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By C. R. Kuzell

GEOGRAPHICAILY the industry of reducing and refining of copper continued to migrate from the .United States during 1931. While this country is losing the predominant position of its copper industry, our engineers still enjoy their pre-eminent position as leaders in the design and operation of the new plants being brought into production in the four quarters of the globe. Progress in the art, as indicated by the construction of the recently completed works, has not been revolutionary. Standard methods have been followed, augmented by novel features in flow-sheets demanded by the results of progress in the art of ore-dressing, and by reason of problems incidental to the peculiarities of the new ores being treated. Design and construction progress has likewise been orderly and without radical departure from less recent practice; new features in primary and secondary equipment being logical developments to provide for special conditions to keep abreast of improve¬ments in accessory apparatus. The new plants are being rapidly brought in without hazardous novelties and for the purpose of yielding the principal ingredient, copper, in as large a volume as it will he possible to market; the old United States procedure. In the meantime the domestic plants are shut down or operating at a very low capacity, threatened with untimely demise unless the metallurgist can extract and develop other valuable products from the ores. International agreements .and tariff protection may be a temporary reviving stimulant, but for many mines the days of profit will be numbered unless the ores can be made to yield as values the ingredients now being discarded. Invisible progress in this direction has been made during the past year, and it is to he hoped that time will prove that the outstanding progress of 1931 was in the development of processes for the ultimate recovery of by-products from copper ores, perhaps even relegating the value of the copper yield to secondary importance.

Jan 1, 1932

By G. C. Wallick

Since the appearance of the paper, "Solution of the Equations of Un-steady State Two-Phase Flow in Oil Reservoirs," by W. J. West, W. W. Garvin, and J. W. Sheldon,' a two-fold investigation of this subject has been carried out. One objective of the investigation has been to deter-mine the feasibility of solving such problems 7on a medium-size com-puter such as the Datatron*, and the other objective has been to in-vestigate the application of such cal-culations to experimental and theo-retical petroleum reservoir research. In the first Datatron calculations, the fluid and rock properties published by West, et al, were used, together with the published equations describ-ing the system. Details of the formu-lation not given in the original paper are discussed in the Appendix to this note. Reference should be made to the subject paper for the complete equations and defintion of symbols. An unexpected result of this in-vestigation was the discovery that the linear solution published by West was in error. Thus, in addition to describing the Datatron solutions and to discussing certain numerical diffi-culties which will be encountered if one uses the published method of solution, the purpose of this note is to indicate the nature of this error. LINEAR FLOW Since the linear case requires a minimum amount of scaling, a fixed-decimal point Datatron program was written for the one-dimensional flow problem and an attempt was made to duplicate the solution described by West. In the case described, fluid was produced at a constant rate, Q, until such time as well pressure reached 0.04. Production was then continued at constant pressure. From the constants and curves given by West it was determined that the ini-tial constant production rate could be approximated by Q = 0.007. An ini-tial dimensionless time step ?t = 0.434 X 10 - "as used, and each suc-cessive time step was doubled until a value of = 0.444 was reached. This constant interval was then used for the remainder of the solution. In subsequent solutions, several varia-tions in the time schedule were em-ployed, including smaller time steps and slower rates of increase in the time steps. In all cases, almost identical results were obtained regardless of the time schedule employed. However, as described below, it was noted that the time schedule had some influence on the rate of convergence of the solutions. As a check on the accuracy of the solution, the cumulative production at each time step was calculated using the two methods described in the Appendix. Satisfactory agreement was observed with the differences in these two values of the order of two parts in 50,000. It should be noted that the mass balance check as described is of questionable value, particularly with regard to the well pressure and saturation. This is especially true in the radial solution where pressure and saturation values near the wellbore would make only a negligible contribution to the numer-ical integration. It is believed, how-ever, that such a comparison is ot value in determining the over-all accuracy of a solution. In comparing the Datatron solu-tion with that published by West it was discovered that in the later stages of depletion, the pressures near the well declined more rapidly in our solution than in the West solution, and that the limiting well pressure of 0.04 was reached at an earlier time than that originally reported. It thus became evident that it would be impossible to duplicate the production schedule described by West and a constant rate of production was maintained until the well pressure was equal to 0.0. A representative comparison of the results published by West with those obtained in this investigation is shown in Fig. 1, which is a plot of GOR as a function of cumulative recovery. These two curves should be in agreement until a cumulative recovery is reached which corresponds to a well pressure of 0.04 — for the Datatron solution, a recovery of approximately 5.6 per cent. Actually, a major disagreement is evident. Subsequent correspondence

