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By William Phillips

This paper was prepared at the request of Capt. A. F. Lucas, Chairman of the Institute's Committee on Petroleum and Gas, as a preliminary discussion of the fuel oils which are used-in the Southwest, giving their composition, uses, prices, etc. Should it afterward be found desirable, these observations may be extended over a larger area. In this connection, J suggest that samples of all kinds of fuel oils, including various tars, from all over the country, be collected and sent to some testing laboratory for examination. We would be glad to extend the facilities of the laboratory of the Bureau of Economic Geology and Technology, at the University of Texas, for this purpose, and to do the work free of all expenses. After such examinations are made and the results classified and studied, it would be possible to suggest specifications under which fuel oils should be purchased, and to arrive at some conclusions respecting the uniformity of such specifications, with due allowance for climatic and other more or less local conditions. While such specifications do not now depart from a common standard anything like as much as specifications for illuminating oils do, yet there are discrepancies which should not exist. Some of these discrepancies are of lessening importance, due to the pronounced tendency to distill every crude oil that can be distilled with profit, and to send residues into the market for fuel oil. Residues from the refineries are of a more uniform chemical composition and physical nature than many of the crude oils that go direct into the fuel market. The lower flash-point oils are disappearing from the fuel trade and in place of them we are getting oils that do not flash below 220° F., nor burn under 320° F. In the Southwest, it has not been long since we had fuel oils that flashed at as low as 110° F., and the average flash point over a considerable period was 151° F. Now we are using oil that flashes at about 30° lower than the former burning point, which was 256° F. In other words, the flash point has risen, on the average, 100°; and the burning point about 75°.

Jan 6, 1914

By Norman Braly

THE following is not offered as an extended paper on the subject of shaft sinking, but more as a description of the present practice of shaft sinking in the Butte district. The Anaconda company is sinking at present the following shafts: Badger State, Moose, Mountain Consolidated, Mountain View, Pennsyl-ania and Tropic. At the High Ore and Leonard mines the shafts are being deepened by sinking two compartments and will subsequently be enlarged to the full working size by raising alongside the compartments which are now being sunk. The following other companies are also sinking shafts: Butte-Alex Scott, Butte and Zenith City, Boston and Corbin, North Butte Mining Co. (at the Granite Mountain shaft), East Butte Co. (at the Pittsmont shaft), and the Rainbow Lode Development Co. Shaft Timbering. The shafts of the Butte district, generally speaking, are of three compartments arranged in a straight line, and the pump compartment is usually arranged for the use of an auxiliary hoisting cage, as well as for carrying pipe lines and electric wires. The Rainbow shaft is practically square, and of three compartments. The length of the shafts outside of timber will vary from 17 to 22 ft., and the width from 6 ft. 6 in. to 7 ft. Fir timber is used in all shaft work. This may be obtained from the Coast, when it is known as Coast fir, or from Idaho or Montana, when it is known as native fir. The native fir seems to be the tougher wood, but it is hard to obtain clear lumber in the lengths and sizes necessary for shaft work. Some of the shaft timber of the Anaconda company has been treated with creosote as a preservative, but this practice has not come into general use, as the ground of the camp is so heavy that the timbers usually crush before they rot. The size of timber most .used for shaft work is 12 by 12 in., while 10 by 12 in., 10 by 14 in. and 12 by 14 in. are used for some of the special pieces around the shaft. The new Leonard shaft is timbered with 14 by 14 in. wall plates throughout.

Jan 8, 1913

By E. C. Bain, E. S. Davenport

When annealed carbon, or low-alloy, steels are suitably heated the ferrite (alpha iron solid solution) and the carbide, of which they are composed, react together to form a single solid solution of carbon (and other elements) in gamma iron (austenite). This reaction may begin only at a certain minimum temperature called the eutectoid temperature. The first ferrite and carbide thus to react together do so in definite proportions known as the eutectoid composition. With further heating the original constituent that is present in excess is dissolved in the austenite until exhausted. Whether the excess constituent be ferrite (hypoeutectoid steels) or carbide (hypereutectoid steels) the procedure is much the same, and upon slow cooling the homogeneous austenite expels the excess of either (proeutectoid ferrite or carbide) until the eutectoid composition is reached. Thus the reaction upon heating is strictly reversed upon cooling and accordingly an equilibrium diagram of this process may be constructed. Such an equilibrium diagram concerns itself only with the temperatures involved and with the proportions and compositions of the phases developed under conditions that suffice to permit equilibrium. Specifically it has no concern for the type or mode of distribution of these constituents nor with any extra-equilibrium conditions such as undercooling, which is almost universally encountered in metals. Actually, however, of great importance is the sluggishness toward equilibration caused by the not inconsiderable time involved in diffusion. Thus instead of the theoretical transformation point AF, we find for a certain rate of heating or cooling the use of the designations Ar and Ac (transformation temperature found on cooling and upon heating). These figures obviously have no particular meaning except as measured and expressed for a definite heating or cooling rate, and represent the apparent temperature displacement caused by an inadequate time allowance. In this paper are presented the results of a study of the time required for the transformation of austenite to ferrite and carbide at a variety of temperatures and also of the time required for the reaction austenite? martensite at the temperatures at which this reaction occurs instead of

