrate grade, minimising recovery by entrainment. With the growing application of these types of cells, not only in cleaner stages (their traditional use) but also in pre-flotation and scalping duties, there is a need to predict their performance more accurately for circuit design and optimisation. No widely accepted model for pneumatic flotation cells is available and their metallurgical performance is often scaled up from laboratory and/or pilot plant testing by assuming fixed values for metal recovery, concentrate grade and concentrate percent solids. This paper describes a new approach to simulate pneumatic flotation cell performance and presents case studies demonstrating its applicability. Keywords: Pneumatic flotation, flotation modelling, flotation circuit design, flotation circuit optimisation, Jameson cell

Jan 1, 2020

1. TRAINING IN ENGINEERING WORK FOR ENGINEERS IN TRAINING 1.1 Registration as engineer in training An engineering graduate who has passed an examination recognised by the Council1 is eligible for registration as an engineer in training in terms of the Professional Engineers&apos; Act2. Application forms can be obtained from the Registrar.

Jan 1, 1979

By A. E. Lewis, J. Du Plessis, J. Chivavava

Current metallurgical processes lack selective and efficient REE recovery from aqueous leach liquors. Antisolvent crystallisation is used to selectively recover pure inorganic salts from aqueous solutions without excessive chemical use in ambient conditions. Yttrium sulfate octahydrate (Y2(SO4)3.8H2O) was used as a representative rare earth salt to investigate the effect of antisolvent selection on the crystal morphology and size distribution of the final crystalline product produced through antisolvent crystallisation. The antisolvents of choice were short length alcohols that are water miscible i.e. methanol, ethanol, 2-propanol and t-butanol. All experimental parameters except the type of antisolvent were kept constant under isothermal conditions. The large reduction in solubility of the salt by the antisolvent creates high supersaturations. This leads to high product yields as observed in this study at > 91%. In general, a variety of morphologies were successfully formed in each individual alcohol including; rods, plates, prismatic cubes and starburst clusters. The powder X-ray diffraction (XRD) analysis indicated that Y2(SO4)3.8H2O was the only species present for all final products. There was a change observed from the starburst and rod structures for the shorter alcohols towards prismatic cubes for the longer alcohols due to different strengths in the solute-antisolvent interaction at the crystal growth faces. The crystal sizes increased with an increase in the alcohol chain length because of the lower supersaturation conditions due to a limit in the mass transfer induced by the bulkier alcohol structure. In conclusion, it has been shown that the particle morphology and size can be greatly affected by the selection of antisolvent even under high supersaturation conditions, contributing to understanding the effect of different antisolvents on the final particle characteristics. The engineering of favoured product characteristics can therefore be achieved by antisolvent crystallisation, presenting a viable alternative process for the recovery of REE. Keywords: Antisolvent crystallisation, rare earth elements, morphology, particle size

Jan 1, 2020

By J. Du, V. Shukla, C. R. Kumar, U. Shivangi, D. P. Chakraborty

During the processing of iron ores, a large quantity of low-grade tailings containing abundant iron values in fine particle size range is produced. Due to the increasing demand in the high-grade ores for metallurgical operations and the stringent environmental conditions imposed on mining activity, it is essential to utilise the waste tailing pond slimes. Recovery of iron values from these tailing ponds not only enhances the life of the existing operating mines, but also enables achievement of a sustainable process. Iron ore tailing/slime (-150 μm) is produced after scrubbing and wet screening of a low-grade iron ore. Tailing beneficiation of iron ore has been addressed either by flotation or wet high intensity high gradient separation. The present paper aims to recover iron values from discarded slime containing assay ~56% Fe, ~6.5% SiO2, and ~7.0% Al2O3 from the Noamundi processing plant by using a high gradient magnetic separator (HGMS). In this process, tailings are effectively processed with a vertical pulsating high gradient magnetic separator to produce a relatively high-grade iron concentrate. The results of investigation indicate that this novel process is capable of producing an iron concentrate assaying ~ 60% Fe and ~3.25% alumina with a yield of ~55 % on weight basis and recover > 65 % iron bearing particles from the tailings. More importantly, the corresponding tailing contained ~45% Fe. It was concluded that this novel process may be effectively used to recover iron values from low-grade hematite tailings. Keywords: Iron ore slime, Fe recovery, HGMS, yield, Noamundi

