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By M I. Pownceby, S Hapugoda, H Han, E Donskoi, L Lu

Iron ore sinter mineralogy and its associated macro-/micro – structure has a strong impact on sinter quality. At a macroscopic scale, sinter ideally consists of a strong, porous matrix which bonds relict ore nuclei together. Microscopically, the sinter matrix (solidified from the melt part of the initial sinter mix) generally consists of complex calcium ferrites known as SFCA, secondary magnetite and hematite grains (precipitated from the primary sinter melt), glass and silicates. Depending on the conditions under which the sinter structure is formed, two major forms of SFCA with different morphologies are generally recognised, ie fine micro-platy SFCA (usually termed SFCA-I) and coarse columnar/prismatic SFCA. Detailed optical microscope observations of several pot grate and compact sinter samples revealed the presence of a variety of fine, coarse and dense forms of both types of SFCA. The controlled cooling tests revealed that cooling rate has an impact on the form and size of SFCA crystals. In this work, EPMA analysis of SFCA compositions from a variety of pot grate and compact sinter samples are presented. Results show that the SFCA types exhibit a broad range of complex compositions within the system Fe2O3/CaO/SiO2/Al2O3/MgO/TiO2/MnO2. Although SFCA-I and SFCA are known to be crystallographically distinct and with different, well-defined ranges in composition, there was considerable overlap in compositions of the experimental sinter samples. The development of SFCA textures depend on several parameters including the local chemistry and conditions including the maximum temperature attained and the cooling rate, within the sinter bed, thereby suggesting that SFCA identification based on morphology alone may be erroneous.

Nov 8, 2021

By R Schiller, E Chait

"Quebrada Blanca is a porphyry copper deposit located in the Eocene-Oligocene belt of Chilean Andes. Operation started in 1994 and original reserves for supergene mineralisation gave a life-of-mine (LOM) of 20 years; however, further exploration extended the LOM until 2020.In order to optimise the remaining five years of supergene resource, geometallurgical variables were added to the geological model including copper recovery and acid consumption. LOM plan encompasses finishing the phase number eight and the future development of phases 11 and 13.LOM lithologies include diorites, quartz monzonites, porphyries, igneous and tourmaline breccias. Main alterations are: potassic, phyllic and argillic. Litho-alterations or geometallurgical units were formed by combining five types of rocks with three types of alteration. According to statistical information, the deposit can be represented with 12 litho-alteration units, which cover 98 per cent of the LOM volume.For copper recovery information, there were 40 mini-columns tests made in order to obtain kinetics curves and copper recovery values after comparing the difference between head and tails grades. Acid consumption was obtained using an automatic neutraliser device-based test.Historical data said that copper recovery in pads after 450 days of irrigation is around 75 per cent of recoverable copper while acid consumption is around 5.5 kg/t.According to this study, lower recoveries are related with potassic-altered diorites and argillic-altered quartz monzonites with values around 60 per cent. On the other hand, best recoveries could be found in potassic-altered quartz monzonites and argillic-altered diorites with values closer to 80 per cent.It was also revealed that the most reactive rock is the potassic-altered diorite with 12 kg/t of acid consumption. This rock covers approximately 40 per cent from phases 11 and 13.The final geomet model was calculated using a program that relates each geometallurgical unit with a specific copper recovery and acid consumption value. Results found after comparing model-obtained information with historical leached pads data were used to validate the model.CITATION:Chait, E and Schiller, R, 2016. Adding Copper Recovery and Acid Consumption Variables to the Geological Model of Quebrada Blanca, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 257–266 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By G W. Wilson

Net solar radiation is one of the key factors affecting evaporation from soil covers on waste rock dumps. While net solar radiation is normally taken as a uniform parameter over a site for modelling purposes, there can be significant variation in the radiation actually received on various portions of the dump. These variations are a function of the difference in the cover slope and direction of exposure. A predictive model has been developed to calculate the actual radiation received on the dump surfaces, based on the net radiation data measured at a weather station on a horizontal surface. The model has been designed for application at any mine site in the world, and can be used in conjunction with evaporative flux models to help estimate the effect of three-dimensional variations in the cover surface exposure. The waste rock dump at Equity Silver site (located in British Columbia, Canada) has a barrier-type soil cover system to limit oxygen entry and water infiltration. This cover has been monitored for the past nine years. The new net solar radiation model has been applied to this site and the predicted net radiation was compared to the values of actual net radiation measured on the site. The predicted variations in net radiation over the three-dimensional surface of the waste rock dump have been applied to calculate spatial variations in potential and actual evaporation (or evapotranspiration) from the site. Preliminary analyses indicate actual evaporation from a site at this latitude could vary by a factor of two for exposed versus shaded slopes. The implications of these variations for the site are considered.

