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By Victor Formanek, Paul Chataignon, Maurice Rey

IT is found when floating oxidized lead ores by sulphidization, that the presence of calcium salts in the water, is usually detrimental and lowers the recovery. This effect is particularly marked in dry countries such as North Africa, where the waters often carry large amounts of calcium sulphate and where the ore may even contain gypsum. The effect of calcium salts is readily visible. Whereas in their absence cerussite is quickly stained brown and then black by sodium sulphide, in their presence the mineral remains very light in color. A similar effect is produced when barium sulphide is used as a sulphidizing agent instead of sodium sulphide. Magnesium salts have little or no effect and even tend to reduce the detrimental effect of calcium salts. A study of this phenomenon indicates that it is due to the precipitation of calcium or barium carbonate in contact with the mineral simultaneously with the formation of lead sulphide. The chemical reactions can be interpreted as: PbCO3 + Na2S = PbS + Na2CO3 CaSO4 + Na2CO3 = CaCO3 + Na2 SO4 They might also be written: PbCO3 + S = + Ca++ = PbS + CaCO3 The precipitation of calcium carbonate can be followed by the lowering of the pH with which it is accompanied. Magnesium carbonate is more soluble than calcium carbonate and usually does not precipitate under the conditions prevailing. It is interesting to note that calcium salts have no effect on anglesite (lead sulphate) because calcium sulphate is soluble, but barium salts hinder the sulphidization of anglesite because of the precipitation of barium sulphate. Remedies When calcium sulphate is present in large amounts, the softening of the water with soda ash is usually too expensive to be considered, but the precipitation of the objectionable calcium carbonate can be prevented in two different ways. One is the use of sodium hydrosulphide instead of sodium sulphide. This salt gives a lower pH than sodium sulphide and does not bring about the immediate precipitation of the calcium which remains in solution as calcium bicarbonate. The other procedure is to add ammonium salts such as the sulphate or chloride which have the property of increasing markedly the solubility of calcium carbonate. Ammonium salts have other effects such as cutting down conditioning time and accelerating flotation. They should be added to the flotation cells rather than to the ball mills. Table I. Effect of Sodium Hydrosulphide and/or Ammonium Sulphate on the Flotation of Two Oxidized Lead Ores Mibladen Ore4 La Plagne Ored Con- Tall- Con- Tail-centrate ing centrate ing Reagents Pb, Pct Pb, Pct Pb, Pct Pb, Pot Without addition of CaSO, NaS 56.0 0.68 43.4 0.65 NaSH 56.2 0.60 40.4 0.50 Na2S + (NH4)2SO4d 54.6 0.56 NaSH + (NH4)2SO4d 43.0 0.66 With addition of 40 lb/ton CaSO4 Na2S 39.4 3.85 34.0 3.21 NaSH 49.1 0.70 49.6 0.88 Na2S + (NH4)2SO4d 53.9 0.85 NaSH + (NH4)2SO4d 51.0 0.45 41.1 0.78 In table I are given the results of representative tests on two different oxidized ores. They show that the strongly detrimental effects of calcium sulphate can be offset by the two procedures outlined above. Sodium hydrosulphide is now used regularly on a mill scale on certain ores. Tests are being carried out with ammonium salts. It should be noted that malachite is subject to influences similar to cerussite. One final word of caution—when the ore is rich in primary slime, which is flocculated by the calcium salts, it may be indispensable to remove these by washing or precipitation with sodium carbonate instead of keeping them in solution by the above methods. Acknowledgment We wish to thank Socibtk Miniere & Metallurgique de Penarroya and Minerais et Metaux, for permission to publish these results.

Jan 1, 1951

By Maurice Rey, Victor Formanek, Paul Chataignon

IT is found when floating oxidized lead ores by sulphidization, that the presence of calcium salts in the water, is usually detrimental and lowers the recovery. This effect is particularly marked in dry countries such as North Africa, where the waters often carry large amounts of calcium sulphate and where the ore may even contain gypsum. The effect of calcium salts is readily visible. Whereas in their absence cerussite is quickly stained brown and then black by sodium sulphide, in their presence the mineral remains very light in color. A similar effect is produced when barium sulphide is used as a sulphidizing agent instead of sodium sulphide. Magnesium salts have little or no effect and even tend to reduce the detrimental effect of calcium salts. A study of this phenomenon indicates that it is due to the precipitation of calcium or barium carbonate in contact with the mineral simultaneously with the formation of lead sulphide. The chemical reactions can be interpreted as: PbCO3 + Na2S = PbS + Na2CO3 CaSO4 + Na2CO3 = CaCO3 + Na2 SO4 They might also be written: PbCO3 + S = + Ca++ = PbS + CaCO3 The precipitation of calcium carbonate can be followed by the lowering of the pH with which it is accompanied. Magnesium carbonate is more soluble than calcium carbonate and usually does not precipitate under the conditions prevailing. It is interesting to note that calcium salts have no effect on anglesite (lead sulphate) because calcium sulphate is soluble, but barium salts hinder the sulphidization of anglesite because of the precipitation of barium sulphate. Remedies When calcium sulphate is present in large amounts, the softening of the water with soda ash is usually too expensive to be considered, but the precipitation of the objectionable calcium carbonate can be prevented in two different ways. One is the use of sodium hydrosulphide instead of sodium sulphide. This salt gives a lower pH than sodium sulphide and does not bring about the immediate precipitation of the calcium which remains in solution as calcium bicarbonate. The other procedure is to add ammonium salts such as the sulphate or chloride which have the property of increasing markedly the solubility of calcium carbonate. Ammonium salts have other effects such as cutting down conditioning time and accelerating flotation. They should be added to the flotation cells rather than to the ball mills. Table I. Effect of Sodium Hydrosulphide and/or Ammonium Sulphate on the Flotation of Two Oxidized Lead Ores Mibladen Ore4 La Plagne Ored Con- Tall- Con- Tail-centrate ing centrate ing Reagents Pb, Pct Pb, Pct Pb, Pct Pb, Pot Without addition of CaSO, NaS 56.0 0.68 43.4 0.65 NaSH 56.2 0.60 40.4 0.50 Na2S + (NH4)2SO4d 54.6 0.56 NaSH + (NH4)2SO4d 43.0 0.66 With addition of 40 lb/ton CaSO4 Na2S 39.4 3.85 34.0 3.21 NaSH 49.1 0.70 49.6 0.88 Na2S + (NH4)2SO4d 53.9 0.85 NaSH + (NH4)2SO4d 51.0 0.45 41.1 0.78 In table I are given the results of representative tests on two different oxidized ores. They show that the strongly detrimental effects of calcium sulphate can be offset by the two procedures outlined above. Sodium hydrosulphide is now used regularly on a mill scale on certain ores. Tests are being carried out with ammonium salts. It should be noted that malachite is subject to influences similar to cerussite. One final word of caution—when the ore is rich in primary slime, which is flocculated by the calcium salts, it may be indispensable to remove these by washing or precipitation with sodium carbonate instead of keeping them in solution by the above methods. Acknowledgment We wish to thank Socibtk Miniere & Metallurgique de Penarroya and Minerais et Metaux, for permission to publish these results.

