One-Dimensional (1D) Reactive Transport with PFLOTRAN: Marl Leaching Case - Progress Status.
Abstract not provided.
Jove-Colon, Carlos F.↗
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One-Dimensional (1D) Reactive Transport with PFLOTRAN: Marl Leaching Case - Progress Status.
Abstract not provided.
Jove-Colon, Carlos F.↗
Reactive Transport Modeling Results Update: EPA1315 OPC Leaching Experiments.
Abstract not provided.
Lopez, Carlos↗
THE UTILITY OF SONARS AND THE SANSMIC LEACHING CODE FOR MONITORING CAVERN SHAPE DEVELOPMENT DURING THE RECENT UNPRECEDENTED OIL VOLUME MOVEMENTS AT THE SPR
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Zeitler, Todd↗
Ex Situ Tailings Leaching and Lateritization with Electrolytic Acid Recycling for Critical Metal Concentration and Mineral Carbon Sequestration
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37 INORGANIC, ORGANIC, PHYSICAL, AND ANALYTICAL CH↗
Pilot-Scale Testing of an Integrated Circuit for the Extraction of Rare Earth Minerals and Elements from Coal and Coal Byproducts Using Advanced Separation Technologies
The primary objective of this project was to develop and demonstrate an integrated pilot-scale circuitry for recovering high-value rare earth elements (REEs) from coal and coal byproducts. The target performance was to produce a mixed REE product with content of at least two percent by weight on a dry mass basis in a cost-effective and environmentally benign manner. During the first nine months of the project period (Phase 2 Budget Period 2), pilot plant construction was completed including all field site startup activities such as permitting, engineering design, procurement/bidding, unit fabrication, site construction, equipment installation, module assembly, safety training, and circuit shakedown. During the remaining 21 months of project period (Phase 2 Budget Period 3), detailed field-testing activities were performed including feedstock sample collection and preparation, exploratory testing, circuit modification, detailed parametric study, and performance optimization. A detailed techno-economic analysis was performed based on the pilot plant testing findings which provided various scenarios for REE production. The project successfully accomplished the proposed target performance by producing mixed rare earth oxide (REO) with greater than 90% purity by weight in a continuous pilot scale operation from two distinctly different coarse refuse materials (i.e., West Kentucky No. 13 and Fire Clay coal seams), and at least three secondary sources (i.e., heap leach process and naturally formed acid mine drainage system). Project partners included the University of Kentucky, Virginia Tech, West Virginia University, Alliance Coal, Blackhawk Mining, Mineral Refining Company, and Mineral Separation Technologies. The pilot scale test facility was constructed at a former mining complex owned by Alliance Natural Resource Partners (Alliance Coal). The site was rehabilitated to accommodate the equipment installation, construction and fabrication, electrical power requirement, water line management and containment. The process units constructed and installed included X-ray sorting unit, crushing and grinding unit, physical separation unit, acid leaching unit, solvent extraction unit, and wastewater management unit. A rare earth mineral concentration unit was constructed as a standalone unit for flexible operation. A detailed environmental assessment and control plan was carried out to identify and quantify any potential impacts of the pilot-scale processing circuitry on the human and eco-system health and well-being. Corresponding mitigation strategies and control measures were provided. A conceptual flowsheet was developed to effectively remove thorium and uranium from high purity rare earth oxide mix or any potential radionuclide enriched stream. The two distinct feedstock materials were secured from the Blackhawk Mining Complex in eastern Kentucky where the Fire Clay (Hazard No. 4) seam is processed. The West Kentucky No. 13 (Baker) coarse refuse material was collected from an active process stream at an Alliance coal preparation plant located in western Kentucky. Characterization analysis indicated that both of feed materials generated from the two sources contained >300 ppm of TREEs on a dry whole mass basis which met the requirements for a qualified feed stock. The two feedstocks were further upgraded using a dual x-ray sorter to prepare the feed material for hydrometallurgical circuit. Thermal treatment on feed material prior to leaching was found to: 1) improve the leaching recovery of REEs, 2) increase the leaching kinetics, and 3) allow the leaching reaction to occur at lower acidity. Roasting at 600°C was selected as the pre-treatment condition for both West Kentucky No. 13 and Fire Clay coarse refuse material. Over 40% of leaching recovery was achieved by roasting West Kentucky No. 13 material having a top particle size of 3 mm in the pilot scale operation using 1.2M sulfuric acid leaching at 75OC. Initial pilot scale testing involved continuous operation of the pilot plant for 94 hours. The leaching unit was operated at solid-to-liquid ratio of 1 to 10 (w/v) using 0.5M sulfuric acid solution at a temperature of 75°C. The continuous solvent extraction circuit utilized rougher and cleaner units with DEHPA and TBP as the extractants. An innovative stripping circuit was developed to accumulate the REE concentration in the stripping solution to a level above 600 ppm. A bleed stream from the recycled strip solution was treated using oxalic acid precipitation which produced a high grade rare earth oxalate. The oxalate product was roasted to remove the oxalate which produced a rare earth oxide product having a purity greater 90%. Due to high concentrations of contaminant ions in the pregnant leach solution (PLS), a modified flowsheet was developed that involved pre-concentration of the REEs using multiple stages of precipitation and redissolution. The advantage of this process was improved removal of contamination before the downstream purification process and a significant cost reduction relative to the circuit that utilized the solvent extraction process. The modified circuitry included processes involving leaching, multistage precipitation, redissolution, and oxalate precipitation followed by roasting of the oxalate product. The circuit produced a mixed REO that was 92.96% pure from the initial test. A detailed parametric test plan was carried out which involved varying key parameters including solids feed rate, acid flowrate, acid concentration, multistage precipitation pH, redissolution pH, oxalate precipitation dosage and pH. The response variables included REE recovery, contaminant recovery, REO product grade and overall chemical consumption. Test results indicated that the acid-to-solid ratio is the key parameter to leaching efficiency as performance deteriorated with an increase in solids concentration. The optimal pH determined for REE precipitation and redissolution was 6.5 and 2.5, respectively. Additional tests were conducted to further improve the flowsheet. Recirculating a portion of the PLS to the feed of the leach tanks improved the leaching performance by lowering the pH of the leaching system and reducing the contamination recovery by shortening the residence time. Moreover, the removal of Al prior to REE precipitation significantly reduced the oxalic acid consumption in the oxalate precipitation circuit. The modified circuit produced over 90% grade REO by