Jan 1, 1958

By Stuart Croasdale

Not long ago I was called upon to conduct some experiments on the treatment of ores from the New Cornelia copper mine, Ajo mountains, Arizona, for the Calumet & Arizona Copper Co. The problem was a very interesting one and contained some unusual features. Incidental with this problem was the utilization of a large amount of low-grade pyrite in the Bisbee mines of the company, provided this could be accomplished by any feasible method. The New Cornelia mine is situated 45 miles south of Gila Bend, the nearest railroad point, on the Southern Pacific Railroad. It is about 160 miles, almost due west, from Tucson. This section is one of the most arid in the United States. The normal annual precipitation, as given by the United States Weather Bureau covering a period from 1870 to 1901, is 3 to 5 in. During June and July, 1913, the records kept by the company gave the following results. These are all I have available, but they show the enormous evaporation that can be expected, which may be a factor in leaching operations on a large scale. Temperature, Degrees F. High Low Average June...................... 112 62 84.0 ' July...................... 115 71 87.8 Evaporation June July Concrete tank, inches............ 11.7 10.72 Iron tank, inches.............. 14.8 13.07 Rainfall, inches................ None 0.77 Cloudy days................... None 10.00 Aside from several springs about 10 miles from the property, the nearest known water in quantity at the time these experiments were made was the Gila river at Gila Bend, 45 miles distant. Part of the developinent of the property of course included drilling for water in the valley, which was in progress at the conclusion of my experiments. The ore deposit itself is unusual. It consists of of intrusive plug of monzonite or granite porphyry. Chalcopyrite, bornite, and magnetite are finely disseminated throughout the magma in sufficient quantities to make the greater part of the mass commercial ore to a certain depth.

Jan 1, 1915

By Eugene Guccione

It would be extremely difficult and unfair to compare the United States to Australia when wondering about which of the two countries offers the best opportunities for running a mining operation. However, if somebody would ask "Where is the best place to mine outside the US?"-I'd answer without reservations: Australia. Some of my Australian mining friends will disagree with me for my giving their country such a high rating. And they'll probably accuse me of being an optimist by telling me, as one of them did, that "the difference between an optimist and a pessimist is that the pessimist is better informed."

Jan 1, 1977

By T. E. Kihlgren, N. B. Pilling, E. M. Wise

Systematic data are presented on the relation of composition of nickel silvers to color, tarnish resistance, hardness and liquidus temperatures, for alloys containing up to 30 per cent nickel and 50 per cent zinc, further modified at certain nickel and zinc contents by 0 to 8 per cent of tin and 0 to 10 per cent of lead. Attention has also been directed, using a still more limited range of alloys, toward a study of tensile properties and such casting properties and characteristics as fluidity, shrinkage and pressure tightness. In some of these, rather broad ranges of composition are dealt with; in others, it has been expedient to focus attention on the Federal Specification 20 per cent nickel silver casting alloy, WWP-541. Introduction The demand for an alloy essentially white in color, which could be cast and worked by simple methods and which possessed strength and corrosion-resisting properties superior to the usual brasses, led to the discovery and early use of cupronickel alloys containing zinc, and sometimes tin, lead and other metals. This type of alloy was generally known as 'German silver until the war, when the term "Nickel Silver" became generally accepted. The element principally responsible for the white color of the alloy and for its strength and corrosion resistance is nickel, which may constitute from 8 to 30 per cent of the alloy. The functions of the other elements will be described later. The manifold uses of nickel silver are largely due to its color, corrosion resistance, ease of casting, fabrication, joining and rather moderate cost. In many instances the alloy is finished by grinding and polishing while in others it may be finished with a silver or nickel plus chromium plate. Electroplates applied to a nickel silver base behave more satisfactorily than those applied to the softer and less resistant alloys, and owing to the white color of the underlying metal the article will retain a pleasing appearance despite local failures of the plating.

