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HISTORY OF PROCESS DEVELOPMENT The manufacture of coke from coal commenced at the beginning of the 17th century by heaping coal into circular mounds in the open, covering with earth, clay or ash and slowly burning the coal in the mounds to produce a heap of coke. Naturally the process was slow in production rate with low coke recovery due to the loss of that part of the coal which was completely burnt to ash. As a natural consequence of this initial carbonisation method, a coke oven design was developed so that the coking chamber, which was constructed of brick, was circular in plan, cylindrical in section with a dome shaped roof - similar in shape to the original coal mounds. As the shape of the chamber was similar to the old dome shaped beehive, the ovens were called beehive ovens. The ovens were built in long single or double rows with retaining walls between each oven and as such were known as a battery of beehive coke ovens. In these early beehive ovens, coal was charged from hand wheeled tubs into a hole into the top of each oven, hand- levelled with a rake and the oven side opening sealed with bricks and clay except for some small holes in the opening to admit limited quantity of air for combustion. Coking commenced by the heat retained in the brickwork from the previous charge. The gases evolved from the coal, together with the admitted air, were partially burnt inside the oven to supply the heat for carbonisation and the surplus gases were burnt to waste at the top charging hole. The heat from this gaseous combustion inside the oven together with the reflected heat from the brickwork of the arched roof of the oven allowed carbonisation of the coal charge to proceed from its surface slowly downwards to the floor of the oven. At the end of the coking cycle (usually 4 or 5 days) when there was no more evolution of gases, as indicated by diminishing gas flame and a drop in oven temperature, the brickwork of the oven opening was removed. The coke was then quenched by inserting a pipe with spray holes to spray water over the coke charge in the oven after which the coke was raked from the oven. The work of charging and levelling the coal, bricking and removal of oven opening, quenching and raking out the coke all required manual labour. As these ovens only carbonised a 5 to 7 ton charge of coal per cycle of 4 to 5 days, to produce 3 to 41/2 ton of marketable coke, the labour required per ton of coke was high.

Jan 1, 1975

Life inside the mining family was regulated by the mine. The whistle shrieked the change of shift to the Cornish, Welsh and German miners digging the alien soil for copper at Burra in the 1840s and the deep quartz gold mines at Bendigo in the 1880s.During the golden years of alluvial mining in the 1850s, and later out on the frontier fields of Kalgoorlie and Jarrahdale in Western Australia, the lure of the mine dominated family life. How could it do anything else? The canvas tent, bark hut or weatherboard cottage in which the mining family lived close by the mine facilitated prompt access. The working man of the family attended to the demands of the mine first, and those of his family afterward. Who then organized and nurtured the mining family? The role of the mining wife was multi-fold. She played the traditional roles of home-maker and comforter, raising her children, tending her house and vegetable garden and supporting those friends and acquaintances who had lost children to disease, or menfolk to the mine. Epidemics of measles, the whooping cough or influenza carried away children in infancy or at too young an age. The death of a husband through accident or disease (pthisis was a likely killer) forced upon the mining wife another role, that of provider. When Elizabeth Vigus' husband died in a Bendigo mine accident in March 1880 Elizabeth set up a grocery business to support her young children.' But in other cases the mining wife supplemented the family earnings by part time work, not in the mines for the Mines Act legislation of 1842 forbade the working of women underground, but in auxilfiary occupations which were in demand.2 Ann Mitchell, a herbalist ,found plenty of work in the goldfield towns of Daylesford and Maldon in the 1860s and 1870s.3 So too did Susan Rule and her daughter Catherine who conducted a sewing and millinery establishment in Beechworth in the 1870s.4 The making of the family's clothes was another task of the mining wife and mother.Until the wonderful invention of the sewing machine, this was a tedious and time-consuming task. James Slater Edmonson bought his wife a Wertheim for the sum of 12 pounds in June 1886, just before leaving her and his chil- dren in the Adelaide suburb of Goodwood whilst he went alone to the goldfields of the West. In October of that year he wrote to Louise telling her how he missed the childrens' wondering "whether Slater wore my old shirt now. I always think of him as he sat on my knee at breakfast in that old shirt; and how pleasant it would be if I could rock both of them and sing as I used to do".5

