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By Warren E. Wilson

THE transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. ENERGY. REQUIREMENTS Carrying out the idea contained in Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: [Solids load in terms of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L Slope of energy grade line by s Mass per unit volume of solids by m, Mass per unit volume of liquid by ml Mass per unit volume of mixture by m,,, We then have as an expression for the solids load: W = pQm¬] The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1945

By C. J. Beevers

Tlze influence of zirconium hydride precipitate mprphology on the fructure of Zr-H alloys tested at strain rates of 10- sec at 20° and - 196°C and at strain rates of -500 sec.-1 at 20°C has been inves-tigated. The critical stage in the embrittlement of these alloys was the fracture of the zirconium hyhide precipitates to form large interval cracks at a stress of 25,000 10 137,000 psi. In specimens con-temping up to 100 ppm H the preciptates existed as isolafed platelets in the zirconium matrix and their fracture as enhanced by low temperature (—196°C) and high strain rate (-500 set-f) loading. These resl~lts lead to the conclusion that the rate of strain hardening of the zirconium metal controls the frac-tzrm of the hydride platelets. Increased hydrogen contents of 500 to 2000 ppm resulted in an increase in the 1)volume fraction and size of precipitates and crlso led to crack propogation in the hydride at both 20 and —196C at the low strain rates (10'* sec-'). Quenching led to refinerment of hydride precipitate size and reduced the degree of embrittlement at — 196°C; this effect is attributed to the absence of fracture of the fine precipitates. The formation of zirconium hydride precipitates has been shown to be the cause of hydrogen embrittlement of zirconium.' Several investigators'-" have studied the problem in broad outline, demonstrating that impact-loading conditions, testing temperatures below room temperature, and increasing hydrogen content produce increased brittleness. For example, Muehlenkamp and schwope2 showed that the transition temperature for notched impact tests was raised from 150" to 350°C on increasing the hydrogen content from 45 to 490 ppm. The degree of embrittlement is also controlled by the morphology of the zirconium hydride precipitates. The solubility of hydrogen in zirconium decreases from 600 ppm at 550°C to 10 ppm at 20"C,~ so that the form of the zirconium hydride precipitates (ZrH2) can be modified by varying the cooling rate through the Zr-aZr + ZrH2 phase boundary.= Forscher showed that quenching a specimen containing 35 ppm H from above 315°C resulted in a reduction in area of 70 pct at -196"C, whereas slowly cooling from above 315°C resulted in a ductility of half this value. The role of the hydride precipitate in the fracture process has been examined by Young and schwartz6 and more recently by westlake.' From metallo-graphic observations on slowly cooled Zr-H alloys they have suggested that cracks are nucleated in the hydride precipitates by the interactions of twins in the zirconium metal with the hydride. The experiments reported in this paper show that crack propagation in the zirconium hydride precipitates is the critical stage in the embrittlement of Zr-H alloys. The factors influencing fracture of the hydride precipitates and crack propagation in the zirconium metal will be related to the problem of embrittlement. EXPERIMENTAL TECHNIQUES The zirconium used in these experiments was produced from an ingot of vacuum arc-melted iodide zirconium forged down to 1 in. diameter at 800' to 850°C. The 1-in.-diameter bar was surface-machined and cold-swaged to 0.375 in. diameter. After a vacuum anneal at 800°C for 1 hr the hardness was 60 to 70 Vdh. The total impurity content was approxi-

