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By L. C. McCabe

THIS paper is a progress report on a study, of the distribution of the banded ingredients (Figs. 1 and 2) in- Illinois screenings and the method of determining their distribution. Proximate analyses and screen tests1 have been reported and a further report on the washability characteristics of the screenings is in preparation. The microscopic study of the screen-ings is not yet far enough advanced to permit a conclusive report on the effect of mining, screening and washing on the distribution of the banded ingredients in the coals of Illinois. However, considerable fundamental data-have been collected as a background for the screenings investigation and for a proposed study of fuels composed of known quantities of vitrain, clarain and fusain, the three common banded ingredients of Illinois coals. These fundamental data and the preliminary information now available from the microscopic study of screenings are broadly indicative of the reaction of the coals to the physical processes affecting them in mining and preparation. Proximate analyses of samples of vitrain, clarain, and fusain from several Illinois mines have been published 2, but those given in Table 1 serve to show the differences that are apparent in proximate analyses. The vitrain samples are consistently low in ash, and the unit-coal value is uniformly lower than in clarain. The average .fuel ratio (F.C./V.M.) of the vitrain samples in the table is 1.57-whereas the clarain average is 1.42. The fusain analyses vary considerably in ash and moisture. In the two samples from this mine the moisture values range from 8.99 to 22.80 and the ash from 4.86 to 16.91 per cent. Fusain has higher fixed carbon than either vitrain or clarain but the percentage is erratic from sample to sample.

Jan 1, 1936

By Thomas V. Falkie

4.3-1. Introduction. DEFINITION OF PRODUCTION SYSTEMS ENGINEERING. Surface mines, like most other mining and manufacturing operations, are usually organized along functional departmental lines. Frequently a mine manager will have his responsibilities divided among production, maintenance, administrative, and engineering functions. Often these are separate departments, but various combinations and variations of these functions can be found in mining companies throughout the world. The production function is responsible for physically removing the ore from the ground, processing it, and shipping the product to its market. The production manager supervises and is responsible for all man-machine systems necessary to extract and process the ore. The maintenance function, which is sometimes included within the framework of production, has the task of insuring the maximum availability and longevity of equipment needed to produce the ore. The administrative function is responsible for accounting, personnel, industrial relations, purchasing, public relations, property, and other related staff functions. For purposes of this chapter, engineering can be described as the function which plans, designs, and supervises construction of physical plant and equipment. In this context engineering includes mechanical, civil, chemical, electrical, and mining engineering. However, another phase of engineering is to prepare plans to insure that the various man-machine components function in an optimum way towards the single most important goal of any enterprise-the maximization of return on the investors' money. Along with the planning, another duty of engineering is to audit all phases of the operation for the purpose of recommending improvements toward this same general goal. These latter two ideas form the basis for a discussion of production engineering and systems engineering. Production engineering is a catchall, problem-oriented function whose duties can include: 1) solution of day-to-day production problems, 2) short-

Jan 1, 1968

By B. W. Prof. Frazier

(Read at the Wilkes-Barre Meeting, May, 1877.) THIS interesting application of the laws of thermodynamics to metallurgical practice has not been discussed by any writer, within my reading, except the late Prof. Gallon of Paris. In his Cours de Machines,* Prof. Callon has given a short but lucid discussion of this subject in the admirably clear style for which he was remarkable. Although the interest attached to the subject is rather theoretical than practical, I have thought that an attempt to explain the mechanical action exerted in the hot-blast oven might not be entirely devoid of utility, especially in the suggestions which it involves. The statement, that the mechanical work performed in heating the blast is, when high temperatures are produced, considerably greater than the work performed by the blowing-engine, may seem incredible to many who are practically acquainted with the subject, appearing, as it does, to conflict with the results of their experience. It is, nevertheless, literally true, and I shall attempt in this paper to prove it, and to explain why it has not rendered itself palpably manifest in practice. Before proceeding to the mathematical demonstration, it will be well to recall to mind the definitions of some terms employed by writers in mechanics and thermodynamics. The actual energy of a moving body, as defined by Rankine, is the work which it is capable of performing against a retarding resistance before being brought to rest, and is equal to the energy which must be exerted on the body to bring it from a state of rest to its actual velocity. The value of that quantity is the product of the weight of the body into the height from which it must fall to acquire its actual velocity. * Vol. I, chap. x, [§ ] 4, p. 312.