Jan 1, 1988

By N Murphy and, S Santomartino, M Cramer

Acid rock drainage (ARD) can have detrimental effects on mining infrastructure, water reuse options and environmental discharge. As a result, some form of short- or long-term treatment is required at many mine sites. Successful treatment depends on a number of factors including the selection of an appropriate treatment technology and its correct implementation. A variety of passive and active technologies are available for dealing with ARD. Passive treatment solutions do not require continuous chemical inputs but take advantage of naturally occurring chemical and biological processes. They are often low cost, site specific approaches and are predominantly used in post-closure situations. Passive systems are well suited to drainage with low acid loads and low soluble aluminium concentrations, and are more effective with highly reduced (ie low oxygen) water. While they are often regarded as permanent fixes, most require some form of ongoing maintenance and all have an engineered life expectancy (ie finite carbonate and organic matter budgets). Passive treatments include, oxic limestone drains, anoxic limestone drains, limestone diversion wells, various reducing and alkalinity producing systems, pyrolusite limestone beds, aerobic and anaerobic wetlands, permeable reactive barriers (organic ¦ carbonate and zero valent iron), slag leach beds, and gas redox and displacement systems (GaRDS). Active treatment solutions require a continuous or semi-continuous activity (eg reagent dosing) to correct water quality issues. They can be engineered to treat virtually any type of drainage or site specific ARD issue, and have no set acid load or redox limitations. Tasks handled by active treatment systems are almost invariably larger and more costly than for passive systems. Active treatment approaches include; pH control, adsorption/absorption, ion-exchange, electrochemical concentration, biological mediation/redox control, flocculation/filtration and crystallisation. A range of generally available and proprietary technologies has been developed around these approaches. Active treatment involving pH correction with alkaline amendments is the most common and cost-effective form of ARD treatment. A variety of fixed plant and portable treatment systems are available. Calcium-based reagents, including quicklime (CaO), hydrated lime (Ca(OH)2) and limestone (CaCO3) remain the reagents of choice for the majority of active ARD treatment worldwide. This is principally due to their widespread availability, non-proprietary nature, and variety of proven dispensing technologies and cost-effectiveness. Limestone, in particular, is a very low cost reagent. However, due to æefficiency of useÆ problems with limestone in passive treatment systems, combined with its very low cost, there has been a recent emphasis on developing active treatment systems that use limestone efficiently. As the requirement for treatment increases, there is a growing emphasis on treatment sludge management and sludge stability. The final choice of a suitable treatment technology is dependent upon water chemistry (eg pH, type of metal contamination, acidity), flow rates, acid load, local natural resources (neutralising materials, topography), climate, site logistics, environmental risk and economics. The benefits and limitations of both passive and active treatment technologies are summarised.

Jan 1, 2003

By Williams RE, Ralston DR

Acid mine drainage is a serious environmental and economic concern, resulting in regulation of the mining industry and significant water treatment costs. An interdisciplinary approach is needed to analyze the many aspects of a complex acid drainage problem. This paper describes the integrated application of techniques from the following fields to the characterization of acid drainage at a large underground lead-zinc mine: 1) hydrogeology, 2) surface stream analysis, 3) structural analysis of fracture distribution, 4) hydrochemistry, 5) geostatistics, and 6) multivariate statistics. The Bunker Hill Mine in northern Idaho, is developed in highly fractured and faulted quartzites and discharges 95 L/sec of water with an average pH of 2.8 and an average zinc concentration of 120 mg/L. Research has been conducted on all aspects of the acid water production and drainage problem. Recharge to the mine has been investigated using geophysical techniques and tracer studies. Fracture mapping and geostatistical analysis of the fracture data have been used to determine characteristics of inflow to the mine workings. Hydrograph analysis techniques have been applied to water flow data within the underground workings. Chemical sampling and analysis have been used to identify acid producing workings, and better understand mechanisms controlling acid water production. Multivariate statistical techniques have been applied to analyze the quality and flow data. Research results from the many subprojects of the Bunker Hill study reveal the importance of an interdisciplinary approach to the analysis of acid mine drainage problems.