Jan 1, 1965

By C. D. Haynes, S. C. Gambrell

The practice of roof bolting in WORLD-WIDE mining activities is very well-known, having been used extensively for over 50 years. The theory of roof bolting, however, is extremely complicated, and only recently have investigators attempted to explain the theory. Various means of stress analysis have been utilized in rock mechanics, most of which reduce rock to an isotropic, homogeneous medium that can be easily analyzed. One of the indirect research methods is the photo-elastic technique, the theory of which is widely known and needs no elaboration. This technique has been used extensively to examine the state of stress around underground openings of various shapes in certain types of strata, and under various loading configurations.'-" In-Situ Mine Roof Trusses The photoelastic technique was employed here to evaluate the state of stress above a mine opening where support is provided by the in-situ mine roof truss. This support configuration, shown in Fig. 1, has gained popularity in recent years, particularly in mines with bad roof conditions." To the authors' knowledge, no scientific determination has been made of the advantages roof trusses may offer over conventional bolting configurations comparable to roof trusses. Ordinary mine roofs usually sag or bulge downward soon after mining. Normally, after some movement the roof begins to stabilize, but the face of this bulging mine rock is usually in tension and will almost always fail unless supported. Because rock fails easily in tension, it is desirable to "preload" the roof, that is, to impress a compressive force on the mine roof to prevent bulging and subsequent tensile failure. Ordinary rock bolting installations, particularly if the bolting is perpendicular to the mine roof plane, do not produce a preloading effect. Although bolting installations squeeze the strata together and produce a beam effect desirable for support, they do not have the advantage of actual uplift. They pull down at the anchorage point with a force equal to the bolt tension which sometimes causes a separation in the plane of the anchorage detrimental to roof strength. In-situ mine roof trusses require the installation of steel tension rods in the roof of the mine much the same as conventional roof bolts, except that the rods are arranged to produce a stress pattern above the upper surface of the opening similar to that of a prestressed beam. The roof rock acts to transmit the compression force. The simplest form of truss consists of two steel rods anchored to the back of holes drilled at an angle of approximately 45 over the mine pillars. These rods are collared near the middle of the room, and the lower ends of the rods are connected by tie rods and turn-buckles. Used in this manner, a small-diameter steel rod can provide essentially the same support as a wide-flange steel beam. By tightening the turnbuckles, an uplift on the roof can be produced. Objective of the Investigation The objective of this investigation was to determine qualitatively the state of stress that exists above a mine roof after the installation of the in-situ mine roof truss vs. that of two angle roof bolts, hereafter referred to as "roof bolts," drilled and placed in the same configuration, that is, 45" above each pillar and connected together in the middle of the room by a bearing member such as a steel beam. Each state of stress was to be identified, insofar as possible, as compression or tension above the mine opening. Model Construction A photoelastic model of a mine opening was coil-structed from -in. thick PLM-4B birefringent plastic and was machined to specification shown in Fig. 2. The dimensions of the opening were chosen to simulate a width-to-depth ratio of 3 to 1. Two % -in. holes were drilled at 45" angles, as shown, to provide for the loading devices. After construction, the model was annealed at 240°F to remove stresses produced during machining.

Jan 1, 1971

By B. C. Blake

IN general, billet-type or long slender ingots are used when it is desired to produce directly from the ingot in one conversion a product of medium or small cross-sectional area. They are designed to eliminate the blooming and billet-mill stages of the usual procedure. As used at Connors, they constitute a means of keeping a continuous rolling mill busy in the production of high-quality small rods, small angles, hoop stock, baling ties and concrete reinforcing rods without the need for a blooming or billet mill. They represent sound steel with the best surface that is commercially practical to use in the rolling of small continuous mill products.

Jan 1, 1947

A record $18.7 billion worth of metals, non- metals and fuels was produced in the U.S. during 1962, according to a year-end estimate from the USBM. The 1962 total value, based on preliminary statistics compiled by the Bureau, was nearly $600 million more than in 1961. Most minerals gained both in value and volume of production during 1962 according to information collected by the Bureau. All mineral fuels except anthracite showed increases. Among nonmetallic minerals- other than fuels-gains exceeded losses by two to one. Metal production held steady as gains and losses balanced.

Jan 2, 1963

By R. G. Burdick, W. R. Peters

Shallow underground voids resulting from early coal mining and other resource recovery activities over the past several decades are now being recognized as a significant cause of ground subsidence problems. Uncertain knowledge of abondoned coal mines also imposes potential hazards in current underground mining operations, since water inundation or the release of methane gas is a serious hazard when mine excavation operations break into an abandoned mine. Requirements for detecting and mapping subversive manmade tunnels in previous work has led to the development of a surface-operated automatic earth resistivity survey system and associated computer data reduction process for military applications. Field demonstration tests of this system have resulted in successful detection of tunnels at depth-to-diameter ratios greater than 10 to one. Under the sponsorship of the Bureau of Mines, a similar high-resolution earth resistivity system was designed and constructed for use as a potential means of detecting abandoned coal mine workings. The hardware and software aspects of third new system are described together with applications of the method for surveying and mapping abandoned coal mine workings. Field tests are presented to illustrate the detection of both air-filled and water-filled mine workings.

