Search Documents

By R. B. Khumalo, T. Chauke, T. Mtshali, F. K. Mulenga, B. Makgate, N. M. Chiloane, T. Moyane, V. C. Madanda

While high production rates are important for the profitability of a mine, the influence of geotechnical engineering cannot be ignored. In block caving, a high-stress level is required for cave propagation; this is one of the core elements for ore recovery. Therefore, stress distribution and geological structures play a major role in the planning and layout design of the mine. This paper explores the influence of stress and geological features on the design of the mine from the feasibility study phase. The geology and stress measurement of the mine were determined through logging and overcoring, respectively. This study found that while production rate is the primary operational consideration, the long-term layout design of the undercut face advance, face angle, drawpoint spacing, and undercutting sequence is mainly governed by the stress orientation and major geological structures. Moreover, widely spaced drawpoints result in isolated draw zones, consequently, resulting in ore loss and early dilution whereas closely spaced drawpoints affect the stability of the major apex pillar. This paper also provides insight into how operations can leverage stress distribution to optimise mine layout and safety in massive mining methods.

Nov 24, 2025

By Felix E. Wormser

REGISTRATION started bright and early Monday morning in the Silver Corridor at the Waldorf-Astoria with Mrs. W. H. Bassett as chairman. What a registration-over 200 ladies! Several joined us from as far away as California and Washington State. Apparently they enjoy our meetings so much that present travel inconveniences didn't discourage their coming. At noon the ladies joined the men at the All-Institute Luncheon and Monday evening we had our informal dance, which is separately reported. Tuesday morning was devoted to our annual meeting at which all Section annual reports were read. At noon we were all the guests of the W.A.A.I.M.E. at a luncheon at the Port Arthur Restaurant on Mott St. in Chinatown. Many of us rode down on the old 3d Ave. elevated railway-the last of four. It, too, is soon to be torn down. The luncheon was delightfully unusual, with huge dishes of delicious Chinese food on each table. Our chairman, Mrs. Read, who lived in China in her youth chose our menu, and she chose it well, for most of us indulged so heartily that as we arose from the table we felt we couldn't eat again for 24 hours. Someone donated sherry wine, and boxes of Chinese pencils as favors. We wonder who? All owe a big vote of thanks for this lovely party to Mrs. T. T. Read and Mrs. H. A. Prosser, chairmen, and Mrs. H. W. Hardinge, Mrs. R. Hursh, and Mrs. R. L. Lloyd, vice-chairmen.

Jan 1, 1944

By Sven Petersen, Berit Lehrmann, Bramley J. Murton, Paul A. J. Lusty

"INTRODUCTIONToday’s metal mining focusses on land based ore deposits which only consider one third of the Earth’s surface. Through an increasing global demand of strategic metals used in the electronic and green industry (Zepf et al. 2014), ‘uneconomic’ sites of lower grade in more technical challenging grounds such as greater depth and/or more remote areas are developed by the mining industry often resulting in higher environmental impact. However, considering the steady growth of the Earth population a shortage of certain metals may occur (Krausmann et al. 2009). In 2010, the United Nations allowed commercial exploration for seafloor massive sulphides (SMS) in international waters with currently six contractor licenses being granted.Although the number of discovered vent sites has steadily increased since the first discovery of hydrothermal venting at the Galapagos Rift in 1977 (Corliss et al. 1979) with more than 600 sites being known today, their economic potential is poorly constrained. So far Nautilus Minerals Inc., a Canadian mining company, is the only company having conducted a NI 43-101 resource estimate for the Solwara site, located in the territorial waters of Papua New Guinea. Other studies using bulk geochemical data from 95 sites published in literature do exist (Hannington et al. 2011, Monecke et al. 2016) suggesting a global resource potential for today’s known SMS deposits of 600 million tons with a median grade of 3 wt.-% copper, 9 wt.-% zinc, 2 g/t gold and 100 g/t silver. However, the geochemical data being used originated mainly from easily recoverable, mainly non insitu surface grab-samples and are under current mineral resource classification schemes such as JORC code or NI 43-101 not even accepted for resource estimations. Furthermore surface samples such as high-temperature sulphide chimney and talus material are not representative for the entire deposits with information regards thickness and continuity of individual sulphide layers obtained through drilling being scarce. In addition, it remains unknown what geological processes occur once hydrothermal activity ceases and whether metal tenor becomes enriched, depleted or disappears."

