In modern coal preparation plants, solid-solid and solid-liquid separation processes used to treat fine coal are least efficient and most costly operations. For example, field studies indicate that the froth flotation process, which is normally used to treat minus (-0.2 mm) fine coal, often recovers less than 65 to 70% of the organic matter in this size range. Fine coal separation processes are also inherently less effective in removing pyrite than that of coarse coal separations. Moreover, while fines may represent 10% or less of the total run-of-mine feed, this size fraction often contains one-third or more of the total moisture in the delivered product. In order to address these issues, several multistage coal processing circuits were set up and experimentally tested to demonstrate the potential improvements in fine coal upgrading that may be realistically achievable using an "optimized" fine coal processing flowsheet. On the basis of results obtained from this research, engineering criteria was also developed that may be used to identify optimum circuit configurations for the processing different fine coal streams. In the current study, several fine coal cleaning alternatives were evaluated in laboratory, bench-scale and pilot-scale test programs. Fine coal processes compared in the first phase of this work included spirals, water-only cyclones, teeter-bed separators and froth flotation. The performance of each technology was compared based on separation efficiencies derived from combustible rejection versus ash rejection plots. The resulting data was used to identify size ranges most appropriate for the various alternative processes. As a follow-up to this effort, a second phase of pilot-scale and in-plant testing was conducted to identify new types of spiral circuit configurations that improve fine coal separations. The experimental data from this effort indicates that a four-stage spiral with second- and fourth-stage middlings recycle offered the best option for improved separation efficiency, clean coal yield and combustible recovery. The newly developed spiral circuitry was capable of increasing cumulative clean coal yield by 1.9 % at the same clean coal ash as compared to that of achieved using existing conventional compound spiral technology. Moreover, the experimental results also proved that slurry repluping after two turns is not effective in improving separation performance of spiral circuits. The third phase of work conducted in this study focused on the development of methods for improving the partitioning of pyrite within fine coal circuits. The investigation, which included both laboratory and pilot-scale test programs, indicated that density-based separations are generally effective in reducing sulfur due to the large density difference between pyrite and coal. On the other hand, the data also showed that sulfur rejections obtained in froth flotation are often poor due to the natural floatability of pyrite. Unfortunately, engineering analyses showed that pyrite removal from the flotation feed using density separators would be impractical due to the large volumetric flow of slurry that would need to be treated. On the other hand, further analyses indicated that the preferential partitioning of pyrite to the underflow streams of classifying cyclones and fine wire sieves could be exploited to concentrate pyrite into low-volume secondary streams that could be treated in a cost effective manner to remove pyrite prior to flotation. Therefore, on the basis of results obtained from this experimental study, a combined flotation-spiral circuitry was developed for enhanced ash and sulfur rejections from fine coal circuits. Finally, the fourth phase of work conducted as part of this investigation focused on evaluating a new mechanical, non-thermal dewatering process called Nano Drying Technology (NDT"). Experimental results obtained from bench-scale testing showed that the NDT" system can effectively dewater fine clean coal products from more than 30% surface moisture to single-digit moisture values. Test data obtained using a pilot-scale NDT" plant further validated this capability using a continuous prototype facility. It was also observed that, unlike existing fine coal dewatering processes, the performance of the NDT" system is not constrained by particle size. / Ph. D.