Jan 1, 2020

By V. Sheikhzadeh, M. R. Sheikhzadeh, A. Haratian, M. Zare

High Pressure Grinding Roll (HPGR) as an efficient grinding technology plays a major role in mineral processing plants. Investigating the effects of this technology on different parts of mineral processing circuits reveals this technology’s capability. In this paper, HPGR grinding effects on an iron ore beneficiation circuit, in comparison with conventional crushing systems were studied. A representative feed sample was taken from an iron ore beneficiation plant, then crushed to minus 6 mm in a laboratory by two methods: HPGR and cone crusher, and finally a concentrate plant circuit was simulated. Plant simulation was performed by doing laboratory tests consisting of grinding, magnetic separation, flotation and filtration on HPGR and cone crusher products, respectively. According to comminution tests results, in particle sizes smaller than 6 mm, a cone crusher product is coarser than an HPGR product. Also, the Bond ball mill work index of the HPGR product was 10% less than the crusher product. Therefore, based on the Bond formula in applying HPGR, ball mill power consumption decreased by 12%. After ball milling to minus 100 microns, magnetic separation tests results showed that the recovery of the cone crusher circuit output was 90.8% and the product Fe grade was 66.4%. For the HPGR circuit these figures were 87.6% and 67.3%, respectively. Therefore, by using an HPGR, a higher Fe grade of the final concentrate was achieved. A flotation test was then performed on the magnetic separation concentrate, where the results showed that the crusher circuit recovery and the sulfur grade of the product were 96.5% and 0.192%, respectively; whereas the HPGR circuit recovery was 97.2% with a sulfur content of 0.185%. Findings revealed that flotation performance in the HPGR circuit was more efficient than the crusher circuit. Vacuum filtration tests on concentrate produced through the HPGR circuit resulted in a moisture content of 12% that was 2% higher than the crusher circuit’s concentrate in the same condition. Keywords: High pressure grinding roll, power consumption, magnetic separation, flotation, filtration

Jan 1, 2020

By M. H. Dehghanpoor, M. Zivdar, M. Torabi

"A hydrometallurgical route was developed to extract metals such as gold, silver, copper, and selenium from anode slime from Sarcheshmeh Copper Complex in Iran. For the extraction of copper by leaching, the anode slime was heated and stirred with 4 M sulphuric acid in the presence of oxygen. Optimal conditions for recovery of copper were temperature 80°C, mixing time 5 hours, and oxygen flow rate 2 L/min, resulting in a copper recovery of 94.47%. Silver and selenium were dissolved by leaching anode slime from which the copper had been removed with nitric acid. For gold extraction, anode slime was heated with aqua regia at 90°, and the gold then extracted with 2-ethyl-hexanol at a concentration of 0.5 M and phase ratio of 4. The extract was then stripped with soda solution at pH=10.5 and a phase ratio of 5. Finally, gold was recovered by precipitation with oxalic acid at a ratio of mass to volume 10 g/L. Total gold recovery was 80%. IntroductionDuring the electrorefining of copper, some metals are accumulated in the byproduct anode slime, which is an important source for recycling and recovery of these metals (Petkova, 1994). In this paper, a hydrometallurgical route is described for the recovery of precious metals from copper anode slime of Sarcheshmeh Copper Complex, which is located in southeastern Iran .Copper anode slime is black or gray powder with a particle size less than 200 mesh. Approximately 2 to 20 kg of anode slime is produced per ton of copper cathode. The slime is characterized by higher amount of Ag, Se, Pb, and Cu compared to other metals and a very low Au content (Khaleghi and Ghader, 2013).The copper anode slimes are composed of 11.69% selenium, 3.93% copper, 5.26% silver, 0.13% gold, and 45% barite, plus lead, tellurium, antimony, sulphur, and silica with small amounts of nickel, iron, zinc, and bismuth. Gold is one of the most important noble metals due to its wide applications in industry and economic importance, so recovery of gold from slime is more appealing than that of other metals. According to the World Gold Council (2013), the demand for gold has increased during the last decade. This increase is due not only to the increasing demands of the jewellery market, but also to the increasing demand for gold in industrial as well as medical applications (Ranjbar et al., 2014)."

Jan 1, 2016

By D. Weijie, S. Kai, Y. Hu, X. Ning, M. Shiyi, W. Junpeng, Z. Guihong, L. Xianjie

Flotation is one of the most important processes in the copper production industry. The copper grade in the concentrate is a pivotal indicator of the flotation process performance. However, it is difficult to achieve real-time measurement through hardware sensors. Based on the production data provided by an industrial operation, a novel soft sensor based on multi-layer perceptron (MLP) is proposed for effective monitoring of this pivotal indicator. The proposed soft sensor uses the nonnegative garrote (NNG) to perform global variable selection for MLP, while a local search (LS) approach is incorporated to improve the model presented by NNG. Simulation results show that the proposed algorithm has higher prediction accuracy and better model simplicity than other algorithms. Furthermore, the variables selected are consistent with the field experience, and the developed model can provide reference of feedback control for the optimisation of the process. Keywords: Flotation process of copper ore, nonnegative garrote, local search, soft sensor, multi-layer perceptron

