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By R. Franfield

This paper describes a situation that many drilling and blasting contractors have found themselves in – a client?s fear of the use of explosives. In the summer of 2007 Blasting Services Ltd was asked to undertake some drilling and blasting work in an abandoned quarry that was being restored for a housing development. Previous quarrying operations had left a series of overhangs that were considered unsafe and required removing. The site contractor considered that the best and safest way to carry out this work was by drilling and blasting. Blasting Services Ltd, with the technical assistance of Exchem Explosives, designed a drilling and blasting regime to accomplish this work with a series of trim blasts. The design took into account the close proximity of a structure in very poor condition and the imposed vibration limit of 6.0 mm/s (0.24 i.p.s.). The faces were profiled, holes drilled and explosives ordered. A few days before the first blast was due to be carried out the client decided that the poor condition of the nearby structure meant that the use of explosives could not be permitted. This decision was made despite the blast design indicating that the vibration level would be well below the 6.0 mm/s (0.24 i.p.s.) limit.

Jan 1, 2009

By Eric B. Sansone

The Federal Coal Mine Health and Safety Act of 1969 requires coal mine operators to sample accurately the amount of respirable dust in the mine atmosphere. A technique approved for such sampling employs a 10 mm nylon cyclone followed by a membrane filter (nominal pore size ????= 5 µm) and a pump. The pump is required to provide a flow rate of 2.0 ± 0.1 liters per minute. Calibrations have yielded a range of flow rates (1.4 to 2.8 1pm) at which the cyclone should be operated to remove the non-respirable portion of the dust. The purpose of this project was to evaluate the performance of the 10 mm nylon cyclone at flow rates of 1.8 and 2.0 1pm using a polydisperse coal dust aerosol. The penetration characteristics of the cyclone were to be compared to the United States Atomic Energy Commission's respirable dust standard. The effect of relative humidity on cyclone penetration characteristics was also to be observed.

Jan 1, 1973

By S. M. Moore

Each year more than 400 coal miners are injured (fatally and nonfatally) by rock falling from between or around roof supports. Many of these injuries can be prevented by the installation of roof screen. However, many coal mines are reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential savings in workers' compensation (W C) premiums that can be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by Illinois and Kentucky were investigated in this study. Using data obtained from the Mine Safety and Health Administration, national and state WC bodies and individual insurance companies (e.g., average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), baseline mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state's actual WC rate-setting formulas, a sensitivity analysis was conducted to determine the total savings in WC costs after a three-year period. Annual savings in WC premiums ranged from $41,000 to $326,000 for the 67-person mine, and $96,000 to $843,000 for the 150-person mine. An economic analysis of the cost of a roof screening program at a 67-person mine was also conducted. The annual cost of a roof screening program here was estimated at $240,000. At this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By M Cooksey, Z Zhu, N Langrené, W Chen, T Tarnopolskaya

"Dynamic decision-making can significantly increase the value of mining projects and deliver better sustainability outcomes under conditions of uncertainty. Despite its benefits, dynamic decision-making is rarely used by the mining industry, possibly because of the complexity of the algorithms used to make sequential decisions under uncertainty and difficulties in visualising the optimal strategies. To achieve better outcomes in practice, decision-makers in industry would benefit from a tool that would help them to understand and implement the results of the analysis.CSIRO is developing innovative dynamic decision support tools that include an intuitive graphical display of the boundaries between the regions of different optimal decisions, which would significantly assist optimal sequential decision-making under uncertainty. The dynamic decision support tool is based on novel stochastic optimal control methodologies that can be used to optimise decisions in situations with multiple uncertain variables and decisions (Tarnopolskaya, Chen and Bao, 2015; Chen et al, 2015; Langrené et al, 2015). The types of market uncertainties that can be considered include commodity prices, interest rates and exchange rate, while the key orebody uncertainties include the reserves, boundaries between ore types and even geometallurgical parameters.The dynamic decision support tools have a number of benefits and can be used to:construct optimal operational strategies and to optimally manage resources projectsgain insight into optimal strategies under different market conditions and project settings compare expected project values under different project settings and to select the setting that provides best sustainability outcomes.This extended abstract demonstrates the use and benefits of the dynamic decision support tools using a synthetic case study of a gold mine.CITATION:Chen, W, Langrené, N, Tarnopolskaya, T, Zhu, Z and Cooksey, M, 2016. Dynamic decisions under uncertainty – a case study on a gold mine, in Proceedings Life-of-Mine 2016 Conference, pp 20–23 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 28, 2016