Jan 1, 2003

By A Woodley, G Keir

The mining industry faces concurrent pressures of reducing water use, energy consumption and greenhouse gas (GHG) emissions in coming years. However, the interactions between water and energy use, as well as GHG emissions have largely been neglected in modelling studies to date. In addition, investigations tend to focus on the unit operation scale, with little consideration of whole-of-site or regional scale effects.This paper presents an application of a hierarchical systems model (HSM) developed to represent water, energy and GHG emissions fluxes at scales ranging from the unit operation, to the site level, to the regional level. The model allows for the linkages between water use, energy use and GHG emissions to be examined in a flexible and intuitive way, so that mine sites can predict energy and emissions impacts of water use reduction schemes and vice versa. This paper examines whether this approach can also be applied to the regional scale with multiple mine sites.The model is used to conduct a case study of several coal mines in the Bowen Basin, Australia, to compare the utility of centralised and decentralised mine water treatment schemes. The case study takes into account geographical factors (such as water pumping distances and elevations), economic factors (such as capital and operating cost curves for desalination treatment plants) and regional factors (such as regionally varying climates and associated variance in mine water volumes and quality).The case study results indicate that treatment of saline mine water incurs a trade-off between water and energy use in all cases. However, significant cost differences between centralised and decentralised schemes can be observed in a simple economic analysis. Further research will examine the possibility for deriving model up-scaling algorithms to reduce computational requirements.CITATION:Keir, G and Woodley, A, 2013. Regional trade-offs between mine water and energy use - a water treatment case study, in Proceedings Water in Mining 2013 , pp 21-28 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

By J Rostami, T Afshar, M Sharifzadeh

Accurate estimation of rock load and selection of types of ground reinforcement to match post-failure behaviour of rock mass are important issues to assure tunnel stability. This paper focuses on the main factors affecting rock load to evaluate the capabilities of the existing models for prediction of rock load. Rock mass properties, project features and excavation method, as well as ground conditions and stress condition were found to be the key factors affecting the magnitude of rock load. Field data from tunnels of about 37.5?km driven in several geological structures in non-squeezing ground was used in this study. An in depth statistical analysis was performed on data to find correlations between the input parameters and measured/observed rock load. A weighting system was proposed to obtain the proper contribution of each factor for estimation of rock load in tunnelling through various ground conditions. The new classification system offered predicted values for rock load that were in a good agreement with the actual field measurements. The high disturbance factor (D>0.4) was the prime factor which controlled the rock load. Rock mass quality was the second key factor influencing rock load, which could be estimated by GSI, GSIresidual , especially when combined with tunnel depth.CITATION:Afshar, T, Sharifzadeh, M and Rostami, J, 2014. Estimation of rock load for tunnels with different disturbance factors, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 3–8 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By Aiming Z, Shenyi S, Xialei W

Zhangmatun Iron Mine is situated in the suburbs of Jinan city. Shandong, China. Because of surface environmental constraints, no suitable site can he selected to dispose of mill waste. The full mill tailings must be disposed of underground. This paper deals with the work conducted at Zhangmatun mine to develop a total tailings backfill system, taking advantage of a double- shaft blade mixer and an activating agitator to prepare backfill. Cement and ground slag are added to tailings for strength development. Two scopes have been filled and the pillars have been effectively mined out during the field trial. The application of innovative total tailings backfill technology is a new concept at the mine. Detailed planning has achieved a close integration of mining, milling and backfilling. Full mill tailings, with or without rock, are disposed of underground in a continuous process to fill, totally and systematically, the cavities formed by mining.

Jan 1, 1998

The Marvel Loch Operations (MLO) plant was treating ores from up to six MLO sources and in addition needed to blend externally sourced ores with its own ores. This presented a metallurgical accounting dilemma, which was subsequently addressed by the design, implementation and testing of an automatic sampling system. The MLO automatic sampler was well designed and has been the subject of previous internal and external reviews. The sampling action is fully automated and the unit is essentially self cleaning particularly when sampling primary ores. The purpose of this paper is to document the rationale behind the design, the specific details of the sampling station, its use for ore purchase purposes and an assessment of the sampling system to meet the needs of the operator.

