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Solid Energy New Zealand Limited owns and operates the Ohai coal field (Wairaki Coal Mining Licence, CML) in western Southland. The Morley Stream passes directly through the middle of the CML and has required diversion in the past to enable open cast mining to proceed. Solid Energy is evaluating two new open cast pits at Ohai. Commissioning these new pits would necessitate further diversion of the Morley Stream. One of these potential new mines is adjacent to an abandoned State Coal Mines (State Coal) pit and there is an opportunity to substantially rehabilitate this old pit as a corollary to the new mine and the stream diversion project.   Past Morley Stream diversions were undertaken by State Coal and so they predated the Resource Management Act 1991 and the degree of community environmental awareness and involvement that has become the norm these days. Not surprisingly the design and construction of the historic stream diversions was based purely upon mining (operational) and engineering considerations. This led to the stream being diverted in a manner that achieved the clinical conveyance of design volumes of water from points A and B.   Solid Energy realised that the proposed new diversions need to adopt more holistic and environmentally sustainable design principles while also including appropriate mining and engineering constraints. This project involved the design of a 1.7 km diversion of the Morley Stream (median flow of about 1000 litres/second) that would enable progressive prestripping and development of the new mines and would also address final rehabilitation (of the stream), geotechnical risks, hydrological and ecological issues. To address all of these factors the project required a multidisciplinary approach. Moreover, it required a clear understanding of how the mining operational constraints could be integrated with the environmental considerations. A significant aspect of the project was how to take account of fishery instream habitat needs, the resource consent requirements and most of all, the involvement of key stakeholders such as the Fish & Game Council, Department of Conservation, the Iwi and local community in the design process. This was achieved through coordinated consultation and stakeholder involvement in a technical design workshop.

Jan 1, 2005

Flotation testing has been done at laboratory scale on hematite mineral and ore samples using stage additions of various hydroxamates. Chemisorption of hydroxamate species onto the hematite surface was inferred from the results, possibly involving both the undissociated hydroxamic acid and hydroxamate ion as maximum recovery occurred close to the pKa of the hydroxamic acid. Flotation testing on a specular hematite ore containing 40 per cent Fe and 42.5 per cent silica gave highest iron grades in concentrate at pH 6, in contrast to published data which suggested pH 9 as the optimum value for ore processing. Using octyl hydroxamate, a rougher concentrate assaying 67 per cent Fe at 51 per cent recovery was obtained using four stage additions of 50 g/t at pH 6. For lower grade concentrates still meeting the chemical analysis for pigment grade micaceous hematite (minimum Fe assay 60 per cent) iron recoveries of 80 per cent were possible for stage additions of 50 g/t octyl hydroxamate in the pH range 6-8.

Jan 1, 1999

A process was developed based on the thermal hydrolysis and precipitation of iron from solutions resulting from the leaching of zinc plant residue. A study was made of the hydrolysis of iron, at temperatures in the range 140-220°C, in both pure solutions of ferrie sulphate and in leach solutions derived from plant spent electrolyte.At 200°C hydrolysis of pure ferric sulphate solutions results in the precipitation of ferric oxide, Fe2O3, at low acid. concentrations, and a basic salt, Fe2O3.2S03.H20, at hIgher acid concentrations. At 180°C ferric oxide is precipitated at low acid concentrations, but at higher acidities. carphosiderite, 3Fe203.4SO.3.9H20, is the stable phase. Frequently a mIxture of ferric oxide and carphosldente IS produced at 180°C, the ratio of phases being dependent on the non-equilibrium experimental conditions.With solutions derived from plant spent electrolyte hydrolysIS at elevated temperatures results in the precipitation of jarosite compounds of the type NaFe3(S04)2(OH)6 The precipitation of natrojarosite jarosite, and ammonio-jarosite was studied in detail. Temperature, free acidity, and the concentration of sodium potassium, and ammonium ions are critical factors which affect the extent of precipitation.A process based on ammonio-jarosite precipitation was tested in a pilot plant.An alternative method of iron removal based on the addition of zinc oxide calcine, to neutralize excess acidity, and the precipitation of ferric oxide was also tested in the pilot plant. Both processes appear economically favourable...

