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By W. Petruk

Mineral liberations that would be obtained by grinding an ore to a specific size can be predicted by analyzing unbroken ore pieces, and mineral liberations in ground ores and mill products can be determined directly by image analysis. Predictions of mineral liberations are made by measuring size distributions and textures of mineral grains by image analysis, and using liberation models. Ground ores and mill products are analyzed to determine the proportion of mineral that is liberated, and the data are used to evaluate mineral behaviour during mineral beneficiation. The liberation data for minerals in a suite of mill products from a circuit are processed by a materials balancing technique to calculate recoveries of minerals in different types of particles throughout the circuit. The liberation-materials balance data define mineral recoveries; define the behaviour of an ore in a concentrator, pilot plant or bench test; and indicate whether recovery improvements are possible and/or feasible.

Jan 1, 1991

By S. K. Kawatra

In most operating coal-cleaning plants, a significant amount of pyrite is recovered in the froth during flotation of high-sulfur coal. Reducing the pyrite recovery first requires that the primary recovery mechanism should be Identified, as different measures are required for reducing entrainment, locked-particle flotation, or true hydrophobic flotation. In this paper, evidence is presented which suggests that hydrophobic flotation is not an important mechanism for recovery of liberated pyrite from freshly-ground coal when the collector is a neutral oil, and that the bulk of the floated pyrite occurs either as a result of simple entrainment or by mechanical locking with floatable coal particles. Column flotation results are also presented which show that significant sulfur reductions can be achieved by reducing levels of entrainment.

Jan 1, 1992

By S. R. Stine

The Cyprus Thompson Creek Mine is located at 2,286m (7500 feet) elevation in the Central Idaho mountains, 61.16 kilometers (38 miles) southwest of Challis, Idaho off State Highway 75. The Thompson Creek molybdenum deposit lies near the eastern border of the late Cretaceous Idaho Batholith within a complexly deformed sequence of Paleozoic sedimentary rocks. Molybdenum mineralization is contained in a composite biotite granodiorite-quartz monzonite stock considered to be genetically related to the Idaho Batholith intrusive event. The ore is being mined by open pit methods and delivered to the concentrator stockpile with an overland conveyor belt system. The concentrator is designed as a 22,680 metric tons (25,000 ton) per day plant at an average head grade of 0.1% Molybdenum.

Jan 1, 1986

Terminology. The term potash originated in "pot ashes" produced by burning wood to ashes in iron pots and then leaching out the soluble salts, chiefly potassium carbonate, from the residue. In modern usage potash is a generic term for various potassium salts used in agriculture and industry. More specifically, it denotes the oxide of potassium, K2O, and the content of potassium in ores and products is stated as weight percent of K2O present. Its salts are commonly referred to as muriate of potash, sulfate of potash, etc. The presence of an adequate supply of available potassium in the soil promotes the health and improves the quality of the plant. insures greater efficiency in photosynthesis, increases resistance to certain diseases, offsets the effect of an oversupply of nitrogen, and helps the plant to utilize soil moisture more advantageously. ("Soils and Men," U.S. Dept. of Agriculture Yearbook 1938.) Occurrence. Potassium ranks seventh among the elements in abundance. It is estimated to constitute some 2.6% of the lithosphere and is found in soluble forms in the oceans, lakes, rivers, and subsurface waters. It is also an important constituent of plant and animal life. Commercially significant accumulations of potassium salts occur in underground deposits formed in past geologic times by evaporation of ancient bodies of saltwater, inland drainage basins, and shallow sea lagoons. The potash content of such waters was the end result of weathering and leaching of insoluble potash-bearing aluminosilicates, such as feldspars, micas, leucitcs, and others. The concentrations of potassium salts in the natural brines of some lakes in the arid and semiarid regions of the continents are also important as sources of potash; the brines of the Dead Sea in Asia Minor, the Great Salt Lake in Utah, and Searles Lake in California typify such concentrations. Every continent has major potash deposits as listed in Table 1 adopted from Cawley.1

Jan 1, 1985

By Brendan Donnelly, Jung-Hoon Sul, Lasitha Piyathilaka, Sanghyun Jeong, Benjamin Probyn

The chains of armored face conveyor (AFCs) are constantly in contact with water and wet coal, which results in a significant proportion of unplanned downtime experienced on longwall equipment due to corrosion. While reverse osmosis (RO) is one of the most prevalent water treatment methods for reducing salinity and dissolved solids in coal mining, its impact on the corrosion of the AFC chains has not been well studied. This work studies the direct effects of RO water on the corrosion of AFC chains. Immersion tests followed by surface micromorphology analysis and corrosion indices indicate that the RO water increased the corrosion rate of the AFC chains due to the reduced tendency of preventive films on the chains.

