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In 1970, the federal Clean Air Act was passed, promulgating limits on the emissions of sulfur dioxide (SO2) from large stationary sources that would affect local air quality. The brunt of this regulation was felt by the electric utility industry, especially those units generating electricity with coal-fired boilers. The law hastened development of a new industry, flue gas desulfurization (FGD), which is composed of three major segments: the electric utility, the FGD equipment suppliers (the scrubbing modules), and the scrubbing chemical suppliers. Dravo's experience suggested the equipment end of the business, but market analysis recommended investigating the chemical processes involved in the removal of the SO2 gas. This resulted in the formation of Dravo Lime Co. to manufacture and market reagents to the FGD industry. This paper will briefly describe the Thiosorbic lime scrubbing process and review Dravo's innovative mine and calcining facility.

Jan 1, 1982

By David R. Mitchell

The mineral pyrite (or marcasite) occurs in coal beds as balls, lenses, veinlets and bands. Several million tons are wasted annually on the refuse dumps from coal mining and coal-preparation activities. Sporadic attempts have been made in the United States to recover this pyrite for manufacture of sulphuric acid but only a few plants have succeeded in recovering it on a commercial scale. Several chemical companies in the East became concerned about supplies of pyrite early in 1943, but by late summer of that year the situation had eased. If for no other reason than national security and self-sufficiency during wartime, methods of recovering pyrite from coal-mine refuse should be worked out and made available to coal-mine operators who are potential producers of pyrite. Eastern sulphuric-acid plants are particularly likely to have their sources of raw material blocked by war conditions, so that special studies to determine where supplies of pyrite can be obtained for these plants from inland sources are desirable. Davis's1 summarization of conditions during World War I gives the following facts: During those years, the shortage of pyrite became acute. Shipments of hand-picked pyrite were made from a number of coal mines in Illinois, Indiana, western Kentucky, Ohio and Pennsylvania. A mechanical concentrating plant for sepa- rating pyrite from picking-table refuse was built at Danville, Ill., and operated for several years, and a small pyrite-concen-trating plant treating washery refuse near Gillespie, Ill., operated for several months in 1918. The Consolidation Coal Co. of Fairmont, W. Va., constructed and operated a sulphuric-acid plant using pyrite recovered from coal-mine refuse. The U. S. Bureau of Mines during 1917-1919 made a comprehensive survey of pyrite-production possibilities from coal mines and estimated that there was a potential yearly production of 435.000 tons from Ohio and Pennsylvania and 928,000 tons from Kansas, Missouri, Iowa, Illinois, and Indiana mines. This estimate included only mines that could produce one per cent or more of their coal production as pyrite containing a minimum of 40 per cent sulphur. PyRite Production According to the Minerals Yearbook,2 approximately 660,000 long tons of pyrite was produced in the United States in 1941, with an average sulphur content of 42 per cent. The average value f.o.b. mines was $3.09 per long ton. Spanish pyrite, 48 per cent sulphur minimum, was quoted at 12 cents a unit c.i.f. U. S. ports, equal to $5.76 a ton for shipments of minimum sulphur content. Most of the pyrite produced in the United States is obtained as a by-product in the dressing of copper and zinc ores. Metal mines in the East making pyrite concentrate for the market or use in company-operated sulphuric-acid plants

Jan 1, 1944

By David R. Mitchell

The mineral pyrite (or marcasite) occurs in coal beds as balls, lenses, veinlets and bands. Several million tons are wasted annually on the refuse dumps from coal mining and coal-preparation activities. Sporadic attempts have been made in the United States to recover this pyrite for manufacture of sulphuric acid but only a few plants have succeeded in recovering it on a commercial scale. Several chemical companies in the East became concerned about supplies of pyrite early in 1943, but by late summer of that year the situation had eased. If for no other reason than national security and self-sufficiency during wartime, methods of recovering pyrite from coal-mine refuse should be worked out and made available to coal-mine operators who are potential producers of pyrite. Eastern sulphuric-acid plants are particularly likely to have their sources of raw material blocked by war conditions, so that special studies to determine where supplies of pyrite can be obtained for these plants from inland sources are desirable. Davis's1 summarization of conditions during World War I gives the following facts: During those years, the shortage of pyrite became acute. Shipments of hand-picked pyrite were made from a number of coal mines in Illinois, Indiana, western Kentucky, Ohio and Pennsylvania. A mechanical concentrating plant for sepa- rating pyrite from picking-table refuse was built at Danville, Ill., and operated for several years, and a small pyrite-concen-trating plant treating washery refuse near Gillespie, Ill., operated for several months in 1918. The Consolidation Coal Co. of Fairmont, W. Va., constructed and operated a sulphuric-acid plant using pyrite recovered from coal-mine refuse. The U. S. Bureau of Mines during 1917-1919 made a comprehensive survey of pyrite-production possibilities from coal mines and estimated that there was a potential yearly production of 435.000 tons from Ohio and Pennsylvania and 928,000 tons from Kansas, Missouri, Iowa, Illinois, and Indiana mines. This estimate included only mines that could produce one per cent or more of their coal production as pyrite containing a minimum of 40 per cent sulphur. PyRite Production According to the Minerals Yearbook,2 approximately 660,000 long tons of pyrite was produced in the United States in 1941, with an average sulphur content of 42 per cent. The average value f.o.b. mines was $3.09 per long ton. Spanish pyrite, 48 per cent sulphur minimum, was quoted at 12 cents a unit c.i.f. U. S. ports, equal to $5.76 a ton for shipments of minimum sulphur content. Most of the pyrite produced in the United States is obtained as a by-product in the dressing of copper and zinc ores. Metal mines in the East making pyrite concentrate for the market or use in company-operated sulphuric-acid plants