By J. L. Kenty, J. P. Hirth

The concept of a critical super saturation, below which the nucleation rate is essentially zero and above which it is essentially infinite, is discussed with reference to vapor-solid nucleation. The necessary and sufficient conditions deduced for observations of this type of behavior are: 1) the nucleation rate must exhibit a sharp dependence on super saturation, 2) the growth rate must be sufficiently large that nuclei become observable in the time period of the experiment, and 3) the number of highly preferred nucleation sites must be small. Experiments reveal that the nucleation of silver on sodium chloride is visually detectable at all experimentally accessible super saturations and does not exhibit critical nucleation behavior. Failure to observe a critical super saturation is attributed to the insensitivity of nucleation rate to supersaturation as a consequence of the particular values of the contact angle and the surface free energy for this system. THE concept of a critical supersaturation, below which the nucleation rate is essentially zero and above which it is essentially infinite, arises naturally in homogeneous nucleation theory. Experimentally this type of behavior has been found by Volmer1 and others for water and other low surface tension liquids, as reviewed by several authors.2'3 The same type of behavior has been predicted and observed for heterogeneous nucleation of solids by Yang et al.4 and others,596 as also recently reviewed.2,7,8 In the work reported here on the heterogeneous nucleation of silver on NaC1, however, no critical super-saturation was found. Similar observations have been made recently for other systems.9-11 These results led to a reexamination of nucleation theory which revealed that there are conditions for which critical behavior is not predicted, either for homogeneous or heterogeneous nucleation. Although heterogeneous nucleation is of primary importance in this paper, some insight into critical behavior for such a case can be gained by considering homogeneous nucleation. Accordingly both types of nucleation theory are reviewed briefly. The requisite conditions for critical supersaturation behavior are then considered. The experimental results for the nucleation of silver on NaCl are presented and interpreted in terms of the theoretical presentation. REVIEW OF NUCLEATION THEORY There are essentially two approaches to nucleation theory, the so-called classical theory involving the concepts of bulk thermodynamics, and the statistical mechanical theory in which nuclei are regarded as macromolecules. The classical theory is based on the work of Volmer and Weber12,13 and Becker and. Doring14 and has been extended by Pound et al.15 The crucial assumption in the classical theory is that the small clusters or nuclei can be characterized by the same thermodynamic properties as those of the stable bulk phase. Thus, the nuclei are assumed to have a surface free energy, y, and a volume free energy of formation (relative to the vapor phase), ,, identical to that of the bulk. For deposition under low super-saturation conditions, the nuclei are large and this assumption is satisfactory. However, in many cases of interest, the nuclei contain only a few atoms and this assumption is highly questionable. The statistical mechanical models originated, for the specific case of a dimer as the critical nucleus, with the work of Frenkel16 and were extended later to larger sizes by Walton,17,18 Hirth19 and, more recently, Ht Zinsmeister. These models describe the nucleus in terms of a partition function, the estimation of which is tractable for clusters of 2 to 10 atoms, but extremely difficult for clusters larger than 10 atoms. Although the classical and statistical mechanical models are expected to apply for the limiting cases of large and small nuclei, both are uncertain for intermediate sizes. In this paper we shall treat only the classical model, recognizing that it is exact only for large nucleus sizes and regarding it as a phenom-enological description for small nucleus sizes. When analyses of experimental data using bulk properties show the nucleus size to be small, the resulting parameters should be regarded as largely empirical parameters describing the relative nucleation potency of the system. Considerable justification for the continued use of classical theory is provided by its general success in predicting nucleation behavior as a function of supersaturation and temperature. We emphasize that the qualitative features of the statistical mechanical models, particularly the critical super-saturation behavior that is central to the present work, are the same as those of the classical model. Of course, potential energy terms and surface partition functions replace the volume and surface energy terms of the latter model. The most recent versions of classical nucleation theory have been extensively reviewed.2,3,7 so that only the results are presented here. For homogeneous nucleation of a condensed phase from the vapor phase, the volume free energy change is ?Gv=vrT = =^ln£ [1] where v is the molecular volume of the condensing species. The supersaturation ratio,

