The October 2019 Special Issue on Critical Minerals, Part I, of the SME Mining, Metallurgy & Exploration (MME) journal is online!
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Episyenites—Characteristics, Genetic Constraints, and Mineral Potentialby E. Suikkanen and O. T. Rämö
Cite this article as:Suikkanen, E. & Rämö, O.T. Mining, Metallurgy & Exploration (2019) 36: 861. https://doi.org/10.1007/s42461-019-00120-9
Abstract
Episyenites are sub-solidus, quartz-depleted alkali-feldspar-rich rocks. They form veins and lenticular bodies in granitoid rocks and migmatites in a late- to post-orogenic or anorogenic setting. Leaching of quartz is usually a response to a flux of weakly saline hydrothermal solution in circulation cells above cooling intrusions, where sufficient fluid-rock ratios and thermal gradients are achieved. Fluid Si-undersaturation is achieved by rapid cooling within the field of retrograde Si solubility or by temperature and pressure increase outside retrograde conditions. Some quartz may also be consumed in metasomatic reactions and in response to pressure fluctuation in sealed episyenite bodies. The small size and overall rarity of episyenites imply that conditions for episyenite formation are not commonly encountered in the crust. In addition to quartz depletion, episyenites record complex histories of metasomatic alteration and hydrothermal mineral growth. Nearly all episyenites have undergone Na-metasomatism, which may have led to the formation of nearly monomineralic albitite, and which is occasionally followed by late K-metasomatism, phyllic alteration, and argillization. Depending on the effectiveness of later compaction, recrystallization and vug-filling episyenites may preserve the macroscopic porosity formed by quartz dissolution and brittle deformation. Vuggy episyenites can act as significant sinks for metals carried by crustal fluids and host many significant U, Sn, and Au deposits worldwide. Rare earth-critical syenitic fenites around alkaline intrusions share mineralogical and genetic traits with episyenites.
Petromineralogical Studies of Late Paleocene–Middle Eocene Phosphate Nodules in the Subathu Basin of Solan District, Himachal Pradesh
by Mohd Shuaib, K.F. Khan, Samsuddin Khan, Shamim A. Dar
Cite this article as:
Shuaib, M., Khan, K.F., Khan, S. et al. Mining, Metallurgy & Exploration (2019) 36: 879. https://doi.org/10.1007/s42461-019-0082-6
Abstract
The Late Paleocene–Middle Eocene phosphate nodules of Solan District of Himachal Pradesh belong to green facies of Subathu Formation of Sirmur group. Field observations indicate sporadic distribution of phosphatic nodules in the yellow brown to rusty brown, buff-colored bleach shales. Phosphatic nodules are dull earthy to dark grayish in color and break with sharp and conchoidal surfaces. Studies of phosphatic nodules using petrological microscope, X-ray diffractometry, and scanning electron microscopy (SEM) attached with energy dispersive X-ray (EDX) reveal that cryptocrystalline apatite (fluorapatite) is the dominant phosphate mineral. Silt sized silica is the dominant gangue, whereas pyrite and muscovite are present as minor gangue. The phosphate minerals occur as oolites, pellets, and cryptocrystalline apatite in the groundmass. Silica veins form a network replacing the original phosphatic matrix. Brown to brownish black color may be attributed by the organic matter in the phosphatic nodules. Completely phosphatized radiolaria and partially pyritized planktonic foraminifera are observed. There is corroboration of organic matter in the form of microbial filaments in the groundmass of apatite. The findings in the present investigation indicate that the different forms of the minerals, their texture, and distribution in the phosphatic nodules might be due to environmental vicissitudes in oxidizing to reducing conditions followed by replacement processes. Presence of microbial filaments may be due to the active role of microorganisms during the formation of these phosphatic nodules.