May 1, 2024

By Diler Katircioglu-Bayel, Oner Yusuf Toraman, Safak Gokhan Ozkan

One of the most energy-intensive processes for producing submicron range calcite is stirred media mill. In the present work, numerous operating parameters such as solid mass fraction, grinding media size, media filling ratio, and grinding time have been investigated using a vertical type stirred media mill. The results are evaluated on the basis of mean particle size, specific surface area, and specific energy consumption. After conducting this study, optimum experimental conditions found to be as 70% media filling ratio, 25% solid mass fraction, 1 mm grinding media size, and 120 min grinding time. Besides, energy savings up to 22% were achieved with the choice of proper media size.

By S. K. Kawatra, J. A. Halt

"INTRODUCTION Iron ore beneficiation needs advances in areas such as process control, automation, burner and furnace designs, silica-free binders, and even alternative iron making routes. While some – like new binders – are only step-changes, all are tools that can help make the ironmaking industry more sustainable. This paper focuses on binders. Binders are added to iron ore fines before they are agglomerated. Binders are used to control the balling process and increase the mechanical quality (e.g. strength) of green-balls and pellets. Agglomeration usually follows thickening and filtration. Good-grade ores, which don't require upgrading, are wet ground to decrease particle size and increase specific surface area. Low-grade ores are often extensively upgraded through wet operations (flotation, dispersion and selective flocculation, magnetic separation, etc.). For both types of ores, thickening and filtration conclude the circuits so that moisture contents are lowered and the ores can be agglomerated. It's been known for a long time that thickening and filtration behaviors are influenced by zeta potential (Sandvik, 1979). Highly-negative zeta potential is good for dispersion while low zeta potential is good for filtration. So, the zeta potentials of iron ore particles are lowered with additions of acid or carbon dioxide, for example, to increase filtration rates and throughputs (Eisele et al., 2005). What happens to particles as their zeta potentials are lowered? Particles aggregate together, effectively forming larger particles. How stable then, are those aggregates during mixing and balling? Are they easily broken up during mixing (if mixers are even used)? Are they broken up during balling? Or, do the aggregates resist breakage so that the availability of ultra-fines, which are so beneficial for agglomeration (Stone, 1968), is decreased? For plants where high doses of bentonite are used as binder, the above questions may be of no concern. The bentonite adds the necessary amount of fines; however, even at those plants, the calcium and magnesium contents in the water affect bentonite's effectiveness (more magnesium and calcium means less effective bentonite). (Engesser, 2003) For plants where new silica-free organic binders are being developed and used, the above questions may need to be considered. The binder is only amongst the iron ore particles for a short period of time – until it burns away or thermally decomposes in the induration furnace – so the liberation, availability and distribution of fine particulates is very critical for the quality of balls and pellets. It was hypothesized that high zeta potential conditions would lead to good pellets, while low zeta potential conditions would lead to relatively bad pellets."

Jan 1, 2018

By Jameel Menashi

Aqueous solutions of Na2Cr04 can be reduced with CO to produce hydrous Cr203 and NaHCO3/Na2CO3. CH3OH, HCOO and HCHO in the presence of CO2 also reduce Cr 6+ and yield similar products. It is shown that the rate of Cr6+ reduction can be represented by the expression -d[Cr6+]/dt=k[Cr6+)1/2[R] where [R] is the aqueous phase concentration of CH3OH, HCOO- or CO. With these three reductants, the activation energies for the reduction processes are comparable and are of the order of 100 kJ/mole. Rapid reduction rates are attained at temperatures in excess of 220°C and, with these three reductants, a green acid-insoluble readily filterable hydrous Cr203 is formed. With HCHO significant reduction of Cr6+ occurs at about 80-100°C and an acid-soluble violet modification of hydrous Cr2O3 is formed. The reduction process is reasonably selective: solutions of Na2CrO4 contaminated with Mo6+ and W6+ can be reduced to yield a Cr2O3 product with significantly reduced contaminant levels. These findings are utilized as a basis for a process for recovering chromium from secondary source materials.