weight from both West Kentucky No. 13 and Fire Clay coarse refuse material in pilot scale continuous test programs. A case analysis model was developed to project the REE and major contaminants concentration in each PLS stream based on the leaching condition, pH cut point, and oxalic acid dosage. A correlation was established using empirical and semi-empirical models. Using the models, chemical consumption required for each stage was predicted based on the projected performance of the hydrometallurgy circuit. After identifying the optimum conditions, validation tests were carried out for the treatment of both West Kentucky No. 13 and Fire Clay coarse refuse materials in the pilot plant. The actual circuit performance and chemical consumptions were very close to the model predictions. Other than the two coarse refuse sources, several secondary feedstocks were also tested in the pilot plant facility. A “heap leach” system was constructed using the coarse refuse material generated from cleaning the West Kentucky No. 13 seam coal. Using the two stage SX rougher and cleaner circuit, a concentrate with a grade >90% REO was produced while recovering >97% of the REEs from the heap leach PLS. Naturally generated acid mine drainage (AMD) from West Kentucky No.13 mine was processed using the multistage precipitation circuit in the pilot plant in a test conducted for a period of 32 hours. The final grade of the mix RE oxide produced from the AMD was 90.84% with an overall circuit recovery of 64%. The primary source of REE was the selective precipitation steps involving iron and aluminum rejection. The hydrophobic-hydrophilic separation (HHS) process was proven to effectively recover coal from fine waste materials. For REM recovery, the HHS process was able to produce concentrates at grades of approximately 1.8% REE on an ash basis; however, recovery values were typically low, <10%, under the optimal conditions determined in the laboratory-scale testing. Staged testing of the pilot-scale HHS process for coal recovery and semi-continuous laboratory testing for REM testing showed that a total concentration ratio of more than 15x was observed for the REM recovery process. A circuit simulation package was developed for REE extraction and purification using a spreadsheet-based platform (Microsoft Excel). The REESim circuit simulation package is configured to track the mass and volume flows of components passing through a series of unit operations specified and configured by the user. The mass rates can then be utilized by the user to determine important performance indicators such as product mass yields, concentrate purity levels, element-by-element recoveries, and so forth. The techno-economic analysis showed that the roasting and leaching operations were the most expensive capital items, each contributing approximately 30% to the total capital cost. One notable contributor to the high production costs was the low REE recovery observed in the pilot scale trials. The product basket price was shown to have a strong influence on the economic viability of the scenarios, with the scandium price being the most significant influencer. Operating cost was shown to be extremely sensitive to REE recovery, REE feed grade, and leaching acid consumption. An analysis of ten different scenarios for a 500 t/h commercial operation revealed that three were economically favorable, producing internal rates of return varying from 27.7% to 33.1% and payback periods of 4 to 5 years. The project successfully developed and demonstrated a process to recover REEs from coal and coal byproducts in a pilot-plant operation which consistently produced over 90% grade REO mix from varies types of feedstocks. Commercialization analysis showed that the technology readiness level successfully achieved TRL 6 at the end of the project and demonstrated the need and the potential for scaling the process to further advance the technologies toward the goal of providing a domestic supply of REEs at a commercial scale.
01 COAL, LIGNITE, AND PEAT↗
Laboratory Evaluation to Increase Effectiveness of Field-Scale Soil Flushing in the Hanford 100 Areas
This laboratory study was initiated to develop an improved technical understanding of the key controlling geochemical and physical processes of mobilizing residual Cr(VI) in Hanford sediments to provide increased effectiveness in soil flushing activities planned for Hanford 100 Areas. This was accomplished by (1) quantifying Cr mass, release rate, and surface phase changes as Cr is leached from different Hanford sediments; (2) maximizing Cr leaching during soil flushing by evaluating different leach solutions and leach strategies; and (3) minimizing leaching of remaining residual Cr after soil flushing ends. To quantify geochemical controls on Cr leaching from sediments, sequential liquid extractions were used to identify aqueous, adsorbed, and solid Cr surface phases (precipitates or Cr incorporated into other phases) before and after water-saturated and unsaturated leach experiments with pH 8 artificial groundwater and with other amendment solutions. Changes in the Cr release rate from sediments were also correlated to changes in Cr surface phases. During unsaturated infiltration, water advection occurs primarily through larger pores, with less movement through smaller grain layers leading to a slower release of Cr trapped in smaller pores next to smaller grains. To quantify these physical controls on Cr leaching from sediments, 10 ft high 1-D infiltration experiments were conducted at differing leach solution application rates, and Cr leaching during application and subsequent residual water flow was quantified. Qualitative 2-D experiments were also conducted to evaluate surfactant addition to increase flow in low-permeability zones. Cr-contaminated Hanford sediments from the 100-H, 100-K, 100-D, and 200 East Areas exhibited a wide variety of leach behavior from rapid release of a small amount of Cr to fast (i.e., minutes) and slow (i.e., hours or more) release of high Cr. Sediments with low labile Cr (< 0.2 µg/g) exhibited fast Cr release but less overall release, and most (>95%) leached from the sediments within a few pore volumes in water-saturated columns. With nearly no sorption of chromate, aqueous and sorbed Cr (as chromate) were released quickly from sediment within the first few pore volumes, resulting in high leach concentrations, fast initial Cr release rates, and changes in pre- and post-leach extractions. In contrast, sediments with high labile Cr (0.2 to 20 µg/g) exhibited a combination of fast and slow Cr release due to multiple Cr surface phases contributing to Cr release from the sediment at different rates with greater overall release. This resulted in elevated Cr effluent concentrations even after dozens of pore volumes were leached. In addition, the Cr release rate decreased with increasing time during leaching. A decrease in labile Cr (aqueous Cr, adsorbed Cr, pH 5 acetate dissolved precipitates) in post-leach extractions compared to pre-leach extractions showed initial aqueous and adsorbed Cr release, followed by dissolution of high-solubility precipitates such as CaCrO 4 (if present), then slow dissolution of calcite that may contain some chromate and possibly BaCrO 4 . The labile Cr in all sediments correlated well with the Cr release rate (r 2 = 0.81). Labile Cr also correlated well with the leached mass in water-saturated columns (r 2 = 0.98), and therefore can be used for prediction of potential removal with flushing applications.