Jan 1, 1984

By Robert H. Richards

(Pittsburg Meeting, March, 1910.) THIS is in part a review paper, indicating the various steps that have been taken in developing hindered-settling apparatus, some of the standard data that have been obtained, and some of the conclusions one is led to as to the effect of it in improving concentrating-methods and machinery, and in part it brings in some unpublished work. CAUSES OF LOSS IN ORE-CONCENTRATION. In making a study of the separation of valuable minerals from the waste gangue, particularly by water concentrators, it is necessary to define, first of all, what are the causes of-loss that are likely to interfere with the concentration. Having made this definition; we can then study to see in what way the machines make their losses, and plan remedies for them in an intelligent way. Losses may be stated to occur as included grains, as flat or long grains, as fine free mineral, and as absolute slime. We will take, these up one by one, as follows: Included Grains.-Fig. 1 shows by a set of diagrams how included grains may occur: first, where the heavy mineral grains attached to the quartz are of large size; secondly, where they are of small size ; and thirdly, where very fine particles of heavy mineral are included in the quartz. We have to decide in regard to included grains whether they are rich enough to go direct to, the furnace, or poor enough to throw away, or whether they lie between these two points and require recrushing as the only means of severing the values from the waste. This last case is the only one requiring special treatment at our hands. We will therefore consider that next. * By mutual consent, this paper has been presented also before the Colorado Scientific Society, and the Canadian Mining Institute, Montreal.

Feb 1, 1911

By E. Martinez

The effect of ultrasonic energy transmitted through a liquid on chrysotile asbestos was investigated. Ultrasonic energy was effective in fiberization of chrysotile crudes as well as standard grades of milled asbestos fibers. Treated samples of several grades of milled fibers had increased wet bulk volumes and decreased settling and filtration rates compared to untreated fibers. Ultrasonics may be a method for fiberization of chrysotile without decreasing length or use of chemical additives that alter surface properties. Cavitation of the liquid on the surface of fiber bundles is believed to be the cause of the fiberization. It was found that ultrasonic energy greatly increased the rate of precipitation of latex on asbestos. When chrysotile asbestos fibers require additional opening or fiberization, the usual method employed is some type of high speed impact or grinding to open the fibers and produce a larger surface area. These methods always cause some degradation in length due to breaking of fibers. In addition, the efficiency of mechanical methods decreases as the diameter of fiber bundles decreases. Chemical methods have been proposed and it is claimed that fiberization down to individual fibers of 200-300 A diam is possible with some wetting agents. However, large amounts of reagent are required, approximately EDWARD MARTINEZ is with Central Research Laboratories, American Smelting and Refining Co., South Plain-field, N.J. TP 63H226. Manuscript, Aug. 12, 1963. Discussion of this paper, submitted in duplicate prior to Mar. 15, 1964, will appear in SME Transactions, September 1964, and AIME Transactions, 1964, vol. 229. 50 pct of the weight of fiber treated, and some of the wetting agent remains on the surface of the chrysotile and cannot be removed by any solvent which does not destroy the asbestos.' This investigation was directed at determining whether ultrasonic energy transmitted through a liquid would be effective in fiberizing chrysotile asbestos. Samples of asbestos crudes and commercially milled fibers were treated; settling and filtration rates of treated and untreated milled fibers were obtained. EQUIPMENT A Westinghouse Corp. ultrasonic cleaning system WEHH was used in this investigation. See Table I for a breakdown of this ultrasonic unit. RESULTS AND DISCUSSION Fiberization of Chrysotile Crudes: Samples of chrysotile crudes were pulled apart with tweezers so that coarse bundles of fiber approximately 1/16 in. in diam were obtained. The bundles were put into a bottle with 200 ml of methanol and the mixture placed into the ultrasonic tank for 30 min at a power output of 1200 w. The test was repeated using water instead of methanol. During the irradiation it was noted that the surfaces of the bundles were being fiberized. The liquid eventually became cloudy with dispersed solids. Fig. 1 is an electron micrograph of the fibers, some reduced to unit dimensions, and a fine material released by the ultrasonic energy. The platy particles in Fig. 1 were identified by electron diffraction and differential thermal analysis as brucite, Mg(OH)2, which must have been intergrown in the chrysotile crudes. Electron micrographs of bulk specimens of chrysotile have shown a fine material filling the interstices between tubular fibers.2,3 Electron diffraction patterns of the fine material collected after ultrasonic treatment indicate that it is amorphous. After decanting the solution, bundles of chrysotile remained somewhat decreased in diameter. Continuing the ultrasonic irradiation of the remaining material resulted in additional fiberization. The opening of the fibers occurred in both methanol and water. Cavitation of the liquid by the ultrasonic energy is undoubtedly responsible for the fiberization. When sound waves are transmitted through a liquid at a