A Field Survey of Rare Earth Element Concentrations in Process Streams Produced by Coal Preparation Plants in the Eastern USA
by Gerald H. Luttrell, Michael J. Kiser, Roe-Hoan Yoon, Aaron Noble, Mohammad Rezaee, Abhijit Bhagavatula, Rick Q. Honaker
Cite this article as:
Luttrell, G.H., Kiser, M.J., Yoon, RH. et al. Mining, Metallurgy & Exploration (2019) 36: 889. https://doi.org/10.1007/s42461-019-00124-5
Abstract
A field study was undertaken to experimentally measure the concentrations of rare earth elements (REEs) contained in the process streams generated by a group of 20 coal preparation plants located in the eastern USA. For each site, representative samples of clean coal product, coarse refuse, and fine refuse were collected. Each sample was then partitioned into preselected size and density classes by wet screening/sieving and float-sink testing. The resultant products were dried and subjected to laboratory analyses to determine ash contents and rare earth element concentrations including Yttrium and Scandium. A detailed analysis of the database generated by this exercise showed that coal-based products from these preparation plants contained significant quantities of rare earth elements. In particular, the coarse refuse streams currently discarded by the 20 plants examined were found to contain a sufficient tonnage of REEs to satisfy the current domestic demand for these important elements. The data also showed a strong positive correlation between ash content and REE concentration, which suggested that the primary association of REEs in eastern USA bituminous coal sources is likely within fractions containing inorganic impurities. This association was well described using a simple power equation relating ash content and REE concentration. The ratio of heavy-to-light rare earth elements was discovered to be significantly higher in organically rich fractions of clean coal, suggesting that mineral impurities intimately associated with carbonaceous matter have elevated concentrations of heavy rare earth elements. A similar trend was observed for a grouping of rare earth elements (Y, Nd, Eu, Tb, and Dy) that are likely subject to near-term supply shortages. Finally, the database showed that a linear correlation existed between La for many REEs of interest, although there were several notable exceptions for some high-value REEs (Lu, Pr, and Tb).
The Occurrence and Concentration of Rare Earth Elements in Acid Mine Drainage and Treatment By-products: Part 1—Initial Survey of the Northern Appalachian Coal Basin
by Christopher R. Vass, Aaron Noble, Paul F. Ziemkiewicz
Cite this article as:
Vass, C.R., Noble, A. & Ziemkiewicz, P.F. Mining, Metallurgy & Exploration (2019) 36: 903. https://doi.org/10.1007/s42461-019-0097-z
Abstract
The conventional rare earth element (REE) industry has historically sought to develop ore deposits where geologic processes have produced mineralized zones with commercially attractive REE concentrations. These deposits are extremely uncommon, particularly in the USA. Given the criticality of these materials and the need for sustainable domestic supply, the current research seeks to leverage other autogenous processes that lead to concentrated REE resources. One such process is the generation of acid mine drainage (AMD). AMD is very common in many coal mining districts and results from the exposure and oxidation of pyrite during mining. During the generation and migration of AMD, liberated sulfuric acid mobilizes several metal ions including REEs. Treatment of AMD is required under U.S.C §1251, the Clean Water Act, and often consists of neutralization, oxidation, and metal hydroxide precipitation. To investigate the deportment of REEs during this process, a field sampling campaign was undertaken, whereby the concentration of REEs in AMD and AMD precipitates was measured directly. In the nine sites evaluated in this study, the REE concentrations of the precipitates varied from 29 to 1286 ppm with an average of 517 ppm among the sampled sites. The individual elements were enriched compared with the associated bulk Northern Appalachian (NAPP) coal material by factors ranging from 3 to 15. Furthermore, the distribution of REEs in all precipitate samples favored the heavy REEs (HREEs) when compared with traditional REE ores. This research represents the first part of multi-part research endeavor to characterize, classify, and determine the practicality of refining REEs from AMD and its by-products.
The Occurrence and Concentration of Rare Earth Elements in Acid Mine Drainage and Treatment Byproducts. Part 2: Regional Survey of Northern and Central Appalachian Coal Basins
by Christopher R. Vass, Aaron Noble, Paul F. Ziemkiewicz
Cite this article as:
Vass, C.R., Noble, A. & Ziemkiewicz, P.F. Mining, Metallurgy & Exploration (2019) 36: 917. https://doi.org/10.1007/s42461-019-00112-9
Abstract
Many modern industries rely on rare earth elements (REEs) to produce products that are essential to both civil and defense applications. In a prior study (Vass et al., 2019), the authors showed that REE grades in acid mine drainage (AMD) and associated byproduct precipitates from AMD treatment (AMDp) warrant evaluation as a feedstock for REE production. The current work extends that effort through a broad survey of 141 AMD treatment sites in Northern and Central Appalachia. In this study, 185 raw AMD and 623 AMDp field samples were obtained and analyzed to assess the REE and major metal concentrations. Results show that an average of 282 μg/L and 724 g/tonne of REEs occur in AMD and AMDp respectively. Additionally, both basins contained similar distributions of REEs, and these distributions tended to favor heavy and critical REEs when compared with traditional REE ore deposits. Geospatial analysis identified a total resource of 340 tonnes stored at the 141 sites sampled in this study. While this analysis did not quantify the basin-wide REE inventory, it does indicate the impact that processing cut-off grades will have on the overall AMDp resource base.