12 MANAGEMENT OF RADIOACTIVE AND NON-RADIOACTIVE W↗
Low-Temperature Plasma Treatment for Enhanced Recovery of Highly Valued Critical REEs from Coal (Final Report)
The principle objective of this project was to develop a novel process flowsheet using low-temperature plasma treatment (LTP) integrated with conventional hydrometallurgical processes to recover rare earth elements (REEs), especially highly valued REEs (i.e. scandium and critical REEs), from coal and coal byproducts. LTP treatment was used to pretreat the solid feed material prior to leaching in an effort to improve the leaching characteristics of the REEs and potentially reduce acid consumption. Initially, a detailed experimental program was performed on a laboratory LTP furnace to identify a set of process parameter values to be used for assessing the enhanced leaching performance achieved on four specific gravity fractions of West Kentucky No. 13 and Fire Clay coarse coal refuse materials. High-temperature oxidation (HTO)) was also evaluated as a means for improving REE leachability on the same fractions of both coal sources. The lixiviants used to assess leachability improvements included: i) untreated water, ii) 0.1 M ammonium sulfate solution and iii) 1.2 M sulfuric acid solution at a temperature of 75°C. The laboratory data was used to develop multiple process flow sheets using HTO only to produce a rare earth oxide mix concentrate or utilizing both LTP and HTO treatment processes to generate a high purity heavy REO mix concentrate and a light REO mix concentrate. A techno-economic analysis was performed on each circuit. Low-temperature plasma oxidation was found to enhance REE leaching characteristics for all specific gravity (SG) fractions of the West Kentucky No. 13 coarse refuse material. The leaching recovery improvement was preferentially achieved for the heavy REEs with higher values obtained when treating the lower density fraction material. For example, the leaching recovery for the heavy REEs in the untreated 1.6 SG Float fraction material was 8% in a 0.1 M ammonium sulfate solution. After LTP treatment for 32 hours, heavy REE recovery increased to 33%. The corresponding recovery value for the light REEs was 10%. These findings indicate a strong affinity of the heavy REEs with the organic material, which is the component being reduced in content during the LTP treatment. Scanning electron micrographs showed the development of a honeycomb structure with increased LTP treatment with higher surface area as confirmed from BET measurements. Higher surface area promotes improved leaching efficiency due to enhance chemical interactions. When using 1.2 M sulfuric acid, the LTP treatment was found to provide a heavy REE recovery increase from 23% to 53% for the 1.6 SG Float fraction which corresponded to an ash content increase from around 20% to 35%. However, LTP treatment provided no improvement in leaching characteristic for the 2.2 SG sink material in the acid solution, which was likely due to the low organic content. The LTP treatment of the Fire Clay SG fractions provided no improvement in REE leaching characteristic, which may be due to the association with REE minerals such as monazite and REE-associated mineral matter such as clays. High temperature oxidation using temperatures in the range of 600-750OC was found to significantly enhance the leaching characteristics of both the light and heavy REEs for all density fractions in both coarse coal refuse materials. Generally, the increase in light REE recovery for the West Kentucky material was from 15% for the untreated material to 90% for the HTO treated material at 600°C. Recovery values dropped significantly when using temperatures below 600°C and above 750°C. The impact of the excessive temperatures was most likely a result of sintering. For the heavy REEs, HTO significantly improved recovery in the lighter SG fractions, which may be due to an affinity with the organic content. Only limited heavy REE recovery improvement was obtained for the denser SG fractions. Another benefit of HTO is the conversion of calcium-based minerals to a very soluble calcium oxide, which is removable using a mild acid solution, and the conversion of Fe minerals to relatively insoluble minerals such as hematite and magnetite. The pregnant leaching solution produced from the leaching of all SG fractions obtained from both coal sources was treated using two stages of solvent extraction procedurally modeling a rougher-cleaner circuit followed by oxalic acid precipitation. Rare earth oxide products having a purity greater than 97% was produced from nearly all of the SG fractions. A novel process flowsheet was developed based on the selective leaching of heavy REEs that was made possible through LTP treatment of West Kentucky No. 13 coarse refuse followed by salt leaching. In the flowsheet, coal refuse material is ground to a top size of 1mm and then treated by LTP. The treated material is mixed into a 0.1 M ammonium sulfate solution and then filtered to produce a heavy REE enriched pregnant leach solution. The heavy REEs are concentrated by two stages of selective precipitation to produce a heavy REE mix oxalate and then roasted to obtain a heavy REO mix concentrate. The solid residues remaining after salt leaching is dried and then subjected to HTO treatment followed by concentration using selective precipitation and roasting to produce a light REO mix concentrate. An alternative circuit was developed using HTO treatment only to treat 100% of the coarse refuse material followed by selective precipitation and roasting to generate a high purity REO mix product. A techno-economic analysis performed for the treatment of West Kentucky No. 13 and Fire Clay coarse refuse based on a hypothetical 500-tph plant revealed that the net present value for all scenarios was significantly negative. Primary reasons include: i) a substantial decrease in the basket value of the final product relative to the feed and ii) high chemical costs relative to the amount of product generated. The basket value reduction was primarily due to little to no scandium recovery due to poor leaching characteristics when using low acid concentrations as necessitated by the low feed grades. Operating costs were primarily influenced by the chemical costs relative to the kilograms of concentrate produced and feed grade. For the Fire Clay coarse refuse material, the total operating cost was estimated to be $354 per kg of REO produced which was higher than the basket value of the REO product ($276/kg). Approximately 37% of the total cost was the sulfuric acid used during the leaching process. A comparative analysis performed to assess the economic viability of using low-temperature plasma furnaces found that market value enhancement for the generation of both heavy REO and light REO concentrates will be needed in addition to significant technology advancements to reduce the cost of plasma furnaces.