Jan 1, 1963

By James Barr

Geology and Mineralogy1 TENNESSEE phosphates are commercially divided into three varieties: Brown, blue and white. The first two only are now of commercial importance. The white phosphates of Perry county are not being mined at present owing to their pockety occurrence. The white phosphate is Fig. I.-DEPOSITS OF 75 PER CENT. BROWN PLATE ROCK UNDERLAIN BY SAND. the highest grade of the three varieties, running often more than 90 per cent. tricalcium phosphate. The brown rock is found in three varieties, grading from the dense almost vitreous gray rock, 76 to 82 per cent. tricalcium phosphate, down

Jan 9, 1914

By Edwin B. Eckel

This paper, based on brief field examination and on data supplied by the operators, records the condition of the Italian mercury industry as of March 1945, not long after Italy's collapse. Except for the once-Italian but now Yugoslavian Idria mine near Trieste, virtually all the known mercury deposits are in the Monte Amiata district in Siena and Grosseto Provinces, 75 miles north of Rome. This district contains the Abbadia San Salvatore. the Siele. the Argus, and the Cerreto Piano mines besides numerous prospects and unproductive mines. Under pressure from Germany the principal mines were highly productive during the early part of the war, but production fell off long before its end. All of the mines were damaged in some way at the German withdrawal but by September 1944 the Abbadia had resumed production and several others were only lacking power by the early spring of 1945. The Monte Amiata district is characterized geologically by a thick series of sedimentary rocks that range from Triassic to Eocene in age . and are surmounted on Monte Amiata itself by a thick cap of extrusive trachyte. Bodies of travertine and of landslide detritus, younger than the trachyte, are rather widespread. The structure is complex and little known but most of the ore deposits seem to lie along or near a major, north-trending fault or on associated cross-fractures. Cinnabar is the only important ore mineral and clay minerals are by far the most abundant products of hydrothermal alteration. A southward increase in the amount of stibnite, relative to cinnabar, suggests horizontal zoning. The grade of ore ranges from 0.60 to 3.0 Pet and the reserves, as compared with those in American mines, are enormous. Even in the face of war-time scarcities, the mines are remarkably self-sufficient and labor is plentiful, but the costs of production are estimated to be not greatly below those of the better American mines. Notes on the principal mines, on the mining and metallurgical methods, and on the manufacture of synthetic cinnabar are included in the paper. Introduction With but few brief interludes, Spain and Italy, including the Idria mine near Trieste, which has passed from Austrian to Italian to Yugoslavian hands in the Past 3 decades, have led the world in the production of mercury, or quicksilver, for many centuries. Despite their enviable position with respect to the supplies of a metal so essential in both war and peace, there is a notable lack of up-to-date information in America on the mercury industry of either country. Duschak and Schuette' and Schuettez,3 give by far the best summaries in English known to the writer, but the latest firsthand report cited by them on the Italian mines was published in 1915.' Having been deeply interested in the geology and economics of mercury for several years, the writer naturally felt most fortunate in being the to obtain enough firsthand information on the Italian industry to fill at least a few of the gaps in our knowledge. The opportunity came during the war, soon after this industry, together with the rest of Italy, had capitulated to the Allies. Many of the facts presented here are thus definitely "dated," but others are of less transitory nature, and even the dated facts should be of interest in showing the strength and