Jan 1, 2017

By C. F. Meyer

The recirculation of mine air in an underground colliery, has always been regarded as a hazardous practice in South African collieries. The Mines and Works Act and Regulations states that fresh air, flowing Into an underground section, must flow out directly into the return airways after being utilized once in the section. One of the largest producing underground collieries in South Africa, Initiated a project on controlled recirculation with the permission of the South African Government Mining Engineer. The project involves the recirculation of 65 m3/s of air inside a ventilation district which serves a total of six production areas. The aim of the project was to increase the airflow to the sections without altering the existing ventilation system. Tracer gas studies were carried out to determine the actual distribution of the recirculated air to the individual areas. The results showed that an effective 54,18 m3/s of the recirculated 65 m3/s reaches the sections. This gives an air utilization figure of 83,4 per cent. Dust surveys were also carried out as part of the project to determine the effect of the recirculation of the return air on airborne dust levels. The results showed that the dust levels remained constant in the main intake and return airways and that no hazardous situation was being created. The project proved to be successful and showed that controlled recirculation can be executed and used as a solution for the supply of air to sections without endangering any lives.

Jan 1, 1993

By Wolfgang Linden

The emerging demands on higher steel grades call for improvements in the operation of production units and a corresponding increase in automated processes, and therefore international steel producers require viable functions. The avoidance of slag carry-over to secondary metallurgy processes is essential for achieving steel grade qualities and reducing alloys in order to compensate slag residues in the downstream unit. The automatic tapping procedure in BOF converters will be supported by the implementation of reliable measurements used in high-end steel plants. Automatic tapping will help to achieve the following: Reduced slag carry-over, improved yield, increased reproducibility, improved working environment and safety and operation with minimum possible crew. The automatic tapping procedure is capable of being implemented in new or retrofitted installations.

Aug 17, 2017

By Frank F. Aplan

In looking into the crystal-ball scene of minerals beneficiation for 1970 and beyond, an impartial observer becomes con- fused as to what has been done and what is possible in the way of improvements. The mining official must balance the burden which is being shouldered by the mineral industry for maintaining our environment in its pristine state on one hand, with the solid advancements made in technology and cost savings on the other hand. Then the official wonders if this is really the criterion for judging overall performance today. As a prospective engineer, Dave Bolles sees it this way: "Specialization has caused many engineers to subordinate everything to doing their job the cheapest and fastest way. In the future this won't be enough. The engineer in the mineral industry will be responsible not only for mining, concentrating and smelting, but also for the resultant wastes- the mine dumps, the tailing dams and the slag piles." (See Drift on page 63).

Jan 1, 1971

By J. A. Putnam, A. D. K. Laird

In laboratory research on the behavior of oil, gas and water in porous materials, no direct method has been devised to measure saturation without disturbing the flow. Indirect methods involving various forms of radiation have been developed for measuring one or two components. The X-ray absorption method for two components has been described previously1-3. The possibility of three component measurement has also been mentioned1. Since two component measurement still requires considerable care to insure accuracy, the fact that three components can be measured satisfactorily is not generally appreciated. The fraction of the energy of an X-ray beam of one wavelength transmitted through a material is given by the transmission factor. in which 1 is the beam's path length through the fluid. p is the fluid density, and 1,. is its mass absorption coefficient. Eq. 1 implies that all parts of the beam pass through equal amounts of the fluid. Practical applications are, therefore, limited to systems in which the fluid is distributed so that I varies negligibly over the cross-section of the beam. Typical variations of the transmission factor with X-ray tube potential arc shown in Fig. 2 from which experimental points have been omitted for clarity. The X-ray absorption of low pressure gases in a core is practically zero so it cannot he distinguished from that a vacuum by instruments that will differentiate between liquids. The transmission factor of a gas in a core is. therefore, unity. Consequently, the saturation of only one gaseous component can be found. If two fluids completely fill a core, their two saturations add to unity and, therefore, only one value of the transmission factor, T, is needed to determine both of them. When two liquids do not completely fill the core, their saturations do not add to unity and transmission factors at two wavelengths must be measured. At one of these wavelengths, 7 for one liquid must be different from T or the other liquid. The change of T for one liquid us the wavelength is changed between the two values must be different from the corresponding change of T in the other liquid. The two values of fluid saturation will also give that of the gas because the three saturations must add to unity. In Fig. 2 it can be seen that there are many large differences of T between various oil solutions and water solutions at any one potential, and between potentials for any one solution. Consequently, conditions for optimum accuracy of measurement of the saturations of a gas and two liquids can be chosen. Once the operating potentials have been selected, however. it is more convenient to cross plot the data from Fig. 2 on transmission factor-solution concentration coordinates for the chosen potentials, as shown in Fig. 3. Fig. 3 was used to plan the experimental procedure for Core N-38-1. Since the voltage control at 45 kv was good, the slopes of the curves in Fig. 2 at this potential would cause little error. At low potentials, however, the voltage control was not good, so 33 kv was chosen because the small slopes of the curves made voltage control less critical. The lowest strength beam that could be measured reasonably at 33 kv corresponded to a minimum T factor of 0.385. Fig. 3 shows that both 4.80 weight per cent of cadmium chloride in water and 24.0 per cent of iodobenzene in crystal oil had this T value, and, consequently, were indistinguishable to the X-rays at 33 kv. Thus, at 33 kv, the liquid saturation was given by the same curve regardless of the relative amounts of oil and water solutions present. This simple calibration curve from which the gas and liquid saturation can be read, is shown in Fig. 4. Fig. 3 also shows that at 45 kv, T for the aqueous solution is 0.610. and that T for the oil solution is 0.220. These values were transferred to Fig. 4 to give the triangular calibration plot from which the brine and oil saturations can be found, after the gas saturation has been determined at 33 kv.