Application of PVC as Novel Roasting Additive in Vanadium Extraction from Stone Coal
by Tiejun Chen, Qiyong Li, Qiang Li, Zhihong Wu, Guowen Dong, Junying Wan
Cite this article as:
Chen, T., Li, Q., Li, Q. et al. Mining, Metallurgy & Exploration (2019) 36: 931. https://doi.org/10.1007/s42461-019-0084-4
Abstract
In order to search for a more economical and effective roasting additive for extracting vanadium from stone coal and solve the question of the disposal of PVC waste, PVC waste was proposed as a novel roasting additive to enhance extraction of vanadium from stone coal and realize the comprehensive utilization of resources. By adding this roasting additive, the leaching rate of vanadium reached 94.5% under the conditions that the PVC waste addition was 10 wt%, the roasting temperature was 800 °C, the roasting time was 1 h, the sulfuric acid concentration was 15 vol%, the liquid-to-solid ratio was 1.5 cm3/g, the leaching temperature was 95 °C, and leaching time was 4 h, which increased by 8.4% compared with that of blank roasting. The thermodynamics calculation showed that the decomposition reaction of mica was performed easier with the participation of HCl produced by pyrolysis of PVC. Meanwhile, The SEM-EDS and aperture test analyses demonstrated that more cracks and voids appeared in the roasted sample by roasting with PVC waste, which could facilitate the release and extraction of vanadium. Additionally, the TG-DSC analyses indicated that PVC could provide large amounts of heat for the roasting process to promote the decarbonization and roasting efficiency of stone coal because of its high calorific value.
H3PO4 Production Process Utilizing Phosphatic Clay as Feed Material
by Matthew Gordy, Robert M. Counce, Patrick Zhang, Rasika Nimkar, Jack S. Watson
Cite this article as:
Gordy, M., Counce, R.M., Zhang, P. et al. Mining, Metallurgy & Exploration (2019) 36: 941. https://doi.org/10.1007/s42461-019-0074-6
Abstract
Waste phosphatic clay presents a difficult disposal problem and a significant loss of P2O5. Recent developments make this study on the use of phosphatic clay as a feed material for H3PO4 production worthwhile. (a) New data from Florida Industrial and Phosphate Research Institute (FIPR) suggest that the “clay” waste can be effectively treated by improved solid separation coupled with flotation to produce feed material for recovering additional phosphate. (b) The use of “OLI Flowsheet” provides capability to predict performance of aqueous-based processes for a range of possible operating conditions. Using the new FIPR data and “OLI Flowsheet” enabled the design of a new process that can potentially extend the usefulness of current P2O5 ore deposits. Predictions using OLI Flowsheet were able to identify process conditions that effectively overcome the difficulty of high Mg content of phosphatic clay. Estimated manufacturing cost in the current study lead to valuation of the P2O5 content of the phosphatic clay recovered/utilized at 60 to 70% of that of the phosphate rock feed material currently used. Phosphatic clay should be considered as a valuable material that is currently given zero value. The new FIPR data and “OLI Flowsheet” enables the design of a new process that can potentially extend the usefulness of current phosphate ore deposits by extending the fraction P2O5 of recovered.
