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By Lilly PA, Kavetsky AP

An empirical, kinematic model is developed which calculates the trajectory of individual blocks of rock in a blast, their point of impact in the muckpile and the final shape of the muckpile. The method is based on the determination of initial block velocity and requires input on stemming length, bench height, subdrill, burden, spacing, energy of explosive per unit length of blasthole, energy attenuation rate and resistance to movement. The model, which will form part of a comprehensive blast simulation package, has application in open cut mining in predicting muckpile geometry and the movement of grade blocks during blasting, which in turn impact on shovel productivity and grade control.

Jan 1, 1986

By Murdaugh W

The application of photogeological techniques and analysis of more advanced remote sensor images can be a cost and time effective tool in the search for desert placer gold deposits. Examples are given of regional analysis using standard false color composit Landsat images and conventional black and white aerial photography. The examples include identification of sites to sample to evaluate the pathways of .alluvial gold migration and identification of potential sites of concentration. The results of these studies developed an increased number of target areas and exploration in proceeding with promising results. In addition to the development of additional target concepts and identification of specific target areas data useful to evaluation of mineability can also be obtained from photogeological investigation. Some indication at least can be obtained of the existence of duricrusts and salines that may hamper mining. Identification of possible sources of water is another important task in the desert environment. In addition indications of the clay content of the material to be worked is very important because of its economic impact on operational costs and overall recovery. The application of photogeology to these questions is discussed.

Jan 1, 1988

By J C. de Vries

When times are tough in the mineral industry û as they have been for some time û typical responses by mining companies are to cut costs and/ or increase production. The latter is seen to spread fixed costs, thereby also reducing unit costs per tonne mined. However, most mineral deposits respond to increased production rates by lowering ore grades, so the reduction in cost per tonne of ore does not necessarily result in a reduced cost per unit of metal produced. The authorsÆ experience is that this process often leads to a mine plan that is a long way from optimal. Producing to the point where marginal cost equals marginal revenue may work well in a manufacturing operation, but it is not necessarily the best strategy for a mining company with a depleting resource. Variations in tonnage and grade with depth, ore reserve optimism and a variety of other factors conspire to produce variations in production rates and grade that often result in suboptimal outcomes and potential mine failure. Added to this is the focus of industry analysts and senior corporate management on such measures as unit costs and ounces in reserves. This is perceived to have the perverse effect of driving the share price down when strategies are adopted that actually improve the value of the mining operation, and vice versa! Many mine failures can be prevented by a close examination of the tonnage grade curve and an understanding of how margins, net cash flows and resulting business risk change with cut-off grades and rates of production. A few relatively simple analyses early in the evaluation of a project can demonstrate the tradeoffs between the main decision parameters, leading to right-sizing of the operation. Once major commitments to major capital items have been made, and reserves have been publicly reported, optimisation of the operation becomes much more difficult than if it had been done earlier. Such investigations can also identify the upside potential and downside risks associated with factors outside the companyÆs control. Many strategies that may enhance value in the good times significantly increase the downside risk in the bad times. What is often called risk analysis typically looks at superficial issues, and does not evaluate the major factors leading to the potential for financial disaster. This paper illustrates some of these major risks, and demonstrates how they may be fully evaluated so that corporate decision makers can make fully informed decisions that take account of the real risks.

Jan 1, 2003

By McGuckin P. H, Logan A. S

The BHP Minerals' Cannington silver-lead-zinc deposit is located approximately 190 km south-east of Mount Isa. The deposit, is hosted by Proterozoic rocks beneath a thin cover of younger sediments, and contains a total estimated resource of 45.3 million tonnes at 11.1 per cent lead, 4.4 per cent zinc and 500 g/t silver. The deposit is separated by faulting to form a shallow, lower grade Northern Zone and a deeper, higher grade Southern Zone (Roche, 1994). Contract development of an exploration decline to test mineralisation in the Southern Zone commenced with the excavation of a boxcut in August 1993. The complex variably deformed Proterozoic sequence interpreted from drill core and intersected in the exploration decline has required a high level of practical rock mechanics input to achieve: ò appropriate support designs, ò a high quality of installation, and ò required development advance schedules. The Norwegian Tunnelling Index (Q System) in conjunction with regular face mapping is used as an operating tool to assist in decision making for ground support. Line mapping, discontinuity analysis and wedge assessments are conducted routinely to optimise ground support requirements in the decline, intersections and all other underground excavations. Throughout the project, an emphasis has been placed on developing a strong working relationship with the contractor to achieve quality assurance program targets.

