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By B. J. du Plessis

Automation of the conventional grinding circuits of Palabora Mining Company (PMC) required that adequate instrumentation and a control system had to be put in place, after which extensive testing had to be done to evaluate different grinding control strategies. The objective was either to optimize the particle size from the grinding circuit, or to increase milling circuit throughput while maintaining particle size at existing levels. Prior to automation of the process as described below the circuits were controlled manually. No on-line information on cyclone feed densities, cyclone feed flow or cyclone overflow particle size was available. This type of control relied heavily on interpretation of the process parameters by the operator and on his ability to make informed decisions on what the set points should be. An integral component of the automation process was the installation of a SCADA package (SysCad SDX) for implementing the on-line control strategy. A self-correcting model for prediction of cyclone particle size was developed and implemented on the SDX. The model is calibrated using values from the particle size analysers while they are operational. If the particle size analysers choke or fail for any reason, the model continues to provide a good estimate of the particle size. This technique ensures that a continuous particle size indication is available compared to the intermittent reading from the instrument. The continuous signal allows particle size to be used as a process variable for control purposes. A strategy was developed and implemented which allows for the production of a controlled grind size from the conventional milling circuit to the flotation circuit. The grinding control strategy allows for the control of particle size from the individual cyclone overflows. This strategy allows the operator of the process to increase or decrease production rate by specifying a coarser or finer grind size required for flotation.

Jan 1, 2001

By R. J. Davidson

Arguments are presented in favour of a new interpretation of the chemistry pertaining to the dissolution of gold, when this takes place in typical gold plant leach circuits where calcium is present in solution. It is suggested that under these conditions the standard chemistry, as it applies in the absence of calcium, is no longer valid. Thus the concept of very soluble sodium and aurocyanide ionic species in solution is replaced by the presence of weakly soluble calcium aurocyanide, which is a covalent coordination complex that is non-ionic. Supporting evidence for this is provided and the implications that the presence of calcium in gold plant circuits may have for operational efficiency are discussed. In particular, because of the formation of this weakly soluble calcium complex, the addition of lime to the leach circuit is questioned. Arguments for the use of a sodium hydroxide/sodium carbonate/sodium bicarbonate buffer solution as a means of controlling pH and of increasing ionic strength in a so-called ?soft? leach are presented. Such a lime-free system, besides benefiting the leaching of gold, would have many downstream benefits in existing CIP circuits. Furthermore, the direct electrowinning of gold from the leach now becomes potentially feasible.

Jan 1, 2007

By M Eggers, H Baxter

This paper discusses the geotechnical condition of bedding shears and small- and large-scale faults and shears in Archean banded iron formation (BIF) bedrock, based on observations of core and televiewer images. Information that can be collected includes the orientation of defects, infill thickness, strength and material type, defect shape and surface roughness.The spacing and continuity of bedding shears may be critical to slope design where performance is structurally controlled with bedding dipping out of a wall. The intensity of development of the bedding shear is a useful indicator of the amount of shearing a bedding surface has been subjected to. This can allow for engineering judgement to be used in adopting peak or residual shear strengths along defect surfaces when developing the design.The spacing and continuity of bedding shears may also contribute to the blockiness and degradation of a rock mass where the structural orientation is otherwise favourable. Additionally the condition, thickness and strength of infill of major fault structures can affect slope design and performance, depending on orientation.Comparison of the condition of structures observed in core with the corresponding in situ televiewer image has been made to assess what characteristics are most accurately captured by each technique and whether the additional information gathered from using the televiewer technique is justified by the cost. The question of whether the data collected by a televiewer is sufficient to ultimately replace the need for core is considered.CITATION:Baxter, H and Eggers, M, 2013. Fault and shear characterisation in archean banded iron formation bedrock - a comparison between core and televiewer data collection techniques, in Proceedings Iron Ore 2013 , pp 47-56 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By A Proudman, J Isles, R Lawolo, G Forster