Jan 1, 1935

By E. M. Wise, T. E. Kihlgren, N. B. Pilling

Systematic data are presented on the relation of composition of nickel silvers to color, tarnish resistance, hardness and liquidus temperatures, for alloys containing up to 30 per cent nickel and 50 per cent zinc, further modified at certain nickel and zinc contents by 0 to 8 per cent of tin and 0 to 10 per cent of lead. Attention has also been directed, using a still more limited range of alloys, toward a study of tensile properties and such casting properties and characteristics as fluidity, shrinkage and pressure tightness. In some of these, rather broad ranges of composition are dealt with; in others, it has been expedient to focus attention on the Federal Specification 20 per cent nickel silver casting alloy, WWP-541. Introduction The demand for an alloy essentially white in color, which could be cast and worked by simple methods and which possessed strength and corrosion-resisting properties superior to the usual brasses, led to the discovery and early use of cupronickel alloys containing zinc, and sometimes tin, lead and other metals. This type of alloy was generally known as 'German silver until the war, when the term "Nickel Silver" became generally accepted. The element principally responsible for the white color of the alloy and for its strength and corrosion resistance is nickel, which may constitute from 8 to 30 per cent of the alloy. The functions of the other elements will be described later. The manifold uses of nickel silver are largely due to its color, corrosion resistance, ease of casting, fabrication, joining and rather moderate cost. In many instances the alloy is finished by grinding and polishing while in others it may be finished with a silver or nickel plus chromium plate. Electroplates applied to a nickel silver base behave more satisfactorily than those applied to the softer and less resistant alloys, and owing to the white color of the underlying metal the article will retain a pleasing appearance despite local failures of the plating.

Jan 1, 1935

By E. H. Dix

The success of the Charpy impact test in the steel industry has led those interested in aluminum alloys to investigate the possible application of this test to aluminum and its alloys. In this paper are set forth the results of a few preliminary tests made at the Lynite Laboratories of the Aluminum Manufactures, Inc. In order to study more thoroughly the significance of the Charpy test, slow-bending tests have been made duplicating the arrangement of the impact machine. Tests have been made on both notched and unnotched specimens of a series of copper-aluminum alloys. Apparatus The standard American-made Charpy impact machine delivering a blow of over 200 ft.-lb. was much too large to test aluminum-alloy specimens of the same type used for steel, so that it was necessary either to adopt a larger specimen or to have a special light-weight hammer made. Since there is no known relation between specimens of different sixes, even though geometrically similar, it was decided to adhere to the 10 by 10 mm. specimen. Accordingly, Sauveur and Boylston, the American makers of the machine, were asked to construct a special light-weight pendulum adapted to their standard machine, this pendulum to have one-tenth the capacity of the standard. The writer wishes here to acknowledge his indebtedness to Mr. H. M. Boylston for the personal interest and care exercised in constructing this pendulum. The machine with the light-weight hammer in place is shown in Fig. I. It is mounted on a concrete foundation with the concrete covering the flange on the base of the machine. Vibration is further prevented by four 5/8-in. (15.8 mm.) foundation bolts. The large standard hammer shown against the wall was also used for some of the unnotched bars that required more than 20 ft.-lb. to rupture. When the light-weight hammer was received, it was carefully cali-

Jan 1, 1921

By American Institute of Mining Engineers

David J. Mack (University of Wisconsin, Madison, Wise.)—Have the authors considered that the effects reported in their paper are readily explainable on the basis of equilibrium grain boundary segregation rather than ordering? It seems to the writer that all of their experiments and observations correlate well with known behavior associated with equilibrium grain boundary segregation. Although at first glance one would tend to dismiss such a possibility because of the similarity of the gold and copper atoms, their sizes differ by 12.8 pct which is in the range where segregation at grain boundaries would be expected. Then, too, the electrochemical differences between the copper, the Cu3Au, and the gold strongly suggest that a segregation mechanism is involved in the corrosion of the Cu-Au alloys. R. Bakish and W. D. Robertson (authors' reply)—The authors would like to thank Dr. Mack for his consideration of the problem of grain boundary segregation as an explanation of the observed phenomena. We had considered the problem in the following terms: Assuming a pure binary Cu-Au alloy and an ideal solid solution, the surface excess, T2= N25 — N2, of the second component (gold) may be expressed in the following way20 1 —exp (—da/kT) where 8 is the difference in surface tension of copper and gold, and a is the surface area per atom. In the present case, 6 is positive which implies that the component in excess at the boundary is gold. Since the only chemically active component in ferric chloride is copper, it does not appear that the segregation of gold is likely to increase the boundary activity. A small concentration of a third component could profoundly affect the situation. Analysis indicated only spectrographic traces of impurities, all of which are soluble in both copper and gold. Quenching experiments, reported in the paper, also tended to eliminate significant effects attributable to a third component.