Jan 1, 1967

By D R. L Vedoy, L Botha, S P. Gumfekar, J B. P Soares

Most of the known in-place reserves of oil sands are found in Canada. These oil sands ores are mined using various processes to extract bitumen. The by-product slurry resulting from the extraction process, known as tailings, is discharged into tailings ponds. Fine kaolin clays are an important ingredient of the tailings because they remain suspended in water for many years. Current polymeric flocculation/dewatering technologies use acrylamide-based polymers such as polyacrylamide (PAM), but sediments produced using PAM-based flocculants hold considerable amounts of water due to the formation of loosely-packed flocs. Anionic polymers are better flocculants due to bridging and the charge-neutralisation mechanism. In this paper, we will compare the synthesis of two flocculants: thermosensitive, anionic flocculants and thermosensitive, hydrophobically modified anionic flocculants. The first flocculant is a copolymer of acrylic acid (AA) and N-isopropyl acrylamide (NIPAM), in which AA is the anionic monomer. The second flocculant is a copolymer of NIPAM, AA and the hydrophobic monomer N-tert-butyl acrylamide. We evaluated the dewatering efficiency of 2 wt per cent kaolin using both flocculants by comparing settling rates at different flocculant dosages, charge neutralisation ability and supernatant clarity. To better understand floc formation, we measured the floc size formed by both flocculants at different flocculant dosages. We propose that the hydrophobic monomer in the terpolymer flocculant helps expel water out of flocs, ultimately decreasing the water content in the sediments.CITATION:Gumfekar, S P, Vedoy, D R L, Botha, L and Soares, J B P, 2015. Thermosensitive, hydrophobically modified anionic flocculants for dewatering kaolin clays, in Proceedings Tailings and Mine Waste Management for the 21st Century , pp 45–50 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By R L. Brathwaite

The epithermal gold-silver deposits wall poster is one in a series of posters on different mineral deposit types being produced by GNS Science to raise awareness of the characteristics of New Zealand mineral deposits and their prospectivity. Epithermal deposits consist of quartz veins, quartz vein stockworks and hydrothermal breccias that are mainly hosted in subaerial andesitic or rhyolitic rocks of Tertiary age. The deposits were formed at depths of typically less than 1 km and at temperatures between 180¦ and 320¦C, although ore grade mineralisation was generally deposited at less than 270¦C.   In New Zealand, epithermal deposits occur in association with late Cenozoic volcanic rocks in the Northland, Coromandel and Taupo volcanic regions. All of the past and present gold production, estimated at greater than 380 t (along with 1300 t Ag) comes from the Miocene-Pliocene Coromandel Volcanic Zone, where about 50 epithermal gold-silver deposits occur over a 200 km long belt. In Northland, silver and mercury have been produced from hot spring type epithermal deposits, and several gold-silver prospects have been identified. Ore grade gold-silver mineralisation is currently being deposited in several active geothermal fields of the Taupo Volcanic Zone, and extinct geothermal systems occur on the western side of the zone. The active systems have provided natural laboratories to help understand the processes of epithermal mineralisation.

Jan 1, 2005

By R G. Moreby, S Bluhm, FávonáGlehn, C Pascoe

Resolution Copper Mine in Arizona, USA, is planned to be a 2000 m deep panel cave mine with virgin rock temperatures exceeding 80¦C in rock with high crystalline silica content, resulting in a very challenging ventilation and cooling system design. This is one of the largest copper orebodies ever found and, with a planned rock production rate of 120 kton/day. The project is currently in its prefeasibility phase and the paper discusses features of the ventilation and cooling system design which includes multiple ventilation shafts with total primary ventilation flow of about 3000 m3/s and both surface and underground refrigeration plants producing more than 140 MW refrigeration capacity. The cooling system features large surface bulk air coolers chilling all downcast air supplemented by an underground secondary air cooler system and a chilled service water system. This will be a very challenging mine to ventilate and cool but this work has demonstrated that it will be technically achievable with the application of existing technology.CITATION:Bluhm, S, Moreby, R G, von Glehn, F and Pascoe, C, 2013. Life-of-mine ventilation and refrigeration planning for Resolution Copper Mine, in Proceedings The Australian Mine Ventilation Conference , pp 125-138 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By K Faure