Jan 1, 1970

By C. S. Lord

"The Regal Silver mine is 5 miles north-northwest of Albert canyon; a station on the main line of the Canadian Pacific railway; 21 miles east of Revelstoke; British Columbia.Development, commenced in 1918, comprises more than 8,000 feet of workings from six adits, one of which extends into the Regal Silver property from the adjacent Snowflake ground. Production has been limited to a little tungsten concentrate recovered recently from a few r..undred tons of ore treated in a small underground mill. The property has received some attention as a possible source of tin. The deposits are large quartz veins containing irregular sulphide bodies. Ore minerals normally make up less than 1 per cent of the veins and include pyrite, galena, sphalerite, scheelite, and stannite. The veins are approximately parallel with the bedding of the enclosing graphitic, siliceous to calcareous, slates and argillites. These strata are of Precambrian age, strike about northwest, and dip northeast. No other rocks are known in the vicinity of the workings . Granitic rocks, intrusive into t'he sediments, outcrop 4 miles to the southwest and 7 miles to the northeast and may underlie the property at no great depth."

Jan 1, 1954

By Ludwig Schmidt

Large quantities of gasoline are lost each year by evaporation from bulk-storage station tanks. The rate of evaporation from bulk-station tanks is influenced by two principal factors: First, the vapor pressure of the gasoline in the tank; and second, the amount of "breathing" of the tank. The vapor pressures of the gasoline govern the amount of the gasoline vapor in the air-vapor mixture above the gasoline in a tank at a given temperature. This is one of the principal factors in determining the total pressure of the air-vapor mixture within the tank. The "breathing" of the tank is caused by the inflow of air from the atmosphere during certain periods of the day and the expulsion of the air-vapor mixtures at other times, caused by atmospheric variations. The pressure built up within a tank and the amount of gasoline in the air-vapor mixture are both a direct function of the temperature existing within the tank. Therefore, two of the most common methods for reducing evaporation losses from bulk-station tanks are: (1) The use of vapor-tight tanks equipped with relief valves and operated at 1 or 2 ounces of pressure and about .5-ounce vacuum; (2) reducing the temperature of tanks and their contents by painting them a light color, by insulation, and in some instances by using water sprays. Water sprays, although effective, usually are too expensive to operate economically. In order to determine and compare the effects of various colors, insulation, and pressure in reducing the evaporation losses of gasoline held in storage in bulk-storage tanks, the Bureau of Mines made observations from May 29 to October 15, 1930, on five horizontal tanks of 12,000-gallon capacity each, which were erected at the plant of Black, Sivalls and Bryson Manufacturing Co., Kansas City, Mo. This report gives in detail the results of these evaporation tests.

Jan 1, 1931

By E. W. Dewing

The solubility of A14C3 in cryolite melts has been determined by holding the melt with aluminum in a graphite crucible for 4 hr at temperature, quenching the crucible in water, and analyzing the frozen salt by CH4 evolution. The solubility is given by log(wt % Al4C3) = 1.298 -1.l74 (wt ratio NaF/AlF3) independent of temperature in the range 985" to 1080°C. Additions ofA1203, .NaCl, CaF2, or LiF up to 10 wt pct do not affect the solubility, provided that LiF and CaF2 are replaced by equivalent NaFin calculating the ratio. The variation with ratio can be accounted for quantitatively on the hypothesis that the Mechanism of solution is Al4C3 + 2AIF3 = 3Al2CF2 since the solubility varies as (aAIF3) 2/3; Mechanisms involving C", C2- , AlC- , A12C2+, and Al2CF+are re-jected. At ratios above 1.5 truces of c;- are also present. It was noticed some years ago in this laboratory that carbon specimens made cathodic in a cryolite bath decrease in diameter over a period of time.' It was also found that cryolite which had been melted with aluminum in a graphite crucible was sometimes yellow, and that A14C3 could be seen intimately embedded in the cryolite crystals. CH, could be recovered by acid attack. There was thus reason to believe that A14C3 had a genuine solubility in molten cryolite, and the object of this work was to measure it. EXPERIMENTAL AND RESULTS A graphite crucible, 2 in. ID and 2.5 in. deep with a close-fitting lid, contained about 150 g of cryolite (plus additives) and about 20 g of 99.99 pct pure Al. It was placed in a vertical Inconel furnace tube, covered with refractory brick to provide thermal insulation, and the top of the tube was roughly sealed with disks of asbestos paper. An alumina tube led argon down to the level of the top of the crucible to expel air, and an alumina sheath with a Pt-10 pct Rh ther-mocouple inside was positioned between the crucible and the wall of the Inconel tube. The tube was heated in an electrical resistance furnace for the desired length of time; the refractory bricks were then removed, and the crucible was lifted out with tongs and quenched in water. After such a crucible was broken open it could be seen that its inner surface was yellow owing to the presence of A1,C3, and sometimes needle-like crystals could be observed (at X75 magnification) growing out into the nielt. It was concluded that Al,C, had been present as a separate phase when the crucible was at tenlperature and that the melt had been saturated: in