Jan 1, 1958

By E. J. KENNEDY

THE annual dinner-dance was held in the large ball room of the Commodore hotel Wednesday evening. A total of 577 were seated at the dinner, over which President Eavenson presided as chairman and toastmaster. In introducing George C. Stone, the James Douglas Medalist, W. R. Ingalls outlined the former's association with the zinc industry. He said, in part In designing and building Palmerton, Stone erected a metallurgical monument and won the title of chief engineer of the New Jersey Zinc Co., with transferal to its New York office in 1900. There lie held forth in dignity, efficiency, and with appreciation until his retirement from active service in 1929. After a few years lie was diverted from the position of chief engineer for his company to the superior one of chief metallurgist and that gave to Stone his next big chance.... This was the conception of a research department and the ability to persuade his directors to adopt it. I shall not pretend to summarize the great things that afterward developed. I shall mention only the institution of the manufacture of lithopone in this country, the charging of the Wetherill furnace with briquets, the mechanicalization of the Wetherills, the accomplishment of continuous distilling, and most recently the introduction of reflux-refining.

Jan 1, 1935

By William Johnson

IT is now recognized that several important types of reactions in metallic systems proceed by the formation of nuclei and the growth of these nuclei. The process, of freezing is a simple example of this, as Tammann pointed out years ago.1Tammann held that the rate of freezing is determined by a rate of nucleation, expressed as the number of nuclei formed per unit volume of unfrozen liquid per second, and a rate of growth of these nuclei, expressed as the linear rate of radial growth in units of length per second. For isothermal freezing the conception is simple; for ordinary freezing, extending over a range of temperature, it is not as simple, for the values of the two constants must change with change in temperature. There is ample evidence that the postulated mechanism is correct even though a quantitative derivation of the rate of isothermal or of ordinary freezing in terms of the two constants has been lacking. In recent years other reactions have been found to proceed in a similar fashion. It has been well established, particularly by Bain,2-8 that the formation of pearlite from the eutectoid decomposition of the solid solu-tion austenite proceeds in such a way, and Polanyi and Schmid, Tammann and Crone, Parnop and Sachs, and others'-1 have shown that the process of recrystallization proceeds in a similar way. Isothermal reaction rates have been determined for eutectoid decom-position by a number of investigators. Bain's2-8 work on the formation of pearlite from austenite is especially valuable in this respect-the type of isothermal reaction curve obtained, as illustrated in Fig. 13, showed an initial slow rate, accelerating to an intermediate maximum rate which then decelerated to the completion of the reaction; similar curves have been obtained by Wever and his collaborators.' Other eutectoid decompositions show similar behavior: for example, the decomposition of the beta eutectoid in the copper-aluminum system, studied by Smith

Jan 1, 1939

By Robert Ammon

THE milling practice at Mascot, at present, consists of dry crushing to 5/8 in., jigging, fine grinding, and flotation. The ore arrives at the mill from two mines, No. 1 mine shaft being located in the mill proper, and No. 2 mine shaft being situated about 2000 ft. east of the mill. No. 2 mine is served by a coarse-crushing plant and aerial tramway to the mill bins. In jigging the No. 2 mine ore, 50 per cent. of the mill feed is rejected by the jigs as a jig tailing ranging from 5/8 in. to 10 mesh; the balance of this ore, after removing a jig concentrate, is ground to flotation feed. In jigging No. 1 mine ore, approximately 80 per. cent. of the mill feed is rejected as a coarse jig tailing, and the balance is ground to flotation feed without the removal of a jig concentrate. The difference in the treatment is due to the fact that No. 2 mine ore carries a higher content of zinc than that from No. 1 mine. Of the total concentrates produced, 30 per cent. are made on jigs and 70 per cent. by flotation. At present, 1350 tons of No. 2 ore and 800 tons of No. 1 ore are treated each 24 hours. The recoveries for a typical week are as follows: Heads 3.55 Jig concentrates 61.30 Jig tails 0.80 Flotation concentrates 60.60 Flotation tails 0.09 Extraction 86.50 The associated minerals are a straw-colored sphalerite, dolomite, calcite, pyrite, quartz, and a small amount of greenockite, which occurs as a stain and sometimes small crystals along with sphalerite, but is comparatively rare. Pyrite is present in small quantities, usually less than 0.5 per cent.; quartz occurs in bands a few inches thick. A complete analysis of the ore is as follows:

Jan 9, 1924

OUTSTANDING symbol of the machine age, the steam shovel needs no introduction. Few individuals there are, in the United States at least, that have not watched with fascination the almost human motion, the more than human strength, the rhythm of the whole steam-shovel cycle of crowding, digging, swinging, dumping, swinging, and crowding again. A mass of steel with some copper and other metals, fed with water and coal, guided by a single alert and dexterous operator, or perhaps two, does the work that would require 50 to 500 men or even more. The largest shovels working in the copper mines have loaded into waiting railway cars as much as 10,000 tons of ore in a single 8-hour shift. With no equipment but hand shovels, it would, in fact, be impossible to duplicate the work with any number of men; first, because they would be in each other's way, and second, because the steam shovel can lift single pieces of rock weighing 5 or 10 tons with effortless ease. Of recent years, internal-combustion engines and electric motors have displaced steam as a motive power in many instances; and of course the detailed design of the shovels has been improved immensely. Whereas the largest dippers used 30 years ago had a capacity of 3 ½ to 4 cu. yd., today the 350-ton shovel used at Chino has an 8-yd. dipper, as has the 320-ton shovel at Chuquicamata. Incidentally, for stripping at coal mines a shovel having an 18-yd. dipper is in use. However, the principle of operation of the modern shovel, regardless of the size and type of motive power, is essentially the same as that of the earlier shovels. It is fair to say that the indispensable of open-cut mining is the steam shovel, or, to be more precise, if perhaps less picturesque, the power shovel. Somehow the electric shovel seems to lack the glamour that

Jan 1, 1933

By Edwin Shorey

IN THE southwestern Wisconsin zinc district, lying in Lafayette, Grant, and Iowa Counties, and the adjoining territory in northwestern Illinois, the orebodies are generally small, and are mined out in a few years. Many of the mines are developed with an expectation of five years or less of actual production. For this reason, as well as for considerations of economy, certain features in the development practice, which have been worked out with a view toward keeping the ratio of time required for, development to that occupied by actual production as small as possible, are of interest. ORGANIZATION The organization of a crew for, this work is as follows: Inasmuch as speed in sinking is desirable, the shaft crew comprises three shifts working 8 hr. each. In the smaller shafts, running up to 6 by 12 ft. (1.8 by 3.6 in.) in size, two men are employed on a shift; in larger shafts an additional man is usually employed. A mine foreman, or ground boss, has general control of the work of the three shifts. He is usually with the crew on the first and the early part of the second shifts, and plans the work to be done for the third shift when it comes on. His schedule, of course, is flexible and when pump changes, or other special conditions warrant, he, arranges his working time to suit the occasion. The surface crew for a sinking operation is made up of two hoistermen, each working 12-hr. shifts. Since the various operations in the shaft must be conducted in succession,. and not in combination, one with the other, separate attendance for hoisting and general plant is unnecessary. Drilling is done when no hoisting is necessary and vice versa. The hoisterman trams out and dumps the excavated material which the crew in the shaft is mucking, as a part of his regular hoisting work.

Jan 8, 1920

By C. P. Proctor, J. T. Crawford, John Griffen

TROUGH washers were among the earliest methods used for concentrating ores; they are referred to by Agricola about the middle of the sixteenth century as already being used while the hand- operated jigging sieve had only recently come into use. For cleaning coal the trough washer apparently was first used in France and Belgium about 184.1. The early trough washers were intermittent and the refuse at least was shoveled out by hand. The invention of a continuous jig washer, by a Frenchman, Berard, in 1848, acted as a stimulus to coal washings in jigs and retarded progress in trough washing in the nineteenth century. Chapman and Mott 1 describe a number of trough washers developed in England and on the Continent from 1850 to 1900, such as the Bell and Ramsay, Ramsay, Bell and Green, Wunderlich, Scaife, McLellan, Lodge, Elliott, Blackett, and Murton. The later of these removed the refuse mechanically, either intermittently or continuously, but in ways that interfered with stratification. THE RHEOLAVEUR A Belgian engineer, Antoine France, is responsible for the development of means to remove refuse materials continuously from a trough washer, while preserving stratification and preventing the loss of coal with the refuse. The methods and equipment devised, which have become known throughout the world as rheolaveur, were initiated at the St. Nicholas mine near Liége, Belgium, about 1912. This installation consisted of two units to wash 20 to 8-mill. (3/16 to 5/16-in.) and 8-mm. to 0 (5/16-in. to 0), each having only two superimposed inclined troughs. The first few "Rheo" boxes on the top trough removed refuse directly and the remaining Rheo boxes discharged impure refuse to the lower trough for rewashing. On the lower trough the first few boxes removed refuse and the following boxes removed a middling product used for boiler fuel or other purposes. The overflow of both troughs was joined to form the washed coal product.

