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By Gerald Edmunds

The liquidus line on the phase diagram for temperature versus composition of a binary alloy system, representing the boundary between the homogeneous melt and the heterogeneous melt plus solid, besides showing the minimum casting temperature, is useful in determining eutectic compositions and the presence of phase changes such as peritectic transformations and intermetallic compound formation. The liquidus often is determined by the cooling curves of thermal analysis, but supercooling frequently occurs, so that the temperature values are low. An alternative method, the one employed here, consists of determining the composition of samples of the melts from heterogeneous alloys that are at equilibrium. With meticulous attention to fundamentals, this method is capable of yielding values limited in precision only by the temperature control and measurement and analyses. Apparatus By starting, on the one hand, with a supersaturated solution and allowing the excess solute to precipitate, and on the other hand with an undersaturated solution and allowing more of the solute to be dissolved, both processes being carried out at the same temperature, equilibrium may be said to have been approached from both directions; that is, from above and from below. If the analyses show the solutions to contain the same fraction of solute when equilibrium has been approached from above and from below, one is assured that equilibrium has been reached in both determinations. In order to carry out such determinations, it is necessary to have equipment for obtaining and maintaining the proper temperatures; for holding the melt, measuring its temperature, and stirring, sampling and analyzing it. The present work required temperatures from 420° to 900°C. The equipment consisted of: A furnace, which was a vertical alundum tube, 13/4-in. diameter inside and 18 in. long, wound with Nichrome resistance wire, surrounded by a 3-in. i.d. alundum tube wound for 5 in. of length near each end, for compensation of end losses. This muffle was contained in an iron shell with Transite end plates, the intervening space being packed with Sil-O-Cel. The outside dimensions of the furnace were 12 in. in diameter by 181/2 in. in height. Heating current was supplied by three independent variable-voltage transformers (Variac), one connected to each end compensator winding and manually controlled, and one connected to the main winding through a series resistance of 5 ohms. This resistance was short-circuited for

Jan 1, 1944

By Edward A. Capillon

The problem of scrap is probably of greater importance in the pro duction of gold, silver and other precious metal alloys than is the case for base metals and alloys. Remelting of gold and silver scrap in the shop is always a costly process because it involves losses by oxidation of thc base metals in the alloy with a consequent increase in the percentage of precious metal. This is most marked in the case of alloys containing the low-boiling and easily oxidized metals zinc and cadmium. Since it is not generally feasible to assay every shop remelt it follows that the possibilities of lost values arc large. So-called rolled gold and silver plate is made by soldering or welding the precious metal alloys onto base metal. Such plate when defective must be sent to the refinery to recover the gold and silver values, thereby involving additional expense in the form of refining charges. It is obvious therefore that gold and silver alloy scrap is an important item in manufacturing cost and must be reduced to a minimum. The amount of scrap resulting from poor metal will depend on a number of factors in the manufacturing operations, not least among which is the melting and casting practice. The notes which follow are a summary of experiments performed at various times to determine what variables in melting and casting are influential in the production of sound alloys. Metals Used in Alloying Pure gold, that is, gold of 24-kt. value is very much too soft for jewelry purposes and it is therefore always alloyed with other metals to increase its hardness and resistance to wear. For the same reason pure silver is seldom used alone but is alloyed with copper, the standard alloy being sterling silver which contains 92.5 per cent. silver and 7.5 per cent. copper. The metals commonly used for alloying gold are silver, coppcr, nickel and zinc. In addition to hardening the gold and producing alloys of different karat values the use of these metals makes possible the production of gold alloys of various colors such as yellow, red, green and white. The gold-silver-copper alloys vary in color from light greenish-white through green, yellow and red depending on the relative amounts of the three metals. Green or yellowish-green colored golds

