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By S. Perez

Core drilling is a fundamental process in the mineral exploration industry. As deep exploration in hard rocks is accelerated due to the depletion of near ground mineral resources, deep-drilling operations need to reduce their reliance in operators experience and begin to implement new technologies in order to increase efficiency and cost-effectiveness of the overall process. Currently, lack of real-time information on the bit/rock interaction is a major problem. Therefore the mechanism behind the acoustic emission, which is a proven technique for remote monitoring, generation in mechanical rock cutting must be understood in order to provide real time information about the rock. This paper aims to identify the influence of mineral composition of granitic rocks on the acoustic emission (AE) during rock abrasivity test with a single contact tool and to evaluate the feasibility of AE as a means to infer Cerchar Abrasivity Index (CAI). Single contact tool will provide precise and accurate description of the source of AE. To provide constant grooving speed Cerchar apparatus was modified to monitor AE during rock/tool interaction. Tests on the modified Cerchar apparatus are carried out under the static load of 70N and cutting speed of 1 mm/s. A Scanning Electron Microscope (SEM) is utilized to visualize the groove in the rock and to identify the mineral composition through Back-Scattered Electrons (BSE) reflected from the rock sample. Based on the microscope images, it was found that biotite shows a larger groove area and plate-like fractures accompanied with burst-type AE signal, whereas plagioclase and quartz, which are characterized by similar hardness and BSE brightness, possess shattered-like fractures and more continuous-type AE signals. In addition, a-value of the Gutenberg-Richter scale has been found to well correlate with Cerchar Abrasivity Index (CAI).

Jan 1, 2015

By E. Wolmarans

The selection of the main mineralised zone (MMZ) concentrator milling circuit has followed a comprehensive route over the preceding 10 years. Many knowledgeable parties have contributed and assisted in developing the milling circuit according to the best available information. The final selection of the flow sheet was completed after DRA conducted a techno-economic study on four possible comminution circuits ranging from conventional crushing and ball milling to fully autogenous (AG) milling. The results from all the above mentioned work indicated that it is possible to treat the MMZ ore at the desired throughputs to the desired grind sizes in a fully autogenous grinding circuit. The flow sheet was designed to treat the main mineralised zone (MMZ) arising?s from both underground and open cast operations and incorporates a fully autogenous grinding (AG) circuit. Commissioning of the 375ktpm concentrator at Nkomati Nickel began in August 2009 with the first ore reporting to the mill on the 15th of September 2009.

Jan 1, 2012

By D. P. Bishop, G. J. Kipouros, R. L. Hexemer Jr., I. W. Donaldson, W. F. Caley

Aluminum powder metallurgy (APM) is an advanced metal forming technology frequently employed to fabricate complex structural components in a near-net-shape fashion. Exhibiting nearly 100% material utilization, APM continues to rise in popularity. Recent research initiatives in this field have emphasized the development and processing of advanced 2xxx and 7xxx series alloys. In particular, APM equivalents to wrought 7075 and 2024 have been sought. An overview of work completed in this area is presented. Included are data on the various production stages (powder compaction, sintering, forging, and heat treatment) as well as the microstructure and mechanical properties of the finished products.

Jan 1, 2010

By Charles Amoah

Abosso Goldfields Limited has been optimizing the comminution circuit at the Damang Mine Operation for the treatment of 100% primary ore. The optimization process started at a base rate of 343 tph when treating primary crushed primary ore, and was increased to above 600 tph with partially crushed primary ore. Abosso Goldfields engaged Orway Mineral Consultants (OMC) to assist with the optimization process in 2009 and the two companies have been working together since, to further refine the process. Continuous improvement was achieved through crushing circuit modifications, changes in operating processes and grinding control system utilization. This paper discusses the improvements developed for the increase in milling capacity and improvements in final product size.

