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By J Stojanovic, V Kasic

Gold-bearing polymetallic sulfide deposits are common in Serbia, and mostly occur in the Karpato-Balkan and, to a lesser extent, the Serbo-Macedonian metallogenic province. These deposits û Coka Marin (Majdanpek), Grabova Reka (Blagojev Kamen), Lece (Medvedja) and Veliki Cip (Boranja) û are dominated by Cu and/or Pb-Zn mineralisation that contains a variety of Pb, Zn, Cu and Fe (¦ Sb, As, Bi) sulfides and sulfosalts, as well as native elements. Gangue minerals are composed of quartz, carbonates and rare silicates and barite. Based on their gold content these deposits can be divided in three groups:1. Polymetallic deposits with a high gold content (>1 g/t).2. Polymetallic deposits with an intermediate gold content (0.1 to 1 g/t), and3. Polymetallic deposits with a low gold content (<0.1 g/t). Gold occurs as native gold or as an alloy with silver. Native gold ranges in size from 5 to 20 ¦m, with a few rare grains over 100 ¦m, as isolated grains or as microveins (in the first group of deposits). Its distribution within the deposit in uneven. Native gold in the second group is smaller in size (1 - 5 ¦m, rarely 20 ¦m), where it is uniformly distributed in a quartz-silicate matrix. The third group is of no economical importance. The subject of research has been the first group of polymetallic deposits.

Jan 1, 2008

The Nena high sulphidation-type copper-gold deposit is composed of andesitic pyroclastic units with minor interstratified lava flows. Epiclastic sandstone and conglomerate, and siliciclastic rocks consisting of mudstone, shale, lithic and bioclastic sandstone and conglomerate are locally present. Pre-mineral diorite intruded and domed the volcano-siliciclastic sequence. Economic copper-gold is hosted by vuggy quartz altered lapilli tuff and tuff breccia. The early hydrothermal alteration stage involved deposition of epithermal banded quartz and chalcedonic silica in localised environments. This alteration stage was locally superimposed by a hydrothermal breccia event. Following the breccia event, magmatically-derived volatiles were channelled through a NW striking dilational fault zone and generated the high sulphidation alteration. The high sulphidation alteration is zoned from a vuggy quartz core with local massive replacement silica through a broad pyritic quartz-alunite halo and finally to a thin illite-smectite zone. This is surrounded by regional chlorite-epidote-smectite alteration. A gangue mineral phase consisting of quartz, alunite and baryte overprinted the zoned high sulphidation alteration and preceded the main copper-gold-arsenic mineralisation. Early sulphide deposition is characterised by marcasite, arsenical pyrite and melnicovenite. Main-stage copper-gold mineralisation involved sequential deposition of hypogene bladed covellite+chalcocite, enargite-luzonite and luzonite+stibioluzonite respectively. The copper sulphides are zoned from bladed covellite-chalcocite in the north (cross-sections 5000 - 5600 N) through enargite-luzonite (cross-sections 4700 - 5000 N) and to luzonite????stibioluzonite in the south (cross-sections 4400 - 4700 N). A late phase of minor acicular covellite and enargite+luzonite overprinted the zoned copper sulphides and associated gold mineralisation. Gold mineralisation commenced at the transition from enargite to luzonite, and continued through the deposition of luzonite-stibioluzonite in both stages. Most gold was deposited during the main-stage mineralisation. Yellow native sulphur was deposited in the remaining vugs and defines the waning stage of the hydrothermal system. Post-mineral dioritic intrusives and feldspar porphyries intruded the volcano-siliciclastic sequence. Finally, supergene chemical processes with associated erosion and uplift generated a 40 - 50 m thick oxide-hosted gold cap. The oxide gold cap is underlain by a 35 - 45 m thick blanket of supergene chalcocite+covellite mineralisation which overlies and occurs throughout the main hypogene copper-gold zone. Supergene gold is related to the distribution of primary gold-bearing sulphides. Therefore, remobilisation and redeposition of gold are considered to be insignificant and very localised.