PT Nusa Halmahera Minerals (PT NHM) is an Indonesian joint venture company owned by Newcrest Mining Limited (75 per cent) and PT Aneka Tambang (Persero) (25 per cent). PT NHM operates two underground gold mines, Kencana and Toguraci at the Gosowong mine site in North Halmahera District, North Maluku Province, Indonesia.Due to several major stope overbreak events at Toguraci, the mining method and sequence was reviewed. The method was changed from a bottom up Avoca stoping approach using waste rock as fill material and leaving sill pillars, to a top down open stope method using paste fill without sill pillars.The revised mine plan originally had stoping progressing from the current level to the base of reserves followed by extraction of the sill pillars in the upper levels at the end of mine life. Geotechnical modelling was conducted on the revised mine plan and it was recommended that the sill pillars be taken early due to the risk of them overstressing and being difficult to recover if left until the end of mine life. The backfill material above the sill pillars is unconsolidated run-of-mine development waste rock or unconsolidated rock fill (URF).Due to earlier attempts to extract sill pillars under URF not being successful in the upper areas of the mine, it was stipulated that the URF above the five remaining pillars must be consolidated prior to stoping. Cost benefit analysis showed this to be justifiable due to the high-grade nature of the ore locked up in the pillars.Grout was initially planned to be injected in one pass through self-drilling anchors (SDA) in a 2.5 m pattern. Initial attempts to do this failed due to the SDA aperture and grout consistency. A modified method was employed using poly pipe installed into open holes that were drilled through the sill pillar and into the URF from below. Repeated application at each location ensured full grout consolidation. Drill holes were used to confirm grout thickness as well as add more grout. Design thickness was based on stress analysis.Despite different URF types being present due to varying quantities of fines, the process was similar for each area. The presence of paste in the URF of one target affected the quantity of grout it received.With a design thickness of 4–5 m of consolidation in the URF, depending on stope width and dip, each consolidated pillar (to date) maintained integrity during the stope operation. Some fall of material was anticipated; however, quality assurance / quality control (QA/QC) during grouting, staged stope extraction and careful blast design minimised fall of material.CITATION:Proudman, A, Isles, J, Lawolo, R and Forster, G, 2017. Rock fill consolidation for the extraction of sill pillar stopes at Toguraci gold mine, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 199–210 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By N Hungerford, A Coote, S Doyle, D Henderson, F Della-Pasqua

A systematic, detailed airborne geophysical programme carried out in the Taupo Volcanic Region (TVR) in 2005 by Glass Earth NZ Ltd culminated in the identification of 21 epithermal gold targets. In this paper we present the follow up ground-truthing results from the first target drilled, Tahunaatara (TAH).   The Tahunaatara target was identified by means of a multi-disciplinary integrated approach (ôdata interventionö) that defined and ranked the target as being highly prospective for epithermal gold mineralisation. A follow-up ground resistivity survey over the target area revealed the presence of a ôresistorö buried beneath an extensive cover of recent volcaniclastic material. Geochemical soil sampling carried out in conjunction with the resistivity survey yielded anomalous gold values.   The host unit at TAH consists primarily of flow banded rhyolitic lavas, volcaniclastic deposits and tuffs of the Ohakuri Ignimbrite with small outcrops of silicified bedded breccias forming local topographic highs. Results from the first orientation drill hole show wide zones of intense silicification, clay alteration and mineralisation. Sulphide mineralisation is disseminated throughout the core in association with hydrothermal replacement and also occurs in association with quartz veins and infill in fractures (sulphide-only veinlets as opposed to quartz + sulphide veinlets). Textural evidence from drill core indicates that hydrothermal fluid flow was multi-staged with potential for multi-stage gold, and other elements. The presence of polymictic hydrothermal eruptive breccias indicates a hydrothermal vent system.   Alteration varies from argillic to silicic, with pervasive adularia and vuggy silicification also present. Chlorite is minor and associated with quartz, illitic clay and adularia. Assay results currently available confirm the presence of anomalous gold values in association with zones of vuggy silica deposition and intense silicification.

