EARLY STAGE METALLURGICAL EVALUATION OF A COMPLEX LOW-GRADE URANIUM PROJECT By Will. Goodall & Julian. Perkins Aura Energy Ltd, Australia Presenter and Corresponding Author Will Goodall
[email protected]
ABSTRACT The development of complex low-grade mineral deposits is becoming an increasingly important target for resource companies. These deposits often present marginal economic opportunities, which are highly dependent on the development of low-cost and efficient processing strategies and the utilisation of by-products to become a viable development target. For exploration-focused companies this presents a critical investment decision stage early in project development on whether to pursue potentially expensive resource definition or look for other targets. The development of the Häggån project, Sweden, has presented a target requiring careful technical evaluation through the early development stages. The project represents a low-grade complex uranium resource, with significant vanadium, molybdenum, nickel and zinc credits. A process of mineralogical characterisation and multi-directional scoping metallurgical test work has been presented, focused on providing the basis to make a confident investment decision to progress the project to pre-feasibility evaluation. Implementation of this process has allowed the development of conceptual flowsheets targeted at efficient recovery of value metals and the rejection of technically ineffective options based on an understanding of the material and its behaviour.
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INTRODUCTION The evaluation of metallurgical process options for complex low-grade ore deposits is a critical activity in the earliest stages of project development. The consumption of readily accessible, simple mineral deposits with time has led to a greater proportion of new project opportunities being discovered in low-grade, high volume mineralized areas. This shift places greater emphasis on the importance of developing an efficient and cost driven processing option at a very early stage to avoid wasting funds and resources on projects that are unlikely to become viable. To undertake metallurgical process development for complex low-grade deposits, an important first step is to gain a fundamental understanding of the mineralogy and then to use that knowledge through the whole development process to constantly direct and optimize the metallurgical testwork program. By taking this approach the technical risk associated with new project development can be reduced, through systematic evaluation of all likely options and by providing valid reasons for the rejection of poor options. The project examined in this study represented an example of a high volume, low-grade uranium deposit that had previously been considered uneconomic. The study will demonstrate how through first understanding the mineralogical and physical nature of the deposit and then systematically evaluating possible process options, a viable process could be identified. This in turn would provide sufficient confidence to progress the development from the scoping stage to give adequate justification for the funding of a pre-feasibility study.
THE SWEDISH ALUM SHALE The Lower Palaeozoic (Middle to Late Cambrian age) shale hosting the mineralisation in the Häggån project is well developed through Scandinavia. The characteristic formation of this shale facies extends from southern Sweden (Skåne) to northernmost Norway (Finnmark). The unit represents a long period of very slow marine deposition and in general is about 20m thick, however this expands to up to nearly 100m in the southernmost regions (Andersson 1985). The term alum shale refers to parts of the black shale formation from which the alum salt KAl(SO4)2.12H2O was extracted over 300 years ago. The name Alum Shale Formation is now used for the entire lithostratigraphic unit throughout Scandinavia. The formation is dominated by black shales with an organic carbon content up to 20% w/w. The formation includes some grey shales and silts, particularly in the lower depths; limestone lenses and discontinuous carbonate beds; and iron sulphide occurring as nodules or thin bands disseminated throughout the formation. The Alum Shale in Sweden and Estonia was mined for uranium through the middle of the 20th century. These operations formed strategic resources for Russian and European nuclear programmes and cost considerations were not a major factor in the decision to exploit them. By today’s economic standards, these operations would be uneconomic and to this date further development has not been pursued.
THE HÄGGÅN PROJECT Exploration drilling of the Swedish Alum Shale in the Storsjön district for uranium mineralisation was originally undertaken in the 1970’s by the Swedish Geological Survey (Andersson et al 1985). The work completed then identified low-grade uranium mineralisation but at the time recovery of uranium from the shale of this region was considered uneconomic and no further work was undertaken.
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Figure 1: Location of Häggån Project within the Swedish Alum Shale Belt, Storsjön District, Central Sweden Early in the new century, as metal prices began to rise, interest in the exploration potential of the Scandinavian shale by private companies began to grow. In 2001, a Canadian company began reexamining the areas originally drilled to develop on the mineralisation identified. In 2006 Aura Energy Ltd (ASX ticker: AEE) obtained permits for exploration of the Storsjön area after reinterpretation of the geological model suggested that the mineralisation might be more extensive than indicated by historic drilling. The initial drilling program met with immediate success and in 2010 an inferred JORC compliant resource of 1.79B tonnes at an average grade of 160ppm U3O8 (cut-off grade of 100ppm uranium), for a total of 631 Mlbs of contained U3O8 was compiled. The mineralisation defined for the Häggån Project occurs in a continuous flat lying sheet ranging in thickness from 20m to 250m. The mineralized Alum Shale outcrops in a number of areas in the district, including several areas in the Project Area. Elsewhere it is overlain by limestone with thickness up to 100m around the resources areas. The mineralized shale is characterized by repeating thrust sheets, which has locally caused the unusually large thicknesses. The geology and geochemistry of the deposit suggested that the Häggån project was a uranium prospect without direct analogies to operating processes elsewhere in the world.