Jan 1, 1944

By D. H. Yardley

GEOCHEMICAL investigation of trace elements in surface materials was begun near Ely, Minn., in 1953 along the basal contact of Duluth gabbro with Giants Range granite (Fig. 1). This article presents data on the distribution of copper and nickel in till and in stream sediments in the area and proposes an explanation for the types of distribution found. The Duluth gabbro, one of the world's largest basic intrusives, intrudes rocks which range in age from Keewatin to middle Keweenawan. Within the test area the gabbro is in contact with granite except for short sections where it is in contact with remnants of iron formation. Sulfide mineralization occurs within the gabbro, near and parallel to the basal contact for a distance of several miles. Schwartz and Davidson' have described the geologic setting of the mineralization. The sulfides, believed to be syngenetic, include chalcopyrite, cubanite, pentlandite, pyrrhotite, and minor amounts of bornite. They occur disseminated in the silicates and as small interstitial masses. The ratio of copper to nickel is about 3.8:1, based on 66 chemical analyses of rock samples from various outcrops (Ref. 1, p. 702, and Ref. 2). Test Procedures: With specified exceptions, all nickel and copper tests were made by the chromo-graph method,:' which measures the intensity of a colored spot formed by a reaction between the metal being determined and special reagent paper. The intensity is then compared to the intensity of spots prepared from samples of known metal content. Details of the test procedure are outlined in another article (Ref. 4, pp. 77,78). All soil samples tested in this investigation to date have been weighed on an analytical balance. However, a volumetric scoop designed to provide about 0.1 g of soil adds to the speed and ease of testing and has been found to give satisfactory results (Ref. 5, p. 531, and Ref. 19). The size of the samples used for the tests was 0.1 g. Whenever such small samples are used there is some question as to whether they are representative of the several grams in the field sample. Many repeat tests of the samples used in this investigation demonstrated that results can be reproduced within the limits of accuracy of the method without formal mixing beyond that inherent in screening the soil fractions. Furthermore, the 0.1 g is probably as representative of the field sample as the field sample is of its area of influence. Hawkes and Lakin (Ref. 6, p. 291), who considered the general problem, compared ground and quartered bulk samples of 500 g with 5-g grab samples. They concluded that "there is no significant loss in accuracy of data by substituting grab samples for bulk samples." The term soil implies somewhat different things to the geologist, engineer, and soil scientist.' For convenience the term as used in this article refers to unconsolidated material (the mantle) overlying bed rock. Sampling Procedure: Samples were taken at 100-ft intervals along north-south traverse lines across the gabbro-granite contact. The soil (till) samples were taken at an average depth of 1 ft, which was below the high-humus surface layer and into clean till. Samples taken at 1-ft intervals down to ledge showed as high a metal content at 1 ft below the air-surface as at greater depths and in two instances were slightly higher. The till at 1-ft depth did not appear to differ from material at greater depths. Total depth to bedrock has been tested at only a few points and where measured varied from 1 to 10 ft. Aerial Distribution Contours and Profiles: Plotting of copper, nickel, and cobalt content in contour form (Fig. 2) shows that anomalous amounts of these metal ions occur in till over and closely adjacent to mineralized areas of the gabbro. Contouring nickel content alone, or the copper content, outlines the same target area. Contours of the copper content provide a more distinct anomaly than nickel because of the higher copper concentration. The traverses are rather widely spread for interpolation; however, drilling has confirmed the target area essentially as shown.