Jan 1, 1984

By C. J. B. Dreyer

In terms of resources of amphibole asbestos only South Africa can be considered. There are only two asbestos producing fields, one in the Northern Cape Province producing only crocidolite and one in the Eastern Transvaal producing only amosite. The banded ironstone sequence in which amphibole asbestos occurs is discussed, and the genesis of the fibre is briefly reviewed. The properties and characteristics of amphibole asbestos are discussed and compared to those of chrysotile. Approximately 300,000 tonnes of amosite and crocidolite are produced annually. Crocidolite has a potential for a fivefold increase of production, which could be sustained beyond the year 2000. An estimate of the resources of the various producing fields in South Africa is made. Mining methods are reviewed and evaluation techniques discussed.

Jan 1, 1978

By Felicia F. Peng

An economic analysis for evaluating coal preparation plants in the Appalachian Region is presented. A complete coal preparation plant which includes multi-streams and/or multi-stages coal cleaning circuits, complete closed water circuits, environmentally acceptable refuse disposal and reclamation systems is used to determine the capital and operating costs. The plant is capable of producing metallurgical grade clean coal or deep cleaned steam grade coal for future power generating plants use. Utilizing a profit maximization as a decision criterion, an optimal yield, and optimal coal clean¬ing circuits, closed water circuits, and disposal and reclamation systems for a given coal characteristic may be selected among the process alternatives by using a process simulation-economic model of the coal preparation plant. A deterministic analysis is used to calculate the discounted-cash-flow interest rate of return (DCFIROR) and project the selling prices, and determine the Return on Investment (ROI) for the cases of 32.65% and 52.74% plant utilization with producing two cleaned products. Sensitivity analyses are performed to quantify the effect of total capital investment, operating cost, run-of-mine cost and yield on an economic criterion-ROI and selling prices. A probabilistic evaluation using the Monte Carlo simulation technique is further made to quantify the risk involved in obtaining the desired profit levels. The results of the economic analysis for a given coal preparation plant configuration and characteristics of raw coal indicates that the selling prices of producing two cleaned products are greatly affected by the change in percent of plant utilization (production schedule), yield, and run-of-mine cost, and total capital cost. The sensitivity analysis also shows that the total capital investments has a three times greater effect on required selling prices than operating costs. The risk analysis shown that the overall economics look very attractive with current coal sale prices and 52.74% plant utilization.

Jan 1, 1979

By ZHONGLING ZHAO, WANTING SONG, GUOZHONG LIU, LEIXIN SHI, ZUI WANG, YUANYUAN LIU, KE GAO, DONGHONG JIANG, WANG LUO, PEICHAO LU, Jianwei Cheng, WANCHENG ZHENG, Gang Zhao, YUNGANG WANG, JIANBING ZHAO

Coal spontaneous combustion is one of the major disasters in coal mines, and it can have devastating effects on the normal mining and production activities of mines. This paper presents the field investigation, field measurement, data analysis, and formulation of fire-extinguishing measures conducted to extinguish the fire from a coal spontaneous combustion incident that occurred in a high-gas coal mine in China.

By R. Hogg

Based on a simplified model for the motion of particulate materials in horizontal, rotating cylinders, approximate expressions have been derived relating the power consumption of a tumbling mill to its dimensions and speed of rotation and to the amount and nature of the charge. The basic assumptions underlying the model are that there is no slip between the charge and the mill shell, that the free surface of the charge remains essentially planar during rotation and that there is no transfer of momentum from the material flowing down the free surface back to the mill shell. The relationships formulated from this model have been shown to be in excellent agreement with experimental results taken from the literature. While the requirements of the model are most closer met at low speeds of rotation, the results appear to be valid even for speeds up to 70-80 percent of the critical speed. Poorer agreement between theory and experiment is found when the charge consists of relatively large particles (balls) and when the filling of the mill exceeds 50 percent of the available volume. Qualitative explanations of these deviations have been offered.

Jan 1, 1971

By T. J. Fenner

Surface and underground mining operations are often hindered by unknown tunnels and stopes and the precise locations of adjacent workings. Mine safety and planning are adversely affected especially in previously undermined areas where historical records are incomplete or non-existent. The results of several Ground Penetrating Radar (GPR) studies for the detection of known mine tunnels in underground coal; salt, and gold mines are presented. A case study to locate unknown, abandoned mine workings at a surface gold mine successfully located all workings to a depth of 11 meters.