A COMPLEX PROBLEM The approach of the authors was to evaluate the prospect without pre-conceived notions of an appropriate process, basing the development on gaining a deep fundamental understanding of the mineralogy and relationship to the project. This philosophy allowed for systematic evaluation of all opportunities. The drilling program completed for the Häggån resource included an extensive assay protocol, in which 34 elements were routinely analysed at 2m vertical intervals. This provided an excellent geochemical database for further evaluation of the deposit. 261
The geochemistry identified that significant potential by-product metals, including vanadium, molybdenum, nickel and zinc were present and should be considered in process development. The primary metallurgical issues apparent from examination of the drill core and the geochemistry included: • • • • • • •
The mineralisation, while extensive, was of low grade and would require a low-cost process to be viable; The deposit contained extensive and numerous calcite veins ranging in size from a few milliimetres to several centimetres, presenting a potential issue for acid leaching; There were a number of limestone lenses up to several metres thick forming internal waste pods within the deposit, which was also overlain by limestone; The mineralized shale material was dominated by very fine grain sized minerals. The presence of carbonaceous material might have a negative effect on processing. Pyrite was visible throughout the mineralisation, presenting a potential issue for carbonate leaching. A wide range of value metals were noted, presenting complexity in defining which should be targeted to add the most value.
Taking these potential issues into consideration it was apparent that the Häggån Project presented a complex opportunity, requiring a fundamental understanding of the key parameters to potentially result in a positive outcome.
PRELIMINARY MINERALOGY The logging of the deposit lithology through the drilling program gave the first indication of mineralogical distribution. The material appeared remarkably homogeneous with consistent textures across the whole deposit. There were two major textures, an organic/mica matrix and quartz with associated calcite. These textures repeated down to very fine sizes. To progress with development of a viable processing flowsheet for the Häggån project, a fundamental understanding of the composition of the Alum Shale was required. A preliminary mineralogical programme was implemented using optical microscopy, coupled with scanning electron microscopy (“SEM”) and laser ablation inductively-coupled plasma mass spectrometry (“ICP-MS”) on several drill core fragments from the initial drilling program. This scoping work gave the first indication that the mineralised shale matrix, consisting predominantly of organic C, pyrite and muscovite was host to the uranium values, along with nickel and molybdenum as potential byproducts. Vanadium was shown to be more broadly distributed and associated with muscovite. It 3+ was speculated that the source of the vanadium might be roscoelite (K(V , Al, Mg)2AlSi 3O10(OH)2), in which a vanadium atom has replaced an aluminium atom in the muscovite crystal structure. The relationship of quartz-calcite veining to the organic/mica matrix can be seen in the photomicrograph presented in Figure 2. This confirmed the proposition from visual logging of the drill core that the organic/mica matrix was interspersed with small-scale quartz-calcite veins.
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Figure 2: Photomicrograph of Pyrite Grain in Organic/mica Matrix, with Association to Quartz-calcite Vein. Image Courtesy of CODES, Tasmania The populations of pyrite were also qualitatively defined with two preliminary populations distributed across the two textural types highlighted in Figure 3.
Figure 3: Photomicrograph of Pyrite Textures and their Associations with the Two Textural Types. Image courtesy of CODES, Tasmania Examination of the two textural types and of individual pyrite grains gave the first clues to the distribution of potentially valuable metals in the Häggån deposit. LA ICP-MS was utilised for trace
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level determination of elemental distribution across the phases identified. An example of the elemental maps for an area of pyrite, also showing the two textural types, is shown in Figure 4.
Figure 4: Trace Elemental Distribution in Pyrite and Shale Textural Types Mapped by LA ICPMS. Image Courtesy of CODES, Tasmania In total 74 spot analyses using LA ICP-MS were taken across a range of phases identified. The key element concentrations are summarised in Table 1. It can be seen that of the potentially valuable products, uranium and vanadium were more closely associated with the organic/mica matrix, while nickel was concentrated within the pyrite. The molybdenum distribution was variable, although the greater fraction appeared to be associated with the organic/mica matrix. It was noted that there was very little association of these elements with the quartz zones. Table 1: Elemental Distribution of Key Elements between Organic/mica Matrix and Pyrite Grains. Spot Analysis undertaken by LA ICP-MS on a Total of 74 spots
The initial mineralogical findings, while not considered representative because they were conducted on spot samples of core, suggested that the valuable metals in the Häggån material were closely related with the organic/mica matrix and not with the quartz. However, it also appeared that pyritic nature of the shale material and the presence of numerous calcite veins may be a key to the process, as both presented challenges in terms of reagent consumption for conventional leaching options. There was still no clear direction on whether an acidic or alkaline leach would be more appropriate for the uranium. There was also no understanding of the nature of the uranium mineralogy.