Jan 1, 1999

By Monger N. R

Factors leading to the decision to use Joy Continuous Miners for pillar extraction in the Burwood Colliery are outlined. A detailed description is given of the Joy 1-CM-2 Continuous Miner, including the power supply to the machine, and the method of cutting coalfrom a face. Modifications found necessary in practice are dealt with.The Joy 10 S.C. shuttle car used in conjunction with the Miner, and alterations made to this unit, are then described, together with Ancillary units for the whole installation such as transformers, motor generator set, air compressor,_ cables for power supply and water supply for dust suppression.Special arrangements were necessary to transport the Miner and Shuttle Cars through the tunnel to the Victoria Tunnel seam, and the method of overcoming the transport difficulties is discussed. Manning of the Continuous Miner unit, and final preparation andsubsequent working of the pillars is outlined. Special attention must be given to maintenance of the machines.It is concluded that the combination of Joy Continuous Miner and Shuttle Car is an excellent method for removing pillars in the Burwood Colliery.INTRODUCTIONThis paper deals with the Burwood Colliery in the RH.P. Group in which some units are engaged in the mechanical extraction of pillars with Joy Continuous Miners.In 1941, an amendment to the Coal Mines Regulation Act of N.S.W. forbade the extraction of pillars by mechanical means unless with the express permission of the Minister for Mines. As this permission was not forthcoming, no mechanical extraction of pillars took place in Australia until 1954 As it was doubtful during most of this period whether mechanical extraction would ever come about, coal generally was won with a view to maximum extraction in the first workings, thus giving rise to centre gannon bord panels and 45° angle cut throughs as shown in Fig. 1.

Jan 1, 1957

By Atalla YM

Steel rolling mill product finishing pro cesses gradually become or often already are unsuited for current production and marketing conditions. Indications are made as to what has been done to (i) rectify situations of this kind and, (ii) to reduce both product finishing time and costs. The principal features of two separate case studies are reviewed, these being (a) the finishing-end of B.H.P.'s Newcastle o.2 Merchant Mill and, (b) a mixed functions finishing yard and warehouses group (known as he Merchant Mills Loading Yard) which has been perating for a number of years. The areas in ach case are identified as a system while the arious functions, principal product streams nd processing equipment are respectively iden- ified as sub-systems and system elements. etails are given of the models used in each Sub-system dependency on the system as a le is traced by means of an information and ision analysis or by time studies and the sonableness of the system elements is deter- ed by computer programming or by simple com- ison methods. The-accumulated data is used derive improved mill finishing-end or Load Yard,functioning concepts according to pro- rules whilst at the same time considering

Jan 1, 1975

By McDonald G. G

In the deformation of a metal or a single-phase alloy all the crystals composing the metal after their shape in a similar manner as the metal in bulk and to about· the same extent. This is not the case in an alloy consisting of two phases of different resistance to plastic deformation. In such an alloy the two phases take part in the deformation approximately in inverse proportion to their hardness, at least at the beginning of the deformation. (1) The hardness difference becomes less in the course of the deformation and thereby the deformation more equal. The phenomena occurring during cold-rolling of some two-phase alloys have been investigated by H. Unckel who made statistical grain measurements of the degree of deformation of the two phases after various reductions by rolling and also microscopical observations of the flow in the neighbourhood of the grain boundaries. Because of the different relative flow of the two phases, the normal flow is disturbed in the vicinity of their boundaries, and some additional plastic flow must take place in these regions, the disturbance being levelled out at some distance from the boundary. The deformation of the two phases present simultaneously in the material should thus be different from that of each phase separately. The investigation of theimportance and the effects of this interaction of crystals of different phases is the subject of this paper.The deformation of crystals occurs as a gliding of crystal layers following crystallographic laws. In consequence of this slip the crystals composing the material tend to assume a common orientation, i.e. some crystallographic axes tend to become parallel to each other in all the crystals. The type of this preferred orientation depends on the type of the crystal and the working process. Any change in the mechanism of deformation results, therefore, in a change in the preferred orientation. Such a change in kind and degree of preferred orientation will, therefore, also occur if the interaction of crystals in a two-phase alloy gives rise to a mechanism of deformation different from that in the free phases.The present paper describes a study of the preferred orientation, the microstructure and hardness in wires of brass drawn to various extents.The development of preferred orientations in the two phases with increasing reduction was followed up in various layers of the wire.EXPERIMENTAL DETAILSThe experiments were carried out with a commercial 5 S.W.G. wire (0,212 in. diameter) containing: 62·7% Cu, 37·3% Zn, very faint traces...

