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By F. W. Schaller

TETRAGONALITY is an important aspect of the models of transformation, strengthening, and tempering of ferrous martensite. When the c/a

Jan 1, 1963

By S. Skowronski

The melting of cathode copper, ususally containing 95.98+ per cent. • Cu, would appear to be a simple matter. Owing to the well known affinity of copper for sulfur, however, so much sulfur is absorbed by the copper during the operation that a long and expensive refining process is required to remove it, and the final product is inferior in purity and lower in electrical conductivity than the original cathode copper. The furnace operation, as now conducted, includes the following stages: 1. Charging the cathodes into the furnace. 2. Melting the cathodes. 3. Oxidizing period—saturating the copper with cuprous sub-oxide and bringing the metal to the condition known as "set copper;'' this is now done by blowing air into the molten bath of metal. 4. Reduction period, otherwise known as "poling," whereby the oxygen previously added is removed by covering the molten bath completely with coke or charcoal and introducing large poles of green wood; this reduces the cuprous sub-oxide, bringing the copper to the "tough pitch" stage ready for casting. 5. Casting Period. Many experiments have been made to obviate the long periods of oxidation and subsequent reduction, with no success. While the copper first cast will be good, as indicated by the set surface of the bars, it is impossible to hold it to the "tough pitch" condition, and long before the whole charge has been poured the copper will become "overpoled," entailing numerous difficulties. Sulfur has always been blamed by the furnace men as the reason why this refining process is necessary, yet the literature on the subject does not sufficiently emphasize how small a proportion of sulfur will ruin copper for practical purposes. The sulfur absorbed by copper during the melting period comes from two sources: (1) the sulfate left on the cathode sheets in the tank houses; (2) the fuel. The percentage of sulfates on the cathodes is variable, and depends on the care taken in washing the cathodes after removing

Jan 1, 1919

By W. W. Odell

Once it was believed that anthracite or coke were the only fuels generally available and suitable for the generation of water gas, particularly so when this gas was made in the generators of standard carburetted water-gas sets such as are commonly used in the manufacture of city gas. During the past 8 years, practice has changed considerably and it is now common to find gas plants operating without anthracite or coke and using bituminous coal in the generator. This change in fuel has appreciably increased the marketability and sales of certain coal producers' products; accordingly it is of interest to coal operators and others to know the reason for this change, the present trend in practice and the adaptability of various coals for generation of water gas. This paper briefly discusses these conditions and some of the factors relating thereto. Bituminous Coal not Ideal but Cheaper than Coke In spite of many changes which have been made in both process and generating apparatus for using bituminous coal as generator fuel more efficiently, in general it is not an ideal fuel for this purpose. It is supplanting coke because it costs less; it is so much cheaper per ton delivered at the gas works that water gas made therefrom costs less per unit-volume than that made from coke, even though the quantity of generator fuel consumed per therm (100,000 B. t. u.) of gas made is greater with coal. This is the reason why so many city gas plants, as well as others, are using bituminous coal instead of coke. Conditions Favoring Change from Coke to Coal The conditions which have made it possible and feasible, as well as economical, to supplant coke with coal may be summarized as follows: (1) Changes in methods of operating and alteration in generating apparatus; (2) lowering of the calorific standards for city gas, and (3) availability

Jan 1, 1927

By Hugh Marriott

MANGANESE and its ores have been recently dealt with in comprehensive papers to the Iron and Steel Institute by Sir Robert Hadfield, and in a series of papers read before the A. I. M. E. at the Cleveland meeting. I have taken the opportunity of adding this mineral, to my records of the "World's Mineral Production and Consumption," as recently pub-lished in the Economist ?. The tables bring out the following significant facts: The world's production has increased by 10 per cent since 1913. India's production has increased by 3 per cent since 1.913, the supply from this source now being 31.1 per cent of the world's output. Russia is now producing at about half its pre-war rate and yields 19 per cent of the world's production. West Africa supplies 12.5 per cent of the world's total output. Brazil's present output is 2.5 times that of the pre-war period, being 11.3 per cent of the world's output.