Jan 1, 1950

By Samuel S. Wyer

Over 8,000,000 people in the United States depend on natural gas for their cooking, heating and lighting service. This service has been made possible only by the investment of large amounts of capital in transportation lines connecting the gas wells to the various communities, thus combining a hazardous mining operation with a public utility service. The effect that making these lines " common carriers" will have on the present and future service to these 8,000,000 or more people is of far reaching ethical and economic importance, and deserves careful consideration. As an introduction, it must be conceded that: 1. Natural gas, although a mineral and obtained by a mining operation—with more uncontrollable and uncertain features to cope with than exist in the mining of coal or other minerals—when served to the public becomes a public utility service. 2. The governmental power to regulate all public utilities in the interests of the public cannot be successfully disputed. 3. Mere suggestion of inconvenience to the utility is wholly irrelevant, as it cannot be considered or allowed to influence in making a defense against governmental regulation. Unfortunately it is not generally appreciated that:.. I . 1. Legislative enactments are not necessarily laws, and cannot abrogate economic laws or alter engineering facts. 2. There is a clear distinction between "common carrier7' and "public utility," although the terms are frequently confused. "A common carrier is one who undertakes for hire to transport persons or goods, or both, from place to place for all persons indifferently, the term implying indifference as to whom shall be served, and an equal readiness to serve all who apply, in the order of their application." On the other hand, property becomes a "public utility" only "when for a particular purpose the public have a right to resort to the premises and make use of them." It is, therefore, important to note that not all public utilities can be regarded as common carriers. The U. S. Supreme Court has already

Jan 1, 1915

By Ross Blake

THE St. Joseph Lead Co. became interested in the Tri-State district in 1921 through acquisition of prospecting and development rights on approximately 20,000 acres of land extending northeastward from the Picher mining field, mostly in Kansas. Prospecting with churn drills was begun the same year and the first commercial ore body was found on the Jarrett tract at the then northern limit of the main Picher field. Operations at this mine began in 1923. Kansas Explorations, Inc., was organized in 1923 by the St. Joseph Lead Co. as a wholly-owned subsidiary to carry on its operations in the Tri-State District from an office at Joplin, Mo. As prospecting of the original acreage proved largely negative, attention was subsequently turned to other sections of the district. Numerous additional leases were acquired and prospected. The Company was successful in developing into operating properties the Jarrett, Isherwood, Robinson, Sullivan and Lotson tracts and, in more recent years, the Ritz, Snapp, Buckingham, Gibson and Jasper mines.

Jan 1, 1947

By Glenn A. Keep

THIS paper is not an announcement of the successful conclusion of the Cerro de Pasco Copper Corporation's pacos-pyrite problems, but merely a description of the commercial-scale, intermittent-roasting operations conducted by the writer, J. H. Lucas and others of the staff. It was prepared by request, for presentation at the February meeting of the Institute. The results obtained in Peru confirm the 1-ton-a-day experiments completed about three years ago at Kearny, N. J. by C. C. Gordon and Mr. Lucas, where it was first demonstrated that a mixture of about 80 per cent. pacos ore with 10 per cent. pyrite ore and 10 per cent. salt could be roasted intermittently in a small 1-ton-a-day Holt-Dern furnace without extraneous fuel, producing porous calcine that could be leached rapidly. Mr. Keep and his staff, by successfully introducing improvements in the construction of the commercial furnace and in their handling of material, have established beyond doubt a low cost for roasting. The other steps in the process of treating pacos and pyrite ores have not as yet been demonstrated as sufficiently conclusive to warrant their acceptance for a permanent plant. Cerro de Pasco Ores The ores of the Cerro de Pasco Copper Corpn. at Cerro de Pasco, Peru, are situated at an altitude of 14,000 ft. The town and mines are served by the Cerro de Pasco Railroad, which operates between Cerro de Pasco and Oroya, the site of the company's copper and lead smelters. Table 1.—Average Analysis of Cerro de Pasco Ores Pacos Ore Pykite Ore Silver, oz............................................... 10.3 12.0 Gold, oz................................................ 0.007 0.01 Lead, per cent........................................... 1.7 1.8 Copper, per cent......................................... 0.19 0.27 Zinc, per. cent........................................... 1.4 1.02 Iron, per cent........................................... 13.4 34.45 Silica, percent.......................................... 61.8 24.5 Alumina, per cent....................................... 2.5 0.62 Lime, per cent........................................... 1.5 0.20 Manganese, per cent..................................... 0.42 0.09 Sulfur, per cent,.......................................... 1.1 34.2