In modern coal preparation plants, solid-solid and solid-liquid separation processes used to treat fine coal are least efficient and most costly operations. For example, field studies indicate that the froth flotation process, which is normally used to treat minus (-0.2 mm) fine coal, often recovers less than 65 to 70% of the organic matter in this size range. Fine coal separation processes are also inherently less effective in removing pyrite than that of coarse coal separations. Moreover, while fines may represent 10% or less of the total run-of-mine feed, this size fraction often contains one-third or more of the total moisture in the delivered product. In order to address these issues, several multistage coal processing circuits were set up and experimentally tested to demonstrate the potential improvements in fine coal upgrading that may be realistically achievable using an "optimized" fine coal processing flowsheet. On the basis of results obtained from this research, engineering criteria was also developed that may be used to identify optimum circuit configurations for the processing different fine coal streams. In the current study, several fine coal cleaning alternatives were evaluated in laboratory, bench-scale and pilot-scale test programs. Fine coal processes compared in the first phase of this work included spirals, water-only cyclones, teeter-bed separators and froth flotation. The performance of each technology was compared based on separation efficiencies derived from combustible rejection versus ash rejection plots. The resulting data was used to identify size ranges most appropriate for the various alternative processes. As a follow-up to this effort, a second phase of pilot-scale and in-plant testing was conducted to identify new types of spiral circuit configurations that improve fine coal separations. The experimental data from this effort indicates that a four-stage spiral with second- and fourth-stage middlings recycle offered the best option for improved separation efficiency, clean coal yield and combustible recovery. The newly developed spiral circuitry was capable of increasing cumulative clean coal yield by 1.9% at the same clean coal ash as compared to that of achieved using existing conventional compound spiral technology. Moreover, the experimental results also proved that slurry repluping after two turns is not effective in improving separation performance of spiral circuits. The third phase of work conducted in this study focused on the development of methods for improving the partitioning of pyrite within fine coal circuits. The investigation, which included both laboratory and pilot-scale test programs, indicated that density-based separations are generally effective in reducing sulfur due to the large density difference between pyrite and coal. On the other hand, the data also showed that sulfur rejections obtained in froth flotation are often poor due to the natural floatability of pyrite. Unfortunately, engineering analyses showed that pyrite removal from the flotation feed using density separators would be impractical due to the large volumetric flow of slurry that would need to be treated. On the other hand, further analyses indicated that the preferential partitioning of pyrite to the underflow streams of classifying cyclones and fine wire sieves could be exploited to concentrate pyrite into low-volume secondary streams that could be treated in a cost effective manner to remove pyrite prior to flotation. Therefore, on the basis of results obtained from this experimental study, a combined flotation-spiral circuitry was developed for enhanced ash and sulfur rejections from fine coal circuits. Finally, the fourth phase of work conducted as part of this investigation focused on evaluating a new mechanical, non-thermal dewatering process called Nano Drying Technology (NDT™). Experimental results obtained from bench-scale testing showed that the NDT™ system can effectively dewater fine clean coal products from more than 30% surface moisture to single-digit moisture values. Test data obtained using a pilot-scale NDT™ plant further validated this capability using a continuous prototype facility. It was also observed that, unlike existing fine coal dewatering processes, the performance of the NDT™ system is not constrained by particle size. / Ph. D.

In order to develop a better understanding of the role of the froth phase in coal flotation, froth stability measurements have been conducted using a continuous flotation cell. The results indicate that the mass flow rate of coal or mineral matter reporting to the product is linearly related to the flow rate of water reporting to the product. This relationship has been used to distinguish the relative contributions of hydrophobic attachment or hydraulic entrainment to the total product recovery. Simple mathematical expressions have also been developed to characterize the cleanability of various coals. A mathematical model based on first principles has been developed to provide additional insight into the complex relationship between the various operating parameters of a flotation cell and the froth behavior. The predicted results compare favorably to actual flotation test data for most conditions. In addition, a simple method was developed to study the fundamental interactions of various frothing agents with coal particles. The results of these studies indicate that the adsorption behavior of frothers in the presence of coal depends on both the physical properties of the frothing agent and the coal. / Master of Science

Conventional drinking water treatment consists of a coagulation, flocculation, gravity separation, filtration and disinfection processes each working individually but also as an interdependent system. One of the main reagents used for drinking water treatment are coagulants that destabilise the suspended particles which results in the formation of flocs. For many years, the coagulant of choice was aluminum sulphate, also know as alum. Alum has slowly been replaced by new coagulants, such as polyalumium sulphates and polyaluminum chlorides, because they yield more consistent plant performance than with alum over the wide temperature range experienced by Canadian treatment plants. Recent research has determined that the alum solubility envelop varied significantly in terms of pH range with temperature, thus cold temperature performance may be improved by adjusting the coagulation pH. Dissolved air flotation (DAF) is now used at some water treatment plants to replace sedimentation because it is much more compact than gravity settling, and it is somewhat better than sedimentation for the removal of algae, organics and operation in cold temperatures. The objective of this thesis is to help operators and managers of drinking water treatment plants incorporating DAF by: a) investigating the cold water turbidity removals of DAF systems using alum, the most economical coagulant; and b) investigating the impact of DAF saturator pressure on the bubble sizes produced and floc removal. This first initiative is based on fairly recent research on the impact of pH on the cold-temperature aluminum solubility. It uses this knowledge about the impact of pH to evaluate DAF treatment of Ottawa River water in cold-water conditions using DAF batch tests. The effect of pH against final turbidity at cold temperatures was first evaluated by increasing the pH of the coagulated water, the higher pH helped attain good turbidity removals. For the coagulant dose tested, good turbidity removals were observed for both warm and cold waters at nearly the same pOH conditions. At room temperature the turbidity removals increase with both increasing flocculation G and flocculation time. While at cold temperatures, when aluminum flocs are known to be much more fragile, the turbidity removals appear to be independent of G and GT. The second initiative studied the relationship between floc size and bubble size in DAF systems by changing the DAF saturator pressure. Increasing the saturator pressure did not significantly decrease the mean bubble size. The flocs attach to bubbles that were significantly larger than the bubbles. The assessment of DAF efficiency based on the unitized effluent floc distribution proved inconclusive, it may be possible that the conditions resulting with the larger mean effluent floc size has a greater removal efficiency since it began with a smaller fraction of small flocs entering the flotation stage.

