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By D Thornton

Geologists use large data sets and spend many hours and much computing time modelling the distribution of minerals throughout a rock mass. Significant investments in drilling, assaying and quality assurance and quality control (QA/QC) are made before sophisticated geostatistical techniques are applied to the collected data. A large amount of research has been done in this field such that the ore control model is a reasonably accurate reflection of the mineral that is actually in the ground. However, after all this effort, the mining engineers then work just as diligently to blast the in situ rock mass to a particle size that can be efficiently handled and further reduced by downstream comminution processes. In open cut mines, this invariably requires large amounts of explosive energy that displaces the ore boundaries from where the geologists originally defined them. Failure to accurately account for blast movement will result in misclassification, whether that is ore to waste; low grade to high grade; sulfide to oxide; or other contaminates û collectively referred to as ore loss and dilution. The financial consequence of getting this wrong is substantial and not well quantified. This diminishes all of the good geological QA/QC done in defining the ore. This paper discusses various technologies used to physically measure blast-induced orebody movement and the difficulties in modelling this with a purely theoretical approach. Rock mass structure, blast energy, design and timing contours all have significant and interrelated effects on blast movement vectors. Measured blast movement has a large variance due to the uncertainly of many of the controlling parameters û arguably dominated by the heterogeneous nature of the rock mass. Although there have been significant advances in blast movement modelling capability in recent years, this unpredictable blast movement limits their use in the context of production ore control. Certainly, without physical field measurements, accurate predictions of movement are near impossible.Limitations of traditional surface markers in determining three-dimensional (3D) blast movement within the orebody are also reviewed, and advancements in accurately measuring these vectors within blasted ore discussed

Sep 26, 2011

The Fiji islands are the world's most ac- cessible Cenozoic remnant arc. They provide occasional subaerial exposures of an arc frag- ment some 750 km long and 80 km wide, or about the size of the NE Japan volcanic arc. The fragment is one part of the former Vitiaz Arc which once connected Fiji to Vanuatu and Tonga. The Fijian portion has the most completely stu- died Tertiary section which contains a history of arc volcanism from the late Eocene to late Miocene (the Early and Mature Arc stages), a rich diversity of subsequent magmatic responses to the most recent episode of arc rifting (Ear- ly Rifting stage), and an example of transition from arc to intraplate volcanism as this rift- ing proceeded (Late Rifting stage). The arc fragment is surrounded by younger backarc ba- sins. The chemical composition of magmas in both the arc and backarc changed with time, but in opposite directions. In the arc, magmas be- came enriched in sialic components from the Early Arc to Early Rifting stages, whereas in the basins the magmas became more similar to mid-ocean ridge basalts (MORBs) with time. In southwest Fiji, the oldest backarc basin has subducted obliquely beneath the remnant arc du- ring the last 3 Ma, resulting in chemically ex- otic calcalkaline volcanism. During the last decade over 150 radiome- tric age measurements and 650 rock analyses, many with trace element and isotopic measure- ments, have been obtained for samples from Fi- ji, Tonga, and the surrounding seafloor (Figs. 1 and 2). These data are summarized, drawing from the publications of Colley and Hindle (1984), Gill et al. (1984), Hawkins and Melchi- or (1985), Kroencke (1984), Malahoff et al. (1984), Scholl et al. (1985) and Whelan et al. (1985), from the specific research papers refe- renced therein, and from the half dozen or so manuscripts currently in preparation by these people and others.

Jan 1, 1987

When, in 1985, Marty McFly is accidentally sent back to 1955 in a DeLorean ‘time machine’ in Back to the Future, he encounters young versions of his parents, and must make sure they fall in love or he will cease to exist. This Conference celebrates 40 years since the inaugural Mill Operators’ Conference in Mt Isa in 1978; over this period, great technological advances have been made in almost all aspects of mill operation, resulting in considerable improvements in grades and recoveries from ever-more complex ores. However, in 2018, the mining industry is facing challenges that might have been difficult to predict in 1978 (or were they?). The computer revolution, the information revolution, and the bio-technology revolution have all contributed to the increase in operating efficiencies, but, combined with the social revolution, they also call into question the existence of the mining industry as we know it. This presentation looks back at the ‘way things were’ in the 1970s, and at what was then considered to be the ‘basics,’ and attempts to sketch what the mining industry might be like in 2058. As Back to the Future Part II demonstrates so vividly, it is much easier to review the past than predict the future. It is difficult to know what form the mining industry will have in 40 years’ time. Nevertheless, as this presentation will argue, it is important to take action now, to ensure that it (like Marty McFly) does not cease to exist. Some courses of action are proposed, including going back to the basics of a very different past.*This is an abstract only. No full paper was prepared for this abstract.*CITATION:Franzidis, J P, 2018. Back to basics or back to the future?, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 3–4 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By J Araya, O Leiva