Jan 1, 1930

By C. H. Frick

BEFORE presenting the main topic, as indicated by the title, this paper will give some of the high-spot history of the anthracite industry. INTRODUCTION The earliest recorded use of anthracite was in the year 1762, when a company of Connecticut pioneers discovered anthracite near the present city of Wilkes-Barre, Pennsylvania. In the early days of the industry, the coal was used only in large pieces, the smaller pieces being discarded to waste piles. As the output of the industry increased during the nineteenth century, uses were developed for somewhat smaller sizes down to and including pea coal. Early in the twentieth century, sizes smaller than pea, such as No. I and No. 2 buckwheat, were developed and used. Accu¬mulations of piles from current mining operations were still increasing; that is, the smaller sizes were still being discarded. It was not until after World War I that preparation of the so-called steam sizes began commercially. This was largely No. 3 buckwheat, which is screened through a 3f 6-in. round-hole screen and over a 332-in. or a M 6-in. round-hole screen. USE OF SMALLER SIZES The earliest recorded experience in burning the small sizes of anthracite for steam production, other than in a previous small experiment, was in the year 1913, when a size known as "boiler fuel" was introduced-a mixture of rice and barley (No. 2 and No. 3 buckwheat). At that time, there was still no method known for burning the still smaller sizes. TABLE I.-Screen Sizes Used for Small Sizes of Anthracite [ ] At the present time, various sizes down to and including No. 5 buckwheat are being prepared (Table I) from current mining and bank-reclaiming operations. The sizes of No. 4 and No. 5 buckwheat and smaller are generally being used for steam production in large boilers. The Pennsylvania Power and Light Co., following World War I, used 680,000 tons of anthracite per year, burned largely on stokers for production of steam, whereas for the year 1943 it used 1,940,000 tons of stoker sizes and 400,000 tons of smaller sizes for grinding and burning in pulverized form. This company has taken a very large part in the research and development of ways and means of burning these smaller sizes of anthracite, including the grinding of the very small sizes. This paper is particularly concerned with the latter feature.

Jan 1, 1946

By Thomas C. Baker

The Fresnillo unit of the Mexican Corporation, S. A., is situated at the old historic mining town of Fresnillo, Zacatecas, Mexico, 33 miles north of the city of Zacatecas and 750 miles south of El Paso, Tex. It is connected with the Mexican Central Railroad by a broad-gage spur 5.5 miles long, belonging to the company. Fresnillo is surrounded by a broad semi-arid valley on the Mexican plateau at an elevation of 7300 ft. The monotony of the valley is broken by a number of low hills. Near the town of Fresnillo is the " Cerro de Proano," an elliptic-shaped hill of gentle slope, rising out of the plain to a height of 325 ft.; the base has a major axis of 4500ft. and a minor axis of 3000 ft. In this hill is the silver-bearing mineralized zone, which comprises past and present mining operations. (In this paper money values are expressed throughout in American currency and the dry short ton of 2000 lb. is used.) The discovery of ore at Fresnillo is credited to the Spaniards, soon after the conquest of Mexico. There are, however, no authentic records to show the amount of work done by them. In 1832, the State of Zacatecas took possession of the property and operated it. Evidently the rich ore above the horizon of the plain had been fairly well exhausted, as an old record states that thirty shafts were used for unwatering, requiring 300 men and 1500 animals. The cost of pumping alone amounted to $150,000 a year. In spite of the fact that cheap convict labor was used, mining costs were reported as $20 and treatment charges as $20 per ton. In 1835, the state was obliged to seek foreign capital, and a company formed in London secured a controlling interest in the property. This new organization sank two shafts, one 11 by 19 ft. and the other 10 by 15 ft., to unwater the mines and extract ore. Large Cornish pumps were purchased in England and hauled overland from Vera Cruz; engines also were installed to operate the arrastres (grinding mills) in the patio process of ore treatment. Substantial buildings were erected and operations were conducted on a large scale, for that period. The patio, the largest in Mexico, was 1150 ft. square. The costs for 1850 were:

Jan 1, 1925

By Richard E. Gray, Robert W. Bruhn

INTRODUCTION Underground coal mining has been practiced in the United States for over 200 years. Much early mining was not as efficient as today and unrecovered coal pillars, often of variable size and spacing, remain to support the overlying strata after mining has ceased. These conditions constitute a potential subsidence problem since the pillars will eventually fail. The possibility of subsidence above an abandoned mine must be anticipated except when total extraction has been achieved, permitting subsidence concurrent with mining, or when large coal pillars which have an adequate factor of safety for long-term support remain in the mine. Coal is found in 37 states and is mined under- ground in 22 states (HRB Singer, Inc. 1980). Underground mining produced 40 percent of the 7.5 x 10" kg. (830,000,000 tons) of U. S. coal mined in 1980 (Keystone, 1981). There are approximately 70,000 abandoned or inactive underground coal mines in the United States and approximately 2000 active underground coal mines (NAS/NAE, 1975). Underground coal mining in the United States will eventually extend beneath 162 x 109m2 (40 million acres). Thirty-two billion m2 (8 million acres) are already undermined (HRB Singer, 1977). Subsidence due to mining has affected more than 8 x 109m2 (2 million acres) in 30 states, 7.7 x 10 9m2 (1.9 million acres) from bituminous coal mining and the remainder primarily from anthracite mining (Johnson and Hiller, 1979). Subsidence from all other underground mines has affected only 6.9 x 10 7m2 (17,000 acres) (Johnson and Miller, 1979). Subsidence appears to be particularly widespread in two states, Pennsylvania and Illinois, although serious subsidence is found in localized areas of numerous other states (HRB- Singer, Inc. 1977). Host underground coal mining has occurred in the eastern United States, where the U. S. Bureau of Mines estimates 1.6 x 109m2 (398,600 acres) in urban areas in 18 states (out of a total 1.7 x 109m2 1418,400 acres]) are threatened by subsidence (Table 1) (Johnson and Miller, 1979). Pennsylvania and West Virginia contain 57 percent of the total threatened areas. The U.S. Bureau of Mines estimates the total cost to stabilize the estimated 1.7 x lo9m2 (418,400 acres) of subsidence prone urban land located above abandoned mines is over $12.5 billion (1978 costs).