Jan 1, 1970

By M. F. Handley

Sequential grid mining has been employed as the primary mining strategy at the AngloGold West Wits Elandsrand Mine since the inception of stoping operations from the sub-vertical shaft system in 1988. Little has been published on sequential grid mining as a mining strategy since, and the purpose of this paper is to describe sequential grid mining, its history, its implementation at Elandsrand, and the outcome to date. The results obtained using sequential grid mining have been evaluated from seismic data, productivity statistics, and rock related accident statistics. Comparisons with neighbouring mines namely Western Deep Levels and Deelkraal, have been made to substantiate the improved conditions observed at Elandsrand, and to compare sequential grid mining directly with longwall mining in similar conditions and at similar depths. Sequential grid mMining is shown to be cost-effective, more flexible, and safer than long wall mining in similar conditions, at least on the Ventersdorp Contact Reef in the western part of the Carletonville Goldfield. More work is necessary to establish it as the favoured mining strategy at depths exceeding 3000 metres, and for reefs other than the Ventersdorp Contact Reef.

Jan 1, 2000

By S Tjerry, S Szylkarski, B Willis-Jones, I Dubinski

Achieving relinquishment of mining sites is becoming increasingly difficult, with a study by Byrne (2013) estimating that out of a set of 73 mining and metals sites around the world, less than ten per cent had achieved relinquishment over the last 50 years. The median forecast duration of time allocated for water management in the closure plans studied was approximately ten years, while the reality at the actual sites was a median duration of 22 years to date and growing.One of many challenges in mine closure planning is managing geomorphic stability of stream diversions. The challenge is to be able to demonstrate to regulators and the community that a proposed rehabilitation solution will be stable over the long term.During the mining operations, local watercourses and surface drainage can be diverted and managed using engineered channels and drainage structures. However, during the post-operations stage, the engineered channel and drainage may require modification to a more natural condition that will facilitate long-term environmental sustainability. Achieving geomorphic stability of a diverted watercourse is difficult, with the majority of stream diversions in Australia failing to meet expected design performance (White et al, 2013). Geomorphic stability is a complex process that is controlled by a combination of factors, including hydraulic flow conditions, geotechnical conditions and vegetation.This paper proposes that a fully dynamic and deterministic geomorphological model can be used to supplement the traditional empirical approaches to assess long-term geomorphic stability. The paper provides an overview of the geomorphological models and how they have been applied in engineering studies for forecasting future evolution of stream systems. The paper discusses the potential application of these models in the mining industry for forecasting and optimising surface water structures for the operations and post-operations stages of the mining cycle.CITATION:Szylkarski, S, Willis-Jones, B, Tjerry, S and Dubinski, I, 2016. Improving the reliability of stream diversion designs in mine closure plans using a morphological tool, in Proceedings Life-of-Mine 2016 Conference, pp 212–215 (The Australasian Institute of Mining and Metallurgy: Melbourne)

Jun 28, 2016

By A. Baydusa, K. Barron

A theory for cutter roof failure in coal mines has been developed. It is shown that the key parameters influencing the likelihood of cutter roof failure are the rock strength, the vertical stress (depth) and the horizontal to vertical stress ratio. The support force necessary to prevent the development of this mode of failure is, in most cases, so high as to make it impractical to apply a force of such magnitude. It is, therefore, better to allow the failure to occur and to design a bolting system to support the dead weight load of the failed block. This is probably best achieved by using a yielding bolt system (such as split set bolts) which is anchored well beyond the top of the potential failure block. The maximum potential caving height determined from this theoretical approach compares favourably with field observations reported in the literature.

