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By S. Partney

Waste clay ponds are a necessary part of the phosphate rock beneficiation process. They are large, averaging a square mile in size; unsightly, in that they protrude 20 to 50 feet (6 -15 meters) above natural grade; stabilization and reclamation is slow; and they have had a history of catastrophic failure. The State of Florida has never published a specific rule regarding the configuration or height of these ponds, but has enforced numerous policies along these lines. The current rules are given and interpreted, along with the reasoning for current policy. In addition, an overview of the history of the policies, and an attempt at explaining the reasons for the changes that have occurred is given. The industry has been forced to make major monetary decisions, having effect for decades on mine economics, in a constantly changing regulatory environment. The Bureau of Mine Reclamation is confident that a direction of regulation has been solidified, but changes are still to be anticipated.

Jan 1, 1998

By J. G. Groppo, A. E. Bland, B. K. Parekh, W. F. Stotts

Many coal preparation plants discard fine coal (generally minus 100 mesh) because it is difficult to beneficiate economically. At the Kentucky Energy Cabinet Laboratory (KECL), the waste streams from several Kentucky preparation plants were evaluated for cleaning with counter-current column flotation. A series of performance evaluations were conducted using a 0.05 m (2 in.) internal diameter (I.D.) 7.9 m (26 ft.) high laboratory column. The objectives of these tests were to assess the performance in terms of recovery and percent ash in clean coal and to optimize the operating parameters relative to combustible recovery and grade. Refuse streams containing 30-50 percent ash were upgraded to 2-8 percent ash with Btu recoveries of nearly 90 percent. A pilot scale column flotation cell was installed at the Bell County Coal preparation plant. This pilot plant utilized a 5.79 m (19 ft.) high and 0.15 m (6 in.) I.D. column to verify the performance data of the 0.05 m (2 in.) I.D. column for two different waste streams. Performance evaluation and pilot plant testing results related to scale-up are discussed.

Jan 1, 1988

By Cloyd D. McDowell

The Federal coal Mine Health and Safety Act of 1969 has been criticized by those who say it goes too far in controlling the activity of private enterprise and by those who say it does not go far enough in eliminating fatalities and accidents in an accident prone industry. Possibly a better evaluation of this legislation lies somewhere in between. From the viewpoint of the coal industry it can be said that through the efforts of those engaged in mining, both the operator and the miner, the year 1969 marked a record year in mine safety with only 190 fatalities and a frequency of 0.84 per million man hours exposure. This outstanding achievement compared to any previous was the result of a determined effort on the part of a concerned industry to improve its reputation as being the most hazardous industry in the nation. After the effective date of the Act, March 31, 1970, the Bureau of Mines assumed complete control of all safety and health activities of the mining industry. Unfortunately, there was no improvement in the safety record of underground mines in 1970 with the total of 255 fatal injuries and a frequency of 1.04 for the year. The number of fatalities declined to 177 in 1971 but the number of underground coal mines had declined from 5,955 to 2,268 during the two year period. Meanwhile, surface mined coal increased from about 38% of total production in 1969 to over 50% in 1971, keeping the annual production of coal at a normal level but involving fewer men.

Jan 1, 1973

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

"The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud.INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1]."

Jan 1, 2018

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud. INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1].

Jan 1, 2019

By Javad Sattarvand, Milad Ghahramanieisalou

As new technologies are introduced to the mining industry, the challenges related to their safe assimilation and the potential changes they bring to mining techniques necessitate adjustments to the current operations. Predictive simulations are critical to understanding unforeseen scenarios and shifting costly changes from the operational stage to the design. The Digital Twin methodology has gained a lot of attention in recent years. Many industries, from manufacturing to engineering and even social sciences, are adopting this approach to understand better the complex systems they are working with. By considering the enormous expenses and risks related to a pilot project, DT could replace that by acting as a prototype in which realistic tests and simulations are carried out with close to zero cost. DT includes a data hub, simulation and analysis tools, and visualization platforms to enable appropriate designs and monitoring plans focusing on unknown areas. A few examples of DT's application in mining include fleet management, Mine-to-Mill optimization (especially D&B), geotechnical digital twin (gDT), etc. When developing regulations, testing the efficiency of the mine layout, design, and magnitude of the restricted area and its access points, human-machine interactions, risk identification, and management processes, examining the efficacy of control systems, and monitoring, DT could help speed up (or completely replace) the process in the physical world.