Boliden’s mineral processing plant in Garpenberg wanted to investigate their zinc-regrinding circuit. The re-grinding circuit had performed subpar when running as the total recovery of zinc was lowered and problem with the dewatering of the final concentrates was also noted. Therefore, the zinc re-grinding circuit is currently not being used. Furthermore, it was noted during investigations about silicate depressant that the zinc circuit had problems with coarse sized sphalerite particles locked with silica. Which could perhaps be liberated with the re-grinding circuit. The purpose was therefore to investigate how a re-grinding step could be used to liberate locked sphalerite particles from gangue. The practical part of the thesis involved lab-scale SMD-mill, re-grinding of the scavenger concentrate and cleaner tailings which are the two streams being fed to the SMD-mill in the plant. The re-grind was done in three fractions: bulk, +90µm and +125 µm fraction with subsequent lab-scale flotation afterwards to identify how the re-grinding effected the flotation results. Also, a small QEMSCAN analysis was performed for the +90 µm fraction to identify how liberation was improved by re-grinding. The flotation trials were performed with two references and two different intensities for the re-grinds.  The results showed an overall increase in grade of zinc in the concentrate with increased grinding for all the trials. SiO2 and MgO content (typical elements for silicate gangue) was also reduced in the zinc concentrate, showing that re-grinding helped liberate locked sphalerite particles. The QEMSCAN results showed that the liberation of sphalerite particles in the +90 µm fractions increased with re-grinding. For the bulk fraction scavenger concentrate the zinc recovery increased slightly, for the +90 µm fraction scavenger concentrate, a slight decrease in zinc recovery could be seen after re-grinding. The decrease in recovery was larger for the cleaner tailings (bulk and +90 µm) and was decreased further with increased grinding. However, for the +125 µm fraction tests, zinc recovery was increased for both scavenger concentrate and cleaner tailing. The references showed an overall high recovery but a zinc grade in the concentrate close to the feed grade of zinc and contained a high grade of SiO2. Indicating that without re-grinding the amount of free sphalerite is low. Overall, the re-grind and flotation tests pointed towards that value can be created for the plant by re-grinding the scavenger concentrate and cleaner tailing.

The core of this project was the development of a column flotation dynamic model that can reasonably predict the changes in the concentrations of all solid and bubble species, along the full column height. A dynamic model of a process is normally composed of a set of partial or ordinary differential equations that describe the state of the process at any given time or position inside the system volume. Such equations can be obtained from fundamental material and/or energy balances, or from phenomenological derivations based on knowledge about the behavior of the system. A phenomenological approach referred to as population balance modeling was employed here. Initially, a two-phase model was formulated, which represents the behavior of the gas phase in a frother solution. The column was viewed as consisting of three main regions: a collection region, a stabilized froth and a draining froth. Experiments were carried out, based on conductivity techniques, for obtaining empirical data for model validation and parameter estimation. After testing the two-phase model, the equations for the solid species were derived. Consideration of the effects of bubble loading, slurry density and slurry viscosity on bubble rise velocity and, therefore, on air fraction is included in the model. Bubble coalescence in the froth is represented as a rate phenomenon characterized by a series of coalescence efficiency rate parameters. Auxiliary equations that help describe the settling of free particles, the buoyancy of air bubbles, and the processes of attachment and detachment, were also developed and incorporated into the model. The detachment of solids from the bubbles in the froth zones was attributed to coalescence, and it was assumed to be proportional to the net loss of bubble surface area. Almost all parameters needed to solve the model equations are readily available. The set of differential equations that comprise the model can be solved numerically by applying finite difference approximation techniques. An iteration has to be performed, which involves calculating the product flowrate at steady state, modifying the tailings rate and solving the model again until a mass balance is satisfied. / Ph. D.