Jan 1, 1999

By T. M. Morris

THE rate of flotation of solid particles determines the percentage recovery of these particles which can be obtained during a given time interval. It is an established fact that the recovery is greatest in the fine sand size range and falls off on either side of this size range. It is also a fact that the history of the development of concentrating devices has been marked by a desire to recover the small particles: tables were developed to treat material too fine to recover in jigs; flotation recovers material too fine to be recovered on tables. As our ores become more complex and require finer grinding, a greater proportion is in the slow-floating size range. It is therefore of economic importance that the effect of size of particle upon rate of flotation be studied. Many studies have been made along this line. The present investigation attempts to quantify the rate of flotation by means of a rate equation, and then compares the effect of size upon the rate of flotation. It was found that the flotation rate constant increases with increase in diameter of particle in the size range smaller than the optimum size particle. The flotation rate constant decreases with an increase in diameter of particle in the size range coarser than the optimum size particle. This suggested two different actions taking place, and these are discussed. Suggestions for further work in order to improve the flotation rate of both fine and coarse particles are made. The usual method of expressing rate is to plot cumulative percent recovery vs time. If we knew the equation of this kind of curve we could replot it as a straight line, and the slope of the line would be an index of the magnitude of the flotation rate. It is easier to compare straight lines than curves. The basis used for obtaining the equation of the recovery vs time curve is the law of mass action. This law states that the "rate of a chemical reaction is proportional to the `active masses' of the reacting substances present at any given time." This, of course, implies a constant temperature. We can modify this for our use by using the term flotation reaction instead of chemical reaction and the term floatable particle instead of active masses. [ ] where C denotes concentration of particles, t denotes time, k denotes a proportionality constant. If the exponent of C is 1, then we have a first order rate reaction. If the exponent of C is 2, then we have a second order rate reaction. For a first order rate reaction, we obtain, after integration, [ ] where k is again a proportionality constant and has the dimension of the reciprocal of time. C, is the original concentration or weight or percentage of floatable particles. C, is the concentration or weight or percentage of floatable particles remaining in the flotation cell after time t. If all particles of the mineral to be floated were capable of floating we could take C, as 100 and C, would be 100 - R, where R is the cumulative percentage recovery of mineral at time t. But it is not usual for all of the particles to be capable of floating, because of surface oxidation, etc. Therefore, we must introduce a factor x which is the percentage of unfloatable mineral. Hence the first order equation is [ ] It has been proposed by various workers1-3 that this equation describes the flotation rate. More recently N. Arbiter4 states that a second order rate equation fits the recovery vs time curve more closely than does a first order rate equation. Morris' has discussed the evidence presented for this statement. It has been concluded that the first

Jan 1, 1952

The Government of New South Wales has granted approval for Hunter Valley Aluminium Pty., Ltd" and Alcan Australia Ltd. to proceed with the building of an aluminum smelter at Lochinvar and the expansion of a smelter at Kurri Kurri. Construction of the smelters is expected to start as soon as the consortium is finalized in early 1982. The smelters are part of a plan to increase Australia's aluminum capacity fivefold to 1.5 million tons by the mid-1980's. CHILE Construction of a $61 million lithium carbonate plant in northern Chile will be started in early 1983 by Sociedad Chilena del Litio Ltdao Production is scheduled to begin in 1984 with an initial capacity of 6,350 tons of lithium carbonate produced from a brine deposit. The firm is owned 55% by Foote Mineral Company, of Exton, Pennsylvania, and 45% by Corporacion de Fomento de la Produccion-CORFO.

Jan 1, 1982

By Robert L. Anderson

A shaft or slope bottom is that portion of a mine that is contiguous to the bottom of the main hoisting shaft or slope. It includes the terminal tracks for loaded-and empty-car storage, locomotive run-arounds, locomotive shops, charging stalls, supply rooms, pump rooms, waiting rooms, and offices for dispatcher, weighmen, and foremen. This paper will deal principally with the arrangement of tracks and the handling of cars on the bottom, but includes other data that may be affected, directly or indirectly, by the bottom lay-out. All of the mines discussed are being operated in comparatively level coal seams.