The mine evaluation process, composed of the designing, planning and valuation of a mine project, is, among other variables, based on an estimated orebody model. However, as the information from the drill holes is limited, it is impossible to know with certainty the mineral content in each block within the estimated block model. Consequently, uncertainty about the value of metal grades, or other geological attributes, arises because of the lack of information at unsampled locations. This is one of the reasons why, in any mine operation, there will always exist a divergence between estimated and real values. Cut-off grade optimisation is the process used to determine the operating mining strategy (long-term production scheduling) that will maximise the total profit value (normally referred as the net present value (NPV)) of a mine operation. This is done by defining a dynamic cut-off grade policy which is constrained by the production capacity of the mining, mill and refining.Real options is a type of contract, written on real assets such as mining, oil and other natural resource industries, that gives you the right but not the obligation to exercise the contract. It is based on operational and managerial flexibility, which allow mine decision makers to react to either good or bad operational and economic events in the future. But, what do geostatistics and cut-off grade optimisation have to do with the application of real options in a mine project evaluation process? The objective of this paper is to answer the previous question by showing how geostatistics combined with cut-off grade optimisation are key tools for corporate/managerial decision makers, which allow them not only to analyse the resource/reserves of their projects, but also to generate operational strategies that can maximise project value by minimising the downside risk due to ore tonnes and grade uncertainties, while achieving operational and physical constraints such as mining and processing capacities. A real copper grade open pit mine project is used as a case study.

Sep 26, 2011

The value derived by good mine planning is often underestimated as many mines steer their attention to the more immediate and more easily recognised successes made by the day-to-day operational æbush fireÆ fighting. Similar to machine maintenance, the efforts made to quickly rectify short-term breakdowns are well recognised; whereas the value derived from the work in planning, scheduling and undertaking preventative maintenance is not so obvious or readily quantified. Too often the focus of the business shifts to building and supporting an efficient operations team at the expense of developing a highly experienced team of mine planning practitioners that are well trained with procedures and purpose-built tools to develop optimum mining strategies. A resource has an inherent value based on its quality parameters and its likely mining conditions. How well that value can be unlocked is governed, to a large degree, by the proficiency of the mine planning practitioners, as they, essentially, set the foundation of the mine business and the path along which the mine will operate. The importance of having a well documented and procedural mine planning methodology that integrates both a thorough technical and economic understanding of the resource cannot be over emphasised. The aim of the mine planning process is to develop a robust mine design and production capability which can best cater for future uncertainty whilst providing the best opportunity to maximise cash flow into the business. The planning process must be transparent and auditable, clearly summarising the process steps and assumptions made to arrive at a justifiable and defendable outcome. To provide maximum value to the business, the planning process must be repeatable and scalable to enable easy replication throughout the different mines and projects within an organisation. The diminishing numbers of highly experienced personnel within our industry, coupled with the growing pressure to clearly demonstrate that appropriate due diligence has been exercised, means that well documented and monitored mine planning processes using purposed built and innovative tools need to be developed to ensure optimum outcomes.

Jan 1, 2009

By M Miller

Operators of nickel laterite pressure acid leach and gold pressure oxidation technology frequently employ a series of steam condensing heaters to preheat the leach feed slurry. The heaters are an economic pinch point û although the thermal performance of the heaters is critical to overall process viability and operating cost minimisation, they are also a significant capital item and must be sized accordingly. Despite this, discussions of the factors underlying heater design are conspicuously absent from technical literature, and existing operations feature a conflicting range of heater sizes and configurations. This variety would suggest that at least some of the chosen heater designs are oversized. As part of its detailed design stage, the Ravensthorpe Nickel Project (RNP) constructed and tested a pilot plant heater to develop an understanding of the heating process and to ensure that the RNP heaters were appropriately sized. This testwork used RNP slurry at design density and viscosity on a 450 mm diameter, 6500 mm high column. The pilot plant testwork demonstrated the intrinsic efficiency of the direct steam condensation heating process, which was essentially independent of slurry density and viscosity. Presence of non-condensable gases in the steam phase decreased the heat transferred in certain circumstances, but heater performance remained consistent with thermodynamic limitations. Stagewise heat transfer within the column was rapid, and suggested either a very high slurry surface area or excellent heat transfer within the slurry phase. While the testwork did not set out to distinguish between these effects, physical evidence pointed to droplet formation in the inter-tray shower zone, giving a high heat transfer area. For scale-up purposes, the key facts were that the heat transfer rate was not limited by using high viscosity slurries and heaters could be designed using the principles applied to the smaller of those in existing operations. This paper discusses existing slurry heater designs, the current understanding from open literature, the pilot plant testwork, results and analysis, and scale-up to commercial sized equipment.