Jan 9, 1927

By F. J. Haller

IN 1943, we found, in the new Mather operation, a very unusual and disappointing condition in the footwall rock where all of our main haulageways were to be located. With the exception of a few hundred feet out of the thousands then planned and since driven, all of the openings require support. In general, the ground is not particularly heavy, at least by our standards, but it must be supported and the back carefully covered. Also, our operations require rather wide plat openings (16 to 18 ft) and rather flat curves at the turn-outs, all of which adds to the problem. When this completely unforeseen condition was discovered, ordinary timbering was the only answer, because of the impossibility of obtaining adequate steel supplies in time due to war shortages. Then by the time steel supplies were obtained and a practice developed, we found ourselves with between four and five miles of main haulage drifts and plat openings of a permanent nature, all of which were supported by hardwood timber, some of which had already rotted to an unsafe degree. Also, our drifting program was adding some two miles per year to the above figure. F. J. HALLER, Member AIME, is Superintendent, Mather Mine, Cleveland-Cliffs Iron Co., Ishpeming, Mich. AIME Columbus Meeting, September 1949. TP 2840 A. Discussion (2 copies) may be sent to Transactions AIME before May 31, 1950. Manuscript received Aug. 22, 1949. Revision received Jan. 17, 1950. To give all of the details of our experiments, studies, and experiences would fill a large book, but we were satisfied that our organization had done an excellent job, starting from scratch. You can appreciate our surprise when we discovered that we had added very little to an existing record. Attention was called to a footnote in a copy of "Carnegie Pocket Companion" year 1923. The note read: "Full information as to uses of H-Beams is given in pamphlet entitled 'Steel Mine Timbers.' " A search of our Engineering Library brought to light a pamphlet printed in 1911 showing extensive use of steel for this purpose. Except for welding in place of riveting, we have added very little to the designs in that book. We have, however, developed an experience which proves, at least to our own satisfaction, that steel mine supports pay off under the proper conditions. We are now convinced, in spite of the relatively high initial cost of materials, that steel sets, because of comparatively low installation cost, are actually cheaper than either treated or untreated wood sets, if they are properly designed and adapted. Several Sections in Use: Convenience guided our first experiments in steel sets, since for a number of years, we had used the 4-in. WF, 10 lb per ft section in 10-ft lengths as safety forepoles along with wood forepoles. This section replaced the 40-lb rail which had formerly been used for the purpose. Having the 4-in. WF sections on hand, it was natural that our first experiments in steel sets involved their use. Later, the rolling of the 10-lb section was discontinued in favor of the much stronger 13-lb section, which is now standard for both legs and caps in areas of no particular weight, but which, none the less, require permanent support. Our experiments involved the use of the 10- and 13-lb sections as legs and 6- and 8-in. I beams, 16 and 23 lb per ft, respectively, as caps. Experience soon indicated that, if the ground were heavy enough, and the opening wide enough to require a cap stronger than 13-lb, the 8-in, 23-lb should be used, eliminating the 6-in. I beam. We now know that the 8-in. WF section of the same weight, which we propose to use in the future, is considerably stronger for the amount of money involved. Comparative Strengths: Tables I to IV giving strengths of several species of timbers as compared with the structural sections that we now use have been prepared in accordance with accepted practice. Experience indicates that failure due to compression perpendicular to the grain of the wood does not destroy the effectiveness of a wood set frequently enough to be a practical factor. Therefore, no table of these values is included.