Jan 1, 1959

By Ernest A. Hersam

The small amalgamation-test of the laboratory is not always reliable as a basis for important decisions as to the character and commercial treatment of ores. The conditions of continuous practice on a large scale are not always reproduced in the brief laboratory-test. Sometimes such a reproduction would be very difficult, or impracticable; sometimes it is not even attempted, and the test is performed in a rough way, without regard to the effects of small variations in the methods pursued. The aim of a laboratory amalgamation-test is to learn how much of the gold or silver in an ore can be economically recovered by amalgamation on a commercial scale. The yield can be increased, perhaps, by additional chemical treatment or by extreme comminution of the ore; but this might cost more than the value of such increase. Moreover, in practice, other operations often follow amalgamation; and these operations may remedy a defective extraction by amalgamation, more economically than could be done by directly increasing that extraction. The test should be performed, therefore, upon a uniformly-sized material, and under conditions that are precisely known and capable of being exactly duplicated, or purposely modified. The crucial feature of the test is the effectiveness of the contact made between the gold or silver of the ore and the amal-gamating-surfme. This may depend upon the temperature; the duration of contact; its force, due to gravity or impact; the hardness, plasticity or liquidity of the amalgam; the roughness or other quality of the amalgam-surface; the scouring action of the ore upon the amalgam, or the suspension and

Jan 1, 1905

By C. W. Hoffman, W. R. Hinken

Probability sizing, a fairly recent development in the field of particle separation, is now under investigation to determine its value to the mining industry. The method employed is probability sizing through a vibrating network of decreasing screen openings. While probability sizing is not new, the possibility of using it for coarse, highly abrasive ores at high tonnage rates has only recently appeared to merit practical consideration. The advantages of probability sizing vs. conventional screening methods are: (1) reduced space requirements for a high capacity unit, (2) fewer moving parts, (3) more versatile product sizing selection, (4) reduced maintenance requirements, and (5) lower power consumption. Screening is a necessary part of virtually every crushing circuit. At some stage in the process, particle separation becomes a controlling factor in the operation. The basic type of screen in use today is the vibrating unit, which usually employs a mechanical vibrator of one type or another to obtain sizing and material transportation across the screen surface. The vibrating screen has its limitations and inherent problems. Operators must live with: (1) high maintenance — worn rods, screen cloth, spacers, cross members, chutes, etc., (2) low capacity in relation to floor space requirements, (3) extreme noise levels, (4) high power consumption, and (5) structural problems in the screens and supporting steelwork. All plant operators have become tolerant of these drawbacks through the years, and little innovation has been developed in screening. The vibrating screens in use today incorporate improved metals, more efficient power sources, better lubricants, and increased manufacturing skill, but the basic machine itself remains unchanged. Screening is a relatively small part of most plants' operating expenditures and gets relatively little attention from research and development. The HiProb sizer is a new approach to the problem of particle separation on a high tonnage basis. Developed by Dr. Fredrik Mogensen (president of Fred-rik Mogensen, A. B. of Djursholn, Sweden), this unit has found its primary acceptance in the sizing of wet or dry materials in the fine ranges. Climax has been working cooperatively with Mine and Smelter Supply Company, the United States licensee, on the adaptation of the sizer to obtain consistent recovery at a high tonnage rate on a — 3/8-in. separation. This was a new application for the sizer and opened up a whole new field of problems. Dr. Mogensen's machine utilizes the control of normal probability of falling particles and is designed to eliminate some of the undesirable features of vibrating screens mentioned earlier. The machine is light, compact, consumes less power than conventional screens and has a low noise level. PRINCIPLES Probability sizing is a process designed to allow solid particles to segregate with a high ratio of the materials reporting into their respective size fractions with a minimum amount of misplaced particles. Sizing is defined in this paper as segregation of dry solid particles into fractions of comparable sizes. Probability is usually estimated as the ratio of the number of ways an event might occur vs. the total number of allowable ways. The probability process described depends upon the fact that particles falling through a network will segregate across the sizing surface even though all openings are larger than the particles. The degree of segregation is dependent upon the particle size and the trajectory applied by the sizing network to each falling particle. Specifically, the process utilizes a maze made up of screen cloth to control the path of a particle in free fall as illustrated in Fig. 1. The usefulness of this probability sizing process is dependent upon the degree of efficiency of an effective maze. Primarily, a maze must be extremely flexible to meet the multi-faceted nature of solid particles and, at the same time, the degree of control must be reproducible so that a desired separation is predictable. Dr. Fredrik Mogensen has formulated a maze that is not only flexible and easily controlled, but, mote importantly, easily adaptable to a working machine.