Jan 1, 1949

By Edwin B. Eckel

This paper, based on brief field examination and on data supplied by the operators, records the condition of the Italian mercury industry as of March 1945, not long after Italy's collapse. Except for the once-Italian but now Yugoslavian Idria mine near Trieste, virtually all the known mercury deposits are in the Monte Amiata district in Siena and Grosseto Provinces, 75 miles north of Rome. This district contains the Abbadia San Salvatore. the Siele. the Argus, and the Cerreto Piano mines besides numerous prospects and unproductive mines. Under pressure from Germany the principal mines were highly productive during the early part of the war, but production fell off long before its end. All of the mines were damaged in some way at the German withdrawal but by September 1944 the Abbadia had resumed production and several others were only lacking power by the early spring of 1945. The Monte Amiata district is characterized geologically by a thick series of sedimentary rocks that range from Triassic to Eocene in age . and are surmounted on Monte Amiata itself by a thick cap of extrusive trachyte. Bodies of travertine and of landslide detritus, younger than the trachyte, are rather widespread. The structure is complex and little known but most of the ore deposits seem to lie along or near a major, north-trending fault or on associated cross-fractures. Cinnabar is the only important ore mineral and clay minerals are by far the most abundant products of hydrothermal alteration. A southward increase in the amount of stibnite, relative to cinnabar, suggests horizontal zoning. The grade of ore ranges from 0.60 to 3.0 Pet and the reserves, as compared with those in American mines, are enormous. Even in the face of war-time scarcities, the mines are remarkably self-sufficient and labor is plentiful, but the costs of production are estimated to be not greatly below those of the better American mines. Notes on the principal mines, on the mining and metallurgical methods, and on the manufacture of synthetic cinnabar are included in the paper. Introduction With but few brief interludes, Spain and Italy, including the Idria mine near Trieste, which has passed from Austrian to Italian to Yugoslavian hands in the Past 3 decades, have led the world in the production of mercury, or quicksilver, for many centuries. Despite their enviable position with respect to the supplies of a metal so essential in both war and peace, there is a notable lack of up-to-date information in America on the mercury industry of either country. Duschak and Schuette' and Schuettez,3 give by far the best summaries in English known to the writer, but the latest firsthand report cited by them on the Italian mines was published in 1915.' Having been deeply interested in the geology and economics of mercury for several years, the writer naturally felt most fortunate in being the to obtain enough firsthand information on the Italian industry to fill at least a few of the gaps in our knowledge. The opportunity came during the war, soon after this industry, together with the rest of Italy, had capitulated to the Allies. Many of the facts presented here are thus definitely "dated," but others are of less transitory nature, and even the dated facts should be of interest in showing the strength and

Jan 1, 1949

By J. W. JESSUH

TO certain people the name of Northern Rhodesia brings only a vague recollection of a distant country somewhere in Africa; to others, it means a big game territory and the opportunity for excellent shooting; to some, a land of sunshine and peace; but to mining men the name signifies the land where the latest discoveries of copper ore are astonishing the mining world and bid fair, in a few years, to make Northern Rhodesia one of the largest copper- producing countries. There is a fascination about this part of the globe that is somewhat similar to southern California, and at the same time there is that feeling of mystery peculiar to a tropical climate. Perhaps it is the isolation, the fear of wild animals, or the influence of an undefinable something which cannot be expressed.