Jan 1, 1941

Following extensive trials on a 100 000 tpa Research and Development facility at Kwinana, Western Australia, Hismelt Corporation (wholly owned by Rio Tinto) are confident of the future commercial success of this technology. The Hlsmelt process uses iron ore fines and non-coking coals to produce molten pig iron suitable for feeding to BOF and EAF steelmaking vessels. The unique process configuration combines direct injection of raw materials into the molten iron bath, high rates of metal and slag mixing, and effective post-combustion of the bath gases to achieve efficient coal utilisation. This is the key to the strong economic performance of the process.

Jan 1, 1999

By Poi G, Madgwick J. C, Davis P, Pannowitz D, Cho D, MacKay B, Chan S. M, Varga R. A, Chun U

Periodic dilute sulphuric acid treatment ofstill, batch-fed, molassesenriched, microaerobic, manganese-dioxide-Ieaching microbial cultures led to marked improvements over untreated cultures. Percentage Mn++ recoveries increased (7X), rate of Mn++ formation improved (3X) and yields of Mn++ per molasses use were more economical (5 X). The yield in terms of molasses used improved with decreasing sugar used (eg 1.17 g Mn++/gram molasses at 0.2 per cent molasses (w/v) feed and 24 per cent (w/v) pulp density MnO2 tailings). The most effective Mn++ leaching in terms ofpercentage recovery (97 per cent) and rate ofleaching (46 mg Mn++/day) was at 0.5 per cent molasses (w/v) and 14 per cent pulp density (w/v) MnO2 tailings.

Jan 1, 1986

By Madgwick J. C

A new strain of Thiobacil/us ferrooxidans (SPK I) has been isolated which has a higher copper to iron release ratio (2.04) than the standard T ferrooxidans (CPBI) (1.12). The diminished iron oxidizing ability enables almost quantitative release of copper (93 per cent) from a chalcopyrite-rich concentrate without interference by iron reprecipitation. In a single batch leach, Bougainville copper concentrate yielded 9.3 grams/litre soluble copper from a 4 per cent w/v suspension in 100 days at 32°C.Solvent extraction recovery of soluble copper is recommended as a final step in a leaching process.

Jan 1, 1983

Column Bioleaching of a Low-Grade Nickel-Bearing Sulfide Ore Yellow Mountain Deposit belonging to GobiMin Inc is located in the Chinese Tianshan and Altay orogenic belts, Hami, Xinjiang Province, NW China. Indicated Mineral Resources of its reserve are estimated at 12 million tonnes averaging 0.40 per cent nickel. Mineralogical examination and chemical analysis indicated that the Yellow Mountain ore is a type of high Fe, low Mg and low Ni sulfide ore: the main sulfide minerals are ~8.0 per cent pyrrhotite (Fe1-xS), ~1.0 per cent pentlandite ((Fe, Ni)9S8), ~0.6 per cent chalcopyrite (CuFeS2), ~0.5 per cent pyrite (FeS2), and ~0.2 per cent violarite((Fe, Ni)3S4); the main gangue minerals are ~48 per cent olivine, ~28 per cent antigorite, ~3 per cent chlorite, ~2 per cent talc, ~2 per cent tremolite, and~1.5 per cent magnetite; the chemical analysis of the ore sample are Ni 0.40 per cent, Co 0.016 per cent, Cu 0.24 per cent, Fe 12.4 per cent, S 5.6 per cent, MgO 9.6 per cent, CaO 1.80 per cent, SiO2 48.5 per cent and Al2O3 10.3 per cent. Preliminary results of the shake flask test showed that pyrrhotite is easily leached and the high concentration of ferric iron would hamper the growth of bacteria which will bring a negative impact on the bioleaching conversely. This work aims to set up a method to prevent the negative impact of the ferric ion on the bioleaching of the ore. Two method were combined to prevent the negative impact of the ferric ion, they are adaptation of the mixed bacteria to iron ion and bleeding the off-solution periodically to control the levels of iron ion (depending on the tolerance of the microorganisms to the ferric ion) in the leaching solution. Using a mixed mesophiles composed of Acidithiobacillus ferrooxidans, Acidithiobacillus thiooxidans and Leptospirillum ferrooxidans under room temperature, a nickel recovery of 93 per cent and a cobalt recovery of 85 per cent were achieved in 280 days column bioleaching process with an average nickel concentration in the leaching solution of 2.6 g/L. The results showed that the negative effect of the ferric ion can be successfully controlled and the nickel can be bioleached successfully from theYellow Mountain ore.