Jan 1, 2012

By Z. F. Zhou, X. Y. Zhou, V. G. Papangelakis, Z. D. Jankovic, S. M. Ulyanov, S. N. Lvov

Recently developed flow-through electrochemical techniques (1,2) were employed for measuring pH in high concentration aqueous solutions and geothermal brines. The pH values of (1) 5.0 mol kg-1 NaCl + x mol kg-1 HCl, (2) 3.0 mol kg-1 of NaCl + 0.5 mol kg -1 of KCl + 1.0 mol kg-1 of CaCl2 + x mol kg-1 HCl (x < 0.1 mol kg-1), and (3) mixtures of Al2(SO4)3-H2SO4-H2O, MgSO4-H2SO4-H2O, and NiSO4-H2SO4-H2O solutions have been measured at temperatures from 100 to 280°C. The obtained experimental data and their thermodynamic analysis demonstrated that the electrochemical system can be used for on-line pH monitoring of high concentration aqueous solutions with an accuracy of ±0.1-0.3 pH units over a wide range of temperatures. Greater deviation exists in real laterite leach solution measurements due to in situ precipitation of Na-alunite. The system is excellent for research and process development work. Further work is needed to ensure robust operation under pressure leach conditions of laterites.

Jan 1, 2004

By Geoffrey Gilbert

Rossland is in the West Kootenay district; 5 miles north of the International boundary and 4 miles west of Columbia River at Trail. The first claim in the district was located in 1887; but there was little activity until 1890; when the showings that later became important were staked and development began. Production increased rapidly from 1; 800 tons in 1894 to a peak of 360; 000 tons in 1903; and averaged 286; 000 tons for the next 13 years. It declined sharply in 1917; and continued to decline until the closing of the main mines in 1928. Between 1933 and 1941 leasers mined a further 137; 000 tons. Total output has been about 6; 200; 000 tons; with average recovery of 0.47 oz. gold; 0.6 oz. silver; and 1 per cent copper. The total value was $82; 500; 000; of which aboµt three-quarters was in gold; the camp still leads the province in total gold produced. Ninety-eight per cent of the tonnage came from a small compact area worked by four inter-connected mines : the Le Roi (39 per cent); Centre Star (25 per cent); War Eagle (24 per cent); and Josie (10 per cent); all now owned by the Consolidated Mining and Smdting Company of Canada; Limited.

Jan 1, 1954

By P. J. Le Roux

"Mining worldwide and definitely in South Africa, is constantly under pressure to reduce its cost structure so as to sustain profitability. In underground gold mines where an open stope mining method is employed, dilution often has a significant effect on the viability of sustaining profits. Target Mine practices Open Stope mining Method and it was found that in some open stopes dilution was in excess of 10%, which has a significant impact on the sustainability of the mine.Dilution in excess of 10% can result in the reduction of the recovered grade from 5,5 to 4,5 grams per ton (g/t). The reduction of 1 g/t in recovered grade results in a potential loss of about ZAR21 Million per month bases on a gold price of ZAR240 000 per kilogram. Based on Life of Mine projections the potential loss of income could be as much as ZAR3,3 Billion. A reduction in dilution would have the opposite effect. There are a number of factors which affect dilution of which falls of ground in open stopes is a major contributor. The falls of ground can be attributed to a number of factors such as beam failure as a result of a larger than normal expected roof area (hydraulic radius too large), poor ground conditions, and poor blasting. The cost of damage to or loss of trackless equipment as a direct result of the falls of ground in open stopes, is very significant. The review of financial figures has indicated that this could be as high as ZAR491 million. This combined with the added cost of transport, hoisting, secondary blasting, milling and plant treatment costs of ZAR293 results in an estimated opportunity loss of ZAR784 million.Currently there is a significant amount of data available in the mining industry which could be effectively used to develop suitable back analysis techniques but to date this has not been used effectively. If dilution can have such an impact on current and future mining ventures then the optimization of back analyses for the prediction of dilution in open stoping could assist significantly in the reduction of dilution in massive open stopes. Rockmass classifications, geotechnical information, blast techniques, blast design, the stress strain environment and hydraulic radius all have an effect on, or play a role in the evaluation of dilution. Each of these factors will be taken into consideration to ultimately determine a measurable or calculated percentage of dilution in massive open stopes.The amount of overbreak in an open stope can be determined by subtracting the planned stope volume in m3 from the actual measured final stope volume in m3, which is obtained from the CMS (cavity monitoring system). This is in turn divided by the planned stope volume in m3 to determine the percentage overbreak. The CMS wireframe is imported into the geological model and its grade re-evaluated. From this the actual percentage dilution for this open stope can be determined. The dilution obtained can result in a major reduction of recovered grade for the open stope, but can also increase the recovered grade. The latter is normally the exception rather than the rule.When analysing data from Target mine the following was achieved:• Using 11 years of data a method of measuring and predicting the percentage dilution in open stoping was developed. This took into account rock mass quality, stress-strain state, and the hangingwall hydraulic radius (size of stope hangingwall exposed).• Implementation of this prediction method resulted in a reduction in falls of ground in open stopes. The benefit of this was a reduction in the damage to mechanised equipment resulting from less falls of ground which had a positive effect on the profit margins of the mine. As a direct result the recovered grade from the open stopes increased due to the reduction in the amount of dilution.• A design criterion, Dilution Stress-Strain Index (DSSI), was developed which allows the user to calculate, with certainty, the stability of the open stope and determine if m"