Jan 1, 1931

By G. R. St. Pierre, A. J. Wilhelem

In connection with the reduction of hematite or magnetite to metallic iron, it appeared desirable to study the rate of reduction of each oxide to the next lower oxide under conditions which excluded all other condensed phases. For this purpose, pellets of Fe,O,, Fe,O,, and oFeOo were prepared. A gas train for the preparation of Ar-Hz-H20 mixtures was employed to establish a continuous flow of a controlled atmosphere over a pellet suspended on a balance. In this manner the following individual reduction steps could be studied: For example, pellets of hematite were reduced solely to magnetite by adjusting the gas atmosphere such that it was oxidizing to wustite. In other words, the reaction was complete when the hematite pellet had been totally converted to magnetite. The pellets were prepared from Baker Chemical ferric oxide. The ferric oxide powder was moistened and handrolled. Air sintering at 950°C produced Fe,0, pellets with a density of 3.65 g per cm3. The pellets ranged in size from 0.6 to 1.2 cm diam. The magnetite and

Jan 1, 1962

By H. E. Strauss

The time-to-fracture test has been applied to pure platinum and to two alloys of platinum under the special conditions of small cross-sectional area of the specimens and of a test temperature above the annealing temperature. While the results are similar in character to those secured with large specimens of steel at temperatures below the annealing temperatures, two differences are noted: (I) an upward inflection in the curves for log stress vs. log time for the alloys, and (2) a sharp break in the curve for log stress vs. log time for pure platinum when single crystals filled the wire. The experiments showed that: (I) platinum and 10 per cent rhodium-platinum maintained their strengths at IIOO°C. better than did to per cent iridium-platinum, and (2) the rhodium-platinum after a time exceeded the iridium-platinum in strength at IIOO°C. Platinum alloys in their service at high temperatures encounter conditions somewhat similar to those that exist in the usual form of the "time-to-fracture"l test with steel. This paper is intended primarily to present results obtained by a "time-to-fracture" test under conditions approximating actual service for platinum (at the temperature IIOO°C. under oxidizing conditions) with three materials, pure platinum (99.99 per cent), a platinum alloy with 10 per cent iridium, and a platinum alloy with 10 per cent rhodium. The experimental procedure differed from that previously used in this type of measurement in that the samples were of small diameter (0.010 in.), were fully annealed, and were tested at a temperature definitely above their annealing points (600°C., IOOO°C. and 1050°C., respectively, for 60 per cent cold reduction and 15 min. heating time2,3). The experimental arrangement was similar to that previously described,' using as heating unit a small, horizontal, platinum-wound furnace, ½ by 12 in., maintained at constant temperature by an automatic controller. The samples were 0.010-in. wires, which had been annealed at 1200°C. by drawing them slowly through the furnace. The time-to-fracture was measured electrically. Elongation was not determined because of the difficulties inherent in the experimental arrangement. Life—that is, the time-to-fracture—was considered the important property in these first tests. No attempt was made to measure the reduction in area, but the types of fracture were noted qualitatively; in the text they are referred to as (I) the normal, plastic, or necked fracture, (2) the knife-edge fracture, and (3) the brittle or intercrystalline fratture with small reduction of area. Results for different constant applied stresses are plotted in Fig. I. In all cases at least two trials were made for each stress, although they cannot always be indicated separately on the curves. The breaking Stresses drop very rapidly to a small frat-286

Jan 1, 1943

By H. E. Strauss

The time-to-fracture test has been applied to pure platinum and to two alloys of platinum under the special conditions of small cross-sectional area of the specimens and of a test temperature above the annealing temperature. While the results are similar in character to those secured with large specimens of steel at temperatures below the annealing temperatures, two differences are noted: (I) an upward inflection in the curves for log stress vs. log time for the alloys, and (2) a sharp break in the curve for log stress vs. log time for pure platinum when single crystals filled the wire. The experiments showed that: (I) platinum and 10 per cent rhodium-platinum maintained their strengths at IIOO°C. better than did to per cent iridium-platinum, and (2) the rhodium-platinum after a time exceeded the iridium-platinum in strength at IIOO°C. Platinum alloys in their service at high temperatures encounter conditions somewhat similar to those that exist in the usual form of the "time-to-fracture"l test with steel. This paper is intended primarily to present results obtained by a "time-to-fracture" test under conditions approximating actual service for platinum (at the temperature IIOO°C. under oxidizing conditions) with three materials, pure platinum (99.99 per cent), a platinum alloy with 10 per cent iridium, and a platinum alloy with 10 per cent rhodium. The experimental procedure differed from that previously used in this type of measurement in that the samples were of small diameter (0.010 in.), were fully annealed, and were tested at a temperature definitely above their annealing points (600°C., IOOO°C. and 1050°C., respectively, for 60 per cent cold reduction and 15 min. heating time2,3). The experimental arrangement was similar to that previously described,' using as heating unit a small, horizontal, platinum-wound furnace, ½ by 12 in., maintained at constant temperature by an automatic controller. The samples were 0.010-in. wires, which had been annealed at 1200°C. by drawing them slowly through the furnace. The time-to-fracture was measured electrically. Elongation was not determined because of the difficulties inherent in the experimental arrangement. Life—that is, the time-to-fracture—was considered the important property in these first tests. No attempt was made to measure the reduction in area, but the types of fracture were noted qualitatively; in the text they are referred to as (I) the normal, plastic, or necked fracture, (2) the knife-edge fracture, and (3) the brittle or intercrystalline fratture with small reduction of area. Results for different constant applied stresses are plotted in Fig. I. In all cases at least two trials were made for each stress, although they cannot always be indicated separately on the curves. The breaking Stresses drop very rapidly to a small frat-286

