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By T. B. Armstrong

"The energy situation is extremely fragile and there is a critical need for energy conservation in the short term and development of new technology in the long term. To keep tabs on energy flow, separate electrical metering is being installed in major industries: The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production. IntroductionThe current energy situation is best described as being extremely fragile. The supply of hydrocarbons is being exhausted, while energy needs remain the same in the industrialized nations and are increasing in the Third World.The petroleum era is expected to come to an end in this century, as there will no longer be enough. Gas, coal and electricity will replace some oil as it becomes scarcer during the transitional period before new sources are gradually developed.There is a critical need for energy conservation in the short term. Conservation will, to a great extent be justified by savings, as it will be less expensive to recover energy that is being wasted than to acquire additional energy.The gradual development of new technology is expected to increase energy self-sufficiency in the long term. New energy sources are attracting a growing interest and are becoming more economically feasible and marketable. These rediscovered energy sources include peat, wood, solar, domestic waste and windmills.The sudden increase in energy prices has forced users to look at the way energy is used and find ways to save it. Major industries have found that using separate meters for each department spurs energy responsibility. Energy trim programs now rank in importance with profitability, productivity and with quality.Energy has become such a crucial cost element that better data are essential for planning. Data acquisition varies with plant size and number of utilities. Meters can be read manually or monitored by computerized energy management systems.The relationship of electrical energy input and production levels for a typical uranium mine will be used to demonstrate two steps to energy management. The first step reduces demand through peak load control; the second reduces the waste of energy by monitoring energy used per unit of production.The direct benefit of saving energy is the money saved."

Jan 1, 1980

By F. C. Bond

SIZE of grinding media is one of the principal factors affecting efficiency and capacity of tumbling-type grinding mills. It is best determined for any particular installation by lengthy plant tests with carefully kept records. However, a method of calculating the proper sizes, based on correct theoretical principles and tested by experience, can be very helpful, both for new installations and for guiding existing operations. As a general principle, the proper size of the make-up grinding balls added to an operating mill is the size that will just break the largest feed particles. If the balls are too large the number of breaking contacts will be reduced and grinding capacity will suffer. Moreover, the amount of extreme fines produced by each contact will be increased, and size distribution of the ground product may be adversely affected. If the balls added are too small, grinding efficiency is decreased by wasted contacts that are too weak to break the particles nipped; these largest particles are gradually worn down in the mill by the progressive breakage of corners and edges. Ball rationing is the regular addition of make-up balls of more than one size. The largest balls added are aimed at the largest and hardest particles. However, the contacts are governed entirely by chance, and the probability of inefficient contacts of large balls with small particles, and of small balls with large particles, is as great as the desired contact of large balls with large particles. Ball rationing should be considered an adjunct or secondary modification of the principle of selecting the make-up ball size to break the largest particle present. Empirical Equation In 19521,2 the author presented the following emerical equation for the make-up ball size: B - ball, rod, or pebble diameter in inches. F = size in microns 80 pct of new feed passes. Wi - work index at the feed size F. Cs - percentage of mill critical speed. S — specific gravity of material being ground. D == mill diameter in feet inside liners. K - 200 for balls, 300 for rods, 100 for silica pebbles. Eq. 1 was derived by selecting the factors that apparently should influence make-up ball size selection and by considering plant experience with each factor. Even though Eq. 1 is completely empirical, it has been generally successful in selecting the proper size of make-up balls for specific operations. But an equation based on theoretical considerations should be used with more confidence and have wider application. The theoretical influence of each of the governing factors listed under Eq. 1 was accordingly considered in detail, as described below, and a theoretical equation for make-up ball sizes was derived. Derivation of Theoretical Equation Ball Size and Feed Size: The basis of this analysis is that the largest ball in a mill should be just sufficient to break the largest feed particle into several pieces, excluding occasional pieces of tramp oversize. In this article the size F which 80 pct passes is considered the criterion of the effective maximum particle feed size. The smallest dimension of the largest particles present controls their breaking strength. This dimension is approximately equal to F. As a starting point for the analysis it is assumed that a 1-in. steel ball will effectively grind material with 80 pct passing 1 mm, or with F- 1000µ or about 16 mesh. The breaking force exerted by a ball varies with its weight, or as the cube of its diameter R. The force in pounds per square inch required to break a particle varies as its cross-sectional area, or as its diameter squared. It follows that when a 1-in. ball breaks a 1-mm particle, a 2-in. ball will break a 4-mm particle, and a 3-in. ball a 9-mm particle. This is in accordance with practical experience, as well as being theoretically correct. Confirmation of this reasoning is supplied by the Third Theory of Comminution," which states that the work necessary to break a particle of diameter F varies as F. Since work equals force times distance, and the distance of deformation before breaka4e varies as F it follolvs that the breaking force should vary as F½ These relationships are expressed in Table I, with a 1-in. ball representing one unit of force and breaking a 1-mm particle. This establishes theoretically the general rule used in Eq. 1 that the ball size should vary as the square root of the particle size to be broken. Ball Size and Work Index: The work input W required per ton" varies as the work index Wi, and the