Jan 1, 1982

By James H. Gardner

It is a striking fact in the domestic oil statistics of 1926 that with an increase in production of 13,000,000 bbl., there was nevertheless a lowering of both crude and gasoline stocks. It was the first reduction in the last eight years. In 1918, the last year of the war, stocks of crude and refined products were lowered 16,000,000 bbl., but in the following seven years stocks were increased 350,000,000 bbl., due largely to the phenomenal crop of major pools discovered in 1923. The big pools of that year coming as they did through some curious coincidence unprecedented in the oil industry, not only piled up a sudden increase of stocks at a time of large imports but they continued to add their quota to an increase of storage until 1926. With an increase in crude production, according to the Oil and Gas Journal, from 760,000,000 bbl. in 1925 to 773,000,000 bbl. in 1926, and an increase in cracking facilities of about 30 per cent., there was a draft on storage of 22,300,000 bbl. of crude and 1,000,000 bbl. of refined. Crude and refined changed from a stock surplus of 29,130,000 bbl. in 1925, to a deficit of 23,300,000 bbl. in 1926, or a differential total of 52,430,000 bbl. With a stock reserve of 541,700,000 bbl. of crude and refined oil in 1925, the amount drawn from storage was not large in percentage but it marked an important contrast with previous years. It meant a year of stability in prices and a sound condition in the industry. The indicated domestic demand increased from 695,120,000 in 1925 to 743,360,000 in 1926, (7 per cent.), 1 per cent. of which was due to California fires. Most of the gain in consumption was due to the increased demand for gasoline. There was little comparative change in the combined imports of crude and refined oil. There was an increase in exports of 17,200,000 bbl. (1.5 per cent). A petroleum balance sheet for 1926 and a comparativl sheet for the last two years, production in the United States by major districts and pools, and a list of imports are given in Tables 1 to 5, inclusive.

Jan 1, 1927

By H. J. Engell, W. Frank, A. Seeger

SWISHER and ~urkdogan' have determined the solubility of oxygen in a iron. They found it to lie between 7 and 10.5 at.-ppm at 881°C. Using earlier work on the permeability of oxygen in bcc iron,' they arrived at the following expression for the diffusion coefficient of oxygen in bcc iron (k = Boltzmann's constant, T = absolute temperature): where The present authors3 have estimated the activation energy QD for the diffusion of oxygen in bcc iron in two different ways from various experimental data.2'4"6 In the first method QD + QS (QS = heat of solution for oxygen in bcc iron in thermodynamic equilibrium with solid FeO) was determined either from data on the reduction of FeO with hydrogen4 or from measurements of the internal oxidation of silicon in a iron.5 Using an estimate of QS from investigations of the oxygen solubility in 6 iron2 (in thermodynamic equilibrium with molten FeO) Q = 0.97 ev or Q~ = 1.06 ev were obtained. From the magnitude of these values it was concluded that oxygen migrates interstitially in bcc iron and that consequently the activation energy QD for diffusion of oxygen has to be identified with the migration energy E~ of the oxygen atoms. Therefore, for a second estimate of QD from the diffusion coefficient of oxygen in 6 iron at 1450" ~,~ we used the theoretical expression:3 (a = length of elementary cube edge, m = mass of interstitially diffusing atom, sM = migration entropy), applicable to diffusion via the octahedral interstices in bcc lattices. Taking sM = 3.8 k (migration entropy of carbon in a iron), one gets Q = 0.90 ev. Hence, in Ref. 3 Q = (0.98 *0.1) ev was proposed as a plausible value for the diffusion energy of oxygen in bcc iron. sM + 9 (9 = solution entropy for oxygen in bcc iron) can be determined from permeability measurements.4 Using the above-mentioned estimate for QS , the following expression for the solubility c (atomic concentration) of oxygen (in thermodynamic equilibrium with solid FeO) in bcc iron was derived: From [4] the solubility of oxygen at 881°C in a iron is calculated to lie between 0.6 and 12 at.-ppm. We see that the solubility determined by Swisher and