Jan 1, 1930

By I. Cornet

INTRODUCTION WITH the greatly expanding use of magnesium during the war, it appeared necessary to the War Metallurgy Committee that the notch sensitivity of magnesium alloy extrusions be further investigated and the influence of various factors upon notch sensitivity be noted. Accordingly, the National Defense Research Committee of the Office of Scientific Research and Development placed a research contract with the University of California, supervised by the War Metallurgy Committee. The work discussed herein was done under "Restricted" Project NRC-21, reported in Dec. 1943 1; it has been released for publication by the O.S.R.D. SCOPE OF THE INVESTIGATION The notch sensitivity of various magnesium alloy extrusions was determined under axial tension, and also under several conditions of nonaxial tension. Hardness, grain size and other metallographic features were observed; stress-strain data were obtained, and the chemical compositions of the alloys were determined, in order to correlate these quantities with the notch sensitivity. Notch sensitivity under axial tensile load was determined for a few specimens which had been cold worked to various degrees by tensile prestraining. Magnesium alloy extrusions studied included Dowmetal 0, X, J, and M. For comparison, extrusions of aluminum alloy 24S-T, cast bars of aluminum alloy I22-T2, and cast bars of magnesium alloy C were also tested, The specified chemical composition of these alloys is given in Table I. All bars tested were within specification limits of composition. Investigation of the tensile notch sensitivity of various magnesium alloy extrusions, under conditions of axial and eccentric load, showed little or no correlation with stress-strain or other conventional data. Since the many uncontrolled variables present might mask significant relationships, one alloy was selected for intensive study. Magnesium 0 alloy extrusions were subjected to experimental heat treatments, and the changing hardness, tensile strength, microstructure, and notch sensitivity observed. EXPERIMENTAL TECHNIQUES AND PROCEDURES Tensile Testing Bars used for axial loading were machined with 3/8 in.- and with 1/4-in. minimum diam, and included specimens with 6o° V notches (Fig I and 2). Each V-notched specimen was roughed out carefully and then finished with a honed tool. These V-notched bars were sampled frequently, sectioned, polished, and examined at 500 X magnification to control

Jan 1, 1948

By Charles Barrett

RECRYSTALLIZATION textures-the orientations of grains after recrys-tallization-have been studied extensively not only because of their metallurgical importance but also because of the information they yield regarding the atomic movements during recrystallization. Recrystallized grains must possess the same orientations as the submicroscopic nuclei from which they have grown, and thus they offer a direct road by which experimenters can penetrate the submicroscopic realm. Many perplexing results have come of these studies, but one of the most striking has been the apparent difference in behavior between single crystals and polycrystalline grains. For example, Burgers and Louwerse have shown' that when single crystals of aluminum are deformed by compression and are recrystallized (at 600° C.) the new grains appear with orientations that are different from the orientations of the deformed crystal in which they grew; but polycrystalline specimens, on the other hand, appear to have a [110] fiber texture both before and after recrystal-lization and thus to retain their texture during recrystallization. This apparent difference between the habits of single crystals and small grains is the subject of the first part of the present research and has been explained by findings reported herein. A similar situation appears to exist for aluminum when elongated or drawn into wires, for single crystals alter their orientation2,3 and polycrystalline wires retain their texture 4,1 6 upon recrystallization after elongation (at least with high-purity alumi-num), and it is possible that the explanation found in this paper for compression can be extended to cover elongation and other types of deformation as well. The second part of this paper deals with the theory of recrystallization textures. In the most successful theory for aluminum recrystallized after compression,1 much weight is given to the orientation of highly stressed nuclei in the deformed material. Taylor's "local distortions" are assumed to exist along the slip planes where displacement has occurred, consisting of fragments rotated in a specific way with reference to the slip plane and slip direction; namely, about an axis lying in the slip plane perpendicular to the slip direction. 1,7,8,9

Jan 1, 1940

By E. S. Davenport

WHEN annealed carbon, or low-alloy, steels are suitably heated the ferrite (alpha iron solid solution) and the carbide, of which they are composed, react together to form a single solid solution of carbon (and other elements) in gamma iron (austenite). This reaction may begin only at a certain minimum temperature called the eutectoid temperature. The first ferrite and carbide thus to react together do so in definite proportions known as the eutectoid composition. With further heating the original constituent that is present in excess is dissolved in the austenite until exhausted. Whether the excess constituent be ferrite (hypoeutectoid steels) or carbide (hypereutectoid steels) the procedure is much the same, and upon slow cooling the homogeneous austenite expels the excess of either (proeutectoid ferrite or carbide) until the eutectoid composition is reached. Thus the reaction upon heating is strictly reversed upon cooling and accordingly an equilibrium diagram of this process may be constructed. Such an equilibrium diagram concerns itself only with the tempera¬tures involved and with the proportions and compositions of the phases developed under conditions that suffice to permit equilibrium. Specifically it has no concern for the type or mode of distribution of these constituents nor with any extra-equilibrium conditions such as undercooling, which is almost universally encountered in metals. Actually, however, of great importance is the sluggishness toward equilibration caused by the not inconsiderable time involved in diffusion. Thus instead of the theoretical transformation point AE, we find for a certain rate of heating or cooling the use of the designations Ar and Ac (transformation temperature found on cooling and upon heating). These figures obviously have no particular meaning except as measured and expressed for a definite heating or cooling rate, and represent the apparent temperature displacement caused by an inadequate time allowance. In this paper are presented the results of a study of the time required for' the transformation of austenite to ferrite and carbide at a variety of temperatures and also of the time required for the reaction austenite ?martensite at the temperatures at which this reaction occurs instead of

Jan 1, 1930

By T. A. Rickard

Among the names which won a world-wide fame during the golden age of the early fifties, Bendigo and Ballarat were to Australia, what the Yuba and Grass Valley were to California. The map of Victoria did not for a long time show the name of old Bendigo;* for this first and more distinctive name was replaced with the more English Sandhurst, just as the alluvial diggings gave place to quartz mining. Towards the close of last pear, steps were taken to give back the old name, as associated with the early days of rich alluvium, and more suggestive than the application taken secondhand from an English military academy. History. The first discovery of gold? in this district, was made in the autumn of 1851 ; but there has never been any certainty as to the day or the man. At that time the country around Bendigo Creek was a part of the Ravenswootl sheep-run, and its resemblance to the Forest Creek district (now Castlemaine), induced the first prospecting. It was late in November when the "rush" broke out; the shepherds left their flocks; the sylvan solitudes were disturbed by * The name of Bendigo is said to have been derived from a hut-keeper on the Ravenswood sheep-run, who on account of his fondness for " fisticuffs " was nicknamed Bendigo, after the prize-fighter of that name. It is not aboriginal, as is often supposed, but Spanish ; and equivalent to our Benedict. There must always be some confusion between the two names of Sandhurst and Bendigo; the town, and with it the gold-field, having been three tinies named. Several other old familiar names have been likewise unfortunately replaced by second-hand English ones; so that an old digger talks of Forest Creek when he means the modern Castlemaine, Mt. Ida for Heathcote, Growler's Creek for Bright, etc. I shall use the names Sandhurst and Bendigo interchangeably. The first discovery of gold in Australia, was made by E. H.Hargraves, February 12, 1851, near Bathurst in New South Wales. In August of the same year the discovery at Buninyong, near Ballarat, inaugurated the first of a series of rushes to the Victoria gold-fields.