01 COAL, LIGNITE, AND PEAT↗
Physical, Chemical, and Mineralogical Characterizations of MSWI Ash Product and Recommendations for Downstream Processing
The primary objectives of this project are to (1) systematically characterize MSWI ash, and (2) based on characterization findings, design preliminary flowsheets for downstream processing. To achieve these objectives, a total of ten tasks were completed, including sample collection, physical separation tests, liberation tests, synthetic MSWI ash preparation, elemental composition analysis, sequential chemical extraction, mineralogical characterization, pozzolanic activity characterization, thermal stability characterization, processing flowsheet design, TEA and T2M, and project performance reporting. Many useful findings and conclusions were obtained from the exhaustive efforts of this project from several different aspects, including: a) Valuable Metals in MSWI Ash: MSWI ash contains a diverse array of valuable metals. Based on potential recoverable values, the most valuable metals present in MSWI ash include Fe, Ti, Mn, Cu, Zn, V, Co, Ni, Sr, Sn, Ag, Mo, and Sc. Some of these metals have been identified as critical minerals by DOE and DOI, suggesting that MSWI is a promising feedstock for critical mineral recovery. Noticeable graphical and seasonable variations in the valuable metal content of MSWI ash were observed. Nevertheless, it was challenging to discern any clear, definitive patterns for conclusions from those observations. Compared with bottom ash, fly ash contains more volatile metals, such as Zn and Sn, but less nonvolatile metals, such as Fe, Mn, Cu, Zn, Co, and Ni. Mineralogical analyses showed that MSWI ash contains a substantial amount of calcium minerals, such as portlandite, lime, gypsum, and calcite. In addition, it was found that different types of valuable metals often exist in the same particles. b) Physical Separation of MSWI Ash: Both dry sieving and wet sieving were performed on MSWI ash. A notable disparity in the size distribution of the same material was observed when using the two different sieving methods. The disparity is due to the agglomeration of small particles. For the valuable metals investigated, no significant enrichment in a specific size fraction was observed, suggesting that it is challenging to preconcentrate the valuable metals through size fractionation. Due to the presence of ferromagnetic materials, such as Fe, most of the materials reported to the magnetic products obtained by dry magnetic separation. However, the enrichment effect is minimal due to the existence of particle agglomerates. Density separation at a cut-off density of 2.7 SG or higher led to noticeable enrichment of selected valuable metals, particularly Ti. The unburned carbon present in MSWI ash was effectively removed by flotation using diesel as the collector. A novel reagent scheme, Na2S plus cationic collectors, that can efficiently beneficiate nonferrous metals plus Co was developed. c) Liberation Tests: The particle size of MSWI ash was effectively reduced by grinding, and as a result, the encapsulated valuable metal particles (if any) were liberated to a certain degree. However, particle size reductions did not noticeably enhance the beneficiation performance using the physical separation methods, primarily due to the inefficiency of these methods in processing fine particles and/or a possibility that insufficient liberation is not a limiting factor for achieving satisfactory physical separation performance. Valuable metals were classified into water leachable, ion-exchangeable, acid soluble, reducible, oxidable, and insoluble forms. It was found that the distributions in the different categories, i.e., the occurrence modes of the valuable metals, were not affected by the particle size. d) Leaching Characteristics of Metals from MSWI Ash: Most of the valuable metals were extracted from the fly ash samples when using 1 M HCl or HNO3 as the lixiviant. The leaching reaction is a very fast process, which can reach equilibrium within the first 5 min. The releasing of Co, Ni and Ag are sensitive to leaching temperature, a higher recovery value could be obtained when using relatively higher leaching temperatures. The leachability of the valuable metals present in MSWI bottom ash is relatively lower than that of fly ash. Leaching recoveries increased with elevations in the acid concentration. Relatively high leaching recoveries were obtained for REEs, Mn, Co, Ni, Cu, and Zn using 1 M HCl or HNO3 as the lixiviant. Elevations in the reaction temperature noticeably increased the leachability of the valuable metals, whereas the leachability was barely influenced by oxidizing and reducing agents. Similar to fly ash, leaching valuable metals from bottom ash is a rapid process, with most of the leaching reaction completed within the first 5 minutes. e) Combusted iPhones: The original structure of iPhones was remained after treating at 400 ºC and 600 ºC, while after being treated at 800℃, the screen bent, and the back cover of iPhone melted. Increasing the combustion temperature to 1000℃, the screen scattered, and most of the components turned into ashes. Combustion enhanced the leachability of REEs, while the leachability of the other valuable metals, except for Zn, was barely affected. Most of the REEs present in the original iPhones occurred as oxidizable forms. With elevations in the combustion temperature up to 600 ºC, the oxidizable REEs were transformed to acid soluble forms. However, further elevations in temperature resulted in decreases in the acid soluble fraction and corresponding increases in the reducible and oxidizable forms. Additionally, combustion temperature also significantly altered the occurrence modes of other metals present in the iPhones. f) Synthetic MSWI Ash: It was found that in the absence of hydrogen peroxide, all the elements except for Si were leached to certain degrees. It is noteworthy that approximately 80% of Zn was leached with 1.2 M HCl. When hydrogen peroxide was added to the reaction system, noticeable increases in the leaching recovery of Fe, Mn, Co, Ni, and Cu were observed. The leaching recovery of Al and Si was barely affected by adding hydrogen peroxide. These results suggested that the majority of Zn in the synthetic MSWI ash existed as metal oxide, a portion of Fe, Mn, Co, Ni, and Cu existed as metal oxide, and Al and Si are associated with glasses which are difficult to leach. Additionally, the remaining Fe, Mn, Co, Ni, and Cu in the metallic form were efficiently oxidized in the presence of hydrogen peroxide. g) Pozzolanic Activity and Thermal Stability of MSWI Ash: MSWI fly ash has higher pozzolanic activity compared to the bottom ash sample, which indicates that the fly ash sample consumed more portlandite because of its smaller particle size as reactivity fundamentally relates to reaction surface area. However, after the recovery of valuable elements, the pozzolanic activity of both the valuable elements fraction and the less valuable elements-rich products decreased significantly, which means that the valuable elements recovery lowers the Ca(OH)2 consumption, thus leading to the low activity of SCM. h) Flowsheet Design for Metal Recovery from MSWI Ash: Based on the results of the comprehensive physical separation and acid leaching tests, circuits that enable the beneficiation of the valuable metals were developed. In these circuits, the valuable metals are recovered into nonferrous, ferrous, and other valuable metal concentrates, which are processed separately in the acid leaching step. The subsequent separation and purification steps are simplified due to the physical beneficiation step. In addition, the overall recovery cost is reduced since physical beneficiation is much cheaper compared with chemical processing. Using different technologies, such as selective precipitation and solvent extraction, a comprehensive hydrometallurgical circuit was designed, and compounds of Cu, Zn, Mn, Co, and Ni with a purity close to or even higher than 95% were successfully generated.