Jun 25, 2023

By Per H. Fahlstrom

The story of the Aitik mine begins in the early thirties. At that time blocks of low grade chalcopyrite impregnations were found in Swedish Lapland not far from the present mine, some 40 miles north of the arctic circle in the vicinity of the iron ore field at Gällivare. The ore prospect was viewed upon several times. By means of geophysical studies and an extensive diamond drilling programme during the end of the fifties and the beginning of the sixties a huge low grade ore body was discovered under the muskeg overburden. In 1965 55 million tons of ore with 0.5 % Cu was known to a depth of 160 ft. In 1970 the ore reserves had increased to 165 million tons with 0.40 g, Cu at a depth of 400 ft. Decision to develop the mine and to start production at an annual production of 2,2 million tons was made in 1965. A mine and concentrating plant, the A section, went on stream in 1968. In 1970 decision was made to expand the production with 3,3 million tons pr year. A second section of the plant, the B section, went on stream during the end of 1972 and the beginning of 1973 The production in the initial plant has been increased to 3,1 million tons pr year and during 1974 the combined A and B sections are expected to treat' about 7,2 million tons of ore. The two sections will treat 10,000 and 13,000 tons per day.

Jan 1, 1974

By Ahmet Tezcan, Zeki M. Dogan, Yalçin Kaytaz

Autogenous grinding in Turkey was first applied to quartzite where the mill is grate discharge type with an internal diameter of 2.99 m, having a length of 3.20 m and a capacity of as to 11.0 tons per hour. The second autogenous grinding was applied to a copper ore at Murgul Cakmakkaya Flotation plant of (KBI) Black Sea Copper Works having three mills 26 feet in diameter and 10 feet in length in close circuit with hydrocyclones. Recently, alterations were made in grinding circust by adding pebble mills. Furthermore, rubber liners and lifter bars were introduced pebble extractor and classifier were new features in the primary mills and circulating load was reduced by 20-50 tons per hour. At the end of these charges, capacity of each mill line was increased to 48 tons per hour with a full capacity of more than 9000 tons per day. Thirdly, autogenous grinding has been executed successfully on the slags of Samsun Smelting Plant of (KBI) Black Sea Copper Works with a mill of 20 feet in diameter and 6 feet on length having pebble mill following autogenous grinding.

Jan 1, 1992

By R. W. Bruce

The normal practice for processing complex sulphide ore has been to grind with steel balls and/or rods in steel mills before flotation. Little thought had been given to the effect of grinding media on flotation selectivity of copper-lead -zinc ore until 1960 when Rey and Formanek(1) published their experiments on sphalerite floatability after grinding in a porcelain mill with ceramic pebbles. The studies by Rey and Formanek had a pronounced effect on the milling of copper-lead-zinc ores in Canada and possibly throughout the world at a time when autogenous grinding was creating a great deal of interest. The impression was created, and possibly still exists with some mill men, that grinding an ore in the absence of steel will result in a loss of selectivity between copper, lead and zinc minerals. This is in spite of the fact Rey and Formanek's experiments showed that the "naturall floatability" of sphalerite was not apparent after grinding in a porcelain pebble mill in an alkaline pulp.

Jan 1, 1976

By Andrew Thomson, Peter Wright

"A section of a London Underground expanded segmental lined tunnel constructed in the 1970’s progressively deteriorated after 2001, with extensive radial joint spalling and mid-segment cracking. As part of the PPP contract Tube Lines Ltd and their partners carried out various analyses of the tunnel, including analysis of laser scans of the tunnel, geotechnical testing through the tunnel walls, geotechnical desk studies, real-time monitoring and even full-scale segment testing. This paper discusses the behaviour of the tunnel lining in this setting, the causes of failure and the lessons that can be learnt from this rare example of a tunnel which was progressively failing in service. It is finally noted that a replacement tunnel lining was installed during night-time possessions and completed in 2015. INTRODUCTION In 2003, Tube Lines Ltd. was established under a Public-Private Partnership (PPP) scheme to provide maintenance services and upgrade projects for London Underground (LU). The Deep Tube Tunnel Knowledge and Inspection Programme comprised the assessment of all tunnels, shafts and other underground structures. This included a number of lengths of tunnel constructed using segmental concrete linings. First used by LU in the 1960s, the savings both in cost and speed of construction led to concrete linings replacing cast iron for new tunnel construction in all but a few cases. During the initial scoping phase of this programme, inspections identified potential issues with the expanded concrete segmental linings on the Jubilee line. The inspections found a number of segments with surface spalling (Figure 1a), which did not appear on the post-construction defect records. These were generally relatively shallow, up to 40mm deep, spread over a wide area of segment. These were different to spalls caused by handling damage, which are typically localised at segment corners. Subsequent inspections over the following years identified an increasing number of segments with these types of spalls as well as an increased incidence of longitudinal mid-segment cracking (Figure 1b). The increasing occurrence of these defects was hypothesised to be caused by tensile stresses developing in the vicinity of the joint."