Jan 1, 1929

Jan 1, 1943

By C. A. Eligwe, F. W. DeVries, A. E. Torma

This study is an investigation on using hydrogen peroxide for tank-leaching of uranium with sulfuric acid. The optimum sulfuric acid concentration was found to be 0.03 mole 11 -'for a 25% pulp density suspension. The order of reaction with respect to sulfuric acid concentration was found to be ns0.75. The best results vith hydrogen peroxide were obtained at a 4.64 molar ratio of 6202 to U in the ore. The order of reaction was found to be zero with respect to hydrogen peroxide concentration. The theoretical maximum rate effect of Hz02 on uranium extraction was derived as Vm = 8.15 x mole &-'min-'. The highest rate experimentally attained was 8.09 x mole kL'min-'. It was found that the uranium extrac- tion was not affected by the pulp density up to the 50% used in this study. However, the final yieldswere significantly higher (over 90%) for leaches carried out in the presence of Hz02 than (less than 90%) for leaches carried out in the absence of H202. The reaction was found to follow a first order course with respect to substrate concentration (pulp density) regardless of the presence of H202. The effect of temperature was more explicit at moderate tempetatures (30 to 55 C) than at lower (10 C) or at higher (70 C) temperatures. At this latter temperature an almost complete extraction was realized even without H202. The apparent activation energies were found to be AEa = 2.68 kJ mole-' in the absence of H202, 0.5 kJ mole-'with Hz02 at 40 C to 70 C. and 4.72 kJ mole-' with Hz02 at 10 C to 40 C. All these activation energy values suggested that the uranium extraction process is diffusion controlled regardless of the presence of H202. Further, the change in activation energy in the presence of Hz02 indicates the possibility of change in the reaction mechanisms at lower and higher temperatures.

Jan 1, 1980

By E. S. Machlin, A. J. McEvily

HE results of an experimental investigation by McEvily and Machlin have indicated that a strong relationship exists between the processes of cross slip and fatigue. Whenever dislocations can cross onto an intersecting plane and expand thereon, as in AgCl crystals, the usual type of S-N behavior is obtained, but when such a process is difficult, as in NaCl crystals, fatigue does not occur in its usual manner, if at all. MgO, because of its crystal structure (NaCl type) and slip systems ({110}<110>) would not be expected to fail in fatigue. However, Cornet and Gorum2 have found something akin to the usual type of fatigue behavior for MgO crystals. The purpose of this note is to attempt to resolve this apparent discrepancy. In Cornet and Gorum&apos;s work as-cleaved crystals were tested under cyclic axial loading, whereas in McEvily and Machlin&apos;s work chemically-polished crystals were tested in reversed bending. Because of the difference in specimen preparation, the major effort of the present work was directed toward comparing the behavior of as-cleaved as opposed to chemically-polished specimens. The effect of type of loading seemed less important, but a check was made on this also. Eighteen bend tests and one axial load test were made. The specimens were cleaved from Norton magnorite blocks, and measured 1 by 3/16 by 1/8 in. To avoid failures in the grips, the thickness of the bend specimens was reduced from 1/8 in. near the ends to 0.04 in. in the central portion. This reduction was from one side of the crystal only. Twelve specimens were polished in orthophosphoric acid to remove the damage due to cleavage. The other six specimens were not chemically polished so as to maintain the as-cleaved condition of the crystal faces which had not been reduced. The axial-load specimen was reduced from both faces to a thickness of

Jan 1, 1962