Mar 1, 2010

By R H DAVIES

Peak Gold Mine is a new gold and base metal mining project being developed by Enterprise Metals (a 100 per cent subsidiary of CRA) at Cobar in the centre of New South Wales. The Cobar district has a long history of copper and gold mining operations stretching back to the 1870s. Extensive gold and base metal mineralisation was discovered at the Peak in 1981 below a series of gold mines which had been active around the turn of the century. Evaluation of the deposit through the 1980s led to CRA's decision to develop the project for production at a capacity of 450 000 tonnes per annum in late 1992. The project has been developed with a green fields approach despite the long history of mining in the local area. This approach has lead to several innovative ideas being trialed at the site which are expected to result in the operation being a low cost producer of gold and base metals.

Jan 1, 1992

The Que River and Hellyer polymetallic massive sulphide ore bodies are located 3 kilometres apart in the Cambrian Mount Read volcanics of Western Tasmania. Both deposits are classic examples of the Kuroko volcanogenic style. Base and precious metal rich hydrothermal solutions rose through a deep fracture system and vented on the sea floor, causing sulphides to accumu- late in adjacent topographic lows. Together the preproduction geological resource is approximately 23 million tonnes at greater than 18% zinc plus lead, with economically significant silver and gold. Exploration by Aberfoyle in the Mount Read volcanics com- menced in 1970. Que River was discovered in 1973 following a comprehensive data review, a regional stream geochemical survey and a helicopter-borne electromagnetic survey. Production began in 1981, and the mine is currently producing 300,000 tonnes per year. The ore is sold to the Electrolytic Zinc Company of Aus- tralasia Ltd mill at Rosebery. Ten years of extensive geological mapping, soil geochemistry, and induced polarisation, combined with continual study of the Que River orebody and its host rocks, formed a strong frame- work for the discovery of the Hellyer deposit, which was found as a result of drilling .a subtle time domain electromagnetic anomaly in 1983. Hellyer is currently being mined at a rate of 250,000 tonnes per year, with the ore being concentrated at Aber- foyle's converted Cleveland tin mill. Treatment of 1 million tonnes per year will begin in early 1989 at a new mill located near the Hellyer site.

Jan 1, 1988

By J Jackson, E Sellers

Blasting related indices were developed to quantify variation in rock strength and structure for each stope, and for each lift of each stope, as part of a Mine to Mill study at an underground open stope mine. The indices were derived from combinations of data from underground diamond drilling because the simplifying assumptions contained in the block model prevented sufficient differentiation in rock mass characteristics to enable adjustments to the blast design.Application of combinations of blasting related indices enabled the identification of stopes with a high potential for inconsistent and oversize fragmentation, underbreak and overbreak. The indices also fragmentation outcomes to deliver increased material flow within stopes, better digability and predict indicative mill throughput. The stopes with issuesCalibration of the results enabled adjustments to the blast design based on the rock mass conditions and accordingly improved the blast outcomes. This enabled improved productivity by reducing underbreak and overbreak, having a more appropriate fragmentation, which reduces unplanned rock handling and improves bogging. The reduction in recovery drilling creates huge opportunities to improve output by reducing production delays and decreasing the overall cost per tonne. By understanding the explosive performance and blast energy distribution in relation to rock mass blastability poor, performing designs in blind upholes and convex stopes were identified.The conclusion of the work is that value improvement, regarding increased revenue and reduced costs, can be achieved across the whole underground Mine to Mill process chain by developing stope based indices from the existing geological information to correctly tailor the blast design to the rock within, and surrounding, the stope.CITATION:Jackson, J and Sellers, E, 2017. Blasting related indices as the key to underground blasting improvements in a challenging rock mass, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 23–32 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By Dobos SK

The Black Ridge Gold Deposit is located in the northern part of the Clermont Goldfield, which is situated at the boundary of the Bowen Basin and Anakie Inlier, Queensland. The goldfield comprises a series of NNW-trending semi-isolated basins, which are tectonically aligned with the Denison Trough, the west-most division of the Bowen Basin (Fig. 1-1).The main stratigraphic units in the deposits in younging sequence are the Anakie Metamorphics, the early Permian Blair Athol Coal Measures, the late Permian(?) White Section, and the Tertiary basalts which are regionally widespread, and intruded by abundant mafic and felsic plugs.