36 MATERIALS SCIENCE↗
Vadose Zone Soil Flushing for Chromium Remediation: A Laboratory Investigation to Support Field‐scale Application
Cr(VI) flushing from the vadose zone to the groundwater (with subsequent Cr(VI) removal in groundwater by pump-and-treat system) is a promising remedial technique that has recently been used at field scale. This laboratory study was conducted to provide the technical basis to design a field soil flushing strategy. The objectives were to (1) quantify the relationship between sediment Cr(VI) and Cr(III) mass and release rates and subsequent Cr(VI) leaching; (2) investigate different methodologies to maximize Cr(VI) leaching, and (3) investigate methods to minimize leaching of remaining residual Cr. Characterization of Cr-contaminated sediments (Hanford Site, WA) exhibited Cr(VI) showed that leach rates that were correlated to different Cr surface phases. Sediments with low leachable Cr(VI) (<2 μg/g) leached Cr rapidly, so slow infiltration of water in a single pulse was sufficient to leach most Cr. In contrast, sediments with high Cr (2 to 200 μg/g) released some Cr(VI) quickly but 10 to 50% Cr(VI) slowly (tens to hundreds of hours). Efficient unsaturated leaching of these sediments required a different infiltration strategy that includes: multiple slow leach pulses with time between flushing cycles; the use of a surfactant to increase Cr leaching from low-permeability zones, and the use of a reductant (Na-dithionite or Ca-polysulfide) in the final leach water was highly effective at decreasing residual Cr leaching. This study clearly demonstrated that the methodology of basing laboratory Cr flushing on parameters such as Cr release mass and rates could be used to improve the efficiency of soil flushing at field scale.
Accelerated Cr Flushing↗
Demonstration of Scaled-Production of Rare Earth Oxides and Critical Materials from U. S. Coal-Based Sources (Final Report)
The project objective was to demonstrate scaled production of high purity rare earth oxides (REO), nominally exceeding 90% grade, from coal refuse sources using innovative technologies that reduce cost and improve environmental outcomes relative to traditional rare earth processing technologies. The project utilized a critical material pilot plant constructed and tested as part of a previous U.S. Department of Energy project. Target performance criteria was a 50% reduction in production cost based on previous optimum values, 150% increase in recovery and greater than 2% concentrates of rare earth oxides, cobalt and manganese. Concentrate production goals were to produce a rare earth mix oxide product at a rate of 200 grams per day having a minimum purity of 50% as well as products of cobalt and manganese having a minimum purity of 2%. A previous pilot plant investigation identified acid cost as the major contributor to an operating cost that made the recovery of rare earth and other critical metals from bituminous coal sources economically challenging. As such, acid cost reduction was a major target using bio-oxidation reactors to produce sulfuric acid from naturally occurring coal pyrite. Based on laboratory data, a bio-oxidation circuit was designed for the pilot plant to produce 7.5 l/min of acid using two 11-m3 (3000 gallon) reactors equipped with 40 hp aerators for air dispersion. The pilot-scale tests revealed that acid concentration equivalent to as high as 1.0 M sulfuric acid could be continuously produced. However, the bio-acid contained exceptionally high iron concentrations, which complicated downstream processing of the pregnant leach solution (PLS). A TEA of the bio-oxidation circuit showed that production cost was approximately $0.13 per kg acid equivalent if produced using a four-day retention time in the reactors. This value represents a 48% reduction from that of purchased bulk sulfuric acid. Calcination (or roasting) studies were conducted on coarse refuse from West Kentucky No. 13 and Fire Clay coal seam sources. The test results revealed the potential to increase recovery by nearly 100% using temperatures between 500°C to 700°C with light REE recovery value being the most improved. Acid baking of the calcined products using sulfuric acid at 250°C increased heavy rare earth recovery from around 40% to 80% while decreasing the acid requirements by over 50%. The existing pilot plant was upgraded for the pilot scale demonstration of REE and CM recovery. The primary feedstocks were West Kentucky No.13 and heap leach pregnant leach solution (PLS) while a secondary feedstock was a lignite waste material from a construction sand operation. The pilot scale operation started with PLS generation through leaching followed by iron and aluminum removal, nominally at 3.3 and 4.5 pH, respectively. Leaching lixiviants used for the test were industrial grade sulfuric acid or bio-acid generated at the pilot scale facility. For most of the tests, the solid feed rate was 200 lb/hr whereas lixiviant was added at 2 gpm to provide an optimal residence time of 45 minutes. Following the contaminant removal step, several different process schematics were tested with the goal of maximizing REE recovery and purity. In the first test, direct processing of aluminum precipitation raffinate for REE recovery using oxalic acid at pH 1.5 was investigated. Overall REE recovery was approximately 45%. Unfortunately, elevated calcium content in the PLS significantly impacted the RE-Oxide product grade. Similarly, high calcium content decreased both the product purity and grades of CM products. As such, a new flowsheet was tested with the same initial process schematic but different precipitation stages for REEs and CMs at pH 6.0 and 9.0, respectively. It was noted that the overlapping precipitation behavior of Co, Ni and Zn with REEs limited the applicability of this process schematic. While this change increased the RE-Oxide grade from 36% in the first test to 87%, the loss of critical metals to the REE cake and bypass of the REEs to the CM cake significantly impacted the recovery of both REEs and CMs. Therefore, the modified process flowsheet combined oxalic acid precipitation stage raffinate and redissolved CM cake filtrate to maximize both the recovery and purity of the products. Consequently, a RE-Oxide product with 85% purity and CM cakes with over 19% Co, 38% Ni, 14% Zn and 9% Mn content were generated with significantly higher recoveries. While the modified process flowsheet improved recoveries and grades, elemental losses observed in separate precipitation and redissolution losses inspired the adaptation of a single precipitation stage at pH 9.0 for both REEs and CMs. This change was anticipated to maximize the REE recovery while minimizing the costs associated with separated redissolution and processing stages. As expected, REE recovery in this new circuit arrangement was over 56% with a product grade of over 87% RE-Oxide content. Similarly, Co, Ni, Mn, and Zn recoveries of 54%, 40%, 67%, and 66%, respectively, were achieved. Unfortunately, the elevated calcium content present in the solution due to its precipitation at pH 9.0 caused a decrease in the CM cake quality. Therefore, the final process flowsheet involved the addition of calcium oxalate precipitation following the oxalic acid precipitation stage, which effectively eliminated calcium contamination of the CM products. Finally, the pilot scale experiments conducted using bio-acid achieved comparable REE leaching recoveries to the conventional sulfuric acid leaching. Elevated iron concentration in the solution caused the co-precipitation of REEs with the iron and aluminum cake, resulting in the REE losses. Furthermore, elevated iron content bypassing the iron and aluminum precipitation stages contaminated the metal sulfide and manganese cake, respectively. A techno-economic analysis was performed based on a commercial facility capable of treating 500 tph of coal-based material. The production cost for West Kentucky No. 13 coarse refuse material ranged from approximately $500-$700/kg of total rare earth oxide whereas the lignite source had significantly lower production cost of $100-$300/kg. The significant difference in cost was due to the easier leaching characteristics of the lignite material and the higher feed concentrations. All process scenarios resulted in a negative net present value (NPV). For the lignite feedstock, laboratory REE leach recovery values were about 30% higher than the pilot plant data. Using the lab leach results, a positive net present value was achieved and the production cost decreased from $100-$300 $/kg to less than $150/kg of total REO.