Jan 1, 1951

By F. J. Haller

IN 1943, we found, in the new Mather operation, a very unusual and disappointing condition in the footwall rock where all of our main haulageways were to be located. With the exception of a few hundred feet out of the thousands then planned and since driven, all of the openings require support. In general, the ground is not particularly heavy, at least by our standards, but it must be supported and the back carefully covered. Also, our operations require rather wide plat openings (16 to 18 ft) and rather flat curves at the turn-outs, all of which adds to the problem. When this completely unforeseen condition was discovered, ordinary timbering was the only answer, because of the impossibility of obtaining adequate steel supplies in time due to war shortages. Then by the time steel supplies were obtained and a practice developed, we found ourselves with between four and five miles of main haulage drifts and plat openings of a permanent nature, all of which were supported by hardwood timber, some of which had already rotted to an unsafe degree. Also, our drifting program was adding some two miles per year to the above figure. F. J. HALLER, Member AIME, is Superintendent, Mather Mine, Cleveland-Cliffs Iron Co., Ishpeming, Mich. AIME Columbus Meeting, September 1949. TP 2840 A. Discussion (2 copies) may be sent to Transactions AIME before May 31, 1950. Manuscript received Aug. 22, 1949. Revision received Jan. 17, 1950. To give all of the details of our experiments, studies, and experiences would fill a large book, but we were satisfied that our organization had done an excellent job, starting from scratch. You can appreciate our surprise when we discovered that we had added very little to an existing record. Attention was called to a footnote in a copy of "Carnegie Pocket Companion" year 1923. The note read: "Full information as to uses of H-Beams is given in pamphlet entitled 'Steel Mine Timbers.' " A search of our Engineering Library brought to light a pamphlet printed in 1911 showing extensive use of steel for this purpose. Except for welding in place of riveting, we have added very little to the designs in that book. We have, however, developed an experience which proves, at least to our own satisfaction, that steel mine supports pay off under the proper conditions. We are now convinced, in spite of the relatively high initial cost of materials, that steel sets, because of comparatively low installation cost, are actually cheaper than either treated or untreated wood sets, if they are properly designed and adapted. Several Sections in Use: Convenience guided our first experiments in steel sets, since for a number of years, we had used the 4-in. WF, 10 lb per ft section in 10-ft lengths as safety forepoles along with wood forepoles. This section replaced the 40-lb rail which had formerly been used for the purpose. Having the 4-in. WF sections on hand, it was natural that our first experiments in steel sets involved their use. Later, the rolling of the 10-lb section was discontinued in favor of the much stronger 13-lb section, which is now standard for both legs and caps in areas of no particular weight, but which, none the less, require permanent support. Our experiments involved the use of the 10- and 13-lb sections as legs and 6- and 8-in. I beams, 16 and 23 lb per ft, respectively, as caps. Experience soon indicated that, if the ground were heavy enough, and the opening wide enough to require a cap stronger than 13-lb, the 8-in, 23-lb should be used, eliminating the 6-in. I beam. We now know that the 8-in. WF section of the same weight, which we propose to use in the future, is considerably stronger for the amount of money involved. Comparative Strengths: Tables I to IV giving strengths of several species of timbers as compared with the structural sections that we now use have been prepared in accordance with accepted practice. Experience indicates that failure due to compression perpendicular to the grain of the wood does not destroy the effectiveness of a wood set frequently enough to be a practical factor. Therefore, no table of these values is included.

Jan 1, 1951

By Peter H. Fletcher

THE ECONOMIC THEORY It will assist in the understanding of the Resource Rent Tax ("RRT") proposals in Australia if the economic theory behind the tax is briefly explained. The idea goes back to the concept of rent put forward by the nineteenth century economist David Ricardo. His analysis was broadly as follows: A tenant who farms land which is just on the margin of economic viability cannot afford to pay his landlord any rent, because the value of the crop produced by such marginal land is equal to the cost of producing the crop, plus the cost of keeping the tenant alive. There is nothing left out of which to pay rent to landlord. A tenant who farms good land can afford to pay his landlord rent -- but the maximum amount of rent that he can be asked to pay is the excess of the value of the crop produced by the good land over the cost of producing the crop and of keeping the tenant alive. Adapting this theory to the taxation of mining projects, modern economists who favour this system of taxation say that a profitable mining (or resource) project produces rent in the Ricardian sense. The amount of the rent is the excess of the profits by the project over the amount of profit which is required to attract the taxpayer to invest in the project in the first place. This explains why the tax is called "Resource Rent Tax". PAPUA NEW GUINEA PRECEDENT Although Papua New Guinea ("PNG") is now independent of Australia, there are still close links between the two countries, and strong similarities between their respective tax regimes. PNG introduced an Additional Profits Tax ("APT") in 1978. In many respects, Australia's RRT proposals follow the precedent set by PNG1s APT. However, in practical terms, not a great deal can be learned at this stage from PNG's experiences since, according to the writer's understanding, no PNG mining project has yet paid any APT. The Bougainville mine is not liable to APT, being subject to tax under its own separate Act, which contains a different formula. Outline of PNG Additional Profits Tax. PNG has two versions of Additional Profits Tax ("APT) - one applying to general mining, the other to petroleum mining. The two versions are very similar. The general mining version is described below. The tax applies on a licence by licence basis. This means that losses on Licence A cannot be offset against profits on Licence B. Expenditure relating to two or more licences, is apportioned "in such manner as the Chief Controller thinks reasonable". Tax is levied on the "accumulated value of net cash receipts" at the rate of (70 - n)% where n=the rate of income tax. The accumulated value of net cash receipts ("av") is calculated as follows: Calculations commence from the first year in which substantial expenditure was incurred on the construction of facilities for the carrying out of mining operations on the licence. For that year the av consists of the total allowable expenditure to date on that lease. For the second year, the av is the av at the end of year one indexed in accordance with the statutory formula set out below, plus the total allowable expenditure in year 2. For year 3, the av is the av at the end of year 2 adjusted by the index factors, plus the total allowable expenditure in year 3. This process is repeated year after year. When the licence starts to earn revenue, the revenue is deducted from total allowable expenditure for that year. If revenue exceeds total allowable

