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By G J. Butler

Since 2005 BHP Billiton Iron Ore has been working with the consulting company Runge to implement the software XPAC Advanced Destination Scheduler (ADS) in life-of-mine studies. The Mt Whaleback mine, being the single biggest open cut iron ore mine in the world, presents some special challenges in scheduling. The biggest challenge has been the slow run time and difficulty in targeting all ore and waste requirements. The high waste to ore strip ratio means that much of the cost of mining is tied up in waste dumping. It is therefore very important to schedule waste efficiently. At Mt Whaleback there is an additional challenge of managing two types of waste: inert and environmentally sensitive pyritic material. This paper shows how the XPAC destination scheduler is configured to achieve efficient and repeatable waste scheduling results at the Mt Whaleback mine. It also discusses proposed improvements to the process. The overall ore and waste schedule is configured to target mining parameters, crusher feed and grades. The schedule is run in two stages. The first stage is ore mining. The waste that needs to be mined is locked in a sequence. The second stage is the waste scheduling. The waste destination scheduler uses three inputs: source waste parcels locked in a sequence, a haul road and a waste dump block model. The process involves setting some rules for dump progression and haul path selection. These can be validated. The schedule is then run to target the shortest travel time. The results of scheduling are graphical and numeric and can be transferred into many different formats for reporting. The graphical result shows face positions in time that can be used for detailed dump design. The numeric results are used to estimate truck fleet size, fuel consumption, labour and equipment financials. The process has been successful in reducing the project time line, increasing the ability to do more scenarios and increasing schedule accuracy, transferability and repeatability.

Jan 1, 2007

The link between exploration success and new ideas from research and development (R&D), although often assumed, is rarely well explained or well illustrated. One issue is understanding the role of innovative ideas in the exploration process, and another is determining the relative contributing factors to an exploration success. A useful measure of gold exploration success over an extended time period and geographic area is discovery rate (θ): Discovery rate θ = gold discovered/exploration expenditure (usefully expressed as tonnes of gold per $1 million). The total gold discovered is strongly leveraged to discovery rate, as any increase in θ typically has a positive effect on exploration expenditure. Hence, improving the discovery rate is a key driver for a sustainable mineral industry, especially for gold. There are very few ways to influence θ beyond innovative R&D ideas and breakthroughs, and well-motivated and trained staff. The Australian gold industry has experienced exceptional exploration success with 9956 tonnes of gold discovered from 1979 to 2003, inclusive. This equates to: Discovery rate θ = 0.9 t/A$1 M Two-thirds of this gold has been found in the Yilgarn craton of Western Australia (6265 tonnes Au). Without this exploration success, there would have been no platform on which to build the current gold industry and to take advantage of elevated gold prices and advances in mining and processing. Several causes of the Western Australian gold success have been proposed, and more than one may be correct. Gold price, and mining and processing advances have all had influence; but exploration success and discovery of gold were essential to add to the miniscule resource base of 1979. A comparison of the exploration methods used to search for gold in the Yilgarn craton in 1979, and 1990, highlights a small number of breakthrough ideas introduced to the industry during the 1980s, and in most cases emanating directly from local R&D. By the 1990s, these ideas were being used in most gold exploration programs in the Yilgarn craton, and are still used today. These six breakthroughs are: epigenetic timing of gold deposit formation (ie after formation of rock sequence), chemical and mechanical role of host rocks, relevance of carbonate alteration haloes, gold dispersion in the regolith, optimal sampling media, and appreciation of landforms in exploration program design. Interestingly, gold exploration science in Western Australia entered the 1980s period without rigid paradigms in place regarding gold deposit formation; there was a balance between descriptive and genetic models, there was scope to research gold deposits from microscopic to global scales and there were highly effective industry research uptake processes in place: it is suggested here that these favourable conditions contributed to the exploration success and industry growth.