Jan 1, 2005

By Kneisley RO

This paper presents a brief overview of current bump mechanics theories and pillar design methodologies, and relates these concepts to experiences at two mines located in a north- central Utah coalfield where different pillar designs were used to control mountain bumps. Experiences gained at the first mine demonstrated the successful implementation of a two-entry, 9.8 m (32 ft)-wide, yield-pillar design. A U.S. Bureau of Mines field study quantified the timing of chain pillar yielding and resulting load transfer from the gateroad. In-mine pillar response, although apparently sensitive to site-specific conditions, compared favorably to estimates derived using two yield- pillar design methods. A second study conducted at another mine located in the same district, but subjected to different geologic conditions, documents the unsuccessful attempts to employ progressively narrower three-entry pillar designs based on successes achieved at the first mine site. This second mine never achieved a true yield-pillar design. Use of pillar widths ranging between 9.1 m (30 ft) and 27.4 m (90 ft) always resulted in violent bumps in the tailgate pillars. The pillars were either too large to yield, or too small to support peak operational loads. Hypothetical gateroad systems were evaluated using both analytical and empirical approaches for configurations comprised of two rows of conventional pillars, and systems incorporating both yield and abutment pillars. Analysis concluded that yield pillars less than 6.1 m (20 ft) wide and abutment pillars ranging between 30.5 m (100 ft) and 39.6 m (130 ft) square would be required to achieve a stable gateroad design. However, results of a field study conducted on a two-entry, 36.6 m (120 ft)- wide abutment pillar concluded that abutment pressures from the first panel overrode the pillar, and that a still larger pillar may be required to preclude bumps in the tailgate during second panel mining. Without the benefit of a demonstrated in-mine success, it is not clear which design would ensure elimination of gateroad bumps.

Jan 1, 1992

By R C. Morris

The channel iron deposits (CID) (Ramanaidou, Horwitz and Morris, 1991; Morris, Ramanaidou and Horwitz, 1993; Ramanaidou, Morris and Horwitz, 2003; Morris and Ramanaidou, 2007) are one of the two major iron ore types mined in the Hamersley Province. The CID provided around 41 per cent of the total of 268 Mt of iron ore mined from the Hamersley Province in 2006. The current CID resource is around seven billion tonnes. The CID deposits occupy numerous meandering palaeochannels in a mature surface that includes Precambrian rocks and ferruginous palaeogene valley fill. These palaeochannels are generally less than a kilometre but can range to several kilometres in width and from one metre to more than 100 m thick. The two currently mined CID areas are located:in the western Hamersley Province in the Robe palaeochannel, and in the east in the Yandi palaeochannel. These two major CID channels extend over 100 km to 150 km, with the Robe system up to 5 km wide. The granular ore facies typically contains ooids and lesser pisoids with haematite nuclei and goethite cortices, with abundant goethitised wood/charcoal fragments and goethitic peloids, minor clay and generally minimal porous goethitic matrix. Post deposition weathering is common and produced secondary facies.Ooids and pisoids were mostly derived from the stripping of a well- vegetated deep ferruginous surface. The peloids were derived both intraformationally from fragmentation and reworking of desiccated goethite-rich mud. Minute wood/charcoal fragments in the soil that were initially replaced by goethite and then dehydrated to haematite, formed nuclei for many ooids and pisoids. In addition, abundant, generally small (in situ by goethite within the consolidating CID. This profusion of fossil wood, both as ooids and pisoids nuclei and as discrete fragments as well as the local presence of kenomagnetite suggests major episodic wild fires in heavily vegetated catchments. The goethitic matrix was the result of chemically precipitated iron hydroxyoxides, resulting from leaching of iron-rich soils in an organic environment.