By A. L. Engel

This publication describes further research by the Federal Bureau of Mines in beneficiating domestic manganese ores. The ores selected for treatment were typical of many manganese deposits in California and Nevada in that they have not been economical to mine because they are low grade or contaminated with siliceous minerals and are not amenable to physical concentration.3 4/ It would be desirable to develop a simple treatment method resulting in nearly complete extraction of manganese from these ores and applicable to small-scale operations. Sulfation reduction appeared to be such a method. The studies reported in this publication were directed toward sawdust or lignin-sulfonates as reducing agents. Sawdust, as waste from lumbering, and lignin-sulfonates, byproducts of woodpulp papermaking, are comparatively cheap and may be available in areas where manganese ores are found. Sugar, oils, and coal were also tried as reducing agents.

Jan 1, 1960

By H. E. Powell

The selective recovery of nickel and zinc from a waste phosphate solution by solvent extraction was investigated by the Bureau of Mines as part of a broad program to recover valuable metals and compounds from industrial wastes, and to reduce the associated land and water pollution caused by dumping these wastes. The waste, which resulted from the phosphation of automboile bodies, yielded 99 percent of both the nickel and the zinc by extraction with dinonyl naphthalene sulfonic acid (DNSA) and di-2-ethylhexyl phosphoric acid (EHPA), respectively. Dilute sulfuric acid was used to recover the metals and regenerate the solvents.

Jan 1, 1970

By R. D. Brown

"AbstractOver the past five years, tax rates in the Canadian mining industry have skyrocketed upward across Canada, as both federal and provincial governments have sought to increase their share of mining revenues.However, the past few years have not only brought high tax rates to the industry, but also a number of important new federal and provincial tax incentives.In large part, these incentives ~ special credits or reduction in tax for carrying out specific exploration, processing or other activities - have helped to moderate the adverse effects of the higher tax rates and have provided inducements for exploration and processing activities which might not otherwise have been undertaken.However, the effectiveness of certain of these incentives can be questioned: our over-all tax structure now consists in large part of punitively high marginal rates with some extraordinarily generous but capricious tax reductions for specific activities and may not be in the best long-term interest of either the industry or the nation.A rational tax structure, with moderate and stable tax rates, together with reasonable incentives confined to activities specifically requiring them, could be a more effective spur to investment than the present tax and incentive jungleIntroductionIN THE YEARS since federal tax reform, the mining industry in Canada has faced a continued escalation in its tax burdens as federal and provincial governments have leapfrogged each other in imposing new and higher taxes on the industry. The result has been that marginal tax rates imposed on mining and processing activities throughout Canada have jumped to hitherto unparalleled levels."

Jan 1, 1977

By Martin A. Elliott

The Bureau of Mines began its investigation of the use of Diesel engines underground to provide technical information necessary to assist this development in the United States. It was considered desirable to furnish such assistance because successful European developments/ in this. field had indicated that Diesel mine locomotives might be safer than electric trolley locomotives. The Bureau has recognized for some time the importance of eliminating the hazards associated with underground haulage systems using bare trolley wires. For example, each year 30 to 40 fatalities occur as a result of electrocution by contact with bare trolley wires. In 1943 and 1944, 101 men were killed and extensive property damage occurred as a result of mine fires started by falls of roof grounding the bare trolley wires.