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LEACHING EVALUATION: PHASE 1 – PRELIMINARY INVESTIGATION The preliminary evaluation of leaching options was initially focused on uranium recovery. This decision was made to target what was potentially the most valuable metal present. However, analysis of the distribution of other metals was included at all stages to allow for the potential recovery of these to be evaluated on the basis of a successful outcome. As there was no clear indication of the preferable direction for development of a leaching process, a program to evaluate leaching by both acidic and alkaline leaching was initiated. The results of preliminary leaching testwork can be seen in Table 2. Table 2: Summary of Leaching Results by Acid and Alkaline Media. Conventional Acid Leach Conditions (P80 of 75µ µm; 65°C; 20 g/L H2SO4; 2% w/w solids); Conventional Carbonate Leach Conditions (P80 of 75µ µm; 90°C; 0.3M Na2CO3; 0.1M NaHCO3; 40% w/w solids) Method
Extraction U (%)
Ni (%)
Mo (%)
Consumption V (%)
H2SO4
Total carbonate
kg/t ore
kg/t ore
Conventional acid leach
93.9%
22.7%
12.2%
4.7%
87
Conventional carbonate leach
76.9%
0%
29.8%
0%
-
9.4
Preliminary evaluation of basic leaching options for the Häggån material gave mixed results. Conventional acid leaching gave excellent extraction of uranium but with high acid consumption of 80-100 kg H2SO4/tonne ore. Alternatively, conventional carbonate leaching returned on average about 10-15% lower extraction than the acid leaching values. The recovery of by-product elements showed that although alkaline leaching was more effective for recovery of molybdenum, no nickel was recovered. Both leaching options returned negligible vanadium recovery, consistent with the proposition from mineralogical analysis that the vanadium was hosted within the muscovite crystal structure. The results of initial leaching showed that uranium could be readily recovered by acid leaching but that the high acid consumption supported the initial assumption from the 1970’s that the Swedish Alum Shale would be a high cost source of uranium.
A FUNDAMENTAL UNDERSTANDING OF MINERALOGY The apparent complexity of uranium recovery by conventional leaching methods highlighted the need to establish a deeper fundamental understanding of the mineralogy at the Häggån project. This was approached with two key goals in mind. The first was to define the uranium minerals present and their deportment. In addition, the gangue mineralogy was highlighted as a major driver for the leaching options and a detailed evaluation of the distribution of key reagent consuming minerals was included. The QEMSCAN system was utilised to examine the uranium deportment, modal mineralogy, mineral distribution and key mineral associations. The organic/mica matrix presented a number of analysis complexities, as the conventional QEMSCAN sample preparation process, in which particles are mounted in an epoxy resin, would result in an inability to differentiate the carbon from the background. To overcome this, a method of mounting the particles in carnauba wax blocks, based on methods utilised in the coal industry (Butcher et. al. 2003), was developed. QEMSCAN analysis located only a small number of discrete grains of a U-Ti-Si mineral, which only accounted for a fraction of the assayed uranium concentration. It was established that the major population of uranium might be present below the detection limits of QEMSCAN and a more suitable technique would be required for determination of the uranium deportment.
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Background Brannerite Mo?? Gangue
Figure 5: False Colour Mineral Map Showing U-Ti-Si Phase within a Häggån Sample Determined by QEMSCAN PMA Analysis. Pixel spacing of 2µ µm In conjunction with the QEMSCAN analysis, a program to characterise uranium deportment utilising field electron gun (“FEG”) electron probe microanalysis (“EPMA”) with the CHIMAGE package using the techniques of Harrowfield et al (1993) was implemented. The use of QEMSCAN gave a statistically valid indication of the gangue mineral distribution but was unable to provide sufficient chemical resolution to enable detection of uranium associated with the organic/mica matrix. This programme focused on identification of uranium minerals and associations within the organic/mica matrix through high resolution wavelength dispersive spectral (WDS) X-Ray mapping. Figure 6 shows an example of the phase-patched map for one of the samples examined. The presence of rare grains of a U-Ti-Si phase was confirmed but, as was indicated by the QEMSCAN results, these were minor and not attributable to the major uranium population. It was suggested that it was most likely that the uranium could be present adsorbed onto the carbonaceous matter at extremely fine size as had been observed by numerous observers from other deposits (Vine 1962; Breger 1974; Landais and Gize 1997). Another suggestion was that some of the uranium could be adsorbed onto the sulphide surfaces, as described by Wersin (1994)
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Figure 6: Phase-patched Mineral Map of Organic/mica matrix Defined by CHIMAGE using FEG EPMA. Red Boxes indicate Location of U-Ti-Si Phases Identified The findings of the FEG-EPMA study, combined with inferred findings from LA-ICP MS and QEMSCAN analysis, showed that the major population of uranium was most likely present as adsorbed species on the surfaces of organic carbon, and possibly also on pyrite grains. A population of discrete grains of a possibly unknown U-Ti-Si phase was identified as a minor, possibly refractory population. Comparison to the leaching recovery results suggested that this was not a major uranium host and further quantification of the mineral composition was not pursued. The cause for the high acid consumption was attributed to the presence of the calcite visually identified in the material, however there was a need to better understand the nature of the calcite distribution to seek an opportunity to reduce its influence on the process. QEMSCAN mineralogical characterisation confirmed the modal mineralogy and the results are shown in Figure 7. This confirmed that two distinct mineralogical domains were present. The organic/mica matrix was dominated by organic carbon, pyrite and muscovite, while the quartzcalcite veins identified from lithological observation were dominated by quartz, calcite and muscovite. The uranium, molybdenum and nickel values were all shown to be concentrated in the organic/mica matrix. This suggested on a micro-level that calcite was not directly associated with the uranium value and potential existed for its removal.
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Figure 7: Modal Mineralogy of Häggån Sample Determined by QEMSCAN PMA Analysis. Pixel Spacing of 2µ µm Calcite (CaCO3) was the only carbonate mineral observed to be present. Figure 8 shows that at a particle size of 80% finer than 2mm the calcite was predominantly liberated, suggesting that for direct acid leaching all the calcite would be available to consume acid. Interestingly, the volume of calcite present was only sufficient to account for a theoretical maximum potential acid consumption of 44 kg H2SO4/tonne ore. The acid consumption recorded in the initial testwork was about double this, suggesting that slow acid consuming minerals such as feldspars and other silicate minerals might have an effect in acid leaching under the conditions tested.