Jan 1, 1985

By A. A. Scheil, S. I. Hoyt

In technological studies on steel considerable emphasis has been placed on the identification of the foreign inclusions, testimony of which is adequately given in the metallographic literature covering methods for their identification. The outstanding accomplishment in this field is the scheme devised by Campbell, Comstock and Wohrman for identifying various oxides, sulfides, silicates and other constituents by systematic etching tests and microscopic examination.' The time involved in carrying out this procedure has led to attempts to secure quicker methods, or at least to shorten the time required to secure essential information. An example of this is the use of dark field illuminatin,2 but when this was tried out by the authors it was found that cavities and pits on the polished surface appeared bright on a dark field and could not be distinguished from small glassy inclusions. Still another means is the Leitz Ultropak, an instrument equipped with optical units for the illumination of the specimen by reflected light to secure an effect similar to dark field illumination. The structures are made visible in exceedingly sharp details by utilizing the inherent properties of the samples, such as refractive index, absorption coefficients and reflection powers. We have had no experience with the Ultropak, but a recent paper by Benedicks and Löfquist3 describes methods for determining the refractive index of (Fe, Mn) O and MnS by the use of this equipment. Presumably this method would be subject to about the same limitations as dark field illumination, though the, feature added by Benedicks and Löfquist appears to offer great promise in inclusion identification.

Jan 1, 1935

By A. A. Scheil, S. I. Hoyt

In technological studies on steel considerable emphasis has been placed on the identification of the foreign inclusions, testimony of which is adequately given in the metallographic literature covering methods for their identification. The outstanding accomplishment in this field is the scheme devised by Campbell, Comstock and Wohrman for identifying various oxides, sulfides, silicates and other constituents by systematic etching tests and microscopic examination.' The time involved in carrying out this procedure has led to attempts to secure quicker methods, or at least to shorten the time required to secure essential information. An example of this is the use of dark field illuminatin,2 but when this was tried out by the authors it was found that cavities and pits on the polished surface appeared bright on a dark field and could not be distinguished from small glassy inclusions. Still another means is the Leitz Ultropak, an instrument equipped with optical units for the illumination of the specimen by reflected light to secure an effect similar to dark field illumination. The structures are made visible in exceedingly sharp details by utilizing the inherent properties of the samples, such as refractive index, absorption coefficients and reflection powers. We have had no experience with the Ultropak, but a recent paper by Benedicks and Löfquist3 describes methods for determining the refractive index of (Fe, Mn) O and MnS by the use of this equipment. Presumably this method would be subject to about the same limitations as dark field illumination, though the, feature added by Benedicks and Löfquist appears to offer great promise in inclusion identification.

Jan 1, 1935

By Bennett R. Bates

Nowhere in the world has cone classification in closed-circuit grinding with tube mills reached as high a state of development as on the Wit-watersrand. In the development of the Far East section the necessity of fine grinding to obtain satisfactory residues gave rise to a situation with which the cone could cope only at a great sacrifice in efficiency of tube-mill grinding. H. A. White1 recognized the limitation of the cone in his study of the theory of tube milling as early as 1915, but it was not until Ewing, Davis and Willey2 carried out their tests at Springs Mines during 1920 that the advantages of mechanical classifiers were demonstrated. During the fall of 1924 the writer supervised a series of tests with Dorr classifiers undertaken by Mr. White, Consulting Metallurgist, at the mill of the Geduld Proprietary Mines, Ltd., Dersley, Transvaal (Far East Rand). The importance of classification to obtain efficient tube-mill grinding was thoroughly demonstrated. The installation of classifiers of the above type that followed the test confirmed the results of the experimental work. Tests with Cones, Straight Type and Bowl Classifiers The test at the Geduld mill was made to determine the relative grinding efficiency of a tube mill when grinding in closed circuit with cones, straight type Dorr classifier and Dorr bowl classifier. The efficiency of a tube mill, as expressed in this article, is based on the tons of minus 200-mesh material produced by the mill. The efficiency is also expressed in terms of tons of minus 90-mesh material produced by the mills to correspond to the standard practice on the Witwatersrand. The results given for the performance of the tube mills grinding in closed circuit with cones (Fig. 1) were taken from previous tests carried