Jan 1, 1995

By S Roberston, G Batt, R Shanahan, B Shah, A Trench, H Bromfield

The geographic imperative of ore systems leaves mining, more so than virtually any other aspect of modern industry, heavily dependent on obtaining and maintaining social licence to operate. A prospective miner cannot arbitrarily walk away from a deposit in search of more favourable community perceptions – concentrations of mineral endowment represent highly specific and fixed elements of the landscape, and their development and extraction thus requires the acceptance and trust of a specific network of landholders and stakeholders, sometimes with diverse and even competing interests in the development and use of the local environment. Weakness or absence of such social licence can bring added costs and complexity to developing and operating a mine – and in extreme cases even result in project failure. Because it must be maintained throughout the life of a project, such social risk becomes a substantial part of overall project risk profile, particularly during routine operations once initial technical and development risks have been largely resolved.CITATION: Bromfield, H, Roberston, S, Shah, B, Shanahan, R, Trench, A and Batt, G, 2018. On equitable distribution of mining wealth and the social licence to operate imperative – an African case study, in Proceedings Complex Orebodies Conference 2018, pp 6–7 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018

Tasker (1985) has detailed the accepted practices for constructing polygonal blocks of influerrce in the estimation of ore reserves. Further he suggests that users might fall into traps such as omission or overlap when determining the area of influence for each sample point.As an example of these types of errors Tasker (1985) describes under "Coverage ofthe Area" how 25 per cent of the exploration area will not be included in an equiangular triangular grid where hexagonal blocks of influence are used.The introductory suggestion of Tasker (1985) of "incorrect procedure" by Rudenno (1983) (and Armstrong (1983)) reiterated around the Figure 2a of Tasker (1985), is based on the brief description of the polygonal area of influence by Rudenno (1983). That description is not too different from that given by King, McMahon and Bujtor (1982).

Jan 1, 1985

Brand se Baai, on the North-Western Cape Coastline of Southern Africa, is the site of the Namakwa Sands Limited Heavy Mineral Sands Mining Operation. The mine is a 14 000 000 tonnes per annum open-cast strip-style mining operation using truck and excavators to mine the sand. The orebody consists of relatively unconsolidated sands reworked by fluvial, marine and aeolian mechanisms. The total orebody includes three primary ore types, namely the Red Aeolian Sands, the Dorbank (a hardcap layer) and the Orange Feldspathic Sands, which forms the bulk of the resource. Mine planning and scheduling is done using a 50 m + 50 m variable bench height system with three primary bench heights. The Aeolian and Dorbank each have a bench that varies with their thickness of up to a maximum eight metres and the Orange feldspathic sand, a standard bench height of eight metres. For a number of years Namakwa Sands used a complex 2D, Datamine Based, macro system to perform the resource evaluation and mine design. As the mining operations expanded it became clear that a complete system redesign was required to cope with the increasing operational pressures coming to bear on the old system. The biggest challenges facing any replacement system are the sheer scale of the mining operation (16 km long and 7 km wide resulting in a minimum of 30 000 individual 50 m + 50 m mining blocks), the size of the geological data files (up to one gigabyte) and representing the benching system accurately (as the benches follow the topography and are not horizontal). In addition to these large-scale issues, the ability to run multiple mine design scenarios was required, an aspect which the old system was unable to perform adequately. Based upon these primary difficulties, a new business process was developed which required a complete rethink of the mine design methodology as well as the creation of a new system to deplete the resource model, produce monthly reconciliation evaluations and create a ultimate pit limit based on the operations mining cost model. Using a system based on the Earthworks Mine2-4D Mine Design and Scheduling System (with its mine design and scheduling specific tools), as well as a small, customised program, massive time-savings have been made. It is now possible to run several life of mine scenarios in the same timeframe that a single life of mine run took under the old system. This paper will present the methodology and business decisions behind the new Namakwa Sands system, and detail the restructuring and development used to deal with the large data set sizes and long processing times inherent in the old system. Future enhancements to the system that will make it more generically applicable to open-cast mining scenarios will also be dealt with.