Jan 1, 1892

By Douglas R. Piteau

Douglas R. Piteau (Engineering Geologist, De Beers Corp., Kimberley, South Africa; Presently Consulting Specialist at Nchanga Mine Open Pit, Chingola, Zambia)-Mr. Lutton is to be commeneded on his analyses and presentation of such practical data which can be used for predicting a preliminary optimum slope angle in rock. It appears that he assumes all slopes analyzed to be straight. Although there are overriding factors which affect the general slope configuration, such as structural discontinuities, ground water, and other environmental factors such as weathering conditions, regional stress, etc.,8 the affects of plan geometry under certain conditions also can be of outstanding importance. In investigation of natural slopes, for example, the effects of plan geometry are not entirely appreciated and it is in this respect that discussion is given. Experience in southern Africa has shown that the overall geometry of surface openings in rock, particularly those with dimensions of the order of 1500 ft or less, i.e., where radii of curvature of the slopes are less than, say, 700± ft, must be considered in rational slope-stability analyses. That is, the slope angle is not only a function of slope height, strength of the mass, and geometry of structural discontinuities, but also is dependent upon the plan geometry of the local slopes and the plan shape of opening. Technically speaking, adequate analysis of the slope must be three-dimensional, the third dimension being that of overall plan configuration and the analysis must incorporate its subsequent effects. However, current theories of slope stability all deal with the slope as a two-dimensional problem, i.e., a slice of unit length of an infinitely long slope is considered for a condition of plane strain. In this the tacit assumption is made that the plan radii of the crest and toe of the slope are infinite and no attention is given to the plan geometry of the slope. Clearly this is not the condition encountered in practice, particularly in open pit mining where excavations, relatively speaking, can be small and where these radii of curvature accordingly can have important affects on the slope stability. This is illustrated by the analyses of Piteau and Jennings 9 of the natural slopes which had developed around pipes of four diamond mines in the semiarid Kimberley area of South Africa. Although these slopes consist of dolerite overlying shale, the values of the slope angles of all the pits were corrected for a height of slope of 320 ft assuming that no dolerite occurred. This brought the data onto a comparable basis. These data were analyzed statistically and a clear empirical relationship, as is shown [in Fig. 4], is seen to have formed between angle of slope and curvature of slope, i.e., the angles of slope are greatest where the radii of curvature are least. Also, it can be seen that the slope of each line is different, indicating that the plan shape of the open pits also are significant, the curve being steepest for the open pits which are more circular in shape in contrast to those which are more elongated. The effect of slope curvature on the average angle of slope for the two extreme radii of 200 ft and 1000 ft for all four open pits is 39.5° and 27.3°, respectively. Hence, the concave curvature effect represents an average increase in slope angle of 12.2° from the essentially straight slope. In quantitative terms in Mr. Lutton's Fig. 2, this represents displacing the curve for a 320-ft pit from the second curve from the bottom (i.e., shale in subtropical climate) to that of the third from the bottom (i.e., quartzite in temperate environment).

Jan 1, 1972

A SELF-RESPECTING miner doesn't wash the cuttin's off his hard hat until he quits his job but, on the other hand, he keeps his lamp clean and a "spot" focus on the work at hand; the former habit is a matter of pride and the latter serves for efficient operation. During 1949, the mining engineer got all the "wet raises" for the problem of cutting costs to meet the drop of metal prices and keeping research work up to schedule when strikes tied up his work were his headaches. We note, not without admiration, that he hasn't washed the muck off his hat in token of surrender but, on the contrary, his light shines bright on the job to be done as can be seen in the following pages. Few people have any conception of how much ore and rock have to be mined and how much sorting it takes to cull a pound of copper from a 1 % ore body. These people would be even more startled to picture the problems faced by the engineer in mining and extracting, profitably, .2 oz of gold from a ton of ore as is done at some of our lode mines. Strikers have sometimes kept our researchers from their pilot plants and research laboratories, but the end of the year shows progress in making ore out of taconite and ferruginous sandstone, with a pilot plant in operation at Aurora, Minn., and plans for one near Birmingham. Unfortunately, it looks as though the mining engineer will have to make John Doe secure in mineral resources in spite of himself. Coal for steel, power, and to warm our homes is still economically possible, in spite of Mr. Lewis, through development of mechanized tools for mining. During the year, the use of continuous mining machines has extended, transport systems are being adopted to keep the ore flowing from the face to the tune of 40 tons per man-shift. Experiments in the gasification of coal are going forward in several places and may eventually extend our fuel reserves by the addition of coal seams now too lean for mining. Liquid fuel from coal, oil shale, or natural gas will become economic when oil consumption exceeds our supply; but, at less than 30 cents a ton direct cost for mining oil shale, the engineer is almost making it competitive under present market conditions. In this age of specialists, the mining engineer is still holding his own as a broad-gauge man. As a matter of fact, he is emerging from a stage of semispecialization as a coal, nonmetallics, or metal mining man to the realization that, with mechanization, mining is fundamentally materials handling whether it is fluorspar, coal, or copper ore. From the coal business come rubber-tired mechanical loaders to the iron mines of Birmingham. Shuttle cars are also finding application in the lead mines of the Tri-State. Open-pit methods are universal. Belt conveyors, trucks, Diesels, and mobile, jumbomounted drills are finding application in all types of mining. In preparation, such methods as flotation,- heavy density, electrical, and centrifugal are finding application in coal, nonmetallics, and metals. These concepts have been recognized by the mining engineer who has kept himself informed and has been quick to seize a method, machine, or a process from another operation and, with ingenuity, applied it to his problems. This, our Annual Review issue, is written by mining engineers. In gathering these separate articles under one cover for publication, we think of it as a trrbute to the tireless energy and skill of these men. They get little of the publicity accorded physicists, medical men, labor leaders, or company executives, but they supply raw materials without which industry would die.