Jan 1, 1969

By W. C. Williamson, J. L. Shugert

DUST control is receiving considerable attention today by most companies as is evidenced by the many articles, papers, and books written on the subject. ~~~t of the larger mining companies have given this subject close study and have established standardized methods of control. W. C. WILLIAMSON and J. L. SHUGERT, Members AIME, are Assistant Ventilation Engineers, Anaconda Copper Mining Co., Butte, Mont. San Francisco Meeting, February 1949. TP 2705 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Jan. 17, 1949. Permissible Limits: The question of permissible dust limits has been considered by medical commit-ties and various national conferences. Dr. Lanza, of the Metropolitan Life Insurance Company, Dr. Le-roy U. Gardner, of the Saranac Laboratory for the study of Tuberculosis, and Dr. Pancoast, of the University of Pennsylvania, have been members of these committees. A statement included in reports of some of these committees is as follows: Concentrations to which the dust must be induced in order to be safe have not been absolutely determined. There is evidence that a concentration of more than five million particles per cubic foot of a highly siliceous dust is dangerous. Therefore, it may be stated that it is desirable to avoid counts of more than five million particles of dust containing free silica to a high percent in working places. Table I gives the safe limits adopted by several state industrial codes. Table I. Safe Limits of Dust Concentration Limit Million Particles State Per Cu Ft Cal~fornia Silica (less rhdn 10 pct) 50 Silica (10 pct to 50 pcr) 10 Sllica (more than 50 pct) 5 New York Frcc silica less than 10 pet by weight of rock formation 100 Frec silica more than 10 Pct by weight of rock formation 10 Oregon Silica (over 40 pct) 5 Silica (20 to 40 pct) 10 Silica (10 to 20 pet) 20 Wisconsin 15 million particles, under 10 nicrons in longest dimension per cu ft of air, when the quartz content of the dust is 35 pct, variations in free silica content will make proportional inverse changer in this standard Close attention to dust concentrations has been given to mining operations, since these operations are frequently carried on at considerable distances

Jan 1, 1951

By J. L. Shugert, W. C. Williamson

DUST control is receiving considerable attention today by most companies as is evidenced by the many articles, papers, and books written on the subject. ~~~t of the larger mining companies have given this subject close study and have established standardized methods of control. W. C. WILLIAMSON and J. L. SHUGERT, Members AIME, are Assistant Ventilation Engineers, Anaconda Copper Mining Co., Butte, Mont. San Francisco Meeting, February 1949. TP 2705 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Jan. 17, 1949. Permissible Limits: The question of permissible dust limits has been considered by medical commit-ties and various national conferences. Dr. Lanza, of the Metropolitan Life Insurance Company, Dr. Le-roy U. Gardner, of the Saranac Laboratory for the study of Tuberculosis, and Dr. Pancoast, of the University of Pennsylvania, have been members of these committees. A statement included in reports of some of these committees is as follows: Concentrations to which the dust must be induced in order to be safe have not been absolutely determined. There is evidence that a concentration of more than five million particles per cubic foot of a highly siliceous dust is dangerous. Therefore, it may be stated that it is desirable to avoid counts of more than five million particles of dust containing free silica to a high percent in working places. Table I gives the safe limits adopted by several state industrial codes. Table I. Safe Limits of Dust Concentration Limit Million Particles State Per Cu Ft Cal~fornia Silica (less rhdn 10 pct) 50 Silica (10 pct to 50 pcr) 10 Sllica (more than 50 pct) 5 New York Frcc silica less than 10 pet by weight of rock formation 100 Frec silica more than 10 Pct by weight of rock formation 10 Oregon Silica (over 40 pct) 5 Silica (20 to 40 pct) 10 Silica (10 to 20 pet) 20 Wisconsin 15 million particles, under 10 nicrons in longest dimension per cu ft of air, when the quartz content of the dust is 35 pct, variations in free silica content will make proportional inverse changer in this standard Close attention to dust concentrations has been given to mining operations, since these operations are frequently carried on at considerable distances