Jan 1, 1990

By T Blenkinsop, A Jenkins, H Moore, L Fisher, N Oliver, J Cleverley, M Nugus, J McLellan, J Hill, N Brunacci

The 2 Moz Vogue mineral resource is a recent discovery at the Sunrise Dam gold mine (SDGM), lying approximately 600 m vertically below the Sunrise Dam open pit. Unlike most other lodes, which are dominated either by steep vein sets or shallow high-strain (shear) zones, it lies broadly in the hinge of a folded felsic porphyry body, where it cuts intermediate volcanic lavas and volcaniclastic rocks and is not specifically concentrated at a lithofacies contact. The mineralisation is controlled by a complex interplay of moderate to shallow, west-dipping high-strain zones that are transected by irregular and discontinuous high-grade, north-west-, north–south- and east–west-trending breccia sheets that initially caused problems with attempts to model broad domains. However, by combining the structural controls and styles with the distribution of gold, arsenic, sulfur and antimony, we have improved opportunities for scheduling, extracting and processing the most appropriate material considering the current mining and processing constraints.The early delineation process for the Vogue lode revealed a significant mineral resource for the SDGM, coupled with the recognition of some potentially difficult downstream implications for mining and processing. The ultimate success of mining at Vogue will be a consequence of combining standardised data collection and geological modelling techniques with the results of the novel innovative techniques applied at an early stage of the project.Innovation (in the truest sense) drives change and application to deliver efficient and effective outcomes. Initial data generation and interpretation in lodes above Vogue provided a sound geological model based on strong structural foundations, leading to ideas that could be tested and validated by stress inversion analysis and modelling simulations. Geomechanical modelling simulations, based on the structural analysis, then provided an insight into both the structural and fluid evolution of the system and highlighted target areas. This promoted support for deeper drilling and sampling, which lead to the discovery of Vogue.With additional data, the structural model for Vogue was re-engineered through combination with geochemical and hyperspectral analysis and detailed geological and geostatistical domaining. Initial poor results from metallurgical tests from a small batch sample refocused efforts to undertake mineral and chemical characterisation and profiling (geometallurgy) and combine this data with the composite geological-structural-geochemical model. The result is a conditional probability model for differentiating mineralised or waste domains, with direct outcomes for planning, drilling and mining.CITATION:Nugus, M, Oliver, N, Blenkinsop, T, Hill, J, McLellan, J, Cleverley, J, Fisher, L, Brunacci, N, Moore, H and Jenkins, A, 2014. Exploration and innovation – the discovery and evolution of the 2 Moz Vogue Gold Resource, Sunrise Dam Gold Mine, Western Australia, in Proceedings Ninth International Mining Geology Conference 2014 , pp 323–340 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By J Yamatomi, Y Tamanishi, S Miwa

The four-dimensional (4D) network relaxation method, originally developed by Akaike in 1999, uses Lagrange relaxation and is known for its ability to include the price change and possible better outputs, but the computation time is huge. However, the continuous advancement in computer technologies allows the use of higher-speed computers with more memory capacity. Therefore, in this paper the authors have tried to implement the method with adjustments and improvements in its algorithm to suit the current computing system. As a result, the computation time has been drastically reduced and more constraints can be included, which makes the scheduling a more realistic one. The case study of an open pit gold mine with 129 500 blocks can be solved considering the capacity constraints in 71 seconds. In previous case studies, the mined ore had to be processed or disposed of immediately. Now, stockpiling can be included as another option. Since the details of the algorithm are still being improved, the computation time is a little longer than the previous cases without the stockpiling option. However, it is found that scheduling with the stockpiling option is possible in a reasonable time. In conclusion, the 4D network relaxation method has become more useful with advances in computer technologies and provides better results.