Jan 1, 2015

By G. P. Demopoulos

The solubilities of calcium sulfate in HCl-CaC1z-H20 solutions up to 100°C were successfully modelled with the aid of OLI software after bringing proper modification to it. Modelling, more specifically, was carried out with the help of the Environmental Simulation Program, V-6.6.0.4 (www.olisystems.com). New model parameters were obtained by regressing the solubility data. A private databank was established by adding the newly defined model parameters and the calcium sulfate hemihydrate phase that is not included in the original OLI database. The McGill-modified OLI software was successfully applied not only to modelling solubilities but also the transformation of the various CaS04 phases, namely dihydrate, hemihydrate and anhydrite. In this paper, the modelling methodology will be discussed along with a comparison of the capabilities of the newly developed model vs the original OLI model.

Jan 1, 2005

Energy benchmarking is widely applied across industries to determine the potential to improve energy performance. Unlike the many technologies for which energy benchmarks and efficiency rating systems are already in maturity, benchmarking of mining processes is still in its infancy. Benchmarking energy usage in mining operations, where energy costs are high and represent a significant portion of overall operating expenses, is challenging due to the uniqueness of each mining operation. Mines vary in terms of geology, geographic setting, ore body geometry, mineralogy which impact the mining method, material handling, processing and waste management. In order to gauge the opportunity for reducing the energy intensity of individual operations and also to compare the energy performance of operations on a global scale, the deployment of new benchmarking methodologies, models and frameworks is necessary. As a step towards the establishment of an accurate framework for benchmarking the energy consumption of mining operations, comminution, one of the more energy intensive mining sub-processes, was selected as the focus of a new energy benchmarking method. The methodologies and processes used to onset fragmentation in a rock mass to the crushing or grinding of rock to a size that is suitable for economic extraction are predominantly dependent on the ore properties. For this reason, site specific factors, such as the size reduction requirements and ore hardness were included in an energy benchmark model so that the energy performance of comminution processes at different operations could be normalized and compared on a global scale. Single particle compression breakage has been identified by several prominent researchers in the field of comminution as the upper limit of energy utilization in size reduction processes. For this reason single particle compression breakage was proposed as the basis of a work index for ores to be used as an input for the benchmark model. Overall, the new method showed that the energy benchmark for mineral comminution can be normalized and subsequently compared for a range of mining operations once the ore hardness has been appropriately characterized.