Jan 1, 1951

Before hostilities between the colonies and Great Britain began in 1775, most of the coal used in the northern colonies undoubtedly came from England, with some supplies for New England coming from Nova Scotia; in New York and Philadelphia the supply came from England with some from the Virginia mines. Table 4 shows data obtained from the Public Records Office, London, for the years 1768-1772. This table supplements Table 3, which records only a few shipments from Richmond in 1768 and 1769; as there was no coal shown in imports coastwise in 1769 in Table 4, and only two small shipments in 1769 in Table 3, there may have been a cessation of shipments then, for some reason not now known. In Table 5 are incorporated all the data in the first two, with a few earlier years, and an attempt is made to estimate the amount used during the period by allowing for coal exported. With the small total tonnages a carryover from one year to another makes a large difference in the annual figures, but over the period the result may be fairly indicative of the amount of coal used before the Revolution. According to present day ideas the quantity of coal used during that war was very small, but apparently it was badly needed for the manufacture of the small amount of munitions made in those troublous days. As Philadelphia was the leading city along the Atlantic coast, and as the southeastern part of Pennsylvania was the leader in making iron, as well as making rifles, it is natural that the first call for coal came from there. At first, procuring supplies such as coal, powder and sulphur was handled by a Committee of Safety, which later became a Council of Safety and later this committee's functions were turned over to a Board of War, functioning with the military organization of the colony. . The organizations of the several colonies were slightly different, but Pennsylvania, Maryland and Virginia followed this scheme generally. Later in the war the Congress began to make arms for the troops, and it is difficult now, with the few records available, to tell when the state functions ceased and those of Congress began. On December 5th, 1775, the Committee of Safety

Jan 1, 1942

By C. H. Frick

Before presenting the main topic, as indicated by the title, this paper will give some of the high-spot history of the anthracite industry. Introduction The earliest recorded use of anthracite was in the year 1762, when a company of Connecticut pioneers discovered anthracite near the present city of Wilkes-Barre, Pennsylvania. In the early days of the industry, the coal was used only in large pieces, the smaller pieces being discarded to waste piles. As the output of the industry increased during the nineteenth century, uses were developed for somewhat smaller sizes down to and including pea coal. Early in the twentieth century, sizes smaller than pea, such as No. I and No. 2 buckwheat, were developed and used. Accumulations of piles from current mining operations were still increasing; that is, the smaller sizes were still being discarded. It was not until after World War I that preparation of the so-called steam sizes began commercially. This was largely No. 3 buckwheat, which is screened through a 3/16-in. round-hole screen and over a 3/32-in. or a 1/16-in. round-hole screen. Use of Smaller Sizes The earliest recorded experience in burning the small sizes of anthracite for steam production, other than in a previous small experiment, was in the year 1913, when a size known as "boiler fuel" was introduced—a mixture of rice and barley (No. 2 and No. 3 buckwheat). At that time, there was still no method known for burning the still smaller sizes. At the present time, various sizes down to and including No. 5 buckwheat are being prepared (Table I) from current mining and bank-reclaiming operations. The sizes of No. 4 and No. 5 buckwheat and smaller are generally being used for steam production in large boilers. The Pennsylvania Power and Light Co., following World War I, used 680,000 tons of anthracite per year, burned largely on stokers for production of steam, whereas for the year 1943 it used 1,940,000 tons of stoker sizes and 400,000 tons of smaller sizes for grinding and burning in pulverized form. This company has taken a very large part in the research and development of ways and means of burning these smaller sizes of anthracite, including the grinding of the very small sizes. This paper is particularly concerned with the latter feature.

Jan 1, 1949

By G. S. Rice

Some form of longwall mining is generally used in Continental Europe; also in Great Britain where the coal is weak and friable, or the coal bed provides material for pack walls and filling, or where the bottom is soft and squeezes up easily, or the roof is pliable, or the bed is thin and brushing provides building material, or the thick multiple seams are mined in layers, such as the 24-ft.seam at Weymss, Scotland, or the 10-yd. pitching bed near Coventry, England. Pillar methods have been retained in British fields where the coal beds are from 5 to 9 ft. (1.5 to 2.7 m.) thick, and free from thick partings or binders or without draw slate, which would provide waste rock for pack walls, or where the roof is hard and requires shooting to bring down, and also the bottom or floor is comparatively hard. The room-and-pillar system, by which probably 95 per cent. of the coal of the United States is produced, is now found in only a few mines in Wales, where it is known as the pillar and (single) stall method. The American room-and-pillar system is not equivalent to the bord-and-pillar or stoop-and-room system. The essential difference between the bord-and-pillar and the average American room-and-pillar system is probably due to natural conditions. The majority of the coal mines in the United States have a depth of less than 700 ft. (213 m.). The bulk of the coal mined in Europe comes from a depth of more than 1200 ft., while the deepest mines in Great Britain, Belgium, and France reach 3500 to 4000 ft. In deep workings wide spans are impracticable, even with the use of abundant timber; accordingly, as the mines in the United States become deeper the rooms will become narrower and the pillars thicker, and the bulk of the coal will be extracted on the retreat. It is generally conceded, without reference to the cost of production, that the larger the pillars left on the first mining, the more thoroughly can the coal be extracted. In Europe complete extraction is generally compelled either by the lessors or by the governments; whereas in this country it has been more largely a question of competitive cost of production, or the support of the surface in the flat farming districts of the

