Search Documents

By J. T. Berling, J. B. Conway

MANY studies have been devoted to the correlation of low-cycle fatigue behavior. Several of these have also been concerned with the identification of an effective approach to the prediction of low-cycle fatigue behavior from short-term tensile data. Perhaps the first such approach was that reported by coffin1 in the form: where is the plastic strain range, Nf is the number of cycles to failure, and c is a constant characteristic of a material. In this analysis by Coffin it was reasoned that in a short-term tensile test the value of Nf could be considered as equal to a , in which case the value of c is interpreted as the tensile ductility, f, measured in such a test. If, therefore, the Coffin equation is assumed to be applicable to a given material, it is a simple matter to predict the low-cycle fatigue behavior. Once the tensile ductility is known, this point is located at Nf = 1/4 on a log-log plot of vs Nf and a line is drawn through this point with a slope of to yield a prediction for the fatigue life. Actually, while this equation is quite useful it is not generally applicable, for experiment has shown that the value of —$ does not reflect the relationship between and Nf for all materials or even for a given material over all temperature ranges. An equation similar to the Coffin equation was prepared almost simultaneously by Manson.2 This expression had the form: In this case the exponent on Nf was not assigned any particular numerical value in accord with the various values of this exponent which had been observed for different materials and at different test conditions. Because of this characteristic it will be recognized that little use can be made of this expression in predicting low-cycle fatigue results. A fairly effective expression for use in predicting fatigue behavior was proposed by Manson3 based on a study of numerous sets of experimental results. This relation (Method of Universal Slopes) makes use of modulus of elasticity, ultimate strength, and tensile ductility, and assumes an exponent on Nf equal to 0.6 in representing the plastic strain range contribution to total strain range. While this equation of Man-Son does not yield completely accurate results it must be recognized as being the best means available at

Jan 1, 1970

By J. R. Campbell

In order to discuss the subject intelligently, it will be necessary to touch briefly on the forms in which sulphur is supposed to exist in coking coal to be carbonized in beehive or byproduct ovens. Sulphur is known to exist in coal as sulphides and sulphates, as can be determined experimentally. Then there is the so-called organic sulphur, i.e., sulphur in combination with the carbon, 'hydrogen, and oxygen of the coal, about which much has been written, but nothing definitely proven in an experimental way. In fact, so far. as we know, there never has been developed a satisfactory and conclusive method in the laboratory for the direct determination of organic sulphur. For most coking coals, it can be safely assumed that the preponderance of sulphur is in the form of pyrite (FeSz). When exposed to comparatively low temperatures during the coking process, it loses its sulphur according to the following chemical reaction: 7FeS2 = Fe7Ss + 6S. It will be seen, therefore, that six out of the 14 atoms of sulphur (or 42.8 per cent.) are expelled, or volatilized. FelSa remains in the coke as pyrrhotite, or magnetic sulphide of iron; this accounts for the highly magnetic properties of the powdered coke. A more commonly accepted reaction is FeS2 = FeS f S, in which the straight sulphide of iron is produced with 50 per cent. volatilization of sulphur. In the beehive methods of coke making, air is introduced in sufficient amounts to carry on the distilling and the coking processes, and the sulphur is oxidized along with the other volatile products: First, into sulphur dioxide (SO2), known by the pungent andsuffocating odoremitted from the trunnel head; second, into sulphuric anhydride (SOa); and, finally, into sulphuric acid, as it comes into contact with the air and moisture. In a properly regulated draft on a beehive oven, there is never complete combustion of the gases. In other words, the coking process should be carried on in a reducing atmosphere, and a low-grade producer gas kept issuing from the trunnel head. For coals of about 30 per cent. volatile matter, the ratio of air to gas is 3½ to 1. In complete combustion, the ratio is 6 to 1, producing an extremely high temperature