At Sams Creek a gold-bearing, peralkaline granite pophyry dike intrudes Ordovician-Silurian metapelite and quartzite. These metasedimentary rocks have undergone three phases of folding: a first phase of recumbent folds (F1), a second phase (F2) of northerly-trending inclined folds with a prominent crenulation cleavage, and a third phase (F3) of steeply plunging folds. The granite porphyry dike has a thickness of up to 40 m and intrudes thin lamprophyre dikes along its contacts. Deflection of the crenulation cleavage around the granite dike, shearing within the dike and its sinuous outcrop pattern indicate that the granite dike has been deformed during F2 and F3. The granite has been hydrothermally altered, with a first stage of magnetite-ankerite alteration. This first stage is cut by a second stage of quartz-pyrite veins and a third stage of sulphide, quartz and ankerite veins, accompanied by pervasive silicification represented by quartz-perthite-albite-rutile¦pyrite-ankerite-sericite assemblages. Lamprophyre is altered to an ankerite-chlorite-sericite assemblage. The sulphide veins are composed of arsenopyrite + pyrite ¦ gold ¦ galena ¦ sphalerite ¦ chalcopyrite ¦ pyrrhotite. The altered granite and the veins have been deformed and recrystallised. The competent sulphide minerals pyrite and arsenopyrite are fractured and brecciated, with fractures locally filled by galena, a ductile sulphide mineral. Vein quartz shows local cataclastic and mortar textures and develops pressure shadows around arsenopyrite. Alteration and mineralisation are confined to the granite porphyry and adjacent lamprophyre. These features together with the style of veining and the sulphide mineralogy indicate that the mineralisation was formed from a magmatichydrothermal fluid that was sourced from the granite. Oxygen, carbon and sulphur isotope values of the vein minerals (quartz, ankerite and sulphides) are also consistent with a magmatic source.

Jan 1, 2003

Wet mining of auriferous gravels at L & MÆs Waikaka operation in Southland creates spoil dumps derived from stripping of a clay rich overburden, tailings from past dredging and the processing of auriferous sands and gravels. Over recent times, spoil dumps have slumped on six occasions, threatening the viability of the wet mining operation. Close examination of slump morphology, coupled with the sequence of material placement and rate of dumping suggested a mechanism associated with undrained failure of a 2 m layer of silt and sand tails forming the base of dumps.Commencing with some information on the geotechnical properties of source overburden, back analysis of failure morphology and sequence and development of undrained loading conditions from first principals, the most likely failure mechanism and factors controlling failure were defined. Based on analysis results, a scheme to stabilise the lowwall dumps was developed and implemented, augmented by continuous monitoring of pore pressures in the basal wet tails, and laboratory testing for confirmation of computed material properties.Lowwall spoil stability has improved since implementation involving changes to the method of spoil loading, and modification of wet tails composition to improve permeability and drainage of the basal wet tails under load. The method and approach to prediction and verification of failure mechanism provide a sound basis for continued mining of deeper auriferous gravels.

Jan 1, 2002

By Boontanjai C, Boyapati ER

Reagents used for batch flotation tests were Aeroxanthate-343 & 350, Dow Z-6, & reagent-4037 (from Benambra) as copper collectors; Na2SiO3 & Na2S2O3as gangue depressants; lime & HCl as pH regulators; and pine oil & MIBC (methyl isobutyl carbinol) as flotation frothers. A Denver Sub A flotation cell of capacity 250 gms was used for the batch flotation experiments. Atomic absorption spectrophotometer was used for copper assays and microscopy for selectivity and liberation studies. Melbourne tap water was used throughout the experiments. Unless otherwise stated, the following parameters apply to the flotation experiments: flotation impeller speed - 1300 rpm, pulp density during flotation- 22%, reagents addition - in 4 stages, total conditioning time for collectors - 8 minutes, total flotation time- 9 minutes. A variable quantity of copper mineral collector according to the test in question was added, then conditioned the pulp for 2 minutes. The froth was removed by hand, scraping the full length of the cell at constant depth every 10 seconds. No frother is required for this copper tailing flotation. Four concentrates were collected after addition of collector and conditioning at 2,4,6 and 9 minutes. The pulp volume was maintained at a constant level to yield a constant froth depth.