Upgrading of Phosphate Fines by Fatty Acid Flotation Using Amylase Enzyme as a Surface Modifier
by A. Yehia, K. E. Yassin, M. Amar
Cite this article as:
Yehia, A., Yassin, K.E. & Amar, M. Mining, Metallurgy & Exploration (2019) 36: 949. https://doi.org/10.1007/s42461-019-00126-3
Abstract
Phosphate fines (size ≤ 45 μm), from the Red Sea Region, Egypt, were subjected to beneficiation process to recover the phosphorite grains from such fines. The sample is characterized by low P2O5 content (18.78%) associated with gangue minerals such as silica (22.77% SiO2) and carbonate (2.01% MgO) indicating its low grade. Phosphate pre-concentrate was prepared by gravity separation, using Falcon, where 63.37% of the phosphate, with 24.94% P2O5, 1.46% MgO, and 10.54% SiO2, was firstly recovered reducing the mass flow to the subsequent beneficiation process. However, the gravity tailing was the feed for the reverse phosphate flotation using bench scale column flotation where oleic acid was used as a carbonate gangue collector and amylase enzyme as a phosphate depressant. Under the appropriate flotation conditions (0.1% amylase, 5·10−4 mol/dm3 oleic acid, and temp. 30 °C), phospho-concentrate assaying 0.61% MgO, 13.14% SiO2, and 27.85% P2O5 with a P2O5 recovery of 55.45%, was finally obtained without the use of expensive depressants, e.g., phosphoric acid or sodium silicate. A tentative flow sheet for the whole process was postulated.
A Micro-Scale Investigation of the Adsorption of Collectors on Bastnaesite
by Jinhong Zhang, Dongbo An, James Withers
Cite this article as:
Zhang, J., An, D. & Withers, J. Mining, Metallurgy & Exploration (2019) 36: 957. https://doi.org/10.1007/s42461-019-00118-3
Abstract
A micro-scale investigation was carried out by applying atomic force microscopy (AFM) to study in situ the adsorption of various collectors, i.e., oleic acid, octanohydroxamic acid (HA), and salicylhydroxamic acid (SHA), on bastnaesite in aqueous solutions. The obtained AFM images show that the surface morphology of bastnaesite changes greatly after it comes into contact with the solutions of the collectors, suggesting that all these collectors can effectively adsorb onto bastnaesite. Increasing temperature can facilitate the adsorption of oleic acid onto bastnaesite. Results from attenuated total reflectance–Fourier transform infrared spectrometry (ATR-FTIR) also show that all of these collectors adsorb strongly onto bastnaesite, with strong absorbance spectra being detected. The ATR-FTIR results confirm those obtained by AFM. In general, hydroxamic acid collectors (HA and SHA) adsorb onto bastnaesite mainly in the form of insoluble metal hydroxamate. This specific adsorption mechanism explains that a high selectivity with a moderate collectivity will be achieved with a hydroxamic acid collector for the flotation of bastnaesite.
Selective Precipitation of High-Quality Rare Earth Oxalates or Carbonates from a Purified Sulfuric Liquor Containing Soluble Impurities
by Ruberlan Gomes Silva, Carlos Antonio Morais, Leandro Viana Teixeira, Éder Domingos Oliveira
Cite this article as:
Silva, R.G., Morais, C.A., Teixeira, L.V. et al. Mining, Metallurgy & Exploration (2019) 36: 967. https://doi.org/10.1007/s42461-019-0090-6
Abstract
The purity of rare earth products depends on the ability of the reagents used to selectively promote the precipitation of the rare earth elements (REE) from impurities present in sulfuric liquors. The present study aims at investigating the conditions required to precipitate REE, either as rare earth oxalates or carbonates, from purified rare earth sulfuric liquor with a low Al3+ and UO22+ and a very high Ca2+, Mg2+, Mn2+, and SO42− impurity content. The rare earth sulfuric liquor was obtained in a previous study by the reaction of a rare earth ore with sulfuric acid, pyrolysis at 700 °C/2 h, cooling at 20 °C, and water leaching. The impurities such as Fe3+, Th4+, and PO43− were completely removed from the liquor in two consecutive neutralization steps. In order to recover the REE, the purified rare earth liquor was then treated with oxalic acid and sodium carbonate. Rare earth oxides (REO) precipitated with oxalic acid displayed a very high purity (99.2% w/w) and low impurity content. In contrast, the rare earth carbonates precipitated with sodium carbonate presented a comparatively lower REO content (68.5% w/w) and 94.2% of purity, but with a higher content of impurities. Better precipitating conditions were achieved at 60 °C, compared with 20 °C, and under stoichiometric reagent consumption, with REE precipitation efficiencies higher than 96%. The dosage above the stoichiometric condition caused the precipitation of calcium oxalates and co-precipitation of rare earth sulfates when oxalic acid is used, and precipitation of calcium, manganese, and uranium carbonates and also a co-precipitation of rare earth sulfate when sodium carbonate were used. An integrated flowsheet to produce a high-quality rare earth product from rare earth sulfuric liquor containing some impurities is proposed.