Jan 1, 2007

By Sorentino C

Mineral valuation may appear to be a special field of financial valuation when in fact it is a subset of general investment decision theory. The only significant difference in mineral valuation is that in the case of valuing an on-going mineral asset, ie a mine, often there is no initial capital outlay involved. One is required to determine what the mineral asset as it stands is worth. But this is the same problem as determining the maximum one could pay today and not incur a loss over the life of the asset. In other words one needs to determine the Net Present Value (NPV) of the asset. The two most widely used discounted cash flow methods, NPV and Internal Rate of Return (IRR) came of age with the rapid adoption of the personal computer and spreadsheet programs. The two methods are directly related as the IRR is a special case of NPV. NPV measures the absolute increase in real wealth that can be expected from a certain investment, which makes it particularly suitable for mineral asset valuation, but ignores the efficiency of the investment. In contrast, IRR gives the expected efficiency of the investment but ignores the increase in wealth aspect. IRR has two serious theoretical weaknesses; multiple IRR's and a very rigid assumption about the rate of return of reinvested proceeds, plus if no initial equity expenditure is involved there is no IRR. The choice of the ""hurdle"" rate can also be a problem and in certain circumstances one can obtain different rankings of mutually exclusive projects using IRR and NPV. Although NPV and IRR supposedly measure financial risk, in fact they only address a very narrow area of risk, namely the time value of money given the real discount rate. They side step the issue of what the real discount rate should be and varying degrees of risk. The classical way to establish the real discount rate is to use the weighted average cost of capital.

Jan 1, 1994

By M J. Ireland, D P. Horne, S J. Simpson

The Sinclair Nickel Project, wholly owned by Xstrata Nickel Australasia (XNA), includes an underground mine, a concentrator, a village, an all-weather airstrip and associated mine infrastructure. The operation is located 70 km north of Leonora in the prospective and historically productive eastern Goldfields region of Western Australia. Xstrata Nickel has successfully developed the Sinclair deposit into a 350 kt/a operation since its discovery in November 2005, with contained nickel metal production to date in excess of 10 600 t, of which 5500 t is from the underground operation.The narrow, continuous, medium-grade Sinclair nickel sulfide deposit accessed by the Phoenix Decline has a shallow 15¦ - 20¦ plunge that is approximately 1000 m long down dip and open at depth to the north. The flat plunge and low-grade nature of the orebody presents challenges regarding the amount of horizontal waste development that can be economically excavated, which in turn affects ventilation, egress, mine services and diamond drilling.The mineralisation is hosted between an ultra mafic hanging wall and a basalt footwall and is being mined using a bottom up, sublevel open stoping method. The stopes are retreated to a central access, filling voids in the process with waste mullock where possible and using cemented mullock fill where end exposure is required. Tight filling the flat dipping hanging wall has proved to be challenging given the angle of repose for the waste mullock fill is steeper than the angle of the hanging wall.Reconciliations of mine performance against the original feasibility study show positive changes and operational optimisations enabling Sinclair to achieve its mining targets in a safe, sustainable and economic manner. This paper discusses the mining operations associated with the first two years of the Sinclair Underground Nickel Project, providing an overview of the various issues encountered and the lessons learnt in the process.CITATION:Ireland, M J, Horne, D P and Simpson, S J, 2012. Overcoming mine design and production challenges at the Sinclair underground nickel project, in Proceedings Narrow Vein Mining 2012 , pp 37-48 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By D A. Williams

The conventional method of transporting mine tailings from a process plant to a tailings storage facility (TSF) is in the form of a low to medium density1 slurry, which can be pumped using centrifugal pumps. Over the past several decades, therefore, mine operators have generally adopted this default approach. There are, however, some convincing arguments why consideration should be given to thickening the tailings slurry to a higher density paste consistency. In particular, the minimisation of environmental impacts, the reduction in water consumption and the cost savings relating to TSF closure make paste an attractive option for the life of mine tailings disposal. There are several environmental impacts that are associated with a conventional slurry TSF, such as land sterilisation, aesthetics, dust, seepage and acid mine drainage. Some of these impacts, such as dusting, may be difficult to mitigate while the TSF is in operation. Others can be influenced by adjustments made to the tailings slurry density at the plant. Unplanned fluctuations in slurry density and/or viscosity can have a significant influence on TSF performance and can exacerbate the environmental impacts relating to the facility. Relatively minor changes in the solids content of the slurry can affect tailings beaching behaviour, degree of segregation, position of the phreatic surface, amount of freeboard, consolidation behaviour, placed (settled) density (and hence TSF capacity) and in situ strength of the tailings. Paste is a non segregating slurry and many of these effects can be minimised at the TSF when the tailings slurry is thickened to a paste. These improvements are manifested by reduced seepage, flexibility in selecting an appropriate final landform, ability to rehabilitate progressively and reductions in plant water usage. It is possible to minimise environmental impacts of a TSF through the use of tailings paste. Ideally the paste production plant should be part of the initial design of the process plant. However, significant benefits can be gained by adding a paste production plant at some time later in the operational life of the mine to reduce environmental impacts and facilitate an environmentally acceptable TSF closure.

