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By W P. Breytenbach

"The phased application of statistical process control (SPC) infers that processes are classifiable as pre-SPC (or phase zero), phase I and phase II processes, with different objectives being pursued in the different phases. Phase zero processes, defined as processes affected by assignable causes of variation, are characterised as follows: out-of-control with unknown autocorrelation structure having an unknown statistical distribution not performing consistently on target with minimum variability potentially incapable of conforming to the product specifications. The objectives of phase I SPC is to obtain estimates of the mean (µ), variance (s2), probability density function (f(X)) and the autocorrelation structure, as well as transforming the process into a phase II process (defined as a process conforming to specifications and performing on target with minimum variability). As in any other type of research or pragmatic improvement intervention, sampling is required is to achieve these objectives. The challenge of deciding a phase I sampling plan is multifarious since it has to cope with the aforementioned effects of assignable causes and provide for rigour estimates of the unknown parameters. Furthermore, deciding sampling plans in phase I invariably presents a classical catch-22 situation: appropriate sampling is required to determine the autocorrelation structure of a variable, but knowledge of the autocorrelation structure is required to formulate an appropriate sampling plan. This paper suggests, for random variables, guidelines to formulate sampling plans, and presents a methodology to estimate the inherent process parameters of process variables and an approach to identify assignable causes in phase I. CITATION:Breytenbach, W P, 2014. Rational sampling and standard deviations for random variables in phase i application of statistical process control, in Proceedings Sampling 2014, pp 141–154 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jul 29, 2014

WaihiÆs Cornish Pumphouse is located on the edge of town in a dynamic and changeable area where history affects the future. As in any successful project, ingenuity and team work have been significant factors in the planning and implementation of the buildingÆs relocation.   Work around the Pumphouse site commenced circa 1900 with preparation to sink the No 5 shaft. Construction of the pumphouse and the steam driven C pump was completed by 1904 with hints of what was to come evident in 1905 as the No5 Shaft collar had sunk 350mm in four years. By 1907 surface cracking developed over the Milking Cow cave area and in 1913 200 mm was cut off the main pumping rod as the ground was sinking at a rate of 25 mm per year.   In 1913 ground around No 6 shaft had subsided 600 mm.It was also the end of an era as steam gave way to hydroelectricity and electric pumps replaced the C Pump. C Pump was maintained as a backup from 1913 to 1929 and then finally scrapped. 25 years of underground caving operations removed approximately 400 000 cubic metres of ore leaving a 7 ha collapse crater that eventually became known as the mine lake.   On the edge of town a tension zone over the Royal Lode began to develop 50 m diameter chimney collapses as large scale blocks relaxed into unfilled stopes in the old Martha underground mine. After the 1999 collapse part of the main street was closed and after the 2001 collapse Hazard Zones were declared, more streets closed and areas evacuated and retired from residential use.   By 1984 the Pumphouse has approximately one degree of tilt and in 2002 tilt on the Pumphouse accelerated briefly and then settled down. Tilt accelerated again late in 2004 and a scoping study was initiated to define protection options for the pumphouse. The Martha Mining Licence requires Newmont to 'Take whatever precautions are necessary to protect the old Cornish Pumphouse adjacent to No5 shaft' and viable solutions needed to found.   Any work on the Pumphouse comes standard with a wide range of stakeholders, governing agencies and legal framework to consider. Land Information New Zealand (LINZ) is owner and administrator of the pumphouse.