Jan 1, 1951

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By R. J. Davidson, M. V. Nkosi, V. Veronsee

When activated carbon is used for the recovery of gold and silver from Witwatersrand gold-plant solutions, severe fouling of the carbon by calcium carbonate is experienced. Also, difficulty is encountered in the retention of the adsorbed silver on the carbon because it is displaced by gold. The acidification of plant pregnant solutions before they are contacted with carbon was therefore investigated as a means of overcoming these shortcomings. The results from batch laboratory experiments on the effect of pH on gold adsorption from plant solutions and from synthetic solutions of high ionic strength indicated reaction rate optima in the region of pH 5 and below pH I. The effect of pH on the equilibrium capacity constant for gold adsorption onto activated carbon was also studied, using both a plant solution and a synthetic solution of high ionic strength. In the case of the plant solution, the results suggested only a marginal change in carbon capacity for gold in the pH range 12 to 6, with significant increases in capacity at lower pH values. In the case of the synthetic solution of high ionic strength, the capacity constant was found to increase steadily as the pH was lowered from 12 to I. It was also found that the gold adsorption capacity of the carbon was significantly better for the synthetic solution. No precipitation of gold from the pregnant solution (gold content approximately 6 g/t) was noted as the pH was decreased to values as low as ,. However, about 25 per cent of the 0,6 g/t silver content was found to have precipitated when a pH value of 3,0 was attained. Further substantial precipitation of silver took place as the pH was lowered further. The investigation was extended by the passing of suitably acidified gold-plant solutions through one or more columns containing activated carbon. Preliminary results showed that the particle size of the carbon had a significant effect on the rate of gold adsorption. No displacement of adsorbed silver by gold was observed to take place at an influent pH of 5. In a typical continuous 'merry-go-round' operation involving three columns in which gold was eluted from the lead column every 48 hours, effluent values averaging 0,001 g/t gold could be consistently achieved under simulated plant-operating conditions. The results were obtained at relatively high velocities of influent solution, resulting in very little hold-up of gold in the carbon-recovery circuit. More than 99,9 per cent of the gold and more than 99,8 per cent of the silver was recovered in the adsorption circuit, while the elution data indicated an average elution recovery of 99,8 per cent of the adsorbed gold and 98,2 per cent of the adsorbed silver. Elution of the carbon was carried out at 91 QC. This involved pretreatment of the carbon with a sodium cyanide solution, followed by elution with deionized water. At a loading flowrate of 66 bed volumes per hour, the indicated take-up of metals on the carbon in the lead column was as follows; 53 k g/t gold, 6 k g/t silver, 57 k g/t copper, 14 k g/t nickel, and I k g/t zinc. Only a trace of cobalt was adsorbed. At an elution rate of 1,7 bed volumes per hour, the lead column was found to be completely stripped in under 5 hours, with concentrations of gold in the eluate reaching 6 to 7 g/l. Besides the excellent adsorption and elution results obtained when acidified pregnant solution was used, no fouling of the carbon was indicated, and the pressure drop across the three columns was considered to be well within practical limits. When compared with conventional zinc precipitation, the carbon process would appear to have considerable economic merit, with strong indications that significant savings would be possible both in capital and operating costs. Further pilot-plant experiments, preferably on a gold plant, would be required before a meaningful cost comparison could be made with the existing process.

Jan 1, 1979

By A. L. Jones

The subject assigned to me of "The Power" Problem of the Mine" is so broad that I have decided to confine myself to the question of the power supply itself, rather than the application of electricity throughout the property. We will look over certain economic aspects of the problem, particularly in the way it affects the mine manager or executive and his attitude toward the power plant. The accepted method of power application to the mine is by electric power, and the experience of each of you is that the job is never complete as you are always undertaking some further electrification which necessitates the stiffing up the power house to care for the additional load. The mine executive, then, is wrestling almost continuously with the problem of his power plant. Either it is overloaded and must be enlarged and rebuilt; or its operating costs are excessive; or it breaks down and limits coal output; or some one wants to sell new and improved machinery, until the manager almost hates the thought of such a restless part of his property. If he purchases power, every day new uses for it seems to be found. In many ways this unrest is due to short-sighted policy toward the problem, and the tendency when the problem of extension has to be met, to dispose of it in the cheapest way and without thought as to where the future is leading. The day when each opening had its own boiler plant, steam hoist, steam driven fans, steam pumps, steam loaders, etc., seems to be far behind us. The recognized method of power distribution today is by electric power, except for rather limited applications of compressed air; most of which is electrically compressed. During fifteen years contact with mines of Colorado and Wyoming the writer has seen nearly every mine power house double or triple in capacity or abandoned in favor of purchased power, worn. out and inefficient machinery thrown out, isolated boiler plants dropped, seam pumping at great depth changed to electric drive, all resulting in greatly reduced labor and fuel costs. Power costs, which in such isolated plants were considered normal, would be unthinkable today. One property where the main power supply was by compressed air from steam driven com¬pressors used under boilers 8.35 per cent of all the coal mined. The cost of power was 12 1/2 cents per ton of coal mined. Another property of four openings, each with its own boiler plant and usual equipment of hoists, pumps, etc., had a power cost of 10.3 cents for each ton mined; and these were days when you did not get very high prices for coal. Another property operating two mines mostly by steam used 11.1 per cent of the coal output under their own boilers and the power cost was 19.45 cents for each ton mined. Each of these properties is now electrified and power costs are from 3 to 4 cents per ton, and the fuel burned runs from 1 to 2 per cent of the output. With the broadening market and high prices for slack coal and the high prices for labor in boiler and engine rooms, the cost of steam is reaching a high figure, and efforts must be made to keep this down, by more efficient steam production and by utilizing this steam in better machinery which will increase the work done by each pound of steam. It must not be overlooked either, that the plant which is very wasteful of fuel is usually a pretty old one and one far from reliable. In a general way the more efficient the plant, the greater its reliability. Ample and reliable power supply means ability to get out coal when there is a market. The increasing use of purchased power is largely due to changing conditions which are tending constantly to push up the cost of power, unless the most efficient machinery is used and large scale production possible. The

Jan 1, 1923

By T. R. A. Gokcen

THE possibilities of separating and purifying metals by high vacuum distillation were examined by Kroll.1 He suggested vacuum treatment for the removal of zinc from the lead produced after Parkes desilverizing. The St. Joseph Lead Co. developed the first commercial vacuum dezincing process at their Herculaneum refinery, as described by Isbell.² The Broken Hill Associated Smelters at Port Pirie, Australia, has, after several years of pilot plant operations and fundamental investigations, developed a continuous process, which will be described briefly in a forthcoming publication." As the full-scale continuous vacuum dezincing plant at Port Pirie is still experimental, publication of full practical details of the plant will be deferred until the unit is operating as a normal part of the continuous refinery. This paper deals only with theoretical aspects of vacuum distillation processes, with particular reference to vacuum dezincing. The method of mathematical analysis is of general interest as it may be applicable to other metallurgical separations which have been investigated recently.4-6 Evaporation Processes At about atmospheric pressure, or higher, most liquids possess a boiling point—a temperature at which any heat put into the liquid is absorbed only as latent heat, not as specific heat. If a steady heat input is supplied, the liquid's temperature rises to this value, then remains constant while bubbles of vapor form beneath the surface. The rate of evaporation is determined solely by the rate of heat transfer to the liquid; the temperature of boiling is determined by the partial pressures of the volatile constituents in the liquid, and the total pressure above the surface. If the rate of heat transfer to the liquid is increased, the temperature remains constant, and the rate of boiling increases. When evaporating metals under vacuum, however, the partial pressures concerned are generally so small that boiling does not occur, because at even a fraction of a millimeter below the surface the hydrostatic pressure is usually too great to permit the formation of a pocket of vapor. In addition, the high thermal conductivity of metals tends to prevent the local superheating which is necessary for bubble formation.' Although this effect is doubtless also exerted when boiling metals at higher pressures, the magnitude will be less because the degree of superheat required to form bubbles is very much less at the higher temperatures involved. Under vacuum, therefore, evaporation of volatile constituents takes place only from the exposed surface, and the rate of evaporation depends upon the surface area, the surface concentration of volatile constituents, the surface temperature, and the partial pressures of volatile constituents immediately above the surface. If the heat input is raised above a certain level, the effect is not to increase the evaporation rate at constant temperature, but to raise the temperature of the liquid until at some higher level an increased rate of evaporation (and thus of latent heat absorption) again balances the heat input rate.? Many substances (including metals) have a very large intrinsic evaporation rate at quite low temperatures—far below their normal boiling points. However, at atmospheric pressure, the large numbers of atoms evaporated are almost completely deflected back into the liquid (or solid) surface by air molecules. Thus the back condensation rate is practically equal to the gross evaporation rate, and the net evaporation rate is practically zero. It can therefore be seen that an overall distillation rate depends not only upon the intrinsic evaporation rate, but also upon the ability of the volatile atoms to move away from the evaporating surface. This movement is facilitated by the provision of a condensing surface close by. Vacuum Distillation: The function of the vacuum above the evaporating surface is to remove foreign molecules, so that the chances of deflection of an evaporated atom back into its source are reduced. When the residual gas pressure is reduced so far that the evaporated atoms have a high probability of reaching a nearby condensing surface without suffering collision with a foreign molecule, the state of affairs is termed "molecular distillation." This process is practiced commercially today for the purification of numerous organic chemical products of high unit value, but not, to the writer's knowledge, for any metallurgical separation. When the degree of vacuum produced in a still is not sufficient to promote molecular distillation, then the evaporated molecules must diffuse through the