Jan 1, 1959

By L. Kaufman

An equation is derived relating AF a", the difference in free energy between austenite and martensite, to temperature and composition in the iron-chrmnium and iron-chromium -nickel systems. This equation is used to calculate at M,for low-carbon stainless steels. A method for utilizing the energetic equations to calculate M, in stainless steels is suggested. In addition, an explanation is given for the anomalous effect of chromium in forming a y loop concurrent with the lowering of M, in iron-base alloys. It has been demonstrated that the effect of alloying elements on the Ms temperature is strongly dependent upon the influence of the alloying element on the thermodynamic properties of the y and a phases. Notwithstanding the fact that alloying elements affect the Ms temperature by changing the lattice parameters, elastic constants, and inter-facial energies of the a and y phases, their primary role is a thermodynamic one. Thermodynamic analyses have been presented for the iron-carbon'-4 and iron-nickele systems in which equations are derived for Fa and FY, the free energies of the b.c.c. (or b.c.t.) and f.c.c. phases. These equations, which can be obtained with the aid of the equilibrium diagrams, can be used to help predict the Ms vs composition curves. This paper deals with a calculation of the thermodynamic properties necessary to compute AFffLY for iron-chromium alloys. In view of the lack of direct information concerning the activity of chromium in the a and y phases of iron, the regular solution approximationa,g was utilized to bridge the gap. After the necessary equations were derived for the iron-chromium system they were used to calculate AFff— Y for stainless steels by combining them with the previous results obtained for iron-nickel alloys. The results of these calculations offer a means of rationalizing the anomalous behavior of chromium in iron-base alloys. The latter depends on the regular solution approximation in contrast to the explanation offered by zener , which is based upon the "magnetic specific heat effect." THERMODYNAMICS OF THE IRON-CHROMIUM SYSTEM If y represents the atomic fraction of chromium in an iron-chromium alloy, then the free energy of the b.c.c. phase Fa, the f.c.c. phase FY, and the difference in free energy between the f.c.c. and b.c.c. phases of the same composition are given through [3] and Fmy represent the free energies of mixing of the a and y phases; while @era -' and uFea 4Y represent the difference in free energy between f.c.c. and b.c.c. chromium and iron respectively. The values of AEFe0ff4Y in the temperature range between 0" and 900K ( .

Jan 1, 1960

By S. C. Sun, T. Chao

The reaction of gaseous sulfur trioxide with different aluminum-bearing minerals for the conversion of their alumina content into water soluble sulfates has been investigated. Theoretical considerations indicate that this reaction is thermodynamically feasible and that preferential sulfatization can be accomplished. The five minerals tested were gibbsite, boehmite, diaspore, kaolinite, and halloysite. In general, over 90% sulfatization could be achieved in 1 hr when using sodium sulfate as an additive at a temperature between 500° and 600°C. It was found that sulfates of the light alkali and alkaline earth metals exhibited some effectiveness in promoting the sulfa-tizing reaction. Selective sulfatization of alumina in a mixture of calcined halloysite and ferric oxide was accomplished. The gas-solid type of sulfatizing reactions have become increasingly important in recent years for the treatment of the oxides of such nonferrous metals as cobalt,' nickel,' manganese,3 zinc, and copper.4-6 Similar reactions were attempted in the present investigation for the recovery of aluminum from the low grade aluminous ores. Unlike the Bayer process in which silica is a serious contaminant, or the acid processes whereby iron usually causes difficulties, the reactions between gaseous sulfur trioxide and various aluminous minerals could hopefully be selective for aluminum. THEORETICAL CONSIDERATIONS Reactions between aluminum-bearing minerals and gaseous sulfur trioxide have been shown respectively in Table I and Fig. 1 to be thermodynamically favorable for temperatures below 1000°K approximately. Calculations of the thermodynamic functions were where ?G, ?H, ?Cp and ?S are changes in free energy, enthalpy, constant-pressure specific heat, and entropy of the system under consideration, respectively; T is the absolute temperature, R the universal gas constant, and K the equilibrium constant. Thermochemical data so far available for these calculations were based on sources given by T. Chao.7 In Table I, the free energy values indicate that reactions between the gaseous sulfur trioxide and

Jan 1, 1968

By G J. Harbort, Z Pokra

It has been observed in several Jameson cell installations where the source for flotation feed is deslime screens, that the recovery of coal particles greater than 0.5 m is not as great as that of finer material. Consequently, a research project was undertaken at a CHPP in the Bowen Basin, Queensland, to assess the possibility of increasing the recovery of coarser particles (+0.5 mm) within the downcomer of the Jameson Cell. The effect of decreasing turbulence and agitation in a commercial-scale downcomer was investigated to assess the effect on the recovery of both coarse and fine coal particles. This paper details the findings of the testwork, summarising the results relating to differences in the operating parameters within the downcomer.