Jan 1, 2006

By M P. Sandy, J B. Albrecht

"Narrow vein mining can be loosely defined as mining operations involving orebodies with a true thickness of less than 4 m. Some authors have limited the term to steeper-dipping orebodies where the ore will rill under gravity. However, ænarrow veinÆ can also legitimately refer to some of the flatter orebodies, such as many of the platinum, chrome and gold deposits of Southern Africa, and several in Australasia including Bullen (at Norseman), Sunrise Dam, Telfer (Patterson, 1993) and the famous æflatmakesÆ at the Emperor Gold Mine in Fiji (Gordon and Harmse, 1993). Where the orebody dip is less than about 40¦ additional effort is usually required to remove the broken ore, including the use of scrapers, hydraulic monitors or remote-operated excavators (Tombs, Buckingham and Hancock, 2008).Working in such narrow orebodies presents particular operational and geomechanics challenges. A range of mining methods have been developed or adapted to exploit narrow orebodies including:cut-and-fill,narrow vein bench or longhole stoping,Alimak stoping,shrinkage stoping, and methods for shallow dipping orebodies (including Airleg and mechanised methods).The paper discusses some of the advantages and disadvantages of each method with a focus on geotechnical aspects including mining sequence options, stress management, ground support requirements, dilution and recovery.CITATION:Sandy, M P and Albrecht, J B, 2012. An overview of geotechnical issues in narrow vein mining, in Proceedings Narrow Vein Mining 2012 , pp 185-194 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 26, 2012

By A D. Pradana, D R. Pradana, Y H. Putra, Z Zulhan, T Hidayat

The high temperature processing of nickel laterite ore is currently dominated by the Rotary Kiln- Electric Furnace (RKEF) technology. The ratio of SiO2/MgO (S/M) in the ore is one of the critical parameters that determines the success of the RKEF process. Incompatible S/M ratios in the ore can cause aggressive chemical interaction between the slag and MgO refractory. The slag-refractory interaction at conditions relevant to nickel laterite ore smelting was investigated in the present study. Synthetic mixtures representing the slag compositions of nickel laterite ore smelting were prepared by mixing MgO, SiO2, Al2O3, CaO, Fe2O3, and Fe and heating the mixtures in steel containers at 1300°C under an argon atmosphere for 3 hrs. Pure MgO crucible were used as containment material in the melting process of the mixtures to represent the MgO refractory. The kinetics of the slagrefractory reaction were investigated using samples with selected initial S/M ratios of 2.0 and 3.0 (wt/wt) (with constant CaO = 3 wt per cent, Al2O3 = 5 wt per cent, FeO = 10 wt per cent) by performing melting at 1500°C in MgO crucibles for 5, 30, and 120 mins under an argon atmosphere. The effect of the S/M ratio in slag on the slag-refractory interaction was also investigated by performing melting of mixtures with different initial S/M ratios of 1.75, 2.0, 2.5, and 3.0 (wt/wt) in MgO crucibles at 1500°C for 120 mins under an argon atmosphere. After the melting process, each sample was quenched in water, mounted in resin, ground, polished, and coated. The interaction between the slags and MgO refractory was then evaluated by analysing the polished samples using a scanning electron microscope equipped with an energy dispersive spectroscopy (SEM-EDS). Thermodynamic simulations using FactSage™ 8.0 thermochemical software were performed based on the experimental slag compositions, refractory, and conditions to support the evaluation process. The results show that a higher S/M ratio in slag leads to a higher tendency for the refractory component to dissolve into the slag. A higher S/M ratio in slag also leads to an increasing extent of solid-state diffusion of iron oxide into the remaining refractory grains. The dissolution of MgO into the slag is rapid; the MgO concentration in the final slag increased by 2.4 wt per cent and 5.7 wt per cent at initial S/M ratios of 2.0 and 3.0 wt/wt, respectively, within 5 and 30 mins of the melting process.