Jan 1, 2007

By M Logsdon, T R. Clark, T I. Mudder

Oxidation of pyrite and other metal sulfides is accelerated by the activity of acidophilic iron- and sulfur-oxidising microorganisms. The ideal biocide for controlling biocatalysed acid rock drainage (ARD) would be inexpensive, inhibitory toward sulfide-oxidising bacteria at low concentrations, relatively non-toxic toward non-target organisms, degradable in the environment outside the zone of application, stable and long-lived in the impacted environment, sufficiently mobile to penetrate into piles of waste rock or tailings, and effective regardless of the metal sulfides present. A potent agent for combating ARD that fits these criteria is the thiocyanate anion (SCN-). The objective of this study was to determine the effectiveness of thiocyanate in controlling or preventing ARD within several types of sulfidic waste rock and tailings. Results were evaluated within a context encompassing process water and solids management strategies employed at different precious and base metal mines. Sulfidic rock was treated with solutions of potassium thiocyanate and subjected to accelerated weathering tests in lab-scale humidity cells, trays, and columns. Certain tests also utilised solid phase copper thiocyanate (CuSCN) to evaluate a ærelease on demandÆ strategy. Leach solutions were analysed for parameters indicative of ARD including pH, redox potential, conductivity, acidity, sulfate, iron, and certain heavy metals. Two larger scale field tests were also initiated by personnel at separate mines using, in the first case, ore containing a pyritic component in large columns and, in the second case, waste rock containing base metal sulfides in lined pits. Humidity cell tests showed low concentrations of thiocyanate solution applied to pyritic waste rock could reduce ARD formation by 90 per cent or more compared to untreated controls, with the residual release of ARD components (eg sulfate, Fe, acidity) probably reflective of strictly chemical oxidation of the sulfides by molecular oxygen. Thiocyanate applied at a concentration of 22 mg/kg to waste rock from a base metal mining operation in large laboratory scale columns reduced acidity and sulfate release by 75 per cent to 80 per cent compared to untreated controls. This extent of reduction in ARD may not meet discharge water quality criteria, but has the potential to significantly reduce lime consumption and sludge production in a conventional neutralisation facility. The use of a solid-phase metal thiocyanate compound (CuSCN) at a dosage of 194 mg/kg pyritic ore resulted in the following per cent reductions in leach solution parameters in comparison to controls: sulfate (78 per cent), acidity (93 per cent), dissolved Fe (99.4 per cent), and dissolved arsenic (99.9 per cent). This compound has the further advantage of being relatively stable until needed, dissolving only in the presence of an acidic and oxidising ARD solution. If this line of investigation progresses favorably, it offers the possibility of an ARD prevention technique that may be applicable in some settings. Mine-site pilot scale tests with pyritic ore resulted in reductions in sulfate (56 per cent) and Fe, As (66 - 68 per cent). In sum, thiocyanate at relatively low concentrations dramatically reduced ARD from sulfidic rock in a simulated heap. The pilot test using waste rock containing base metal sulfides has begun only recently. A preliminary process design and cost estimate was made for fullscale application of thiocyanate to waste rock at a hypothetical gold mining operation producing 5000 tpd of ore and 10 000 tpd of sulfide containing waste rock. The combined capital and operating costs for the full-scale thiocyanate application facility were about $US 0.25/tonne of waste rock and $US 0.50/tonne of ore. Although thiocyanate is readily available commercially, it also occurs in process solutions associated with certain types of mining processes. Such solutions after selective treatment might be recycled to sulfidic waste rock or tailings, rather than being treated for thiocyanate degradation.