Sep 13, 2010

In the minerals industry allocating capital for setting up a new project or for upgrading an existing facility usually requires that a compromise be found between the initial cost of equipment, its operating performance and on-going operating costs. To maximise the return on the investment usually requires that various possible combinations and layouts of equipment be assessed. This task is both difficult and time consuming. Computer simulation can be used as a means of determining the expected operating performance of a range of alternate equipment configurations. Simulation is a particularly important technique for analysis as it is one of the few approaches that allows the analyst to take into account the variability inherent in a mining process. While computer simulation is a well established technique its use, until recent years, has been limited to large organisations with access to substantial computing facilities. However, with the current availability of low cost computing platforms combined with easily understood animated displays, simulation is now becoming more widely used across the industry.

Jan 1, 1990

By J Lapwood, S Pal, M Pascoe

The devastating failure of Dam I at the Corrego do Feijao iron ore mine near Brumadinho in Brazil on 25th January 2019 left at least 270 people dead or missing. In the immediate aftermath of this event, many mining companies commenced reviews of their tailings and water storage facilities (T&WSFs), and Rio Tinto was no exception. Rio Tinto has an interest in over 250 T&WSFs across 47 sites in 13 countries as owner-operator, as managing JV partner, and through non-managed partnerships and other ownership arrangements; T&WSF experts from Rio Tinto’s Surface Mining Centre of Excellence visited the sites in order to complete Technical Risk Reviews of each facility. These reviews looked beyond the day-to-day operational aspects of T&WSF management, and the site’s compliance with local regulations and Rio Tinto’s internal D5 Standard – Management of tailings and water storage facilities, in order to provide a strategic, second line of assurance perspective. The reviews focused on design and construction; operational controls; risk control framework; and competencies and capabilities of personnel involved in the design, operation and closure of the facilities. Additional activities were carried out in parallel with these reviews, including the development of a consistent approach to dam break studies; the public disclosure of information on Rio Tinto’s tailings storage facilities and their management; and contribution to external programs such as the initiatives being co-ordinated by the International Committee on Mining and Metals. This paper provides an overview of the tailings and water storage assurance work undertaken by Rio Tinto’s Surface Mining Centre of Excellence Tailings Validation Taskforce to date; work that will continue to improve governance and assurance for all of Rio Tinto’s T&WSF.

Jul 1, 2021

By Dinuzzo P, Wychie J, Desoe C

The Selwyn Gold-Copper Mine is located 150 km southeast of Mount Isa within Proterozoic units of the Mount Isa inlier (Figure 1). Mineralization in the mine area is associated with strongly sheared ironstones and iron rich schists at the base of the mid- Proterozoic Staveley Formation. The ironstone horizons outcrop as a series of dominant ridges over a 10 km strike. Economic gold-copper mineralization has been defined in 4 separate zones within a 3.5 km length of the outcropping ridges. A current combined geological resource for the deposits is 5.3 million tonnes of 5.0 g/t Au and 2.02 Cu. Weathering of the deposits has resulted in a pronounced vertical zonation of mineralogy, especially copper, within the ore zones. This has produced 3 separate ore types which require different treatment techniques. The categorization is based on copper content and zones are termed leached (surface), oxide and sulfide. Ore removed from the pits will be mainly leached and oxidized with minor sulfide near the pits final depths, while the underground ore will be from the sulfide zone with minor oxide in some areas.