Jan 1, 1954

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

Introduction Surface mining is a business, and the objective of most businesses is to make as much money as possible within certain responsible constraints. Businesses normally measure their potential to make money by estimating the discounted cash flow rate of return on new projects or the net present value for ongoing projects. These estimators of profitability are calculated using the following well-known equation: [ ] where i is the year, n is the total number of years, A, is the after tax net cash generation in year I, r is the discount rate and NPV is the net present value. In its simplest form, when the value of r makes the NPV equal to zero, the discounted cash flow-rate of return (DCF-ROR) is defined. The maximum economically recoverable reserve is identified at the point where the NPV or DCF-ROR is at a maximum. Depending on the measure used (DCF-ROR or discount rate) for calculating the NPV, the definition of recoverable reserve will be different for exactly the same deposit. Maximization of the measure of profitability and the economically recoverable reserves should not be separated from the mining method and sequence of exploiting the reserve. The surface mining methods may involve multiple-pass development of the resource, mining to the ultimate limit in one pass or a combination of the two methods with backfill. Multiple-pass development When planning a multiple-pass mining operation, it is important to determine the width of pit expansions that are appropriate for a specific situation. Production personnel typically want wide, open work areas where operations can be set up to run at peak efficiency. While this is understandable, the trade-off between highly efficient production operations and an overinvestment in stripping should be carefully weighed. As long as safety is not compromised, a few cents per bank cubic yard increase in operating cost due to lower productivity by working in more confined spaces may well be offset by a decrease in the overall short-term stripping requirements. A mini study on productivity and cost should help quantify this cost before deciding on minimum phasing requirements. The bigger challenge associated with establishing smaller work areas is securing operations and management buy-in to marginally lower productivity when reporting results. When cost reduction is the important management goal and everyone is focusing on unit costs, one sometimes loses sight of the goal of total cost management and, even more importantly, maximum profitability. The first pit should be defined by the first grouping of recoverable mineral that has the highest after-tax unit value and forms a feasible mining unit. The assumed mining of the first pit should be followed by the assumed mining of the second most profitable practical pit or pit expansion. If this process of grouping and sequencing is followed until the profitability approaches zero, the general sequence required to extract the maximum economically recoverable reserve is identified. The curve in Fig. 1 is constructed for a rate of production, R, and a quality management strategy, Q, by calculating the NPV for a series of sequences, with each sequence assuming that the operations are terminated on successively lower profit pits. These values are plotted to form the curve in Fig. 1. If the operation ceases on a pit that has too high of a profit, opportunity is lost, and the NPV is lower than the maximum. The NPV curve peaks and starts to decline before the zero-profit pit is assumed mined. This occurs because the money invested in advanced stripping would have made more money invested at the discount rate than by mining the next pit expansion. Therefore, the maximum economically recoverable reserve is defined by the pit limit corresponding to the maximum NPV. There may be an opportunity to refine the definition of the "ultimate pit" around the maximum-value point. The precise definition of the ultimate pit early in the mine life may or may not be important today, depending on how far off in the future these limits will be reached. By the time an operation must decide on its last increment of economically recoverable reserve. much of the

Jan 1, 2000

By G. P. Larson

We define a specialized mineral market as follows: Specialized markets occur where a low volume of a given mineral is used to convey a large benefit to a specific product. Sales of these value added minerals to small markets generally require a different marketing approach than bulk or commodity markets for traditional minerals. Specialized markets are differentiated by several characteristics as listed in Table 1. [Table 1-Characteristics of Specialized Markets 1. Long-term sample evaluation 2. Small volume markets 3. Highly controlled properties 4. Customized specifications 5. Technical selling effort] Many of the specialized markets require long term evaluations before sales can be made. This can vary greatly depending on the type of industry and ultimate end-use of the final product. In some cases, complete evaluations may take from several weeks to years. Length of sample evaluation has no relationship to size of market. That is, the total market may be quite small or fairly large (e.g. from a few hundred pounds to a few million pounds). Uniformity of value added minerals from lot to lot is an extremely important characteristic for these markets. Variations that are of no consequence to one industry (or application) cannot be tolerated in another. This has led to customized specifications and test procedures for specific markets and applications. Because of the greater uniformity and customized specifications, the need for greater technical selling effort is required. This is particularly true in the initial developmental phases and contacts as well as in maintaining the account after the initial order. A small entrepreneurial organization can be used to advantage to achieve goals of sales and profit to these types of markets. The large commodity or bulk producer's sales and marketing organizations are not ordinarily equipped to give the necessary service or technical assistance to the specialized markets. The organization usually required for this type of marketing is given in Table 2. [Table 2- Specialized Marketing Organization Requirements 1. Flexible approach to technical and marketing problems 2. Versatile background in a broad range of industries 3. Laboratory backup when necessary 4. Marketing expertise 5. Technically competent management] In developing specialized markets, a flexible approach to both technical and marketing problems is necessary in understanding the needs and requirements of the industry and the applications involved. This flexibility is also helpful in adapting existing products to new market areas. Having knowledge of a range of industries as broad as possible can be very helpful. It often happens that a product developed in one area or industry can lead to its application in a different area with excellent results. Laboratory assistance for a wide range of tests and test procedures is an important asset in any development program for specialized markets. Laboratories can also assist in competitive evaluations. Marketing expertise and a management with some degree of technical understanding are necessary in to assign proper cost/ benefit ratios to a proposed new product or application of existing products. This is important since the total market volume can be small and the time required for initial contact to first sale can often be lengthy. Also, management's technical and marketing expertise is required in evaluating all possible options that can be considered viable when an application is presented. Once an organization of this type is in place, it is then possible to determine the needs of the marketplace, including when various minerals might have to be customized to fill those needs. There are generally two methods of marketing, as noted in Table 3. [Table 3- Method of Marketing 1. Product Driven -Develop a product with certain desirable properties and try to find a market to utilize those properties. 2. Market Driven -Know market well enough that when a need arises, it is possible to identify a product that will satisfy the need.] These approaches are not mutually exclusive and can be used at the same time. Marketing a value added mineral generally requires the "market driven" approach. Hence, the importance of a knowledgeable management as outlined above. To determine where needs and opportunities exist, several approaches can be used. Table4 lists some of the methods employed.