Jan 1, 1878

By John Gross

The building of reverberatory furnaces (Fortschaufelungsofen) where ordinary brick, fire-brick and iron are comparatively cheap, is quite a different matter from the building of such furnaces in isolated camps, where proper material is only to be obtained at high cost and with long delays. Time is always a large factor in metallurgical operations, and the metallurgist may find himself in a position where it becomes necessary to erect apparatus and have it running in less time than it would take to obtain material from outside. It is here that he must make use of makeshifts. In Mexico, one of the most important of these is the alluseful adobe. Employed from prehistoric times, it still has its sphere of application in a surprisingly large number of instances requiring an article that can be quickly and cheaply produced. It is quite possible to erect serviceable reverberatory furnaces with no other materials than adobes, stone and wood. The old Mexican smelting shaft-furnace was an example of what can be done with adobes, and the large number of ruins of this class of furnace scattered through the mining districts of Mexico are silent testimony to the genius of a people who demonstrated their capability of adapting material at hand to their needs. The present paper will describe briefly the construction of an adobe reverberatory furnace which is giving good results, can be quickly erected at a low cost, and, if properly built and handled, will last a long time. The material being simply sun-dried, it naturally follows that the construction must be quite heavy. The adobes should be evenly made, with just sufficient straw to hold them together, and not too large in size (9 in. wide, 18 in. long and 4 in. thick

Jan 1, 1902

By Clarence Zener

Theoretical considerations by one of the authors have ledl to the prediction that the dynamic internal friction of annealed metals has a broad maximum at a certain grain size. This prediction they have verified2 for alpha brass. In the present paper the theory is extended to include the precise manner in which the internal friction varies with grain size on either side of the maximum. New experimental data are presented to test this extended theory. The physical basis for the type of internal friction that is being investigated lies in the interplay between stress and temperature. Just as local fluctuations in temperature give rise to local fluctuations in stress, so likewise fluctuations in stress give rise to fluctuations in temperature. But temperature gradients give rise to necessarily irreversible thermal currents, and hence to internal friction. The theory for this type of internal friction has been developed quantitatively,³ and has been experimentally verified,4,5 for the particularly simple case of the transverse thermal currents accompanying transverse vibrations. Of more fundamental importance are the intercrystalline thermal currents accompanying all types of vibration. The variation in stress across grain boundaries arises from the elastic anisotropy and at least partial random orientation of the individual crystallites. An exact quantitative treatment of the internal friction due to these thermal currents is of course out of the question. However, much information can be obtained from general arguments.1,6 Thus when the frequency of vibration f is low, or when the grain size d is small, adjacent grains remain nearly in thermal equilibrium, and vibration proceeds isothermally with little internal friction. Again, when the frequency is high, or the grain size large, the heat flow per cycle is small, the vibration proceeds adiabatically with little internal friction. The internal friction will become a maximum in the transition region. The degree of adiabatic-

Jan 1, 1940

By Clarence Zener

Theoretical considerations by one of the authors have ledl to the prediction that the dynamic internal friction of annealed metals has a broad maximum at a certain grain size. This prediction they have verified2 for alpha brass. In the present paper the theory is extended to include the precise manner in which the internal friction varies with grain size on either side of the maximum. New experimental data are presented to test this extended theory. The physical basis for the type of internal friction that is being investigated lies in the interplay between stress and temperature. Just as local fluctuations in temperature give rise to local fluctuations in stress, so likewise fluctuations in stress give rise to fluctuations in temperature. But temperature gradients give rise to necessarily irreversible thermal currents, and hence to internal friction. The theory for this type of internal friction has been developed quantitatively,³ and has been experimentally verified,4,5 for the particularly simple case of the transverse thermal currents accompanying transverse vibrations. Of more fundamental importance are the intercrystalline thermal currents accompanying all types of vibration. The variation in stress across grain boundaries arises from the elastic anisotropy and at least partial random orientation of the individual crystallites. An exact quantitative treatment of the internal friction due to these thermal currents is of course out of the question. However, much information can be obtained from general arguments.1,6 Thus when the frequency of vibration f is low, or when the grain size d is small, adjacent grains remain nearly in thermal equilibrium, and vibration proceeds isothermally with little internal friction. Again, when the frequency is high, or the grain size large, the heat flow per cycle is small, the vibration proceeds adiabatically with little internal friction. The internal friction will become a maximum in the transition region. The degree of adiabatic-