Jan 1, 1948

By J. W. Langhans, B. D. Blake

The U. S. Bureau of Mines has been investigating the application of thiosulfate to recover gold directly from low grade refractory carbonaceous ores for the past four years. Statistical experimental methods were used to determine the effects of various leaching parameters on both gold extraction and thiosulfate consumption. Seven variables were investigated, including thiosulfate, copper, ammonia, sulfite, and sulfate concentrations, leaching time, and leaching atmosphere. This information was used in face-centered cubic (FCC) surface response experiments to define the optimum leaching conditions for a low grade ore. Two models were generated from the data to predict gold extraction and thiosulfate consumption as a function of the leaching parameters. The models predicted a gold extraction of 67 pct and reagent consumption of 2 kg [S2032/mt ore at 0.25 M S2032-, 0.0014 M Cu2+, 0.27 M NH,OH, 0.00625 M SO;-,] and leaching for 16 h. Actual tests resulted in 62 pct gold extraction with 0.93 kg S2032- consumed/mt ore. Subsequent column leaching tests have further demonstrated the viability of applying this technology to recover gold from carbonaceous ores without first pretreating the material. Gold extractions in the column tests ranged from 69 to 78 pct with thiosulfate consumption ranging from 0.4 to 5.4 kg/mt ore. These results compare to <5 pct gold extraction using standard cyanide leaching conditions.

Jan 1, 1996

By K. C. Corin, B. McFadzean, J. G. Wiese

The Bushveld Complex of South Africa contains almost 90% of the world’s reserves of platinum group minerals (PGMs). In the flotation of PGMs, there are significant challenges arising from the need to treat ever-decreasing grades of the relevant ore deposits. The major challenge in the flotation of these ore bodies is the control and management of the gangue minerals, particularly silicates such as orthopyroxene and plagioclase feldspar which are often rimmed with talc which makes them naturally floatable. It has been shown that various polysaccharide depressants such as CMC and guar have different properties in terms of depressing the gangue minerals. Since the PGMs are often associated with sulphides, copper sulphate is widely used as an activator in PGM flotation but can inadvertently activate the gangue minerals and reduce the recovery of PtTe2 which accounts for up to 40% of the Pt in the Platreef ore body. Depressants also reduce the mass of solids reporting to the froth and can thus destabilize the froth. This effect on the froth can be mitigated by using higher frother dosages or water of higher ionic strength. In summary due care needs to be taken to carry out site test work to develop an optimum ratio of collector, frother, activator and depressant to ensure that the highest grades and recoveries of the PGEs are obtained while reducing depressant dosage as much as possible. Chromite recoveries can be reduced through the application of gravity separation or reducing entrainment through reduced water recovery. INTRODUCTION The Bushveld Complex located in the north of South Africa contains the world’s largest deposit of Platinum Group Minerals (PGMs) (1). The so-called upper Critical Zone of the Complex hosts the largest concentration of platinum group elements (PGEs) in the world. The Zone hosts the Upper Group Chromitite No.2 (UG2) reef and Merensky Reef as well as the Platreef mineralization of the northern limb of the Bushveld Complex. The Merensky pegmatoid Zone contains the base metal sulphide grains and associated PGMs. Of particular importance to the topic of this paper is the occurrence in these reefs of various gangue minerals. The Merensky Reef consists predominantly of orthopyroxene (~60%), plagioclase feldspar (~20%), pyroxene (~15%), phlogopite (5%) and occasional olivine. Secondary minerals such as talc, serpentine, chlorite and magnetite have widespread occurrence. Of these, talc is the only mineral to show significant natural floatability (2,3). Although present in small amounts talc is closely associated with pyroxene and imparts significant floatablity to the high concentrations of pyroxene present in these ores (4). The base metal sulphides consist predominantly of pyrrohtite (~40%), pentlandite (~30%) and chalcopyrite (~15%). The major platinum group minerals are cooperite (PtS), braggite [(Pt,Pd)NiS], sperrylite (PtAs2) and PGE alloys. The UG2 consists predominantly of chromite (between 60 – 90% by volume) with lesser silicate minerals (5 – 30% pyroxene and 1 – 10% plagioclase). Other minerals present in minor concentrations, can include silicates, oxides and base metal sulphides. Total PGE varies by locality but on average range between 4 and 7 g/t. The PGMs present in the UG2 Reef are highly variable, but generally the UG2 is characterised by PGE sulphides such as laurite (RuS2), cooperite (PtS) and braggite [(Pt,Pd)NiS. The base metal sulphides consist predominantly of chalcopyrite, pentlandite and pyrrhotite. The Platreef is located in the northern limb of the Bushveld Complex. This reef consists of a complex assemblage of rock types, with pyroxenites, serpentinites and calc-silicates being the most abundant. Base metal mineralization and platinum group element (PGE) concentrations are found to be highly irregular, both in value as well as in distribution. The predominant platinum group minerals (PGM) in the mined area of the Platreef are the PGE tellurides, alloys, arsenides and sulphides. T