Jan 1, 1988

This paper is intended to continue the subject in the light of recent work, from where it was left at the conclusion of the discussion on the excellent paper by the late Mr. Noel Haszard. The position then obtaining is briefly summarised below:-The points which received general agreement at that time were:-'(1) That the gold occurs, probably in an elemental form, in a state of very fine division in the pyrite.(2) That the aim in roasting of Kalgoorlie flotation concentrates should be to obtain the maximum porosity possible. The gold tellurides were amenable to treatment with either a' solution of high lime and cyanide strength combined with fine grinding, or bybromocyanidation without re-grinding.(3) That a low roasting temperature is essential during the elimination of the first atom of sulphur, if the final calcine is to give a good extraction by cyanidation.(4) Even with the best temperature conditions of roasting then determined, and after a period of cyanidation of the calcines until no further extraction was obtainable, the lowest calcine residues obtainable were equivalent to approximately 1·8 dwts. per long ton, based on an original concentrate of pure pyrite.There was some division of opinion at the time as to whether the temperature of the later stages of roasting was significant, and also as to the reason why the initial roasting temperatures had so marked an effect.

Jan 1, 1939

By K Voigt, B Bolton, C Rumble

The Ok Tedi copper and gold mine is owned and operated by Ok Tedi Mining Limited (OTML) and is located in the Star Mountains of the Western Province of Papua New Guinea. Waste materials (both sulfidic and carbonate materials) from mining operations, comprising 55 Mt of waste rock and 31 Mt of tailings per annum, are discharged into the headwaters of mine area creeks where erosion and sediment transport carries significant proportions of the total waste into the Ok Tedi and Fly River systems. The tailings are potentially acid forming (PAF) while the waste rock has been strongly acid consuming with an excess of acid neutralising capacity (ANC) due to the high proportion of limestone waste. Tailings and waste rock are eroded into the river system and transported to the lower Ok Tedi and floodplains where deposition takes place. Because of concern due to vegetation die back, a dredging operation commenced in March 1998 to remove the sand fraction from the main river channel. OTML identified a potential future ARD risk from the stockpiled sands due to reduce limestone production at the mine and sulfide/carbonate mineral segregation in the river system and in the dredge sand stockpiles. This paper describes the structure and framework of an empirical model developed to assist OTML in planning ongoing mining operations to control ARD. Modelling identified the need to provide additional mined limestone to prevent future ARD generation from material deposited in the mine area creeks. With the additional limestone, the ARD risk in waste rock storage areas and in the Ok Tedi and Fly River sediments is predicted to remain low due to the combined effects of high ANC, high alkalinity of natural waters and high moisture contents (mostly fully water saturated). However, monitoring and modelling indicate that the ARD risk from the stockpiled sands will increase substantially during the remaining years of the operation. OTML have initiated programs to identify and evaluate options to mitigate this risk.

Jan 1, 2003

Until recently, nickel laterite projects have been largely ignored, due to their supposedly higher treatment costs compared with sulphides, and to their perceived processing difficulties. A number of laterite projects in the 1970s were abandoned because of high capital and operating costs, while others experienced severe technical difficulties. Conventional wisdom had it that laterites would not be exploited until most of the sulphide orebodies had been depleted. Why then, have Australian mining and exploration companies started a resurgence of interest in the treatment of nickel laterites? It all began with the Bulong project, which has been under study since 1987, while more recently two further projects, Murrin Murrin and Cawse, have been initiated and are being fast-tracked to take advantage of the current high nickel and cobalt prices. A further project in NSW, Syerston, has also been investigated at bench-scale, but no further development has yet been announced. All these projects have the following advantages: 1. Good location near a major mining centre and,reasonably close to power, water, roads, accommodation and a developed service industry. 2. Straightforward, low cost mining. 3. Low cost of acid or sulphur. 4. Good to very good cobalt credits. In addition, there have been significant technological developments in the design and operation of high pressure autoclaves and high rate thickeners, as well as solvent extraction and electrowinning technology. This has helped to decrease capital costs and technical risks, and has allowed the development of alternative flowsheets to the existing laterite operations (Taylor, 1996).