Jan 1, 1970

By J. E. Tobey, David R. Mitchell, J. H. Kerrick

DETAILED knowledge of purchase specifications established by coal consumers is essential to the successful design and operation of a coal-preparation plant. ANTHRACITE Specifications should be considered by both the consumer and the producer as guide posts outlining a certain path for both to follow. Plants differ, therefore specifications must differ to meet the varied requirements. Specifications are not of themselves a "cure all" or an assurance of complete satisfaction in fuel utilization. Specifications should not be regarded as a license to engage in an orgy of buying any and all coals that may happen to fall within their limits. This would lead to the indiscriminate mixing of coals which may vary in their burning characteristics. Different coals of equal grade tend to be individualistic in their burning characteristics. Rather, specifications should be used to help simplify coal selection, to reduce the actual number of coals used over a given period; in other words, to have a stabilizing influence on coal or customer turnover. In the final analysis, the use of too many coals per year in a plant is costly to both the consumer and the producer. Too many sources usually result in higher operating and maintenance costs for the consumer and higher sales costs for the producer. Finally, the statement included in the report accompanying the com¬prehensive Coal Selection Charts prepared by the Coal Committee of the National Association of Purchasing Agents, and released in 1938, is significant. This Committee worked diligently for a period of 15 years seeking to find a scientific and reliable method for selecting coals. The two following paragraphs are quoted from its report: More important is the fact that no ready-made classification or set of specifications can possibly replace the judgment and resourcefulness of an experienced combustion engineer or fuel technologist, nor the usual need for full scale plant tests under the exact conditions of use before a final decision is made ...

Jan 1, 1943

By J. E. Tobey, David R. Mitchell, J. H. Kerrick

DETAILED knowledge of purchase specifications established by coal consumers is essential to the successful design and operation of a coal preparation plant. ANTHRACITE Specifications should be considered by both the consumer and the producer as guide posts outlining a certain path for both to follow. Plants differ, therefore specifications must differ to meet the varied requirements. Specifications are not of themselves a "cureall" or an assurance of complete satisfaction in fuel utilization. Specifications should not be regarded as a license to engage in an orgy of buying any and all coals that may happen to fall within their limits. This would lead to the indiscriminate mixing of coals which may vary in their burning characteristics. Different coals of equal grade tend to be individualistic in their burning characteristics. Rather, specifications should be used to help simplify coal selection, to reduce the actual number of coals used over a given period; in other words, to have a stabilizing influence on coal or customer turnover. In the final analysis, the use of too many coals per year in a plant is costly to both the consumer and the producer. Too many sources usually result in higher operating and maintenance costs for the consumer and higher sales costs for the producer. Finally, the statement included in the report accompanying the comprehensive Coal Selection Charts prepared by the Coal Committee of the National Association of Purchasing Agents, and released in 1938, is significant. This Committee worked diligently for a period of 15 years seeking to find a scientific and reliable method for selecting coals. The two following paragraphs are quoted from its report: More important is the fact that no ready-made classification or set of specifications can possibly replace the judgment and resourcefulness of an experienced combustion engineer or fuel technologist, nor the usual need for full scale plant tests under the exact conditions of use before a final decision is made.

Jan 1, 1950

By J. E. Tobey, David R. Mitchell, J. H. Kerrick

DETAILED knowledge of purchase specifications established by coal consumers is essential to the successful design and operation of a coal-preparation plant. ANTHRACITE Specifications should be considered by both the consumer and the producer as guide posts outlining a certain path for both to follow. Plants differ, therefore specifications must differ to meet the varied requirements. Specifications are not of themselves a "cure all" or an assurance of complete satisfaction in fuel utilization. Specifications should not be regarded as a license to engage in an orgy of buying any and all coals that may happen to fall within their limits. This would lead to the-indiscriminate mixing of coals which may vary in their burning characteristics. Different coals of equal grade tend to be individualistic in their burning characteristics. Rather, specifications should be used to help simplify coal selection, to reduce the actual number of coals used over a given period; in other words, to have a stabilizing influence on coal or customer turnover. In the final analysis, the use of too many coals per year in a plant is costly to both the consumer and the producer. Too many sources usually result in higher operating and maintenance costs for the consumer and higher sales costs for the producer. Finally, the statement included in the report accompanying the comprehensive Coal Selection Charts prepared by the Coal Committee of the National Association of Purchasing Agents, and released in 1938, is significant. This Committee worked diligently for a period of 15 years seeking to find a scientific and reliable method for selecting coals. The two following paragraphs are quoted from its report: More important is the fact that no ready-made classification or set of specifications can possibly replace the judgment and resourcefulness of an experienced combustion engineer, or fuel technologist, nor the usual need for full scale plant tests under the exact conditions of use before a final decision is made . . .