Jan 1, 2019

By C. Leticia Hernández, P. Martín Reyes, D. Iván Alejandrox Reyes, L. Miguel Pérez, T. Julio Cesar Juárez, G. Mizraim Uriel Flores, T. Raúl Moreno, R. Aislinn Michelle Teja

This study shows the characterization of tailings from the mining district of Coscotitlán, State of Hidalgo, Mexico. Chemical analysis by Spectroscopy of atomic absorption (AAS) allowed to quantify the concentration of Fe, Mn, Zn, and Pb as the most abundant elements in the mineral sample; in addition to the presence of Sn, Te, Bi, Ni, Co, and Cd. The presence of precious metals as the Ag (64 g ton−1) and Au (0.65 g ton−1) were identified. The greater distribution of Ag was detected in particles with diameter of 74 lm. Species as pyrite (FeS2), galena (PbS) and hematite (Fe2O3) were observed in a matrix of quartz (SiO2) by Polarized Optical Microscopy (POM), these species were also identified by X-ray Diffraction (XRD). Both the elemental and mineralogical composition of tailings was corroborated by the technique Scanning Electron Microscopy - Energy Dispersive Microanalysis (SEM-EDS).

Mar 1, 2018

By Batric Pesic

The interactions of mineral-pyrite with T. ferrooxidans were studied by using a cyclic voltametry technique. The interactions with bacteria were examined during the entire growth stage of bacteria (fermentation), The pyrite ?surface invariably drastically changed &apos; its. properties at the second day of fermentation. Beyond two .. days of fermentation, the cyclic voltamograms were insensitive to convective diffusion produced by stirring. The product layer was examined by SEM,. X-ray diffraction and chemical analysis. The SEM study revealed that bacteria populated the pyrite surface at an extremely high density levels. The high density of bacteria, and the solid reaction products formed on the pyrite surface created conditions for pore diffusion which explained why the CVs became .insensitive to convective diffusion in solution (stirring). The x-r:ay diffraction study confirmed jarosite as a product layer. A mechanism of T. ferrooxidans cells serving as nucleation sites for jiarosite formation is proposed.

Jan 1, 1991

By J. Girard, V. Aubé, M. Barakate, S. Bellec, M. R. Eddahabi, Y. Chaik, M. Deshaies

"The activated carbon used in the carbon-in-leach (CIL) process to absorb gold is subject to breakage. This carbon breakage can be minimized, but it will always be necessary to deal with fine carbon particles, known as carbon fines. These fines contain a significant amount of gold and mining operations generally struggle with the treatment of this product. Since IAMGOLD’s Essakane mine is located in Burkina Faso, shipping carbon fines out of the country was not considered a viable solution due to the logistical burden and legal procedures in place. A benchmarking study demonstrated the complexity of treating carbon fines on site. IAMGOLD partnered with the host government to find an innovative approach to economically recover the gold locked in these fines. This paper describes the new circuits commissioned in July 2016 built to improve carbon fines collection, incinerate the carbon fines and leach the gold in the ashes produced by incineration. These new circuits successfully demonstrate that it is possible to extract more than 94% of the gold content in the carbon fines. This is profitable for the company as it improves the overall mill recovery rate by 1.0% to 1.5% and significantly decreases gold inventory build-up on site.INTRODUCTION The activated carbon used in the carbon-in-leach (CIL) to absorb gold is subject to breakage. This breakage can occur in the CIL tanks, the stripping circuit, the regeneration process or anywhere throughout the pumping and handling processes of carbon through the circuit. Fine carbon particles are invariably created and, because of their small size, present a large specific surface which favors gold absorption (Hilsher et al., 2014). At the very beginning of operations at Essakane, the carbon fines were recovered by means of a thickener and a filter press, but the circuit was undersized and inefficient. Subsequently, a decantation basin was used to settle the fines before the thickener. The settled fines were shovelled from the basin into bags. All these bags were stored in the backyard of the plant. The average tonnage of carbon fines recovered monthly was 23.5 tons with a gold grade at 1938 g/t."