Sep 26, 2011

By D Pateman, F de Fredrick

The Bentley underground mine, which is located 60km north of Leonora, Western Australia, is owned and operated by Round Oak Minerals Pty Limited. A high-grade Zn-Cu Volcanogenic Massive Sulphide (VMS) orebody, Bentley was discovered in 2008 with production commencing in 2012. Long-hole open stoping is employed to mine the steeply dipping ore body, and in recent years the Avoca method has been adopted. This requires all development in a particular area of the mine to be completed prior to stoping commencing in a bottom up sequence. Voids are commonly backfilled with loose Run-of-Mine (ROM) waste loose rockfill (LRF); or alternatively cemented rockfill (CRF) can be used. Stopes in the 3860 AOS and 3900 AOS panels were located below previously extracted Avoca zones, with LRF situated immediately above them. These high-grade stopes were originally designed with rib and sill pillars to maintain control of the existing backfill during extraction. Due to changing factors over time, including mining economics, alternative options to achieve full extraction of the stopes were reviewed, including: • Mining small successive full height stopes which would involve drawing a blanket of LRF down with the extracted ore (stoping under waste); • Stabilisation of the existing LRF above the stopes and implementing a primary/secondary stoping sequence, with the primary stopes backfilled using a cement fill. An assessment of these options was undertaken which determined that the second full extraction method was the most beneficial to the operation, and could be achieved using a cement grout (Minova’s FB200 grout) to stabilise the LRF in conjunction with implementing a sequence where the primary stopes would be backfilled with cemented aggregate fill (CAF). Footwall drives were developed adjacent to the rockfilled stopes, allowing for the installation of horizontal grouting conduits, as well as providing access to the top of the planned 3860 and 3900 level stopes for the placement of CAF. An interval at the base of the previously rockfilled stopes was grouted at multiple horizons to achieve the required thickness, with a comprehensive quality assurance/quality control (QA/QC) process undertaken. Primary stopes were CAF backfilled via an arrangement of downhole winzes, linking the top of primary stopes to cross-cuts developed off the associated footwall drives. CAF was delivered underground through an open reticulation network that was extended to the specific levels. Overall the successful implementation of the revised mining method effectively increased ore recovery while maintaining control of overbreak and dilution, providing significant financial upside. This was achieved while maintaining effective controls in light of the changes to the risk profile. Stabilisation of the existing rockfill has allowed the extraction of stopes to full height, including the originally designed sill pillars, representing an average 23% increase in ore tonnes on the initial design, with dilution from the stope crowns averaging less than 1%. The use of CAF in association with primary/secondary stoping, eliminating the need for rib pillars, has yielded an average 47% increase in ore tonnes compared to the base case mine design.

Mar 15, 2021

By P Holtham, K C. Runge, E Tabosa

Traditionally, comminution and flotation have been surveyed, analysed and optimised separately. This often results in less than optimal outcomes because of the complex interactions between the two processes. Flotation recovery is a strong function of the valuable mineral particle size distribution with lower recovery of ultra-fines due to poor flotation kinetics, optimum recovery for the intermediate sized particles and lower recovery for the coarser particles due to poorer liberation. Thus, to optimise flotation recovery and grade it is usually best to grind to a fine size. This requirement needs to be balanced by the cost of grinding finer and the revenue that results from an increase in grinding throughput, which often results in a coarsening of the flotation feed.This paper demonstrates a technique which can be used to predict the effect of a change in feed grind size on flotation performance. It can be used to determine the optimum grind size for a particular operation which optimises total circuit profitability. Analysis is relatively easy to perform, as a minimum, requiring only size by assay data of the feed and product streams of the process. Case studies will be presented that demonstrate the usefulness of this technique in feed grind applications. Potential use of the technique for regrind evaluation is discussed. The limitations associated with this technique in certain circuit scenarios will also be outlined.CITATION:Runge, K C, Tabosa, E and Holtham, P, 2014. Integrated optimisation of grinding and flotation circuits, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 77–84 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By S Raval

The estimation of greenhouse gas emissions from coalmines is increasingly becoming a priority for both operators and regulators. At the same time, involved scientific challenges in the methane measurements is generating significant interest from global research community. Various technology providers are also actively engaged in improving their sensors and platforms to address the measurement challenges and capitalise the market opportunities. This presentation highlights rapidly evolving multifaceted landscape of coalmine greenhouse gas emissions. It covers the advances in sensors for the detections, including ground-based, airborne, and satellite-based systems as well as models for the emission estimations, including attempts to reduce the involved uncertainties.