Aug 21, 2024

By N K. DavT, A Coggan

The performance of in situ, shallow water covers in controlling acid generation was evaluated at four recently decommissioned pyritic uranium tailings sites at Elliot Lake, Ontario, Canada. The tailings are acid generating, having a pyrite content of approximately five to ten per cent and a very low to non-existent alkaline buffering capacity. Adverse market conditions led to the closure of all operating uranium mines in the early- to mid-1990s, followed by rehabilitation and decommissioning of mine and waste management facilities. In order to control acidic drainage, all new and active tailings areas at Denison, Quirke, Panel, Spanish-American and Stanleigh mine sites were re-engineered to provide in situ submersion of these tailings under shallow water covers. The tailings impoundment dams were upgraded and reinforced or in some cases reconstructed, to minimise seepages and provide a minimum 1 m depth of water cover. With the exception of Stanleigh mine, which was closed in 1996 and rehabilitated during the late-1990s, the water covered Tailings Management Areas (TMAs) at other sites have been in operation for a period of six to nine years. The surface water quality at these sites is maintained at near neutral pH conditions by periodic addition of lime slurry to the established water covers. Additional treatment of effluent is provided as required to meet the discharge water quality standards. The performance of the established water covers at the sites was evaluated by analysing the annual equivalent limestone consumption data. The results were compared with the performance of an older, decommissioned Nordic TMA, which was rehabilitated using revegetation techniques. The water covers at these sites are performing as designed and acid generation has been reduced to low levels. Based on the total amount of yearly equivalent limestone required to maintain water quality in the water cover and at the effluent treatment plant, it is concluded that acid generation rate at the Denison TMA has reduced to less than 0.15 per cent of its pre-water cover conditions during normal operation and rehabilitation periods. In comparison to the revegetated Nordic TMA, the acid generation rates at Denison and Panel TMAs have decreased to less than 1.6 per cent of that at the Nordic TMA. The acid generation rate at the Quirke TMA was comparable or only slightly lower than that at the Nordic TMA because of its unique terraced cell configuration, having tailings in five hydraulically interconnected cells at different elevations. Historic acidity and oxidation reaction products, accumulated during normal operating and pre-water cover periods, are flushing from upstream locations to downstream cells, thus requiring much higher equivalent limestone additions for acid neutralisation and effluent treatment.

Jan 1, 2003

This paper describes the technical development of the Newman beneficiation plant over its 23 years of operation and how this development relates to the mine ore reserves of high-grade and low-grade ore. It also discusses the importance of the beneficiation product grades in the overall blending of shippable products from a range of mines to optimise overall ore resource utilisation. Also discussed is the importance of the ore reserve reconciliations in resource utilisation. Data is provided on the capacity and quality improvement results of the plant upgrade, carried out in 1999. This shows how the original capacity upgrade targets have been exceeded and the additional quality benefits that have been obtained from this capacity.

Jan 1, 2002

By Jarvis ST

Colliery spoil is the major constituent of colliery spoil 'rockpaste' which has been used to infill abandoned limestone mines in the West Midlands of England since the early-1980s. The other constituents of rockpasle are pulverised fuel ash, lime and water. A key property of the rockpaste is its ability to flow within the mine workings over considerable distances before setting. This paper describes the work carried out to identify suitable sites as sources of colliery spoil for making rockpaste and goes onto consider the flow properties of the material and the monitoring methods used on-site. A full-scale trial comprising mulling a 6600 nt' section of an abandoned mine was conducted prior to the mulling of the Littleton Street Mine which had a volume of about 500 000 m'. As well as describing the methods used for monitoring the movement of rockpaste material, a description is also given of a dip-meter developed specifically for the purpose of measuring the level of paste in the mine. This paper shows how a colliery spoil based paste has been successfully rued for infilling abandoned mine workings. The monitoring systems developed for use in the two case studies presented in this paper are now being used extensively for infilling other abandoned mine workings. To-date over I I mines with a total volume exceeding two million cubic metres have been backfilled.