Jan 1, 1943

By S. A. Falconer

D. A. Dahlstrom—Mr. Falconer has done an admirable job of proving the adaptability of the cyclone to the beneficiation of a very difficult size range in the preparation of coal. The addition of the cyclone to other coarser size methods makes it economically feasible for the coal operator to rigorously control his marketed products at a high and uniform quality level down to the 48 and 100-mesh size. Furthermore, it should be pointed out that by utilizing the cyclone as a deslimer or classifier the —3 200-mesh fraction, usually containing large percentages of undesirable material, can also be quickly and cheaply removed from the clean coal. Advantages of such a step are several: (1) Reduction of ash content, especially where clays and slimes are serious. (2) lower moisture content of the final coal. (3) easier mechanical and thermal de-watering of the fine coal due to higher cake and bed permeabilities. The author has indicated that adjustable control means should be installed on the cyclone operating on a heavy-media slurry. This also holds true for the cyclone acting as classifier, deslimer or preliminary dewatering agent and is worthy of special emphasis. Such controls, while not necessarily automatic, must be simple, rapidly adjustable, and require a minimum of attention. Three methods available to the operator are (1) throttling valve on the feed line, (2) back pressure valve on the overflow, and (3) adjustable diameter underflow nozzles. The first two, while fulfilling the specifications, suffer the disadvantage of restricting capacity if severe throttling is necessary. The third method has little effect on capacity and easily maintains an underflow of correct moisture content. Several designs are possible, but two find the widest application. The first was developed by Robert Piros of The Truax-Traer Coal Co. and uses a simple casting containing two holders for underflow nozzles which can be rotated about an axis parallel to the cyclone centerline. These holders retain the underflow nozzle flush with the conical section walls. When a change in nozzle diameter is required, the new size is placed in the spare holder and rotated into place. A second design utilizes a soft rubber cylindrical tube which is compressed or released from three sides, causing the inner diameter of the tube to decrease or increase. This is achieved by using a machine-threaded,

Jan 1, 1951

By S. A. Falconer

D. A. Dahlstrom—Mr. Falconer has done an admirable job of proving the adaptability of the cyclone to the beneficiation of a very difficult size range in the preparation of coal. The addition of the cyclone to other coarser size methods makes it economically feasible for the coal operator to rigorously control his marketed products at a high and uniform quality level down to the 48 and 100-mesh size. Furthermore, it should be pointed out that by utilizing the cyclone as a deslimer or classifier the —3 200-mesh fraction, usually containing large percentages of undesirable material, can also be quickly and cheaply removed from the clean coal. Advantages of such a step are several: (1) Reduction of ash content, especially where clays and slimes are serious. (2) lower moisture content of the final coal. (3) easier mechanical and thermal de-watering of the fine coal due to higher cake and bed permeabilities. The author has indicated that adjustable control means should be installed on the cyclone operating on a heavy-media slurry. This also holds true for the cyclone acting as classifier, deslimer or preliminary dewatering agent and is worthy of special emphasis. Such controls, while not necessarily automatic, must be simple, rapidly adjustable, and require a minimum of attention. Three methods available to the operator are (1) throttling valve on the feed line, (2) back pressure valve on the overflow, and (3) adjustable diameter underflow nozzles. The first two, while fulfilling the specifications, suffer the disadvantage of restricting capacity if severe throttling is necessary. The third method has little effect on capacity and easily maintains an underflow of correct moisture content. Several designs are possible, but two find the widest application. The first was developed by Robert Piros of The Truax-Traer Coal Co. and uses a simple casting containing two holders for underflow nozzles which can be rotated about an axis parallel to the cyclone centerline. These holders retain the underflow nozzle flush with the conical section walls. When a change in nozzle diameter is required, the new size is placed in the spare holder and rotated into place. A second design utilizes a soft rubber cylindrical tube which is compressed or released from three sides, causing the inner diameter of the tube to decrease or increase. This is achieved by using a machine-threaded,