Jan 1, 1990

By G Chitombo, G Lewis

The knowledge of the expected run of mine (ROM) fragmentation during the feasibility study of a green field site is essential for the design of downstream crushing and milling processes. However,the prediction of blast fragmentation can be a challenging process due to the uncertainty related to the blast design and rock mass parameters. In general,a number of possible blast designs are considered in a feasibility study,with factors such as bench height and blasthole diameter influencing the design. Rock mass parameters are normally derived from a limited number of geotechnical diamond drill cores. Although diamond drill core can provide detailed information concerning intact rock strength (unconfined compressive strength and point load),it provides limited structural information to predict the in situ block size distribution of the rock mass. The in situ block size is an essential input to the prediction of the fragmentation distribution, in particular the coarse end. This paper outlines the methodology of blast fragmentation modeling undertaken for a green field feasibility study at the Riska gold deposit in Indonesia. The favoured milling process for the feasibility study was dump leaching,with no crushing of the ore material extracted from the pit. For this reason,blast fragmentation was a critical issue to be addressed by the study. A range of blast designs were considered with bench heights and blasthole diameters ranging from 4 m to 7 m and 76 mm to 102 mm respectively. Rock mass data was obtained from 19 diamond drill cores across the deposit (total drill length approximately 2200 m). Intact rock strength was estimated from qualitative strength descriptors,while the in situ block size distribution of the rock mass was estimated from the Rock Quality Designation (RQD) of the core. The methodology of fragmentation prediction involved a number of steps including the estimation of rock mass parameters from diamond drillhole data including the estimation of the in situ block size, delineation of blast domains within the deposit,and the assessment of the impact of various blast design parameters and rock mass variation on the predicted fragmentation distribution. The modelling process for the Riska green field site highlighted the need for additional rock mass data to be collected post feasibility stage and once the development of the mine commenced.

Jan 1, 2004

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 R M. Smart, A T. Harrison

Backfill systems using unconventional cementitious binders have been successfully used in relatively deep and hard rock tabular mines for longwall, scattered stoping and shaft pillar removals. Two unconventional cementitious systems are discussed: A ground granulated blast furnace slag (GGBS) with an activator containing lime. This has been used with a total tailings æstableÆ slurry at a relatively low dosage of less than seven per cent, by mass, based on dry solids. The room and pillar stopes supported with this fill were up to 20 m high and the tailings were relatively high in pyrite. A GGBS binder plus a silicate based activator and gelling agent that virtually eliminated all water drainage (and hence post filling shrinkage). The system used a slurry, but the addition of the activator created a final placed product that had the behaviour of a paste. The system has been successfully used for almost ten years in the following applications: stoping widths (mining heights) above 1.6 m; and shaft pillar removals in seismically active areas at the end of the life of the shafts. This paper discusses the technical merits and costs of the systems and compares them to other fill systems (including true paste systems). the acceleration (normally g, 9.81m/sec2) is increased to a significantly larger (but generally unmeasured and unknown) value, as shown in Figure 1.

Jan 1, 1998

By T Haskell, Y Joe, S Dickie

A current trend amongst producers is the increase in dewatering of iron ore fines (<0.5 mm), due to an increase in wet beneficiation processes used to upgrade lower quality ore deposits, increased water content in orebodies located below the water table, and increasing focus on transportable moisture limits (TML). Vacuum filtration is commonly used for dewatering of iron ore fines, and as with all filtration processes performance decreases with particle size. Significant improvements are possible with the use of filter aids, usually anionic surfactants and flocculants. Anionic surfactants reduce the surface tension of the associated water, which is then removed more readily leading to improved filtercake moistures. Anionic flocculants used correctly can improve cake permeability, in turn, decreasing cake formation times and increasing filtration rates. In addition to increased throughput, savings and benefits for producers can include: moisture limit penalties, improved handleability, along with energy and water savings.The effect of filter aids on two iron ore fines samples from the Pilbara region were evaluated using a lab scale vacuum filtration unit. The first sample showed a reduction in the filtercake moisture content by 2.3 per cent weight/weight (w/w) with the use of BASF’s Drimax® 1234. For the second sample a reduction in the filtercake moisture of 3.4 per cent w/w along with an increase in filtration rate of 20 per cent was achieved by using a combination of Magnafloc® 5250 and Drimax® 1239 NF, as both filter throughput and cake moisture were considered important by site.CITATION:Haskell, T, Dickie, S and Joe, Y, 2017. The use of filter aids on iron ore fines, in Proceedings Iron Ore 2017, pp 183–186 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