Jan 1, 1937

By R. C. Woodhead

During past years the most difficult problem presented to the coal preparation engineer has been the cleaning of fine coal, The problem of handling fine coals is more involved than the coarser sizes, because the higher surface moisture contents complicate screening, and the larger percentages of fine near size material complicate separating problems. Low gravity high capacity specific gravity separations were not possible with equipment available at that time. Coal preparation engineers did develop machines that would effect a separation of the fines, however, to maintain a high efficiency, the point of separation had to be a relatively easy one. The more difficult separations were made in the coarse coal sections of the washeries, and the fines were generally cleaned at higher specific gravities in equipment such as tables, feldspar jigs, rheolaveurs, etc. Clean products from both processes were mixed, resulting in overall lower ash contents, During the last several years, however, we have been confronted with increasing demands ZGP more precise separations of finer coals to still further reduce ash contents. A study was begun and a program instigated to improve the cleaning of fine coal, The result of this endeavor was the development of the new cycloidal heavy medium cyclone.

Jan 1, 1964

By V. K. Alves, P. Lacoste-Bouchet

"In order to improve ball milling efficiency in pellet plants, Companhia Vale de Rio Doce (CVRD), experiments new technologies in Santa Luzia (M.G. – Brazil) research center facility. Pilot grinding tests have been conducted with small grinding media, size distributed from 5 to 12 mm (millpebs). Tests have been conducted on an ore from CVRD’s Usina 5 Nibrasco, in Vitória, Brazil. A sample of the regrind circuit feed was obtained and used in comparative grinding tests with two types of grinding charges: 100% balls and a blend of 70% balls and 30% millpebs by weight. According to the balls charge, the grinding media mixture has showed the following benefits: reduction of the power draw, increase of the mill capacity, reduction of the circulating load and a slight increase of the ore blaine. Following the results obtained, CVRD is going to further explore the regrinding potential of millpebs.INTRODUCTIONCompanhnia Vale de Rio Doce (CVRD) is the largest mining company in Brazil. CVRD is involved in iron, base metals, gold and bauxite industries. Moreover, CVRD operates several pelletizing plants in Brazil and in other countries.CVRD’s Center for Mineral Development (CDM) is equipped with a comminution laboratory where studies are carried out that are aimed at the optimization of existing plants at several mining sites as well as the technical and economical feasibility of new projects. The laboratory is equipped with a complete pilot plant with several mills and classifiers, enabling continuous grinding tests in grinding circuits of almost any capacity and characteristics.In January 2004, CVRD undertook to conduct pilot tests at CDM for evaluating the effect of small grinding media on ball milling efficiency. Generally speaking, it is not unusual for operators to add oversize grinding media in a conventional regrind mill. This fact originates mainly from prohibitive prices of grinding media below 25 mm due to a production capacity decrease of grinding media producers. The size of the grinding media is an important variable affecting ball milling efficiency. Use of oversize balls decreases the number of contact surface between which particles can be caught and broken."

Jan 1, 2005

By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger's and Kick's laws (Bond, 1952). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed. The work input is given by the following equation: [111W = 1 Owi( FP FF(1)] where W is the The energy input to the mill, (kWh/st); F is the Feed size based on 80% passed, (µm); P is the Product size based on 80% passed, (µm); and W; is the Work Index, (kWh/st) Methods to determine the Bond Work Index Various methods have been proposed by Bond to determine the Work Index. There are some problems with the Standard Bond Grindability Test: •Requiring a special mill. •Requiring a relatively large amount of sample. •Being relatively time consuming and tedious. [ ] During the recent years, numerous attempts have been made to install faster and easier methods to determine the Work Index. A complete description of these methods is presented by Yap et al. (1982). Determining the Bond Work Index by a new size ball mill During the recent years, researchers at the University of Tehran determined the Bond Work Index, using the Anaconda simplified methods by its Denver 305- x 127-mm (12 x 5-in.) laboratory wet ball mill. The equation used to calculate the Work Index for the above mill containing 16.12 kg (35.5 lbs) of steel balls is as follow: [1 _ 1-1W;=0.54x [P [)(2)] In addition, an attempt has been made to determine the Bond Work Index similar to standard Bond ball mill procedure but by a different ball mill, smaller and in consequent portable and less laborious than the Bond ball mill. Method description Just as the standard Bond grindability test, this method is a locked-cycle dry grinding and screening process that is carried out until steady state condition is reached. Test equipment The test is conducted in a special ball mill, called the New Size Ball Mill (NSBM). The mill is scaled down with a coefficient of two-thirds from the standard Bond ball mill. It measures 200 x 200 mm (8- x 8-in.) inside (instead of 300 x 300 mm or 12 x 12 in. in Bond ball mill). Just like the Bond