Jan 1, 1951

By Eldon Dunlap

RECENTLY the producing branch of the petroleum industry has shown a considerable and growing interest in the quantitative determination of the water, oil, and gas content of cores as it relates to estimation of reserves, to providing a, basis for judging as to the feasibility of applying various secondary-recovery and pressure-maintenance processes, and in general to the operation of producing properties. For years it has been known that oil, gas and water coexisted in the pores of many semidepleted oil measures. It is now generally believed that oil-bearing beds contain connate water in amounts varying up to 50 per cent of the pore space. However, only lately has reliable direct evidence of the extensive occur-rence of connate water in oil sands been published. Very little work has been done to ascertain the effect of this water on the productivity and ultimate recovery of oil from wells. It was the intent of this research to study the influence of this indigenous water on the flow of oil through unconsolidated sands, with the hope that the results could be applied qualitatively to the flow of oil in underground reservoirs partly saturated with water. Obviously, ultimate recovery estimates based on the void volume of the reservoir will be in error unless a, conscious or unconscious correction is made for the volume of water occupying a portion of the effective pore volume of the reservoir- Evaluation engineers in the past have applied arbitrary saturation and recovery factors that may inadverently have compensated for the failure to consider specifically the water content. Nevertheless, the evidence available indicates that these factors need revising if each is to be independently correct and truly representative. Although the mere fact that the connate water renders part of the pore volume of the reservoir incapable of containing oil is of primary importance to the operator, it is also of interest to know whether and to what extent the presence of varying amounts of interstitial water influ-ences the permeability of the sand to oil.

Jan 1, 1938

By T. F. Archbold, R. E. Grace

The use of interference colors is the simplest experimental way to determine the approximate film thicknesses of oxides formed on liquid metals. A large amount of data under various conditions can be obtained easily and quickly. The oxidation of liquid lead has been studied by several investigators,1-5 but all of these studies have been on thick oxide scales. The purpose of- this work was to investigate the range 0 to 2000A film thickness. By using electron diffraction, Jenkins concluded that the surface of the film on molten lead is yellow orthorhombic PbO with the (110) planes parallel to the surface. This was confirmed by Shimaoka and Yamai7 who also used electron diffraction from the melting point of lead to 730°C. The rate of oxidation of pure liquid lead in air was measured in the temperature range 453" to 643°C by observing the rate of formation of interference colors on static lead baths. Approximately 170 g of lead (99.999 pct) was placed in a refractory dish having a working surface of 13.8 sq cm. This crucible was placed in a deep-well resistance furnace open to the atmosphere. The times at which the various interference colors appeared, after wiping the molten metal surface clean, were determined at 453", 533°, 600°, and 643°C. Elapsed times were determined with a stopwatch and all colors were estimated by eye. Bath temperatures

Jan 1, 1959

By Nicholas Kowall, K. L. Clark

Maximum furnace efficiency and close control of final steel composition demand that the steel melter be able to follow closely the variations in the carbon content of the bath. For many years, the fracture test has been used by melters for estimating carbon, because it requires only a very short time to pour a sample into a chill mold, cool it in water, break it with a sledge hammer, and observe the fracture. After long experience gained from the observation of such fractures, most melters are able to estimate the carbon content of steel with amazing accuracy, particularly in the lower carbon ranges. Notwithstanding the utility of the fracture test, there has been widespread interest in other methods for rapid carbon determination to supplement it and to ensure against possible errors in human judgment. Several methods, based upon measurement of magnetic properties as functions of carbon content, have been described in the literature1-6 and have been found to give reasonably reliable results in plant practice when scrap composition is known and controlled. A difficulty encountered in the use of analyzers of the magnetic type is that alloying elements do change the relationship between carbon content and magnetic properties and it becomes necessary to establish calibration curves for all the expected combinations of alloy residuals that are not oxidized in the bath. The work reported here is of an investigation of a method that is independent of alloying elements normally present in constructional steels, being based upon the relatively well-known fact that in plain carbon and low-alloy steels martensitic hardness is a function of carbon content alone. Work oF Other Investigators In I938, Burns, Moore and Archer published & curve? that showed that the hardness of martensite is a function of carbon from 0.10 to 0.50 per cent and is independent of grain size or alloy content, at least within the rallge found in construc-tional S.A.E. compositions. Some of the deeper hardening types investigated, however, contained carbon to a maximum of only o.2o per cent. Using this principle, Kern8 developed a method of rapid carbon determination for S.A.E 43xx steels, His procedure consisted of chill-casting a test bar that was large enough to give a a that could be prepared for Brinell indentations. The test piece was quenched immediately after solidi fication was complete and hardness readings were taken, so that the carbon content could he read from an experimentally determined curve giving the relationship between as-quenched hardness and carbon. Experimental Work Although Kern's work had been done with a relatively deep-hardening steel, it