Jan 1, 2005

By Matthew S. Sanchez, Richard J. Lee, Matthew D. Zock, Drew R. Van Orden

"Current Environmental Protection Agency (EPA) risk models and asbestos fiber counting techniques for monitoring airborne exposures treat all fibers as being equivalent and ignore the effect of dimension on toxicity. While this may have been appropriate when modern analytical methods were not available, the measurement world is now fully capable of making the appropriate distinctions. The importance of width as a determinant of carcinogenicity has been recognized for nearly 30 years within the published literature. Stanton,1 in a landmark study of a suite of fibrous materials, found that durable mineral fibers with widths less than 0.25 mm and lengths longer than 8 mm were causally associated with the induction of tumors. Pott,2 in a later study, postulated a fiber toxicity model that agreed reasonably well with the Stanton findings, but extended the concept to include varying toxicity as a function of length and width. Wylie et al.3 reanalyzed samples used by Davis et al.4 and found that tumor generation was strongly and inversely correlated with fiber width. Crump5 testified increasing fiber length was correlated with disease occurrence. Lippman6 reviewed a number of studies and suggested that mesothelioma was closely associated with long, thin fibers. Berman et al.7 published a risk model, which found that disease was correlated almost exclusively with fibers longer than 40 mm and width less than 0.25 mm. Recently, Loomis et al.8 reanalyzed textile mill samples and found that aspect ratios greater than 40:1 were highly correlated with disease. Incorporation of width into risk models has been limited by the availability of data from operating asbestos plants and, in particular, the observation of an asbestos plant for which there is no observable mesothelioma. There have been only limited studies published on the dimensions of asbestos fibers, with a few serving as the primary references in other literature.9_13 A rare opportunity to collect detailed concentration and dimensional data of airborne crocidolite from an operating asbestos-cement plant using modern analytical procedures arose, providing a chance to supplement the existing literature on airborne fiber dimensions. The asbestos-cement plant is located in the eastern end of Cochabamba, Bolivia along Highway 4. Cochabamba is a relatively large city located in the Andes Mountains, approximately 250 km southeast of the capital La Paz. The city sits in a valley that runs east-west with mountains to the north and south. The crocidolite used in the products made at this facility is mined in central Bolivia, approximately 75 km northwest of Santa Cruz. The crocidolite is a fibrous magnesioriebeckite 14 (blue asbestos). Crocidolite fibers from Bolivia were the subject of prior reports suggesting the fibers are generally longer and thicker than crocidolite from Australia or South Africa.9 At the cement plant, the run-of-mine crocidolite was first pulverized using a hammer mill and then sieved on vibratory sieves to produce fibers of limited size ranges for use in the cement. The sieved crocidolite is mixed with other components, including recycled cement product, to produce a slurry that is vacuum filtered into sheet form. The sheets are then manually molded into a corrugated shape, air dried, and sawed (using circular saws) to accepted dimensions. The sawed sheets are cured in soaking pits for 28 days before being shipped to the consumer. Prior to shipping, the sheets are finished by a light hand sanding followed (at customer request) by applying one or more coats of paint. A detailed description of the cement plant and the industrial hygiene results can be found in Van Orden et al.15"

Jan 1, 2012

By H. H. Reamer, B. H. Sage

The formation volume and the relative volume of the liquid Phase of mixtures of Condensate and gas from five different parts of a field in the Louisiana area have been established experimentally. These studies included gas-condensate ratios as high as 25.000 cu ft/bbl at pressures up to 5,000 psi. The gravities of the tank condensates were between 53.9° and 63.9° API. The compositions of the condensate and gas samples inves-ligated are presented in tabular form. The formation volume. the relative volume of the liquid phase. and the specific volume of the mixtures are available. A rough correlation of the formation volume as a function of pressure. temperature, and gas-condensate ratio has been presented. This correlation i-considered applicable only to the range of conditions covered in this investigation for systems of similar nature and com-position. INTRODUCTION Often a knowledge of the influence of temperature. Pres sure, and relative quantities of condensate and gas upon the volumetric behavior of fluid*. from a given field has been employed in estimating the Conditions existing in the reservoir. The early work of Beecher and Parkhurst' and the more rrcent Studies of the Dominguez' and San Joaqnin Valley Field;: are examples of experimental studies of this character. Correlation of the volumetric behavior of naturally occurring hydrocarbon mixtures has Proved To be feasible' . at low gas-conden-sate ratios-; for fluids*. from a single geographic area. Progres 5 is being made in the Prediction of the volumetric behavior of distillate systems at high gas-condensate ratios and pressure.. However. the estimation of the retrograde". ' dew-point pressure still remains uncertain and probably can best be established by experiment. For this reason it appears of interest to make available experimental information about the volumetric behavior of mixtures of condensate and gas from five different part.; of a single field. This material shows a rough similarity to data for mixtures of condensate and gas from San Joaqnin ' Valley Fields. In the present paper. 110 attempt has been made to coordinate the available information on tbe volumetric behavior of fluids from condensate fields. It appears hopeful that progress can be realized by the utilization of recently developed equa-tions of state.!' However. the primary problem in any such correlation is the proper identification of the components and constituents of naturally occuring systems. Additional experimental information may be required to permit the determination of the values of the constants of the equation of state for References given at end of paper. Manuscriptt received in the office of the Petroleum Branch June 11. 1951. *For present purposes low gas- condensate ratios will he considered as those below 3,000 cu ft/bbl, and the region of high gas-condensate ratio will be limited to compositions containing more than 5,000 cu ft of pas per bbl of condensate. Mixtures Containing between 3,000 and 5.000 CII ft OF gas per bbl will be considered of intermediate gas-condensate ratio.