01 COAL, LIGNITE, AND PEAT↗
Characterization and recovery of rare earth elements and other critical metals (Co, Cr, Li, Mn, Sr, and V) from the calcination products of a coal refuse sample
Here, the impacts of calcination on the leaching recovery of rare earth elements (REEs) and other critical metals from a Baker seam coal refuse material were investigated. Acid leaching test results showed that leaching recovery of the REEs, especially the light REEs, was increased significantly from around 24% to 80–90% after calcination under 600–750 °C for two hours. By considering the major element leaching kinetics and the mineralogical changes of the dominant minerals caused by calcination, two mechanisms were proposed for the improved leaching performance: (1) decomposition of difficult-to-dissolve REE-bearing minerals into soluble forms; and (2) liberation of REE-bearing minerals encapsulated in clays after calcination due to the dehydration and disintegration of the layered clay structure. Sequential chemical extraction tests showed that a significant fraction (around 75%) of the REEs in the raw material were present in an insoluble form. After calcination under proper temperatures (600–800 °C), the majority of the insoluble light REEs (LREEs) were converted into soluble forms. Changes in the modes of occurrence of the LREEs and phosphorous based on leaching rate and sequential leaching correlated well as a function of the calcination temperature. This finding in addition to SEM-EDX results tends to indicate the potential that the LREEs occur primarily as rare earth phosphates such as monazite in the untreated coal refuse material. The acid leaching and sequential chemical extraction results show that co-extraction of REEs and other critical metals such as Co, Mn, and Li is possible using a calcination-acid leaching procedure.
01 COAL, LIGNITE, AND PEAT↗
Economic Extraction, Recovery and Upgrading of Rare Earth Elements from Coal-Based Resources
The overall objective of this project is to produce a rare earth elements product with greater than 8% rare earth elements from coal-based resources using an economically viable and environmentally benign processing methodology. The objective of the coal processing work will be to acquire appropriate coal feed and produce a viable clean coal product, a concentrated REE feed stream, and a pyrite stream (where available) that can be used to enhance leaching. The objective of the bio-oxidation portion of the project is to utilize bacteria to oxidize sulfide minerals and ferrous ions to ferric ions to accelerate leaching and remove the sulfides to prevent future acid mine drainage and related liabilities. The objective of the solution conditioning work is to remove iron and control leaching solution conditions to facilitate extraction of REEs without significant thorium extraction. The objective of the column leaching work is to engineer/model and simulate heap leaching to demonstrate low-cost extraction of REEs from coal-based resources. The objective of the solvent extraction work is to selectively extract and recover through controlled stripping of REEs that are solubilized through leaching. The objective of the precipitation portion of the project is to recover the REEs from the solvent extraction stripping solutions and dry them to achieve the final product of rare earth elements that is > 8 % REEs. The objective of the technical, economic, and environmental analysis is to determine the overall viability of the processing approach that is demonstrated in this project. This project encompasses a range of technologies that are currently in industrial practice that are applied and engineered to produce rare earth element product from coal feed sources. The potential sources include large coal feed resources in active coal mines in addition to coal waste. This technology involves the utilization of advanced coal processing technology that can be used to deliver clean coal for the market as well as rare-earth-element-bearing non-coal rock that is of the correct size for heap leaching applications in addition to providing concentrated sulfide minerals (for mid to high sulfur coals) for cleaner coal and for enhanced bio-oxidation to accelerate leaching of REEs from the non-coal rock. The removal of the sulfide minerals cleans the coal, accelerates subsequent REEs extraction, and it removes the future potential for most acid-rock drainage. (For low sulfur coals some pyrite may be purchased if needed.) The processing method also utilizes bio-oxidation to enhance ferric ion production to enhance leaching, while also consuming the sulfide mineral and its associated environmental liability. The project includes a technical, economic, and environmental analysis to facilitate an appropriate assessment of commercial viability for this processing technology.
01 COAL, LIGNITE, AND PEAT↗
Extraction Kinetics of Rare Earth Elements from Ion-Adsorbed Underclays
Citric acid has been identified as an environmentally sustainable organic acid capable of leaching up to ~30% of easily accessible REEs from underclay material. An analysis of the leaching profiles was performed to discern the reaction rates, extraction efficiencies, and potential leaching mechanisms of REEs and cations of interest from ion-adsorbed underclays. The initial leaching stage follows a slow intraparticle diffusion mechanism followed by a second stage controlled by a mixed diffusion regime. The leaching profiles of Ca and P were similar to those of REEs, suggesting that REEs are most likely derived from mineral surfaces such as hydroxyapatite or crandallite rather than predominately from underclays. Fitting to a modified diffusion control model found diffusion-controlled leaching to be the primary mechanism whereas non-diffusive mechanisms made up about 22% of the extracted REEs. Gangue cations associated with underclays had less non-diffusive leaching than REE species, indicating that their leaching kinetics may be dominated by diffusion from within the material or potentially from product layer formation. Fitting to Boyd plots further indicated that REEs were leached following intraparticle diffusion control. These results have important implications for the development of more efficient and sustainable methods for extracting REEs or critical minerals from alternative feedstocks.