Jan 1, 1943

Discussion of the paper of Robert Henry Lee, presented at the New York Meeting, October, 1913, and printed in Bulletin No. 82, October, 1913, pp. 2515 to 2522. J. E. JOHNSON, JR., New York, N. Y.:-This subject is of great and increasing importance, as the exhaustion of the deposits of ore of ideal character compels the iron industry to turn more and more to those which are less desirable on account of physical structure as well as analysis. Thus there has come about the use of the Mesabi ores in this country and also of the Minette ores in Europe, which produce far greater percentages of flue dust than did the lumpy ores first used. It is a matter of common knowledge that the utilization of this flue dust, containing commonly 50 per. cent. or more of metallic iron, has been a problem very difficult and tedious of solution. The number and variety of the solutions now offered is a sufficient proof of this. . I have no intention of going into the merits or demerits of any of these processes as such, but there are certain considerations concerning the effect of their product on the operation and fuel economy of the blast furnace which I believe to be new and which I desire to bring out. There are two methods, so different as almost to be direct opposites of each other, whereby an ore may exercise an unfavorable effect on the economy of the blast furnace. The first of these is well known and results from the ore being refractory and highly resistant to the action of the reducing gases of the blast furnace. This results in its being incompletely reduced when it reaches the zone of fusion and in the reduction being finished by the direct action of carbon in the lower portion of the furnace. This is so well understood as to need little elaboration or comment. There is, however, another action which may be equally as fatal to the economy of the furnace as the foregoing, and as this point is, to the best of my knowledge, not generally appreciated, it is one upon which I desire to lay particular stress. This consists in the ore being, in a sense, too reducible.

Jan 12, 1913

By R. E. Hummel, J. W. Koger

IT is generally accepted that the martensitic start temperature (Ms) can be determined by resistivity measurements and is that temperature where the resistivity vs temperature curve on cooling first deviates from a straight line.1-4 Both Titchener and ever' and pops5 have used resistivity measurements to follow the martensitic transformation in b1 Cu-Zn and ternary Cu-Zn-based b-phase alloys, respectively. Actually because of the nature of their measurements the temperature of minimum resistivity (M) was used by these authors to determine "Ms" . Also, due to experimental difficulties in detecting M, with resistivity measurements, Pops and Massalski 8 determined M, by optical microscopy. The present communication is concerned with high-precision, low-temperature resistivity measurements which reveal a definite M, temperature which differs from the temperature of minimum resistivity (M). It is also shown that M is merely the result of the geometry of the percent-transformation curve and the temperature coefficient of resistivity of the alloy. In other words, it will be demonstrated that M has no apparent significance (except that it denotes a temperature at which some percent of martensite can be detected in a given alloy). Alloy specimens were made by melting 99.999 pct Cu and 99.9999 pct Zn in sealed and evacuated quartz tubes and homogenizing them at 601PC for 300 hr. After the alloy was cold-rolled into foils suitable for resistivity measurements, it was quenched in ice water from 871PC. The chemical analysis showed a zinc content of 38.8 wt pct. The cooling rate during the martensite transformation was 1°C per min and was controlled with a thermistor. Electrical resistivity measurements were made using a K-3 type potentiometer; the reproducibility of the resistivity values was better than 0.03 pct. Temperatures were measured with a thermocouple to an accuracy of better than 0.2° C. Fig. 1 shows the resistivity vs temperature curves of the martensitic transformation in B1 Cu-Zn for both cooling and heating. The starting point of the martensitic transformation (M,) was taken as that