Jan 1, 1952

By H. P. Greenwald

THIs paper brings forward a discussion that was prepared for the meeting of the Coal Division in Chicago in 1938.1 War in Europe less than a year after that meeting, followed by our defense preparations and entry into the conflict, caused general cessation of large-scale experimentation on the subject, and it is best to confine this paper to events in the United States during the five-year period since the preceding paper was written. The views of members of the Coal Division's Committee2 on Rock-dust-ing for 1943 were obtained before this paper was written and the manuscript was reviewed by them. Extent of Rock-dusting Statistics on the use of rock dust in the United States during the years 1930-1933, appearing on page 97 of Volume 139 of the Transactions are extended in Table I herein through the year 1941. The results through 1933 did not appear encouraging, but there has been improve-ment since that date. It has always been true that the larger operations were the ones that used rock dust. In 1941 only 10 per cent of the mines in the country used rock dust, but they produced 57 per cent of the coal, employed just under 50 per cent of the miners, and accounted for 52 per cent of the man-hours. Unfortunately, collection and compilation of these statistics are slow procedures, and figures dealing with the tremendous war expansion of the industry are not likely to be available for a year or more. In 1941 the five leading states in regard to percentage of man-hours worked in mines using rock dust were: Utah, 92.1; Wyoming, 79.3; Alabama, 77.4; New Mexico, 72.5; West Virginia, 65.3. The first four states were the same as in 1933, but West Virginia displaced Washington in fifth place. On the basis of pounds of rock dust used per ton of coal, the five leading states in

Jan 1, 1944

By H. P. Greenwald

THIs paper brings forward a discussion that was prepared for the meeting of the Coal Division in Chicago in 1938.1 War in Europe less than a year after that meeting, followed by our defense preparations and entry into the conflict, caused general cessation of large-scale experimentation on the subject, and it is best to confine this paper to events in the United States during the five-year period since the preceding paper was written. The views of members of the Coal Division's Committee2 on Rock-dust-ing for 1943 were obtained before this paper was written and the manuscript was reviewed by them. Extent of Rock-dusting Statistics on the use of rock dust in the United States during the years 1930-1933, appearing on page 97 of Volume 139 of the Transactions are extended in Table I herein through the year 1941. The results through 1933 did not appear encouraging, but there has been improve-ment since that date. It has always been true that the larger operations were the ones that used rock dust. In 1941 only 10 per cent of the mines in the country used rock dust, but they produced 57 per cent of the coal, employed just under 50 per cent of the miners, and accounted for 52 per cent of the man-hours. Unfortunately, collection and compilation of these statistics are slow procedures, and figures dealing with the tremendous war expansion of the industry are not likely to be available for a year or more. In 1941 the five leading states in regard to percentage of man-hours worked in mines using rock dust were: Utah, 92.1; Wyoming, 79.3; Alabama, 77.4; New Mexico, 72.5; West Virginia, 65.3. The first four states were the same as in 1933, but West Virginia displaced Washington in fifth place. On the basis of pounds of rock dust used per ton of coal, the five leading states in

Jan 1, 1944

By C. A. Eligwe, F. W. DeVries, A. E. Torma

This study is an investigation on using hydrogen peroxide for tank-leaching of uranium with sulfuric acid. The optimum sulfuric acid concentration was found to be 0.03 mole 11 -'for a 25% pulp density suspension. The order of reaction with respect to sulfuric acid concentration was found to be ns0.75. The best results vith hydrogen peroxide were obtained at a 4.64 molar ratio of 6202 to U in the ore. The order of reaction was found to be zero with respect to hydrogen peroxide concentration. The theoretical maximum rate effect of Hz02 on uranium extraction was derived as Vm = 8.15 x mole &-'min-'. The highest rate experimentally attained was 8.09 x mole kL'min-'. It was found that the uranium extrac- tion was not affected by the pulp density up to the 50% used in this study. However, the final yieldswere significantly higher (over 90%) for leaches carried out in the presence of Hz02 than (less than 90%) for leaches carried out in the absence of H202. The reaction was found to follow a first order course with respect to substrate concentration (pulp density) regardless of the presence of H202. The effect of temperature was more explicit at moderate tempetatures (30 to 55 C) than at lower (10 C) or at higher (70 C) temperatures. At this latter temperature an almost complete extraction was realized even without H202. The apparent activation energies were found to be AEa = 2.68 kJ mole-' in the absence of H202, 0.5 kJ mole-'with Hz02 at 40 C to 70 C. and 4.72 kJ mole-' with Hz02 at 10 C to 40 C. All these activation energy values suggested that the uranium extraction process is diffusion controlled regardless of the presence of H202. Further, the change in activation energy in the presence of Hz02 indicates the possibility of change in the reaction mechanisms at lower and higher temperatures.