Jan 1, 1989

By W Polkowski, M Tangstad, M Wallin, J Jiao

Fe-26Si-9B (wt per cent) alloy has been identified as a potential high-temperature phase change material (PCM) due to its attractive properties, such as its high latent heat of fusion and low volumetric change during solid/liquid transition. For the successful utilisation of this alloy into thermal energy storage (TES) systems, the development of a cost-effective production method is essential. Presently, the Fe-26Si-9B alloy is produced by mixing FeSi alloys with either pure boron element or FeB alloys. However, the use of pure boron is financially prohibitive, and the carbothermic reduction results in high greenhouse gas emissions and high energy consumption in the production of FeB alloys. In this regard, our study proposes a silicothermic reduction method to produce Fe-26Si-9B PCM by using FeSi alloys and B2O3-based oxides. Accordingly, the influence of various parameters on the production process was investigated, including operating temperature, holding time, B2O3 content in the added oxides, and initial Slag/Metal (S/M) ratio. Based on the experimental results, the optimal parameters for producing FeSiB alloys with over 9 wt per cent boron were determined. Consequently, it was documented that the FeSi alloys and the added oxides enriched with 50– 65 wt per cent B2O3 should be subjected to temperatures ranging from 1550–1650°C, maintain an initial S/M ratio exceeding 1, and ensure a holding duration beyond 1 hr. Moreover, the energy consumption of this process was estimated to be ~1.86 MWh/t metal and the mass loss was lower than 7 per cent. Therefore, silicothermic reduction offers a sustainable approach for producing FeSiB alloys with a boron content above 9 wt per cent.

Jun 19, 2024

By E Hernandez, I Agustian, W Kusumohastuti

PT Freeport Indonesia is completing the first underground gyratory crusher excavation at their Deep Mill Level Zone (DMLZ) mine located in the mountains of Jayawijaya, District of West Papua, Indonesia. DMLZ is an underground block cave mine with an expected output of 80 000?t/d and a scheduled production start in 2015.In order to deliver crushed ore to the mill through the existing conveyor belt, the DMLZ block cave mine will construct two 1000?kW 1.6?×?2.3?m gyratory crushers with their own respective ore flow conveying systems underground. Each crusher is composed of top, middle, and bottom chambers with an ore bin. This system includes a series of high capacity conveyors approximately 5 km in length including transfer stations and support facilities.The development methodology required a combination of lateral development by jumbo drilling machines and vertical development by mechanised raise climbers (MRC) due to the complexity of the chamber design.This paper will discuss the project scope, design and development methodology, ground conditions, innovative survey controls, and the challenges faced due to the remote location.CITATION:Hernandez, E, Kusumohastuti, W and Agustian, I, 2014. Excavating the first underground Gyratory Crusher at the Deep Mill Level Zone mine, PT Freeport Indonesia, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 43–50 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