Jan 1, 1998

By T A. Short, P L. Brown

Kestrel Mine co-disposes around 1.5 Mt of fine (2SO4/t and to date around 9 Mt has been deposited. The mine plans to co-dispose of approximately 39 Mt of material, creating a dump with a maximum height of 35 m and a disturbance footprint of 260 ha. Benign materials suitable for rehabilitation are severely limited and no successful rehabilitation strategy has yet been demonstrated. A rigorous measurement and modelling program is resolving outstanding issues associated with acid rock drainage (ARD). The objective of the program is to develop and validate a strategy for dump rehabilitation that will control ARD and allow for mine closure. The strategy includes using a soil cover to reduce water flux into the dump to control the overall oxidation rate and to control effluent release. Monitoring of the oxygen distribution and water flux in the uncovered material commenced in February 2001 and monitoring in waste under a trial cover commenced in February 2002. Carbonaceous and sulfidic oxidation rates have been measured. Initial modelling indicates that the carbonaceous material has the general effect of slowing the rate of pyrite oxidation, lowering the maximum sulfate production rate and extending the time to oxidise all the sulfides. The combined approach of measurement and modelling provides a sound basis for assessing the performance of the trial cover and predicting the long-term quality and quantity of effluent at the base of the rehabilitated waste dump.

Jan 1, 2003

By L B. Zhou, S Li, B L. Wang, S B. Li, Q Wang

The rapid development in ChinaÆs metal mine resources has caused a series of mine environment problems such as serious ecological damage and environment pollution, and secondary geological disasters. The paper has reviewed ChinaÆs metal mine development, including current status and environmental situation, analysed existing problems in mine development and environmental management, and proposed corresponding countermeasures and suggestions.CITATION:Wang, B L, Li, S B, Li, S, Zhou, L B and Wang, Q, 2012. Current situations, problems, countermeasures and suggestions for ChinaÆs metal mine development, in Proceedings Life-of-Mine 2012 , pp 353-356 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

Mining operations in the Obuasi mine of AngloGold Ashanti Limited began in 1897 using mainly gravity concentration and amalgamation to recover the gold from predominantly high-grade quartz ores. The mine is primarily an underground operation, although some surface mining still takes place. Ore is currently processed by two main treatment plants: the Sulfide Treatment Plant (STP) for underground ore and the tailings treatment plant for tailings reclamation operations. STP was commissioned in February 1994 to treat 180 000 t/m of double refractory gold ore and it hosted the then world's largest commercial scale bacterial oxidation plant which uses Gencor's BIOX« technology. Gold is recovered by gravity and flotation, with the regrinding of the sulfide concentrate before bio-oxidation to liberate locked gold in sulfi de minerals. The BIOX« product is leached using the carbon in leach (CIL) process. Gravity gold recovery is a major component of STP operations and its contribution to overall plant recovery is about ten to 13 per cent. Gravity recovery has undergone series of technological upgrading, since the very first circuit which consisted of a Knelson concentrator and a Gemeni shaking table which was incorporated into the milling circuit at the start of the plant. Presently, each of the gravity circuits which consists of Knelson concentrator(s), cyclones and Gekko InLine Reactors (ILR), are located at the Milling and BIOX« circuits. More recently a new gravity recovery facility comprising of four 48 inch Knelson concentrators and cluster of ten inch cyclones have been incorporated at the tail of the flotation circuit. An additional ILR will be commissioned in 2010 to treat the concentrate generated from the new Semi Autogenous Mill #2 (SAG 2) Knelson: concentrator tower under construction. Since the inception of STP, significant gravity recoverable gold has been recovered via Knelson concentrators and the employed circuits have seen a number of modifications and upgrades, motivated by ore mineralogical changes and recovery concerns. The pioneering adoption of the ILR technology in 2001 and 2003 was a critical factor in enhancing the improved recovery benefit of the gravity circuits. Therefore the objective of this paper is to highlight the mine's experience in the gravity recoverable gold process with emphasis on plant modifications and gravity equipment upgrade. Operational challenges and the contribution of gravity gold to overall plant recovery are also discussed.