Jan 1, 1996

By Bellairs P, Stewart JC

On 20 December 1983 at 8.40am blast- hole 57 of Blast RN177 prematurely detonated follwed by a violent deflagration of blast- hole 59 at 10.40am. The blastholes were 380mm diameter drilled in black pyritic Mt McRae shale. The blastholes were loaded with ANFO and primed. Blasthole 57 was stemmed and blasthole 59 was not. A poss- ible mechanism which could lead to such a detonation is described. preliminary tests showed the top 5m of a bench had the highest temperature which could not be influenced by a range of treat -ments. Small scale laboratory tests show- ed that zinc oxide (ZnO) and magnesium carbonate (Mg2CO3) inhibited the reaction. A series of tests using AN with and without inhibitors loaded into the collars of blast -holes drilled in the top 6m of the bench and unstemmed is described. An analysis of the time to reaction of the holes loaded with AN against temperature shows that the decomposition of the AN in contact with the black pyritic shale is probably a sulphuric acid catalysed react- ion. All the inhibitors tested can neut- ralise the sulphuric acid in the shale. 10% ZnO appears to be an effective inhibitor of the reaction providing the AN/ZnO (90/10) is kept out of contact with surface material whose temperature can be as high as 60¦C in summer. It is concluded that ANFO/ZnO (90/ 10) should be now tested in stemmed holes.

Jan 1, 1986

By M H. Randell, J B. Haines

The Honeymoon Uranium Project comprises the Honeymoon Deposit,Goulds Dam Deposit and Yarramba Prospect in the Southern Lake Fromeregion, South Australia. Both the Yarramba Prospect and HoneymoonDeposit (which includes Honeymoon and East Kalkaroo orebodies) arelocated in the Yarramba Palaeovalley. The Goulds Dam Deposit is about75 km north west of Honeymoon, in the Billeroo Palaeovalley. Exploration for sediment-hosted uranium began in the area in the late1960s, culminating in the discovery of Honeymoon and Goulds Dam inthe early 1970s. In 1982, a 25 L/s demonstration plant was built atHoneymoon to confirm suitability for in situ leaching. The project was put on hold in 1983 due to changes in government policy. Southern Cross Resources Australia Pty Ltd acquired the project in mid 1997. Roll-front deposits form from a migrating geochemical cell, anadvancing reduction-oxidation interface between oxygenateduranium-bearing groundwater and its reduced aquifer host. The source of metal is uranium-anomalous granites, which were eroded fromsurrounding ranges. The weathered granites and resultant sediments are stripped of uranium by oxidised groundwaters, to form solutions carrying uranyl carbonate complexes. The solutions percolate down-slope through permeable sand zones until contacting a reduced environment where uranium precipitates. Uranium mineralisation occurs interstitially between and as thin coatings on sand grains, usually in the form of uraninite or coffinite. The palaeovalleys (previously termed palaeochannels) are incised intounderlying Cambrian/Precambrian basement rocks and filled withsemi-consolidated, largely un-cemented, Tertiary sediments of the Eyre Formation. The late Palaeocene to middle Eocene Eyre Formation is the basal unit of the Tertiary succession in the Callabonna Sub-basin of the Lake Eyre Basin. Further to the north, the Lake Eyre Basin overlies the Jurassic-Cretaceous Eromanga Basin, which comprises much of the Great Australian Basin. Eyre Formation sediments are in turn overlain by the late Tertiary Namba Formation and terrestrial Quaternary units. At Honeymoon, the Eyre Formation includes immature, pyritic,carbonaceous sands and gravels interbedded with lignite andkaolinite-illite-montmorillonite clays. It was deposited by braidedstreams during epeirogenic uplift of the Olary Ranges, accompanied bysubsidence of the Lake Eyre Basin. The Eyre Formation comprises aconsistent stratigraphy of interbedded sand and clay units that arecontinuous over the Honeymoon Deposit but with local variationindicating some hydrological connection between aquifers.In the Honeymoon orebody, uranium occurs within coarse-grained,pyritic sands where the basal sand pinches out between overlyingcarbonaceous clay and the southern palaeovalley slope. Its characteristics are typical of the well-developed, planar limb of a roll-front uranium deposit. Zonal changes related to the reduction-oxidation interfaces across the deposit are readily observable. The orebody extends for nearly 1000 m along the valley margin, is 400 m wide at its maximum and averages about 6 m in thickness. The East Kalkaroo orebody also lies near the southern margin of theYarramba Palaeovalley. It extends discontinuously for approximately 3.5 km and averages 200 m width, with significant uranium present in the Middle as well as the Basal Member. The Billeroo Palaeovalley extends northward for 40 km sub-parallel tothe western side of the Benagerie Ridge, is up to 5 km wide and hasgently dipping banks. Eyre Formation within the palaeovalley isconcealed beneath 90 m of Quaternary and Namba cover, isapproximately 40 m thick, and is sub-divided into three upward-finingunits, which are remarkably similar to the Yarramba Palaeovalleysequence.