Sep 1, 2024

A research and testing facility has been developed at Simtars to allow the spontaneous combustion testing of coal on a large scale. In this adiabatic reactor, the coal particle sizing is closer to that which could be found in the goaf and therefore the reaction is more likely to simulate the underground situation. This allows aspects of the spontaneous combustion process and methods for detecting a heating to be investigated which cannot be done in an operational mine. Up to 18 tonnes of run of mine or crushed coal is loaded into the reactor and allowed to self-heat while the temperature and gas profiles within the coal pile are monitored. The gases evolved from the coal are analysed using a fixed gas monitor system and gas chromatography to determine which of the common mine fire indicators and ratios can be used to predict the onset and progress of coal heatings. Benchmarking of the inherent propensity of a coal to spontaneously combust at both the large- and laboratory-scale is also possible by comparing the sample against previously tested samples. It has been found that the coal particle size affects both the time it takes for the coal to react and the location of the developing heating. The heatings generate approximately one metre from the inlet but in the crushed coal case do not progress to an advanced heating stage until the æhot spotÆ migrates towards the coal surface where there is an increased oxygen supply. Physical properties such as the compressive strength of the coal also affect the heating size and the time it takes for a heating to develop. Higher strength coals appear to oxidise more slowly and therefore take longer to self-heat as oxidation of the coal is the heat-generating step in the spontaneous combustion process. The best indicators of the state of the heating were found to be CO make and GrahamÆs ratio which are independent of air flow. Investigation of a potential spontaneous combustion should begin as soon as the CO make or GrahamÆs ratio exceed the background levels. The results from the tests undertaken on a number of Queensland and NSW coals are reported here together with the implications for the mining industry.

Jan 1, 2005

By G Cavanough, A Olsson, J Rock, A C. Torrance

There are various methods currently in use to determine detonation performance of bulk explosive products including velocity of detonation (VOD), high speed photographic analysis, fragmentation analysis and unconfined detonation of samples. For assessment prior to detonation the current practice is limited to a measurement of cup density. This paper presents additional techniques that have been developed to provide measurement and assessment of bulk explosive products. Techniques applicable prior to detonation include a mobile laboratory to allow on-site chemical and physical analysis of sensitised products and ‘in-hole density’ measurement to monitor the in-hole density during sleep time. Assessment of detonation performance is achieved using detonation pressure sensors. These sensors allow measurement of detonation pressure, average VOD, propagation time and magnitude of the pressure induced in the explosive column by detonation of adjacent blastholes.CITATION:Cavanough, G, Torrance, A C, Rock, J and Olsson, A, 2015. Measurement and assessment of bulk explosive products, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 477–484 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By C I. Edwards

Phosphorous is typically present in iron ore either in the mineral apatite, Ca5(PO4)3OH, or associated with the goethitic fraction of the ore. In this paper, we present results that show that phosphorous associated with the goethitic fraction can be removed with a heating pretreatment around 300 176;C for 1 h, followed by a caustic leach with 1 - 5 M NaOH, or a leach with sulfuric acid at pH 0.5. The heating pretreatment is considered to dehydroxylate the goethite to form a haematite intermediate phase, ‘protohaematite’, that reacts during the leach making the phosphorous accessible to the leachant. In the acid leach, ten to 15 per cent of the sample was dissolved to obtain a leach product with 0.08 per cent P. In the caustic leach, as no weight loss occurs, it is proposed that iron species associated with the protohaematite have either not dissolved or have reprecipitated at the high pH values of the leach.For an ore containing apatite as the source of phosphorous, a leach with sulfuric acid dissolved the apatite, lowering the phosphorous levels in the sample. The sample treated contained carbonate minerals (including siderite) that also reacted with the acid, and in practice a balance between the dissolution of the apatite and the carbonate minerals may be needed to minimise acid consumption.

Jul 11, 2011

By P Revell, G Wilkie, P Walters, L Keeney, F Faramarzi

Grade Engineering® is a value-driven strategy aiming to maximise benefits by exploiting intrinsic grade variability of ore deposits. It utilises a multitude of separation techniques to remove low grade material before being introduced to post-blast size reduction processes, i.e. crushing and grinding. At an operational scale, Grade Engineering® techniques can be implemented at the drill-and-blast stage, to control the degree of fragmentation in accordance with the spatial distribution of valuable minerals across an ore domain through differential blasting. It can also exploit intrinsic textural and breakage characteristics of a given ore type to reject uneconomical barren material at coarse sizes, which is known as Natural Deportment. Grade Engineering® can also be employed in the deliver stage of the mining value chain through sensor-based bulk and particle sorting. Each strategy is therefore highly dependent on the responsiveness of an orebody to available techniques. That is, the extent to which Grade Engineering® drivers would be capable of unlocking potential values of currently un-amenable orebodies depends on how detailed the heterogeneous nature of ore domains can be described. This paper considers exploring possible linkages between deportment characteristics of an ore and its competence when broken in a range of energy inputs at a particlebased level. Accordingly, potential opportunity may lie in exploring whether the intensity of input energy (or degree of fineness) can be used as a lever to improve or ‘unlock’ ore amenability to Grade Engineering® techniques. Implications from such a detailed approach offers new opportunities to enhance Grade Engineering® gains through engineering ore competence variability.

Nov 10, 2020