Aug 1, 2010

By S Sims, A Stradl, G Hope, B Lacouture

Column flotation cells have been installed in numerous base metal operations around the world. The majority of these cells utilise conventional air-only spargers to introduce air into the bottom of the column. The recent development of instruments to measure the bubble characteristics in these columns has provided a renewed understanding of column behaviour. These new tools provided insight into why the columns at Red Dog Mine had never performed up to expectations. After efforts to optimise the spargers failed to substantially change the bubble size and air efficiency, alternate sparger systems were investigated. The Metso Minerals CISA Microcel sparger system appeared promising and was selected for a full-scale plant trial. The Microcel sparger system was originally developed for the coal industry at the Virginia Centre for Coal and Minerals Processing. In October 2003, a Microcel was retrofitted into one of two 3.66 m diameter flotation columns in the zinc retreat circuit at Red Dog Mine. The operation of the two different sparger systems in parallel allowed a detailed comparison. This paper discusses the performance of the Microcel based on the data collected during several detailed surveys. Bubble size measurements carried out in the pulp zone using the McGill University bubble viewer showed a significant difference in bubble size. The mean Sauter diameter of the bubbles decreased from 3.4 mm for the jetting-type sparger to 1.9 mm for the Microcel sparger. The overall recovery and the recovery by size fraction for both valuable and gangue minerals were compared. Paired t-tests demonstrated that the Microcel column produced a higher concentrate grade (0.6 per cent zinc absolute) and a higher unit recovery (2.8 per cent zinc absolute) than the existing Canadian Process Technologies Inc (CPT) SlamJet column. These improvements provided a payback period of 1.5 months for the $109 000 investment.

Jan 1, 2005

By T Laing, M A. Abdalla, P Hagan

Open pit operators are facing a number of challenges in the area of blasting and blasting processes, such as wall control, quality of fragmentation, explosives consumption and vibration. Remaining financially viable is one of the most significant objectives for mining professionals as they work through the blasting challenges within the mining chain of processes. This objective is such a significant challenge, primarily due to the high consumption of materials and the direct consequences on the downstream processes in the mining work flow, including loading, hauling, crushing, milling and mineral concentration. This paper reviews and discusses a technique to achieve this objective, namely air-decking. The performance in a number of mines around the world is reviewed focusing on the economic contributions of air-deck blasting. Various scenarios are investigated and simulated, covering both conventional and air-deck blasting, using JKSimblast software. The scenarios explore a range of issues and found that optimum results could be achieved with a minimum amount of explosives. The results obtained from the simulations were verified with blasts at Rixs Creek Coal Mine, Australia. Based on the verification results the simulations were then modified to more accurately model the actual conditions and parameters on site and to ensure that the models derived meet the production and control requirements at the mine site. Ultimately, overall analysis and comparisons were applied to the outcomes of the conventional and air-decking blasts revealing a reduction in material consumption of around 25 per cent with consequential savings and up to a 30 per cent reduction in ground vibration.CITATION:Abdalla, M A, Laing, T and Hagan, P, 2012. The contribution of air-deck blasting to open pit mining economics, in Proceedings Eighth Open Pit OperatorsÆ Conference , pp 7-12 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 18, 2012

By F D. I Hodgson

Twenty-seven observation test pits were excavated at four different spoils dumps at three collieries in the Witbank Coalfield. The holes were dug into the upper three metres of the spoils to ascertain if any weathering profiles, variation with age or influence of topsoil covering could be gauged. Samples from these holes were also subjected to laboratory testing (ABA, humidity cells and mineralogy on several selected samples). The test pits showed that there was a good correlation between in situ paste pH-values and ABA results for samples taken from the pits. At all the collieries, no correlation between age of spoils and in situ acidity could conclusively be drawn. The depth of soil cover also appears to be a minor consideration in determining the eventual acidity of the spoil. In the pits where acidity was found, the weathering was evident when contrasted to areas still buffered. It is unclear if the weathering leads to the acidity or vice versa, as the soil covering appeared to be similar. The spoils proved to be extremely variable with regard to particle size and material on a micro and macro scale. Evidence from most of the holes suggested that there was enough water contained in the fine material in the spoils to support the initiation of acid generation. Conclusions from this investigation are that the mineralogy of the spoils is the most important factor in determining the eventual occurrence of acidic hotspots, and that prior characterisation and selective handling will provide the best management practice for these spoils.