Jan 1, 1922

By W. L. Chu

It was observed that chromized coatings on low carbon steel could be embrittled by cooling slowly through the 475°C range. A higher chromium coating was more susceptible to embrittlement than a lower chromium coating. Coating structure and composition could be correlated with the chromizing process and prior carbon content of the steel. IT was recently observed that during the fabrication of rectangular tubings made with chromized thin gage 1010 steel, a high percentage of these parts were fractured readily in a brittle manner under the light impact-forming load employed. Ensuing investigation indicated the cause for such brittle failures to be the so-called 475°C embrittlement of the chromized coatings. The experimental work and results are briefly reported. Rectangular tube sections of normalized AISI C1010 steel, which had a cross-section of 0.5x1.5 in. and a wall thickness of 0.030 in., were chromized to yield coatings having several ranges of thickness, varying from 0.0005 to 0.002 in. Essentially, the chromizing process is similar to that described by Samuel and Lockington,' which consists of heating the parts in a sealed container at about 955°C for 4 hr in intimate contact with a chromizing mixture* samples in a hydraulic press, and the other employed a simple beam bending test of the flat sides of the tubing walls. Neither, however, was able to differentiate the relative brittleness of samples known to include both the brittle and tough types, since the impact-brittle samples were found to deform readily in a manner similar to the tough ones. Both the tough and brittle samples were studied by reflection microscopy. Observation showed that the microstructures of the steel cores were similar in content and distribution of inclusions as well as in the ferritic grain size. Results of Filar micrometer measurements showed that coating thicknesses+ of the brittle samples averaged 0.0007 in., whereas those of the tough samples averaged 0.0016 in. The micrographs of Fig. 2 illustrate the structural differences between coatings of both types of samples. By dissolving the steel cores in nitric acid, coatings of both the brittle and tough samples were isolated. Simple bending tests showed that coatings of the brittle samples cracked readily at about 90" bend, whereas those of the tough samples could be bent nearly 180° without cracking. The coatings as well as the steel cores were analyzed for carbon and

Jan 1, 1956

By Philip W. Henry

There is a difference of opinion among engineers on the subject of depreciation in general, and still more on its application to any given case The committee which was appointed by the American Society of Civil Engineers to make a report on Valuation of Public Utilities, states "There is no subject connected with valuation about which there are more diverse views than those relating to depreciation;" and in the discussion which took place upon the presentation of the preliminary report of this committee in January, 1914, that part of the report dealing with depreciation called forth more discussion than any other. From this discussion it was evident that the word "depreciation" was not always used in the same sense, so it is quite important that some definition be adopted. As this paper relates to the proper method of treating depreciation by corporations or individuals operating oil properties, both of which must make a return of annual net income to the government in connection with the income tax, it will be in order to adopt the definition of the Internal Revenue Bureau. Under date of Mar. 29, 1910, the Commissioner of Internal Revenue of the Treasury Department stated "Deduction on account of depreciation of property must be based on lifetime of property, its cost, value and use." Again, on Form 1035, "Return of Annual Net Income" for miscellaneous corporations (Section 2, Act of Congress approved Oct. 3, 1913), a deduction from gross income is allowed for "Total amount of depreciation for the year." This item is defined more particularly as follows: "The amount claimed under Item No. 5 (b) for depreciation should be such an amount as measures the loss which the corporation actually sustains during the year in the value of buildings, machinery, and such other property as is subject to depreciation on account of wear and tear, exhaustion, or obsolescence. The amount taken credit for on this account in order to be allowable should be so entered on the books as to constitute a liability against the assets of the corporation. The amount claimed under this item should not cover losses in the value of stocks and bonds. Decrease in the book value of securities owned, so far as such decrease represents a decline in the actual value of such securities, should be deducted under item 5 (a) of the return. "Where depreciation of physical property is made good by renewals, replacements,