Jan 1, 1992

By Caroline Gerwig, Joseph R. Wickline, Zach Agioutantis

In longwall mining, the progressive caving of the roof allows stored energy to dissipate over an extended time. However, in some cases, the roof strata above the longwall panel are unusually strong and do not break with the movement of the working face. Instead, the roof material bridges over the expanse of the longwall, forming an unstable beam. Later, when this beam breaks, the stored energy can release suddenly, causing a seismic event to propagate to surrounding mine workings. The focus of this paper is to estimate the size of beams that form over longwall panels to the seismic energy recorded by the mine. Analytical calculations are used to verify two-dimensional models of longwall roof beams to determine the strain energy and gravitational energy developed, which is then released upon strata failure to surrounding mine workings. Models are developed using RS2 to model the stress, strain, and deflection along the bottom of the longwall roof beam to calculate gravitational and strain energies. Gravitational energy is shown to be the main driving factor in energy release, except for the beams with low heights. Analyses allow the determination of beam characteristics based on seismic data. The analysis is refined using swelling factor to compare various beam drop heights. The models show an increase in seismic magnitude for higher drop heights. However, the actual data provided by the Buchanan mine displays a decrease in seismic magnitude with increasing drop height, indicating that energy is dissipated into the broken gob material as drop height increases.

Jul 1, 2023

By Edward Holloway

The basis for the selection of the ultimate pit for an open pit mining operation is generally opaque. Some form of value-based analysis will generally guide the final decision; however, the value will often gradually plateau whilst the ultimate pit size continues to increase. There is no established framework to guide specialists in the selection of the ultimate pit, with such decisions regularly made on a somewhat arbitrary basis. The key component that is missing from the analyses typically used to select the size of the ultimate pit is risk. Once risk is quantified within the analysis framework, the decision-making range is significantly reduced, and the process for selecting the ultimate pit becomes more consistent, replicable, and defensible. This paper provides an approach to including risk in the process used to select the ultimate pit.

Feb 21, 2024

By P. Bissert, M. Klein, L. Yan, D. Yantek, R. Matetic

"U.S. Mine Safety and Health Administration (MSHA) regulations require underground coal mines to use refuge alternatives (RAs) to provide a breathable air environment for 96 hours. One of the main concerns with the use of mobile RAs is heat and humidity buildup inside the RA. The accumulation of heat and humidity can result in miners suffering heat stress or even death. MSHA regulations require that the apparent temperature in an occupied RA not exceed 95 °F. To investigate this, the U.S. National Institute for Occupational Safety and Health (NIOSH) conducted testing on a 23-person tent-type RA in its Experimental Mine in a test area that was isolated from the mine ventilation system. The test results showed that the average measured air temperature within the RA increased by 9.4 °C (17 °F) and the relative humidity approached 94 percent at the end of a 96-hour test. The test results were used to benchmark a thermal simulation model of the tested RA. The validated thermal simulation model predicted the volume-weighted average air temperature inside the RA tent at the end of 96 hours to within 0.06 °C (0.1 °F) of the average measured air temperature. IntroductionIf an accident occurs in an underground coal mine, miners who fail to escape from the mine can enter a refuge alternative (RA) for protection from adverse conditions, such as high carbon monoxide levels. One of the main concerns with the use of mobile RAs is the potentially adverse thermal environment inside an RA from the metabolic heat of the occupants and the heat released by the carbon dioxide (CO2) scrubbing system. Moreover, the humidity within the RA will increase through occupants’ respiration and perspiration, and from the chemical reaction within the CO2 scrubbing system. The accumulation of heat and humidity can result in miners suffering heat stress, heat stroke or even death."