Jan 1, 2001

By J D. Bryant

The Ernest Henry Copper-Gold mine, the third largest copper deposit in Australia, is located 35 km NE of Cloncurry in the Mt Isa-Cloncurry mineral district of NW Queensland. Ernest Henry Mining Pty Ltd (EHM), a wholly owned subsidiary of Xstrata Copper Australia, operates the mine. Approximately 10.2 Mt of ore grading 1.12 per cent Cu and 0.56 g/t Au were processed during the 2001 - 2002 financial year. The Ernest Henry plant produces copper-gold concentrate, which is trucked to Mt Isa for smelting. The Ernest Henry deposit is a member of the diverse Fe-oxide-(Cu-Au) class of deposits. The orebody is hosted within a sequence of moderately SSE-dipping, intensely altered Paleoproterozoic intermediate metavolcanic and metasedimentary rocks. Copper (chalcopyrite)-gold mineralisation occurs mainly within the magnetite-biotite-calcite ¦ pyrite matrix of a 250 m + 300 m pipe-like breccia body. The breccia pipe comprises two anastomosing lenses separated by a lower grade, clast-supported breccia zone. Several anastomosing brittle faults affect the macro-geometry of the orebody. Copper grade is proportional to the degree of matrix support in the breccia. The July 2002 Ore Reserve was 88.7 Mt @ 1.0 per cent Cu and 0.5 g/t Au. Resource estimation at Ernest Henry is based on a large database of predominantly diamond drill hole information. Drill spacing is typically 50 m + 50 m. Ordinary kriging is used to estimate Cu, Au, SG and IST (oxidation indicator) into model parent cells of dimension 20 m E + 20 m N + 16 m RL. The block model is comprised of five geological-geochemical domains represented by 3D wireframe surfaces. Zonal control, where samples and blocks are assigned the same domain number, is used to control grade estimation. The geometry of the domains defines hard grade boundaries across strike and soft boundaries along strike A Deep Drilling Project was initiated in August 2002 with the initial aim of testing the down-plunge extension of the orebody to a vertical depth of over 1 km. The encouraging intersections to date have resulted in a significant increase in the mineÆs open pit Mineral Resource together with the identification of an underground Mineral Resource. Added benefits of the project included an expanded and updated geological legend and an improved understanding of the host geological sequence.

Jan 1, 2003

By M Robotham

In April 2013, Rio Tinto Kennecott Copper’s (RTKC) Bingham Canyon mine experienced what is arguably the world’s largest ever in-pit slope failure. The failure initiated along a major, continuous, low-strength bedding fault, named the Manefay bed, and comprised approximately 145Mtonnes of rock and waste dump material. Initial East wall slope deformations were detected some months prior to catastrophic slope collapse by RTKC’s slope monitoring systems. However, the failure mechanism was different to that which had historically been experienced in the prior more-than-100 years of open pit mining. Failure occurred as two separate events, the initial one likely consisting of a frictional failure of rock materials along the Manefay bed itself, before a second phase of failure occurred, which included a large mass of old rail dump material. The combination of these events led to a much larger failure run-out than had been anticipated by the RTKC team which led to loss of, and damage to, mining equipment and infrastructure as well as significant interruption to mining activities. Due to the diligent monitoring of the slopes no safety impacts were experienced at the mine, and as such the details and learnings from this failure can readily be shared with the broader mining and geotechnical industries to avoid similar recurrences elsewhere. This is an extended abstract only. CITATION:Robotham, M, 2018. Bingham Canyon Mine – from adversity to excellence, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 3–4 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By J Player