Jan 1, 2003

By Millett R

In January 1989, Western Mining Corporation (WMC) acquired the Perseverance and Rocky's Reward orebodies and associated infrastructure from the Agnew Mining Company (AMC). AMC had closed the operation in August 1986 and it had been on care and maintenance since that time. To return the operation to full production as quickly as possible it was planned to mine the Rocky's Reward orebody by opencut mining and treat this ore whilst the underground Perseverance workings were rehabilitated. Removal of overburden from Rocky's Reward thus commenced immediately with a target of April 1989 for delivery of the first ore to the concentrator. Extensive rehabilitation and upgrading of the concentrator commenced in late January, and continued in stages for an 18 month period. The first stage consisted of overhauling all existing equipment and completing a crusher upgrade commenced by AMC prior to the mine closure. During this period, it was necessary to recruit a workforce and re-establish all the required support structure. In late April the plant commenced commissioning trials on mullock. The first ore was crushed on the 11th May 1989 and concentrate produced the following day. From May 1989 to June 1990 the plant was producing nickel concentrate whilst further upgrading continued. During this period, the flotation capacity was tripled by installation of a bank of eight Wemco 144L flotation cells in rougher-scavenger duties and two 12 metre by 2 metre diameter columns in cleaning duties. The existing vacuum drum filters were replaced by two Larox PF32 pressure filters, eliminating the need for solar drying and allowing the flash drier to be fed direct from the filters. These, along with a major borefield expansion and further work to the crusher increased the nominal plant capacity from 600,000 to 1,000,000 t/yr. The plant achieved the 1,000,000 t/yr rate in April 1990.

Jan 1, 1991

By Olofsson O, Durance M-V, Villeneuve J

Modelling of the Sala Agitated Mill (SAM) using a simplified solution to the population balance approach has been carried out. The rate of breakage is modified by the mill power, providing predictions of the resulting product size distribution for different grinding conditions and mill sizes. The power of the mill is estimated by a physically derived function, using an assumption that the mill power consumption mainly arises from the ball to ball friction within the mill. The grinding efficiency, depending on the media ball size and pulp density is approximated in the model. Experimental work on grinding a complex sulphide ore using a pilot scale SAM 7.5 has been carried out. Good agreement has been found between the data from the test campaign and the simulated values. The model has also been validated with experimental data from a pilot grinding of calcium carbonate and from an industrial mill. The model is included in the USIM PAC simulation software.

Jan 1, 1995

By Lawrie DC

Finely ground beryl was decomposed by mixing with ground limestone and soda ash andheating at 900¦C for,4-6, hours.The sintered material was then crushed and leached by boiling under reflux with 5N sulphuric acid for up to 4 hours. The liquor containing Be, ; Al, . Fe and, other elements as sulphates, was separated from the residue which consisted mainly of silica and gypsum, either' by filtration or by decantation washing.The clarified leach liquor.was first treated to remove Fed}: , by contacting with a primary amine-kerosene solution in2 counter-current stages. The iron free liquor wa,s then passed to a 4-stage counter-current extraction unit where it was con-tacted with 0.5M di-(2 ethyl hexyl), phosphoric acid (EHPA) in kerosene. The'operating conditions were such that at least 99 p'br cent of the Be and only 7' per' cent of the Al- was, extracted The ,B6`= and. Al were ;stripped from the organic phase by sod-ium hydroxide solution of suitable concentration to form soluble "beryllate" and "aluminate" respectively. Most of the Be was subsequently precipTitated a's Bo(OH)2 by-hoating the alkaline solution to 90C, the Al remaining in solution. After filtra-tion and washing, the Bo(OH)2 could be ignited directly to BeO at about 1000 C or used as a feed to further purification stages aimed at producing "nuclear grade" BeO.A small coaAter-current extraction unit was operated con-tinuously using liquor derived from the sintering and leaching stages. Data obtained in batch laboratory tests were confirmed. A product obtained by ignition of the washed "beryllate" precip-itate contained more than'99 per cent BeO.