Jan 1, 1993

By Jonathan Ludlow, Robert A. Jankowksi

Introduction A longwall operator can make few changes to increase output, significantly reduce respirable dust, and decrease power consumption. Reducing drum speed, and thereby cutting with increased pick penetration, is one. This article defines the benefits of deep cutting in terms of reduced dust production and power consumption. It also identifies the practical aspects of high pick penetration in terms of shearer performance and coal loading. Before examining some practical aspects of reducing drum speed and looking at the theoretical background, it is worthwhile to summarize what is meant by high penetration and deep cutting, and what potential benefits and pitfalls may be expected. Deep cutting (in the sense of high penetration rather than wide web) can be defined in one or more of the following ways: • Cutting with an average pick penetration distance higher than that used in the past. • Cutting with a pick penetration higher than the longwall operator would have used if the advantages of deep and slow cutting were not considered. • Cutting with a well-designed shearer drum below 40 rpm. All these definitions are slightly arbitrary. They are given to provide a basis for discussion and to make the point that any move towards deeper, more efficient cutting can result in operational benefits. The benefits of deep cutting appear in many different areas. The most noticeable benefit, provided suitable instruments are available, is the reduction of airborne respirable dust. During an experiment on a longwall in the Pittsburgh seam, a nearly four to one reduction in dust levels was seen when drum speed was halved. Not all studies have shown such a big reduction, but it seems that some benefit is almost always obtained when drum speed is reduced. Production rate and specific power consumption are also affected (in a positive sense) by reducing drum speed or increasing pick penetration. Although these changes may not be as spectacular as those in dust level, they contribute to the economic return of the longwall operation. Similarly, improved washability through fines reduction may have a beneficial economic effect. Cutting with shearer drums operating at lower speeds does have some possible deleterious impacts that an operator should be aware of. For example, cutting reactions - loads imposed on the picks by the coal being cut - will be increased as a deeper cut is used. Steps must, therefore, be taken to ensure the stability of the shearer and provide an adequate haulage effort. These increased cutting reactions also result in higher loads on the power transmission system (gearboxes, ranging arms, pick boxes, etc.) from the shearer motor(s) to the pick tip. These higher loads must be anticipated and provided for with the necessary hardware. In particular, extra haulage power must be provided with low drum speeds, since haulage effort required increases roughly in proportion with pick penetration. Because the drum will be rotating more slowly or will have fewer picks, the load on shearer components will also be more variable. If suitable, robust equipment is not used, this increased vibration will decrease reliability. Benefits of Deep Cutting Lower dust levels, decreased specific power consumption, and improved product washability are the most noticeable benefits of reduced drum speeds. Although the benefits will vary greatly with mining conditions and the type of coal, some examples of what can be expected are described below. Reduced Dust Levels Figure 1 shows principal results of a study on the effects of reduced drum speed conducted on a longwall in the Pittsburgh seam (Ludlow, 1981). This figure shows that average dust production was reduced by about 70% when drum speed was halved. By making some assumptions about such quantities as coal density, it is possible to apply this proportional reduction to the quantity of respirable dust liberated per ton of coal mined. When this is done, two kinds of results are obtained: • At 70 rpm, about 1 g (15 gr) of airborne respirable dust is created for every ton mined (roughly one part per million). At 35 rpm, only 0.28-0.37 g/t (3.9-5.1 gr per st) of coal mined become airborne respirable dust. • At 35 rpm, nearly four times the amount of coal may be mined before the compliance level is exceeded, compared with 70 rpm.

Jan 3, 1984

By J. J. Gilman

Examination showed that characteristic cleavage step patterns are observed on the cleavage surfaces of undeformed, slipped, bent, twinned, compressed, and indented zinc crystals; and the effect of temperature is discussed. Dimples were seen to produce cleavage steps in a treelike pattern in otherwise undeformed crystals. The steps seem to originate when cracks intersect screw dislocations. IT has been known for a long time that the path of fracture in polycrystals may be discontinuous (see Jaffe, Reed, and Mannl for review). Recently, Kies, Sullivan, and Irwin2 have proposed, and given evidence, that crack propagation is discontinuous within individual crystals as well. Other evidence has been given by Low.' When discontinuous cracks within a crystal join together to make a macrocrack, the lamellae between each set of two cracks are torn somewhere, forming small cliffs. These cliffs appear as lines when the cleavage surface is observed microscopically.4,5 The lines have been called vein, tree, and riverlike markings by various authors, and they have sometimes been mistaken for fissures. The descriptive term cleavage steps is used in this paper. Cleavage steps vary in height over a wide range of values, from molecular dimensionsG to lor. and larger. Kies, Sullivan, and Irwin,2 as well as George,' have shown that the gross cleavage step patterns for plastics, polycrystalline metals, and for mono-crystals are sometimes similar. Thus, they depend mostly on the mechanical variables that prevail during cleavage and are relatively insensitive to the structure of the material. For example, parabolic markings2,7,8 sometimes result when cracks open up ahead of, and not coplanar with, the main crack front. If the advance crack has the same velocity as the main crack, their intersection line is a parabola, otherwise it is a hyperbola or an ellipse. The patterns are strongly affected by differences in crack velocities. This results in chevron patterns which point to the place of origin of the main crack. It is the purpose of this paper to demonstrate the existence of a mechanism of cleavage step formation which is a continuous rather than a discontinuous process. Also, certain characteristic step patterns are described, and the strong effect of temperature is shown. The specimens were zinc monocrystals (grown from 99.999+ pct pure metal). These were cleaved at room temperature and at — 196°C. Results and Discussion Cleavage step patterns are highly variable from point to point on a given specimen, as well as from one specimen to another. Although the patterns shown in the photographs are typical, they have been selected for graphic illustration. Figs. la and lb compare undeformed crystals that were cleaved at —196 °C and room temperature, respectively. Cleavage at room temperature (Fig. lb) resulted in a higher density of high steps (dark black lines) and enhanced the visibility of the fine background markings. Deformation by simple slip caused no marked change in the step patterns until the glide strain reached about 1.0. But, as Fig. lc shows, the density of high cleavage steps was greatly increased by large glide strains. Corrugations lying perpendicular to the slip direction may also be seen in Fig. lc. These are caused by deformation bands. The cleavage resistance of the crystal of Fig. lc was very high compared to undeformed crystals (estimated by the force on a needle required for cleavage). Striking and varied cleavage step patterns were observed on bent crystals. Two characteristic patterns that were observed on crystals bent at 25°C, and cleaved by reverse bending at —196°C, are shown in Figs. 2a and 2b. The first, Fig. 2a, consists of V-shaped lines similar to the parabolas of other materials2,7 Fig. 2b shows a pattern that is the equivalent of Fig. la, consisting of faint background lines with a few higher step markings. Cleavage of bent crystals at room temperature resulted in Figs. 2c and 2d. Now, the cleavage step lines show a strong tendency to follow one of two perpendicular paths. In Fig. 2c (bent once), many of the cleavage step components that lie parallel to the bend axis are assembled into irregular lines. In Fig. 2d (bent twice), the cleavage steps again tend to consist of two perpendicular components, but neither of the components is assembled into lines. Also, the step density is higher.