Jan 1, 1926

By H. F. Yancey, M. R. Geer

The importance of the size composition of coal is reflected in the difference in price of the various sizes of the same coal and in the large number of primary sizes and mixtures of sizes produced by the modern preparation plant to meet the demands of a market that in late years has grown increasingly conscious of coal size. The importance of size in the industrial and domestic uses of coal has its counterpart in the role played by size composition in the laboratory examination of coal. Investigation of many of the physical properties of coal in their relation to its uses requires an evaluation of size composition. In fact, the Standards on Coal and Coke issued by the American Socaiety for Testing Materials in 1936 contains 12 standard methods of testing that involve some use of size composition. The screen analysis is, of course, the standard method of determining size composition, whether the material is coarse run-of-mine coal or fine pulverized fuel. Custorriarily, the analysis is reported as the weight percentages retained on or passing each of a series of testing screens, the number of which may range from only two or three to as many as 20. When more than two or three sieves are used, which is usual, the number of weight percentages obtained is so great as to make the analysis difficult to comprehend. Direct comparison of analyses made 011 the same sieves is hard, and comparison of analyses covering the same range of sizes but made on different sieves is next to impossible. Consequently, a frequent purpose of sizing tests-—-correlation of size with properties and charactcristics related to size-—is rendered difficult. Less cumbersome means of expressing size composition have been attempted with various degrees of success. An average particle size, calculated by any of several formulas, is useful for some purposes, also the weight percentage retained on a single sieve finds some application as a

Jan 1, 1938

By M. R. Geer, H. F. Yancey

The importance of the size composition of coal is reflected in the difference in price of the various sizes of the same coal and in the large number of primary sizes and mixtures of sizes produced by the modern preparation plant to meet the demands of a market that in late years has grown increasingly conscious of coal size. The importance of size in the industrial and domestic uses of coal has its counterpart in the role played by size composition in the laboratory examination of coal. Investigation of many of the physical properties of coal in their relation to its uses requires an evaluation of size composition. In fact, the Standards on Coal and Coke issued by the American Socaiety for Testing Materials in 1936 contains 12 standard methods of testing that involve some use of size composition. The screen analysis is, of course, the standard method of determining size composition, whether the material is coarse run-of-mine coal or fine pulverized fuel. Custorriarily, the analysis is reported as the weight percentages retained on or passing each of a series of testing screens, the number of which may range from only two or three to as many as 20. When more than two or three sieves are used, which is usual, the number of weight percentages obtained is so great as to make the analysis difficult to comprehend. Direct comparison of analyses made 011 the same sieves is hard, and comparison of analyses covering the same range of sizes but made on different sieves is next to impossible. Consequently, a frequent purpose of sizing tests-—-correlation of size with properties and charactcristics related to size-—is rendered difficult. Less cumbersome means of expressing size composition have been attempted with various degrees of success. An average particle size, calculated by any of several formulas, is useful for some purposes, also the weight percentage retained on a single sieve finds some application as a

Jan 1, 1938

By F. R. Hensel

THE alloy under discussion is a nickel-silicon bronze and is one of many age-hardening or precipitation-hardening FIG. I.-EFFECT OF QUENCHING AND AGING TEMPERATURES ON HARDNESS OF TWO TYPICAL NICKEL-SILICON BRONZES, NO. 2351 AND No. 2362. copper-base alloys. The hardening con-stituent is a nickel-silicon compound.1 The quenching temperature necessary to pro-duce hardening by a subsequent aging treatment is a function of the composition and extends from about 700'C. to ()50'C.; the aging temperature range extends from 400° to 600°C. The work carried out by the authors of this paper was confined to cast alloys having compositions within the ranges of 1.5 to 3.00 per cent Ni; 0.4 to 0.75 per cent Si; balance, copper.