Prem, Priscilla↗
AN ACID BAKING APPROACH TO ENHANCE RARE EARTH ELEMENT RECOVERY FROM BITUMINOUS COAL SOURCES
Rare earth elements (REE) are a group of 17 elements typically classified as light and heavy rare earth elements, which play a crucial role in developing the latest technologies for energy, defense, and medical sectors. Even though REEs have been found in more than 200 minerals, only bastnaesite, monazite and xenotime are commercially exploited for REE extraction. However, the recent exponential increase in REE demand has spurred countries such as the United States into research for the extraction of REEs from secondary sources such as coal, acid mine drainage, and coal ash. Several coal sources (e.g., Fire Clay seam coal) across the United States have been identified to contain elevated concentrations of rare earth elements (>600 ppm), and various researchers have investigated the feasibility of both physical and hydrometallurgical extraction techniques for rare earth concentration and subsequent extraction. However, both physical and direct leaching were concluded to be inefficient for RE beneficiation and extraction due to low recoveries. Alternatively, thermal treatment provides promising means for RE recovery from bituminous coal sources. The positive impact of thermal treatment/calcination was established to be due to the decarbonization and dehydroxylation of the clays, which released entrapped rare earth elements within the dominant minerals and converted them into an acid-soluble form. Nonetheless, the improvement in recovery was limited to the light REEs (LREEs) with an insignificant increase in the heavy REEs (HREE). It was demonstrated that the light and heavy REEs in the material were associated with difficult-to-leach minerals such as monazite, xenotime, and zircon, which were not decomposed by simple calcination due to their high thermal stability. Hence, roasting in the presence of chemicals was necessary to ensure the decomposition of those REE-containing minerals. As such, this study was focused on the acid baking treatment of bituminous coals with an aim to enhance REE recovery, especially HREEs. Based on the presence of REE minerals like monazite and xenotime, three pre-leach treatment methods, i.e., 1) roasting, 2) direct acid baking, and 3) acid baking after roasting were investigated. Roasting tests at 600 ⁰C revealed that the recovery of light REEs (LREEs) was enhanced while the recovery of HREEs remained relatively unaffected. LREE and HREE recovery values of 38.3% and 21.3%, respectively, were achieved using a 50 g/L sulfuric acid solution at 5% solid concentration and a solution temperature of 75 ⁰C for 2 hours. Comparatively, direct acid baking at 250 °C provided substantial increased LREE and HREE recovery values to approximately 49.4% and 53.0%, respectively, using an equivalent acid dosage. Recoveries were maximized to 77.0% and 79.6% for LREE and HREE, respectively, by roasting followed by acid baking. Similar results were obtained from the treatment of a second bituminous coal source. Due to strong correlations between REE and Al recovery values, tests were performed on kaolinite and illite, which were prominent clay minerals within the source coals. These experiments revealed that the REE recovery improvements were likely a result of dehydroxylation of clays and subsequent release and decomposition of REE-bearing minerals such as monazite, xenotime and zircon. Subsequently, a parametric study was conducted to identify the impact of acid baking parameters on rare earth element recovery. The factors investigated using a three-level statistical experimental program were acid baking time, acid solution concentration, baking temperature, and acid solution-to-solids ratio which were found to significantly impact REE and contaminant element (Al, Fe, and Ca) recovery. An increase in baking temperature up to around 250 ⁰C improved the light and heavy REE recovery values by more than 50 absolute percentage points relative to performances achieved when direct leaching. As aforementioned, acid baking was needed to both decompose the clay minerals and liberate the REE minerals, which allowed access for the acid to solubilize the REEs. Acid concentration of the solution used for acid baking was studied as a means of minimizing the amount of acid needed to achieve a target REE recovery. However, thermo-gravimetric and differential scanning calorimetry analysis (TGA-DSC) of sulfuric acid under oxidizing atmosphere revealed that the addition of water decreased the evaporation temperature, which explains the lower REE recovery values obtained when using acid concentrations less than 100%. Using pure sulfuric acid at an acid-to-solid ratio of 0.8:1 resulted in recovery values of around 70% for both LREEs and HREEs. The decomposition reaction time was relatively quick with 65% of the TREEs recovered within the first 10 minutes. Following the identification of optimum operating conditions through the parametric study, a systematic leaching study was carried out to examine the impact of leaching parameters, such as solid-to-liquid (S/L) ratio, temperature, and time, on REE recovery using acid baking conditions of 1:1 acid to solids ratio at 250 °C for 30 minutes. The solid-to-liquid ratio was varied from 1-20% by weight at 25 °C, 50 °C, and 75 °C solution temperatures. The results indicated that reaction time and solution temperature considerably impacted the recovery of heavy and light REEs. Interestingly, LREE recovery reduced from 68% at 5% S/L to 58% at 20%S/L, whereas HREE recovery of 78% remained unaffected. The decrease in the LREE recovery was determined to be due to La and Ce precipitation, likely through isomorphic substitution with calcium in gypsum. The kinetic data indicated that 67% LREE and 77% HREE recovery could be obtained within the first 15 minutes of the reaction, suggesting fast reaction kinetics. Furthermore, raising the solution temperature from 25 °C to 75 °C increased the LREE and HREE recovery from 60% and 32% to 67% and 79%, respectively. The kinetic modeling results demonstrated that the rate-limiting step in the LREE dissolution was diffusion and chemical reaction, whereas the HREE extraction was controlled by only chemical reaction. The leaching study concluded that using 20% S/L at 75 °C for 15 minutes maximized LREE and HREE recovery. The lab-scale precipitation study showed that Fe and Al in solution could be removed at pH 4.5 followed by REE precipitation at pH 6.0 using 6 mol/L NaOH. Finally, the bench-scale data was used to develop a process flowsheet for REE recovery from low-grade bituminous coal sources using acid baking. Finally, based on the proposed flowsheet, a concentrated RE-cake obtained through selective precipitation at pH 6.5 was re-leached using HCl at pH 1.5. The resultant leachate was used to identify the impact of various operating parameters on REE recovery and purity with an aim to