Jan 1, 1956

By Orville R. Lyons

This paper presents a method whereby the amount of misplaced material and the difficulty of the separation can be used to compare coal cleaning equipment of all types, from effectiveness and capacity standpoints. The correlations presented do not include all types of equipment currently available, but the method can be used to evaluate any make or type of coal cleaning equipment, both old and new. THE relative performance of coal washing equipment, or the effectiveness with which any type or make of equipment removes impurities from coal, has been most difficult to evaluate in the past. The most widely used yardstick is the Frazer and Yancey efficiency formula developed in 1922,' but Yancey in a later article states that "washers treating coals of different density composition or operating at different densities of separation cannot be compared directly on the basis of this criterion."' Prior to and since 1922, a variety of other methods has been used for comparison purposes, including the distribution curve, the error area, and the "ecart probable" or probable error. Yancey and Geer in discussing these methods conclude, "Performance can be evaluated in a number of different ways, with the choice of the proper method to use being dictated by the objectives of the investigation and the data available."' It is true that performance can be evaluated in a variety of ways, but if the equipment is to be evaluated on an effectiveness basis, there should be only one universal comparison method. Varying methods have been used because one universal comparison method has not been found or developed. In the article previously quoted, Yancey and Geer state in clear terms the primary concept for a universal comparison method: "One of the simplest, and certainly one of the most obvious evaluations of washery performance is the quantity of sink material in the washed coal and the float material in the refuse. If the washery products are tested at the density at which the washing unit is operated, the sink in the washed coal and the float in the refuse represent material that has been misplaced." The quantity of misplaced material was used as a criterion of washery performance by Lincoln in 1913," by the United States Bureau of Mines in 1938,' by Hancock in 1947," and by the national French research agency Cerchar in recent years.' In 1950 Andersone proposed the use of this criterion as an efficiency value to replace the Frazer and Yancey formula. However, none of the above-mentioned investigators used the misplaced material concept in a manner that would provide universal coal-cleaning equipment comparisons. The Correlation Theory The ideal coal cleaning process would treat all sizes and would make a perfect separation at any given specific gravity. All material lower in density than the desired value would report in the coal product and all material higher in density would report in the refuse product. Unfortunately, no known cleaning process achieves this goal and there seems little likelihood that any process yet to be invented will do more than approach it. When coal is treated in volume under operating conditions, it is impossible to avoid mechanical entrapment, fluctuations in throughput and effective gravity of separation, and the creation of turbulent currents, even when a true heavy-liquid bath is used and the feed is closely sized and contains little intermediate gravity material. This being so, it is possible to appreciate the difficulties inherent in trying to obtain a perfect separation when treating a wide range of sizes and a feed containing high percentages of intermediate material, using turbulent currents to help create the effective separation gravity, under operating conditions which normally tend to be on the overload side. When coal is separated from refuse in any coal cleaning equipment, some refuse always reports to the coal and some coal to the refuse; the writer therefore assumed that there should be a relationship between the total amount of misplaced material produced by any given piece of equipment and the difficulty of separation as represented by the percentage of near gravity material in the feed. With small amounts of near gravity or k0.1 material in the feed there should be less misplacement of material than would occur with large amounts of near

Jan 1, 1953

By Arnulf Muan, Hobart M. Kraner

Ferromanganese alloys react with aluminu-silica brick in blast furnace hearths and cause the formation of new phases with low refractoriness and consequent failure of the refractory lining. The nature of these reactions is explained on the basis of petrographic observations, thermodynamic data, and Phase equilibrium considerations. Animproued hearth design resulted from these findings. The production of ferromanganese in blast furnaces presents many problems which are not encountered in the production of iron in the same furnaces. One of the most serious of these problems is the rapid failure of refractories caused by their reaction with the ferromanganese metal. The present paper describes such a "hearth attack", its symptoms, causes, and cure. Following an introductory discussion of operating practice in ferromanganese blast furnaces, results are presented of a petrographic examination of samples from a furnace which had to be shut down because of refractory failure. The causes of this failure are then analyzed on the basis of petrographic observations combined with thermodynamic and phase equilibrium data available in the literature. The latter type of data were subsequently used as a guide in recommending improvements in hearth designs which have since proven themselves successful in practical operations. I) FERROMANGANESE BLAST FURNACE PRACTICE It is considered good practice to blow in a furnace on iron and operate it in this way for a period of time preliminary to producing ferrornanganese. During this time, iron generally penetrates the bottom refractories six or more feet below the original working surface of the hearth. Not only does this metal pervade the joints to these depths, but the pores of the refractories are usually also filled by molten iron. The high throughput of the iron furnace provides ample heat to maintain a high temperature in the hearth refractories. This, together with a high ferrostatic pressure, causes porous clay brick of the hearth to shrink. During subsequent use of such a furnace in ferromanganese production, the iron in the pores of hearth brick if; replaced to some extent by ferromanganese. The replacement is usually not complete. This is probably due to the lower metal throughput and lower prevailing temperatures resulting from this in the brick of the hearth bottom. When the ferrornanganese furnace is banked—or when the temperature in the furnace falls due to a furnace delay—-manganese oxidizes even though the atmosphere is largely CO. Considerable expansion accompanies this oxidation of manganese from metal to oxide. The hearth refractories are then subjected to the disintegrating forces of the volume expansion. These forces may in extreme cases be large enough to lift the furnace off its mantle supports as was described in a previous paper.' When cooling takes place slowly, however, and the temperature remains at a fairly high level, the refractories undergo fluxing action by the MnO. This latter reaction and the reduction of SiO2 of the clay brick is the general subject of the present paper. The observations to be described herein were made on a ferromanganese furnace located at the Johnstown plant of Bethlehem Steel Co. The furnace was lined entirely with clay refractories and had operated on iron for 266 days and on ferromanganese for 96 days prior to being taken off because of a hearth wall failure. II) PETROGRAPHIC EXAMINATION AND CHEMICAL ANALYSIS The photograph reproduced in Fig. 1 shws a cross section of the upper part of an 18 by 9 by 4-1/2 in. bottom block removed from the blast furnace. This brick came from a depth of approximately 6 ft below the original working face of the bottom. It will be noted that the 4-1/2 in. dimension is almost the same as when the brick entered the furnace. Three characteristically different zones marked I, II, and III on the print of Fig. 1 are apparent. Results of chemical analyses of the original brick and samples from each of these zones are listed in Table I. Petrographically the various zones were characterized as follows: The central zone marked I was essentially un-reacted brick. The microstructure consisted of fine needles of mullite, and the pores in this entire area were filled with iron. This is characteristic of

Jan 1, 1962

By G. R. St. Pierre, S. W. Gilby

The equilibrium vapors established over Fe-Cr-Ni alloys containing up to 50 pct Cr at 1600°C were determined by collecting the effusate from thoria Knud-sen cells. From these data, the activities of iron and nickel in the ternary alloys have been calculated. In a separate set of measurements, a mass spectrometer was used to determine the activities of the three components in a series of Fe-Cr-Ni alloys. In binary Ni-Cr melts, the activity coefficient of chromium referred to pure solid chromium as standard state is less than one at low chromium contents but is greater than one for melts containing more than 30 at. pct Cr. For Ni-Cr melts containing less than 10 at. pct Cr, In ?Cr = 0.15-0.92 N2. Ternary activity diagrams for Fe-Cr-Ni alloys are presented. In iron-rich alloys, the interactions between the three components are negligible; however, in nickel-rich alloys, the three principal interactions have significant positive values. THE activities of components in binary Fe-Ni alloys have been measured both in the solid solution1-3 and liquid solution4-6 temperature ranges. The determination for Fe-Ni melts have been made by three different methods: transpiration,4 Knudsen-cell effusion with analysis of an accumulated condensate,' and Knudsen-cell effusion with direct measurements of relative component intensities by mass spectrometric methods.6 The results obtained by the three different methods in three different laboratories are in excellent agreement and considerable confidence is placed in the activity coefficients for Fe-Ni melts. At 1600°C, the activity coefficients in dilute solution are ?°Ni = 0.66 and ?oFe = 0.40 for the pure liquid components as standard states. Measurements made on solid solution Fe-Cr alloys7-10 indicate that the component activities are slightly greater than ideal. The measured activities in iron-rich Fe-Cr liquid alloy11-16 indicate almost ideal behavior for this system. The recommended activity coefficient of chromium in dilute solution in liquid iron at 1600o C, ?oCr, is 1.18 when referred to pure solid chromium as standard state and 1.04 when referred to pure undercooled liquid chromium as standard state." The conversion factor of 1.14 is based on a melting point of 1857°C and an enthalpy of melting of 4047 cal per g at. for chromium.21,37 The activity coefficient of iron in dilute solution in liquid chromium has not been measured. Activity measurements on solid Ni-Cr alloys have been made by several investigators.17-20 During the review of this manuscript, Fruehan12 published the results of a study on liquid Ni-Cr alloys at 1600°C. He used galvanic cells22 with ZrO2(CaO) as the electrolyte and Cr/Cr2O3 for the reference electrode. His results indicate that ?oCr in nickel at 1600°C is about 1.01 when referred to pure solid chromium and about 0.89 when referred to pure undercooled liquid chromium as standard state. The activity coefficient of nickel in liquid chromium is unknown. The compositions of vapors freely evaporating from some molten Fe-Cr-Ni alloys have been measured by Lyubimov, Granovskaya, and Berenstein.23 Their results, combined with known vapor pressures, indicate strong positive deviation for chromium in iron-rich alloys and strong negative deviation for chromium in nickel-rich alloys. EXPERIMENTAL PROCEDURES This research was undertaken to determine the activities of components in ternary Fe-Cr-Ni melts containing up to 50 at. pct Cr. At the initiation of the research, we decided to use a Knudsen-cell method wherein a portion of the effusing vapor is condensed on a cold plate.' The composition of the equilibrium vapor for each alloy composition is determined in this method by holding the system under constant conditions for a period of time sufficient to collect enough condensate on the cold plate for chemical analysis. The activities of components in ternary alloys can be calculated from such vapor composition determinations by Eq. [1].24