Circulation of Sodium Sulfate Solution Produced During NiMH Battery Waste Processing
by Antti Porvali, Vivek Agarwal, Mari Lundström
Cite this article as:
Porvali, A., Agarwal, V. & Lundström, M. Mining, Metallurgy & Exploration (2019) 36: 979. https://doi.org/10.1007/s42461-019-0086-2
Abstract
Hydrometallurgical recovery of rare earth elements (REE) from NiMH battery waste can be performed using sulfuric acid leaching followed by selective precipitation as double salt (REENa(SO4)2·H2O) by adding Na2SO4 as a precipitating agent. The formed double salts can then be further converted with NaOH solution to form REE hydroxides. However, present literature has paid little attention to how to minimize the consumption of reagents in this recovery process. In the current study, a process model was built using HSC Sim software, and metamodeling of the simulated flowsheet was conducted with the design of experiments (DOE), after which the basic functionality of the flowsheet was verified by experimental work, starting from sulfuric acid leaching of spent batteries. As the process is based on chemical circulation internally in the process, the objective was to investigate the accumulation of the elements in solution, specifically Na+, K+, SO42− ions and impurities liberated in the NaOH conversion step. It was found out that metamodeling could be utilized in obtaining first estimate of the process behavior, prior to any experiments. Flowsheet functionality with circulated solution was confirmed, obtaining at best 96.9% La recovery.
Direct Pressure Alkaline Leaching of Scheelite Ores and Concentrates
by Patrícia Leitão, Aurora Futuro, Cristina Vila, Lurdes Dinis, Anthony Danko, António Fiúza
Cite this article as:
Leitão, P., Futuro, A., Vila, C. et al. Mining, Metallurgy & Exploration (2019) 36: 993. https://doi.org/10.1007/s42461-019-00109-4
Abstract
The average global recovery rates for processing scheelite ores by physical methods are normally in the range 50 to 70%, and it is problematic to attain a minimum grade of 60% WO3 in the concentrates without penalizing further this yield. Direct hydrometallurgical processing of low-grade concentrates or medium- to high-grade ores may become an interesting global economic alternative as it allows increasing the recovery rate, avoiding major penalizations due to the grade of the concentrate and increasing the added value of the final product. Leaching of scheelite by sodium carbonate presents several advantages such as its selectivity regarding the dissolution of troubling metals like iron and aluminium, allowing for a simple and flexible process diagram. However, it has the disadvantage that it is performed at high temperature and pressure with a high consumption of reagents. The feasibility of leaching is appraised by the extraction yield and selectivity, but also by the operating cost. Considering these aspects, a relevant objective of the research was to transform alkaline pressure leaching into a feasible processing alternative directly applicable to both medium- and high-grade ores and simultaneously to low- and high-grade concentrates produced by physical methods. The research used ore samples and low-grade concentrates from a typical skarn orebody as well as high-grade concentrates from a chalcopyrite-scheelite orebody. Different operating conditions were tested. Temperature, pressure and reagent concentration showed to be fundamental, but not exclusive, in determining the leaching behaviour. The operating conditions are milder than the ones generally proposed in the literature or currently in use in industry, but with higher carbonate consumptions. The research was oriented towards potential industrial implementation allowing for the application of optimization methods and straight scale-up to pilot-plan design and project.