Jan 1, 1916

By Lewis A. Riley

I desire to submit, for the consideration and information of the members of the Institute, the following data, drawings, and tables, showing what I believe will be interesting information with regard to the diamond drill, and its uses in developing mineral lands. I would first say that the results given have been obtained during the past two years by means of two drilling machines belonging to the Lehigh Valley Coal Company, operating on their coal lands in the Mahanoy, Lehigh, and Wyoming regions. 1st. A tabulated statement, giving location of borings, date, progress, total cost, itemized cost per foot, etc. 2d. Section showing strata bored in Nos. 5 and 10 holes at Mt. Carmel, see Plate VI. 3d. Section showing strata bored in No. 19 hole at Hazleton, see Plate VI. 4th. Photograph and description of drill No. 2. From the first it will be seen that the total number of test-holes bored up to May 1st of this year is 24; 9 of these are in the Mahanoy region, 6 in the Lehigh region, and 10 in the Wyoming region, in all of which cores have been taken for the whole distance. The total distance bored is 9902 feet. The majority of the holes have been put down for the purpose of proving the lower veins of the coal series, and have had to encounter the harder rocks of the coal formation. Much of the distance has been through the lower conglomerates, going in some cases through the coarse egg conglomerate, the foundation of the coal deposit, and to the green sandstone and red shale which underlie it. These lower sandrocks and conglomerates are among the hardest known to geologists, and present a formidable resistance to the exploring drill. This fact must be taken into consideratiotl in comparisons of cost of drilling, and when considered will still further improve the very favorable results I am able to show in my statement of costs. In illustration, I will cite the work of one bit when boring in the upper or softer rocks of the Wyoming region, where 280 feet were obtained without resetting, while in the Lehigh and Mahanoy regions, in the lower and harder rocks, bits set with the same quality of diamonds would only bore from 3 to 20 feet before being worn out. In this connection it may

By D. Mizer, J. B. Clark

In a recent investigation of the entire magnesium-yttrium phase diagram, Gibson and Carlsonl report the maximum solid solubility of yttrium in magnesium as 8.0 wt pct Y at 1050°F (565°C). However, independent studies of the precipitation in magnesium-yttrium alloys, 2 conducted prior to the publication of the above phase diagram, indicated a greater solid solubility of yttrium in magnesium. Accordingly, the magnesium-rich region of the diagram was re-investigated to determine more accurately the liqui-dus, solidus, and solvus of the magnesium solid solution in this system. Metallography and thermal analysis were chosen as the best corroboratory methods for accurately determining this region of the diagram. A magnesium-yttrium master alloy was prepared from yttrium of 99.9 pct purity and singly sublimed magnesium of 99.9 pct purity. Subsequently, the twelve magnesium-yttrium alloys shown in Fig. 1 were prepared using standard alloying procedures for magnesium. Metallographic samples of the alloys were heat treated under an SO, atmosphere in an electric furnace capable of control within ±3oF of the desired temperature. Alloy samples for the metallographic determination of the solvus, and for the precipitation studies were solution treated at 1000° F for 16 hr. Subsequent aging times for the solvus determination varied from a minimum of 6 hr for the higher temperatures up to 816 hr for lower temperatures. Periods of 30 min at temperature were used to generate liquation for the metallographic placement of the solidus and eutectic horizontal. The liquidus and the eutectic horizontal were determined by thermal analysis. Simple time-temper- ature heating and cooling curves were obtained from 10-g alloy samples heated and cooled in a well-insulated electric furnace. An Iron-Constantan thermocouple was placed in direct contact with the sample. The thermocouple did not contaminate the alloy sample since metallographic examination showed that the alloy melt did not wet or attack the thermocouple. The heating and cooling rates were approximately 5 to 8OF per min. The standard metallographic procedures for magnesium were used in polishing and etching the magnesium-yttrium alloys studied. Glycol etchant* was etchant was used to reveal precipitate in the solvus determination. The redetermined magnesium-rich region of the magnesium-yttrium diagram is shown in Fig. 1. The

Jan 1, 1962

By S. H. Davis

The pyrites used in making sulfuric acid in the United States have been largely imported from Spain and Canada, the Spanish imports amounting to nearly 1,000,000 tons per annum in the pre-war period. The greatly increased use of sulfuric acid and the cutting off of these Spanish imports, incident to war conditions, brought about a threatened shortage of sulfur supplies during the war period. The bituminous-coal mines of ccrtain districts have, for many years, furnished a small tonnage of pyrite in the form of coal brasses. A mechanical concentrator at Danville, Ill., for a number of years, has been treating hand-picked lump pyrite and coal from the picking belt and from the mines and a small plant near Gillespie, Ill., for a few months has been recovering pyrite from washery refuse. Many mines throughout Illinois, Indiana, Western Kentucky, Ohio, and Pennsylvania have shipped an occasional car of the hand-cleaned lump pyrite. Only a very small percentage of the available pyrite has been recovered in this way, however, as usually the miners throw such lumps into the gob with slate and other impurities. It has been estimated that the western Indiana coal field could furnish more than 100,000 tons of pyrite per annum. The present production is very small. The possibility of furnishing the domestic trade with pyrites recovered as a byproduct from coal-mining operations appears attractive but there are certain features difficult to overcome. Pyrite, to be used in acid making, must meet with certain requirements as to size and purity. Lump ore for grate burners should be under 3 in. (7.6 cm.) and over 1 in. (2.5 cm.) in diameter. Fines for use in mechanical roasters should be under quarter mesh. The material should be high in sulfur, free from arsenic and phosphorus, and as low in carbon as possible. The pyrite obtained from coal can be made to meet all the above requirements, but it is difficult to remove all the carbon. The pyrite in coal occurs as bands and nodules of varying thickness and size and of comparative purity, but mixed with this is more or less web sulfur. The web sulfur carries admixed coal, which may make the concentrate run up to several per cent. carbon. This makes the concentrates subject to firing, causes heavy consumption of niter, and lowers the acid-plant capacity, due to dilution of gases. The hand-cleaning methods