Recovery of tin and copper from concen trates containing cassiterite, stannite and chalcopyrite was studied in laboratory-scale tests designed to simulate continuous operation of the matte fuming process in which tin is recovered as SnS fume while copper accumulates in the matte. Charges made up of 300g of mixtures of the concentrate and siliceous flux were added at a uniform rate in pelletized form to crucibles containing 300g of matte and 50g of slag (both of which were recovered from previous tests), while blowing through the matte with air or oxygen-nitrogen mixtures at a rate of 3 to 5 1/min. Accurate material balances enabled the effects of process operating conditions on tin recovery and sulfur utilization to be deter- mined. When operated correctly, the matte fuming process gave tin recoveries exceeding 90 percent in spite of wide variations in the copper content of the matte and in the operat- ing conditions. Deterioration of tin recovery was associated with undersaturation of the slag with respect to silica, and with high partial pressure of sulfur dioxide (greater than approx- imately 0.2) brought about by oxygen enrichment of the blowing air. The utilization of sulfur varied widely in low grade mattes. Copper in excess of approx- imately 10% promoted formation of SO2 by the commonly accepted reaction: FeS + 1-120 2= FeO + S02 while in copper-free and lower grade mattes a substantial fraction of the oxygen was consumed in the alternative reaction FeS + i 02 = FeO + z S2.

Jan 1, 1980

The Pongkor Gold Mine employs 678 staff. The mine is owned by the Indonesian state company PT Antam Tbk and is located 100 kilometres south west of Jakarta, in Java, Indonesia. Mine and infrastructure development started in 1992 following intensive geological exploration over 15 years. First gold production was in June 1994. The mine comprises three main orebodies: Ciguha, which is nearly mined out, dips at 75¦ from surface to 150 m. Kubang Cicau, dipping at 80¦ from surface to 230 m, and averages 3.5 metres thick at a grade of 10 g/t Au and 128 g/t Ag; and Ciurug, dipping at 60¦ from surface to 300 m and averages eight metres thick at a grade of 14 g/t Au and 172 g/t Ag. Present global ore reserves are 4.6 million tons at a grade of 12.3 g/t Au and 133.6 g/t Ag. Exploration is continuing for new veins. Initial mining was by manual cut and hydraulic fill at Ciguha, followed by Kubang Cicau. In 1996, it was proposed to convert Ciurug to a mechanised cut and fill operation. Two single boom jumbos and three load-haul-dumps (LHDs) were purchased with the assistance of a gold loan from NM Rothschild and Sons. Implementation of this upgrade was delayed by both a national economic crisis and an invasion of the concession by illegal miners. It was not until May 1999 that raise bore vent rises and access drift infrastructure commenced. Major underground development started in 2000 and is continuing. Production cut and waste fill stoping started in July 2001. The change from manual to mechanised methods has required a total change in work practices and attitudes. However, the process has resulted in a production increase from 600 tpd to 1200 tpd. A three-month compliance test (a condition of the financing loan) was successfully completed, from July to September 2001, with an average production of 700 tpd at Ciurug and 500 tpd from Kubang Cicau.

Jan 1, 2002

By T Smith, J Hatley

With the increased focus on the need for more uranium mines, this paper presents the design process currently underway for the development of the Millennium deposit in northern Saskatchewan, Canada. This deposit, located 630 - 750 m below surface, will be accessed by two shafts and will be mined by blasthole stoping methods at less than 500 t/d. Mining grade will approximate two per cent U3O8 and mine life is expected to be six to seven years. The project is a joint venture between Cameco, AREVA and JCU. This paper discusses in detail methods used to predict worker radiation exposures, the methods used to monitor the work environment and procedures unique to the uranium industry necessary to minimise worker radiation exposures.

Jan 1, 2008

By Orr C. W

Continuing the systematic investigation of the influence of elements on the properties of lead(1). If the effects of additions of copper (up to 0.3%) on Port Pirie lead have now been determined in so far as creep rate is concerned.The results in the present paper contain also a survey of the influence of copper on the micro-structure of lead, a feature which has not appeared in the previous papers of this series. Further, it has now become known that creep is but little related to the more usually determined properties such as hardness and tensile strength. Since, however, these properties are of considerable importance in connection with the fabrication and use of this metal. Brinell hardness determinations have also been made and are here included.It is becoming more and more realised also that to be useful, creep data should refer to stresses of the order of those used in practice. This is due to the fact that as soon as deformation takes place, the properties of the metal are automatically changed. Lead does, of course, creep during industrial use, and probably (in lead chamber construction for example) extensions of the order of 5%...