Selective-Reductive Leaching of Manganese from Low-Grade Manganese Ore Using Tannic Acid as Reductant
by Erik Prasetyo, Eti Purwaningsih, Widi Astuti
Cite this article as:
Prasetyo, E., Purwaningsih, E. & Astuti, W. Mining, Metallurgy & Exploration (2019) 36: 1003. https://doi.org/10.1007/s42461-019-00115-6
Abstract
Selective-reductive leaching of manganese from low-grade manganese ore using tannic acid as a reductant in acidic media was demonstrated by XRD, SEM-EDX characterization, and batch leaching studies at room temperature. The pH change in the liquid phase before and after leaching revealed that selectivity of manganese over iron was due to the hydrolysis of iron. The acidity in the liquid phase during leaching was dictated by two mechanisms: proton-consuming reaction (manganese oxide reduction by tannic acid) and proton-producing reaction (ferric ion reduction by tannic acid to produce ferrous ion). Based on the leaching studies, the more prevalent reaction between these two was determined by initial leaching conditions: tannic acid concentration, sulfuric acid concentration, liquid-solid ratio, temperatures, and agitation time. Kinetic studies revealed that maximum recovery was attained within 6-h agitation, which indicated a product layer diffusion process.
Effect of Base Types on the Properties of MgO Particles Obtained from Dolomite Ore
by Mahmut Altiner
Cite this article as:
Altiner, M. Mining, Metallurgy & Exploration (2019) 36: 1013. https://doi.org/10.1007/s42461-019-00122-7
Abstract
This study aimed to produce magnesium oxide (MgO) particles with different properties from dolomite ore using an experimental procedure comprising four stages: sample preparation (S1), HCl leaching (S2), precipitation (S3), and calcination (S4). Three different base sources (NaOH, KOH, and NH4OH) were used as precipitant in the third stage to obtain magnesium hydroxide [Mg(OH)2] from a leachate solution, which was obtained in the second stage. Next, Mg(OH)2 particles generated by different alkali sources were calcined at various temperatures from 600–1000 °C for different durations (1–5 h). The effect of these base types on the properties of each product was investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM), Brunauer–Emmett–Teller (BET), atomic force microscopy (AFM), and wet chemical analyses. The experimental results indicated that each product was identified as periclase (MgO) mineral, which was produced with a specific surface area (SSA) of 4.49–44.54 m2/g depending on the production conditions. The surface roughness of the MgO particles increased with increasing calcination temperature. SEM analyses showed that MgO particles produced at a temperature of 600 or 800 °C were amorphous, indicating that the process was not influenced by the base type, but MgO crystals were smooth when the calcination temperature was 1000 °C. Finally, it was determined from all experimental findings that MgO particles produced via the addition of NaOH have superior properties (such as higher SSA and lower surface roughness) compared with MgO particles produced with KOH or NH4OH as the alkali source. These properties led to great improvement of its usability in industry.
Leaching Process and Mechanism of Weathered Crust Elution-Deposited Rare Earth Ore
by Zhengyan He, Rong Zhang, Wenrui Nie, Zhenyue Zhang, Ruan Chi, Zhigao Xu, Ming Wu, Jun Qu
Cite this article as:
He, Z., Zhang, R., Nie, W. et al. Mining, Metallurgy & Exploration (2019) 36: 1021. https://doi.org/10.1007/s42461-019-00116-5
Abstract
The leaching processes and mechanisms of weathered crust elution-deposited rare earth (RE) ores with ammonium sulfate were investigated for their high efficiency and low consumption exploitation. The leaching process is a typical reversible heterogeneous ion exchange reaction in liquid–solid systems. The raw ore and leached tailings were analyzed by SEM and FTIR. Results indicated that no dissolution or structural changes of the ores occurred during the leaching process. The migrations of RE3+, Al3+, and NH+4 in the leaching process were studied under various liquid–solid ratios. With increasing liquid–solid ratio, the content of NH+4 adsorbed on clay mineral by physical effect decreased, NH+4 adsorbed by ion-exchange increased, and the content of exchanged RE and Al increased until reaching equilibrium. The adsorption capacity of the RE ores for NH+4 remained about 0.61–0.63 mg/g. It was difficult to desorb NH+4 from RE ores with washing water alone. The leaching mechanism of Al was the same as RE, but the leaching of Al slightly lagged behind RE. At the optimal liquid–solid ratio of 0.4:1, the leaching efficiencies of RE and Al were 95.51% and 90.99%, respectively, and 56.65% of NH+4 remained in the leachate. After impurity removal and RE precipitation, the leachate can be reused by replenishing fresh leaching agent. The aim of this study is to enhance the leaching process of RE, restrain the leaching of impurities, and decrease the consumption of leaching agents.