Jan 1, 2016

By R. Hildebrandt, K. Cybulski, G. V. Goodman, Z. Dyduch, M. L. Harris, M. J. Sapko

Past research showed that bituminous coal dust remains dry and dispersible in the presence of moisture. Rock dust must disperse with the coal dust to effectively inert a propagating coal dust explosion. Non-treated rock dust readily absorbs moisture, limiting its dispersibility, while anti-caking treated rock dusts can remain dry. The National Institute for Occupational Safety and Health contracted the Central Mining Institute in Poland to conduct large-scale testing in their Experimental Mine Barbara to determine if a treated rock dust can be as effective as non-treated rock dust in attenuating or quenching coal dust explosions under the same experimental conditions. INTRODUCTION One of the goals of the National Institute for Occupational Safety and Health (NIOSH) is to conduct research to reduce the risk of mine disasters and provide workplace solutions to reduce the risks associated with accumulations of combustible and explosible materials; the most common form of which is the generation of coal dust during the mining process and its subsequent distribution downwind. Dispersible rock dust is a primary defense for preventing coal dust explosion propagation in underground coal mines, and its properties are defined in 30 CFR 75.2. “Pulverized limestone, dolomite, gypsum, anhydrite, shale, adobe, or other inert material, preferably light colored, 100 percent of which will pass through a sieve having 20 meshes per linear inch and 70 percent or more of which will pass through a sieve having 200 meshes per linear inch; the particles of which when wetted and dried will not cohere to form a cake which will not be dispersed into separate particles by a light blast of air; and which does not contain more than 5 percent combustible matter or more than a total of 4 percent free and combined silica (SiO2), or, where the Secretary finds that such silica concentrations are not available, which does not contain more than 5 percent of free and combined silica.” An earlier NIOSH investigation of rock dust revealed two significant concerns with the supply of rock dust to U.S. coal mines: 1) insufficient particles <200 mesh (75 μm) and 2) all rock dusts when wetted and dried formed cakes and were not easily dispersed with a light blast of air (1). Past research by the U.S. Bureau of Mines and others showed that bituminous coal dust remains dry and dispersible in the presence of moisture (2). Rock dust must also be dispersible in concert with the coal dust to effectively inert a propagating coal dust explosion (3, 4, 5, 6). Non-treated rock dust, however, readily absorbs moisture, limiting its dispersibility, while a rock dust treated with long-chain fatty acids (such as stearic acid) can remain dry and dispersible. Stearin-treated rock dust has been used in British coal mines (7) and is commonly used in Polish coal mines.

Jan 1, 2019

By Carl-Henrik Knudsen, Rolf Tore Arnesen, Eigil R. Iversen

Recovery of iron and copper from acid mine drainage was studied in pilot scale at a continuous treatment plant at the mine site of the King&apos;s mine in Roeros, Norway. The average load on the plant was 1 m3/h. The process tested in this project is a three-step process consisting of: 1) Fe(OH)3 precipitation at pH 3.5, dewatering of iron sludge and transport to ferric chloride plant; 2) after filtration precipitation of residual metals at pH 9, concentration of metal sludge by flotation; 3) extraction of metals from the sludge with sulphuric acid followed by recovery of copper by electrolysis. The copper could be used for local production of different crafts, but this condition needs a further evaluation. The process economy at a full scale plant is evaluated.