Jan 1, 1963

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 Mudge S. T

Tin mineralisation occurs at Renison Bell, near Zeehan in Tasmania, as a group of cassiterite-bearing massive pyrrhotite bodies in Cambrian sedimentary rocks. Mineralising fluids from the underlying Pine Hill Granite have replaced carbonate rocks with massive sulphides and cassiterite. The discrete blocks of massive pyrrhotite have high density and high magnetic susceptibility contrasts with the comparatively barren country rock. They are strongly remanently magnetised and are excellent geophysical targets. A detailed geophysical model of the mine was created from a database comprising 1:2500 scale geological mapping and geological logs of approximately 4000 drill-cores. The model comprises 93 parallel cross-sections, each 1.5 km long, spaced 20 m apart. Each cross-section is comprised of up to 150 polygons which represent all known individual rock types and orebodies. The magnetic response of the model was computed and the residual between the model and the observed response, measured by a low-level helicopter magnetic survey, revealed several anomalies unaccounted for by the model. Drill testing of these anomalies led to the discovery of several hitherto unknown massive sulphide occurrences. Surface gravity data of the mine and surrounding district were enhanced by applying digital terrain corrections. The corrected Bouguer anomaly revealed a hitherto unknown positive anomaly associated with the ore zones of the Renison mine. The gravity anomaly of the model was also computed. Unfortunately, the limited resolution of the observed data prevented the delineation of ore targets. The terrain-corrected Bouguer gravity data were also used to model the shape of the underlying non-magnetic Pine Hill Granite. A computer inversion method was implemented to build a 3D block model of the granite. Results agreed with previous conventional 3D modelling techniques but yielded improved resolution of the granite form. The project involved development of new computer software to build and edit the large and complex 3D geophysical model of the mine and to compute the model's magnetic and gravity responses. Software was also developed to build and compute the gravity response of a 3D block model of the underlying granite.

Jan 1, 1997

The Porphyry Gold Mine (150km NE of Kalgoorlie) was reopened in November 1983 with production commencing in April 1984. Original reserves quoted for the open pit were 822000 tonnes at 5.2 g/t Au based on wide spaced diamond drill holes. Close spaced grade control drilling has revealed some loss of tonnage and a significantly lower grade. In the area of the pit mined to 30 June 1985 this grade control reserve estimate contained 67% of the gold in the original ore reserve with the grade being 19% lower. It is anticipated that the final grade control reserve estimate will contain around 80-85% of the gold contained in the original reserve. There has been close agreement between the Mill Head grade (4.19 g/t Au) and the grade control predicted grade (4.2 g/t Au) confirming that dilution is not a major problem. After initial poor agreement (attributed to "lock-up" in the circuit) the bullion calculated head grade steadied out at 93% of the grade control prediction. This difference is considered to represent the geostatistical bias in the drill hole/crushed rock sampling. Gold recovery has been in the range 93-95% (compared with bullion calculated grade). The net result is that the recovered grade of ore mined to 30 June 1985 is 70% of the original ore reserve grade.

Jan 1, 1986

The first stage in process design for coal preparation is a study of the coal seam and its associated features followed by an examination of the mining method to.be employed and the coal market requirements. Normally benefication is based upon the differences in relative densities of the coal constituents and asso- ciated ash forming components. Since initial liberation is obtained at a relatively large particle size, the size to which coal is crushed is more subject to the market requirements than to liberation studies. It is possible to estimate energy requirements for comminution and recoveries and product grades can be cal- culated following the interpretation of standard laboratory tests. Numerous units are available for the separation and concentration of the components of run of mine coal in- cluding jigging, dense media separators, and for the finer particle size fractions, flowing film separators, dense media cyclones and froth flotation.

Jan 1, 1975

By J E. Everett

When ore is being selected from a regularised block model, a common procedure is to accept as ore any block whose iron content exceeds a cut-off value, and whose content in each contaminant analyte is below an appropriate cut-off value. This method defines a quadrant within the multidimensional analyte space, such that any block lying within the quadrant is accepted as ore and any block lying outside the quadrant is rejected as waste. The cut-off levels can be adjusted to maximise ore tonnage at a target grade (or analyte vector). When the method is applied to a project comprising multiple pits with systematically differing grades, it is found that ore tonnage at target grade is maximised if different cut-off quadrants are used for the different pits. This is apparently illogical, because it implies that an ore block from one pit may be accepted as ore, but an identical block from another pit is rejected as waste, even when ore from both pits is to be blended into the same product and there are no production cost differences between the pits. The paradox is an example of the ætheory of the second bestÆ: if there are multiple conditions for optimality and one condition is broken, then obeying the other conditions may not be the best policy. The problem lies with the quadrant cut-off method, which is shown to be inconsistent even for selecting ore from a single pit. A criterion that is inconsistent cannot be optimal. It is shown that a linear composite cut-off function provides maximum ore tonnage, and that a single composite cut-off criterion is consistent and optimal over multiple pits. Further, it is shown that a composite cut-off function not only maximises ore tonnage at a specified composition, but also can be used to select ore so as to maximise project value, if marginal costs and values are available.CITATION:Everett, J E, 2013. Selection of ore from multiple pits, in Proceedings Iron Ore 2013 , pp 185-192 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