Jan 1, 1940

By J. Chipman, G. G. Hatch

T. ROSENQVIST*—It is a pleasure to see the excellent way in which the experimental part of this work has been handled. There seems to be little doubt that the distribution data obtained corresponds most closely to thermodynamic equilibrium under the prevailing reducing conditions, namely equilibrium with graphite and one atmosphere CO pressure. The desulphurization curves in Fig 10 show the same general feature as the curves given by Holbrook and Joseph, but the distribution ratios are from 20 to 40 times greater—undoubtedly due to a closer approach to true equilibrium. In the theoretical discussion, the authors calculate a theoretical distribution (S) ration -jg-. which they find to be about 50 times greater than the experimental. The deviation is so great that the basis for their calculation needs a more thorough examination. The authors base their thermodynamic calculation on free energy expressions where diluted solutions of FeS and CaS are used as standard states. (The activity coefficient in diluted solutions is taken to equal unity.) Such a standard state will change when the nature of the solvent is changed. Taking the free energy of the reaction [FeS] ? (FeS), Eq 2, which is derived from the distribution of sulphur between an iron and a FeO-melt, it is very unlikely that the free energy of this reaction will be the same for a distribution between pig iron and a calcium silicate slag. Therefore a more fundamental basis for the thermodyuamic calculations seems needed, where all thermodynamic equations are referred to unambiguously defined standard states. The most natural standard states for CaO and CaS are the pure solid substances at the same temperature. As standard state for sulphur in iron, pure liquid FeS can be used. This rules out Eq 2 [FeS] ;=s (FeS) because ?F° = 0. The standard equation will then be: FeS, + CaO6 + Cgraph ?Fei + CaS8 + CO. vFo1773 = 25,000 cal It would be more universal and also simpler to refer the escaping tendency of sulphur in liquid iron to the corresponding H2S/H2 ratio which can readily be determined experimentally. As standard state a gas mixture H2S/H2 = 1/1 can be used. (This corresponds at the temperature of liquid iron closely to one atmosphere S2 vapor.) Thus the standard equation for the sulphur reaction can be formulated as follows: H2S0 + CaO3 + Cgraph ?H2o + CaS8 + COg The standard free energy of this reaction has been calculated from the best available data to AF°m3 = —35,000 cal. This gives for the equilibrium constant at 1500°C Now, the solubility of CaS in blast furnace slags has been determined by McCafferey and Oesterle* and corresponds at 1500°C to about 10 pet S (varying somewhat with the composition of the slag.) If the activity of CaS is assumed linear between 0-10 pet as curve 1, (see Fig 11), then acaO = 0.1 (S); (S) being wt. pet sulphur in the slag. For a diluted solution of sulphur in an iron melt saturated with carbon, the ratio H2S/H2 is, according to Kitchener, Bockris and Liberman,f about 0.01 [S], [S] being wt. pet sulphur in iron. Substituting these values in the expression for Kp we find The value 2.103 is only 4 times greater than the experimental coefficient found by Hatch and Chipman, but the value is very sensitive to a small error in AF°. A better agreement with the experimental distribution coefficient can be obtained if one assumes the activity of CaS to run like curve 2 (Fig 11). This (S) will give a lower theoretical W, value, a value which varies with (S) exactly as Hatch and Chipman learned. Such a shape of the activity curve, which corresponds to a positive deviation from Raoult's law, is actually to be expected from the fact that liquid silicate and sulphide phases usually show incomplete miscibility. A closer agreement between experimental and theoretical data can not be expected before we have more complete data for the individual activities of CaS and CaO in the slag. The activities acaS and Ocao referred to the solid phases as standard states, are exact defined quantities contrary to the somewhat undefined expression "free lime," and they are independent of any theory for the constitution of liquid slag. J. CHIPMAN (authors' reply)—The authors wish to thank Mr. Rosenqvist for his very interesting and useful thermodynamic addition. Curve 2 of his figure offers the needed basis for explaining the increase in the ratio (S)/[S] with increasing sulphur content. Attention is called to an error in the printed paper: Fig 2 and 3 are reversed. M. TENENBAUM*—In the figures showing the relationship between excess base and sulphur distribution (Fig 6, 7 and 9) the slope of the curve tapers off in the negative basicity range. Somewhat the same thing is observed with open hearth slags. In that case, the fact that some sulphur distribution between slag and metal is obtained with negative basicity is interpreted as indicating some dissociation of the lime silicate compounds whose existence in oxidizing basic slags has been used to explain various observed phenomena with regard to other slag-metal reactions. In the case of the blast furnace slags, the reduced slope of the sulphur distribution curve with decreasing excess base is attributed to the amphoteric effect of alumina. Has the possibility of other explanations been investigated ?