Jan 1, 1920

By Dorsey Hager

NORTH -CENTRAL Texas has recently become a center of interest for the oil men of America. The bringing in of the McClosky well at Ranger, Eastland County, and the shallow pool at Brownwood, Brown County, in 1917, has stimulated interest in this area to fever pitch. Oil men from all over the United States are now investing there. The area of present interest is shown by the accompanying map (Fig. 1). The money spent in leases runs into millions of dollars. It is no exaggeration to say that a strip of country 200 miles (321 km.) long and 125 miles (201 km.) wide, comprising some 15,000,000 acres (6,070,310 ha.)., has been leased practically solid at a cost for rentals and bonuses of at least $1 per acre. The test holes contracted for will certainly number 400, at a cost of at, east $5,000,000. From what has been done in the past six months, $20,000,000 at least will be spent. To pay, returns on this amount of money, new production to the extent of at least 12,000,000 bbl. must be obtained. At present, the production from new fields will not average over 5000 bbl. per (lay; three wells at Ranger are producing 3000 bbl.; 250 wells at Brownwood produce 1000 bbl.; and the Gray well, Coleman County, is as yet an unknown factor. However, at Ranger there is every indication of developing a good pool covering from 1500 to 2000 acres (607 to 809 ha;), more or less, but the wells are deep, 3400 to 3800 ft. (1036 to 1158 m.), and cost $35,000 to $40,000 to drill. Large wells are necessary to pay for such expensive holes. As new wells are drilled, the gas pressure will be lowered rapidly, and large production need not be expected. For those oil men who expect a second Cushing or an Eldorado, Ranger holds little of promise. At Brownwood, Brown County, there are some 250 shallow wells (depths from. 200 to 350 ft.) averaging 4 to 7 bbl. per day. There is a chance of an extensive producing area, for these shallow sands to the southwest, and the opening of several thousand acres of shallow oil territory, and also some promise of deeper oil horizons in the Ranger horizon, but probably all under 2500 ft.

Jan 6, 1918

By Sheldon Rideout, E. L. Kamen, B. S. Lement, Paul A. Beck, W. D. Manly

The 1200°C isothermal sections of the following ternary phase diagrams were investigated: Cr-Co-Nil Cr-Co-Fe, Cr-Co-Mot and Cr-Ni-Mo. In all these systems the u phase was found to form extended solid solutions. Two new ternary phases of unknown structure were identified. The existence range of intermediate phases of transition elements was correlated with 3d electron vacancy concentrations. THE occurrence of extensive ternary d solid solutions in the Cr-Co-Ni and in the Cr-Co-Fe systems was described in a brief note.' More recently d- solid solutions were also found in the Cr-Fe-Mo system by two independent groups of investigators.2,3 The existence of binary d phases in the V-Co and V-Ni systems, predicted in ref. 1, was experimentally discovered recently by Duwez and Baen and independently by W. B. Pearson, J. W. Christian, and W. Hume-Rothery.5 Duwez and Baen stated two general conditions for the occurrence of the d phase in alloys of the transition elements: "(1) The atomic diameters of the two metals do not differ by more than 8 per cent. (2) One of the two metals has a body-centered cubic structure while the other is face-centered cubic, at least in one of its allotropic forms." In the course of an extensive research program conducted at this Laboratory under the sponsorship of the NACA, several ternary systems, containing d solid solutions and also certain new ternary phases, were investigated. The present report describes the investigated portions of the 1200°C isothermal sections of the following ternary phase diagrams: Cr- Co-Ni, Cr-Co-Fe, Cr-Co-Mo, and Cr-Ni-Mo. It was found that the range of existence of the d phase in these four ternary systems, as well as in one of two ternary systems which were investigated by others, can be fairly consistently described in terms of electron vacancy concentrations. This supports the suggestion of Sully and Heal" that the d phase is an electron compound. Experimental Procedure The ternary alloys were prepared by vacuum induction melting in zirconia and alumina crucibles. Lot analyses of the metals used are given in Table I. Specimens of all alloys were annealed at 1200°C in an atmosphere of purified 92 pct helium and 8 pct hydrogen mixture. Alloys consisting mainly of