Jan 1, 1970

By Marshall Pease, R. J. Brandon

A UTILITY that has a large consumption of coal must insure an adequate and sound supply of fuel. The Detroit Edison Co., which has an annual coal consumption of about four million tons and spends approximately $32 million a year for coal, including freight charges, has developed a program for fuel procurement and for evaluation and selection of fuel for reliable and efficient plant operation. Fuel Procurement Coal Purchasing Division: The Fuel Supply Div. of the Purchasing Dept. combines all of the procurement functions in one group, which must maintain adequate stocks of fuel. In addition to its usual purchasing duties, the division also governs transportation, follow-up, invoice and freight bill, in cooperation with the Accounting Dept. The Fuel Div. is comprised of the fuel agent, an assistant fuel agent, a coal buyer and five coal clerks, who follow the movement of each car, initially approve freight bills and invoices, and file claims whenever there is a shortage of one ton or more. The fuel agent reports directly to the purchasing agent of the Company, and all major programs are planned jointly with the chief purchasing officer. The apparent individuality of the fuel section is necessary because of the tremendous volume of coal cars handled each day, sometimes as many as 500, which must be handled promptly to complete the fastest possible move from the mines to the plants. Determining Annual Coal Volume: Through the combined studies of a Production Dept. Load Committee and the Controller's office, accurate predictions of coal requirements for any given year are provided the Fuel Dept. from 12 to 15 months in advance. This is divided into the requirements of all individual power plants and central heating plants. This in turn determines the quantity and quality for plants whose specific fuels vary with the type of equipment installed. In the Detroit area, which has a high industrial load, early estimates of annual coal consumption usually are resolved at the end of the year within 4 or 5 pct of the original plan. Operating in a highly developed industrial area eases the task of estimating primary output; and, with the residential demands increasing in a steady and fairly well-defined pattern, the overall coal schedules are not subject to radical changes during any given year. Selecting Suppliers: At any time, but especially during or before an emergency, the coal supply factor must be made secure. This is particularly true in the procurement of utility fuel. There are always extreme quantities of so-called "bargain" coal available during the buyers' markets, such as recently prevailed. These opportunistic offerings may be considered a means of averaging down overall price, rather than as a steady and dependable supply source. Coal is purchased on contract from mine operators and sales agents who have proved reliable. They are not opportunists who desert for higher dollars in times of duress; they do not fail to fulfill contracts when markets rise or overship when markets dip. The progressive operator today who is willing to expend capital to improve quality and service deserves much more consideration than a matching of short-term pennies. When a company has a high volume of annual needs, all phases of the mine supply must be considered. The mine must be able to produce the quality required at a fair price and be able to sell oversizes of coal in enough volume to screen sufficient nut and slack. It must crush coal and sell mine run at prices in line with competitive nut and slack and be willing to do so when there is no demand for prepared sizes. Determining Price: The public utility is in direct competition with any of its customers who can, if savings are guaranteed, generate their own power. Today, to a greater extent than ever before, private industry must do a better overall job than Government-controlled operations. The price of coal must be realistically balanced with the price paid by almost all industries and the railroads as well. The supply-demand ratio in coal is not difficult to discern. There is access at all times to Govern-

Jan 1, 1952

By James R. Hindman

Vermiculite is the mineralogical name generally applied to a group of hydrated, trioctahedral ferromagnesian aluminum platty or micaceous silicate minerals. In microscopic particles the mineral as mineral is differentiated from montmorillonite by a higher Si:AI ratio of approximately 3:l and a higher layer charge and higher cation exchange capacities. Commercial vermiculite, the vermiculite which is mined, beneficiated, and discussed in this chapter, is an important industrial mineral which is formed by the alteration of macroscopic particles of biotite and iron-bearing phlogopite. In natural form it has the size and shape of mica, while possessing the ion exchangeable properties of zeolites and some clays. Within the internal structure of vermiculite lie water molecules, which when rapidly heated to high temperature, transform into steam causing the vermiculite particles to increase in volume. This process of thermal exfoliation produces a lightweight product which finds use in various construction products, agriculture, horticulture, and other industrial applications. In exfoliated form it has the low density and biological inertness of expanded perlite, while maintaining a chemically active surface. Most of the uses of vermiculite are for the thermally exfoliated material. In addition, the cation exchange potential of vermiculite and its layer-silicate structure are properties which can be used as the basis for numerous products including intumescent coatings and gaskets, inorganic films, and the treatment of nuclear and other toxic waste. When chemically and physically modified, it possesses some properties common to organoclays and synthetic compounds. In high purity concentrates and manufactured products, these properties can be altered and optimized in applications to produce superior performance. GEOLOGY Mineralogy Vermiculite was first described as a mineral by Thomas H. Webb in 1824. Webb believed vermiculite to be a variety of talc and described its diagnostic property of thermal exfoliation thusly: "If subjected to the flame of a blowpipe, or that of a common lamp, it expands and shoots out into a variety of fanciful forms, resembling most generally small worms having the vermicular motion exact." Observing that with one exception, all commercial vermiculite has formed by the alteration of iron-bearing phlogopite, a simple definition of vermiculite might be the alteration product of trioctahedral micas, macroscopic particles of which display marked exfoliation when heated rapidly to temperatures above 374°C (the critical temperature of water). The first vermiculite described was found near Worcester, MA, but it was later pointed out by Brush (1866) that vermiculite had been known for some time in Japan as a children's novelty. Because the techniques for mineral analysis were somewhat primitive com- pared to those routinely available today, and because the structure and crystal chemistry of the micas and related minerals were not widely understood, the early literature generated at least 20 "new" minerals which were nothing more than variations in the color and chemistry displayed by vermiculite. Trade names such as Dugarlite, Mandolitem, Strong-LiteTM, and Zonolitem are sometimes used to denote commercial vermiculite, thermally exfoliated vermiculite and their products. Vermiculite particles resemble mica and vary in color from colorless through tan and green to black, depending on the chemical composition. Most physical and optical properties are close to those of biotite, with perfect basal cleavage, hardness between 1.5 and 2.5 Mohs, and specific gravity between 2.2 to 2.8. Fine-sized particles will feel talcose, especially when wet. However, for commercial vermiculite the diagnostic test is that of thermal exfoliation. Vermiculite can display a wide range of compositions, depending on the composition of the original mica and the change of chemistry during weathering and ion exchange. Chemical data for some commercial vermiculites are given in [Table 1]. There are numerous analyses for vermiculite given in the literature, and the analyses of vermiculites from commercial deposits will fall within the range indicated by the average analysis in the table. Many analyses separate the adsorbed water (H2OP) from crystalline water and water as hydroxyl (H20+). The basis of separation of adsorbed water is generally considered to be 100°C. However, hydrate water within the vermiculite interlayer begins to be lost at 75°C and lower. Therefore, measured water content and loss on ignition are combined in the table as H20[total].