Big Bell Mine started experiencing relatively large mining related seismic events, and accompanying rockbursts, from February 1999. Twelve rockbursts damaged access development from February 1999 to June 2000, with severe damage extending for up to 100 m along footwall drives. Five seismic events associated with the rockbursts were recorded by the AGSO, measuring 1.9, 2.2, 1.7, 2.4 and 2.2 on the Richter scale. These events affected 485 and 460 footwall drives in August and November 1999 and 535 footwall drive in June 2000 (drive level nomenclature is equivalent to the depth below surface and RL). After the third event (July 1999) it was apparent that the seismicity was not a once-off phenomenon. This (sudden) onset of seismicity required management to implement a series of precautionary measures:to ensure the safety of the workforce, to maintain continuity of production, and to reduce rockburst damage rehabilitation. The measures included extended re-entry periods and exclusion zones in high-risk areas, improved rockburst resistant support systems in areas prone to damaging seismic activity and improved support systems in areas prone to shakedown damage. A full-scale minewide ISS seismic system was ordered in October 1999 and commissioned in February 2000. A portable eight-channel CSIR-Miningtek system was installed at short notice as an interim measure in August 1999. This CSIR system covered the northern producing stopes, where the majority of the rockbursts had occurred. The objective of the seismic monitoring was to monitor relative activity and to determine trends with regards to time and location. MAP3D numerical modelling was used to investigate the relationship between seismicity and various combinations of stress components to assess the relative seismic risk in both current and proposed working areas. This information is used to determine where and at what stage in the mining cycle the rockburst resistant support is required. Analysis of the seismicity has determined two distinct modes of seismicity û small-scale localised, high frequency events in high stress areas and larger events related to shear along previously intact foliation surfaces. This paper includes a summary of the seismic history, seismic data analysis, support system design, mine design adjustments and management action to minimise the effects of seismic activity at Big Bell Mine.

Jan 1, 2000

By K S. Smith, B P. Jackson, P J. Lamothe

The upper Arkansas River south of Leadville, Colorado, USA, contains deposits of fluvial tailings from historical mining operations in the Leadville area. These deposits are possible non-point sources of acid and metal contamination to surface- and ground-water systems. We used stable-metal isotopes to help ascertain metal retention and release mechanisms that influence metal transport from the deposits to shallow ground-water systems. To accomplish this, we excavated an intact core from a small fluvial tailings deposit and performed laboratory column experiments to examine the amount of metals leaching through the core. Deionised water was continuously applied to the top of the core, and effluent was collected from the bottom of the core for a period of about two months. Three saturation conditions were used, including unsaturated, partially saturated, and fully saturated. Reducing conditions developed upon partial and complete saturation of the core. During the unsaturated leaching phase, core effluents were acidic (pH 2.8 - 3.5) and contained elevated metal concentrations. During a portion of the unsaturated leaching phase, stable-metal isotope spikes and conservative tracers were simultaneously applied to the top of the core. The isotopes provide a means to delineate metal behaviour within the core. During the unsaturated leaching phase, the isotope spikes did not exhibit breakthrough with simultaneously added conservative tracers. However, some of the stable-metal isotope spikes did break through when reducing conditions occurred in the core (during the partial- and completesaturation conditions). In this paper, we use Cd as an example of the utility of stable-metal isotopes in geochemical studies. In the case of Cd, after the core was partially saturated, the 111Cd spike was released as evidenced by a change in the Cd isotope ratios in the effluent. This release continued during the fully saturated leaching phase, however, the total Cd concentration did not increase. These results suggest that the 111Cd spike was retained inside the core during the unsaturated leaching phase, and only partially released as reducing conditions developed. Results from this core-leaching experiment indicate there is a large reservoir of water-soluble material within the fluvial tailings deposit, which yields elevated metal concentrations and high acidity, and which may degrade adjacent ground- and surface-water quality. Use of stable metal isotopes in this study facilitated the determination of different metal-retention processes, metal-release processes, and metal sources in the fluvial tailings deposit in response to changing geochemical conditions.