Jan 1, 2006

By Emad Iskander

The seepage forces generated around pressurized water tunnels in rock are investigated using the analysis tools of geomechanics. The theoretical background of the two-phase material behavior is explained and the equations of motion and continuity are presented. Applications on two case studies typically occurring in practice deal with the required minimum rock cover for a pressurized tunnel and the separation between two tunnels carrying different heads. The rock mass behavior as described by the Mohr-Coulomb and Hoek-Brown material models is investigated by numerical analyses and the results are compared. The effect of key material parameters on the safety factors and the associated modes of failure is discussed.

Jan 1, 2008

By C. H. Schack, K. C. Dean

An investigation was made to devise and evaluate a flexible flotation method for beneficiating typical tactite types of calcareous scheelite - bearing ores ; these scheelite - bearing ores are available in large tonnages in deposits containing 0.2 to 0.3 pct tungsten trioxide . The objective was to achieve a high recovery of the tungsten in concentrates , containing a minimum of 10 pct WO3 , that would be suitable for chemical processing into high- quality synthetic scheelite . The principal problems encountered and resolved by bench- scale flotation tests on ore samples from five different deposits were to define the conditions and reagent combinations essential for ( 1 ) achieving selective flotation of a small amount of scheelite from a large amount of calcite and ( 2 ) controlling the effect of small amounts of calcium , magnesium , and iron salts in the ores and water used . The flotation process was evolved from batch tests and continuous runs in a pilot - scale mill with a capacity of 100 lb per hour . The flotation process comprised ( 1 ) removal of sulfide minerals , ( 2 ) conditioning the residual ore pulp with sodium carbonate , quebracho , and trisodium phosphate , and ( 3 ) rougher flotation and single - stage cleaning of a scheelite concentrate . In a pilot plant run , over 91 pct of the tungsten was recovered from ore containing 0.27 pct WO3 , in a concentrate containing 11.2 pct WO3 .

Jan 1, 1964

By G. S. Rice, J. W. Paul, M. W. VONBERNEWITZ

This bulletin abstracts and reviews essential details of reports on 59 fires in different coal mines in the United States, describes the circumstances of origin and the methods of controlling or extinguishing the fires, and points out the lessons taught. all, the fires caused great damage to property and killed nearly 400 men. The data have been compiled from the records of the Bureau of Mines, which cover most of the major mine accidents in the United States since 1.908. The names of the mines concerned and of the persons connected with the management of those mines have been omitted in accordance with the usual policy of the bureau. The bureau hopes that this bulletin will be of service to all persons interested in lessening the fire hazard at coal mines and to all who may be called on to fight a mine fire.