Jan 1, 1994

By Haydn H. Murray, Fred G. Heivilin

The "Hormite Group" was proposed for palygorskite (attapulgite) and sepiolite for their complex magnesium silicate composition and elongate crystals (Martin-Vivaldi and Robertson, 1971). These minerals occur in close association with each other and more complex structural variations may exist (Bailey, 1972). In 1862 Savchenkov used the name palygorskite to describe a mineral from the Palygorsk locality (Hay, 1975), near the Ural Mountains. Ovecharenko and Kukovsky (1984) mention that when mountain leather deposits were prospected in the Palygorsk Division mine it was assumed this unusual mineral was a variety of asbestos. Early mineralogists used the terms "mountain cork" or "mountain leather" when referring to palygorskite. Robertson (1986) mentions that it appears palygorskite was known since Theophrastus&apos; time, ca. 314 BC. J. de Lapparent used "attapulgite" for clays from Attapulgus, GA, and Mormoiron, France, because he thought them different from palygorskite, but the two types were proved to be the same (Bailey et al., 1971). The name attapulgite is still used for the Florida and Georgia deposits when the crystal length to diameter ratio does not exceed 10:1(Merkl, 1989). Georgia palygorskite clays are of much shorter length compared to classic palygorskite. In 1847 Glocker first used the name sepiolite which was called "Meerschaum" by Werner (1788) and Hauy (1801) namedit "Ecume de Mer." Brochant (1802) described low density and white magnesium silicates adding the name Talcum Plasticum and Ecume de Mer. In the Meigs-Attapulgus-Quincy district palygorskite (attapulgite) commonly occurs in two distinct forms referred to as short length palygorskite (Meigs Member) and long length palygorskite (Dogtown Member) (Merkl, 1989). Long length palygorskite crystals (> 10 pm) are rarely observed in the Meigs and Dogtown Members, but when present are in association with dolomite crystals. The short length form is usually less than 2 pm in length and has a low magnesium content whereas the long length form has a high magnesium content and a length greater than 2 pm. The distinctions in morphology are not only important because of the relationship to the origin of the deposits, but also in relation to activity in causing membranolytic activity related to data on palygorskite samples from 9 locations ranging from relatively inert to active in work reported by Nolan et al. (1989). The > 10 pm lengths amounted to only 51 of 17,401 fibers sized. The shortest lengths (< 0.5 pm) were relatively inert. This study pointed out that surface activity, morphology, and chemical differences may be distinctly different within the definition of palygorskite, or for that matter for any individual mineral so that health and other properties must be measured because the name alone does not necessarily indicate uniformity. Palygorskite (attapulgite) fuller&apos;s earth was first sold for drilling mud in 1941. The market for this use expanded slowly and has maintained a level of 7 to 10% of the total US production during the last few years. Most of the fuller&apos;s earth sold for drilling mud comes from the southern part of the Meigs-Attapulgus-Quincy district of Georgia and Florida. Palygorskite clays produced in this area are superior to most other fuller&apos;s earth for mud used in drilling salt formations, but because of high water loss, they are inferior to bentonite where the rocks drilled contain no saltwater. According to Oulton (1965), more than 90 different grades of fuller&apos;s earth are produced. Some of these grades are used for pharmaceuticals designed to absorb toxins, bacteria, and alkaloids; for treatment of dysentery; for purifying water and dry cleaning fluids, dry cleaning powders and granules; for the manufacture of NCR (no carbon required) multiple copy paper; for the manufacture of wallpaper; and as extenders or fillers for plastic, paint, and putty. Fuller&apos;s earth mined near Ellenton, FL, was used for making lightweight aggregates for the construction of concrete barges during World War I1 (Calver, 1957). Still other uses of fuller&apos;s earth and its suitability for uses in new products are outlined by Haden, Jr., and Schwint (1967), Haden, Jr., (1972), and Haas (1970). One special use of fuller&apos;s earth is as a carrier of platinum catalysts that are made in the United Kingdom from sepiolite clays mined in Spain. Other uses of sepiolite fuller&apos;s earth (Chambers, 1959) are similar to those of the palygorskite (attapulgite) type mined in the United States.

Jan 1, 1994

By W. O. Philbrook, R. S. Kaplan

Rates of oxidation were studied for levitation-melted samples of Fe-C, Fe-Si, Fe-C-Si, Fe-C-P, Fe-C-Mn, and Fe-C-S alloys oxidized by mixtures of either 5 or 10 pct O2, in helium at several temperatures in the range 1850o to 2050°C. The concentrations of the oxidized elements decreased linearly with time. Mass transport in the gas phase was found to be the rate-controlling step. Ilowever, a mulliple -film model was necessary to account for the flux of iron and manganese vapor counter to the oxygen flow. THE advent of the basic oxygen and spray-refining processes of steelmaking has generated considerable interest in the kinetics of reactions between oxygen-bearing gases and liquid iron alloys. The object of this research was the determination of the role of mass transport in the gas and metal phases and of surface chemical reactions in the overall oxidation process. Of the reactions for elimination of elements dissolved in liquid iron, the oxidation of carbon has been investigated most extensively. Studies by Fuji and Ura,1 Fuji,&apos; and Gunji et a1.,3 done with pilot-scale basic oxygen furnaces, indicated that mass transport of gaseous oxygen was rate controlling when the metal carbon content was greater than 0.15 to 0.30 pct. When the carbon content was less than this critical amount, diffusion of carbon dissolved in the iron was thought to be rate limiting. L. A. Baker et a1.4,5 levitated samples of iron-carbon alloys and oxidized them in CO-CO2, CO-C02-He, and O2-He atmospheres at 1660°C. These investigators found that the metal carbon concentration decreased linearly with time and their samples exploded when the carbon content fell below about 0.8 pct. They concluded that the rate-limiting step was the transfer of oxidant to the gas-metal interface. A model based on this rate-limiting step predicted decarburization rates within a factor of two of those experimentally observed. The explosion of the samples was attributed to a shift in rate-controlling step from gas-phase mass transport to transport of carbon dissolved in the metal. L. A. Baker and R. G. ward6 and R. Baker7 used freely falling droplets to investigate oxidation kinetics. In both papers, the authors speculated that flame fronts may have been partially responsible for the lack of better agreement between experiment and theory. Swisher and Turkdogan8 oxidized liquid iron having an initial carbon content of 0.97 pct with CO-CO2 and CO-C02-Ar mixtures at 1580°C. They concluded that a surface reaction was the rate controlling step. Iron-silicon alloys have been oxidized by R. Baker7 and Filipov and Martynov.9 R. Baker 7 also studied iron-carbon-silicon alloys falling as liquid droplets through oxygen. He found that the rate of both carbon and silicon oxidation varied inversely with the drop radius. When the initial silicon concentration was greater than 6 pct, the samples cooled during their fall. Manganese oxidation has been studied by R. Baker,7 and by Filipov and Martynov.9 ward10 investigated the loss of manganese from liquid iron under vacuum. He found that mass transport of manganese vapor was rate limiting in argon atmospheres of 0.02 torr to 700 torr pressure. Iron-sulfur alloys were desulfurized with H2-He and H2-Ar mixtures by Peters et al.11 A model based on the transport of H2S from the gas-metal interface into the bulk gas gave fair agreement between the predictions of the model and the experimental results. EXPERIMENTAL TECHNIQUE Essentially, the technique consisted of using helium-oxygen gas mixtures to oxidize levitated samples of liquid binary and ternary iron-base alloys. The fused silica reaction tube is shown in Fig. 1. The coil design is similar to that of L. A. Baker et al.4&apos;5 A Tef-lon-sheathed, ferritic stainless steel lifter (B) and an alumina pedestal (C) were used to support the sample in the coil prior to levitation and for the recovery of the sample. Samples were inserted into