Jan 1, 2002

By C Yeung, N Swa, D Gribble, R Reupana

BlastLiteÖ is a low density explosive developed by Dyno Nobel and has been in use at BHP Billiton Iron OreÆs Mt Whaleback Operation since March 2004. The Mt Whaleback Operation utilises BlastLite because it offers an attractive alternative to ANFO in soft to medium strength rock types. Its use also reduces the amount of liquid fuel oil consumed and is well aligned with BHP BillitonÆs corporate Energy Excellence initiatives to reduce greenhouse gas emissions. This paper presents the:results of characterisation work completed on-site as a fourth year mining engineering thesis, and application and benefits of the product at Mt Whaleback. One of the key features of low density products is that the detonation is non-ideal. That is, the explosive does not react fully in the detonation driving zone. This manifests itself in that the product is highly dependent on charge diameter and confinement conditions. Dyno Nobel has access to the Vixen suite of detonation codes for characterising the ideal and non-ideal behaviour of explosives. The objectives of this work were to use the Vixen_i code to characterise the detonation performance of BlastLite, in order to understand the role of the bulking agent and to allow modelling of the explosive. The detonation performance was established by detonating a series of unconfined charges at various diameters and measuring the velocity of detonation (VoD). The work program consisted of sourcing and preparing cardboard tubes of various diameters (80 mm to 305 mm). The tubes were loaded with explosive at the BHP Billiton Iron Ore Mt Whaleback test range and detonated one at a time. Over the past three years BlastLite has been used in various applications across the site. Initial findings illustrate that shovel productivity from BlastLite shots was comparable to the baseline ANFO blasts. The study results were also backed with visual inspection and feedback from shovel operators. A work program is currently in place to develop different blends of BlastLite and TITAN« BlastLite to tailor the explosives to various blasting applications. The trials at Mt Whaleback are leading to future opportunities of BHP BillitonÆs satellite orebodies.

Jan 1, 2007

By C J. Berry, P H. Briggs

Assessment of metal mobility, acid rock drainage production, and toxic effects from weathering of mineralised materials (ie mine-waste piles and outcropping alteration/ore zones) is key in evaluating and mitigating metal mining sites. Many metal- and acid-generating materials at these sites are sulfur minerals (sulfides to sulfates) with differing reactivities, weathering characteristics and metal contents. Our working hypothesis predicts that data on the types and quantities of sulfur minerals, coupled with comparison of their sulfur isotopic compositions to that of aqueous sulfate in the surrounding environment, can be used to evaluate primary pathways of weathering reactions at mining sites. We developed a sequential extraction scheme to separate the phases of sulfur in mineralised materials. Operationally defined phases, including monosulfides, disulfides, elemental sulfur, water-soluble sulfo-salts, acid-soluble sulfates, and resistant sulfate phases such as barite, are quantified gravimetrically and/or instrumentally. Several sets of experiments were designed to test the effectiveness of this extraction scheme. First, a variety of ore sulfides diluted in an inert matrix was analysed with the scheme. Results showed that molybdenite and arsenic-bearing sulfides do not behave as predicted from just their sulfur oxidation states, indicating that samples containing these minerals need special treatment for complete phase recovery. Second, a natural mine-waste composite standard was analysed by the scheme. The sum of the phases was equal to the total sulfur concentration analysed independently. Each phase had a similar, but distinct sulfur isotopic composition. The results were verified with a laboratory prepared æproxyÆ having the same mineralogical composition as the standard, but with each phase having a very different sulfur isotopic composition. The sulfur isotopic compositions of the recovered phases from the æproxyÆ showed that the scheme effectively separated the sulfur with only minor cross contamination of phases. The speciation and isotopic analyses were applied to both mine-waste piles and naturally occurring alteration zones at nine mine sites. A case study from one site comparing the sulfur isotopic composition of minerals in an alteration zone and dissolved sulfate in proximal streams is presented. This case study shows the application of our data to differentiating sources of sulfate in the surrounding environment.