Jan 1, 1940

By Clarence Wright

THE average lead and zinc content of the ores mined and milled in the Joplin district is low as compared with that of other lead and zinc deposits throughout the United States. Because of this fact and of the high grade of the concentrates, the ratio of concentration is unusually high. Most of the operating companies have aimed to produce as high a grade of concentrate as possible in order to obtain the highest price for their product. They have done this at the sacrifice of the percentage of zinc recovered. The average percentage of zinc recovery is from 60 to 70 per cent. For percentage of recoveries from actual mill tests, see tables. The concentration is commonly effected by crushing the ore to h-in. (12.7 mm.) size and subjecting it to a roughing and cleaning over two or more large Cooley jigs, of the Harz type, the finer sands being treated on the usual types of sand and slime tables. The treatment of slimes by flotation is being tested at several of the larger mills. COARSE CONCENTRATION When the ore. has been hoisted from the mine it is dumped onto a chute leading to a grizzly made of iron bars, usually heavy rails, spaced 4 to -5 in. (101 to 127 mm.) apart. The oversize is caught on the grizzly, the mineralized pieces are broken with sledge hammers so as to pass through, and the boulders that are practically barren are sorted out by hand and trammed to the waste-rock pile. The proportion of waste rock thus sorted out varies from a fraction of 1 per cent to 15 per cent., depending on the nature of the rock.

Jan 10, 1917

Jan 1, 1881

By Oliver Bowles

THE inorganic non-metallic minerals, that is, the non-metallics exclusive of coal, oil, gas and related minerals, constitute the basic raw materials for a number of essential industries. It is estimated that the value of the raw materials in this group produced in 1919, together with certain of the primary products ordinarily manufactured at the mine, exceeded $800,000,000. They are employed chiefly in the building, agricultural, and chemical industries. It has been estimated, by the Bureau of Mines, that approximately 00,000,000 worth of non-metallic minerals and their products were used in the building trades during 1918; with the great increase in building since that date and the increased cost of materials, the value of non-metals used in such industries during 1920 probably exceeded $500,000,000. The chief of these materials are cement, building stone, crushed stone, sand and gravel, clay products, lime, gypsum, glass materials, slate, talc and soapstone. In many of the producing industries of this great group, the waste is so great that the expense of handling it is an important item in the production cost. The utilization of waste is receiving unusual attention in many of the non-metallic industries at present. The most obvious advantages of such utilization are: (1) Conservation of mineral resources, (2) the removal of waste heaps that in many places obstruct development, (3) manufacture of new and useful products, (4) reduced production costs through elimination of a large part of the item of waste-disposal expense. In many industries there is keen competition between producers of various products used for the same purpose, so that advantage will accrue to the manufacturer who is able to reduce his production costs.

Jan 1, 1921

By Stanislaus Skowronski

THE melting of cathode copper, usually containing 99.98+ per cent. Cu, would appear to be a simple matter. Owing to the well known affinity of copper for sulphur, however, so much sulphur is absorbed by the copper during the operation that a long and expensive refining process is required to remove it, and the final product is inferior in purity and lower in electrical conductivity than the original cathode copper. The furnace operation, as now conducted, includes the following stages: 1. Charging the cathodes into the furnace. 2. Melting the cathodes. 3. Oxidizing period-saturating the copper with cuprous sub-oxide and bringing the metal to the condition known as "set copper;" this is now done by blowing air into the molten bath of metal. 4. Reduction period, otherwise known as "poling," whereby the oxygen previously added is removed by covering the molten bath completely with coke or charcoal and introducing large poles of green wood; this reduces the cuprous sub-oxide, bringing the copper to the "tough pitch" stage ready for casting. 5. Casting Period. Many experiments have been made to obviate the long periods of oxidation and subsequent reduction, with no success. While the copper first, cast will be good, as indicated by the set surface of the bars, it is impossible to hold it to the "tough pitch" condition, and long before the whole charge has been poured the copper will become "overpoled," entailing numerous difficulties. Sulphur has always been blamed by the furnace men as the reason why this refining process is necessary, yet the literature on the subject does not sufficiently emphasize how small a proportion of sulphur will ruin copper for practical purposes.