Figure 8: Liberation of Calcite by Screen Fraction, Determined by QEMSCAN, Normalised by Fraction to 100%. Liberation Classes are based on the Exposed Surface Area of Calcite to Background
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LEACHING EVALUATION: PHASE 2 – DEALING WITH CALCITE The outcomes of preliminary diagnostic leaching testwork and detailed mineralogical characterisation were that the presence of calcite (CaCO3) and slow acid consuming minerals would make the use of an acid leaching process uneconomic unless their effects could be mitigated in some manner. Using the fundamental knowledge gained through mineralogical characterisation two initial options were explored. • •
Eliminate calcite as a problem by revisiting alkaline leaching Remove calcite from the equation by differential flotation
A further programme of bench-scale testwork investigating alkaline leaching was undertaken. This work applied a range of conventional alkaline leaching conditions but failed to substantially add to the recovery of uranium. Recovery improvements were achieved through use of pressure carbonate leaching, however, reagent consumption was excessive due to the presence of pyrite and the anticipated operating cost would be prohibitive. Direct physical removal of the liberated calcite grains by flotation was evaluated by a series of bench-scale tests. The calcite particles showed insignificant response to direct flotation under a wide variety of collectors and conditions, even though the liberation and particle size distribution suggested this might be a possible solution. This led to the investigation of bulk sulphide and organic carbon flotation, with rejection of calcite to the flotation tail. Results of this work showed that calcite could be effectively rejected to a tailing, however losses of uranium were prohibitively high. The outcomes of attempts to control the influence of acid consuming minerals had been unsuccessful and it appeared that the historic findings that the Häggån material represented a high cost source of uranium were correct. The preliminary metallurgical testwork had shown that acid leaching was effective but with a high acid cost, while alkaline leaching was less efficient at recovery of uranium. The fundamental mineralogical analysis program had identified the reason for high acid consumption was the presence of significant liberated calcite but all attempts to minimise the influence of this had been unsuccessful.
A BETTER UNDERSTANDING The primary issue with the use of acid leaching was the high cost associated with acid addition and the pyrite represented a possible source of acid within the ore itself. Rather than look at options to remove the source of acid consumption the focus was changed to investigate methods to utilise the contained acid value of the material to mitigate the cost impact. Re-examination of the geochemical data showed that on average the material should be net acid generating, providing an excellent basis to pursue this option. Generation of sulphuric acid requires a means of oxidation of the pyrite. A number of options were available for this. Flotation of a pyrite concentrate with subsequent oxidative roasting was considered. This was estimated to be too expensive, given the probable high capital cost, the low grade of material and the high throughputs required. The remaining option was to try bacterial leaching to catalyse the oxidation of pyrite as described by Lowson (1975). The bacterially catalysed oxidation of pyrite proceeds by the mechanism described in Figure 9. The process is well suited to uranium leaching as the oxidation of pyrite produces ferric iron, which acts as the oxidant in the uranium dissolution reaction. Precipitation of ferric iron to ferrous sulphate produces the sulphuric acid required for the uranium leaching reaction. In the presence of appropriate bacteria the process is autocatalytic, allowing for sustained leaching.
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Figure 9: Bacterial Oxidation of Pyrite (from: Stumm and Morgan 1981) Examination of operations that treated similar material to the Black Shale at Häggån identified the Talvivaara polymetallic operation in Finland. This project effectively utilises a bacterial heap leaching process to recover Ni, Cu and Zn from material similar to the Häggån material but with pyrrhotite as the dominant sulphide mineral, rather than pyrite (Schippers et al 2007). The successful implementation at the Talvivaara operation, demonstrated that bacterial leaching could be a technically and commercially viable option. A preliminary scoping program for bacterial leaching using shaking bottles on 3g samples of material showed positive uranium recovery and strong oxidation of pyrite. The results of this program are summarized in Table 3. Table 3: Bacterial Shaking Flask Leach Test Results. Leaching Conditions (P80 of 2mm; 45°C; 3g Ore; Starting pH 1.8; Leach Time 42 days; Biotic) Method
Bacterial shake flask leach
Extraction U (%)
Ni (%)
Mo (%)
V (%)
83%
62%
37%
0.6%
Following the success of the shaking flask tests a bacterial agitated tank leach program was undertaken. The results are summarised in Table 4. Even without optimisation of the process to promote ferric iron hydrolysis and generation of acid, the acid consumption was reduced from 87 kg H2SO4/tonne ore to 60 kg H2SO4/tonne ore. This program was then further developed to investigate the effects of promoting ferric precipitation for sulphuric acid generation and it was shown that the process could be made acid generating. Table 4: Summary of Bacterial Tank Leaching Results. Conditions with no Fe hydrolysis (P80 of 2mm; 35°C; 5% w/w Solids; Starting pH 1.6; 44 days; biotic); with Fe hydrolysis (P80 of 2mm; 35°C; 5% w/w Solids; Starting pH 2.5; 44 days; Biotic) Method
Extraction
Consumption
U (%)
Ni (%)
Mo (%)
V (%)
H2SO4 (kg/t)
Bacterial tank leach (no Fe hydrolysis)
87%
56%
45%
2.7%
60
Bacterial tank leach (Fe hydrolysis)
83%
45%
18%
-
-10.7
Preliminary bottle roll acid leach tests at large particle size had indicated that heap leaching may be a possible option and a program of small-scale bacterial column leach tests was undertaken. The bacterial column leach tests were undertaken using 40cm high columns with a diameter of 100mm. Columns of this size were utilised to efficiently define the amenability of the material to bacterial heap leaching, as a precursor to the use of larger columns for parameter definition.