Jan 1, 1968

Four years of intensive exploration effort by Newmont Mining Corp. came to a climax May 27 in Carlin, Nevada, when local, state and federal officials joined with company officers at a dedication ceremony saluting the start-up of the firm's wholly- owned subsidiary, Carlin Gold Mining Co. Before an invited audience at the plant site, Newmont's President Plato Malozemoff and Nevada's Senator Howard D. Cannon pushed a lever and watched the new melting furnace rotate until a brilliant orange-red stream of molten gold poured forth from its mouth. Two 1200-oz bricks of bullion were poured that day, and they are the first of many that will lift this mine into the number two slot among U.S. gold producers and fourth among all gold mines on the North American continent. Although Newmont Mining first entered the gold mining business in 1929 in Grass Valley, California, the Company had operated no gold producers since the well-known Government directive L208, issued in 1942, closed down U.S. gold mines for the duration of World War 11. Now it is back in the business in a big way-a $10 million mine and plant capable of producing 200,000 oz. per year. All the gold from the Carlin operations is earmarked for the U.S. Mint in San Francisco.

Jan 7, 1965

By E. Northcott, F. N. Oberg

By electrostatic separation, course Florida pebble phosphate, too lou-grade to find a ready market, can be upgraded to a satisfactory saleable product. Pebble running from 60 pct bone phosphate of time may be upgraded to 73 to 77 pct BPL depending on the mineralogical. structure of the pebble and the nature of the silica innpurity. High-grade concentrates from Florida pebble phosphate deposits have long been recovered by wet methods. When work was begun in the field, only the coarse pebble (generally +1 mm) was re- covered. In the 1920's a froth flotation process was developed to separate quartz from the finer phosphate fractions, but this method proved unsatisfactory for treating the intermediate fraction (—1 mm +35 mesh). Consequently various other processes were adapted, including agglomerate tabling and spiral separation of reagentized feed. Today the three processes are used—washing and screening for coarse pebble, tabling or spiraling for intermediate sizes, and flotation for —35 mesh material.

Jan 1, 1959

By W. L. Dumoulin

A RATHER detailed description of the entire plant and leaching process was given in a paper recently presented to the Institute,1 so this paper will cover briefly only the crushing practice of the New Cornelia Copper Co. for the year, 1918. The ore, which is mined by steam shovels and loaded in side-dump cars, passes through two crushing plants. These crushing plants were constructed to crush the ore required by a leaching plant of 5000 tons daily, capacity, during a crushing period of 16 hours. The primary plant reduces to less than 3 in. (76 mm.) and the secondary plant, to ¼ in. (6.35 mm.), which is sufficiently fine for satisfactory percolation of solution through the ore in the leaching tanks, and to give good extraction. The ore is then conveyed to the leaching tanks by a system of 28-in.. (71-cm.) belt conveyors. On the way to the leaching tanks, it passes through an automatic sampling plant, where a 1-per cent. sample is taken. From the primary crushing plant, the ore is conveyed by two 36-in. (91-cm.) belt conveyors to a 10,000-ton steel storage bin, with a reinforced-concrete flat bottom. The ore discharges from the bottom of this bin onto four 20-in. belt conveyors, which deliver it to the four units of crushers in the secondary crushing plant. There is no storage bin between the mine and the primary crushing plant. The ore breaks very coarse, is very hard, and contains a great many boulders, some as large as 4 ½ by 4 1/2 by 10 ft. (1.3 by 1.3 by 3 m.). Jams form in the bowl of the coarse crusher in the primary crushing plant, as a consequence, and are freed by means of an immense steel hook operated from a 40-ton electric traveling crane.

Jan 8, 1919