Jan 3, 1918

By H. F. Yancey

THE jig may be termed the standard coal-washing machine. Although exact figures on the relative tonnages of coal treated by the various coal-cleaning processes are not available, it is known that a large part of all the coal given special cleaning treatment is handled by jigs. This wide use of the jig warrants a special study of its performance. In this paper are presented the results of an analysis of work done by a three-compartment jig in treating bituminous coal from the Thick Freeport bed in Allegheny County, Pa. Certain systems of washing coal require sizing before washing, while others treat unsized coal. The prevailing practice, however, for the preparation of coking coal from the more easily washed coal is to treat an unsized or natural feed at about -1 in. size and smaller. Under such conditions, the minimum size of particle that is beneficiated has long been and is still a matter of discussion and conjecture. In a careful laboratory investigation Jüngst1 concluded that the practical limit of settling of fine coal is reached at from 0.02 to 0.008 in. On the other hand, Hancock2 has found that single-compartment jigs, used in preparing coal for the domestic market, usually allow most of the heavy (refuse) particles smaller than 1/4 in. to go into the washed coal. He also states that it is impossible with the ordinary coal jig to make any noticeable improvement in the quality of the coal finer than 20 mesh (about 1/30 in.). Because of its importance in coal cleaning, the work described in this paper was undertaken for the purpose of obtaining information on this subject, as a part of the general coal beneficiation program being carried out by the Bureau of Mines.

Jan 4, 1925

By Thomas Fraser, H. F. Yancey

The jig may be termed the standard coal-washing machine. Although exact figures on the relative tonnages of coal treated by the various coal-cleaning processes are not available, it is known that a large part of all the coal given special cleaning treatment is handled by jigs. This wide use of the jig warrants a special study of its performance. In this paper are presented the results of an analysis of work done by a three-compartment jig in treating bituminous coal from the Thick Freeport bed in Allegheny County, Pa. Certain systems of washing coal require sizing before washing, while others treat unsized coal. The prevailing practice, however, for the preparation of coking coal from the more easily washed coal is to treat an unsized or natural feed of about — 1 in. size and smaller. Under such conditions, the minimum size of particle that is beneficiated has long been and is still a matter of discussion and conjecture. In a careful laboratory investigation J?ngstl concluded that the practical limit of settling of fine coal is reached at from 0.02 to 0.008 in. On the other hand, Hancock2 has found that single-compartment jigs, used in preparing coal for the domestic market, usually allow most of the heavy (refuse) particles smaller than 1/4 in. to go into the washed coal. He also states that it is impossible with the ordinary coal jig to make any noticeable improvement in the quality of the coal finer than 20 mesh (about 1/30 in.). Because of its importance in coal cleaning, the work described in this paper was undertaken for the purpose of obtaining information on this subject, as a part of the general coal beneficiation program being carried out by the Bureau of Mines.

Jan 1, 1925

By Edwin C. Holden

In making the original estimates for a report recommending the cyanidation of Ymir stamp-mill tailings, the writer was unable to find in any of the standard works on cyanide-practice certain data applicable to the local conditions; nor has he as yet seen quite the same practice recorded in the rapidly growing literature of cyanidation. To present these figures, together with a few novel features of local practice, the present paper is offered. The ore treated is obtained from a fissure-vein of steep dip, occurring in a slate country. The ore-body is lenticular in plan, swelling from 4 ft. in width at one end to a maximum of 42 ft., and tapering irregularly to less than 18 in., where payvalues cease. In the wider portions of the fissure the ore is mixed with more or less horse-matter from the walls. The clean ore is white, massive quartz, with from 8 to 12 per cent of sulphurets; the latter being pyrite, sphalerite and galena, the predominance being in the order given. The precious metal tenor is variable, the gold more closely following the pyrite, the silver the galena. The present mill-supply is practically all from below the oxidized zone. The unavoidable admixture of horse-matter and stope-filling with the ore as mined, affects more or less the character of the mill-feed, which sometimes contains from 10 to 25 per cent of slate. During the last three months of 1902, the mill-feed carried 0.3977 oz. gold and 1.903 oz. silver per ton of 2,000 lb., 2.65 per cent lead and 2.92 per cent zinc. The iron, unfortunately, was not determined. The Ymir mill and mine have been fully described elsewhere;* suffice it to say of the milling that the crushed ore,