Jan 1, 1980

By R. H. Sweetser

There are probably as many variations of the method for blowing-in blast-furnaces as there are furnace superintendents. That some of these variations are poor practice is shown by the troublesome and sometimes disastrous experiences of many furnace-men, even within the last few years. It is the object of this paper to put on record a blowing-in practice that has gradually developed to the point where the results are sure, safe, and satisfactory; and also to bring into discussion several variations that are questionable. The blowing-in of a modern blast-furnace is one of the greatest of metallurgical operations. On its degree of success depends much of the future life of the lining and the subsequent behavior of the furnace; and it involves possibilities of danger to the men engaged in the work. The objects to be attained are the heating of the interior of the furnace to the high temperature necessary for smelting iron; the starting of the smelting process; and the gradual movement of the stock in its downward course. The most critical points are the placing of the kindling-wood, the " bringing down " of the gas, the closing of the iron-notch, and the first cast. The lighting of the first gas, the opening of the cinder-notch for the first flush, and the running of the first cast are anxious times for the superintendent and the furnace-men. Drying-Out the Furnace. As soon as a furnace-lining is completed it should be thoroughly dried out, whether it is to be blown-in soon or not; and the longer it is dried, the better for the furnace. At least two weeks is desirable; but good results can be had with 10 days of firing. The crudest and least desirable method of drying is to build a wood fire in the bottom of the furnace, throwing the wood in through a tuyere-arch, and using the iron-notch for

Jan 1, 1913

By Klimpel. R. R.

Acknowledgments I am honored to be named as a recipient of the Antoine M. Gaudin Award, and I would like to briefly single out four groups who deserve a special thank-you relative to this award. The first group consists of Professor Kawatra, the many international support references that he gathered and the Gaudin Committee itself. I know such things don't come without considerable effort by many individuals, and I'm grateful for their support. The second group consists of the management of The Dow Chemical Company, the managements of various mining companies and, in recent years, the management of G. S. Technologies, all of whom supported me, financially and otherwise, in my grinding, flotation and other particle-processing work. Operations people are the most overworked and under recognized people in the world, whether they be in the mining or chemical industry. Obviously, without extensive industrial support, especially from operating personnel in the mining industry, almost nothing about the work contained in this lecture would have, or could have, happened. The third group consists of a number of coworkers who have played key roles in specifically implementing the new flotation-product portion of this lecture. Of these, Bob Hansen, Don Leonard, Basil Fee, Simon Isherwood, Fernando Cones and Marty Meulendyke deserve special recognition. They have not only been dedicated, but they have been great to work with. Also in this third group are many members of the international academic community, especially those at Penn State. Professor Leonard Austin must be singled out because he took me as a young mathematically oriented chemist and taught me to be an effective process-oriented engineer. Without this experience, I would not have had the over 30-year career as an engineer, which I have loved. In addition, Len planted the size-reduction-research virus in my blood, which I can't seem to get rid of. I also need to mention Professor Frank Aplan, whose obvious love of mineral processing and froth flotation helped to create my interest in flotation. Thus, for those of you who don't like my approach to flotation, it's all Frank's fault. The fourth, and certainly not the least, group to thank is my wife, Charlene, and our two sons, Richard and Steven. No one could have a more supportive family. In every sense of the word, my wife has been my partner from the beginningwe were married at the end of my freshman year in college. She has sacrificed much for me, and I will never really be able to express the depth of my gratitude for Charlene's encouragement, loyalty and support over the past 36 years. General comments on froth flotation The industrial process of froth flotation: Where is it going? I'm here to tell you loud and clear that froth flotation is alive, well, and growing. The inherent characteristics of high mineral throughput per unit of capital, low operating cost per unit of feed, high mineral specificity and high operating flexibility have guaranteed froth flotation an important past and a growing future. As with any evolving technology, there are always trends at the industrial level that are primarily driven by economics. The current trends include: •the movement towards larger and larger equipment, •more alternative equipment designs (e.g., the column cell),