This opportunity to present a few thoughts on the question of professional education and training comes at a most appropriate time. In Britain, a great deal of energy is being expended in this, so-called, post-Finniston era, to create a new look for the status of engineers and engineering by the Engineering Council. Naturally, the tune and the tempo are being called by the big institutions which essentially serve the manufacturing industries. However, the Engineering Council in Great Britain also includes the mining and processing institutions such as the Institution of Mining and Metallurgy (IMM) and so there is a chance - if it is taken - for us to contribute to the debate and the determinations that emanate from it. The IMM, of course, caters for the professional status of those engaged in the mining and minerals industries and includes many who do not see themselves as engineers. Nevertheless, the debate is important to these other groups whose livelihood is inextricably bound up with those who are engineers and there are many facets of the engineering policies being shaped that will, I am sure, eventually be enshrined in the requirements for professional status across all the various disciplines in our industry.

Jan 1, 1986

By M Ridd

The Pasminco Century Mine (PCML) is located 250 kilometres north-north-west of Mt Isa in Queensland. As one of the worldÆs largest zinc mines, PCML comprises mining and milling operations at Lawn Hill, concentrate filtering and shipping facilities at Karumba in the Gulf of Carpentaria, and a 304 km slurry pipeline connecting the two operations.Mineralisation at Century occurs in the upper part of the Lawn Hill formation and is comprised of siltstones, shale and sandstone. Pyrite occurs immediately above the mineralised sequence, and throughout the footwall carbonaceous shale forming the lower proportion of the deposit. Zinc predominantly occurs as sphalerite (ZnS) and lead as galena (PbS). These sulfides are very fine-grained and may oxidise rapidly upon exposure. The pyrite, sphalerite and galena are the main potential sources of acid rock drainage (ARD) or heavy metal leachate at the mine site.To minimise the generation of ARD in the waste rock dumps, they are constructed in a specific manner and waste rock is classified into the following three classes based on competence and acid forming/consuming capabilities:Class 1: Competent rock, non-acid forming or acid-consuming material. Class 2: Non-competent, non-acid forming or acid-consuming material. Class 3: Acid-forming material. The generalised Waste Rock Dump Design for current operations comprises an outer zone of Class 1 material and inner zone with Class 2 and 3 material. Limestone is the Class 1 material used for both its structural support and acid neutralising capacity. Reserves of limestone at the mine site are limited. Net acid generation (NAG) tests confirm that a mix of limestone and shales can be net acid consuming.Field based and laboratory based trials have been established at PCML to quantify blending levels of limestone and shale that will produce leachates both neutral in pH and low in metal/salt concentrations. This paper will discuss the first six months results (of a two-year sampling program) from the trials.

Jan 1, 2004

By N LaBranche

The resurgence of Silicosis and Coal Worker’s Pneumoconiosis (CWP) in coalmines has placed the spotlight back on the management of worker exposure to particulate matter across all mining sectors (Coal Workers Pneumoconiosis Select Committee, 2017). There already exists a substantial body of knowledge on particulate matter and its impacts upon human health (National Academies of Sciences, Engineering and Medicine, 2018). The contribution of silica to respiratory disease is well known. However, it is also clear that there are major gaps in our understanding. In part, those gaps relate to the contribution to adverse health effects of mineralogical constituents in the dust (LaBranche et al, 2021a). It is also becoming evident that the size and shape of the particulate matter can affect its potential impact on human health. The University of Queensland has developed a methodology for characterising respirable and inhalable dust samples using scanning electron microscopy. Multiple samples from numerous mining operations have already been analysed using the UQ methodology. The process has been successful in showing the variety of mineralogical components and particle size distributions present in various areas of the mines.