Jan 1, 1950

By W. N. Poundstone, P. C. Thakur

Introduction Horizontal drilling in coal mines is a relatively new technology. The earliest recorded drilling in the United States was done in 1958 at the Humphrey mine of Consolidation Coal Co. for degasification of coal seams. Spindler and Poundstone experimented with vertical and horizontal holes for several years. They concluded in 1960 that horizontal drilling in advance of underground mining appeared to offer the most promising prospect (for degasification) but effective and extensive application would be dependent upon the ability to drill long holes, possibly 300 to 600 m, with reasonably precise directional control and within practical cost limits (Spindler and Poundstone, 1960). Mining Research Division of Conoco Inc., the parent company of Consolidation Coal Co., began a research program in the early 1970s to achieve the above objective. The technology needed to drill nearly 300 m in advance of working faces was developed by 1975 and experiments on advance degasification with such deep holes began in 1976. Preliminary results of this research have already been published (Thakur and Davis, 1977). To date nearly 4.5 km of horizontal holes have been drilled for advance degasification and earlier results were reconfirmed. In summary, these are: • The greatest impact of these boreholes was felt in the face area where methane concentrations were reduced to nearly 0.3% in course of two to three months from original values of nearly 0.95%. • The methane concentration in the section return reduced to 50% of its original value immediately after the boreholes were completed, indicating a capture ratio of 50%. • The total methane emission in the section (rib and face emission plus the borehole production) did not increase but rather gradually declined with time. • Initial production from 300 m deep boreholes in the Pittsburgh seam varied from 3 m3/min to 6 m3/min but then slowly declined as workings advanced inby of the drill site (well head) exposing a larger surface area parallel to the borehole. Encouraged by these results, it was decided to design a horizontal drilling system that would be mobile and compatible with other face equipment. A mobile horizontal drill can be divided into three subsystems: the drill rig, the drill bit guidance system, and borehole surveying instruments. The drill rig provides the thrust and torque necessary to drill 75- to 100-mm diam holes up to 600 m deep and contains the mud circulation and gas cuttings separation systems. The drill bit guidance system guides the bit up, down, left, or right as desired. Borehole surveying instruments measure the pitch, roll, and azimuth of the borehole assembly. Additionally, it also indicates the thickness of coal between the borehole and the roof or floor of the coal seam. Thus, it becomes a powerful tool for locating the presence of faults, clay veins, sand channels, and the thickness of coal seam in advance of mining. In recent years, many other potential uses of horizontal boreholes have come to light, such as in situ gasification, longwall blasting, improved auger mining, and oil and gas production from shallow deposits. The purpose of this paper is to describe the hardware and procedure for drilling deep horizontal holes. The Drilling Rig [Figures 1 and 2] show the two components of the mobile drilling rig: the drill unit and the auxiliary unit. The equipment (except for the chassis) was designed by Conoco Inc. and fabricated by J. H. Fletcher and Co. of Huntington, WV. The drill unit. It is mounted on a four-wheel drive chassis driven by two Staffa hydraulic motors with chains. The tires are 369 X 457 mm in size and provide a ground clearance of 305 mm. The prime mover is a 30-kw explosion-proof electric motor which is used only for tramming. Once the unit is Crammed to the drill site, electric power is disconnected and hydraulic power from the auxiliary unit is turned on. Four floor jacks are used to level the machine and raise the drill head to the desired level. Two 5-t telescopic hydraulic props, one on each side, anchor the drill unit to the roof. The drill unit houses the feed carriage, the drilling console, 300 m of 3-m-long NQ, drill rods, and the electric cable reel for instruments. The feed carriage is mounted more or less centrally, has a feed of 3.3 m, and can swing laterally by ± 17°. It can also sump forward by 1.2 m. The drill head has a "through" chuck such that drill pipes can be fed from the side or back end. General specifications of the feed carriage are: [ ] The auxiliary unit. The chassis for the auxiliary unit is identical to the drill unit but the prime movers are two 30-kW explosion proof electric motors. It is equipped with a methane detector- activated switch so that power will be cut off at a preset methane concentration in the air. No anchoring props are needed for this unit. The auxiliary unit houses the hydraulic power pack, the water (mud) circulating pump, control boxes for electric motors, a trailing cable spool, and a steel tank which serves for water storage and closed-loop separation of drill cuttings and gas.

Jan 1, 1981

By R. C. Ferguson, Harlowe Hardinge

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

* A loan, perhaps as great as $100 million, from American insurance companies is about negotiated for development of Labrador iron ore properties. The program will call for a total expenditure of up to $200 million, including, as well as all mine development work, the power facilities, housing for mining personnel, and auxiliary equipment. * Mobile air compressors and a water tank, mounted on a 30-ton rubber-tired trailer and towed right up to the drilling crews, eliminates the need for expensive underground water and air pipe lines in the experimental oil shale mine operated by the Bureau of Mines near Rifle, Colo.