Jan 1, 1996

By H. S. Wang, Y. P. Gupta

Diffusion of zinc and germanium in liquid silver is measured in the temperature range 975" to 1400°C by use of the capillary-reservoir technique. The ex-perinzental results are discussed in terms of the fluctuation and the "hole" theories of liquid diffusion, assunzing the solute as a point charge equal to the difference in the valences of the solute and solvent atoms. The additional coulombic interaction term due to the excess solute charge is evaluated in terms of the Thomas-Fenni approximation. The theoretical and experimental results agree with both the fluctuation and "hole" theories of liquid diffusion. No perceptible effect on solute diffusion in liquid silver of the size difference between the solute and solvent atoms/ions is apparent. DIFFUSION in liquid silver of electropositive solutes indium, tin, antimony, and homovalent gold has been previously reported by Leak and Swalin1 and Gupta,' respectively. These solutes have nearly the same atomic size but differ principally from one another in excess valence (+2, +3, +4, respectively). The observed differences in the temperature dependence and the rate of solute diffusion were explained in terms of an added coulombic interaction between the solute and the solvent atoms. In the case of diffusion of gold in liquid silver, the data were found to be consistent with an effective valence for gold of -0.85, as proposed by Le Claire.3 Recently, Wang and Gupta4 have reported a study of diffusion of ruthenium, iron, and cobalt in liquid silver. The results of diffusion of these solutes and their temperature dependence were found consistent with the hole theories and the fluctuation theories of liquid diffusion, if ruthenium, iron, and cobalt were diffusing in liquid silver with relative valences -2, -1.3, and -0.5, respectively, instead of -3, -3, and -2 as suggested by the relative position of these elements in the Periodic Table of Elements. Furthermore, the predicted valences for these solutes were observed to be consistent with the sparsely available information on the magnetic properties of some of the dilute alloys of copper and silver with these solutes. In theoretical calculations of solute diffusivity, the solute was assumed as a point excess charge equal to the difference in the valences of the solute and solvent atoms. Both the assumption of a point excess solute charge and the use of the Thomas-Fermi approximation to calculate the potential of a solute charge are less precise for the case of electronegative solutes. Therefore, it was not possible to infer from the data on ruthenium, iron, and cobalt an effect, if any, on solute diffusion of the relative differences in either the Period or the size of the solute and solvent atoms or the solute atoms themselves. A study of the diffusion of Period 4 elements, specifically zinc and germanium was, therefore, undertaken since zinc and germanium, respectively, have relative valences of +1 and +3 in liquid silver and the data obtained on diffusion of these solutes could be compared with the available data on diffusion of Period 5 electropositive solutes, indium, tin, and antimony in liquid silver. From such a comparison, it was hoped to learn something about the effect, if any, of the ion size or the structure of the ion core on impurity diffusion in liquid silver. EXPERIMENT The diffusion of tracer zinc and germanium in liquid silver was studied by the capillary reservoir technique, modified adequately for use at high temperatures. In principle, capillaries containing liquid silver

Jan 1, 1970

By John Ward Smith

This paper describes and evaluates a method that the Federal Bureau of Mines used to determine the ultimate composition of the organic material in 10 samples of oil shale from the Mahogany zone of the Green River formation in Colorado and Utah. Since interest has centered on the Mahogany zone, the richest section of the deposit, these 10 samples were chosen to provide a broad representation of this zone with the objective of obtaining generally valid ultimate composition values. The most probable organic composition for the Mahogany zone and the 95-percent confidence limits for each component average (in weight-percent of organic material) are: Carbon 80.5 ± .4 Hydrogen 10.3 ± .1 Nitrogen 2.4 ± .1 Sulfur 1.0 ± .1 Oxygen 5.8 ± .5 As only minor variations in composition were observed, the organic material of this particular bed is quite similar over a wide geographic area.

Jan 1, 1961

By K. K. Dehn

This study focused on determining the stability of thin shotcrete fill fences at a hard rock mine in northwestern Ontario during increased pour rates of paste backfill. Determination of the stability of the fence is needed for both personnel safety and to avoid failure of a fence. Failure of a fill fence during backfill activities would have severe impacts on production and potential hazards to nearby workers. Approved fill fences at the mine are a slender design of 10.5 centimeters of shotcrete applied to a rebar framework with burlap lined ?bedspring? backing. Approved mine standards dictate that fill rates may not exceed 1.5 meters of paste depth per 10-hour shift when pouring behind a flat fill fence. To decrease the backfill cycle time the study looked at fence stability during increased pour rates of 2.5 meters paste depth poured within a 10-hour shift. An instrumentation design was developed including pressure cells, strain gauges, and tilt-meters that were built into experimental fences to observe the effects of increased pour rates on lateral wall pressures, strain development in the rebar and shotcrete reinforcement, and physical deformation patterns along the face of the fence. The data was analyzed for overall stability.