Jan 1, 1944

By K. L. Clark, Nicholas Kowall

Maximum furnace efficiency and close control of final steel composition demand that the steel melter be able to follow closely the variations in the carbon content of the bath. For many years, the fracture test has been used by melters for estimating carbon, because it requires only a very short time to pour a sample into a chill mold, cool it in water, break it with a sledge hammer, and observe the fracture. After long experience gained from the observation of such fractures, most melters are able to estimate the carbon content of steel with amazing accuracy, particularly in the lower carbon ranges. Notwithstanding the utility of the fracture test, there has been widespread interest in other methods for rapid carbon determination to supplement it and to ensure against possible errors in human judgment. Several methods, based upon measurement of magnetic properties as functions of carbon content, have been described in the literature1-6 and have been found to give reasonably reliable results in plant practice when scrap composition is known and controlled. A difficulty encountered in the use of analyzers of the magnetic type is that alloying elements do change the relationship between carbon content and magnetic properties and it becomes necessary to establish calibration curves for all the expected combinations of alloy residuals that are not oxidized in the bath. The work reported here is of an investigation of a method that is independent of alloying elements normally present in constructional steels, being based upon the relatively well-known fact that in plain carbon and low-alloy steels martensitic hardness is a function of carbon content alone. Work oF Other Investigators In I938, Burns, Moore and Archer published & curve? that showed that the hardness of martensite is a function of carbon from 0.10 to 0.50 per cent and is independent of grain size or alloy content, at least within the rallge found in construc-tional S.A.E. compositions. Some of the deeper hardening types investigated, however, contained carbon to a maximum of only o.2o per cent. Using this principle, Kern8 developed a method of rapid carbon determination for S.A.E 43xx steels, His procedure consisted of chill-casting a test bar that was large enough to give a a that could be prepared for Brinell indentations. The test piece was quenched immediately after solidi fication was complete and hardness readings were taken, so that the carbon content could he read from an experimentally determined curve giving the relationship between as-quenched hardness and carbon. Experimental Work Although Kern's work had been done with a relatively deep-hardening steel, it

Jan 1, 1944

By H. T. Van Ellis

(Butte Meeting, August, 1913.) THE following is a brief description of the cost accounts in effect at the. mines of the Anaconda Copper Mining Co. The accompanying chart, Table I., shows the distribution of labor, materials, and redistributable accounts to the various operating and repair accounts. OPERATING. Development. Owing to the fact that every mine has its peculiarities in formation and methods, the question of arriving at what charges should be included under the head of " Development " is one that has occasioned considerable thought and comment, and in the cost sheet of the Anaconda Copper Mining Co. this, account has been sub-divided to show the expenses of breaking ground, tramming, timbering, shaft sinking, up-raising, and cutting stations and skip chutes. It has always been the policy of the company to charge the entire cost of this work to operations monthly, whether the mines are producing or not, also the cost of the opening up of new mines; and only surface construction and additions are charged to capital, excepting original machinery installed for operating pumping plants. Explanations of the charges to the various sub-accounts follow: Breaking Ground-Work performed in driving cross-cuts from the station, or cross-cuts from the drifts, and diamond-drill work, for the purpose of discovering or developing ore bodies. Tramming.-Cost of tramming waste produced by cross-cuts mentioned above. When waste is hoisted to another level to be used for filling, the charge to this account stops at the station, and when hoisted to surface the charge includes tramming from the collar of the shaft to the dump. Timbering.-Cost of timbering the cross-cuts mentioned above includes the cost of handling and framing such timbers, but does not include the wages of station tenders and top carmen.