Aug 1, 2013

By Roy O. Ifill, W. Charle Cooper, Alan H. Clark

"Queen’s UniversityIn representative, run-of-mine samples from the Panel and New Quirke mines of the Elliot Lake uranium district, Ontario, brannerite, ametamict uranous titanate, occurs in microscopic grain-aggregates which display wide variations, both in textural habit and in the relative proportions of brannerite, titania polymorphs and/or “uraniferous titania”. As is widely documented, much of the brannerite occurs as trellis-like arrays of laths and needles pseudomorphing rutile oranatase. In some cases, the laths and needles are cemented by coffinite. Brannerite, with an average composition of (U0.629Th0.039Ca0.20)(Ti2.199Fe0.13)O69contains =3 wt% Th and is distinctly Ti-rich relative to the ideal composition, UTi2O6. Much of the Si reporting consistently in electron microprobe analyses (between 1 wt%and 5 wt%) is tentatively attributed to contamination by the quartz-sericite matrix, and a minor proportion is attributed to the presence of coffinite intergrowths. “Uraniferous titania” (average par-tial composition by weight: 10.8% U, 0.3% Th,32.9% Ti) is considered to represent an intermediate stage in the conversion of titania to brannerite.The leaching behaviour of brannerite in its different modes of occurrence was studied qualitatively by the rotating-disc (polished section) method in H2SO4and HCl solutions at 25° to70°C. Redox potential was not measured. The progress of the dissolution of the individual brannerite crystallites and grain-aggregates was monitored by optical and scanning electron microscopes. Regardless of its morphology and texture (laths or needles, reticulate or blocky),brannerite is not readily leachable. Regardless of the process conditions, the rate-controlling step appears to be the initial leach-pit formation. These pits expand radially through the aggregate as leaching proceeds. This dissolution mode is independent of the relative amounts of brannerite and titania, i.e. the type of grain-aggregate. Secondary coffinite intergrowths are readily leachable and enhance the overall leaching kinetics of brannerite by accelerating leach-pit formation. These data provide, for the first time, information on the controls on the leaching of the uraniferous phases which dominate the lower-grade ores in the Elliot Lake district."

Jan 1, 1996

By Jan W. Merks

The objective of metrology in mineral exploration is to derive unbiased confidence limits for masses of metals contained in reserves and resources. Sets of test results for gold and silver determined in pairs of interleaved bulk samples give unbiased confidence limits for the central value of each metal. The same sets of test results show how to verify spatial dependence by applying Fisher?s F-test to the variance of a set and the first variance term of the ordered set. A significant degree of spatial dependence gives a higher degree of precision for the central values of gold and silver. Plotting variance terms of an ordered set against the variance of the set and its lower limits of asymmetric 99% and 95% confidence ranges gives a sampling variogram. Sampling variograms put in plain view where spatial dependence in sampling units or sample spaces dissipates into randomness. The application of statistical methods to test results for sets of ordered core samples within boreholes, and for sets of ordered ore volumes within profiles, plays an important role in defining reserves and resources.

Oct 1, 2009

By H. Thalenhorst

For an advanced mineral exploration project that has a ?nugget? issue making the assessment of the deposit grade difficult, a bulk sampling program using a sample tower can provide information that is otherwise not obtainable prior to actual production. This information includes the verification of predicted resource grades for the bulk sample tonnage (and, indirectly, of the methods and parameters used for the grade estimate), confirmation of the physical and grade continuity of the mineralization, and geotechnical data for a future mining operation. The information gathered will improve the input of a feasibility study, particularly with respect to expected mining grades, which are the key economic parameter for most mining projects. The bulk sample rejects can be used for pilotscale metallurgical testwork. Bulk sampling is also useful in cases of potentially biased drill-hole assay results and in the reliable determination of ore grade before shipment of ore from production or development operations to an external beneficiation plant. This paper presents a practical bulk sampling and sample preparation protocol, lists design features for the sample tower, examines the sampling errors that will be encountered, provides information on quality control procedures, and looks at the costs and benefits of such a program, based on the experience of undertaking bulk sampling programs during the period of 1985 to 2008.