May 5, 2022

By N Madani, N Battalgazy

Geostatistical simulation has been widely used in mineral resource classification of metalliferous deposits based on international standards such as JORC code which is one of the most extensively used around the world for reporting the mineral resources and ore reserves alongside with objectives of the companies (Sadeghi, Madani and Carranza, 2015). However, the modeling process becomes challenging whenever the relationship between the variables of interest is complicated and cumbersome to model. This characteristic sometimes manifests itself in complex spatial and global cross-correlation measures, deeply depending on the formation and genesis of the deposit. The objective then is to generate the models that not only reproduces these significant characteristics, but also, it leads to a better understanding of a mineral resource estimation and successive decision-makings. In this respect, there are different approaches taking into account the spatial correlation structures, for which can be classified as independent simulation, co-simulation, and factor-based methodologies. In this paper, three techniques: Turning Bands Independent Simulation (TBSIM), Turning Bands Co-Simulation (TBCOSIM) and Projection Pursuit Multivariate Transform (PPMT); from those aforementioned approaches are selected to model two cross-correlated variables (Iron and Aluminum Oxide) belonging to an Iron ore deposit located in Brazil, in order to examine which methodology better reproduces the corresponding complexity in terms of local correlation measure. CITATION: Battalgazy, N and Madani, N, 2018. Mineral resource assessment in complex ore deposits based on international standards, in Proceedings Complex Orebodies Conference 2018, pp 79–81 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018

By E Kunz

Decreasing the gas content in the worked seam by degassing before the start of winning has proven its worth in some areas of German hard coal mines. On the one hand, it removes the risk of gas/coal outbursts and on the other hand it reduces the gas inflow into the longwall. The underground experiences of the last 30 years have led to a æhit listÆ of coal seams with respect to their suitability to predegassing. Trials in the DMT laboratory and also results gained in the deposit show that the suitability of a seam for predegassing depends on its rank. In seams with volatile matter of 20 - 30 per cent, high efficiency ratings are achieved as a rule, with lower ratings both in seams with lower and with higher content of volatile matter.If the rock pressure is unusually high, the efficiency rating of the predegassing is lower, even in seams which would normally be suitable for this process. Under unfavourable conditions, for example below or above residual pillars, in-seam boreholes cannot be drilled successfully and the method is not practical. But if the worked seam was already stress relieved, the predegassing has an optimal efficiency. Latest investigations show that the distance between the in-seam boreholes can be extended. But in that case, the time necessary to reach the same reduction of the gas content increases considerably.

Jan 1, 2005

By K Nguyen

The advances in computerised processors and digital microscopic imaging have enabled more studies that attempt to quantify textural information from rock images. However, a texture classification system designed for geometallurgy has always been challenging, as it combines general image analysis techniques and specific texture classifications in terms of the size, shape and orientation of mineral particles. In this paper, a novel texture analysis technique using grey level radial shell association (GLRSA) and spectral angle mapper (SAM) to discriminate between different geological images is outlined. The technique can be applied in routine analysis and characterisation of mineralogical and textural attributes which may correlate with variability in mineral processing behaviours. Using this texture analysis technique, a set of small texture reference images, a texturelet reference library, has been identified from a suite of drill core images. Rock types are classified according to the proportion of each texturelet type recognised. The library is site specific.CITATION:Nguyen, K, 2013. A new texture analysis technique for geometallurgy, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 187-190 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By S L. Simpson, J Waworuntu, B M. Angel