May 1, 2007

By Mike Horn, Keith Meade, Marvin L. Smallwood

Technological changes in programmable logic controllers (PLC)'s and the need to improve the architecture of Mine Wide Monitoring Systems (MWMS) has resulted in the design of a PLC based system at the Virginia Division of Island Creek Coal Company. The system which is presently in the installation phase, is a joint effort of Island Creek Coal Company, Westinghouse Electric Supply Company and Mining Controls Incorporated. The system uses an ethernet based network for graphic display of system data operating under the Genesis software package. Telemetry for the system utilizes the Seimens/Westinghouse L1 network operating over two twisted pairs. Remote telemetry uses fiber optic links and modems operating over standard telephone lines. 500 and 2000 series PLC's provide distributed data processing for the system with hierarchial processing operating under the Genesis software on the ethernet network for individual mine site monitoring along with the remotely located division office monitoring. System control and data acquisition is provided for underground MWMS operations and on the surface for a vast methane gathering system that is used for degasification of the mines and collection of pure methane for sale to the Columbia Gas Company as natural gas. The inability of industrial MWMS companies to keep up with the vast changes in technology, along with economic considerations associated with expanding the existing MWMS created the need for the development of the PLC based system to replace the older industrial monitoring systems.

Jan 1, 1993

By Oliver Bowles

WHILE vast quantities of limestone and dolomite. are used in metallurgy, the estimated production in 1926 being 23,860,000 tons, there are many problems connected with their use which have not received adequate study. The literature of metallurgy is notably lacking in comprehensive discussions of fluxing or furnace stone. Approximately 900 lb. of limestone is used for every long ton of pig iron produced in the blast furnace, but this important constituent of the charge receives little attention compared with the intensive study that has been applied to ores and fuels, the other important constituents of the charge. In an effort to elucidate some of the doubtful points, the writer has made a wide review of metallurgical literature, and has had the disappointing experience of finding not more than a paragraph or two on metallurgical stone' for each two or three thousand pages of literature reviewed. Furthermore, the limited information supplied is in some instances contradictory and even absurd. For example, a correspondent in a reputable metallurgical magazine classifies magnesia and silica together as impurities in flux "that must be melted and run off as slag." From the furnace operator's viewpoint it is desirable that as complete knowledge as possible be obtained concerning the effects of chemical composition and physical character of the stone on its action in the furnace and on the quality of the metal produced. A better knowledge of these features is advantageous also to the quarry operator in that it enables him to work more intelligently in producing stone that will best satisfy the metallurgists' requirements. The following pages comprise a brief compilation of data now available. Much of the information is well established and probably acceptable to most metallurgists. On some points there is room for debate, and the chief purpose in presenting a paper at this time is to arouse discussion which will throw more light on controversial points and thus assist the author in preparing a more comprehensive and accurate discussion of metallurgical limestone problems.

Jan 1, 1928

By Oliver Bowles

WHILE vast quantities of limestone and dolomite are used in metallurgy, the estimated production in 1926 being 23,860,000 tons, there are many problems connected with their use which have not received adequate study. The literature of metallurgy is notably lacking in comprehensive discussions of fluxing or furnace stone. Approximately 900 lb. of limestone is used for every long ton of pig iron produced in the blast furnace, but this important constituent of the charge receives little attention compared with the intensive study that has been applied to ores and fuels, the other important constituents of the charge. In an effort to elucidate some of the doubtful points, the writer has made a wide review of metallurgical literature, and has had the disappointing experience of finding not more than a paragraph or two on metallurgical stone for each two or three thousand pages of literature reviewed. Furthermore, the limited information supplied is in some instances contradictory and even absurd. For example, a correspondent in a reputable metallurgical magazine classifies magnesia and silica together as impurities in flux "that must be melted and run off as slag." From the furnace operator's viewpoint it is desirable that as complete knowledge as possible be obtained concerning the effects of chemical composition and physical character of the stone on its action in the furnace and on the quality of the metal produced. A better knowledge of these features is advantageous also to the quarry operator in that it enables him to work more intelligently in producing stone that will best satisfy the metallurgists' requirements. The following pages comprise a brief compilation of data now available. Much of the information is well established and probably acceptable to most metallurgists. On some points there is room for debate, and the chief purpose in presenting a paper at this time is to arouse discussion which will throw more light on controversial points and thus assist the author in preparing a more comprehensive and accurate dis¬cussion of metallurgical limestone problems.