Jan 1, 1927

By B. Cousins

Copper, nickel, zinc and other ?penalty? metals are found with gold together in nature, and have a similar chemical affinity for cyanide used in leaching operations4. As a result, solutions that come from the heap leach pads contain a significant concentration of penalty metals along with the desired gold. In solution, gold and cyanide form the complex Au(CN)2- that is readily adsorbed onto activated carbon. Penalty metals as individual metal ions also load onto activated carbon and, depending on the concentration of cyanide and the pH of the solution, can form various cyanide complexes. Some of these complexes can also load onto carbon while some are significantly inhibited from loading. Copper, for example, can form the dicyanide complex Cu(CN)2-, the tricyanide complex Cu(CN)32-, and the tetracyanide complex Cu(CN)43-. At a low pH and a low concentration of cyanide, the dicyanide complex predominates. At a higher pH and a higher concentration of cyanide, the tetracyanide complex predominates. Without excess cyanide, however, individual copper ions can remain present. Activated carbon is used in gold extraction because it is highly selective for gold and less so for other metals, such as copper3. However, copper has the ability to adsorb (?load?) onto activated carbon and block the adsorption of gold when it is in the dicyanide form and in a high enough concentration. When copper is in the tetracyanide form, copper loading is significantly inhibited, allowing for the maximum gold recovery. In order to ensure the tetracyanide complex is formed, a combination of both available cyanide and elevated pH must be present. In the South Area of Newmont Mining Corporation?s Gold Quarry mine in Carlin, Nevada, copper and nickel loading becomes a problem when certain acid generating zones of the heap are leached. Copper, nickel, zinc and other penalty metals become leachable and the lower pH uses up more of the cyanide and creates carbon loading conditions for the metals. This decreases the carbon?s ability to extract the gold from solution, decreasing the recovery and increasing the concentration of gold in the barren solution that leaves the carbon columns. In this study, operational data from March through to May of 2007 is presented to show how Newmont developed a quick fix to lower pH and elevated metal concentrations that were causing low gold loadings on activated carbon at their South Area Leach (SAL) heap leach operations. Laboratory tested were also conducted to examine the chemical kinetics of metal loading and the individual effects of increased cyanide concentrations. THEORY Copper I ions form four different complexes with cyanide in solution: the monocyanide species CuCN, the dicyanide species Cu(CN)2-, the tricyanide species Cu(CN)32-, and the tetracyanide species Cu(CN)43-. The species are formed from the following reactions: [Cu)(?+-+ K1 (1) Cu ß2 (2) --+?+23)(3CNCuCNCu ß3 (3) --+?+34)(4CNCuCNCu ß4 (4) ] where K1 equals 1019.50, ß2 equals 1024.03, ß3 equals 1028.65, and ß4 equals 1030.35. Along with these complexing reactions, hydrogen cyanide and water also dissociate in solution, as per the following reactions: [H?+-+ KH (5) OHOHH2?+-+ KW (6)] where KH has a value of 109.24 and KW has a 1014 value2. These six chemical equations compete with each other, with domination of one over the others dependent on the CN- and H+ concentrations. Each complex also has different properties when it comes to its loading behavior on activated carbon. The dicyanide copper complex and the monocyanide copper complex have an affinity for adsorption onto activated carbon that is greater than the affinity of gold cyanide for adsorption. They also have a lower solubility in water, which enhances adsorption. The tricyanide copper complex and the tetracyanide copper complex are water soluble and have an affinity for loading adsorption on activated carbon that is less than the affinity of gold cyanide for loading adsorption1. There are three prime reasons for this phenomenon. First, the extra cyanides mask the metal ion more, which blocks the main adsorption characteristic of the complex. Second, the complexes are also larger, making it harder to physically fit into the active sites on the carbon. This helps make the latter two complexes kinetically inhibited from adsorption, freeing carbon active site to load with gold cyanide complexes. Third, it is also important to consider the expressions of the mass and charge balance. These balances describe all the overall concentration of copper and cyanide in solution, as well as marking overall charges of the ions in solution4. This is important because significantly changing the charge from positive/neutral to more negative is another reason for activated carbon to reject adsorbing the higher cyanide complexes. The speciation of copper and other penalty metal cyanide complexes in solution is dependent on the pH. At lower pH values, increased hydrogen ions (H+) can out-compete metals for cyanide ions. Lower free cyanide in the system leads to the lower cyanide species of copper cyanide predominating, with a noticeable concentration of the free metal, monocyanide and dicyanide species of copper cyanide. At higher pH values, hydrogen ions are significantly reduced, reducing the kinetics of HCN formation. More free cyanide remains in the system, allowing for more complexing. With this higher amount of cyanide in the system, the tetracyanide species of copper cyanide predominates, also with a noticeable concentration of the tricyanide species of copper cyanide. Higher copper concentrations, however, require more

Jan 1, 2009

By V Morand, W R. H Ramsay

Structural, mineralogical, and textural data from the Tarnagulla Goldfield demonstrate that the area has undergone three significant phases of deformation. The first phase of deformation (D1) predates the intrusion of the Tarnagulla pluton, which has been dated using K/Ar as 391 ¦ 5MY. We regard this pre-granite deformational event as the major regional deformation. Gold-bearing quartz veins have been deposited in dilational sites on west-dipping reverse faults. These sites demonstrate a geometry related both to faulting associated with D1 and the plunge of the F1 fold axes. This evidence, together with chronology of unmineralised ? D2 and D3 faults, the absence of quartz veining associated with D3 events, and the deformation of carbonate-quartz-biotite-sulphide aggregates within Sl by later S2, lead us to conclude that the bulk of the mineralisation was essentially syn- or late D1 . d18O values for fluids in equilibrium with vein quartz from Central Victoria have indicated equivocal support for either a magmatic or metamorphic origin. Our evidence supports a pre-magmatic origin for mineralisation at Tarnagulla and hence tends to favour a metamorphic fluid source.