Jan 1, 2017

By A. Tordoir, S. S. Kanchibotla, W. Rogers, E. Engmann, B. Bisiaux, S. Ako

Blast-induced rock mass displacement can have a significant impact on ore loss and dilution. The mischaracterization of the grade boundaries both prior to, as well as a result of, blasting can lead to significant financial losses due to ore loss and dilution. Ore dilution occurs when waste material is miscategorized as ore and sent for processing or stockpiling, diluting the run-of-mine head grade and recovery. Ore loss takes place when valuable mineral is miscategorized as waste and sent to the waste dumps, thus never being processed and recovered. Extensive research has been conducted at the University of Queensland over the last fifteen years to understand the impact of blast movement on ore loss and dilution. Recent outcomes of this research have led to the development of innovative tools and techniques to measure and model blast movement. These techniques have been implemented at the Ahafo gold mine in Ghana to understand blast movement dynamics and to implement solutions to reduce blast-induced ore loss and dilution. These solutions are validated through controlled trials and have been incorporated into site standard operating procedures.

Jan 1, 2012

Editor?s note: Each year, the June issue of Mining Engineering features an industrial minerals review. Several people put in a fair amount of time in developing the material for this issue, all the while doing their own jobs. Thank you to the industrial minerals annual review editor, to the Industrial Minerals and Aggregates Division technical committee chair and vice-chairs, and to the authors of the individual commodity profiles. Jim Norman, member SME, of Tetra Tech, is the industrial minerals annual review editor. Candace Trimble, member SME, of Oil-Dri Corporation of America, is chair of the Industrial Minerals and Aggregates Division?s Technical Committee.

Jan 1, 2012

By G Byrne

Long-term water management is typically ignored in mine closure planning. A review of 73 mine closure plans dating from 2007 to 2013 shows that, on average, relinquishment is proposed to be achieved in 11 years. A separate review of 57 mines in actual closure shows that only five have so far achieved relinquishment with an average closure period of 14 years after production cessation. The remaining 91 per cent of sites in closure have an average closure duration to date of 21 years - and still going.A common issue among those sites in closure is water management, with a further common issue being acid and metaliferous drainage (AMD). Of the closure plans that were reviewed only one site had allowance for long-term (&gt;50 years) of water treatment. Of the sites in actual closure 25 per cent have reported that they expect such long-term water treatment, with many acknowledging that in perpetuity treatment is probable.Clearly there is a gap between what is being allowed in closure plans and the reality of closure. There are a number of factors leading to the optimism of the closure plans, including unrealistic assumptions, pressures to reduce closure cost provisions and the technical complexity of many AMD issues.Long-term water management is a reality that needs to be acknowledged and addressed. If addressed early in the mine life cycle the impacts for mining companies can be reduced.CITATION:Byrne, G, 2013. Long-term water management - the forgotten legacy of mine closure, in Proceedings Water in Mining 2013 , pp 163-168 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

The Rio Tinto Iron Ore group currently estimates grade attributes for mineral resources using ordinary kriging (OK). OK produces a smoothed globally unbiased estimate, but does not capture local variability, which can have significant impact on mine planning and scheduling. The dimensions and the multivariate nature of iron ore deposits make point simulations of whole deposits impractical. With a direct block simulation method the size and speed limitations of point simulations can potentially be overcome, resulting in a viable method for capturing local variability and for providing risk measures into the mine plan and schedules. Some iron ore deposits are highly folded (post mineralisation) and require preprocessing using an unfolding algorithm to properly define the continuity parameters and directions of anisotropy before simulation can proceed. Block co-simulation proceeds in unfolded space and since mine planning require a block-model with regular block-sizes; a regular block-model is unfolded first. This paper discusses a block co-simulation approach that can be used to simulate the irregular blocks resulting from the unfolding process. A section of an iron ore deposit is used for testing the block co-simulation algorithm and the simulation results are compared against the Isatis direct block simulation method. Available blasthole data is also used to validate the behaviour of the block co-simulation algorithm.

Sep 26, 2011