Jan 1, 1915

By F. S. Buffington, Morris Cohen

THERE is little quantitative information available concerning the effect of applied stress, in both the elastic and plastic ranges, on diffusion kinetics. Accordingly, a program has been undertaken to determine the rate of self-diffusion in high purity iron, with the specimens subjected to stress during the course of diffusion. Results are reported here for body-centered cubic iron at a diffusion temperature of 890°C. The iron used was 99.97 pct Fe, the principal impurities being 0.002 pct C, 0.008 pct 0, and 0.02 pct Si. The surface-decrease method of self-diffusion analysis was employed, and Fe55 Mn x-radiation, 2.9 years half-life) was the tracer element. Details of specimen preparation and equipment are given elsewhere.' Disks of the high purity iron, % in. diam x 1/4 in. thick, were annealed at 890°C for 18 hr in dry hydrogen to produce large grains of a ferrite. The faces of the disks were then surface ground, and possible grinding stresses were relieved by annealing at 890 °C for 45 min in vacuum. A layer of radioactive iron 0.00186 cm thick was electroplated on one face of each disk. The self-diffusion coefficient was calculated from the ratio of the radiation intensities emanating from each of these faces, as measured before and after diffusion.' The heating was carried out in a vacuum furnace (10-5 mm Hg) controlled to ±1.5°C. The specimens were loaded in compression between two quartz disks during the diffusion runs by means of a lever system operating through a vacuum seal. Thus one quartz disk was pressed directly against the radioactive face of the specimen, minimizing the vapor loss of radioactive iron. Steady-state creep occurred in all instances where permanent deformation could be detected, but the specimens did not exhibit appreciable transverse barreling until natural strains of 0.2 to 0.3 were attained. Essentially, the diffusion took place under uniaxial stress. The plastic strain in each case was obtained at the diffusion zone by measuring the diameter d there before and after the run: e = 2 In — do Let Du equal the self-diffusion coefficient which would have prevailed at the given temperature if no stress had been applied. This was calculated from the temperature-dependence equation which had been previously determined: —59,800 D, = 6.2 exp ( I sq cm per sec* [I]

Jan 1, 1953

By Harry W. Antes

The density changes that occurred on annealing cold-worked poly crystalline molybdenum, over the temperature range of 870" to 1630 SC, were determined. Increases in density were observed for recovery and the beginning of recrystallization. An initial decrease in density was observed during re-crystallization. A second decrease in density was observed after recrystallization was completed, as the microstructure changed from columnar grains to equiaxed gvains Both decreases in density were associated with an increase in gain boundary area to volume ratio. The second decrease in density was observed in the annealing range that immediately precedes secondary recrystallization as reported in the literature VARIOUS techniques have been used to identify the recovery, recrystallization, and grain growth processes that occur when cold-worked metals are annealed. As early as 1928 work was conducted that indicated measurable density changes occur when cold-worked single crystals of brass are annealed.' More recently, work by Murphy exhibited a corre- lation of density with microstructural changes observed on annealing cold-worked polycrystalline a brass.2 Other investigators have performed similar work on a brass and copper.324 In reviewing the literature, it was apparent that density work had been limited to fcc materials. Therefore, it was the purpose of this investigation to determine the density changes that accompany annealing the cold-worked bcc material, molybdenum. Two degrees of cold work, 30 and 90 pct reduction of cross-sectional area by rolling, were selected for this study. EXPERIMENTAL PROCEDURE High-purity vacuum arc-cast molybdenum was used in this study, the principal impurity being carbon in the amount of 80 ppm. The final rolling reductions were made at 1200°C in grooved rolls to provide 5/8-in.-diam rod. Density specimens 0.500 in. diam by 1.5 in. long were machined from the bar stock. The specimens were ground to the final dimensions to provide a smooth surface. Density measurements were made using a hydrostatic technique similar to that used by ~urphy.' This type of equipment permitted weighing in liquid immediately after the air weighing. During each weighing the temperature of the air and liquid was controlled to within 0.0l°C. This was accomplished by sealing the analytical balance in an air-tight brass chamber and surrounding it by an insulated wooden case. Preheated air was circulated in the air space between the brass chamber and wooden case by means of a blower. The temperature of the

Jan 1, 1964

By T. T. Field, M. Gell

MUCH has been learned about plastic deformation behavior in crystals through the use of dislocation and slip band etchants. Slip in nickel and nickel-base alloys has previously been revealed by various etching techniques.l-5 The techniques for nickel-base super-alloys4,5 were successful only in the Udimet 700 composition in specific heat-treated conditions. The objective of this work was to develop a more generally applicable slip-band etching procedure. The nominal compositions of the alloys used in this study are listed in Table I. The low carbon Mar-M2OO alloy referred to in the text is the same as listed in Table I with the exception that the carbon content is less than 100 ppm. Mar-M200 and Udimet 700 were solutionized at 2250" and 2150°F, respectively. The aging treatment for both cast alloys was 32 hr at 1600°F and for wrought Udimet 700 was 4 hr at 1975"F, 24 hr at 1550°F, and 16 hr at 1400°F. The new slip-band etching technique consists of a chemical etch followed by a short electropolish, Table 11. Fig. l(a) shows the best result obtained on an internal section of a plastically deformed single crystal of Mar-M2OO when using a conventional slip-band etching procedure which consists of electropolishing followed by etching. Ghostlike traces corresponding to planar slip bands are observed, but the contrast between the bands and the background is weak. Fig. l(b) shows the same field using the new slip-band etching technique. After the etching step of this technique, the same ghostlike traces of Fig. l(a) are observed, but the subsequent electropolishing accentuates the slip bands and the subgrain boundaries while the rest of the matrix is being polished. Longer electropolishing obliterates the slip bands. The new slip-band etching technique makes use of the etchants and electrolytes that are normally used for a given material. Because of this, it should be possible to slip-band etch other nickel-base superal-loys by using existing etchants and electrolytes for a material and following the procedures outlined in the "Comments" section of Table 11. All the illustrative examples given below show internal plastic deformation in single crystals of regular and low carbon Mar-M200 in the solutionized and