Jan 1, 2003

By D J. Williams, E M. Rykaart

Soil cover systems are routinely used to reduce acid rock drainage (ARD) from mine waste rock dumps and tailings impoundments. A technology for the analysis, design, instrumentation and monitoring of barrier-type (oxygen limiting) and store/release-type covers has rapidly developed and the number of covers under construction has grown extensively during the past ten years. Mine operators are currently investing in this relatively new technology with the expectation that these covers will reduce long-term liability with respect to acid generation, drainage and treatment costs. While the design life for cover systems may extend into hundreds of years, we have less than ten years of construction experience and proven performance. Some cover designs appear to be functioning well; however, the performance of many more is now being called into question. The paper presented herein evaluates the integrity of cover systems from a global perspective. Soil cover systems for the control of ARD are being constructed at mine sites located within high rainfall equatorial regions, deserts ranging from low to high latitude, humid temperate regions, and cold-continental climates. While the technology for analysis and design of soil cover systems is well developed, factors that compromise the performance of cover systems (in addition to the paramount issue of climate) relate to material selection and availability, cost and construction practices. Long-term performance is also strongly affected by physical stability (ie erosion), volume change (ie cracking, desiccation, settlement and freeze/thaw cycling), vegetation, soil evolution and ecological stability. This paper provides a review of the application of soil cover systems for the control of ARD. Cases where cover systems have been constructed with both success and failure are illustrated. General criteria and recommendations for cover selection, design, construction, and maintenance are provided.

Jan 1, 2003

By R Pax

Feed ore variability, the major disturbance into a mineral processing plant receives considerable attention operationally, and at research and strategic levels to ensure that the throughput of the ore are maintained at all times whilst encouraging a sustainable approach to recovering the valuable minerals from ore bodies. Effort and insights from mineralogical and processing characterisation (eg. ore hardness) of sub-ore body domains allows sequencing for processing to ensure consistency and throughput for the size sensitive grinding unit processes. Optimising the blast fragmentation has benefit as does pre-concentration using ore sorting. Blending of ore domains with different characteristics is common for operations. In recent years a number of technologies have become available for on-line and more rapid assessment of ore and have been discussed in a recent paper by Cameron et.al. (2016) as an opportunity for further utilisation of mine-to-mill (and mine to product) strategies. Current concepts have focused on extensive characterisation of ores so that current control practices can be realised and made more effective with relatively slow measurements and actuation. CITATION: Pax, R, 2018. Dynamic concepts for mine to mill operation and optimisation, in Proceedings Complex Orebodies Conference 2018, pp 124–125 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018

The Olympic Dam copper/uranium deposit was discovered in 1975 and subsequently commenced production in 1988 at a rate of 1.5 Mtpa using the sublevel open stoping mining method. In the following five years the operation underwent two optimisation programs that have increased mine output to the current rate of 3.4 Mtpa of ore. The pre-expansion requirements were annual development of approximately 10 000 m and the placement of around 500 000 m 3 /yr of cemented backfill. Previous to 1995 all hoisting was carried out by the Whenan shaft. The Robinson shaft was commissioned during that year to hoist increasing development mullock tonnages and to allow for increased ore hoist via the Whenan system. In 1996 WMC committed to expanding the output of the operation from 85 000 tpa (1996/97 budget) copper metal to 200 000tpa. The corresponding increase in mine ore production requires expansion to a nominal rate of 9.0 Mtpa. To achieve these increased tonnages, major hardware installations will include sinking of a third shaft capable of hoisting in excess of 9.5 Mtpa, a rail haulage horizon utilising automatic trains, expansion of the existing backfill plant to a production capacity of over 1.8 million m 3 /yr, increasing long term development rates to 15 000 m/yr and increasing the production drilling and blasting rates in line with the higher stope production output. This paper provides detail into the intended hardware (shafts, equipment, etc) of the mining system, and outlines the intended long-term operation and management of the operation. Of particular note is the background requirement of ensuring existing production targets are met to provide critical cashflows into the project during the period of expansion.