The major technological advances of the last 30 years have profoundly reshaped mining practices. Machines keep getting bigger while productivity continues to increase and manning levels drop. These trends will continue and the introduction of breakthrough technologies will continue to accelerate, driven by a continued focus on lowering the cost of production, improving the productivity of workers and machines, and ensuring safer working environments. At the same time, global consolidation of mining companies has impacted on technology development in a range of ways. There are many potential technologies being proposed, some already operating in Australian mines. The automation of underground load-haul-dump (LHD) trucks for example, continues the process of removing workers from the most dangerous operating environments. Global positioning systems fitted to draglines, fibre optic communication channels, and 400 tonne trucks are only some examples of what lies ahead. However, where we need to return focus is to the technology of management û which is also one of the two key interrelated themes of this keynote address. Have we, as an industry, adequately been able to keep up with technological advances? Are our management systems as competent as the machinery we employ? The other key theme is the processes of globalisation and consolidation. What are the impacts on our industry? What are the implications of these processes, particularly for the ways in which we manage our people? My paper is divided into several components:some comments on the coal industry and the dramatic restructuring we have seen in recent years; in the context of this restructuring, an overview of my company and its evolution; my views on the role of contractors; some comments on the consolidation also occurring in the mining machinery and equipment, technology and services sectors; some observations on two major areas of technology û drill and blast; and trucks, shovels and excavators; the implications of consolidation and new technology for our employees and management; and finally some comments on research programs.

Jan 1, 2001

Coal mining in New Zealand is principally a State-owned and controlled operation where the State produces 70 per cent of the coal in New Zealand and 65 per cent of coal mined in the Waikato. The State's flexibility in production quality and price makes them tough competition. Coal in the Waikato varies in quality considerably throughout the region making the high quality coal deposits as in the Huntly district a highly sought after commodity. The high quality of coal is often offset against transport costs as, for example, Maramarua coals can compete effectively in the Auckland market and dollar per calorific value basis.

Jan 1, 1984

Minerals Industry - 5 Year Outlook AusIMM sponsored project Positioning the Minerals industry for the 21st Century Does this seem right for New Zealand?  To ascertain the industry's perceptions of issues and capacity to address them: Industry context and issues Capabilities of professional staff Access to professional staff Balancing supply and demand Industry must do it. Cannot wait for others.   Some observations... Negative image - constrains capacity to influence community and government Major repositioning of professional staff competencies required Existing organisational arrangements and capabilities are under pressure to change Difficult to access professional staff of required capability Facing significant external challenges - with real concern about capacity to address them Prevailing culture not suited to needs of the future High level felt need for change The industry has a choiceà We can define where we want to be and act decisively Or We can respond to the future as it unfolds.

Jan 1, 2000

Gold was first discovered on the Shotover by Thomas Arthur in 1862, and soon there were over 3000 miners working on the river. It has often been called the richest river in the world. Most countries probably have one of these, but it was certainly the richest in New Zealand. In the early days some miners collected as much as 200 oz/d from a sluice box, and the early river gravels must have been extremely rich before the gold bearing terrace gravels were sluiced into the river. Otago produced about 10 Moz out of the NZ total of over 30 Moz. Most of this was alluvial and possibly one third or more came from the Shotover. The extremely cold winter and danger of flooding made it a dangerous river to work. During 1863 when heavy rain melted snow in the high country, the river rose 35 ft overnight and drowned many miners camped on the banks. These hazards still exist and dominate the method of working. The early hand miners cleaned out much of the riverbed, and the advent of the hydraulic elevator enabled deeper parts of the river to be worked over. Extensive sluicing of high terraces also took place particularly around Skippers. Several companies have attempted operations on the Shotover. Many of these were distinctly unsuccessful, such as the suction dredge at Maori Point about 1930 and the Skippers Limited fluming in about 1934. At least two attempts were made to divert the river with the Sandhills cut above Skippers and the Oxenbridge Tunnel down near Arthur's Point.