Jan 1, 2000

I found the article most interesting and timely. I consider it imperative that mine case studies like this one are undertaken and published so that the practicing mining engineer and geologist can obtain a better idea of what is the preferred estimation and grade control technique to use, given the site-specific geology, grade distribution, etc. Mining companies, with operating gold mines, need to be encouraged to either initiate such a study in-house or support a research project by financing a graduate student at a local university. The latter method is preferable since it not only provides research support but also an excellent training opportunity for young mining engineers and geologists. Ideally, what is needed is exploration drill-hole, blasthole, and mine production data that can then be used to assess the efficiency of different estimation and grade control methods by comparing the results obtained using the drill/blasthole data with the actual production data. It is hoped that more mining companies will realize that the rewards from this kind of research greatly outweigh the sometimes perceived problem with releasing data to a university. I noticed that the authors used the blasthole data to represent the actual values, the validity of which relies heavily on the two key assumptions with regards to blasthole assays and in-pit ore-waste selection procedures. Could the authors comment in more detail on these assumptions and what criteria they used to check the validity of the assumptions? Since the article was published, have the authors had the opportunity to compare their estimates with the actual production data? It would be interesting to compare other estimation methods with OK and IK. For example, how well did the polygonal method used at the mine compare with the results given in the paper? I would like to mention that B.L. Kwa completed a similar study (Kwa and Mousset-Jones, 1986) using exploration drill-hole, blasthole, and production data from the Alligator Ridge mine in Nevada. The study compared a variety of estimation methods, including IK, and the final results are shown in Fig. 1. The IK method exhibits the smallest difference between estimated and production results and is closely followed by the indicator moving window (Ind MW) method. Reply by Y.C. Kim We do concur with Professor Mousset-Jones with respect to the merit of such case studies performed at a local university. In answer to his request for comment on the validity of the two key assumptions made in our paper, we can only say that these are reasonable assumptions that are very difficult to validate. The assumption of blasthole assays being accurate can be partially validated by a careful analysis of the nugget component of blasthole variograms. The assumption that the in-pit ore-waste selection procedure in use was an effective one can only be validated through comparison between the mine estimates of ounces of gold versus the actual recovered ounces of gold at the mill. Neither one of these two assumptions has been validated to date due to several reasons. For example, the mine went through a change in management since the study. Similarly, no compelling reasons have been presented to perform detailed variogram analysis of the blasthole data. We do, however, believe that the obtained results (i.e., indicator kriging performs better than OK in a highly skewed deposit) will remain the same in relative terms, regardless of how valid these two assumptions may be. The most recent com¬parative study by D. S. Lindsey of the Montana College of Mineral Science and Technology also shows a similar ranking between ordinary kriging versus probability kriging, which is an extended version of indicator kriging.

Jan 1, 1987

By Robert R. Dorfler

We ought not look back unless it is to de- rive from past errors, and for the purpose of profiting by dear bought experience. - George Washington During the mid 1960&apos;s, the application of automation was introduced to Climax by two process control companies. In 1969 Climax decided to seriously consider improved control of selected areas of their milling process after preliminary research had been completed on two rougher ball mill grinding circuits. In 1971 the Climax Mill Department developed a direct digital control scheme for rougher milling, designed a remote control center for handling process water, planned to automate the central steam heating facility for energy conservation, and convert other plant areas to automation. All twenty rougher grinding mills and the industrial water system were automated by 1977. The idea of automating central heating and other areas was abandoned although some fuel conservation items were installed at the heating plant. During this period from 1971 to 1978, the mill capacity was expanded from 39,000 metric tons per day to 43,000 metric tons per day and cyclone classification was introduced to rougher milling. Ore quality changes degraded the milling recovery from over 90% to less than 85%. This report describes the procedure the mill used in developing the automatic process and the occurances that affected the timing and fina1 control logic. The report deals primarily with the automation of the ball mill grinding, classification and flotation process as this was the bulk of the seven-year project.