Jan 1, 1904

By F. L. Grammer

The analyses given in Table I. were made several years ago at my request at a plant using Lake ores. They are of two furnaces, one making basic, the other Bessemer pig— they gave 6 casts each in 24 hr., also two flushes of cinder between each cast. They are offered simply for the sake of reference as, so far as I know, it is unusual to find so frequent full analyses made and none are on record. The variations in manganese-content in the first and second flushes on the basic furnace are interesting, if no more than a freak—the first flush being invariably higher in manganese. The iron-content of the two slags from an average of 12 flushes shows: Basic. Bessemer. Per Cent. Per Cent. Total iron...................................... 2.85 1.70 Combined iron................................ 0.72 0.52 Free iron....................................... 2.13 1.18 The greater iron-loss in the basic slag, while known, has never heretofore been so clearly set forth. This may modify the arrangement for handling slag; heretofore usually determined by proximity of the dumping-heap. If a nearby heap is used on contiguous territory, in the past, tilting molten-slag pots were used. If the Pittsburg flood-records for 30 years are a fair criterion the government regulations were inadequate, and molten slag was dumped in such a manner as to throttle the Monongahela at the time of the spring thaws. At Pittsburg the slag, where it is carried to a distance, is frequently granulated in pits and then removed in cars to which it is conveyed by clam-shell buckets. This practice is also quite customary where the slag is intended for the manufacture of slag cement. The production of 500 tons of cast-iron daily means about 300 tons of- slag and at 2.113 per cent. equals 6.4 tons

Jan 1, 1914

By J. A. Roberson, A. M. Adair

A plate of an 18 pct Ni (250 grade) maraging steel was homogenized, cold rolled, and re crystallized to produce a fine grained, homogeneous structure. The strength properties were then measured as a function of cold rollinq prior to aging. Material in the solution heat treated condition exhibited substantial strengthening from cold rolling. The strength increase and the rate of strengthening were proportional to the amount of cold work. This benefit, of up to 50 ksi, was retained during aging for 3 hr at 900°F. The structures produced by cold rolling and aging have been systematically investigated by transmission electron microscopy. The dislocation debris produced by cold work influences the morphology and distribution of precipitate particles as well as the rate of precipitation. An essential feature of the high nickel maraging steels is a martensitic transformation from y to a with the resultant production of a supersaturated bcc solid solution. This solution decomposes to form precipitates during a subsequent aging treatment.' The martensitic transformation in low carbon Fe-Ni alloys and the mechanical properties of these alloys have been studied extensively.2"4 The carbon content of maraging steel is kept very low to avoid the embrittling Fe3C phase and to minimize the acicular, or twinned martensite common to high carbon steels. Maraging steels are usually strengthened only by heat treatment, but cold working is also quite effective in this respect as was shown by Decker, Eash, and Goldman,' Kula and Hickey,5 and the present work. There is, however, some disagreement between the present results and those of Kula and Hickey. Because of their unusual combination of strength and toughness, the maraging steels are suitable for applications where reliability at high stress levels is required. The strength level is achieved through a combination of solid solution hardening,' substructure hardening,3 and precipitation hardening,6 of which the last is the most important. The toughness of these steels has been reported to be a result of the absence of brittle grain boundary precipitates,7 and the fact that the matrix is a massive martensite rather than a twinned martensite.' One of the problems in the manufacture and use of maraging steel has been its tendency to develop laminations in sheets parallel to the rolling plane.9,10 The present investigation was intended to determine whether or not an improvement in the microstructure of this steel by thermomechani-cal processing would improve mechanical properties. I) EXPERIMENTAL PROCEDURES A plate ($ in. thick) of an 18 pct Ni-250 grade maraging steel* was obtained and examined metallo- "Analysis, wt pct: 18.1 Ni, 8.6 Co, 5.3 Mo, 0.4 Ti, 0.07 Mn, 0.05 Cr, 0.05 Si, 0.002B. 0.1 Cu. 0.05 Al. 0.002 Zr, 0.016 C, balance Fe. graphically. It was found to contain laminations as described by previous investigators. This plate was then annealed at 2100°F for 24 hr to promote homogeneity. Pilot studies on small coupons had shown that this was the minimum temperature required to accomplish homogenization in 24 hr. Since this treatment caused excessive grain growth, the steel was subsequently cold rolled and recrystallized. Cold rolling was accomplished at about 5 pct per pass to a total reduction of 66 pct. The standard solution annealing treatment (1 hr at 1500°F) was adequate to cause recrystallization of the cold worked steel. The recrystallized steel was then cold rolled and ground by various amounts to produce specimens with uniform thickness and selected amounts of cold rolling. All tensile specimens with less than 90 pct cold reduction were ground to a thickness of 0.050 in. Specimens with 90 pct cold reduction were 0.025 in. thick. All