Jan 1, 1952

Discussion of the paper of HOWLAND BANCROFT, presented at the Salt Lake meeting, August, 1914, and printed in Bulletin No. 91, July, 1914, pp. 1767 to 1769. THEODORE SIMONS, Butte, Mont. (communication to the Secretary*). -While preparing maps and models for use in mining litigations the writer was frequently confronted with the problem referred to in Mr. Bancroft's paper. His chart gives the apparent angle of dip with which a vein should be plotted on a vertical section not at right angles to the strike of the vein. For a quick determination of this apparent angle the writer has used an exceedingly simple graphical method, which he believes may prove helpful to engineers engaged in practical work. It is submitted, not as a criticism of Mr. Bancroft's paper, but as a means of solving the same problem when no chart is available and when calculations would consume too much time. The writer's method requires merely a few lines marked off on a piece of tracing cloth, as shown in Fig. 1: Draw line be on any convenient scale. At b erect a perpendicular, ba. By means of a protractor cut off on ba a number of angles within the probable range of variation of dip, so that a line connecting a and c, for instance, forms an angle of 60° with be, etc. With b as a center and be as a radius, draw a circle through c. This completes the tool for solving the problem. Fig. 2 represents the line of strike of the vein and db the trace of the vertical plane on which the vein is to be plotted with the apparent angle of dip. A° is the angle which this plane makes with the strike. Fig. 3 shows the application of the method: Place line cb of tracing over line db of Fig. 2 and move fore and back until line of strike becomes tangent to the circle. Connect d with the point (a) that marks the true angle of dip as found by measurement in the mine (60° in the case illustrated). The angle adb is the apparent angle of dip (C°) to be used in the vertical section. Proof.-As stated in Mr. Bancroft's paper, the apparent angle of dip is: tan C° = sin A° tan B° (1)

Jan 12, 1914

By R. A. Pallanch

Sodium sulfite as a zinc retarding agent in the selective Hotation of complex lead-zinc sulfide ores was discovered by the writer in the summer of 1919 in the testing laboratory of the U. S. Smelting, Refining & Mining Co.'s gravity concentrator at Midvalc. Utah. It proved to be the first reagent to give consistently good results in the laboratory on the cornpanj's Bingham ore. In December, 1919, C. A Lemke, concentrator superintendent, recommended that a flotation unit be added to the existing gravity mill to handle the mill slimes, using sodium sulfite as the zinc depressant Owing to the poor lead and zinc markets and the unfavorable royalty schedules offered by the Minerals Separation Co., the project fell through at this time. In 1921, a number of laboratory tests were made at Midvale. on other ores, among them ore from the Sunnyside mine at Eureka, Chlo. Both sodium and calcium sulfites were used with succcss on this ore. Early in 1923, further tests were made at Eureka, shortly after the Sunnyside company had resumed operations after a two-year shutdown. AS a consequence sodium sulfite was given a trial in the mill. The results showcd sufficient improvement over previous work as to warrant continuous use of this reagent in the lead circuit Exeept for a week or so in 1926 and for several weeks in April, 1927, when sodium cyanide and zinc sulfate were given trial runs in this mill. sodium sulfite has been used ever sinee.. Sodium sulfite was chosen rather than the calcim sulfite brcauscl of its gre1ater sollubility, making it possitdr. to introduee the reagent as a solutiun. This Solution is introdueed into the Marry ball mills along with the original feed of ore,. The consumption of sodium sulfite in the Sunnyside mill varies from 1/4 to 12 lb per ton of ore, the smaller amount probably being nearer the average. In February, 1924, laboratory work at Midvale demonstratcd conclusivcly that good results could be obtaincd on the company's Bingham ore and on most of the custom ores tested by using sodiuin sulfite as the zinc depressant, xanthate as the lead enhancer in the lcad circuit, followed by copper sulfate and additional xanthate to float the zinc Sodium sulfide was used to float the iron pyrite after the lead and

Jan 1, 1928

By J. R. Campbell

IN order to discuss the subject intelligently, it will be necessary to touch briefly on the forms in which sulphur is supposed to exist in coking coal to be carbonized in beehive or byproduct ovens. Sulphur is known to exist in coal as sulphides and sulphates, as can be determined experimentally. Then there is the so-called organic sulphur, i.e., sulphur in combination with the carbon, hydrogen, and oxygen of the coal, about which much has been written, but nothing definitely proven in an experimental way. In fact, so far as we know, there never has been developed a satisfactory and conclusive method in the laboratory for the direct determination of organic sulphur. For most coking coals, it can be safely assumed that the preponderance of sulphur is in the form of pyrite (FeS2). When exposed to comparatively low temperatures during the coking process, it loses its sulphur according to the following chemical reaction: 7FeS2 = Fe7S8 + 6S. It will be seen, therefore, that six out of the 14 atoms of sulphur (or 42.8 per cent.) are expelled, or volatilized. Fe7S8 remains in the coke as pyrrhotite, or magnetic sulphide of iron; this accounts for the highly magnetic properties of the powdered coke. A more commonly accepted reaction is FeS2 = FeS + S, in which the straight sulphide of iron is produced with 50 per cent. volatilization of sulphur. In the beehive methods of coke making, air is introduced in sufficient amounts to carry on the distilling and the coking processes, and the sulphur is oxidized along with the other volatile products: First, into sulphur dioxide (SO2), known by the pungent and suffocating odoremitted from the trunnel head; second, into sulphuric anhydride (SO3); and, finally, into sulphuric acid, as it comes into contact with the air and moisture. In a properly regulated draft on a beehive oven, there is never complete combustion of the gases. In other words, the coking process should be carried on in a reducing atmosphere, and a low-grade producer gas kept issuing from the trunnel head. For coals of about 30 per cent. volatile matter, the ratio of air to gas is 3 ½ to 1. In complete combustion, the ratio is 6 to 1, producing an extremely high temperature