Jan 1, 1995

Up to the end of World War n, flat-back cut and fill stoping with or without square sets was the universal method of ore extraction. Rapid changes in practice occurred during the post-war years; first, ore extraction was fully mechanized; secondly, the stope design for new levels, notably No. 17 Level Zinc Corporation Ltd. and No. 19 Level New Broken Hill Consolidated Ltd., was modified to suit this mechanization; and thirdly, sub-level and shrinkage stopes were developed in orebodies where cut and fill methods would be uneconomical. The following is a brief survey of the changes that have occurred-most of which were inaugurated during the last five years.ROCK DRILLINGDuring 1951 open stoping was investigated in detail. In order to reduce steel handling problems, forged steel onepass detachable bits were introduced, but were soon superseded by the "Swedish method". Investigation showed that when using a bar and arm mounted 3-in. drifter the overall drilling time was approximately 6 min. per ft. compared with 4 min. per ft. using the light jackhammer-airleg combination with 1-in. hex. by 4-in. collared shank tungsten carbide-tipped steel.Although competitive with tipped steels on a cost per foot drilled basis (excluding miner's time), detachable steel bits proved most unsatisfactory for hard drilling conditions.Rapid blunting necessitated frequent bit changes and thus caused lost drilling time. This, together with rapid gauge...

Jan 1, 1957

By Bengt H. Fellenius

THIS PAPER deals with the analysis of results from axial testing of vertical single piles, i.e. the most common field test per-formed. Despite the numerous tests which have been carried out and the many Papers which have reported on such tests and the analysis thereof, the understanding of pile test loading in current engineering practice leaves much to be de-sired. The reason for this is that the engineers have concerned themselves with mainly only one question: "Does the pile have a certain least capacity?", finding little of practical value in analysing the actual capacity and the pile-soil interaction. This Paper aims to show that engineering value can be gained from elaborating on a pile test ? during the actual testing in the Held, as well as in the analysis of the results. The first portion of the notes make use of an earlier Paper by the author (Fellenius. 1975). However, additional views and recent literature have been incorporated.

Jan 1, 1988

By A. M. Milev

A simulated rockburst was conducted underground at Kopanang Mine, previously called Vaal Reefs No. 9 Shaft. The rockburst was simulated by means of a large blast detonated in solid rock close to a crosscut sidewall. The simulated rockburst involved the design of the seismic source, seismic observations in the near and far field, high-speed video filming, a study of rock mass conditions such as fractures, joints, rock strength etc. Knowledge of the site conditions before and after the simulated rockburst was also gained. Some of the important findings are listed below: ?Two areas of damage were identified on the blasting wall: (i) an area of high-intensity damage in which ground velocities of 3.3 m/s were recorded by an accelerometer, which was subsequently ejected in a block of rock, (ii) an area of low-intensity damage where ground velocities of 1.6 m/s were recorded by an accelerometer which remained on the tunnel wall. ?High-speed filming revealed rock fragments being ejected from the wall at velocities in the range of 0.6 m/s to 2.5 m/s. The measurements were taken in the area of low-intensity damage. As planned, no gas pressure was directly involved in damaging the wall of the tunnel. ?The attenuation of peak particle velocities for the main blast in the near field (6 m to 30 m), as a function of distance R, was found to be proportional to R-1.7. ?The simulated rockburst was recorded as a local magnitude ML= 1.3 by the Klerksdorp regional seismic network. ?Peak particle velocities from nearby blasting, recorded on the tunnel wall after the simulated rockburst, were about six times greater then peak particle velocities recorded from the same source area before the simulated rockburst.