Jan 1, 2003

It is usually the case that the short-term plannerÆs role is occupied by the younger site mining engineer. ThatÆs because mine scheduling is generally seen as a mundane and thankless task. But, short-term planners play an extremely important role in maximising the value of the mining resource, generally the companyÆs biggest asset. A look at any mining operation today will expose the competing drivers behind the operational and strategic plans. Operations are too often faced with pushing short-term constraints. It is the responsibility of the short-term engineer to answer the question of æWhere to next?Æ. In answering this question the young mining engineer is in an environment overflowing with urgency and constraints. Hasty decisions made at this time may have significant adverse consequences to the long-term value of the resource. Long-term strategy must address a broad and expanding number of issues that are generally overlooked by the short-term planner. It takes a holistic, rather than the prescriptive approach of short-term planning to determine the best strategic direction for the operation. To better understand the consequences of decisions made at the operational level short-term planners need better vision of the whole business. The key to improving the mining business is a planning process that integrates the short-term planners role with the overall mining strategy. Future developments in the mine planning process will focus on bridging this ædisconnectÆ between the operational objectives and the long-term goals. Improvements will come from the tools which planning engineerÆs use, and the planning methodology surrounding them. There will be significant hardware improvements to improve the speed with which a short-term planner can evaluate plausible options. Software that maximises the use of the improved hardware will be developed. Complimenting these technological improvements, conceptual methodologies of mine planning will improve to allow strategic planning to be brought within the focus of the short-term planning engineer. In todayÆs mining industry the young mining engineer is embarking on a career where business processes are paramount. This paper discusses the ædisconnectÆ that occurs today between the operational planning and long-term strategy and how this is driving developments in mine planning today.

Jan 1, 2003

By M Darvall And R Fare, D Dormer

The Argo Archaean lode gold deposit is hosted within a structurally complex shear zone system, 70 km south of Kalgoorlie, and 3 km west of the crustal-scale Boulder û Lefroy Fault. Argo is an active underground mine currently producing at a rate of 120 koz pa with a Probable and Proven Reserve as at December 2005 of 2.19 Mt at 5.94 g/t Au (419 koz), a reduction on the December 2004 Probable and Proven Reserve of 4.58 Mt at 5.74 g/t Au (845 koz). This Reserve is hosted in six separate gold bearing lodes within the Argo shear system. Presently, a further six subeconomic shears are recognised. The 2005 Reserve figure represents a reduction of 426 koz compared to the 2004 Reserve as a result of a thorough review of the geological interpretation/resource model, and engineering factors applied to the Resource/Reserve, and depletion due to mining. Year on year Resource/Reserve figures at Argo had gradually increased as additional information on the system became available and mining methods were increasingly resolved with early production. The portal for Argo underground mine was cut in June 2002 within the highly profitable Argo open pit. The proposed æAvocaÆ bottom-up mining method required extensive lead time between development and the commencement of stoping in April 2004. During development considerable additional complexity in the Resource was encountered. Additional structures were identified, and complex splay interactions affecting the geology impacted on development. Intensive concurrent exploration in the footwall to the main lode during this period was the main cause of the increase in the Resource/Reserve figures during early development. As development progressed, lower grade continuity was observed than was previously interpreted. At the same time, drilling of deeper portions of the deposit and subsequent reinterpretation of the geological model significantly reduced strike lengths of economic mineralisation. Radical changes to the mining methods were required to improve productivities and to satisfy geotechnical considerations. The changes altered the engineering factors applied, which resulted in a reduction in the Reserve figures.

Jan 1, 2006

By R Austerberry, S Green

Highly weathered ground, limited collar options and the need to achieve a desired outcome at the lowest possible cost led to the need to adopt a construction approach that had no established design methodology. The newest ventilation shaft at Mandalay Resources’ Costerfield Operations had to be established to a depth of 30 m through poor ground best described as ‘sticky swelling clay’, some of which was contained within an old backfilled pit.The ability to excavate a traditional presink was not possible due to limited space and the introduction of men into the shaft was to be avoided due to the associated safety risks and corresponding increase in cost. Piling in conjunction with raise boring and remote fibrecrete application was identified as a possible solution that could achieve the desired outcome at an acceptable level of risk.Construction of the piles was challenging and required continual refinement until an acceptable level of productivity could be achieved. Limitations on concrete delivery was taken into consideration to ensure hole stand-up time was acceptable and re-drills were not required. The adoption of a set process from alignment, drilling, concrete placement and testing ensured the desired design was achieved. Mine operators should consider the use of piling as a method of shaft stabilisation for the portion of raise bored shaft that passes through the weathered zone.The importance of quality control and attention to detail through the entire shaft construction process cannot be underestimated as issues such as insufficient geotechnical investigation, pile and pilot hole deviation, concrete and fibrecrete quality, as well as delays in the application of fibrecrete can place the success of the project at risk.This paper outlines key learnings from a client’s perspective with regards to the concept, planning, design, execution and completion of a 3.1 m diameter raise bore shaft through both weathered and highly structured ground at the Augusta Mine.CITATION:Green, S and Austerberry, R, 2014. Piling and raise bore construction – key learnings from a client’s perspective, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 35–42 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By A Baensch, E Grimsley