Jan 1, 1950

INTRODUCTION Mineral explorationists range throughout the world looking for mineral deposits which can subsequently be made into mines. Mines are not found! Mines are made, usually at great cost and not often without significant risk. The process of making a mineral deposit into a mine involves a great deal more than a complete appraisal of the deposit. The making of a mine demands detailed feasibility studies of many factors, any one of which, if shown to be negative, could prevent the deposit from becoming a successful mine. Total feasibility analysis involves at least three major fields of study: (1) economic analysis, (2) financial analysis, and (3) political and environmental analysis. Economic analysis is essential for proper decision making at many stages in the evaluation of a mineral deposit. It is the yardstick by which the proposed mining project and alter- native investment opportunities are measured. Economic analysis is concerned with costs and potential profits and can be applied at various stages in the evaluation of a mineral deposit to determine if the project merits the additional funds necessary to bring about each succeeding stage. The nearer any deposit approaches to becoming a mine, the more carefully the economic analysis should be made. Financial analysis has reference to how, where, and at what cost the investment funds can be obtained to finance the proposed mine project. Political and environmental analysis has reference to the many intangible factors that can affect the making of a successful mine. These factors loom larger today than ever before, due to the profusion of government regulations and various environmental constraints now imposed upon the mineral industry both by law and by the weight of public opinion. Ultimately, many of these intangible factors can be resolved back into the economic analysis as an added cost to mine development, but more and more often they have become insurmountable obstacles to any development of certain deposits. The role of financial analysis and the analysis of intangible factors must never be underestimated when making a feasibility study upon which to base investment decisions; however, the emphasis of this text will be upon economic factors. THE PARAMETERS OF MINERAL ECONOMIC ANALYSIS The two major elements of economic analysis are: 1) The potential earnings or annual generated cash flow of the project. 2) The investment costs necessary to realize such earnings. If the difference between the two elements is large enough, i.e., if the return on investment from the cash flow is high enough to attract investors, the project can be said to be financially feasible. The mineral economic analysis must take into consideration many parameters, some of which are known or knowable within certain limits, and others of which are unknown or unpredictable within certain limits and for which estimates must be made. Because both known and estimated parameters enter into the same analytical computations, it is important for the analyst to realize the nature of each when considering the results of an analysis. Some parameters such as the grade of mining block, although estimated, are said to be known within mathematically defined limits. Other parameters, such as the future price for minerals produced, etc., can only be estimated or "guestimated" within reasonable limits for which no precision can be assigned. Known or Knowable Mineral Economic Parameters Known or knowable parameters are: 1) Tonnage and grade of valuable minerals for the total mineral body and for individual mining blocks. 2) Metallurgical recovery of valuable mineral. 3) Minimum operating life of mineral de- posit at any specified production rate. 4) Royalty schedule, property payments, etc. 5) Sales-purchasing agreements, smelter schedules, marketing contracts, etc.

Jan 1, 1980

By Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951

By Jack White

MUFULIRA Copper Mines, Limited, conducted tests on the concentrator in 1942 to determine which type of ball mill, that is, Conical Trunnion overflow or Cylindrical Grate discharge, produced the greater quantity of finished product at the lowest cost. All the data which follows applies to Mufulira ore and to the particular degree of grinding produced with the classifiers installed at Mufulira concentrator. JACK WHITE, Member AIME, is Concentrator Superzntendent, Mufulira Copper Mines, Ltd., Mufulira, Nortltern Rhodesia. Sun Francisco Meeting, February 1949. TP 2729 B. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Feb. 28, 1949. Test Procedure: The test was carried out using ball mills No. 1 to 9 (conical trunnion overflow) as the standard with which to compare No. 10 (cylindrical grate discharge mill). The conical trunnion overflow ball mills are 10 ft diam by 72 in., and the cylindrical grate discharge mill is a No. 99. Each mill runs in closed circuit with a Dorr type DSFX classifier 8 ft wide by 30 ft long. The feed is considered as being identical to all ten units. The figures for the test were derived as follows. Dry Tons Milled: For No. 10 unit (cylindrical grate discharge mill) this figure was taken from the ball mill weightometer, which was corrected for error, and then for moisture content. The weightorneter was checked twice a week during the test, and the error did not exceed 3 pct at any time. The moisture was obtained from a 24 hr moisture sample taken by the ball mill operator hourly, and placed in a closed bucket. The sample was cut where the conveyor drops the feed into the ball mill chute. The Mill Sampling Department made a daily moisture determination on this sample. To obtain the dry tons on units No. 1 to 9 the No. 10 figure was deducted from the month's dry tons milled figure, which, in turn, was taken from the crusher weightometer tons corrected for weightometer error and moisture. All tons reported in this test are short tons of 2000 lb. Ball Mill Hours: This was taken daily from the Shift Boss reports giving actual running hours for each unit. Screen Analyses: Coarse Heads (Ball Mill Feed): The monthly composite coarse heads screen analysis, done by the Mill Sampling Department, was used for the minus 200 mesh in the feed to the mills. F. X. Classifier Overflow: The ball mill operators took an hourly 24 hr sample of the overflows of units No. 1 to 9. This was a hand sample, but cut with a special cutter, and is believed to be very reliable. On No. 10 unit, an hourly 24 hr sample, and an hourly shift sample were cut in the same way. These five samples were split, filtered, and dried by the Research Department and were then wet screened on a 200 mesh screen. The average of the three shifts samples on No. 10 unit had to check the 24 hr sample within 1.5 pct to be acceptable. A weighted composite was cut each day from both 24 hr samples, and at the end of each month it was screened as follows: Wet screened through 325 mesh, the plus 325 then screened on a rotap for 15 min through 65, 100, 150, 200, and 325. This was done in duplicate. All splitting, drying, and screening of samples was done by one of the Mill Research staff, to insure accuracy. F. X. Classifier Overflow Densities: These were taken from the Shift Boss reports. Densities were taken by the ball mill operator every hour. Circulating Loads: These were taken a number of times during the test both by direct method (weighing a measured portion of classifier rake sand, and deducting moisture) and by calculation using the formula:

Jan 1, 1951

By W. Crafts, J. L. Lamont

Electrolytic extraction of carbides from quenched and tempered steel and their examination under the electron microscope were found to be helpful in understanding the mechanism of secondary hardening in alloy steels1 and the same technique has been applied to isothermally transformed steels. A preliminary survey of the utility of the method has indicated that it has promising possibilities for investigating the characteristics of pearlite and bainite. The examination of a few carbon and chromium steels has suggested that pearlite is formed with carbides of two varieties representing the lamellar and nonlamellar forms and that bainite appears to form with a structure analogous to, but significantly different from, martensite. Procedure The method used for examining the carbides was similar to that described previously.1 The chromium steels used in the investigation were made from an Armco iron base in an induction furnace and were not treated with grain-refining deoxidizers. The composition of the steels is given in Table 1. Three-inch square ingots were forged and rolled to 1/8-in. thick sheet for use in the isothermal studies. After rolling, the sheets were sandblasted to remove the rolling scale. Specimens approximately 3/4-in. wide and l1/2-in. long were prepared from the 1/8-in. sheet. They were heated in a salt "bath at 2100°F (1150°C) for 1/2 hour, transferred to other salt baths at 1300°F (704°C), 1000°F (538°C), 800°F (426°C), or 600°F (315°C), held for various periods up to 100 hr and finally quenched in water. After quenching, the specimens were cut in half, one half being submitted to microscopic examination and the other half being used for the electrolytic extraction. The isothermally transformed specimens were submitted to examination under the light microscope to determine the degree of transformation and type of structure and were photographed at 2000 X. Carbide residues were obtained by electrolyzing in 10 pet hydrochloric acid and collecting the residue in glycerine to minimize attack by the acid. As pointed out in the earlier paper, the carbides are attacked to some degree during electrolysis. This attack is relatively minor on the coarser carbides but may be quite severe on the fine bainitic carbides. The residues were washed free of the glycerine with water and alcohol. The alcohol was removed by repeated washings of amyl acetate, and the amyl acetate suspensions were transferred to stoppered vials. Debye-Sherrer X ray diffraction patterns of the residues were prepared using a chromium target. A selection based on the microstructure and X ray data was then made of samples considered most illustrative of the progress of transformation. Electrolytically extracted residues from these samples were submitted to examination by W. D. Forgeng and A. C. Jenkins on the electron microscope at the Research Laboratories of The Linde Air Products Co. and electron micrographs were made at X 5000 and enlarged to 25,000 X. The micrographs obtained with both the light and the electron microscopes are mounted side by side in the accompanying figures. Although no effort was made to establish the T-T-T diagram for the chromium steels, the times for holding in the salt bath were adjusted with the intention of making the samples represent different stages in the transformation at temperatures of 600, 800, 1000, and 1300°F. The degrees of transformation and results of X ray examination are shown in Tables 2—4. The carbide designations refer to orthorhombic or trigonal carbide types and are not meant to imply that the compositions are exactly as indicated. X ray Diffraction Data The occurrence of Fe3C and Cr7C3 in these steels was found to be consistent with published data on comparable chromium steels. Steels A and B with relatively low ratios of carbon to chromium contained Cr7C3 in the pearlitic structures and FeuC in the

Jan 1, 1950

By Paul Gordon

The distribution coefficient, k, of interest in zone refining is generally defined as the ratio of the solid to the liquid solubilities of one element in another at the normal melting point of the solvent. In practice, distribution coefficients are frequently calculated from solubilities at temperatures well below the melting point of the solvent, e.g., at a eutectic temperature. This is done because of the general lack of accurate data on the liquidus and solidus curves in most phase diagrams. The practice is equivalent to assuming that the liquidus and solidus curves are straight lines in the temperature range between the melting point of the solvent and the temperature at which data for the solubilities are known. It is recognized that the k values so derived may be in serious error; we have found evidence in recent zone-refining operations that a case in point is silicon in aluminum. The phase diagram found in the literature for silicon in aluminum indicates a eutectic on the aluminum-rich side at 12.5 at. pct Si and at 577°C,1 and a solid solubility of silicon in aluminum at this temperature of 1.59 pct.2 It also shows the liquidus and solidus as nearly straight lines from 577°C to the melting point of aluminum (660°C)1,2 The k value obtained from this data is 0.13. This is almost exactly the same as the k value (about 0.14)' obtained in a similar way for copper in aluminum. Yet we have found by mass-spectrograph analyses* of three bars of zone-refined aluminum that copper is efficiently removed from aluminum by zone refining but silicon is not. Two of the three bars of aluminum were zone-refined in our laboratories, the other in the laboratories of AIAG Metals, Inc. The former two bars, designated bars PG2 and PG3, were 30 in. long by 1 in. in diameter during refining; each was processed in a high-purity graphite boat* and kept under refining stage of the process consisted of thirty unidirectional zone traverses at rates of about 1 to 2 in. per hr, with a zone length of 1 to 2 in. At approximately the middle of the refining period a 6-in. piece of the impurity-laden finishing end of the bar* was cut off and replaced with new start- ing material. Following refining, the purest 20-in. section of the bar (obtained by discarding approximately 2 in. of the starting end and 8 in. of the finishing end of the bar) was leveled and then analyzed. The AIAG bar was 16 in. long and about 1.2 in. in diameter; it was purified by eight passes of a 2-in. zone at a rate of 1.6 in. per hr, the bar being contained in a sintered alumina boat lined with loose alumina powder of the highest purity. Data for the silicon and copper contents of the three refined bars are given in Table I. Derived data for the efficiencies of the zone-refining operations are listed in Table 11. It is clear from these two tables that the copper is removed much more efficiently—by a factor of 10 to 50 times—from aluminum during zone refining than is silicon. For example, in bar PG3 the thirty zone passes reduced the copper content from 10 to 0.02 ppm, i.e., by 500 times, whereas the same treatment lowered the silicon content by only a factor of about 16. It seem virtually certain, therefore, that the distribution co efficient, k, for silicon in aluminum is well above that for copper and is very likely not much below unity. This also implies that the solidus line on the aluminum-rich side of the A1-Si phase diagram is very probably appreciably curved near the aluminum melting point since the phase-diagram data indicate such curvature is unlikely in the liquidus.