May 1, 2010

RÉSUMÉ L’estimation la plus précise possible des coûts de minage est d’extrême importance pour les études de préfaisabilité afin de diminuer les risques financiers des projets miniers. Les méthodologies existantes d’évaluation des coûts de minage ne permettent pas de les évaluer avec des marges d’erreur acceptables tout au long de l’évolution des projets miniers. Elles visent à évaluer les coûts unitaires de minage des études préliminaires de faisabilité, pour les principales méthodes d’exploitation, en prenant pour l’hypothèse que ces coûts dépendent principalement du taux d’extraction journalière. Si cette hypothèse peut être justifiée pour l’évaluation du coût d’opération, elle est fort discutable pour l’évaluation des coûts de minage. Pour démontrer que cette hypothèse est inadéquate nous avons analysé la méthode d’exploitation la plus populaire sur le continent nord-américain c’est à dire celle par chambres vides avec retrait horizontal (Hr – Horizontal retreat), ou « Longhole » et celle avec retrait vertical (VCr - Vertical Crater retreat). Il ressort de cette analyse que les coûts unitaires de minage dépendent principalement des conditions géotechniques du massif rocheux et de la largeur du gisement et non pas du taux d’extraction journalière du minerai.

Jan 1, 2010

By Z. B. Karshigina, Ye. G. Bochevskaya, E. A. Sargelova, Z. S. Abisheva

At the present time of depletion of reserves of high grade rare earth deposits, it has become necessary to involve low grade raw materials in the production. Among the rare-earth deposits in Kazakhstan, the most advanced one is Kundybai. The mineralogical composition of ore is represented by the minerals: a-quartz SiO2, kaolinite Al2(Si2O5)(OH)4, muscovite KAl2(AlSi3)O10(OH, F)2, albite Na(AlSi3O8), rutile TiO2, and hematite Fe2O3. It contains: 59.1–68.5% SiO2, 19.1–28.4% Al2O3, 7.2–10.3% Fe2O3, 0.51–0.84% TiO2, 0.10–2.32% CaO, 0.39–2.10% MgO, and some other elements. Rare earths are located in the clay part (kaolinite, goethite) in the ion-adsorbing form, and they are represented by the main ore mineral churchite - YPO4·2H2O. In the ore the amount of rare-earth elements (total REEs) varies from 210 to 600 ppm. Elements of the cerium group predominate (La, Nd, Ce, Sm, Pr) and the yttrium group is represented by Y, Gd and Dy. Extremely thin intergrowth and accumulation of REEs in clay minerals cause difficulties in their recovery using flotation and gravitation beneficiating methods. The purpose of this work is the development of an extraction method for the main mineral components of ore with transfer of the total REEs into the solution with high extractive efficiency. Results of the experimental work into high-temperature processing (hereinafter referred to as sulphatization) of ore by sulphuric acid with subsequent dual-mode water leaching of the sulphate product (sinter) are presented in this manuscript. The influence on the ore sulphatization process of some physical and chemical parameters: liquid-to-solid (L/S) ratio (0.36–0.86 cm3/g), temperature (100 – 500°?), concentration of sulphuric acid (3.0–18.62 mol/L) and duration (0.5–2.5 h) was studied. Optimal technological parameters of the process were determined to be: L/S ratio 0.36–0.46 cm3/g; temperature 500°?; concentration of H2SO4 9.0 mol/L; soaking time 1.5 h. Under these conditions it is possible to obtain extractions of ~ 85% REE, as well as 58% aluminium and 10% iron. XRD and IR-spectroscopy analyses were used to determine the material constitution of sulphated cakes obtained under optimal conditions. It was found that the primary minerals are quartz and muscovite; however different sodium aluminium silicates and albite minerals are present as well. In the future, it is proposed that the leach residue cakes are processed into precipitated silica (“white soot”).