Jan 1, 1970

By S. R. Maloof, W. R. Mitchell, J. A. Mullendore

Preliminary experimental evidence is presented to show that the metallic impurities aluminum, iron, and silicon, and beryllium oxide as found in commercially pure hot-pressed beryllium powder can be reduced to lower Concentrations by Zone -Purification techniques. The reduction in the concentration of aluminum to extremely low levels (10 PPm) is noteworthy, since earlier work demonstrated that aluminum is a major factor contributing to the hot-tearing of beryllium during fusion welding. On the basis of present findings, a method is suggested for producing beryllium metal of improved weldability. EXPERIMENTAL research now being conducted on the purification of beryllium indicates that several of the metallic impurities and beryllium oxide can be reduced to lower concentrations by employing standard horizontal zone-refining techniques. For the preliminary work, commercially-pure hot-pressed beryllium having a purity of 97.84 pct was used. The chemical composition is given in Table I. The hot-pressed beryllium was machined into bars 5/8 by 5/8 by 4 to 6 in. in length. Beryllium-oxide boats which had been conditioned by prior heating in a vacuum of 10"5 mm of Hg at 1100°C for 3 hr were used to hold the beryllium while it was zone melted. The boat with its charge, along with a smaller beryllium-oxide boat containing scrap beryllium to be used for gettering purposes, was loaded into a quartz tube and evacuated to 10-6 mm of Hg. Alter alternately evacuating and flushing three times with purified argon, the furnace tube was backfilled to a pressure of 1/6 atm of A and sealed off by means of a vacuum-tight stopcock. It was then removed from the pumping equipment and mounted on the traveling carriage of the zone-refining apparatus. Heating was accomplished with a 10 kw 450 kc induc- tion heating unit using an induction coil consisting of two turns of 3/16-in. diam copper tubing. Before starting to zone melt, the atmosphere in the furnace tube was gettered by heating the scrap beryllium in the smaller boat to above its melting point and holding at temperature for several minutes. The carriage was then shifted so as to locate one end of the beryllium ingot in the large boat within the induction coil. A molten zone approximately 3/4 to 1 in. was established by loading directly into the ingot. The molten zone was moved very slowly along the length of the ingot. Before each pass was made the procedure for evacuating, flushing, and backfilling, and so forth was repeatedI' Speeds of 1 1/2 and 3 in- per hr were used with up to twelve zone passes. The appearance of a typical zone-refined ingot is shown in Fig. 1. After zone refining, the grains in the ingot appear to be elongated in the direction of zone travel, whilst before, they are essentially equi-axed. A microexamination revealed the presence of a network in the beryllium grains as seen in Fig. 2. Laue spots in back reflection pictures taken of the sample were not sharp, indicating the presence of subgrains in the structure. As can be seen, however, many of the network boundaries cross beryllium grain boundaries. It has been shown by Martin and Moore1 that beryllium solidifies as a bcc structure which then transforms to the hexagonal structure at about 1250oC.It is believed that the observed network is merely the outline of the bcc structure which formed from the melt and are not subgrain boundaries. In order to eliminate surface contaminants, approximately 1/16 of an inch of material was removed from the outer skin of each ingot after zone refining.