Jan 1, 2003

By Giedl JG

Technical applications of computers within the coal industry are expanding rapidly with an increase in the number of users and user locations. Collinsville Coal Company Pty Ltd is an example of a remote and growing user company which has a Telecom link to MIM's ICL mainframe computer in Brisbane. The Planning Department is one of the technical user departments at Collinsville, with a variety of VDU, plotting and screen graphics hardware supported by the MIM Computing network. Current applications involve the storage, retrieval and manipulation of borehole information, plan and map production, volume calculations from survey data and mine ventilation modelling for underground Operations.

Jan 1, 1983

The purpose of this paper is to look at the costs of tailings disposal, and in particular the comparative costs of different disposal methods in different situations. Costs are often the hidden design parameter. We know that they are important but sometimes feel that it is not very respectable to talk about them in public. It is better from a public relations point of view for it to appear that the decision-making process is driven by environmental factors or best practice rather than looking for the cheapest option. But this presumes that the two are mutually exclusive and this need not be so. These issues are discussed later in the paper. It is also commonplace for the scale of the mining operation to change with time, typically involving increased throughput rates and total tonnage. New ideas in tailings disposal make their appearance. The regulatory attitude can change. Having made a decision about the disposal method at the time of original feasibility study and design it is easy to suppose that the mine is æstuck with itÆ. But that also need not necessarily be so. Careful costing of options may reveal that a switch in method could be cheaper in the long run. The designer will often tend to think in terms of initial earthworks and site development costs. Aspects of power consumption, labour, and periodic campaigns to move tailings pipelines and carry out embankment raises are either overlooked or not given the same emphasis. This can distort the evaluation process. It is productive to think in terms of dollars or cents per tonne as a basis of comparison, not just the total sums. Tailings disposal can cost from as little as 10 - 20 ó/t to over $1.00/t which is a big range. There may be good and unavoidable reasons for the high figure but there would need to be to justify being so far over the industry average. At the same time, mining and processing costs, at least for gold, are around $15 - 20/t. Viewed from this perspective tailings disposal will always be a relatively small component of the annual budget.

Jan 1, 2000

The sulphide ore occurring in the North Mine at Broken Hill is of a complex nature, consisting of an intimate mixture of argentiferous galena, marmatite (zinc-iron sulphide), and gangue minerals. A typical analysis of the ore is:Lead 15.2%Zinc 12.0%Iron 5.3%Manganese 0.8%Copper 0.1%Alumina 0.3%Lime 0.9%Sulphur 11.1%Insoluble 53.8%Carbon dioxide 0.3%Undetermined 0.2%Silver 9 oz. per tonTotal 100.0%The metallurgical practice for the recovery of the lead concentrate and zinc concentrate from the crude ore is briefly described in the following pages.(1) COARSE CRUSHING AND FINE GRINDING OF THE CRUDE OREThe crude ore from the mine, containing lumps to the size of, roughly, 10 in. x 8 in., is tipped over grizzlies with the bars spaced 3 in. apart; the oversize gravitating to a...