Jan 11, 1950

By H. E. Lee, P. C. Feddersen

Greenockite is the only known cadmium mineral of importance. It occurs rather universally, in minor concentrations, as a secondary mineral in sphalerite deposits. The world's cadmium output is obtained through the processing of metallurgical by-products, largely from the treatment of residues from electrolytic zinc, retort zinc and lithopone plants. These sources are supplemented by the processing of fumes from lead and copper smelting operations. The development of modern selective flotation practice in the decade 1920-1930, which permitted the economical mining of complex lead-zinc ores, resulted in significant increases in the quantities of cadmium entering lead smelting systems. Being closely related to zinc as to occurrence, properties and production, most detailed description of cadmium recovery methods are to be found recorded in connection with zinc metallurgy. Other than occasional articles pertaining to particular operational procedures, literature offers but little in the nature of a balanced survey of lead smelter cadmium recovery practices. While many of the basic operations described for the recovery of cadmium from zinc by-products are applicable to the treatment of lead plant products, the inherent problems involved differ widely. In general the cadmium content of related by-products from routine lead smelting operations is present in lower concentrations, exists in a less soluble state and is associated with both a greater quantity and a greater variety of detrimental impurities. To cope with these problems, lead smelter practices are found to follow the general outline: Preparatory Processing 1. Concentration operations 2. Sulphation operations Cadmium Plant Processing 1. Leaching operations 2. Purification operations 3. Sponge precipitation operations 4. Metal recovery operations 5. Refining and casting operations As in the case of related zinc plant operations, cadmium recovery practices at lead smelters are not standardized. They not only vary as to type, but also extent. Depending upon prevailing conditions, lead smelter cadmium operations range from simple concentration campaigns, for the purpose of sufficiently "up-grading" products for shipment elsewhere, to complete processing steps for the production of refined metal. Preparatory Processing The cadmium content of lead smelter receipts is low and, as a rule, proportionate to the zinc content; the usual range of cadmium contained being of the order of 0.01-0.05 pct. Were it not for the low boiling point of cadmium, such small concentrations would, no doubt, be lost in the large tonnages of slag, metal and other smelter end products. However, the ready volatility of cadmium and its compounds at prevailing lead smelting temperatures results in its concentration in fractional portions of fume collected. This collected fume comprises a circulating load within the smelter system. Thus, the cadmium content of blast furnace fume* increases with each successive circulation until an equilibrium value is reached when the sum of the cadmium losses, due to handling and in slag, waste gases and other end products, becomes equal to the intake as ore. With ore receipts averaging, say 0.03 pct cadmium, the concentration value obtainable, through fume circulation in a routine manner, approaches 10-12 pct. In such operations, cadmium concentrations in blast furnace fume of from 3-5 pct are readily .attainable. However, the concentration gain beyond this range, with each additional circulation, is progressively decreased as a result of mounting losses occurring through handling and in end products. Therefore, to avoid excessive cadmium loss and to enhance the ultimate concentration anttainable, it is customary practice to isolate blast furnace fume at some intermediate cadmium content for special concentration procedure. The most common type cadmium concentration "campaign" involves special smelting operations wherein a relatively high portion of blast furnace fume at 3—6 pct cadmium is incorporated into the sinter charge. This type practice is roughly illustrated by the diagram on p. 111. Cadmium fume collected from lead blast furnace operations is not amen-

Jan 1, 1950

By A. M. Starfield

The study of strain waves in rock over the past decade has, for the most part, been an investigation related, but not applied, to rock blasting; the design of rock blasts has proceeded on a basis that is independent of strain wave research. This duality is reflected in the paper by Dr. Clark.1 While the greater part of the paper is devoted to a discussion of spherical waves in elastic and visco-elastic media, it is in an entirely unrelated section, the section on scaling and cratering, that the practical approach to rock blasting is found. Work done by the U. S. Bureau of Mines, as described in the paper by Duvall, Atchison, and Fogelson,2 has established a correlation between crater dimensions and the peak compressive strain in the rock. Strain wave measurements have thus been used to compare explosives but have not, as yet, been applied to the planning of a production round once an explosive has been selected. The optimum burden, for instance, is still determined either by experience, empirical formulae (e.g. that of Langefors 3), or a cratering experiment and the theory that has been developed around that proposed by Livingston.4 Even then there is some ambiguity in applying deductions from short, concentrated explosive charges to the long charges often used in practice (although Bauer 5 reports doing this with considerable success) and wave theory has until recently thrown no light on this problem, being confined, with only a few exceptions, to spherical charges. The hesitation in applying strain wave theory has probably been due to two questions that have not been answered satisfactorily: just how important is the strain wave in rock breakage, and if it is important, under what conditions will the wave break rock? This paper is an attempt to answer the first question, to avoid the second question by an admittedly oversimplified assumption, and then to investigate the implications of this assumption for blasting with long cylindrical charges.

Jan 1, 1967