Jan 1, 1962

By R. J. Sherman, M. R. Achter

Creep rates have been measured for nickel at 1200°F in air and in vacuum, and related to the depth of surface cracks in the specimens tested in the two environments. The surface cracks were observed to grow faster in air than in vacuum in both short-time and long-time tests. Crack depths, however, could be correlated with creep rates only in the short-time tests. Although no cracks were observed in a relatively brittle Ni-Cr-A1 alloy creep tested at 1300°F, the appearance of alloy specimens fractured after interruption in creep indicated that there was a decrease of intergranular cohesion resulting from the diffusion of atmospheric gases into grain boundaries. PREVIOUS investigations1-3 have demonstrated that the effect of atmosphere on the high-temperature strength of nickel and nickel-base alloys may be reversed by changes in the stress or temperature of the test. At high stress and low temperature the creep-rupture life is longer in vacuum than in air, while at high temperature and low stress the relative strengths are reversed. A mechanism to explain this behavior was proposed based on two competing processes. Oxidation hardens and strengthens a metal, but adsorption of gas, by lowering surface energy, reduces the work required to propagate an intergranular crack. The controlling process is determined by temperature and stress. Basic to this mechanism is the concept that environment determines the ease of propagation of intergranular cracks and thus affects creep rate and rupture life. To explore the mechanism in greater detail, this investigation was undertaken to determine the effect of atmosphere on crack initiation and propagation rates in nickel1 and a relatively brittle Ni-Cr-A1 alloy.3 EXPERIMENTAL PROCEDURE The materials were the same as used previously. Nickel of 99.8 pct purity had the following impurity content: 0.02 pct Co, 0.03 pct Fe, 0.02 pct Mg, 0.03 pct Si, and 0.008 pct C. Rods were cold swaged from 1/2 in. to 3/8 in. in diam and machined into speci- mens with gage sections 1/4 in. in diam by 1-1/4 in. long. They were annealed in vacuum at 1600" F for 1 hr to achieve a grain size of 0.07 mm. The alloy was composed of 19.1 pct Cr, 3.6 pct Al, 1.7 pct Si, 0.57 pct Fe, 0.10 pct Mg, 0.027 pct C, 0.004 pct S, and the balance Ni. Sheet specimens of 0.500-in. width and 1.25-in. gage length were machined from 0.033-in.-thick strip and annealed in vacuum at 2000° F for 2 hr, resulting in a grain size of 0.036 mm. Large amounts of an intergranular phase were introduced as a result of the anneal. Creep tests on both materials, using equipment previously described by Shahinian; were conducted in air and in a vacuum of 1 x 10-5 mm Hg. They were continued to rupture and interrupted at various stages of the test. Although extensions were measured from dial gage readings of movement of the lever arm, it is considered that valid comparisons of creep rate may be made in this manner.l Two parameters were measured metallograph-ically* on longitudinal sections of nickel specimens interrupted during creep—the mean depth of surface cracks 2, and the density n as the number of cracks per unit length of surface. They were combined into one parameter by the relation p = n-d [1] As used here, p is equal to the total of crack depths per unit length of surface. RESULTS Nickel. It is well known that the incidence of intergranular cracking is increased with increasing rupture life. Indeed, it is evident in Tables I and I1

Jan 1, 1962

R. Culver, Colorado School of Mines I appreciated Professor Starfield's presentation, because his approach to blasting research reflects my feelings that a lot of practical work can be done toward making short-term advances that can be used by the blasting industry. We should not forget this in our attempt to obtain solutions to such idealized problems as a mathematical theory of cratering. The blasting industry would like very much to have the efficiency of the basic blasting processes improved from the range of 20 percent up into the range of 50 percent. And many of these problems should be approached and attacked in the laboratories. Blasting research at the Colorado School of Mines for the past two or three years has been aimed at improving the pre-splitting process, because we feel it will become increasingly important to the construction industry. We have attempted to develop a unified theory of how pre-splitting works; and the more we investigate, the more variables we find. We may never obtain an entirely satisfactory theory of pre-splitting, but we have developed a tentative explanation. One of our first conclusions was that we are making very poor use of the explosive in current pre-splitting practice. But to appreciate this we must first consider how pre-splitting works. A series of parallel boreholes are loaded with low density charges which are detonated simultaneously. (Delayed detonation has been tried, but most people agree that simultaneous detonation is best, and this is what we found.) Stress waves traveling out from each borehole interact with the stress waves from the two adjacent boreholes influencing the direction of the fractures that occur and propagate from each borehole. It has been suggested in private communication with Mr. Wilbur Duvall of the U. S. Bureau of Mines, that the fracture between the adjacent boreholes is a result of the net particle velocities above and below the pre-split plane which tend to set up a net tension along the pre-split plane. Regardless of what the exact mechanism is that causes the fracture to propagate from one borehole to the next, we know that it, in general, relies upon an interaction of the stress waves. However, if there is a large compressive stress normal to the pre-split place, it can practically eliminate any possibility of obtaining the desired pre-split by preventing the desired fractures from getting started. Aside from obtaining a fracture along the pre-split plane, the other major aim of pre-splitting is to minimize the fracturing that occurs around the borehole. We do this by drilling a large hole and putting a small charge in the middle of it. This results in a badly decoupled charge which is very inefficient; most of the energy is used in compressing the air in the borehole. Out of discussions about how pre-splitting works, we decided that perhaps there was some other explosive geometry which would make more efficient use of the explosive, and result in optimizing the pre-splitting process. Starting in the laboratory with small plexiglas blocks, we found that a fully-coupled explosive charge of 5 grains/in. of PETN gave random fractures extending about I in. from the borehole. A 1/4 in. plexiglas plug fit into a 1/4 in. hole in a similar block with the same fully coupled charge gave random fractures extending out only about 1/2 in. due to the energy absorbed in fracturing the plug. If the plug was split longitudinally but otherwise using the same geometry, it was possible to drive a fracture about 2 in. in either direction from the borehole, with an even smaller amount of fracturing around the borehole.