Jan 1, 1916

By F. E. Bash

PREVIOUS to the war, this country depended on Europe for its supply of a number of alloys of great importance in the manufacture of electrical apparatus and equipment. When this source was cut off shortly after the beginning of hostilities and the small stocks in the United States were exhausted, either inferior substitutes had to be used or the manufacture of the apparatus suspended. Since electrical equipment was of primary importance in the prosecution of the war, the substitutes had to be used until the composition and processes of manufacture of the alloys could be determined. Included among these alloys was one generally known as manganin, which is composed chiefly of copper and manganese and which, beside having a low-temperature coefficient of resistance, has a low thermal electromotive force against copper. These qualities make it valuable for the manufacture of electrical instruments and machines in which it is used in the shape of wire, rod, and strip. As a result of these circumstances, an investigation was started to deter mine how manganin could be made on a commercial basis so that a product of uniform quality and the necessary electrical characteristics could be produced. The object was to procure an alloy with temperature coefficient of resistance not greater than T 1 X 10-5 per ° C. and a thermal e.m.f. against copper no greater than that of the imported material. It is characteristic of manganin that the temperature-resistance curve is not a straight line but has a peak, which varies in position with different samples; it is most desirable, however, to have it at a temperature of 25° C. if possible. The coefficients of manganin from all records available ranged from a large negative to a large positive value. The composition as reported by different authorities is given in Table 1. Prof. Matthew A. Hunter, of Rensselaer Polytechnic Institute, made up a number of small melts of manganin in a graphite crucible and the one that he reported as having the best coefficient gave the following analysis: Copper, 83.6 per cent.; nickel 2.5 per cent.; manganese,-13.6 per cent.; iron, 0.3 per cent. An analysis of a sample of German manganin gave the following: Copper, 82.6 per cent.; nickel, 3.8 per cent.; manganese, 12.8 per cent.; iron 0.7 per cent.

Jan 9, 1919

By Edward Keller

ALTHOUGH the subject matter contained in this paper is presented under a new title, its writing was induced by the work of S. Skowronski and K. W. McComas.1 A discussion of the latter is here combined with the presentation of new and independent data. The physical constants quoted by the authors of that paper, such as melting and boiling points, as also the chemical reactions of the substances involved, are so well established that the numerous confirmatory experiments and their application to metallurgical conditions leave little or no room for criticisms. On the contrary, they will greatly aid to dispel any lingering skepticism as to possible considerable losses of copper accruing when the metal is melted in the presence of its own and, probably, other chlorides and when no provision is made for the recovery of volatile products. Exception may be taken, to some degree, to that part of the work under the caption Melting Drillings of Copper Mixed With Sodium Chloride. These experiments should have been extended in order to make the implied reactions clear and the results convincing. The latter are stated to be, "that when sodium chloride is present above certain minimum quantity relative to copper some of it is found to volatilize in the melt while another part reacts with the metallic copper to produce the volatile cuprous chloride, and that When smaller quantities of sodium chloride are present the reaction is complete and all chlorine is found to have volatilized as cuprous chloride." Metallic copper and sodium chloride are mentioned as the sole reagents in this chemical process, acting, according to the authors, in a neutral atmosphere of carbon dioxide. The chemist, quite naturally, will inquire into the formula according to which that reaction takes place and find that it would present itself as follows: 2Cu + 2NaC1 = 2Na + Cu2Cl2 It would appear that the truth of this formula will be challenged and a more rational explanation of the reaction sought. Apparently, carbon dioxide is neutral relative to metallic copper in its molten state; whether it remains neutral in the presence of sodium chloride is quite another question. Does it not at that high temperature become an oxidizer and do not the conditions of the reaction change to those expressed in the following formula? 2Cu + 2NaCl + 2002 = Na2CO3 + CO + Cu2Cl2