Jan 1, 1998

By Weatherstone N

Ironstone outcrops on the Savage River were discovered by State Government Surveyor C. P. Sprent in 1877. The limonitic outcrops were initially regarded as the surface expressions of pos- sible tin-bearing lodes, the tin deposits at Mt. Bischoff to the east having been found in 1871. By the turn of the century, shafts, adits and trenches had been developed on the deposits in the unsuccessful quest for gold, silver and base metals. The deposits were first reported as a possible source of iron ore by Tasmania Department of Mines (TDM) geologists in 1919. Modern exploration began in 1956 with the area registering the best magnetic response during an aeromagnetic survey of North- western Tasmania by the Bureau of Mineral Resources (BMR). The area had previously been reserved by TDM in anticipation of the lifting of the iron ore embargo. Reconnaissance Geologi- cal mapping (TDM) and grid-based magnetometer surveys (BMR) indicated that resources of iron-bearing material could approach hundreds of millions of tons. This provided the incentive for two diamond drill holes funded by TDM which demonstrated iron mineralisation in substantial tonnages, suitable grades and, fol

Jan 1, 1988

Digitally enhanced Landsat MSS images are being used'in delineating the structural and tectonic features in the Arabian Shield as part of a project funded by King Abdulaziz City for Science and Technology (KACST) of Saudi Arabia. The delineated features together with statistical modeling and structural analysis are to assist in an improved understanding of the evolution and mineralization of the Arabian Shield. Specific emphasis is to be given to relating the extracted linear features to the location of known mineral occurrences and areas of potential geologic hazards. A number of textural (spatial) and spectral enhancement techniques have been applied, to the MSS data covering selected areas, in selecting the most optimal enhancement product for visually delineating the structural geologic .features.

Jan 1, 1988

By F Shi, W Valery

WMC Fertilizers (WMCF) operates a Beneficiation Plant at Phosphate Hill in Queensland, Australia. The plant consists of two-stages of crushing, followed by a wet scrubbing and desliming circuit. Ore is then fed to a tertiary crusher and a milling/classification circuit, with the final product going to a swing-lift thickener and storage tanks. From the slurry storage tanks, the phosphate rock is fed to a horizontal belt filters which dewater the phosphate rock ahead of a hemi-hydrate phosphoric acid plant. The water soluble recovery of phosphate in the phosphoric acid plant is dependent on the moisture content in the filter cake, as this governs the wash water addition to the hemi-hydrate filters. Since commencement of operation, the moisture content of the filter cake fed to the phosphoric acid plant has been higher than design reducing the phosphoric acid plant recovery and at times reducing the throughput through the phosphoric acid plant. The moisture content in the filter cake can be decreased by reducing the generation of fines in the milling and classification circuit (classified as û38 micron material). This paper identifies the causes of fines generation in the Phosphate Hill beneficiation plant. It discusses the options that were assessed to minimise fines generation, and outlines the results to date of circuit changes undertaken at Phosphate Hill.