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The results of the 40cm column bacterial leaching are shown in Table 5 and indicate that almost comparable uranium recovery to bacterial tank leaching could be achieved, even with an increase in the top crush size from 2mm to 25mm. This unexpected and exciting result demonstrated the potential to utilise a coarse-crushed low-cost bacterial heap leach operation for the Häggån material. Importantly, while the recovery was maintained the acid consumption was drastically reduced, with the total consumption was down to 0.4 kg H2SO4/tonne ore. It is thought that with further work, significantly better conditions for acid generation might be possible. Table 5: Summary of Bacterial Column Leach Results. Test Conditions (P80 of 25mm; 50°C; -1 Starting pH 1.6; 102 days; Column 400mm H x 100mm D; Irrigation 1.2ml min; biotic) Method
Bacterial column leach
Extraction
Consumption
U (%)
Ni (%)
Mo (%)
V (%)
H2SO4 (kg/t)
85%
66%
22%
0.7%
0.4
The potential reduction in impact of acid consumption as a major cost factor on the project showed that through fundamental understanding of the Häggån material a viable processing option could be developed. The final step of the scoping investigation was then to optimize the use of the process to maximize the value. This was a major step forward for the Häggån project, demonstrating that through careful and systematic evaluation founded in a thorough fundamental understanding of the mineralogy, a lowcost processing alternative could be developed, transforming the project from a perceived high cost source of uranium. All the preliminary indicators examined to date have shown that the material is ideally suited to bacterial heap leaching, with good permeability and an accessible source of pyrite for bacterial growth.
DEFINING ECONOMIC VIABILITY An economic evaluation of three process flowsheet options was undertaken, based on the testwork results. The purpose was to establish if there was a sufficiently strong case to support the expense of moving on to a prefeasibility study. Bacterial heap leaching or agitation leaching were felt to be the main contenders, taking into account the potential for by-product generation, but conventional acid agitation leaching was included as a base case from which to measure the amount of advantage gained by the other two. All of the flowsheets examined utilised the same mining and downstream processing assumptions to allow direct comparison to be made. The recovery of nickel and molybdenum as by-products to uranium were included in the evaluation but vanadium was excluded at this stage because testwork had indicated that a separate process stage might be required, subsequent to uranium extraction, and much further testwork would be needed. The outcomes of scoping economic evaluation were: 1. Conventional acid leaching for recovery of uranium and co-products returned a significantly negative net present value (NPV) for the project 2. Bacterial agitated leaching for recovery of uranium and co-products returned a negative net present value for the project (NPV) 3. Bacterial heap leaching for recovery of uranium and co-products returned a significantly positive net present value (NPV) for the project. These results confirmed the understanding that conventional acid leaching was not a viable option for the Häggån Project. Significantly, the analysis showed that utilisation of a bacterial heap leaching flowsheet could transform the project into an attractive low-cost uranium project, with significant upside potential from by-product metals.
CONCLUSIONS The program undertaken for the evaluation of the Häggån Project for the recovery of uranium, along with nickel and molybdenum by-products showed that by using a careful systematic approach and by paying close attention to gaining a fundamental understanding of the material characteristics, a viable process option could be identified. The material was known to be low-grade and complex, with conventional leaching options being shown to return a negative value for the project. Through thorough evaluation of the mineralogy and using that knowledge to direct and explain testwork
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results, a previously un-tested leaching option was examined and shown to transform the project both technically and economically. This step forward has opened up a major uranium resource that had previously been considered uneconomic, potentially transforming the approach to recovery of metals from similar deposits.
ACKNOWLEDGEMENTS The authors would like to thank Aura Energy Ltd for permission to present results relating to the Häggån Project. The authors would also like to acknowledge the contributions of Prof Ross Large, The University of Tasmania; Dr Mark Pownceby, Dr Aaron Torpy and Dr Colin McCrae, CSIRO Process Science and Engineering; Dr Bob Ring, Dr John Lawson and Dr M. Ovinis, Dr C.H. Quan, Dr M. Baker and Dr I. Datta, ANSTO Minerals; Dr Keith Quast and Dr Scott Abbott, The Ian Wark Research Institute; and Dr Ralph Hackl, Felicity Perot, Jian Li, D Collinson, D Shiers and Dr Helen Watling, The Parker Cooperative Research Centre for Hydrometallurgy, for their contributions to the testwork and interpretation of results that allowed the project to be a success.