Jan 9, 1919

By Thomas Clagett

THE Pocahontas coal field comprises the area in Tazewell County, Va., and Mercer and McDowell Counties, W. Va., in which Nos. 3 and 4 Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305m.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now, about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, io which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks off bone, each from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 25 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficiently well defined to exert an influence on the mining is rare. Overlying strata are slates, shales, and sandstones, the sandstones predominating, extending from .30 to 1000 ft. above No. 3 seam. The strata directly above the. draw slate of No. 3 seam, or "black rash" of No., 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the

Jan 5, 1922

By C. Ernest Birchenall, Gerd M. Rosenblatt

The pressure and molecular weight of the antimony vapor over solid solutions of tin in antimony, containing 97.4, 94.9, 93.0, and 90.7 at. pct Sb, have been measured from 445° to 545°C by the torsion-effision method. The vapor is all Sb, within experimental error. A solidus curve which differs somewhat from the accepted curve has been calculated using the measured antimony activities. Metallographic examination of additional annealed alloys substantiates the accepted solidus. The solid solutions, which show negative deviations from ideality, probably do not deviate greatly from regular solution behavior. HANSEN' considers the Sb-Sn phase diagram to be well established with the exception of the boundary of the (Sb) phase.* Although there have been no pre- Their data can be combined with data for the solid alloys to calculate the solidus curve of the (Sb) phase. This paper presents measurements of the pressure ant1 molecular weight of antimony over Sb-Sn solid solutions containing 97.4, 94.9, 93.0, and 90.7 at. pct sb. Measurements were made by simultaneous use of the volmer3 torsion and Knudsen, effusion methods. The results of these measurements have been combined with the vapor pressure of pure antimony5 to obtain the activity of antimony in the primary solid solutions. The solidus curve calculated from these data differs somewhat from the accepted curve and is rather uncertain. Metallographic examination of Sb-Sn alloys was carried out to clarify the position of the solidus of the (Sb) phase. EXPERIMENTAL The apparatus and technique have been described previously.5 The measurements were made with the same effusion cells and torsion wire used in the measurements on pure antimony. The torsion measurements were calibrated with mercury at 30" using a mercury vapor pressure of 2.96 x 10-3 mm Hg determined by concurrent, absolute, weight-loss experiments. As in the measurements on pure antimony, an extraneous vapor was observed to distill off during the first run after the cell had been loaded with A fresh sample of alloy and during the first minute or two of every run. The change in composition after a series of runs with one sample was of the order of 0.1 at. pct and thus completely negligible in the calculation of the thermodynamic properties. The alloys were prepared from pieces of Merck antimony and Matheson, Coleman, and Bell reagent-

Jan 1, 1962

By Valentin R. Garfias

The world's production of petroleum during 1926 is estimated at 1,096,000,000 bbl., an increase of about 29,000,000 from that of 1925 and somewhat larger than the average during the last 4 years, which is about 1,049,000,000 bbl. The daily average during the year may be estimated at 3,000,000 bbl. of which the United States produced 2,123,000 and other countries 877,000. The outstanding developments during the year have been the decline of the Mexican fields, the counterbalancing of this decline by the increase in Venezuelan and Russian production, and increases in Roumania and Colombia. During 1926 the United States, Mexico and Russia retained their relative positions as the three leading producers, while Venezuela probably outranked Persia for fourth place, Roumania passed Dutch East Indies for sixth, and Colombia moved from sixteenth to the eleventh position. It further appears that should present conditions develop normally throughout 1927, it is not improbable that Mexico will be outranked by Russia as the second oil-producing country and that the increased yield of Colombia will place this country in the eighth or ninth position, and ahead of India. The United States and Mexico should produce in 1926 about 79 per cent. of the world's total and the aggregate of the seven leading countries should be close to 95 per cent. of the total. It is interesting to note that about 85 per cent. of the total comes from fields on the American continent, the United States accounting for about 71 per cent., and Spanish-American countries and Trinidad for 13 per cent. Production of countries other than the United States and Mexico has gradually increased during the last four years and now represents about 21 per cent. of the total. The United States easily retains its long standing leadership with a total yield of 775,000,000 bbl., or an average daily yield of 2,123,000, representing about 71 per cent. of the total, which is about equal to that of the previous year. An increase in production of about 11,000,000 bbl. is indicated by the statistics so far available. It will be noted that the American production in the last 4 years has averaged about 746,000,000 bbl. yearly and that this uniformity, as well as the marked increases