Jan 1, 1928

By M. R. Khalesi, Y. Kianinia, A. Seyedhakimi, F. Soltani

"The feed ore to Agh-Darreh gold processing plant (Takab City, Iran) contains 2 ppm Au, 7 ppm Ag, and 120 ppm Hg. After cyanide leaching of the feed at pH 10.4, 80% of the total mercury (in the form of cinnabar and metacinnabar) goes to the tailings dam. In this research, separation of mercury from the tailings by flotation was investigated. The effects of collector type and dosage, pH, and the number of cleaner stages were studied. The results showed that after two stages of cleaning a 40–62% recovery of Hg at a grade of 14.3% Hg is attainable. Recycle water from a test in which 50 g/t amyl xanthate and 30 g/t pine oil was used in the flotation stage showed no adverse effect on the leaching and adsorption of gold onto activated carbon.IntroductionMercury, most commonly found in the nature as the mineral cinnabar (HgS), is a hazardous substance because of its mobility and toxicity (Kyle et al., 2012). Roasting is the wellestablished treatment method for mercury ores, in which the sulphur is burned and volatilized and mercury recovered subsequently by cooling and condensation. However, for low-grade ores, flotation is mainly applied for primary concentration due to its comparatively low costs, low environmental hazards, and flexibility regarding feed variations (Bulatovic, 2007). Particularly, cinnabar can be easily recovered by flotation using pine oil and a collector such as xanthate (Bulatovic, 2007; Crozier, 1991). Floatability of cinnabar is reduced due to surface oxidation when it is exposed to the atmosphere after crushing. This problem can be eliminated by adding copper sulphate and lead nitrate or acetate. The best pH for cinnabar flotation is in the range of 6.5 to 8.5 according to Erspamer and Wells, (1954), who reported that cinnabar and stibnite can be floated with long-chain xanthate collectors and an alcohol frother, with cumulative recoveries of 97.8% and 94.7%, respectively. Lead acetate was used as an activator."

Jan 1, 2017

By David Dvorkin, Thomas C. Anderson

INTRODUCTION A major function of the Mine Safety and Health Administration (MSHA), Denver Safety and Health Technology Center (DSHTC), Ventilation Branch, is to provide engineering analyses of underground mine¬ ventilation systems for coal, metal, and nonmetal in¬spection departments in accordance with field surveys and data analyses. Health and safety problems such as the control of diesel contaminants, dust, methane, radon, and radon daughters all are dependent upon ventilation. It is necessary for the ventilation engineers at DSHTC to have a complete understanding of the various and highly complex ventilation systems used in the nation's mines. Only through the proper assembly and analysis of data can the engineer determine the causes of prob¬lems and make suitable recommendations. Furthermore, the engineer must have the ability to predict the results of any such recommendations. In the study of the ventilation system at any indi¬vidual mine, it is common practice to collect data on airflow quantities and pressures at key points throughout the mine. However, these data give only a single point¬-in-time view of the system. As the mine develops and progresses, the system undergoes numerous changes; as the mine becomes more extensive, the ventilation re¬quirements increase, and the costs rise. For proper utilization of the available resources, it is necessary for the engineer to have a sound engineering basis for deci¬sions on how best to ventilate the mine. To evaluate the system as a whole, particularly re¬garding possible system changes, it has become necessary to rely on computers to perform the calculations in¬volved in analyzing the network. This chapter sum¬marizes the implementation of such computer capabili¬ties at DSHTC, as well as the services that are available from DSHTC. ANALOG SIMULATIONS AND ANALYSES OF VENTILATION NETWORKS As early as 1954, the US Bureau of Mines (USBM) in Pittsburgh, PA, was using an analog computer as an aid in analyzing ventilation systems. The Mcllroy Fluid Network Analyzer used filament-tube resistance elements and an electrical power supply to simulate mine-ventila¬tion systems electrically. This system is shown in Fig. 1. Analog simulation has proven to be a practical tool, providing fast analyses of complicated problems after the initial setup has been completed. Advancements in electronic technology over the past 24 years have made possible the construction of an analog computer that has streamlined several operating features of the previous machines. The DSHTC Ventilation Branch now has an improved electronic analog computer to assist in the analyses of ventilation systems. General Features As shown in Fig. 2, the analog computer used at the DSHTC Ventilation Branch has three "element" cabi¬ nets, a console, and a printer. The cabinets contain the fan elements and the airway-resistance elements that are used to "program" an analysis problem; each element is independent of the others. The console contains the "patchboard" and the controls necessary to operate the analog computer. The patchboard connects the inde¬pendent elements into the circuit appropriate for simu¬lating the ventilation system. The total system is controlled from a panel on the console, while the value corresponding to each element is set on the individual controls located on the face of each element. The control panel has three visual output displays that give the quantity of flow, the pressure of the output junction with respect to a datum, and the differential pressure across each element. The control panel also allows manual selection of the element to be monitored, or automatic scanning and printing of values for the entire system. Capabilities The analog computer can simulate an existing mine¬-ventilation network. This is accomplished by collecting field data at the mine to give a complete pressure-¬quantity network. Resistance values for each airway branch may be determined from the quantity of flow in the branch and from the pressure differential between the end-point junctions of the branch. The basic rela¬tonship to determine the resistance is expressed as: H=R •Q2 where H is the pressure differential, R is the resistance, and Q is the quantity of air in the flow. After the resistance values are known for all of the airways in the circuit, the elements on the analog com¬puter can be set to the corresponding values. The elements are designed to reflect the basic relationships, with electrical potentials corresponding to pressures and

Jan 1, 1982