Eloise Copper Mine is situated 56 km east-southeast of Cloncurry, Queensland, Australia. The deposit was discovered by BHP Minerals in 1986 during follow up diamond drilling of EM anomalies associated with major aeromagnetic anomalies. The deposit is overlain by 60 m of Mesozoic cover. Amalg Resources NL commenced mining during 1995 by means of decline access and long hole-stoping. The total mineral resource comprised 1.77 million tonnes at 4.10 per cent Cu as at the end of June 2003. Proved and probable reserves are 930 000 tonnes at 3.97 per cent Cu. Mineralisation is hosted within a strongly foliated Proterozoic meta-sedimentary sequence comprising arenites and schists. The metasediment sequence also contains a coarse-grained amphibolite body possibly representing an early intrusion of gabbroic composition. Economic mineralisation characteristically occurs as a number of steeply plunging lenticular bodies with strike lengths of between 100 and 200 m and maximum widths approaching 25 m. The principal orebody, the Levuka (B) Lode, displays a down plunge extension exceeding 1000 m and is open at depth. Ore mineralogy is almost exclusively chalcopyrite with the principal gangue phases consisting of pyrrhotite + quartz + carbonate. A biotite + hornblende + magnetite assemblage represents locally intense mafic alteration of the metasedimentary host sequence. Magnetite is present in variable amounts and may be locally significant in some of the smaller satellite orebodies. Within the orebodies the sulfides display a range of textures from massive through to vein like. Sulfides commonly display brecciated textures at a range of scales. Post-mineralisation faulting has severely dislocated the orebodies, resulting in a complex arrangement of fault bounded ore blocks. These faults display considerable variability in regard to strike, dip and amount and direction of throw. A recent review of the fault system by means of large-scale structure underground mapping, re-interpretation of drill core and 3D modelling has led to a significant improvement in the understanding of these structures. A significant upgrade of the geological database has accompanied the structural review. The overall improved understanding has aided the mine design processes with particular regard to decline and access development and pillar location. This has also impacted on the design of drilling programs and exploration for additional satellite ore zones.

Jan 1, 2003

By Davey KJ

Complex sulphide ores are characterised by complex texture of the valuable minerals. Any assessment of the performance of differential flotation circuits treating such ore should include analysis of the effects of poor liberation of the valuable minerals. Selected samples were taken from balanced size by size surveys of the flotation section of a concentrator treating an ore body containing fine grained galena and sphalerite in a matrix of iron sulphides which is mostly pyrite. liberation measurements were made using the QEM*SEM system on four or seven size fractions. The measurements indicated that less than 40 per cent of the galena and less than 60 per cent of the sphalerite in the flotation feed was liberated. The final concentrates were largely composed of composites of valuable minerals and iron sulphide however there was a small but significant fraction of liberated iron sulphide. Using the sized mineral flows obtained from the original balances and the liberation measurements, a satisfactory liberation balance was obtained. This enabled the effect of liberation on the flotation performance to be examined; not only for the entire circuit but also in the rougher, scavenger and cleaning sections. In the zinc circuit the effect of a regrind mill could be assessed. Recovery of galena in the lead circuit was limited by the recovery of composite lead particles. Recovery of liberated galena exceeded 95 per cent in all size fractions while the recovery of composite galena exceeded 85 per cent below 38 ¦m but fell sharply above 38 ¦m. Examination of the behaviour of the liberated and composite galena in the various sections of the lead flotation circuit revealed high recoveries of both liberated and composite galena particles in the roughing and scavenging sections but very low recoveries in the final two cleaning stages. Similar trends to those observed in the lead circuit applied in the zinc circuit although the performance of the liberated sphalerite also fell off at sizes greater than 38 pm in this case. Thus the zinc recovery was limited by the flotation of both the liberated and the composite sphalerite. The effect of regrinding was to increase the recovery of sphalerite mainly by reducing the amounts of sphalerite in less liberated sizes rather than by increasing the liberation of the sphalerite within a given size Fraction.