Jan 1, 2001

By P. Oyanader

"In 2001, the Catholic University of Chile undertook a maintenance benchmarking study of six open pit copper mines having mill capacities varying between 18,000 t/d and 156,000 t/d, and collectively responsible for 58% of Chilean copper production. This paper describes the methodology used to conduct the study, as well as the overall results. Key performance indices were selected to measure the effectiveness, efficiency, and development of the maintenance process. Using a balanced scorecard philosophy, these indicators were divided into client satisfaction, financial, internal process, and learning and growth indices.Maintenance was found, on average, to be responsible for 44% of mine production costs. Percentage planned maintenance of equipment fleets was found to be low by world standards, averaging 35%, 56%, and 44%, respectively, for blasthole drill, shovel, and haul truck fleets. Fleet availabilities were found to be significantly influenced by the percentage of planned maintenance achieved, while maintenance cost per equipment was found to decrease non-linearly with increases in percentage planned maintenance. Investment in technical training (including planned maintenance practice) was found to be low by global standards. The paper concludes that fleet availability and maintenance costs in Chilean open pit operations could be considerably improved by improving and/or developing maintenance planning standards, and investing in training personnel in planned maintenance techniques"

Jan 1, 2005

By Moraka N. V., van Rensburg M. Jansen

"SynopsisGoing beyond transformation claims contained in employment equity scorecards and industry compliance reports, this article provides qualitative insight into the initiatives employed and challenges experienced by mining companies in a quest to transform the mining industry. Perceptions expressed during in-depth interviews with 10 senior executives showed that the assumption that mining companies are reluctant to transform is erroneous. Results from this study suggest buy-in and commitment to transformation. This article describes specific initiatives undertaken by mining companies to transform. The most notable initiatives include staff recruitment efforts to appoint historically disadvantaged South Africans (HDSAs), staff development initiatives, as well as community development. The findings furthermore contextualize the challenges experienced by industry participants in their quest to transform. Despite accusations that industry participants are not taking responsibility for the implementation of the transformation agenda, government needs to recognize that they too have a role to play and need to appreciate and assist in the current challenges experienced by the industry.IntroductionWe need a mining sector that works. Mining employs over half a million people. It is the biggest earner of foreign exchange in our country. It also contributes about 20 billion rand directly to tax revenue. Mining also makes a far larger contribution as a buyer of goods and services, and a supplier of inputs to other sectors of our economy and other economies around the globe’ said President Jacob Zuma during the 2014 State of the Nation Address (Zuma, 2014). However, despite apparent commitment from government, enforced compliance with social and labour plans, regulations and Mining Charter targets, the South African mining industry is slow to gain local and international investor trust (Deloitte & Touche, 2013; Mashego, 2013).Industry players acknowledge that their context has changed – creating various challenges (Davis, 2014). In the past few years, such challenges included subdued commodity prices, increased working costs, constrained infrastructure, and high labour costs, coupled with poor levels of productivity, strained labour-management relations, an uncertain regulatory environment, and the inevitable maturation of the industry (Deloitte & Touche, 2013; Davenport, 2014). Mining affected communities also became more cognisant of their rights, and politicians more vocal in their expectations from their relationships with the mining giants (Davis, 2014). One of the main issues that contributed to the difficult operating conditions was claims from organized labour and Minister Shabangu that this industry is too slow to transform. Specific claims were made in terms of employment equity, fair salaries and wages for mineworkers, housing and living conditions, health and safety issues, and general working conditions (Limpitlaw et al., 2005; Miningmx, 2013; Shabangu, 2010). These claims were followed by what Gill Marcus, Reserve Bank Governor, explained as ‘unprotected strike action [that] has escalated into an uncontrolled, violent and unlawful landscape, led by a mob mentality in the absence of formal and recognised leaders’ (Mavuso,2013)."