Jan 1, 1983

Iron formation-hosted iron ore deposits account for the majority of current world iron ore production and consist of three classes: unenriched primary iron formation with typically 25 to 45 wt per cent Fe; martite-goethite ore formed by supergene processes, with abundant hydrous iron oxides containing 60 to 63 wt per cent Fe; and high-grade hematite ores thought to be of hypogene or metamorphic origin overprinted by subsequent supergene enrichment with 60 to 68 wt per cent Fe.Individual iron ore deposits range from a few millions of tonnes to over two billion tonnes at >64 wt per cent Fe, although most are within the range of 200 to 500 Mt. In the Hamersley Province of Western Australia, martite-goethite ores are largely developed in the Marra Mamba Iron Formation, although high (>0.08 per cent P) phosphorous mineralisation is also well developed in the stratigraphically higher Brockman Iron Formation. Whilst the vast majority of high-grade microplaty hematite ore is best developed in the Brockman Iron Formation, this paper provides the first textural evidence of locally significant microplaty hematite mineralisation in the Nammuldi Member of the Marra Mamba Iron Formation, in the Chichester Ranges at the Christmas Creek, Cloud Break and Mount Nicholas prospects. Petrographic studies have identified a high-grade Fe texture composed of nanometre scale plates of hematite in mineralised sections of primary microplaty hematite deposits in the Pilbara and elsewhere, well below the normal depth of weathering or dehydration. The population of nanometre to micrometre scale hematite plates is interpreted to represent various stages of nucleation, crystallisation and progressive growth of hematite from the primary ore-forming fluid in areas that were once iron-rich carbonates or silicates in the BIF.

Jan 1, 2005

By J Davidson

Ion exchange using strong base resins is a well established technology for the recovery and purification of uranium from sulfuric acid leach liquors. Production of uranium at the Beverley Mine in South Australia includes acidic in situ leaching, ion exchange and uranium precipitation. The salinity present in the groundwater results in a relatively high concentration of chloride in the leach liquor reporting to the ion exchange plant. The impact of chloride on the ion exchange process is discussed in this paper. The results presented show that an increased chloride concentration in the lixiviant solution feeding the ion exchange plant leads to increased uptake of chloride by the resin during the loading cycle. At the highest chloride concentration studied (7.5 g/L Cl) chloride took up to 60 per cent of the resin loading sites. The impact of the chloride can be summarised as follows: for a given uranium concentration in the lixiviant feed solution, the uranium loading decreased with increasing chloride concentration; a consequence of decreased uranium loading is an increase in the frequency of elution, with an associated increase in elution reagent costs; and for a given elution efficiency, the uranium concentration in the barrens is expected to increase with increasing chloride concentration.A few options were examined to counteract the impact of chloride: an increase in elution efficiency was targeted to reduce the amount of uranium returning to the loading circuit; increasing the pH improved uranium loading, but this was found to also increase iron extraction; and preliminary screening of alternative resins showed that the higher the theoretical ion exchange capacity of the strong base resin, the higher the uranium loading.Work is continuing on uranium recovery from saline leach liquors, specifically investigating the performance of alternative resins.

Jan 1, 2005

By D David

Clearly a plant designed only to treat average ore at the nameplate rate will fail to achieve nameplate in any typical year. To insert the necessary capability to achieve nameplate it is common process engineering design practice that plant must be able to treat an ore with 80th percentile value (hardness or competence values for example) at the nameplate rate. In the authorÆs experience, apart from a couple of notable cases, this has been applied reasonably well as a principal, and with widespread success. However, in a number of recent instances it became clear that the use of the 80th percentile number would have resulted in significant under design of the plant. This paper makes the case that the 80th percentile, as a principal, has can have serious flaws and its use needs to be assessed on a case by case basis. The discussion in this paper is the first step in developing a new design principal, and the associated methodology for selection of a design value, that will ensure plants are designed to achieve their nameplate capacities.CITATION:David, D, 2013. Is an 80th percentile design point logical?, in Proceedings MetPlant 2013 , pp 6-12 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013