Jan 1, 2003

There is growing acknowledgement that successful organisations, including mining companies, achieve their competitive advantage by better managing their employees, customers and other stakeholders. The fact that I have been asked to address this conference is perhaps evidence of the recognition of the importance of ensuring the improved management capabilities of the personnel within Australian mining companies. As many of you will be aware, prior to my appointment to the role of President of Minera Alumbrera Limited (MAA), the operator of the Bajo de la Alumbrera project, my career had been more generally focussed on industrial relations and human resources management. My appointment was met by surprise in some quarters, as it is unusual for a person without the technical background of a mining discipline to manage a world-class mining operation. However, it was recognised early in the life of the Alumbrera project that the operation did not lack technical expertise or a committed workforce; rather what was required was innovative management of the project and its operational resources, particularly given that we were operating with totally ægreenÆ workforce in a country without a large-scale mining culture. It is my experience that a particular shortcoming of mining companies is that they have tended to ægrowÆ managers based on their technical expertise rather than people management skills. Historically, many companies have assumed, or at least hoped, that their better technical people would develop the requisite skills to take on the challenges of management. Such philosophy has not always met with success. Whilst the topic for my presentation was originally published as People in the Mining Industry, within the context of the overall theme for this plenary session, People: Education for the Future, it is a subject with an extraordinary breadth of possibilities for discussion. Therefore, I have chosen to redefine my presentation to People Management Issues in Offshore Mining Projects,an issue which I consider to be an important area within the overall topic and one which has particular currency given the increased tendency by Australian miners towards offshore exploration. It is also relevant from the perspective of continued rationalisation of our industry, the recognition that there will be more large-scale, global corporate mining entities and that significantly more mining professionals will become æcitizens of the worldÆ. We therefore need to prepare our future leaders for these challenges. The topic also allows me to highlight the importance of an holistic management approach as the establishment of a new project in a foreign environment puts into stark relief the need for the application of the full range of competencies, rather than the narrow ætechnicalÆ approach to doing business. I hasten to add, however, that the mining industry is not alone in this regard and I was interested to read an article in the Weekend Australian of 22/23 January 2000, which reported on the first key speech of former Westpac CEO, Bob Joss, as Dean of the Stanford University Graduate School of Business. He said: à We know that when our students go out into the world, if they have to struggle, it will not be because theyÆre not smart enough, or they donÆt have enough good ideas. It will be because they struggle over managing themselves and their relationships to other people. The article then went on to say that Bob Joss foreshadowed an initiative to build alliances between the Business School and other Humanities departments within Stanford University to draw on their learning. Other business schools have also recognised this need, albeit somewhat late in the day, as have the most enlightened of modern corporations; in summary, successful businesses have recognised that in an age where competitive advantage through, for example, technological breakthroughs is very short-lived, it is the quality of management that will provide the leading edge. By using the Alumbrera project as a case study, I will identify certain of the challenges with which companies and individuals may be confronted in the establishment of a greenfields project outside Australia. However, many of the issues raised will have direct application in the context of the development of an Australian operation. My objective is to provide an indication of areas which should be considered at the formative stages of a projectÆs development, with particular focus on the issues of people management. I would expect that such a study would be of benefit to companies, education institutions and individuals in preparing themselves for similar experiences.

Jan 1, 2000

Background A recent development in many industries,' including the mining industry, has been ever increasing acceptance and adoption of extended (12-hour) shifts and compressed work cycles compensated by more time off work by employees between cycles. In the Western Australian mining industry development of `long-distance-commute' (LDC) coincided with adoption of 12-hour rotating shifts. Preludes to LDC were the `construction-style' rosters with two crews plus allowance for annual and sick leave relievers working the statutory 13 consecutive shifts followed by one shift off. Crews rotated day to night shift weekly or at fortnightly intervals. Average hours of 72 to 78 per week were common. The author had personal experience working these rosters in the East Pilbara of Western Australia from 1977 to 1982. Workers tolerated these rosters until a suitable combination of frustration, boredom, despair and the `Five Grand Fever' triggered them to `pull the pin' and head for relaxation in Bangkok, Manila, Nimbin, or Margaret River according to their personal tastes.