Jan 1, 1927

Discussion of the paper of DORSEY HAGER, to be presented at the Colorado meeting, September, 1918, and printed in Bulletin No. 138, June, 191S, pp. 1109 to 1118. WALLACE E. PRATT, Wichita Falls, Tex. (written discussion*).¬Mr. Hager has touched upon interesting problems related to the broad subject with which he has entitled his paper. As to the existence of the important structural feature in the buried Bend series, which Mr. Hager constructs from well-logs, and designates as the "Bend Arch," there seems to be general agreement among a number of observers. Indeed, M. G. Cheney1 and Robt. T. Hill2 have each published structural maps of the Bend Arch that do not differ essentially from the map shown in Mr. Hager's Fig. 1. The reflection, in the overlying, unconformable Cretaceous, of certain folds in the Pennsylvanian, with which the petroleum accumulations has to do, is an important observation on the part of Mr. Hager, since, as he says, much critical territory is covered by Cretaceous, and it is highly desirable that the underlying Pennsylvanian structure be mapped, if possible. If the folds in the Pennsylvanian are of post-Cretaceous age, then it should be possible, in spite of the intervening unconformity, to determine approximately the position of buried folds beneath the Cretaceous overlap. The idea of post-Cretaceous folding would not be out of accord with the suggestion by Carroll H. Wegemann that the folding of the Permian beds in southern Oklahoma was still in progress as late as the Tertiary.3 But most observers have related the folds in the Pennsylvanian and Permian beds at the surface, both in North Texas and southern Oklahoma, to the process of. emergence of the land area at the end of the Permian. It seems more probable that the reflection of Pennsylvanian folds in the unconformable Cretaceous which Mr. Hager observed is due to long continued persistence of stresses which were largely relieved at the close of the Permian, than that the principal period of folding was post-Cre-taceous; this being the case, it seems likely that only folds of greatest intensity in the Pennsylvanian will, be decipherable in the Cretaceous overburden, and that many promising structural features are so concealed by the Cretaceous as to escape detection except by drilling.

Jan 8, 1918

By Graham John Mitchell

In May, 1919, a crosscut on the 1400-ft. level of the Briggs mine, a Calumet, & Arizona property at Bisbee, Ariz., penetrated a deposit of pyrite and chaleopyrite that had replaced quartzite and limestone and was cut, especially in the quartzite, by numerous fractures. As the drift advanced through these fractured portions, they were found to be free from copper-sulfate minerals although '(sooty" chalcocite was present in the fractures and also as coatings and replacements of pyritc. Howcver, sulfate soon made its appearance in certain fractures, filled them, and began to form stalactites. Several centers of incipient growth were noted for subsequent study, observations in the drift at frequent intervals during the past 17 months affording the opportunity of determining the rate of formation of the stalactites. The one to attain the greatest growth during the period of observation measured 27 in. long and averaged 1 in. in diameter. A close study showed it to be composed of small interlocked crystals of chalcanthite radiating from a small central tube-like opening. The trielinic form of the crystals was evident with prism and pinacoid faces recognixable. Assuming that the supply of copper-sulfate solution was constant (luring the 17 months—-and such an assumption is supported by the fact that the number of drops of excess solution falling, per minute, from the tip of the stalactite at the times of observation checked closely—thc stalactite, it has been estimated, grew at the rate of 1.58+ in. per month. DISCUSSION James I?. Kemp," New Yolk City.—J. S. Curtiss, in the monogr:tpll on Eureka, Ncv., say:: that a stalactite formed in limestonc grew 5/8 in. in three weeks and continued to grow as long as it had a drop of water or solution at the lower end. Another grew at the rate of 3/8 in. in the same time. In the case cited by Mr. Mitchell, the growth was about 136 in. in three weeks. A year ago we had an instance of the formation of cubes of galena on a railroad spike. A mine in northeastern Oklahoma had been idle about two years but on account of pumping of neighboring properties, the circulation of water was active throughout the drift. A cube of galena,

Jan 1, 1922

By P. W. Davies, B. Wilshire

PREVIOUS investigations on the effect of stress changes on the high-temperature creep and fracture behavior of metals have been confined mainly to the testing of complex alloys.172 Most of these alloys are structurally unstable during creep because of such effects as recrystallization, overaging, and so forth, so that the results are of limited application. The only pure metals on which observations have been made appear to be lead3,4 and aluminum5 tested under conditions which allow rapid recovery to take place. The present work describes the effect of a single stress change on the creep and fracture properties of a metal in which recovery is very slow. A Ni 0.1 at. pct Pd alloy was used so that grain growth and recrystallization did not occur during creep testing. Specimens of this material were heat-treated to give a uniform grain size of 10 grains per mm. Creep tests were carried out at 500°C using Denison T.47 machines; details of the creep-testing procedure have been given elsewhere.' All specimens were tested at a constant initial stress of 5 tpsi which gives, in an uninterrupted test, a minimum creep rate of ~10-3 pct per rnin and a life of -7000 min. Specimens were loaded under these standard conditions and after 1500 min, when secondary creep was well-established, the stress was changed to either 8, 7, 6, 4, 2, or 0 tpsi for a period of 180 min. After this time the original load was reapplied and the tests continued to fracture. The results are shown in Fig. 1. No detectable creep occurred during the period of reduced load, even at 4 tpsi, and on reverting to 5 tpsi the creep rate returned to its original value immediately. Increasing the stress to 8, 7, or 6 tpsi resulted in an instantaneous extension followed by a period of transient creep leading to a new steady creep rate. On returning from these higher stresses, the resulting creep rate at 5 tpsi was very low compared with the original value. Furthermore, the higher the stress during the stress change, the lower was the final creep rate. It is significant that the rupture life is comparatively unaffected by any of these stress changes.