Jan 1, 1933

By H Kusumi, A Nishio, K D. Halim, Y Yun

In Japan, evaluation of rock mass is crucial in mountain tunnels that are constructed using the NATM (New Austrian Tunneling Method) to determine the support pattern that in turns determine the final support structure of the tunnel. The current evaluation method is highly dependent on the observations of on-site expert engineers that grades the rock mass in four levels across nine criteria: A. Condition of tunnel face. B. Condition of excavated surface. C. Compressive rock strength. D. Weathering. E. Spacing of discontinuity. F. Condition of discontinuity. G. Orientation of discontinuity. H. Spring water. I. Water degradation. However, in recent years, the number of said experts are decreasing rapidly due to labour shortage caused mainly by an aging population and declining birth rates. This is one of the biggest challenges that the construction sector currently faces, and this may affect the efficiency of tunnel construction projects in the future. Thus, this study proposes the implementation of a deep learning tool called the convolutional neural network (CNN) to quantitatively evaluate rock fractures on tunnel face. In addition, Gradient-weighted Class Activation Mapping (Grad-CAM) is applied to visualise CNN, as well as verify the applicability of CNN to the evaluation of rock fractures. In other words, by creating heat maps of tunnel face images, CNN is utilised to evaluate rock mass in categories that are associated with discontinuities. The CNN model is most accurate when six convolution layers, four pooling layers, and three dense layers are used. This yields an accuracy of 75.0~88.0 per cent. Thus, the CNN model is verified to be feasible method to evaluate rock fractures.

Nov 29, 2022

In the first half of 1987 two four metre diameter raise bore holes were drilled at the Olympic Dam Project for ventilation purposes. The first hole, 476 m x 4.0 m diameter was completed in 96 days, 35 days longer than scheduled. The second hole, 407 m x 4.0 m diameter was completed in 39 days, 2 days ahead of schedule. Both holes were drilled with a Robbins 121R rig using a Wirth head that consisted of 23 face cutters and 5 side cutters, supplied by a contractor. After the sites were located a grouting program was carried out to help seal off an aquifer between 160 and 180 m below the surface. Cavities discovered in the Andamooka Limestone were filled with concrete. Because of the overlying dune sands and clay both sites were pre-sunk to the competent Andamooka Limestone.

Jan 1, 1988

By K L. C Gontalves

In 1999, the Mineral Development Centre of CVRD (CDM) acquired a flotation device referred to as Mini Pilot Plant (MPP), which is able to run continuous flotation tests with drill core samples. Since then this equipment has been used to define the flotation circuit and process conditions during the pre-feasibility phase of CVRDÆs copper sulfide projects. Until now, the MPP results had not been validated for an industrial flotation operation. The present work consists of a comparison and correlation of the results from a flotation MPP and an industrial plant. In order to make this comparison, an industrial plant was completely sampled and the feed ore was tested on bench scale and on a MPP. The metallurgical and mass balances of each test were used for the comparison.

Jan 1, 2005

By M J. Nugus, J G. McLellan, T G. Blenkinsop

High fluid pressures and deformation have been instrumental in the formation of high-grade gold lodes at the Archaean granite greenstone 12 Moz Sunrise Dam gold deposit in Western Australia. Historically, predicting the location of these lodes has been difficult due to the irregular nature and high-grade nugget characteristics of the deposit. Recent exploration successes and the discovery of the Vogue system have led to an improved knowledge of the deep extents of the mine. This new structural and gold distribution data has enabled a new 3D interpretation of the system and has provided a better insight into the links between deformation, fluid pressure evolution and gold deposition. Here we employ 3D geomechanical and coupled fluid flow modelling to better constrain areas of high strain partitioning, rheological failure and focused fluid flow, which provides a substantially increased predictive capacity for further within-mine exploration. The localisation of deformation in and around the main felsic porphyry and major shallow shear zones provides critical insights into fluid pathways that have contributed to the distribution of fracture networks and ultimately gold deposition during late stage deformation events. Understanding the relationship between deformation events, fluid pressure evolution and the partitioning of strain is paramount in understanding the distribution of gold and for focusing on future target areas.CITATION:McLellan, J G, Blenkinsop, T G and Nugus, M J, 2013. Fluid pressure, geomechanical evolution and predictive targeting of the Sunrise Dam gold deposit, Western Australia, in Proceedings World Gold 2013 , pp 419-426 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013