Jan 1, 1979

By Son Hoang Nguyen, Thanh‑Hai Pham, Duoc Van Tran, Dung Thi Kim Nhu

Apatite ore plays a critical and irreplaceable role in fertilizer production and is being upgraded in quality in northern Vietnam. The old tailing of the Bac Nhac Son beneficiation plant (Vietnam) has a significant apatite amount in the form of coarse unliberated particles. These millions of tons of tailing ore have 6–8% of P2O5 content, which can be retreated to provide fertilizer factories. This work presents the results obtained in recovering apatite from this tailing in the laboratory. By the combination of the processes of classification, preliminary flotation, re-grinding and cleaning in the flotation column, an apatite concentrate of 30.68% P2O5 can be obtained with the total recovery of 37.76% from the initial tailing of 7.56% P2O5. The study results not only provide an appropriate technical solution for recovering the apatite ore from tailing ore but also have outstanding meaningness in the recovery of mineral resources under the context of a circular economy.

May 24, 2023

By R. Bhaskar

The results of an experimental study on the dispersion and deposition characteristics of fine coal dust in mine airways are presented. Experiments were conducted at various flow velocities using bituminous and semi-anthracite-coal dust which had a large proportion of dust in the respirable size range. Cross sectional variations in the concentration of dust at various points in the airway were also measured. Instruments used in the experiments included gravimetric samplers and impactors. Concentration measurements were made both for the respirable and total airborne dust. The experimental design, sampling procedures and results are discussed. The study is aimed at providing a better understanding of the temporal and spatial behavior of dust particles in underground mine airways.

Jan 1, 1986

By L. A. Pursell

The case history, which is the subject of this paper, involves the expansion of a lime plant in Texas. The major new production unit is to be a 545.45 tonne per day coal fired rotary kiln. Auxiliary facilities include additional quarry haulage equipment, modifications to the crushing plant, a hydrator, storage facilities, and a loading system. The existing plant consists of a high calcium limestone quarry, two coal fired rotary and 10 gas fired shaft kilns, nth crushing, screening, stockpiling, hydrating, storing and loading facilities. The proposed modification would increase quarry volume from 545,000 tonnes per year to 805,000 tonnes per year. The plan also includes the addition of dolomitic lime to the product line which requires the purchase and handling of 186,000 tonnes of screened, washed dolomite yearly. A decision to undertake such an expansion was previously a simple and straight forward process. Such a decision now faces a major and complicating obstacle which can delay the project, raise the expected cost, or prevent it entirely.

Jan 1, 1980

By A. House

Newmont Mining Corporation owns and operates the Phoenix Mine near Battle Mountain, Nevada, USA. The existing milling facility has operated since 2006. Following the startup of the milling operation, and coupled with the upswing in the copper market, Newmont realized that Phoenix needed a process option to treat lowgrade transition and oxide copper ores. This meant that potentially revenue generating ore was being mined as waste. With no system in place to treat the transition and oxide copper, work began to determine the most feasible approach. After evaluating multiple options, the flow sheet was set for a run-of-mine (ROM) heap leach pad and solvent extraction/electrowinning plant (L-SX/EW). This paper will discuss the synergies and challenges of developing a copper L-SX/EW plant at the Phoenix copper-gold mill.

Jan 1, 2010

By H. M. Park

Risk is a major factor in all mining activities, arising from many internal and external variables. The authors? research has resulted in a systematic method for the quantification of risk in mining projects (Park, 2012). This is the second part of a three-part series. The first part was published in the October 2013 issue of Minining Engineering magazine and can be found here at www.me.smenet.org. This paper discusses methods of assessing the overall risk associated with those variables, and presents a statistical model based on a survey of 31 experts. The third paper, to be publised at www.me.smenet.org on Oct. 14, presents an evaluation of the statistical model, using data from previous mining investment decisions in which the authors have been involved.

Oct 1, 2013

By W. W. Wen

The objective of this investigation was to study the electrokinetic behavior of oxidized coals to indicate the feasibility of separating oxidized coals from ash materials and pyrite by an mine flotation process. The effects of soluble inorganic electrolyte products from coal oxidation were also examined. The electrokinetic behavior of coals of different rank and degree of oxidation were studied by an electrophoreses technique. Generally, increasing the oxidation increases the negative value of the zeta-potential and decreases the isoelectric point (IEP). Flotation results indicate that the flotation recovery increases with decreasing pH values, and that with ferrous sulfate at pH 6.5 the flotation recovery is optimum while the pyrite was depressed. Comparing the flotation results with the electrokinetic results verified that in the presence of ferrous sulfate, the maximum hydrophobicity of oxidized coal with dodecylamine and frother is near its IEP which is around pH 6.5 to 7.0.

Jan 1, 1976