REFERENCES 1. Andersson, Astrid, Dahlman, Bertil, Gee, D.G., and Snäll, Sven, 1985, “The Scandinavian Alum Shales: Sveriges Geologiska Undersoekning, Serie Ca: Avhandlingar och Uppsatser I A4, NR 56, 50p. 2. Butcher A, Gottlieb P, Miller G, French D, Cropp A, Gupta R, Sharma A and Wall T (2003) th “Automated measurement of coal and mineral matter by QEMSCAN” 12 International nd th Conference on Coal Science, 2 -6 November 2003, Cairns, Australia 3. Harrowfield, I.R., MacRae, C.M. & Wilson, N.C. (1993) Chemical imaging in electron microprobes. In Proceedings of the 27th Annual MAS Meeting, Microbeam Analysis Society, New York, 547-548. 4. Vine, J.D. (1962) Geology of uranium in coaly carbonaceous rocks. U.S.G.S. Professional Paper, 356-D, 113-170. 5. Breger, I.A. (1974) The role of organic matter in the accumulation of uranium: the organic geochemistry of the coal-uranium association. Proc. Symposium I.A.E.A. Athens, p. 99-124. 6. Landais, P. & Gize, A.P. (1997) Organic matter in hydrothermal ore deposits, in Barnes, H. L. ed., ”Geochemistry of Hydrothermal Ore Deposits 3rd ed.”, New York, John Wiley and Sons, p. 613-656. 7. Schippers A, Sand W, Glombitza F and Willscher S (2007) “Talvivaara black schist bioheapleaching demonstration plant” Advanced Materials Research (20-21) pp. 30-33 8. Wersin, P., Hochella Jr., M.F., Persson, P., Redden, G., Leckie, J.O. & Harris., D.W. (1994) Interaction between aqueous uranium (VI) and sulfide minerals. Spectroscopic evidence for sorption and reduction. Geochimica et Cosmochimica Acta, 58, 2829-2843. 9. Schippers (2007) 10. Lowsen R (1975) “Bacterial leaching of uranium ores - a review.” Australian Atomic Energy Commission Research Establishment, Lucas Heights 11. Stumm, W. and Morgan, J. (1981). Aquatic Chemistry, First Edition. John Wiley & Sons, New York, NY.
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URANIUM RECOVERY AS A BY-PRODUCT AT TALVIVAARA MINE By 1
1
2
3
Pertti Pekkala, Annika Hämäläinen, Ken Gullen, Erkki Paatero 1
Talvivaara Mining Company Ltd., Finland 2 Cameco Corporation, Canada 3 Outotec Oyj, Finland Presenter and Corresponding Author Pertti Pekkala
[email protected]
ABSTRACT The Talvivaara Mine, located near the City of Sotkamo in Finland, is the first bioheapleaching operation for nickel production in the world. Production of nickel along with the by-products of zinc, cobalt and copper as metal sulphides started in October 2008 at the Talvivaara mine. The ore also contains a low concentration of uranium. Talvivaara and Cameco are working together to design, construct and commission a uranium extraction plant to recover the uranium from their production leach solution. Outotec is the main technology supplier for the solvent extraction and precipitation process areas. The pregnant leach solution (PLS) from the existing metals recovery plant will enter the uranium plant with a uranium concentration of approximately 20 ppm. A novel solvent extraction process will be used to selectively extract the uranium from the PLS into an organic solution. Sodium carbonate will be used to strip the uranium from the organic phase back into an aqueous phase. The yellowcake will be precipitated from the aqueous phase with hydrogen peroxide, settled in a thickener and then dried before being packaged in steel drums. The low uranium concentration in the PLS, the high volume throughput and environmental considerations were the primary factors in designing the processes and selecting the major equipment. The design engineering is complete, the building is under construction, all the major equipment has been ordered and installation of the equipment has started. Commissioning of the new uranium plant will start in early August of 2012. The plant will produce 350 tonnes of uranium per year, when the full nickel production is reached.. The process for obtaining all the permits and approvals to operate the uranium extraction plant from the European Commission, the Government of Finland, the Finland nuclear regulator (STUK Radiation and Nuclear Safety Authority) and the Finnish environmental authorities is underway and is expected to be completed by the end of the second quarter of 2012.
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INTRODUCTION The Talvivaara Sotkamo mining operation is located 460 kilometers Northeast of Helsinki and 350 kilometers South of the Arctic Circle, near the City of Sotkamo in central Finland (see Figure 1).
Figure 1: Location of the Talvivaara Sotkamo Operation The Salivary ore body is one of the largest known nickel sulphide deposits in Europe. There are two large ore deposits associated with the mining operation, Kuusilampi and Kolmisoppi, containing 1,121 M tonnes in measured and indicated resource categories that will provide approximately 46 years of operation at full-scale production. Full-scale production will comprise of 50,000 tonnes of nickel, 90,000 tonnes of zinc, 15,000 tonnes of copper and 1,800 tonnes of cobalt annually. Open-pit mining in ore in the Kuusilampi deposit began in April 2008 with the first metal sulphides produced in the metal extraction plant in October 2008. The average mineral content of the ore in the two deposits is 0.23% nickel, 0.50% zinc, 0.13% copper and 0.02% cobalt. A biological heap leach process was developed to provide an economical method to extract the low-grade metals from the ore. Following the start-up of the metal recovery plant, it was discovered that a very low concentration of natural uranium at 18 to 20 ppm, on average, was being co-extracted with the other metals in the biological heap leach operation. A technical evaluation found that the uranium could be economically extracted from the pregnant leach solution (PLS) and sold commercially, even at this low concentration. Cameco Corporation, with its headquarters in Canada, is one of the world’s largest uranium producers with operations in Canada, the United States and Kazakhstan plus a mine development project in Australia. Cameco and the Talvivaara Mining Company have established a commercial arrangement whereby Cameco will assist Talvivaara to design, construct and commission a uranium extraction plant at the Talvivaara Sotkamo Mine as well as assist with operating procedures, training and regulatory approvals. Uranium production from the plant will be purchased by Cameco under a long-term contact with Talvivaara. For the design of the solvent extraction and precipitation process areas, Outotec is the main technology supplier, which includes process design and supply of the major process equipment. This paper describes the design and operating parameters for the plant as well as provides some background on the environmental benefits, regulatory approval process and the current status of construction.