Jan 1, 2016

By Champoux. G., E. Ghali, W. Zhang, G. Monteith, A. -M. Lafront, G. Houlachi

The efficiency of the electrowinning process of zinc depends on the performance of oxygen evolution reaction, anode quality and corrosion resistance of lead anodes. Original surfaces of rolled commercial lead-silver anodes were examined using conventional polarisation methods in acid zinc sulfate electrolyte with manganese sulphate at 38 °C. The self-discharge method (potential decay) following 5 h polarisation at 50 mA cm-2 was considered to evaluate electrochemical activity, oxygen overpotential and anode stability. The influence of silver content is compared to that of the microstructure, the fabrication process and the composition of the corrosion products. Polished anode samples have also been examined for microstructural and corrosion performance studies. The experiments show that the duration of the plateau at 1600 mV versus SHE, the potential level (active corrosion state/stationary state) and the rate AEco/At during the decay could be a good indicator to compare the corrosion rates of the anodes in industrial zinc electrolytes.

Jan 1, 2008

By P. M. Daignault

"Fording Coal Limited is one of Canada's largest producers of metallurgical coal for the export market. Medium- to highvolatile coals are mined at its Fording River operations in southeastern British Columbia. This paper discusses some of the mine design and production related problems encountered, and how they have been resolved. Other aspects of the operation such as geography, geology and standard mining methods employed are also covered.IntroductionFording Coal Limited was formed in 1968 to mine and market metallurgical coal resources held by the company in the East Kootenay region of southeastern British Columbia<1>. The company, at its Fording River operations, holds approximately 19,635 hectares of coal lands, having an estimated resource of 3.5 billion tonnes of metallurgical coal. Total reserves are approximately 520 million tonnes.Fording Coal is owned 600Jo by Canadian Pacific Enterprises Limited and 400Jo by Cominco Limited. Cominco itself is owned approximately 540Jo by Canadian Pacific Enterprises. The company, originally staffed and managed by Cominco, has been independent since January 1978, with its own board of directors, management, staff and employees. The head office is in Calgary, Alberta. The Fording River operations are located in British Columbia, approximately 70 km north of the Crowsnest Pass, near the Alberta/ British Columbia boundary, only 150 km by air and 320 km by road from Calgary.A contract was signed in 1969 with the Japanese steel mills to provide 45. 7 million tonnes of metallurgical coal over fifteen years, commencing in 1972, at an annual rate of 3 .05 million tonnes per year. The mine is presently expanding annual production capacity from the original 3.0 million long tons (3.05 million tonnes) to in excess of 5.0 million tonnes.Concurrent with this production expansion, a vigorous marketing effort has diversified Fording's customers to include additional countries in the Pacific Rim area. Three cleaned coal products are now marketed compared to the single ""standard blend"" product originally shipped to satisfy the Japanese contract. The product specifications are listed in Table I."

Jan 1, 1985

By Jason Mewis

Do you have it figured out? ? If so, I want to network with you! ? There?s a lot to learn as an industry The more we share the better we?ll all become at dealing with this subject If you don?t have it figured out ? You?re among the majority in the in0dustry ? Project Risk is a broad and complex issue that is dealt with through many processes ? And it?s not intuitive either ? And you don?t want to confuse one process with Another