"The Batu Hijau copper/gold mine, operated by PT Newmont Nusa Tenggara, is located in the south-western part of Sumbawa Island, Indonesia, and is situated ~10 km from the south coast at an altitude of ~450 m above sea level. The mine became fully operational by year 2000 and processes approximately 120 000 t of ore per day that contain an average of 0.53 per cent copper and 0.4 per tonne (g/t) gold. Tailings management at the mine includes a deep-sea tailings placement (DSTP) system that discharges tailings through a pipeline via an outfall located at a depth of 125 m at the head of the submarine Senunu Canyon (Figure 1).The environmental impact assessment undertaken prior to approval of the DSTP evaluated both on-land tailings storage facility (TSF) and DSTP options for tailings management. Negative attributes of the on-land TSF included removal of approximately 2300 ha of forest/agricultural land, the potential relocation of >2000 people from their communities, the high average annual rainfall (>2500 mm) and potential for earthquakes presenting challenges for water management and creating a long-term risk of failure. The final choice and Indonesian Government’s approval of DSTP was also influenced by the close proximity of the deep submarine canyon to the mine. The Batu Hijau mine is one of the larger examples of DSTP in the world and deposits tailings at the greatest depth. The depth of the main canyon increases to >2000 m within 20 km of the coast, and to >3000 m approximately 50 km from the coast.The DTSP at Batu Hijau must comply with an environment management plan (EMP) and environment monitoring plan of permit stipulation that approved its commencement in 2000; updated in 2010 and 2015. The environmental management objectives were to avoid impacts to highly productive components of the ecosystem such as coral reefs, mangroves, surface waters and fisheries, and confinement of impacts to areas of low biological productivity. In the vicinity of the DSTP outfall and within the Senunu Canyon, the management plan predicted significant adverse impacts from the DSTP on the marine ecosystem in the form of: reduction in sea water quality due to elevated turbidity and dissolved copper concentrations burial of benthic organisms a reduced habitat for demersal fish.Recovery of the pelagic and benthic ecosystem to preoperating conditions was predicted to occur within the first two years after operations cease. The CSIRO has performed due diligence studies to independently verify compliance with permits and evaluate new monitoring objectives (Apte et al, 2004; Simpson et al, 2010; Angel et al, 2015), and has performed assessments of environmental monitoring data on the tailings footprint and ore stockpile processing against the above environmental objectives (Simpson and Angel, 2015).CITATION:Angel, B M, Simpson, S L and Waworuntu, J, 2016. Long-term monitoring and targeted assessments to improve management of mine wastes – the Batu Hijau project, in Proceedings Life-of-Mine 2016 Conference, pp 81–84 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 28, 2016

By Y Tang, G Si

The analysis of rock fragment sizes plays a pivotal role in various stages of geotechnical and mining engineering. Traditional case-based methods often falter in addressing the variability in rock fragment images due to fluctuating geological conditions and imaging environments. To counter this challenge, AI-based models necessitate training on expansive and diverse data sets, encompassing a wide range of conditions and fragment types. However, the creation of such data sets is often hampered by financial and environmental constraints. This paper introduces a novel approach for rock fragment recognition, which leverages the power of AI for data augmentation, automated annotation and prediction. The methodology integrates three key components: Style-Based Generator Architecture for Generative Adversarial Networks (StyleGAN) for data augmentation, the Segment Anything Model (SAM) for automated annotation, and the multiscale version of you only look once (YOLO-MS) real-time instance segmentation model. The proposed method begins by training a new StyleGAN model using the original data set. It then delves into exploring the latent vectors of different rock fragment images to generate specific ‘style representations’. These latent vectors form the foundational elements of each image in a high-dimensional space, encapsulating critical features such as size, shape, texture, and colour variations. The next step involves conducting quality checks on the generated images to ensure their realism and suitability as training data. Subsequently, the augmented image set is annotated using modified SAM module, followed by training the YOLO-MS real-time instance segmentation model. The performance of this model, trained on the augmented data set, is compared with a model trained on the original data set, particularly in terms of generalisation abilities in new scenarios. The results exhibit a significant enhancement in accuracy, robustness, and generalisability of the model trained on the augmented data set, underscoring the potential of StyleGAN as a powerful tool for data augmentation in geotechnical engineering image analysis.

Sep 1, 2024

By G Johnson, G W. Wilson

The settlement and drainage behaviour of co-mix, or mixtures of waste rock and tailings, are examined in a column study. Three co-mix profiles were observed for settlement, drainage, and pore water pressure response. A fourth profile with waste rock and no tailings was also observed for settlement. Co-mix consolidated faster than predictions for tailings alone, and maintained a saturated void space relative to waste rock alone. Practical benefits of co-mix as a waste disposal technology may include faster time to consolidation than tailings, a reduction in total settlement of placed material compared to tailings alone, and a reduction in acid drainage by limiting the flow of water and oxygen through maintenance of a saturated void space.

Jan 1, 2003