Jan 1, 1970

By D. M. Warner, Samuel Daniels

In this paper are given the mechanical properties, determined by the Engineering Division, Air Service, U. S. A., of the 94 per cent. aluminum, 5 per cent. copper, 1 per cent. silicon alloy as 8and-cast and as heal-treated commercially to Air Service Specification No. 11,300. This particular alloy was tested both in the cast and in the heat-treated condition. Tension and tension modulus values were obtained from machined and from un-machined test specimens; while data for shear, compression, impact, torsion, and for Brine11 and Rockwell hardness were found from machined bars. Specific-gravity tests were made on sanded, un-machined specimens. The metallography of the alloy is useful as a control to heat treatment and presents some interesting peculiarities in regard to the two iron-bearing compounds. The demand for an aluminum alloy, suitable for sand castings, that may be heat treated to meet requirements of high strength, ductility, shock and corrosion resistance, and ready machineability has led to the commercial development of a proprietary material that contains from about 4 to 5 per cent. of copper, 1 per cent. of silicon, and such impurities as iron, up to a maximum of 0.75 per cent., and manganese, in quantity usually less than 0.1 per cent.' Magnesium is not generally present, for it tends to nullify the beneficial influence of the added silicon by forming magnesium silicide. The function of the silicon is to minimize casting difficulties and, structurally, to favor the precipitation of (insoluble) iron-bearing skeletons rather than the needles, which prevent full realization of tensile properties from heat treatment by stimulating intergran-ular fracture. The amount of silicon necessary to counteract the ill effect of iron is equal to or slightly more than the content of iron; the excess of silicon over this ratio, under certain conditions of heat treatment, enters with CuAl2, the principal hardening constituent, into solid solution and contributes somewhat to the strength of the alloy without appreciably impairing its ductility. The schedule of heat treatment of this alloy depends on the design (cross-sections) of the castings and on the nature of the tensile properties desired. Essentially, the process consists in heating the material at about 950" F., quenching in oil or in cold or boiling water, and artificially aging

Jan 1, 1926

By James A. Beck

Phosphorus is essential to all life processes and therefore to the existence of man. In this role, there are no substitutes for phosphorus. In a commercial sense, phosphorus and its compounds are important to our entire economy, both in agricultural and in nonagricultural applications. The world's phosphate requirements are met by mining phosphate rock. Two general types of phosphate ore deposits are commercially important-igneous and sedimentary. Igneous deposits generally are more difficult and more expensive to mine than sedimentary deposits and are of importance mainly in the USSR, North Vietnam, and South Africa. Approximately 15% of current world production comes from these deposits. Over 80% of world production is from the pellet sedimentary deposits of Florida, North Africa, and elsewhere. Guano, the second type of sedimentary deposit, is produced in South America but is declining in world importance. Production Production of phosphate rock was reported in 31 countries during 1972. However, about 80% of world output is produced in three countries: the United States, the USSR, and Morocco (Table 15.2A.1). Throughout the past 20 years, between 75% and 80% of world output has been produced by these three leading countries. Between 1948 and 1972 world production increased at an average rate of 10% per year, from 21.5 to 103.9 million tons. Growth was quite steady at 6% per year through 1963, but in the 5 years from 1963 to 1968 production increased 11 1/2% per year. Since 1968 production has increased about 91/2 % annually. Since 1948 several new phosphate mining ventures have been developed in Senegal, Togo, Israel, and North Vietnam. In the same period, only one important phosphate deposit, on Makatea Island, has been mined out and closed. A number of other phosphate mining ventures are in various stages of planning and development. The most important new venture in phosphate mining is in the Spanish Sahara. New mining capacity is expected also in Syria, and serious investigation is in progress in Peru, India, Turkey, and Australia. Major expansion of existing operations is expected in Morocco, Russia, Tunisia, Israel, and Jordan. Consumption Historical data on world consumption of phosphate rock are not available in detail, but world consumption of phosphate has closely paralleled world production in the past two decades. The United States is the world's largest user of phosphate rock, accounting for 30%

Jan 1, 1976

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947