Jan 1, 2015

By George F. Crocker

Introduction The Helen mine and the Victoria mine of Algoma Ore Properties, Limited, in the Michipicoten district, Ontario, are 120 miles north of Sault Ste. Marie, on the north shore of lake Superior. They are on the Algoma Central railway between Hawk Junction and Michipicoten Harbour. Between them, these mines have produced 10,757,113 tons of iron ore. From 1900 to 1918, 2,823,369 gross tons of brown ore was shipped, and from 1939 to the end of 1949, 7,933,744 tons of siderite was produced from the open-pit operations and underground development. The finished product, known as 'Algoma sinter', is shipped to blast furnaces both in Canada and the United States. The iron-ore bodies are a part of a tabular shaped area of iron formation that extends over a length of more than 10,000 feet and has maximum width of 1,000 feet. The two siderite orebodies, known as the Helen and Victoria, are separated by two diabase dykes. The Helen orebody is 1,600 feet long with a maximum width of 250 feet and the Victoria is 1,200 feet long with a width comparable to the Helen. These orebodies are considered to be a replacement of the original rock by iron carbonate. The introduction of the iron-1bearing material probably occurred after the development of the major folded structures. The open-pit mining of siderite started in the Helen ore body in 1939 and continued until 1945. During this period, 3,853,395 tons of ore was mined. Open-pit mining started in the Victoria body in 1946 and is still being carried on. Underground development commenced in September, 1947. Diamond drilling and surface sampling had shown the Helen deposit to average 35.61 per cent Fe and 7.28 per cent SiO2, and the Vi.otoria deposit 33.73 per cent Fe and 10347 per cent SiO2?

Jan 1, 1950

By J Cornejo, A Pinochet, C Caviedes

The ore extraction rate and the incorporation of new areas are key operational parameters. In massive deposits mined by caving techniques they both control the cave back geometry, which is mostly responsible for the induced stress state around productive levels. Early detection of deviations from admissible geometries will prevent stability problems related to geomechanical hazards, helping to improve safety and working conditions and reduce the potential for lost reserves.This paper describes a mine planning tool currently in use at the El Teniente mine. This tool involves the control of the cave geometry based on rules that have been defined as a result of different studies and experience collected during the operational history of the mine.This mine planning strategy achieves an extractive mining process with lower risk than conventional planning methods, ensuring ore reserves extraction if followed as planned while reducing unexpected changes in mining sequences.CITATION:Cornejo, J, Pinochet, A and Caviedes, C, 2016. Geometric control for strategic mine planning at El Teniente Mine, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 191–196 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By S M. Zhang, Y X. Han, J Liu, Y J. Li

"Compared with conventional roasting and magnetic separation processes, higher reduction temperatures and stronger reducing atmospheres are used in reduction technologies to reduce iron oxides to metallic iron, which is often liberated during reduction and amenable to magnetic separation. Reduction and magnetic separation of an Australian refractory iron ore was therefore studied as a function of reduction temperature, time and coal addition. It was found that during reduction, the particle size of the reduced iron increased with reduction temperature and time. Subsequent magnetic separation tests were further conducted on the reduction products. A reduced iron powder with a Fe grade of 85.67 per cent was obtained at a Fe recovery of 87.82 per cent under optimum reduction conditions where a metallisation rate of 85.82 per cent was achieved. The mechanism involved in reduction is discussed based on XRD and SEM examinations.CITATION:Han, Y X, Liu, J, Li, Y J and Zhang, S M, 2013. Upgrading an Australian iron ore by reduction and efficient magnetic separation, in Proceedings Iron Ore 2013, pp 397-402 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

Council for 1996/97 President Dr N.A. Barcza President Elect R.P. Mohring Vice Presidents Senior J.R. Dixon Junior M.H. Rogers Immediate Past President D.A.J. Ross-Watt Honorary Treasurer J.A. Cruise

Jan 1, 1997

By C. Stoneback

The attapulgite industry saw a year of solid growth after a weaker 2003. Demand was up between 5 percent and 6 percent driven by almost all of the key markets. Construction continued to show record new housing demand, assisted by continued low interest rates. That drove strong growth in construction related markets like coatings, tape joint compounds and other construction materials applications. The agricultural markets rebounded modestly to increase sales for suspension fertilizers, animal feeds and pesticide carriers. Oil drilling also showed strong growth due to high oil and gas prices. Sales to molecular sieve applications grew due to petrochemical and construction demand. The export markets were assisted by further depreciation of the dollar. The only weak area was in nonagricultural granular sorbents, where other sorbent materials continued to dominate.

Jan 1, 2005