TALVIVAARA SOTKAM OVERALL DESCRIPTION The Talvivaara Sotkamo Mine is a large, open cut mining operation producing up to 100,000 tonnes of ore per day. The ore is hauled to a central crushing facility where it is crushed to a nominal size of 250 mm. The crushed ore is transferred to three stages of secondary and tertiary crushing and screening to achieve a final particle size of 8 mm. The ore is then agglomerated
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using recycled leach solution to consolidate the fines with the coarser ore particles before it is conveyed to a stacker to place it on the primary heap leach pads. Figure 2 provides an overview of the metals recovery plant with the primary biological heap leach pad in the background.
Figure 2: Overview of Metals Recovery Plant & BioHeap Leach Pads There are four primary biological heap leach pads. Each pad is 400 meters wide by 1,200 meters long with ore stacked eight meters high, covering about 210 hectares in total. The ore remains on the primary bio-heap leach pad for 18 months where it is irrigated with slightly acidified raffinate solution and ventilated with forced air. The irrigation solution is continuously circulated in the exothermic leaching process to build up the metals concentration in the PLS. After 18 months, the primary heap is reclaimed and transferred to a secondary heap leach pad using reclaim equipment and a conveyor where it is again placed with a stacker. New ore is stacked on the primary heap immediately behind the reclaiming equipment to maintain a continuous leaching operation. Leaching of the secondary heap leach pad continues for 36 months. Once the leaching is complete, the secondary heaps will be covered and revegetated to blend in with the natural surrounding geography. Approximately 10% of the total irrigation circulation solution flow from the primary and secondary heap leach pads is fed to the metals recovery plant at a flow of 1,800 cubic meters per hour. Hydrogen sulphide is produced on site and used along with pH control to selectively precipitate each of the metal sulphides. As shown in the schematic in Figure 3, copper sulphide is precipitated first, followed by zinc sulphide. In the current operation, following zinc sulphide precipitation, the pH is adjusted with limestone to the correct level for gypsum precipitation. Gypsum is formed and dewatered in thickeners before being pumped to a gypsum storage pond. Nickel–cobalt sulphide is precipitated following the neutralization circuit. After nickel–cobalt precipitation, aluminum and iron are removed from the barren acid solution before it is returned to the collection pond near the primary heap leach pads. The uranium extraction stage will follow zinc sulphide precipitation when it begins to operate, as shown in Figure 3. Each of the metal sulphides that are currently precipitated are settled in thickeners, dewatered on filters, placed in containers and shipped to smelters and refineries.
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Figure 3: Schematic of the Talvivaara Sotkamo Processes
URANIUM PLANT PROCESS DESCRIPTION The PLS flow from zinc sulphide precipitation will be fed to a PLS pond upstream of the uranium extraction plant. The PLS pond will provide a capacity buffer for the uranium plant and serve as a settling pond for any sulphur particles or particles from post precipitation of zinc sulphide to minimize any crud formation in solvent extraction. PLS from the pond is pumped at a flow of 1,800 cubic meters per hour to the two Outotec VSF® solvent extraction mixer-settlers in series. The uranium at a nominal concentration of 20 ppm in the PLS will be extracted using an organic phase made up of 5% di-2-ethylhexyl phosphoric acid (D2EHPA) with Cyanex 923 as phase modifier in an aliphatic hydrocarbon diluent. Pilot tests have demonstrated a recovery above 95% and high selectivity resulting also to low impurity levels in the final product In order to reject the co-extraction of mainly iron and aluminium the loaded organic from the solvent extraction mixer-settlers will be circulated to a loaded organic tank from which the solution will be pumped to the downstream process circuit where it will undergo two stages of scrubbing with sulphuric acid to remove iron and aluminum to prevent accumulation of these impurities. Following scrubbing, the organic will be washed with water in a mixer-settler to remove entrained acidic droplets from the loaded organic. Next, the uranium will be stripped from the organic phase into an aqueous phase using sodium carbonate in two mixer-settlers. Following stripping, the barren organic will be washed with sodium hydroxide in a mixer-settler to prevent accumulation of impurities. In order to avoid extractant loss to the raffinate, the saponified D2EHPA will be transformed to the acidic form in an acidification mixer-settler before being pumped back to the solvent extraction mixer-settlers. The uranium-rich sodium carbonate from the stripping stage will be pumped through three pH adjustment reactors where the solution pH will be adjusted to the optimal acidity for precipitation. The solution from the third pH adjustment reactor is fed to a pH adjustment settler to remove entrained organic. Any remaining organic traces still in the aqueous phase will be separated from the solution in a downstream after settler and returned to the organic wash stage.