Jan 1, 1938

By J Deraisme, P Epinoux, O Bertoli

"Société le Nickel exploits the Dôme lateritic nickel orebody at its Tiébaghi operations in New Caledonia. The site geology has an obvious bearing on recovery and metallurgical performance, with controls that are extremely complex at all scales, including lithology, oxydo-reduction conditions, mineralogy and multivariate geochemistry. In this context, establishing an adapted recoverable resource estimation method that is efficient and transparent proves an interesting challenge. The method must address the key notions of: support effect – the exploration data set can only warrant the estimation of large panels of 20 × 20 × 3 m3 that are much larger than the selective mining units (SMUs) of 5 × 5 × 3 m3 information effect – the selection at the production stage uses estimates based on information that is much denser than that available for mine planning.The problem is rendered more complex in this type of deposit by the fact that the additive variables used for the estimation (metal accumulation and ore tonnage) are not the ones used to establish the selection at the mining stage (ie the nickel grade of SMUs, which is the ratio of accumulation on tonnage). Eventually, the selection criteria not only involve Ni grades, but also other elements such as Al2O3, Fe2O3, MgO and SiO2.The solution presented in this paper is to construct an SMU multivariate simulation in the saprolitic horizon and construct an efficient post-processing aimed at producing multivariate recoverable resource estimates. The efficiency constraints imposed by the number of blocks to be simulated (1.5 million) lead to resorting to direct block simulations. The underlying multivariate discrete Gaussian model, which had its validity tested beforehand, is put to good use to mimic the selection process that happens at the mining stage by offering the ability to simulate a composite value located at random within each block.The paper also presents the application of a scenario reduction algorithm to pick a representative subset of a few simulations to help appraise the risk attached to the downstream (reserve optimisation and mine sequencing) phases of the project. The original version of the paper (see Deraisme, Bertoli and Epinoux, 2014) used an early version of the scenario reduction plug-in (named S2RM) built in Isatis® software, where measuring the distance between the initial set of scenarios and the reduced set was based on Ni only. The implementation presented here is for the second generation of S2RM, where measuring the distance between the initial and the reduced set of realisations is based on all study variables. The benefits of the measure adapted to the true multivariate nature of the problem are highlighted in this updated version of the paper.CITATION:Bertoli, O, Deraisme, J and Epinoux, P, 2014. Multivariate block simulations of a lateritic-type nickel deposit and post-processing of a representative subset, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 99–108 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 24, 2014

By Creasey J. W, Huntington J. F

Faults cutting the Lithgow and Katoomba seams in some collieries of the Western Coalfield can be classified, on the basis of spatial pattern, into three principal groups: meridional swarms of closely spaced minor faults and associated joints, with throws of less than 0.5 m and with exceptional faults of up to 5 m; isolated interswarm strike-slip faults in the Lithgow seam and widely spaced strike-slip faults in swarms in the Katoomba seam; and low-angle normal faults associated with the Lidsdale-Lithgow seam split.Orientations of faults and coal cleats in the coalfield are presented on stereoplots and histograms, as a basis for an analysis of fault dynamics. The low-angle normal faults were found to have been formed only along the Lidsdale-Lithgow seam split, and before the high-angle faults. The origin of the high-angle normal faults in the meridional swarms is discussed and it is suggested that basementcover interactions must have occurred. This suggestion is supported by geophysical evidence. Low-amplitude monoclines in the coal seam, and the spacing of swarms 0.9 to 4.5 km apart, may be due to the underlying basement structure.It is clear that postextensional compressive deformation of swarms in the coalfield gave rise to shortening by minor thrust development, and probably also caused the synchronous strike-slip movement on pre-existing normal faults. There is evidence to show that isolated interswarm strike-slip faults in the Lithgow seam, and widely spaced strike-slip faults in the Katoomba seam, may have originated synchronously with such thrusts. The axis of maximum principal compression ('71) for the thrust event was inferred to be near horizontal and aligned approximately E-W. Recent rock mechanics measurements of the present day stress field indicate a similar stress configuration.

Jan 1, 1981