Jan 1, 2020

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 J. de Korte, Z. Loubser

"SynopsisThe aim of a blending stockpile is to minimize the natural variation in properties of a material. Two simulation models were written, for coneshell and chevcon stacking, and used to investigate the influence of different stockpile parameters on blending efficiency. The coneshell model is used to show that blending efficiency is highly sensitive to stacked height, with lower output variances observed for higher capacities. Output variance in both models also decreased when the stacker movement increment was decreased. For chevcon stacking, increasing stacker speed significantly increases the blending efficiencies. Increasing the length of the blending tail also reduces output variation, but this parameter would need to be weighed against the accompanying decrease in buffer capacity.IntroductionEven in the most uniform deposit of materials in nature, material properties will vary. The degree of variation can be measured, and usually resembles a normal distribution. The aim of blending and homogenization operations is to minimize the standard deviation of this normal distribution (De Wet, 1994). In general, the reduction in variation will result in a more efficient and cost-effective downstream process. Cement plants need a precise blend of raw materials to ensure desired kiln performance; boilers used in power generation are optimized for fuels of a constant specification; and even though the separation density in a coal washing plant can be adjusted, a constant product quality cannot be produced from highly variable feed (Denny and Harper, 1962). The process efficiency, product quality, and environmental compliance depend on consistency of characteristics in the material fed (Kumral, 2005) Many industries involved in"

Jan 1, 2015

By I. Mc Dougall

Water-cooled copper tap-hole blocks provide an effective means of extending the life of tap-hole refractories, as the lower temperature of the refractory material inhibits wear and other damage and promotes the formation of a protective frozen slag layer. This is particularly beneficial during prolonged tapping cycles. The copper block forms an integral part of the lining, and provides mechanical support to the refractory, facilitating rapid replacement of tapping channel bricks without disturbing the surrounding lining. Safety considerations are an important design factor, and the positioning of the water channels is optimized to prevent leaks although the copper is not in direct contact with the molten fluids during normal operation. Water-cooled copper blocks designed by Tenova Pyromet have successfully been utilized on metal, matte, and slag tap-holes for a number of processes, including nickel, platinum group metals, and mineral wool.

Jan 1, 2014

By S. Sekhar, S. Bhargava, A. K. Godugu, J. L. Porathur

Cemented backfilling has enabled underground hard rock mines to extract orebodies more safely with improved ore recovery. It is important to estimate the minimum required strength parameters for a backfill mix by optimizing the binder percentage to enable cost-effective and safe stoping operations. We conducted three-dimensional numerical modelling to study the effect of various stoping parameters on the stability of a free-standing backfill wall. A Mohr-Coulomb material model was used for the backfill material and the rock-fill interface. A strength reduction technique, excluding friction angle, was employed to arrive at a minimum stable strength value for the backfill. For a given combination of strength values, the stability state of the backfill wall could be demarcated using a displacement and yield zone tracking method. The numerical modelling results are compared to some earlier theoretical models. From the simulations, a predictive equation is developed to predict the required strength parameters for a backfill mix to ensure a stable free-standing wall. Examples are given of the successful use of the predictive equation at some underground hard rock mine sites in India.

May 2, 2022

By E. Maritz

"The localized uniform conditioning (LUC) technique was proposed by Marat Abzalov in 2006. The technique converts conventional uniform conditioning (UC) grade-tonnage curves into single grade values attached to each smallest mining unit (SMU). This is achieved by ranking the SMUs within a panel in increasing order of their grade (based on the direct kriging of SMUs). This ranking is then used to localize the conventional UC grade-tonnage curves for each panel by dividing them into classes and computing their mean grades, which are assigned to the SMUs. The quality of this localization process will depend heavily on the validity of the grade patterns predicted by the direct kriging of the SMUs. Abzalov noted that where the distribution of data available for the direct kriging of the SMU is characterized by strong short-range variability, the advantages of using the LUC approach may be more limited. Consequently, a study was undertaken to determine how valid the predicted grade patterns of a typical Birimian-style gold deposit (with high nugget effect and strong short-range variability) might be expected to be. This was determined by comparing the direct SMU kriging ranking (based on sparse data) with the grade control model ranking (based on close-spaced data and the best available estimate of the deposit). The results showed a satisfactory correlation between these rankings and it was concluded that, although the grade patterns predicted by the direct kriging of the SMUs may be less meaningful for deposits exhibiting strong short-range continuity, there is nevertheless a convincing relationship with the actual (or best available) rankings. Therefore, the LUC technique is still considered to be useful for this style of deposit. IntroductionFor adequate technical and financial evaluation of a project, attempts should be made to estimate the recoverable resources – the portion of the in situ resource that can be economically extracted by mining. To achieve this, the estimates of the tonnage and grade of the mineralization should be produced above a given economic cut-off and should take into consideration the proposed mining selectivity."

Jan 1, 2016