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Uranium is precipitated as UO 4. 2H2O with hydrogen peroxide in precipitation reactors. The flow from the last precipitation reactor will be pumped to a thickener where the yellowcake will be settled and pumped to a thickener underflow tank. The thickener underflow will then be washed in a single stage wash tank before being fed at around 5% solids to a product centrifuge. The concentrate containing about 70% solids will be fed to a batch conical vertical dryer with an internal o helix for mixing where it will be dried to less than 1% moisture at a temperature of 200 C The dried yellowcake will be emptied from the drier directly into 210 litre steel drums in packaging system. The packaging system will be maintained under a slightly negative pressure to contain any yellowcake dust. The gases will be processed through a scrubber system before they are released from the building. A schematic of the process flowsheet for the uranium extraction plant is shown in Figure 4.
Figure 4 Process Flowsheet of the Uranium Recovery Plant at Talvivaara Site The uranium extraction plant will produce about 350 tonnes of uranium per annum when the metals recovery plant reaches full-scale production. With a uranium concentration in the ore of 20 ppm, recovery of the uranium would not normally be economically feasible. Since the uranium is co-extracted with the other metals in a cost-effective biological heap leach operation, the uranium can be extracted from the PLS for very little additional cost and sold commercially to provide an additional revenue stream for Talvivaara.
PROJECT STATUS At the time of writing this paper, the design engineering is complete, all the major equipment has been ordered, the construction of the building structure is complete and installation of the large process equipment has started. Figure 5 shows a picture of the uranium extraction plant building under construction. Figure 6 was extracted from the 3D model of the plant created by Poyry.
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Figure 5: Construction of the Uranium Extraction Plant Building
Figure 6: Outotec VSF® mixer-settlers.for uranium extraction.
ENVIRONMENTAL BENEFITS The Talvivaara uranium extraction plant is different from other typical uranium production facilities in that it will not produce any tailings related to the operation. In fact, extraction of the uranium will have a positive environmental benefit. As mentioned previously, the PLS stream from the primary and secondary biological heap leach operations contains about 20 ppm of uranium. In the current operation without a uranium extraction plant, the flow from the zinc sulphide precipitation stage goes to a neutralization circuit, which is followed by the nickel-cobalt precipitation stage and final neutralization stages for removal of aluminum, iron, and other impurities. Uranium in the PLS gets co-precipitated in the neutralization stages. The precipitated impurities are thickened and sent to a gypsum pond where they are stored. Although the uranium is present in very small quantities in both the PLS stream and in the gypsum, it will be beneficial environmentally to remove the uranium rather than placing it in the gypsum pond.
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REGULATORY PROCESS & STATUS There are numerous government departments or agencies that must provide a permit or approval to extract and ship uranium out of Finland, especially if it is shipped outside of the European Union. The following approvals must be obtained: 1. Ministry of Employment and the Economy in Finland – Operating permit to extract uranium. 2. Ministry of the Environment in Finland – Permit to operate the plant. 3. STUK - Radiation and Nuclear Safety Authority– Operating requirements, radiation protection and uranium security. 4. European Commission – Articles 41 and 37. 5. European Commission – Euratom Nuclear Material Safeguards. Talvivaara received the operating permit to extract uranium from the Finland Ministry of Employment and the Economy in March 2012. Talvivaara submitted the environmental assessment for the uranium extraction plant in December 2010. Approval of the project from the Finland Ministry of the Environment is expected before the end of 2012. Talvivaara has worked closely with the Finland nuclear regulator, STUK, to identify all the operating requirements, provide radiation protection of the workers and the public and provide for uranium security in the plant and during shipment of the yellowcake. The European Commission is required to approve the project under two Articles. An application under Article 41 was submitted to notify the European Commission of the investment in a new installation to concentrate natural uranium (yellowcake). Talvivaara received approval under Article 41 in January 2012. An application was submitted under Article 37 to demonstrate that the operation of a facility to process nuclear material will not contaminate the water, soil or airspace of a Member State. Talvivaara expects to receive approval under Article 37 in April 2012. Talvivaara has submitted the Basic Technical Characteristics (BTC) of the uranium extraction plant to Euratom Nuclear Material Safeguards of the European Commission. The BTC contained an overview of the operation to produce yellowcake. Talvivaara updated the BTC in March 2012 to reflect any changes in the plant since the original submission. Once the plant commences operation, Talvivaara will be required to submit to the European Commission a monthly production inventory change report, an annual material balance report and physical inventory listing and an advance notice of eight working days of any shipment of uranium outside of Europe or to France or Britain.
SUMMARY The nickel / zinc ore in the Talvivaara deposit contains a very low concentration of uranium. Since the uranium is co-leached with the other metals in the biological heap leach operation, it is economical to extract the uranium from the pregnant leach solution and sell it commercially as yellowcake, providing an additional revenue stream for Talvivaara. Pilot testing has demonstrated that a solvent extraction process using a 5% solution of the organic solvent D2EHPA with a Cyanex 923 phase modifier can effectively and selectively extract the uranium from a PLS containing less than 20 ppm uranium. Downstream operations will be standard uranium metallurgical processes for yellowcake production. Talvivaara, Cameco and Outotec are working together to design, construct and commission the solvent extraction plant within Talvivaara’s existing metals recovery plant for nickel, zinc, copper and cobalt. Engineering is complete and construction is well advanced so that commissioning of the plant is expected to begin in August 2012.
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There have been a number of government and agency permits and approvals required for the uranium processing facility. All the approvals are expected to be received before the end of 2012. Although there were a number of approvals required, the process was not overly onerous. Once the plant begins to operate, it will provide an environmental improvement by removing the very low concentration of natural uranium that is currently mixed with the gypsum precipitates.
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