Jan 1, 1967

By Donald E. Garrett

The processing of brines from Searles Lake, California, yields one of the world's largest supplies of borax and boron chemicals, second probably only to the massive deposit at Boron, California. Two companies produce from this deposit, the American Potash & Chemical Corp. and the West End Chemical Co., a division of the Stauffer Chemical Co. Two processes are employed in recovering borax from the brines; one is used by both companies and consists of carbonating the brine to form sodium bicarbonate, removing this, and then cooling the liquor to crystallize a crop of borax. Only American Potash uses the second process, evaporating the brine to concentrate it and remove (crystallize) much of its NaCl, Na2CO3, Na2SO4, and NaLi2PO4 content, quick cooling the remaining liquor to remove KCI, and then seeding the filtrate to remove borax. Both methods will be briefly described. Searles Lake is a large evaporite body formed during the most recent two Sierra Nevada ice stages, the Tahoe and the Tioga. Runoff waters from the Sierra Nevada Mountains collected in the large Owens Valley drainage basin and from there overflowed into the Indian Wells basin and then into Searles Lake. In each Lake, evaporation took place. To a large extent the Owens and Indian Wells lakes acted as settling basins, for much of the mud, and insoluble calcium and magnesium precipitates were deposited there. Recent volcanic activity in the mountains to the north and east of Owens Lake had scattered considerable mineral-rich volcanic ash and debris throughout the drainage area, and several hot springs were formed that contained boron and other salts. The present flow rate and concentration of several springs, and the known age of at least the Tioga zone, can fairly well account for all of the boron present in Owens and Searles Lakes.1 The tungsten content can be accounted for in a similar manner.2 Most of the salts thus originally present in the fresh water runoff-from ice, snow, and rain-accompanying these two ice stages were ultimately deposited in Searles Lake. During the Tahoe period, from 25 to 40 ft of massive, mixed salts were deposited, along with six interspersed layers of mud (four inches to two feet thick) representing comparatively brief interglacial periods. Before the final desiccation of the lake, from 10 to 14 ft of new mud was deposited, representing the Tioga stage. The crystals subsequently deposited were from 70 to 90 ft thick, and form the present lake surface. The extent of the salt body is about 34 square miles, of which 12 square miles are exposed salt; the remaining area is covered by recent alluvial deposits (chiefly mud). The present lake is generally dry on the surface but floods to a shallow depth following heavy rains. The crystal body of both zones consists of continuous clusters of salts, and the voids between crystals (about 45 pct of the total volume) are occupied by brine. The level of this brine never recedes more than several inches below the salt surface. The brine pumped to the plants for processing is generally saturated with all of the solid phases present in that area, and thus is a complex mixture containing about 35 pct solids (see Table 3). The brine composition varies somewhat at different areas of the lake and from upper to lower structures, so wells are drilled in the areas and to the depths desired. A casing is cemented in the hole to within a short distance from the bottom, and pumps with about 14-ft suctions are mounted on the casing to pump the brine into the plants. The evaporation process of the American Potash & Chemical Corp. treats upper structure brine, evaporating all of the water from

Jan 1, 1960

By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958