Nov 1, 2011

By D. J. Wesolowski, D. A. Palmer, P. Bénézeth, S. Hilic

The equilibrium solubilities of gibbsite, Al(OH)3, and boehmite, A1OOH, have been extensively studied in our laboratory from 30 to 90°C, and 100 to 300°C, respectively, to ionic strengths of 5 mol.kg-1 (NaC1) over a wide range of pH. The resulting solubility constants have been treated with consistent empirical models that incorporate appropriate experimental data taken from the literature. These studies were initiated in part to understand and quantify scale formation in geothermal fields as a result of fluid re-injection from electric power generating plants. Most of the experimental results were obtained through application of a unique potentiometric method that allows for rapid, precise measurement of the in situ pH and allows the approach to equilibrium to be monitored from under- and super-saturation. The experimental program was extended to deal with more basic solutions (e.g., five molal NaOH), in which the dissolution of boehmite was measured for application to the refining of bauxite by the Bayer process. More recent results in strong caustic solutions at temperatures below 90°C were obtained to investigate the dissolution of aluminum-containing sludge from holding tanks at the Savannah River waste reprocessing site. In this case the solubility enhancing effects (kinetic and equilibrium) of triethanolamine were also established and quantified. Finally, potentiometric solubility experiments were conducted (50°C in 0.l mol-kg-1, NaCl) on aluminum alloy sheets to identify mainly the nature of the aluminum oxide/hydroxide surface layer formed under these conditions.

Jan 1, 2004

By J. Mrozowski, R. Kurowski, A. Chmielarz, Z. Szolomicki

PbSO4 is a commonly encountered component of the materials used in lead recycling (e.g. battery paste) and in by-products of metallurgical processes (e.g. residues from zinc hydrometallurgy, dust from copper smelting). Smelting of battery paste results in SO2 emission and in generation of significant amounts of hazardous ‘white slag’. Those adverse phenomena can be eliminated by removal of sulphur from the paste. Lead-bearing by-product materials are of polymetallic character with usually low Pb content, which makes lead recovery technologically difficult and uneconomic. However it can be successfully achieved by extraction of lead sulphate from solid phase with triethylenetetramine (TETA) water solutions. That compound has capability for selective solvation of a whole lead sulphate molecule. Virtually pure lead carbonate can be precipitated by means of saturation of lead-loaded aqueous TETA solution with carbon dioxide, and complete regeneration of the extractant is achieved by fixing of sulphate ion into gypsum through operation of calcium hydroxide suspension. The paper presents results of the lab and pilot scale studies into process of battery paste desulphurisation and recovery of lead from the residue of zinc calcine leaching.

Jan 1, 2014

By J. E. Quinn, S. Wolstencrof, K. H. Soldenhoff

Ion Exchange (IX) using Strong Base Anion Exchange (SBA) resins is conventionally used to purify uranium from sulfuric acid pregnant leach solutions (PLS), particularly when uranium grades are relatively low. However, the presence of chloride in the PLS interferes negatively with conventional SBA based IX processes, by reducing the uranium loading on the resin. Chloride can be present in the PLS due to the use of saline bore or seawater in leaching, or chloride containing leach/IX reagents. ANSTO has been examining strategies to combat this problem for several years. ANSTO previously reported that a recently developed high-capacity SBA resin was capable of extracting uranium efficiently from saline leach liquors produced at Boss Resources’ Honeymoon in-situ recovery mine in South Australia. The resin was tested successfully in a 2017 Field Leach Trial carried out onsite in the existing wellfield area. While the piloting was extremely promising, the elution was more challenging, with eluate tenors limited to 2 g/L U3O8 using NaCl. This is due to the resin’s strong affinity for the uranium complex, which makes it difficult to elute efficiently by standard methods. In this paper, we describe a novel elution process developed for strong base resins that significantly improves upon conventional methods. The process, developed by ANSTO, is based upon conversion of the extracted uranyl sulfate species to uranyl carbonate, followed by elution with sodium chloride. The loading and elution process was demonstrated in continuous locked-cycle piloting conducted at ANSTO in 2019. The results were excellent, with eluate tenors of 8–9 g/L U3O8 achieved from Honeymoon PLS containing 50 mg/L U3O8, with very little transfer of impurities. In subsequent test work, the eluate was shown to be amenable to uranium peroxide precipitation. The results represent a huge step forward for the purification of uranium from saline leach liquors.

Jan 1, 2020