Jan 1, 1965

By G. T. Smith, R. C. Moyer

Cadmium, along with other impurities such as lead, gallium, germanium and indium, is characteristically found associated with zinc ores, the average ratio of zinc to cadmium being about 200 to 1. The increasing demand for cadmium has led most zinc producers to develop a means for its recovery. The separation and recovery of this byproduct constitutes an important development in the metallurgical field. Zinc sulphide concentrates, principally from domestic sources, are processed at Donora Zinc Works and they assay approximately 60 pct zinc, 30 pct sulphur, 0.37 pct cadmium, and 0.8 pct lead. Concentrates are roasted to convert zinc sulphide to zinc oxide. Sulphur evolved as sulphur dioxide is used to manufacture 60º Baume sulphuric acid by the lead chamber process. The roasted calcine is sintered to remove the remaining sulphur and volatile impurities. Fume from the sintering machines is collected by Cottrell pre-cipitators and sent to the Cadmium Department for cadmium removal. The sintered product, a crude zinc oxide, is then mixed with coke and salt and charged into horizontal retorts for reduction to zinc metal. Furnace residues are then passed through a Waelz kiln which recovers practically all of the remaining zinc as an oxide. This oxide is collected by Cottrell precipi-tators and returned to the sintering plant for further processing. The following discussion covers operating procedure, and equipment used in the Cadmium Department for the manufacture of cadmium and its byproducts. (See attached Flow Sheet, Fig 1.) The production of cadmium may be divided into several steps, each of which will be discussed separately. These steps are as follows: 1. Raw Materials—Sinter Plant Cottrell dust, acid sludge and sulphuric acid. 2. Sulphating—Separation of lead. 3. Purification—Separation of impurities other than zinc. 4. Precipitation—Separation of cadmium from zinc. 5. Briquetting—Preparation of sponge for charging to furnace. 6. Smelting—Distillation of cadmium metal. 7. Remelting—Casting into commercial shapes. 8. Recovery of Rare Metals—Separation of germanium, gallium and indium. Raw Materials The raw material, or Sinter Plant Cottrell dust, from which the cadmium is manufactured, is recovered by Cottrell units collecting fume from two Dwight-Lloyd sintering machines. An average analysis of this dust is 36 pct zinc, 1.5 pct sulphur, 17 pct lead, and 5.5 pct cadmium, with associated elements in smaller amounts. A lesser tonnage of material removed from the sintering machine fans, and called fan cleanings, is also treated for cadmium. It is lower grade material and analyzes approximately 41 pct zinc, 1.2 pct sulphur, 14 pct lead, and 4 pct cadmium. Cottrell dust and fan cleanings are delivered to the Cadmium Department by monorail crane in bell-bottomed dump buckets. Acid sludge and 60" Bauine sulphuric acid are received from the Acid Department. The acid sludge is a mixture of lead sulphate, sulphuric acid and zinc sulphate, and is recovered mainly from the floors of the acid chambers or storage tanks. The acid sludge is delivered to the Cadmium Plant in a lead lined bucket. The 60º Baume sulphuric acid is delivered through a lead pipe to a lead lined storage tank. Sulphating Four well ventilated lead lined sulphating tanks, 8 1/2 ft in diam by 12 ft deep, are used. These are steel tanks lined with 20 lb lead on the bottom and 16 lb lead on the ides, equipped with Hastelloy side agitators. 15 hp motors drive 24 in. propellers at 450 rpm. There are water connections to stuffing boxes to keep sludge from entering the bearings. The tanks are covered with removable wooden lids and are connected to marine plywood

Jan 1, 1950

By Donald L. Katz

The compressing of gas into a partially depleted gas drive oil reservoir to bring the contents to a single phase miscible with gas is proposed as a process worthy of serious study. The compressed gas and vaporized oil would be recovered by cycling. This paper explores the Possibility of recovering oil from reservoirs of 6,000 ft or more by this method. INTRODUCTION Many new methods for increasing the recovery of oil from reservoirs are under consideration in the various research laboratories. One method which should be considered for recovering oil From partially depleted gas drive fields is to compress gas into the reservoir until the contents become a single phase miscible with more gas. Natural gas then may be cycled through the reservoir to recover the oil. This procedure ideally will eliminate the capillary effects of the liquid ill the porous media and should permit recoveries similar to those obtained in cycling operations, probably of the order of 80 per cent. This paper explores the information now available for the prediction of the performance of such a project and discusses the problems likely to be encountered in recovering oil by cycling. For many years dry natural gas has been passed through oil reservoirs to recover natural gasoline which would vaporize from the oil. Recently a process of recovering more of the high boiling hydrocarbons has been described by employing high pressures at which the heavier hydrocarbons have a higher volatility. Nevertheless. this reported method does not eliminate the differential nature of the vaporization process and capillary forces are still present in the reservoir. The process under discussion in this paper would minimize differential vaporization and would remove capillary forces which are responsible for low oil recoveries. Higher gas pressures are required to accomplish this removal of the miniscus, but it does not follow that more gas is required for oil recovery. The process is limited to reservoirs deep enough to permit the use of the required pressure — Probabably 6,000 it and deeper. DESCRIPTION OF THE PROCESS Briefly, the process is to compress natural gas into an oil reservoir which is partially depleted so that it has an oil saturation in the range of 40 to 60 per cent of the hydrocarbon pore space. The injected gas will pass through the channels formerly carrying gas to the wells to build up the pressure and dissolve in the oil phase. This injection will continue without production from the field until the reservoir reaches a single phase. If one were to select unique conditions such that the gas composition, the reservoir oil composition and the Per cent saturation would bring the mixture in the reservoir to the critical temperature and pressure simultaneously the phase behavior accompanying injection is shown by Fig. 1. The oil-gas mixture in the depleted reservoir has a critical temperature at C,. Addition of natural gas to the mixture will give a new critical temperature for each addition. Mixture No. 3 is shown to have its critical temperature at the reservoir temperature and reservoir pressure at P: the critical pressure. Under ideal behavior, the entire reservoir would be at iti critical temperature and pressure and the entire hydrocarbon content would be at a single phase. If the reservoir content were a binary system of methane and a heavy constituent. it would be miscible with more methane. However, for complex mixtures the critical locus does not confine the two phase region, and the dew point curves of Fig. 1 occur above the critical locus. If fluid which was injected into the rezervoir at the critical point were of the