Jan 1, 1993

The Ranger Project Area covers 78.6sq km and is approximately six km east-west by 14km north-south (Fig. 1). The area is known to con- tain two uranium orebodies, Ranger No. 1 and No. 3, two uranium occurrences, Anomalies 4 and 5 and three significant radon anomalies, Ranger 18, 19 and 48. The Ranger No. 1 and No. 3 Orebodies con- tain approximately 120 000 tonnes of U308 in highly chloritised metasediments of the Lower Proterozoic Cahill Formation. The orebodies occur as suboutcrop and exhibit significant radiometric anomalies. These occur within approximately 400m vertically of the projected unconformity with the sandstone of the Middle Proterozoic Kombolgie Formation. Oxidised and reduced uranium minerals occur in quartz-rich schists containing variable muscovite, chlorite and sericite and in the chloritic matrix of chert breccia. Secondary uranium minerals namely saleeite, metatorbernite and gummite associated with earthy iron oxides occur from the surface to 50m depth in No. 1 Orebody and up to 80m depth in No. 3 Orebody. Below the oxidised zone, uranium occurs as pitchblende finely disseminated throughout sequences of quartz muscovite chlorite schist up to 150m thick and as open space fillings in zones of structural deformation. The uranium-bearing schists directly overlie up to 300m of dolomite and magnesite containing minor interbedded schist units and are themselves overlain by approximately 2 000m of more micaceous schists of the upper part of the Cahill Formation. Regional metamorphism has occurred to at least amphibolite facies and has been followed by retrograde metamorphism to a greenschist facies equivalent. The metasediments are intruded by granitic pegmatites and dolerites. Uranium is associated strongly with chlorite of which there are at least five textural varieties. Not all of the various generations of chlorite have associated uranium. Graphitic schist horizons occur which are often less than one metre thick and discontinuous. These are best developed at the base of the quartz chlorite schist sequence where uranium grades up to 1 000ppm U may extend for several hundred metres beyond the main ore zone. Uran- ium is not common in the carbonate sequences and only occurs in chlorite segregations in disrupted zones. The unoxidised ore contains sulphides, mostly pyrite, which are low in abundance (

Jan 1, 1984

By C J. Ptacek, J G. Bain, D W. Blowes

The passive interception and in situ treatment of groundwater using permeable reactive barriers (PRBs) has, within the past decade, become an accepted remedial approach to the control of plumes containing dissolved organic and inorganic contaminants. Since the first applications of the technology in the early-1990s, there have been several field-scale PRBs that have provided excellent treatment of dissolved metals in groundwater. Additional commercial applications and demonstration projects continue to be developed. Three field-scale applications and supporting laboratory columns have shown that several thousands of milligrams per litre of sulfate, more than 1000 mg/L iron, and several tens of milligrams per litre of other metals can be removed from groundwater by PRBs that promote sulfate reduction. The reactive materials, which incorporate various forms of organic carbon, promote microbially mediated sulfate reduction, the generation of hydrogen sulfide, and the subsequent precipitation of sparingly soluble iron and other metal sulfide minerals. Effective performance has been achieved by PRBs installed at a mine-tailings impoundment in 1995 (Sudbury, Ontario) and at an industrial facility (Vancouver, British Columbia). These PRBs have removed iron and other metals from groundwater. Similar materials were also used to create reactive layers within test cells directly in a tailings impoundment in 1999 (Timmins, Ontario) with the objective of removing iron and sulfate from the pore water before it migrates from the tailings impoundment. Zero-valent iron (ZVI) has been shown to provide excellent removal of chromium (VI) in a PRB, which was installed at the US Coast Guard facility in Elizabeth City in 1996, by reduction and precipitation reactions. Laboratory tests have also shown that ZVI reduces As(V) to As(III), which is subsequently removed from solution by co-precipitation with iron oxyhydroxides. In groundwater containing sulfate, ZVI and mixtures containing organic carbon can promote sulfate reduction and the subsequent precipitation of sparingly soluble arsenic and other metal sulfides. A trial PRB system was installed in 2001 and was shown to attenuate arsenic and other electroactive metals in groundwater at an industrial facility in northwestern Ontario. Basic oxygen furnace (BOF) slag, a waste material generated by the steel industry, has shown excellent potential for the removal of phosphate, elements such as arsenic and water-borne pathogens from groundwater in field and laboratory tests.

Jan 1, 2003