The demands for copper and rare earth elements (REEs) in the U.S. will keep rising due to their applications in green energy technologies. Meanwhile, copper production in the U.S. has been declining over the past five years due to the depletion of high-grade ore deposits. The situation for REEs is worse; there is no domestic supply chain of REEs in the U.S. since the demise of Molycorp, Inc. in 2016. Studies have shown that the rejected materials from copper and coal processing plants contain significant amounts of valuable metals. As such, this rejected material can be considered as potential secondary sources for extracting copper and REEs, which may help combat future supply risks for the supply of copper and REEs in the U.S. However, the valuable mineral particles in these resources are ultrafine in size, which poses considerable challenges to the most widely used fine particle beneficiation technique, i.e., froth flotation. A novel technology called the Hydrophobic-Hydrophilic Separation (HHS) process, developed at Virginia Tech, has been successfully applied to recover fine coal in previous research. The results of research into the HHS process showed that the process has no lower particle size limit, similar to solvent extraction. Therefore, the primary objective of this research is to explore the feasibility of using the new process to recover ultrafine particles of coal, copper minerals, and rare earth minerals (REMs) associated with coal byproducts. In the present work, a series of laboratory-scale oil agglomeration and HHS tests have been carried out on coal with the objectives of assisting the HHS tests in pilot-scale, and the scale-up of the process. The knowledge gained from this study was successfully applied to solving the problems encountered in the pilot-scale tests. Additionally, a new and more efficient equipment known as the Morganizer has been designed and constructed to break up the agglomerates in oil phase as a means to remove entrained gangue minerals and water. The effectiveness of the new Morganizers has been demonstrated in laboratory-scale HHS tests, which may potentially result in the reduction of capital costs in commercializing the HHS process. Furthermore, the prospect of using the HHS process for processing high-sulfur coals has been explored. The results of this study showed that the HHS process can be used to increase the production of cleaner coal from waste streams. Application of the HHS process was further extended to recover the micron-sized REMs from a thickener underflow sample from the LW coal preparation plant, Kentucky. The results showed that the HHS process was far superior to the forced-air flotation process. In one test conducted during the earlier stages of the present study, a concentrate assaying 17,590 ppm total REEs was obtained from a 300 ppm feed. In this test, the Morganizer was not used to upgrade the rougher concentrate due to the lack of proper understanding of the fundamental mechanisms involved in converting oil-in-water (o/w) Pickering emulsions to water-in-oil (w/o) Pickering emulsions. Many of the studies has, therefore, been focused on the studies of phase inversion mechanisms. The results showed that phase inversion requires that i) the oil contact angles (θo) of the particles be increased above 90o, ii) the phase volume of oil (ϕo) be increased, and iii) the o/w emulsion be subjected to a high-shear agitation. It has been found that the first criterion can be readily met by using a hydrophobicity-enhancing agent. These findings were applied to produce high-grade REM concentrates from an artificial mixture of micron-sized monazite and silica. Based on the improved understanding of phase inversion, a modified HHS process has been developed to recover ultrafine particles of copper minerals. After successfully demonstrating the efficacy and effectiveness of this process on a series of artificial copper ore samples, the modified HHS process was used to produce high-grade copper concentrates from a series of cleaner scavenger tails obtained from operating plants. / Doctor of Philosophy / Recovery and dewatering of ultrafine particles have been the major challenges in the minerals and coal industries. Based on the thermodynamic advantage that oil droplets form contact angles about twice as large as those obtainable with air bubbles, a novel separation technology called the hydrophobic-hydrophilic separation (HHS) process was developed at Virginia Tech to address this issues. The research into the HHS process previously was only conducted on the recovery of ultrafine coal particles; also, the fundamental aspects of the HHS process were not fully understood, particularly the mechanisms of phase inversion of oil-in-water emulsions to water-in-oil emulsions. As a follow-up to the previous studies, emulsification tests have been conducted using ultrafine silica and chalcopyrite particles as emulsifiers, and the results showed that phase inversion requires high contact angles, high phase volumes, and high-shear agitation. These findings were applied to improve the HHS process for the recovery of ultrafine particles of coal, copper minerals, and rare earth minerals (REMs). The results obtained in the present work show that the HHS process can be used to efficiently recover and dewater fine particles without no lower particle size limits.