Jan 1, 1952

By V. Patarini, M. Salkind

BECAUSE of its high strength and stiffness coupled with low density, continuous boron filament is showing great promise as a potential reinforcement for both metals and nonmetals. Of primary consideration for structural applications is the fatigue behavior of boron fiber composites. Composite fatigue behavior is dependent upon three factors: 1) fiber fatigue behavior, 2) matrix fatigue behavior, and 3) the effect of the interface and fiber ends. The purpose of this note is to describe the behavior of boron filament in rotating beam fatigue. The filament, described by Talley' and Galasso, Salkind, Kuehl, and Patarini,' was made by the hydrogen reduction of boron trichloride on a hot tungsten filament. Testing was performed using a commercial rotating beam wire fatigue tester* described by Votta.~ One end of the test sample is held in a chuck in a synchronous motor. The sample is bent around in a "U" shape, and the other end inserted into a carbide bushing. The maximum strain occurs at the bottom of the "U" where the radius of curvature is a minimum. The strain is adjusted by adjusting the length of the specimen and the location of the bushing. Samples were rotated at 3600 rpm and an automatic shutoff device used to determine the number of cycles to failure. The measurements were accurate to the nearest 360 cycles. The fatigue behavior is summarized in Fig. 1 for fifty-one samples of 4.8 x 10"3 in. diam fibers and ten samples of 3.2 X 10"3 in. diam fiber. The two sample groups represent two different filament production runs and two different nominal size groups. Both groups exhibited nominal tensile strengths of 400,000 psi. No significant difference in the fatigue behavior of the two sample groups is apparent. The scatter in the data is not unexpected for a brittle filament which exhibits a tensile strength distribution with a standard deviation of 10 to 15 pct. Because the test procedure provides a known outer fiber strain in the sample, the maximum stress was calculated assuming a Young's modulus of 58 x 10' psi. This value was selected as being representative of several nominal fiber sizes.2 All of the samples which failed exhibited sudden, catastrophic fracture, and no permanent deformation. One sample was stopped after 1.66 x lo8 cycles at *200,000 psi. The fiber exhibited no apparent damage or deformation. It is apparent that the fatigue resistance of boron filament is quite good. It can therefore be speculated that the fatigue behavior of boron fiber composites would be controlled by the "weaker links", namely matrix fatigue resistance and interface behavior. In fact, preliminary investigations have revealed that this is true for boron-reinforced aluminum? It should be noted that the data presented herein may not necessarily reflect the tensile fatigue behavior of boron filament. It is well-known that boron filament has a radial residual stress pattern with the outer surface in ~om~ression.~ Because of this fact, tensile failure is normally nucleated in the interior of the filament.5 Since the data presented herein is for a rotating beam test in which only the outer surface of the fiber is stressed to the maximum value, it would be incorrect to assume that the failure mechanism and fatigue behavior would be the same if the fiber were tested in tensile fatigue. The fact that many of the data are well above the nominal 400,000 psi tensile strength would tend to support this argument. The authors would like to express their appreciation to the United Aircraft Research Laboratories for supporting this investigation and for permission to publish the results. A portion of this effort was supported by the Air Force Materials Laboratory under contract AF33(615)-2125.

Jan 1, 1968

By R. F. Mitchell, G. F. Dillon, A. F. Armington

In this paper the present methods of boron preparation are discussed with emphasis on the iodide intermediate. Several methods of boron triiodide preparation were investigated, the most satisfactory method being the preparation from the elements. A yield of 60 pct based on iodine is found for this method. Zone-refining, sublinzation, distillation, and fractional-recrystallization results are given for the boron triiodide, with distillation producing the best results. Several methods of decomposition are also discussed. The purity of the resulting iodide boron approximates the puritjl of the best available bromide boron, ELEMENTAL boron has been suggested as a possible high-temperature semiconductor by several workers,' but has never been exploited as a device material. The principal difficulty with boron is the unavailability of pure crystalline material from which accurate information on electrical properties could be obtained. Most boron crystals presently produced are very imperfect in structure and of questionable purity, particularly in relation to carbon content. Several methods of boron purification have been reported in the past few years, all of which produce material with a purity better than 99.0 pct. Most of these methods have been the subject of a recent review by Starks' and will not be elaborated upon in this report. However, of all the purification methods reported, only those using halide intermediates produce a purity approaching that usually required for electronic applications. The bromide intermediate has been most used for the preparation of high-purity boron.'y3 This material is easily prepared from the elements and can be effectively purified by distillation and decomposed by hydrogen reduction. The resulting product contains a trace of silicon, and an uncertain amount of carbon, believed to be approximately 100 ppm. It has been recently reported by Russian workers that high-purity boron can be produced using an iodide intermediate.' In the work cited, boron is reported free of spectro-scopically determined impurities, but the reported detection limits are quite high (- 100 ppm). In addition, no data is given on the carbon content of the product. In the present paper, boron triiodide was chosen as the intermediate for several reasons. First, carbon tetraiodide, which could be a major impurity, is reported to be unstable above 200" and hence could be easily separated from the boron triiodide by its decomposition, since temperatures greater than this are attained during the purification process. The low melting point (45°C) and boiling point (20B°C) and the relatively high vapor pressure of boron triiodide should allow the application of any physical purification technique to the material. The possibility of a lower decomposition temperature than that used for the bromide process is another advantage which might cut down extraneous contamination. The iodide, however, does have certain disadvantages when compared with the bromide, the production cost being the most significant. This is the result of a higher halide-to-boron weight ratio, which means more intermediate must be employed. The cost of the starting material is also greater for the iodide. This cost could be considerably reduced however, by recycling the iodine byproduct. Epitaxial growth of boron, which is probably the best method of producing junctions, should be possible with either the bromide or the iodide as the intermediate. However, this has been attempted only on a limited scale up to the present time because of the unavailability of good crystals. The authors have been relatively successful in laying down a heterotaxy of boron on silicon, however. This will be reported at a later date. Several crystalline modifications have been reported for boron, four of which are now believed to be real.' These are the a! and 0 rhombohedral and the a, and 0 tetragonal. In general, the a! rhombohedral is the low-temperature form and the B rhombohedral is the high-temperature form. The tetragonal structures are detected between 1000"