Jan 1, 2003

By Travis G. A

The Mons Cupri Cu-Pb-Zn deposit is situated within a complex Archaean volcanic sequence dominated by rhyolitic fragmental rocks. This sequence is underlain by a 150 m group of rhyolitic ash flow tuffs with minor conglomerate and sandstone, and is overlain by approximately 180 m of dacite and andesite flows. The Mons Cupri sequence itself consists of several elongate or circular rhyolite domes, flanked by rhyolitic fragmental rocks ranging up to 400 m in thickness away from the domes. The lithology of clasts (typically 1 cm to 1.5 m in size) provides strong evidence that the flanking fragmental unit was derived by explosive eruptions associated with the rhyolite domes. An approximately 15 m thick shale horizon, bleached and silicified at its base, overlies both the rhyolite domes and the frag- mental unit. A south-dipping chloritic pipe-like body, with near-surface dimensions of 1,000 by 250 m and narrowing at depth to produce a cone-shaped cross section, lies within the rhyolitic frag- mental rocks about 250 m north of the largest rhyolite'dome in the Mons Cupri vicinity. Its east west elongation parallels that of the rhyolite dome. Ghost fragments indicate that the material in this fissure is a severely chloritized derivative of the rhyolitic frag- mental unit. The chloritic pipe is overlain

Jan 1, 1973

Tertiary laterite is extensively developed in the Darling Range area, but there is doubt that. the high relief areas along the Darling Scarp have been peneplaned. Three distinct laterite types are distinguished which are termed transported, introduced and residual laterites to avoid ambiguity. Transported laterites are recemented (primarily) lateritic materials and introduced laterites are chemical deposits of bog iron type. Residual laterites are formed in situ from the host material which can be either decomposed bedrock or detrital' clay, and the two varieties formed from these sources are termed residual A and residual B laterites. The genetic requirements for residual laterite formation are (1) host material, (2) porosity and permeability, (3) rainfall, (4) temperature, (5) relief and drainage, (6) vege- tation and (7) time. Previous theories on the origin of laterite appear to be inadequate and incompatible with field observations. Field investigations and drilling in the Darling Range area disclosed lenticular section and thickest laterite development on slopes and thin or oc- casionally no laterite development on crests. Lateritisation is considered a surface feature and the end product of the chemical weathering cycle. The relief and rain-water acting on a kaolinitic host material are con- sidered the major contributing factors in the

Jan 1, 1973

The QSL continuous lead smelting process incorporates two basically different pyro- metallurgical operations: roast-reaction, smelting of galena concentrate, followed by carbothermic reduction of some by-product lead oxide slag. Both steps are performed simultaneously in separate compartments of a long horizontal cylindrical reactor, which is lined with fused chrome-magnesite bricks. This converter-type reactor contains a liquid bath consisting of a shallow layer of lead bullion with a superimposed layer of slag, the lead content of which is gradually reduced from about 50% to discardible levels. Metal relative to slag, and slag relative to the gas phase are in continuous countercurrent motion. A lead well is arranged at the flue end of the reactor, and an overflow tap for slag at the opposite end. Tonnage oxygen for oxidation and smelting of the charge, and sub- stoichiometric mixtures of pulverized coal and oxygen for reduction and heating of the slag, are directly blown into the bath using fluid-protected Savard-Lee injectors. High gas-liquid momentum transfer ensures intensive local mixing, and enables rapid mass and heat transfer.

Jan 1, 1984

By B Muller

The disposal of tailings has become a greater issue due to the changing economics of mineral extraction. The processing of lower grade deposits and tougher environmental regulations has led to the selection process for tailings disposal options becoming more complicated. Current selection methods are too time consuming or too easily manipulated for inexperienced users to use successfully. This paper examines a different method for the initial selection of a tailings disposal system, using a numerical system that was introduced into the mining method selection field. The numerical method system was compiled by reviewing previous work done on the positive and negative aspects of the tailings disposal methods used by the coal and metallurgical mining industries. Assessing the common trends and design aspects of each of the methods was undertaken and a ranking was designed to explain how well suited the tailings disposal methods are to a situation. The tool successfully analysed sites using common data to indicate methods for disposal that would have provided minimal technical problems, low environmental risk and been practically possible. The tool also proved successful in identifying critical areas that would need further investigation. This new system was tested on some disaster case studies from around the world in order to show that it could have successfully selected a different method for tailings disposal that would have prevented the disaster.