Jan 1, 1970

By John Brownson

Ge-Si N-N heterodiodes hare been built recently which show promise as high-speed logic devices. Low-resistivity germanium is deposited on silicon substrates held at temperatures above the germanium melting point and an alloyed heterojunction formed. The diodes are fabricated from this material using conventional mesa techniques. Over-all device qualify has been greatly improved by the use of epitaxial silicon substrates, that is, a subsirale consisting of a film of thin, high -resistivity silicon on a thicker Piece of low-resistirily silicon crystal. A semimpirical device design equation has been devised which predicts switching performance fairly well within the range of resistivitics and geometries employed. Switching times 01 0.9 ns and PIV &apos;s of —20 v are typical of the better 10-ma diodes. Switching times as low as 0.5 ns have been observed for 10-ma diodes and as low as 2.8 ns for 200-ma diodes. With further development it should be possible to improve switching speed by a factor of four. IN our laboratory, Ge-Si heterodiodes have been built recently which show promise as high-speed logic devices. The diodes have been built using a process reported in some detail in an earlier publication.&apos; The process in short involves alloying a thin film of low-resistivity germanium into a silicon substrate. the germanium having been transported and deposited by the well-known Theuerer halide reduction process.2 This process contrasts with that reported earlier by Oldham who used an essentially conventional direct epitaxial deposition technique.3 Oldham&apos;s process, unlike ours, requires cleavage of the silicon sample inside the reactor furnace the instant prior to deposition. Our process has the advantage of yielding large-area macro-scopically plane heterojunctions. Heterodiodes in general have several interesting properties. Without attempting to summarize the theory of heterojunctions.4 two of these properties which have direct relevance should be mentioned. First, N-.V and P-P heterojunction diodes rectify alternating current. Provided that either the band gap or the electron affinity of the two materials is different (which in general is the case), electronic barriers will exist in the junction band structure. When the device is sufficiently forward-biased, the barriers shift so as to permit current flow, and when it is reverse-biased. the barriers block current flow. Second, for N-N and P-P heterodiodes. the foward-conduction process involves majority carriers only. Consequently. when such a device is turned off. there is no minority-carrier storage and hence the diodes switch on and off quite rapidly. As we have reported earlier. N-P and N-N Ge-Si heterodiodes have been built in this laboratory. As one would predict from theory, the N-P heterodiodes were rather slow in turning off while the N-N&apos;s were quite fast. Similarly, the turn-off time of the N-P&apos;s increased rapidly with forward current while that of the N-N&apos;s was independent of forward current. Both N-N&apos;s and N-P&apos;s were moderately photosensitive. Photocurrents resulting from normal room illumination were as high as 1 µa at 1 v for small-area devices. The most disappointing parameter of the early N-N devices was reverse leakage.&apos; The reverse characteristic was invariably "soft" and leakage at only several volts reverse bias was intolerable. The reverse characteristic was greatly improved by using higher-resistivity silicon substrates. Leakage becomes large at voltages about one third of the avalanche voltage one would expect of a P-N homo-junction formed on silicon of the same resistivity. Thus leakage could be reduced to reasonable limits by using unconventionally high-resistivity silicon. Medium-power diodes formed on 30 ohm-cm substrates leaked typically 15 to 50 µa at 100 v reverse bias. The use of high-resistivity silicon substrate material, however, created a new problem. In order to have an acceptable forward-conductance characteristic. the area of the diodes had to be increased. With this increase in area, the capacitance increased. Switching time for N-N heterodiodes is essentially a linear function of capacitance. Thus when the resistivity was raised. so was the switching time. A SEMIEMPIRICAL HETERODIODE DESIGN EQUATION As a result of our experiments with a limited range of diode geometries and resistivities. a semi-

Jan 1, 1965

By Orville R. Lyons

This paper presents a method whereby the amount of misplaced material and the difficulty of the separation can be used to compare coal cleaning equipment of all types, from effectiveness and capacity standpoints. The correlations presented do not include all types of equipment currently available, but the method can be used to evaluate any make or type of coal cleaning equipment, both old and new. THE relative performance of coal washing equipment, or the effectiveness with which any type or make of equipment removes impurities from coal, has been most difficult to evaluate in the past. The most widely used yardstick is the Frazer and Yancey efficiency formula developed in 1922,&apos; but Yancey in a later article states that "washers treating coals of different density composition or operating at different densities of separation cannot be compared directly on the basis of this criterion."&apos; Prior to and since 1922, a variety of other methods has been used for comparison purposes, including the distribution curve, the error area, and the "ecart probable" or probable error. Yancey and Geer in discussing these methods conclude, "Performance can be evaluated in a number of different ways, with the choice of the proper method to use being dictated by the objectives of the investigation and the data available."&apos; It is true that performance can be evaluated in a variety of ways, but if the equipment is to be evaluated on an effectiveness basis, there should be only one universal comparison method. Varying methods have been used because one universal comparison method has not been found or developed. In the article previously quoted, Yancey and Geer state in clear terms the primary concept for a universal comparison method: "One of the simplest, and certainly one of the most obvious evaluations of washery performance is the quantity of sink material in the washed coal and the float material in the refuse. If the washery products are tested at the density at which the washing unit is operated, the sink in the washed coal and the float in the refuse represent material that has been misplaced." The quantity of misplaced material was used as a criterion of washery performance by Lincoln in 1913," by the United States Bureau of Mines in 1938,&apos; by Hancock in 1947," and by the national French research agency Cerchar in recent years.&apos; In 1950 Andersone proposed the use of this criterion as an efficiency value to replace the Frazer and Yancey formula. However, none of the above-mentioned investigators used the misplaced material concept in a manner that would provide universal coal-cleaning equipment comparisons. The Correlation Theory The ideal coal cleaning process would treat all sizes and would make a perfect separation at any given specific gravity. All material lower in density than the desired value would report in the coal product and all material higher in density would report in the refuse product. Unfortunately, no known cleaning process achieves this goal and there seems little likelihood that any process yet to be invented will do more than approach it. When coal is treated in volume under operating conditions, it is impossible to avoid mechanical entrapment, fluctuations in throughput and effective gravity of separation, and the creation of turbulent currents, even when a true heavy-liquid bath is used and the feed is closely sized and contains little intermediate gravity material. This being so, it is possible to appreciate the difficulties inherent in trying to obtain a perfect separation when treating a wide range of sizes and a feed containing high percentages of intermediate material, using turbulent currents to help create the effective separation gravity, under operating conditions which normally tend to be on the overload side. When coal is separated from refuse in any coal cleaning equipment, some refuse always reports to the coal and some coal to the refuse; the writer therefore assumed that there should be a relationship between the total amount of misplaced material produced by any given piece of equipment and the difficulty of separation as represented by the percentage of near gravity material in the feed. With small amounts of near gravity or k0.1 material in the feed there should be less misplacement of material than would occur with large amounts of near

Jan 1, 1953