Jan 1, 1996

By Mining Geophysics

Modern geostatistics has eliminated much of the guess-work in ore reserve estimation, and can calculate levels of confidence for the reserve estimate. Geological modelling software has brought about a revolution in the communication of complex geo-variables to mining engineers, metallurgists, surveyors and managers. Mine planning and scheduling can now avail itself of a veritable toolbox of enabling technologies.

Jan 1, 1994

Sixty years after its inception, after an alternation of periods of admiration and rejection by the industry, Sampling Theory can now be seen as going through its golden age. A brief history of its successive transformations and progresses achieved by its author, Pierre Gy, and his followers, reveals the theoretical as well as practical difficulties that had to be vanquished to reach this point. Its current level of acceptance and corresponding trends are discussed and illustrated, with an insight on its increasing real role in the industry and future prospects for its further development.

Jan 1, 2008

Spectacular high-temperature hot springs (æblack smokersÆ) on the deep ocean floor are producing economic concentrations of base and precious metals. Mining these seafloor massive sulfide (SMS) deposits is about to become a reality in 2010, subject to the timely issuance of permits. Three Australia-based private companies and two government agencies inJapan and Korea have exploration licenses or applications covering vast areas of the seabed in the western Pacific targeting SMS deposits. Technological and financial challenges for recovery of seafloor sulfide deposits are being met and mining these may be less environmentally deleterious than mining on land. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By A Markham

Stream diversions are a common feature of mine sites worldwide. From an operational perspective, a fundamental requirement of a stream diversion is that it should convey water flows between two points in a safe, controlled and predictable manner. On the other hand, rivers and waterways are naturally dynamic systems, and many of their natural values are achieved through their ability to alter their forms and processes over time in response to flow events - often in a potentially unsafe, uncontrolled and unpredictable manner. In order to ensure that diverted river reaches achieve the dual objectives of short-term operational stability and longer term physical equilibrium with surrounding landscapes, alteration of form and process must somehow be understood and incorporated into the design process. However, performance targets for waterways are not always easy to define and assessing what is good and bad in such variable systems can be difficult. This paper discusses the role of the stream diversion from the operational, environmental and regulatory/relinquishment perspectives, and how stream diversions must achieve physical stability in the short-term, but dynamic equilibrium in the long-term.CITATION:Markham, A, 2012. Mine site stream diversions - A balance of competing objectives, in Proceedings Life-of-Mine 2012 , pp 215-218 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

By Horiszt G

his report is going to point to that the effect of the mining dewatering on the area of Transdanubian Mountain Range has reached a phase when the cost of dewatering and providing damages can seriously jeopardize a profitable mining operation. Up to 60s the interest of mining had been dominant only, but since the begining of 80s the conceptions of industrial development and the value of enviroment protection have been changed. The natural karstwater balance of the TMR can be presented (Figure Land 2.) and the present situation of it points to that the descharge has exceeded the reacharge in 50 %. During the past 10 years the decreasing of the stored karstic water resourches was 1.1 km . The decreasing of the spring yields reached 80-82 % of the natural ones. (Figure 3-4.). Conservation of the present discharge may lead up to the totaly drain of the karstwater regime. A computer aid control system had been to develop by the bauxite mining and using of it the yield of dewatering decreased and a moderate rising of the water table has started since 1986. A recunstruction was made at the biggest thermal-medicinal lake its springs have had their source in 40 meter depth inside a spring cave at the bottom of the lake. Earlier the spring yield decreased due to the dewatering activity of mines. As a result of using of the CAC system and the providing works at the thermal sprino,the karstwater balance has got in the state of equilibrum, the first cost was decreased and spring yield was increased and the enviromental limits were adverted from the way of mining.

Jan 1, 1988