Jan 6, 1913

By O. D. Sherby, J. E. Dorn

SEVERAL years ago Zener and Hollomon1 suggested that the flow stress of metals might be related to the temperature and strain rate in accord with the functional equation: s=s(eeh/rt) [1] at the same state. The Zener-Hollomon relation also contains the significant tacit implication that the energy for activation, AH, is substantially independent of the state of the material. The utility of this method for correlating creep data with tensile data was illustrated in a preliminary reportL on the effect of alloying additions on the secondary creep rates of binary a solid solutions in aluminum as shown for Al-Mg alloys in Fig. 1. Not only do the secondary creep and ultimate tensile properties for each alloy correlate well on this plot but in addition the activation energy, AH = 17,900 R cal per mol. is identical for the various compositions. Similar types of curves correlating the secondary creep rates at 422°K (300°F) with ultimate tensile data were also obtained for a series of dilute a solid solutions of copper, germanium, zinc, and silver in aluminum. In all cases the activation energy for creep, AH, was found to be equal to about 17,900 R cal per mol independent of the type of alloying element or its concentration. The coincidence of the activation energies for these alloys is probably due to the fact that the activation energies for creep are rather insensitive to small composition changes. Alloying additions, as shown in Fig. 1, however, increased the stress necessary to obtain equivalent values of ee?H/RT. Thus the stress level in curves of the type represented by Fig. 1 gives the relative creep strengths of the various alloys. Inasmuch as these correlations between creep and tensile data were obtained for creep tests at 422°K, it appeared advisable to ascertain the range of secondary creep rates and temperatures over which correlations between creep and tensile data could be made on the basis of the Zener-Hollomon relationship. If, for example, the activation energy changes with different ranges of creep temperatures, the utility of the proposed analysis would be severely weakened. But if AH is constant not only for all dilute alloys, but also over wide ranges of temperature and secondary creep rates, confidence will be developed not only in the broad utility of the Zener- Table I. Chemical Analysis':' and Grain Size of Alloys Mean Grain Residual Impurities. Wt Pct Diam- Alloying Atomic-eter, Element Pct Si Fe Cu Mg Mn mm Aluminum 99.987 0.003 0.003 0.006 0.001 0.25 Magnesium 1.617 0.003 0.004 0.006 0.26 Copper 0.101 0.003 0.003 0.0006 0.001 0.29 Zinc 1.616 0.003 0.005 0.007 0.001 '0.26 * The authors express their appreciation to the Aluminum Company of America for the preparation and chemical analyses of these alloys. Hollomon relationship, but also in its theoretical justification. The following investigation was instituted in order to ascertain the range of validity of the Zener-Hollomon relationship when it is applied to creep and tension data. Materials, Equipment, Technique In view of rather extensive correlatable data now available on the plastic properties of a series of binary a solid solutions of magnesium, copper, germanium, zinc, and silver in aluminum, these alloys were again selected for the present investigation. Previous reports have already covered the tensile properties of these alloys at subatmospheric temperatures," and at elevated temperature," as well as their fatigue properties at atmospheric temperatures, nd their creep properties at 422°K.' Resistivity measurementsa as well as precision lattice constant determinations? evealed that these alloys are a solid solutions. After homogenization. 0.100 in. thick sheets were rolled to 0.070 in. in thickness and recrystallized to about the same grain size. Since the details of these treatments were reported previously," they will not be repeated here. In order to appropriately reduce the scope of creep tests in this investigation only the typical alloys shown in Table I were investigated. All creep specimens were machined such that their tensile axes were in the rolling direction of the 0.070 in. sheet stock. The gage section was 2 in. long and 0.500 in. wide with a reduced section of 3.50 in. Creep strains were measured by means of rack and pinion type extensometers with a sensitivity of 0.005 in. per division. Readings were estimated to 1/10 of

Jan 1, 1953

By H. A. Sawin, C.M. Romanowitz

Placer dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were *designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry's history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. Deep Dredging At Hammonton, calif., yuba consolidated Gold Fields has, within= the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 80 ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7½-cu. ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1943

By C. M. Romanowitz, H. A. Sawin

Placer dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were *designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry's history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. Deep Dredging At Hammonton, calif., yuba consolidated Gold Fields has, within= the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 80 ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7½-cu. ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1943

By Eldon N. Dunlap

Recently the producing branch of the petroleum industry has shown a considerable and growing interest in the quantitative determination of the water, oil, and gas content of cores as it relates to estimation of reserves, to providing a basis for judging as to the feasibility of applying various secondary-recovery and pressure-maintenance processes, and in general to the operation of producing properties. For years it has been known that oil, gas and water coexisted in the pores of many semidepleted oil measures. It is now generally believed that oil-bearing beds contain connate water in amounts varying up to 50 per cent of the pore space. However, only lately has reliable direct evidence of the extensive occurrence of connate water in oil sands been published. Very little work has been done to ascertain the effect of this water on the productivity and ultimate recovery of oil from wells. It was the intent of this research to study the influence of interstitial water on the flow of oil through unconsolidated sands, with the hope that the results could be applied qualitatively to the flow of oil in underground reservoirs partly saturated with water. Obviously, ultimate recovery estimates based on the void volume of the reservoir will be in error unless a correction is made for the volume of water that occupies a portion of the effective pore volume of the reservoir. Evaluation engineers in the past have applied arbitrary saturation and recovery factors that may inadverently have compensated for the failure to consider specifically the water content. Nevertheless, the evidence available indicates that these factors need revising if each is to be independently correct and truly representative. Although the fact that the connate water renders part of the pore volume of the reservoir incapable of containing oil is of primary importance to the operator, it is also to his interest to know to what extent the presence of varying amounts of interstitial water influences the permeability of the sand to oil.

Jan 1, 1938