Sep 1, 2016

By R. R. Klimpel

The industrial processing of ores containing precious metals such as gold is generally very mining site specific, e. g. McQuiston and Shoemaker (1985), vanZyl et. a1. (1988), Fuerstenau and Hendrix (1990), and Hausen et. a1. (1990). The variety of various mineral processing techniques employed is extensive and includes size reduction, size classification, gravity separation, roasting, cyanidation, chemical and/or bacterial leaching, ion exchange and/or solvent extraction, and froth flotation. Gold occurs in the earth's crust on the average in very low concentration estimated by various sources to be between two and five parts per billion. For a specific ore deposit to be considered as an economic source of gold, generally a concentration of near one part-per million (one gram / Mton) or more is required. From a mineralogical viewpoint, gold often occurs in its native or free form. However, it can also occur as gold compounds of telluride, antimony, and selenium contained within either sulfide minerals such as pyrite, arsenopyrite. and pyrrhotite or, to a less extent, within silicate, oxide, and carbonate minerals. In many locations, alloying of gold with silver is found. The optimal processing method for recovering gold is determined by many factors such as the mineralogy of the gold bearing minerals as well as the associated mineralogy of the gangue minerals, the type of liberation profile of the gold bearing minerals upon' size reduction, and the particle size of the gold itself. Free or native gold particles can vary in size from the extreme of large nuggets to very finely disseminated gold associated in a complex sulfide mineral matrix. Typically free gold particles of several hundred microns or larger are efficiently recovered by one or more of a number of gravity based techniques due to the large specific gravity differential of gold (19.3 to 15) versus typical gangue minerals (5 to 2.7). When gold is associated with a sulfide (usually pyrite) matrix, historically most industrial plants involving size reduction and subsequent liberation have operated in a size range of 50 to 80 percent less than 75 microns. Grinding finer than this has been normally not economical unless the ore has a high gold content in the feed ore.

Jan 1, 1998

By D. Naden

Sulphuric acid leaching is the most common method of recovering uranium from its ores and in the following discussion various alternative low cost routes are investigated for recovering uranium from milled ore acid leach liquors. Historically uranium has been recovered from dilute impure leach liquors by packed bed ion exchange (IX), solvent extraction (SX) or a combination of both. All three process routes as commonly practised require complete separation of leached solids and liquids, as well as extensive solids washing in order to reduce soluble uranium losses to a minimum. Separation and washing are costly and often require equipment covering large areas of land. The size and capital cost of the post leach solids washing operation is controlled by the ease of washing the slimes fraction of the ore and the key to the problem of cost reduction, as many others have realised, lies in the ability to separate the slimes fraction of the ore from the more easily washed sands, to wash the sands in low cost equipment and to recover uranium from the slimes stream. Sands and slimes are usually defined in this context as the material larger than, and smaller than, 325 mesh (44 microns). The proposed flowsheets which minimise solid-liquid separation and washing have usually incorporated ion exchange (Refs 1,2,3), and resin in pulp is in fact practised in several mills in the USA (Ref 1). The resins used in these flowsheets and [cranium] recovery throughout the world have been strong base resins which load rapidly and are moderately selective for the uranyl sulphate complex anion over the range of pH which is conveniently achieved in an acid leach operation. Strong base resins are, however, not as selective as weak base resins and weak base liquid reagents owing to the adsorption of competing anions, some of which find their way into the product while others act as resins poisons. Ion exchange using strong base resins must therefore be followed by a more selective weak base SX purification step if pure yellow cake is required.

Jan 1, 1976

Silver bearing bulk sulfide concentrates from several Peruvian mines were autoclaved in a 19 liter (five gallon) laboratory autoclave in Lima, Peru. Cyanidation of autoclaved concentrates following a hot lime jarosite destruction step extracted over 98 percent of the silver and over 95 percent of the gold. Autoclave operating parameters such as time, temperature, pulp density and autoclaved solids recycle were investigated and reaction rates determined. The problems of excessive heat generation, silver solubility during autoclaving, and formation of pellets containing unreacted sulfides were addressed. The data presented will be used to design a continuous autoclave pilot plant to pro-vide reliable scale up data for a 50 -100 ton per day facility.

Jan 1, 1992

By QUANWEN PU, Yonggang Wei, Bo Li

With the advancement of industrial science and technology, more and more catalysts are used. The spent catalysts produced are solid waste products, which cause environmental pollution if they are not treated. At the same time, the spent catalysts contain some precious metals such as platinum in concentrations that are often higher than in their ore resources. Therefore, the recovery of valuable metals in spent catalysts is a hot research topic at present. However, due to the complex structure of the spent catalysts, dealing with them is a lengthy process. The first step of recovery, enrichment of the spent catalyst, becomes particularly important. In this paper, a pressure solution method is investigated for this purpose, in which the alumina carrier is dissolved and platinum is enriched in the leaching residue.