Jan 1, 1995

By M. C. Fuerstenau, J. D. Miller

The response of calcite to flotation with saturated fatty acids and alkyl sulfonates of various chain lengths is presented. The amount of collector re-quired for flotation is shown to decrease systemati-cally with increasing carbon content from 8 to 12-carbon atoms per molecule. This systematic flotation response is correlated with the solubilities of the calcium-collector salts, which were established with a nephelometer. The 14-carbon collectors deviate from this systematic behavior, and no flotation is possible with the 16-carbon homologs. As determined by infrared analysis, the mechanism of collector adsorption involves a specific chemical reaction be-tween the collector and the surface. A reaction of the form CaCO3(s)+ 2 RCOO"=*=Ca(RCOO)2(s) + C03= is suggested. The pH of the system is also shown to be a most important parameter as it controls the carbonate, calcium, and collector ion concentrations in solution and thus controls flotation response. The importance of hydrocarbon chain length of collectors was recognized early, and various studies have been made on this aspect of flotation. Wilkinson, ' for example, noted that as the chain length of alkyl xanthates is increased, the amount of collector required for flotation of sphalerite is decreased. The same effect has been shown in the flotation of galena.2 Similar results were obtained when malachite3 and calcite4 were floated with various fatty acids. DeBruyn5 has shown that a similar phenomenon occurs when quartz is floated with amines at constant pH. Recently, Fuerstenau, Healy, and Somasundaran6 correlated this phenomenon with the van der Waals cohesive free energy of association of the hydrocarbon chains. It seems unlikely, however, that chain association could be involved in the sulfide and carbonate systems with the collector additions employed, especially with the short-chain xanthates. Rather, it would seem that the decrease in collector concentration with increase in chain length noted for flotation is related to the solubility of the metal-collector salt. The purpose of this investigation is to verify this premise and also quantify this phenomenon if possible. Calcite was selected for study with both alkyl sulfonates and saturated fatty acids as collectors. EXPERIMENTAL MATERIALS AND METHODS The collectors used in this investigation ranged from 8 through 14 carbon atoms per molecule. The fatty acids had a reported purity of greater than 99%. The sodium alkyl sulfonates were provided by the Marathon Oil Cot, Littleton, Colorado, and had the

Jan 1, 1968

By David Burgess, Hershiel H. Hayden

Experiments at the Somerset mine, Colo., under an agreement between US. Steel Corp. and the US. Bureau of Mines are described. A test was made of the tube bundle method of mine air sampling which was developed for European longwall mines; some 300,000 running feet of tubing provided continuous monitoring at 38 locations. Carbon monoxide concentrations were obtained with standard deviations of less than 1 ppm and oxygen concentrations with standard deviations of less than 0.2%. A base level of CO index (ppm CO evolved per percent 0, removed) was determined to be zero to 20 through most of the passageways of the mine. However, in one working section where the coal was initially hot (about 180°F within a pillar) the CO index was demonstrably higher, 120 trending upward to about 60 when smoke appeared and the section was sealed. The data are discussed in terms of laboratory experiments on CO evolution from coals.

Jan 1, 1977

By H. F. Walther

Basic electric-are furnace steel production mainly involves the use of two separate types of slags. The first, known as the "melt down" or oxidizing slag, which is in most cases removed from the furnace mechanically, and the heat finished under the second or deoxidizing slag, which is commonly known as the "finishing slag.'' The melt down slag is formed by the addition of lime or limestone to the charge, which combines with the oxides formed during the melting period and the product of any bottom erosion that may take place.

Jan 1, 1935