Jan 1, 1965

By W. H. Smith

IN the course of an investigation on chromium containing minor amounts of alloying elements, the information herein has been obtained on the crys-tallographic orientations resulting from arc melting, rolling, and annealing. The chromium used in this investigation was electrolytic flake obtained from the Electro Metallurgical Co. This material was heated for 20 hr at 1600°C in hydrogen with a dew point of about — 100°C. The oxygen and nitrogen content of the chromium after this treatment was 0.0003 pct 0 and 0.005 pct N. Five pounds of this chromium were arc melted in a water-cooled copper crucible. After arc melting, the oxygen and nitrogen contents were 0.002 and 0.025 pet, respectively. A section of this ingot is shown in Fig. 1. By means of a back-reflection Laue technique, the crystal habit of the cast ingot was determined. As expected, it was found that the [loo] was approximately the axis of the columnar grains. In this respect, chromium behaves like other body-centered-cubic metals in that a cubic axis is normal to the cold surface of the mold. A slab cut from this ingot was rolled at 600°C as follows: four 30 pct reductions were made with intermediate anneals at 1200°C for 15 min. Rolling was done with the [loo] of the columnar grains in the rolling direction. Following the final anneal, an 80 pct reduction was taken at 600 °C. The final sheet thickness was ten mils. By polishing and etching, the thickness was reduced to four mils and the grain orientation determined by the transmission Laue method using molybdenum radiation with a zirconium filter. Fig. 2 is the (110) pole figure for this material with the ideal (100) [011] and (111) [112] orientations indicated. The orientation is typical of most cold rolled body-centered-cubic metals having the [llO] of the grains approximately in the rolling direction and the (100) planes approximately in the plane of the rolled sheet. There is also evidence of some 10) [1 orientation. The grain orientation of sheet rolled 40 and 80 pct in the same direction as before at 600°C followed by recrystallization at 12000 C for 15 min was determined by a back-reflection Laue technique. A random texture was found in both cases. This 1200°C 15 min anneal was found to give complete recrystallization and some grain growth.

Jan 1, 1956

Cortez Gold Mines operated a conventional 2,100 mtpd (2, 300 stpd) cyanidation plant until the 5 million ton ore body was worked out in 1973. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME 1975.) Sub-grade material was dump leached at the mine -mill location in 1971, and the resulting pregnant solutions were added to the feed in the plant's Merrill-Crowe precipitation unit for gold recovery. In early 1973, however, Cortez put into operation their Gold Acres mine which was located about 13 krn (8 miles) to the west of the 'Cortez mine. Economics at that time dictated the heap leaching of the lower grade portion of the ore body with the higher grade being trucked to the Cortez plant for milling. Leach ore was moved directly from the open pits to prepared leach pads located on a gently sloping terrain of 570 grade. Preliminary pad preparation at the leach site included removal of vegetation which was composed of sagebrush and grading of the base. Pads were comprised of mostly minus 74 micron (200 mesh) materials and were made of clay-silt or slime tailings spread on the pad area in thin 50 to 75 mm (2 to 3 in. ) layers, compacted with a vibrating roller compactor to a depth of about 380 mm (15 in.). This was followed with a 100 mrn (4 in.) layer of coarse rock to prevent erosion on the pad and to allow solutions to move laterally under the heaps. Leach solutions normally traveled the length of the heaps and were collected in plastic-lined ditches from which they flowed by gravity to the pregnant solution storage pond. Both pregnant and barren ponds were constructed in the same manner as the leach pads - being lined with compacted tailings or clay silts. The Gold Acres facility required pregnant solution storage of 6 million liters (1.6 million gal) and barren storage of 3.8 million liters (1.0 million gal) designed to handle two to three days of heap drainage or to retain run-off from a major rainfall. Various heap heights were tried including 3, 6, and 9 meters (10, 20, and 30 ft), with a 6 m (20 ft) height selected. Ore was dumped by trucks in a single 6 m (20 ft) lift using a D8 Caterpillar dozer for

Jan 1, 1981