Sep 26, 2011

By Percival T. J

Precious metal deposits containing fine-grained disseminated or dispersed gold-silver form by reactions between metal-bearing hydrothermal fluids and sedimentary host rocks. Host rock permeability and chemical composition exert a primary influence on the fluid movement necessary for any primary dispersion halos to form around ore bodies, and determine whether a disseminated style orebody itself will even form. Where the mineralized rocks are relatively unfractured the margins of the orebodies exhibit abrupt decreases in gold content, and strong corresponding decreases in the contents of other metals. Primary dispersion patterns of this type are difficult to recognize and commonly are only a few meters (or less) in width. If the host rocks were well shattered and broken before mineralization, remnant amounts of the hydrothermal fluids may be dissipated in fractures lateral to or above the ore zones. This causes geochemical anomalies to form with clays and silt- size silica within the fractures over distances of up to at least 30-50 meters away from the main orebodies. The tendency for large primary dispersion halos not to form around these deposits explains the disappointing results from so many drilling programs where no multi-gram gold-grade material was even found: if large areas and apparent large volumes of rock contain one gram or less gold that likely is the "orebody" and not a halo effect around high grade mineralization.

Jan 1, 1988

By Caress ME, Knight MD

The eruption 75,000 y ago of the Younger Toba Tuff ignimbrite and associated ash-fall ranks as the greatest known volcanic event of the past million years. The ignimbrite is known to cover 21,000 km2 (it may extend farther under alluvium of the coastal plain) and forms a low-angle depositional shield centered on the Lake Toba caldera. The ignimbrite is largely nonwelded and shows systematic outward decreases in lithic and crystal contents and clast sizes. Coarse lithic lag breccias that occur at places 85 km apart point to two or more vents on a fissure having this length. The extensive associated ash-fall traceable as far as Malaysia and India is evidently a co-ignimbrite ash but could have a phreatomag-matic component. The Lake Toba caldera measures95 x 35 km and is >1 kn deep. It is not a single caldera but comprises several contiguous depressions and grabens diversified by two re-surgent domes. Earlier parts of the caldera are related to at least two earlier ignimbrites including the Older Toba Tuff dated by Chesner at 0.5 Ma. The caldera is elongated parallel to, and is only 25 km from, the great Sumatra strike-slip fault, and it can be considered almost certain that the caldera is in part a kind of pull-apart structure. The ignimbrites are very crystal-rich. There is no reason to think basaltic magma was involved in their emplacement. Together with ash-falls they have a total volume of several thousand cubic kilometers. They can be categorized as originating from partially crystallized granitic magmas that erupted because tectonic pull-apart occurred several times across the pluton. Small rhyolitic lava domes on fault-lines comprise the youngest volcanic products.

Jan 1, 1990

By D Cliff

Despite advances in technology, the Australian underground coal mining industry continues to have incidents involving spontaneous combustion. Managing these incidents relies heavily on understanding the state of the event as it develops and controls are applied. Gas analysis plays a major role in aiding this interpretation. This interpretation is based upon laboratory testing of coal samples under conditions very different to those found in an underground coal mine. This paper will explore the links between laboratory data and actual events. It will highlight the difficulties of using laboratory data. It will outline the relevance of laboratory data and how it can be usefully applied to manage spontaneous combustion events as well as in establishing relevant trigger action response plans to facilitate early detection and control of such events. The paper will identify some of the factors that need to be applied when adapting laboratory data to apply to the underground coal mine. It will outline the need for comprehensive knowledge of the local conditions and gas concentrations to be expected at all locations in an underground coal mine. Further, it will demonstrate the need for mine environment monitoring to go beyond simply measuring gas concentrations but also to include reliable atmospheric pressure, temperature and air velocity monitoring.CITATION:Cliff, D, 2013. Gas analysis - Linking the laboratory with reality, in Proceedings The Australian Mine Ventilation Conference , pp 277-282 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013