Jan 1, 1964

By T. Szaplonczay, D. E. Light

The auto-oxidation of sulfur dioxide to form sulfuric acid in commercially produced barren and recycle liquors, containing 1.66 and 4.56 gpl Fe respectively, from Elliot Lake was studied. A gas mixture, containing 3.5% SO2 and air was introduced into these liquors through a frit at a rate of 6.0g SO2 per hour per liter of solution. The barren and recycle were heated to 60°C. At the end of the first 24 hours of the process the concentrations of acid produced were 158 and 220 gpl H2SO4 for the barren and recycle respectively. Corresponding conversion efficiencies were 70 and 83%. In the period from 24 to 36 hours the rate of acid production decreased so that the H2SO4 concentrations at the end of 36 hours were 180 gpl for the barren and 260 gpl for the recycle; average conversions for the initial 36 hours were 55 and 68% respectively. Many investigators1'2'3 have considered the oxidation of sulfur dioxide to form sulfuric acid in aqueous solutions containing metallic ions. A summary of the work performed on this subject prior to 1946 has been presented by Keyes4 who described the auto-oxidation process as follows. Keyes also suggested that in this second reaction ferric ion serves as a catalyst for the conversion of sulfur dioxide to sulfuric acid. Neytzell-de Wilde and Taverner5 have reported on the use of auto-oxidation to produce a 4 to 5% sulfuric acid solution for use in a cyclic leach process at South African uranium plants. This paper describes investigations carried out to determine the feasibility of using the auto-oxidation process to produce sulfuric acid solutions of as high concentration as possible using commercially produced uranium plant barren and recycle liquors from the Elliot Lake region of Ontario. These liquors contained 1.66 and 4.56 gpl respectively of iron, chiefly in the ferrous state. This investigation is of particular interest to the uranium producers in Elliot Lake where the ores contain 3 to 4% sulfur, as sulfides. Flotation of these sulfides and roasting of same to produce SO, bearing gases have been successfully demonstrated using conventional methods.6'7 Barren and recycle solutions, as described previously, are also produced in large tonnages by these uranium producers. The experimental method and apparatus used to examine the auto-oxidation process will be described first. Then the tests carried out on laboratory prepared aqueous solutions containing ferric sulfate to determine optimum conditions will be presented. The next section will describe the tests using the commercial barren and recycle liquors. A final section will summarize the conclusions of the work. EXPERIMENTAL APPARATUS AND PROCEDURES Two types of solutions were used in the tests, a laboratory prepared aqueous solution containing definite amounts of ferric sulfate, and commercially produced iron bearing solution obtained from a uranium leaching plant in the Elliot Lake area. Two liters of a solution to be examined were placed in a 3 in. diam by 30 in. long glass cylinder with a cross-sectional area of 7 sq in. A mixture of air, and SO2 supplied from a lecture bottle for convenience, was introduced into the bottom of the vessel through a standard 30 mm glass frit, 700 sq mm in area. The temperature of the apparatus was controlled by a heating tape and variable transformer. Exhaust gases were vented to atmosphere. At the end of the test period the gas flow was shut off and the solution was analysed for H2SO4, Fe ++, and Fe +++, by conventional means. Four variables were examined using the laboratory prepared aqueous solution of ferric sulfate. First, the effect of SO2 concentration, from 1 to 1076, in the feed gas mixture was considered. Second, the temper-

Jan 1, 1964

By D. B. Smith, John Chipman

THIS investigation was undertaken as a prerequisite to the study of sulphur activities in the liquid system Fe-Cu-Ni, a continuation of the work of Sherman, Elvander, and Chipman,¹ using the same equipment and the same experimental technique. Temperature measurements are made with a disappearing filament optical pyrometer, sighting through a prism onto the surface of the melt which is heated by induction. To obtain true temperatures by this means, a calibration is necessary, since, in the absence of true black-body conditions, the observed temperature is always lower than the true temperature. This calibration is actually an emissivity correction, since the emissivity is a function of both temperature and melt composition. The original furnace was used with a deeper metal bath to provide sufficient depth of immersion for the thermocouple which was inserted through the sampling hole. The atmosphere was a 50-50 mixture or argon and hydrogen passing through the furnace at a rate of 1000 ml per min, and preheated by means of a molybdenum resistance coil. It was assumed that the rapid gas flow effectively eliminated any error which might have resulted from the presence of metallic vapor. The charge in each case consisted of about 300 g of metal, contained in a high purity, dense alumina crucible. At each melt composition, readings of the optical pyrometer and of a thermocouple immersed in the molten metal were taken simultaneously. Calibration curves of optical vs. true temperature were then made for each melt composition investigated. Thermocouples were made of 0.010 in. platinum and platinum-10 pct rhodium. They were checked against the melting point of chemically pure copper. Silica protection tubes with an outer diameter of approximately 1/8 in. were used. The depth of immersion of the tip in the molten metal was approximately 1 14 in. The life of these small thermocouples was relatively short, varying from approximately 1 min at 1550°C to 10 min or longer at 1300°C. Melts containing iron corroded the protection tubes more than those without iron present. Measurements taken while heating checked those taken while cooling, indicating the absence of a temperature gradient between the molten metal and the thermocouple. Results The temperature range of the. measurements on copper-free metals extended from the melting point to 1600°C. For copper the upper limit was 1380°C, and the data were extrapolated linearly for comparison with observations on the other metals and alloys at 1535°C. Examples of typical calibration curves for various melt compositions are shown in Fig. 1. These curves, besides providing a calibration system for the given experimental conditions, also give a basis for the calculation of true emissivities at the effective wavelength of the pyrometer, 0.65 micron. The starting point of the calculations is the value of the emissivity of iron at its melting point. Taking the melting point of electrolytic iron under hydrogen as 1535°C, and using the corresponding emissivity value of 0.43, as determined by Dastur and Gokcen,² the emissivities can be determined as functions of composition (and temperature, if desired). The relationship used is the Wien equation: ln(Ea)=C2/?(1/T1-1/Ta where E is the emissivity; a, the transmissivity of the optical system; C,, the fundamental constant, 14,380 micron-degrees; A, the wave length of light used (0.65 micron); Tt, the true temperature, OK; and T., the apparent temperature, OK. At the melting point of iron the observed temperature was 1358"; this gives 0.62 for the transmissivity, which compares with 0.65 in the similar system used by Dastur and Gokcen. The emissivities of all the compositions investigated were determined at 1535°C, after obtaining from each calibration curve the corresponding temperature. These results

Jan 1, 1953