Pyrite (FeS₂) is the major source of sulfur in various coals, and its efficient removal has proven to be a more difficult task than expected. Flotation is generally considered to be the most practicable process for the preparation of coal fines. However, even this technique is usually unable to remove more than 50% of pyrite from a 65-mesh coal sample, which is the typical feed to flotation. There are three major reasons for the low separation efficiency of liberated pyrite from coal by flotation. They include self-induced hydrophobicity of pyrite caused by superficial oxidation, nonselective hydraulic entrainment of pyrite particles into froth product, and incomplete liberation of pyrite from coal that results in composite coal-pyrite particles, i.e., middlings. The present study was undertaken to address problems associated with these recovery mechanisms of pyrite and develop techniques to enhance pyrite rejection in coal flotation. To better understand self-induced hydrophobicity of pyrite, chronoamperometry and voltammetry on freshly fractured electrodes were used to explore incipient oxidation and reduction of the mineral. Voltammetry on rotating ring-disc electrodes (RRDE) was carried out to provide information on soluble species and kinetics of oxidation and reduction processes. X-ray photoelectron spectroscopy (XPS) was used for chemical identification of oxidation products. Galvanic coupling with sacrificial anodes was investigated as a practical method to cathodically protect pyrite and prevent its oxidation. Microflotation tests were conducted under controlled potentials at different solution pH's, and the results were correlated with electrochemical studies. The feasibility of improving pyrite rejection by controlling its surface chemistry was tested in flotation experiments conducted with a 2"-diameter microbubble flotation column and a conventional 5-liter Denver flotation cell. Effects of froth stability on the microbubble flotation of coal were studied with an objective of minimizing hydraulic entrainment of pyrite. The operating parameters were systematically varied to study their effects on water recovery which was used as a measure of froth stability. It has been demonstrated that the upgrading of coal in a flotation column can be significantly improved when froth stability is properly controlled. In an attempt to enhance the rejection of pyrite in middlings, various column circuits were experimentally examined and theoretically analyzed. The effect of circuit configuration on the overall circuit performance was evaluated by separation efficiency and separation curves. It has been shown that the overall separation efficiency of column flotation is rather insensitive to circuitry due to the unique characteristics of the unit flotation column, i.e., the addition of the wash water into the froth. / Ph. D.

A flotation column has been developed Incorporating the use of fine air bubbles (less than 100 microns) to remove ash-forming minerals from micronized coal. The microbubble generator used In this work has been characterized and found to yield a very narrow size distribution. Microbubble column flotation tests have been conducted to study a series of operating variables such as time, bubble size, feed rate, feed point, feed percent solids, column height, bubble number concentration, make-up water addition and countercurrent wash water addition. The results show that i) fine air bubbles are Inherently better suited for floating small particles; ii) both ash and recovery rates Increase with Increasing feed rate, distance of the feed point from the tailings port, feed percent solids and bubble number concentration; iii) taller columns result In Improved recovery and ash rejection; and iv) the countercurrent wash water addition minimizes the entrainment of mineral matter to the froth product. Proper control of these parameters makes It possible to produce super clean coal (< 2% ash). / M.S.

The objective of this study was to evaluate various multi-stage circuit arrangements that may be used to improve the column flotation of micronized coal. Laboratory flotation tests were performed with two different samples of Pittsburgh No. 8 seam coal. The first coal, Coal A, was ground to two different particle sizes and subjected to both column and conventional flotation. These tests were performed to obtain an initial understanding of the operational behavior of the column process and to compare the results with those of conventional flotation. The second coal, Coal B, was used in the actual testing of three different column circuit arrangements. The experimental test results were compared to simulated results obtained using a rate-based flotation model constructed in the present work. Several hypothetical flotation circuits were also examined using the simulation model and experimental flotation rate data. The circuit test results showed that each of the different circuit configurations possessed specific advantages in terms of throughput capacity, combustible recovery, ash rejection and sulfur rejection. However, the overall performance curves for each circuit were all found to fall on or just below the maximum separation curve predicted using the release analysis technique. Also, the simulated results in almost all cases predicted better results than what was actually obtained. This discrepancy was attributed to the inability of the rate-based model to adequately describe restrictions associated with the carrying capacity of the column froth. / Master of Science