maximize REE precipitation efficiency while minimizing the oxalic acid dosage. The operating parameters for this investigation were oxalic acid dosage, iron (III) contamination, solution pH and temperature. The resultant model suggested that oxalic acid dosage and reaction pH are the most significant factors for the REE precipitation efficiency, followed by the interaction of oxalic dosage and Fe concentration. Test results indicated that increasing the oxalic acid concentration from 0g/L to 80g/L improved the REE precipitation efficiency from approximately 4.2% to 95.0%. Furthermore, raising the solution pH from 0.5 to 2.5 considerably enhanced the precipitation efficiency from 0.0% to 98.9%. A solution temperature elevation decreased REE recovery, which indicated an exothermic reaction between REEs and oxalate anions. Finally, a high level of Fe contamination adversely impacted REE precipitation efficiency. The speciation analysis revealed that the dominant iron species in the solution system were Fe-(C₂O₄)₃³⁻, Fe-(C₂O₄)²⁻, and Fe-(C₂O₄)⁺, which consumed the majority of the oxalate anions
rare earth elements, acid baking, high-temperature↗
On silica's roles in controlling americium migration in contaminated sediments
Two studies in the early 1980s described the leaching behavior of americium (Am) disposed as part of acidic high-salt processing wastes from the Hanford Site’s Plutonium Finishing Plant to nearby ground sediments. Here these batch leach experiments showed that the Am concentrations followed a linear log [Am] versus pH relationship with a slope of –1. Column leach experiments in the second study, however, did not follow this relationship and only ~30% of the americium desorbed even after extensive column leaching. Here, the 1980s research is re-examined along with previously unpublished information and, in light of recent published work, a plausible mechanism is proposed to explain these phenomena. Amorphous silica in the contaminated sediments is postulated to be the substrate responsible for both the exchangeable Am available for leaching and the retained low-leachable Am made evident in the column leach experiments. The exchangeable Am 3+ in the contaminated sediment leach experiments behaves with pH dependence similar to that observed for uptake onto amorphous silica of sodium (Na + ), calcium (Ca 2+ ), barium (Ba 2+ ), cadmium (Cd 2+ ), uranyl (UO 2 2+ ), ferric (Fe 3+ ), chromic (Cr 3+ ), cupric (Cu 2+ ), plumbous (Pb 2+ ), uranium(IV) (U 4+ ), plutonium(IV) (Pu 4+ ), zirconium (Zr 4+ ), analogue lanthanide (gadolinium, Gd 3+ , europium, Eu 3+ , and lutetium, Lu 3+ ) and curium (Cm 3+ ) ions as well as other studies with Am 3+ . Correspondingly, the residual low-leachable Am 3+ revealed in the column leach experiments is attributed to incorporation of Am 3+ within amorphous silica by dynamic Am 3+ sorption and silica precipitation processes.
58 GEOSCIENCES↗
Understanding the Degradation of La 1−x Sr x FeO 3−δ (0 ≤ x ≤ 1) Perovskite Oxides during the Oxygen Evolution Reaction in Alkaline Solution
Perovskite oxides are an emerging class of highly active catalysts for the oxygen evolution reaction (OER); however, their electrochemical stability remains poorly understood. Here, we report a systematic evaluation of the OER activity and stability of La 1−x Sr x FeO 3-δ perovskites in 1 M KOH. Their initial OER activity first increases with increasing Sr content (fromx= 0 to 0.8), and then decreases when the Sr content is increased to 1. Their stability evaluated by monitoring the element leaching from the electrodes show that La does not leach at a detectable rate, but Sr and Fe leach substantially. The leaching of Sr occurs at similar rates under open circuit potential (OCP) and OER potential, suggesting a nonelectrochemical dissolution process. The leaching of Fe is, however, strongly dependent on the electrode potential. More Fe leaching is observed under the OER potential than OCP. Additionally, the electrode with higher initial OER activity leaches more Fe. These results indicate that OER facilitates the dissolution of Fe from the electrode. The leaching of Fe, in turn, is considered responsible for the activity loss of La 1-x Sr x FeO 3−δ during OER. This study brings new insight into the degradation mechanism of La 1-x Sr x FeO 3−δ and their related perovskite oxides during electro-oxidation processes.
Electrochemistry↗
Strategies to Recover Easily-Extractable Rare Earth Elements and Other Critical Metals from Coal Waste Streams and Adjacent Rock Strata Using Citric Acid
Maximum rare earth element (REE) extractability from a sample of Middle Kittanning coal seam underclay was previously demonstrated at ~30% of the total REE content with a citric acid solution (Montross et al., 2020). This report further refines the mechanisms of citrate application for leaching of coal seam underclays using an organic acid lixiviant and evaluates strategies to begin to scale this process to industrially relevant volumes. This study evaluates the suitability of citrate leaching solutions for downstream recovery and separation of the REEs via oxalic acid precipitation. The applicability of citric acid solutions to recover ion-adsorbed metals from clay surfaces was evaluated by leaching a prepared sample of kaolinite with known amounts of ion-adsorbed lanthanum. Citrate systematics was further defined through experiments varying the solution pH. Up-scaling investigations proceeded with percolation leaching columns and stirred tank reactors. Two embodiments were run in percolation leaching: a continuous flow-through process and a multi-stage saturated process. The results showed that in certain limited scenarios, low pH citrate solutions could be used to target REEs mineralized within secondary calcium phosphate minerals. Results from the upscaling attempts encountered difficulties in extraction efficiencies. For the percolation columns, hydrodynamic flow was insufficient to recover the extractable REE content fully. For the stirred barrels, mixing inefficiencies likely inhibited the extractability of the REEs. Oxalic acid precipitation experiments also highlighted difficulties in downstream operations. The REEs were removed from the citrate solution with high efficiency, but were accompanied by equally efficient removal of calcium in orders of magnitude of greater amounts. Overall, for citrate leaching to be effective, future studies will have to: 1) carefully consider the feedstock application (i.e., likely better suited for secondary mineralization of calcium phosphate minerals), and 2) overcome the technical barriers of large-scale processing, such as inefficient hydrodynamic regimes due to small crush sizes needed and difficulties in downstream separation and purification. Alternatively, citrate solutions may be better recommended as leaching amendments to other leaching solutions to enhance REE complexation in the solution.
01 COAL, LIGNITE, AND PEAT↗