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By E. F. La Vigne

On Aug. 10, 1884, the Retsof Mining Co. began the sinking of an 18 by 12-ft. shaft at Retsof, N. Y. Rock salt was reached in 1885, in September, at 1008 ft. below the surface. The first salt was shipped in the fall of 1885. A timber breaker was built in 1886. In July, 1895, there was a consolidation of the Retsof Mining Co., the Greigsville Salt & Mining Co., the Livonia Salt Co., and the Lehigh Salt & Mining Co. These consolidated salt companies are now known as the Retsof Mining Co., a subsidiary of the International Salt Go. Eventually all of the mines were closed except the mine of the Retsof Mining Co., at Retsof, Livingston County, New York. On July 6, 1921, about 1/2 mile south of the original shaft arid plant of the Retsof Mining Co., a new 9 by 28-ft. concrete-liried shaft was started. The shaft is 1063 ft. deep from the top of the shaft collar to the bottom of the salt bed. At the same time a reinforced-concrete breaker and other buildings were erccted. This new plant began operations on Dee. 15, 1923. It was designed for a capacity of 3000 tons in 8 hours. Geology The workable rock-salt bed is 1063 ft. below the surface and 326 ft. below sea level, and varies in thickness from 9 to 10 ft. The total thickness is greater than that but is contaminated with shale above and below the pure mineral. The characteristics of the formation are shown in Tables 1 and 2. The bed dips uniformly 0.87 per cent to the south. Prospecting As the salt deposil in New York State is known, very little prospecting has been necessary. Before the sinking of the Retsof No. 1 shaft, a well was drilled to prove the existence of salt. No. 1 shaft was sunk where the well was located. The several salt wells drilled in Livingston Courlty proved generally the existence of salt throughout the county. (See U. S. Geol. Survey Bull. 669: Salt Resources of the U. S.)

Jan 1, 1938

By E. F. La Vigne

On Aug. 10, 1884, the Retsof Mining Co. began the sinking of an 18 by 12-ft. shaft at Retsof, N. Y. Rock salt was reached in 1885, in September, at 1008 ft. below the surface. The first salt was shipped in the fall of 1885. A timber breaker was built in 1886. In July, 1895, there was a consolidation of the Retsof Mining Co., the Greigsville Salt & Mining Co., the Livonia Salt Co., and the Lehigh Salt & Mining Co. These consolidated salt companies are now known as the Retsof Mining Co., a subsidiary of the International Salt Go. Eventually all of the mines were closed except the mine of the Retsof Mining Co., at Retsof, Livingston County, New York. On July 6, 1921, about 1/2 mile south of the original shaft arid plant of the Retsof Mining Co., a new 9 by 28-ft. concrete-liried shaft was started. The shaft is 1063 ft. deep from the top of the shaft collar to the bottom of the salt bed. At the same time a reinforced-concrete breaker and other buildings were erccted. This new plant began operations on Dee. 15, 1923. It was designed for a capacity of 3000 tons in 8 hours. Geology The workable rock-salt bed is 1063 ft. below the surface and 326 ft. below sea level, and varies in thickness from 9 to 10 ft. The total thickness is greater than that but is contaminated with shale above and below the pure mineral. The characteristics of the formation are shown in Tables 1 and 2. The bed dips uniformly 0.87 per cent to the south. Prospecting As the salt deposil in New York State is known, very little prospecting has been necessary. Before the sinking of the Retsof No. 1 shaft, a well was drilled to prove the existence of salt. No. 1 shaft was sunk where the well was located. The several salt wells drilled in Livingston Courlty proved generally the existence of salt throughout the county. (See U. S. Geol. Survey Bull. 669: Salt Resources of the U. S.)

Jan 1, 1938

By James Yancik

The Howe Coal Co. is mining the Oklahoma Lower Hartshorne bed which ranges in seam thickness from 38 to 42 in. and pitches approximately 7° to the northeast. The raw feed to the cleaning plant averages about 20% ash and 1.10% sulfur, being reduced in the cleaning process to 4.8% ash and 0.9% sulfur with a free swelling index of 8 or 8%. This coal rivals the Pocohantas seam in chemistry and coking characteristics but has better fluidity than the Pocohantas seam coal. The coal is presently being prepared for the Japanese steel market in an 85-tph pilot plant designed and constructed by Roberts & Schaefer. The cleaning-plant flowsheet is discussed in detail.

Jan 1, 1972

F. W. Matthiessen, who, with E. C. Hegeler, of La Salle, was one of the creators of the zinc industry in the-United States, was born in Altona, Schleswig-Holstein, Germany, Mar. 5, 1835. He was one of a family of remarkable brothers, two of whom were long absociated in this country with the sugar and glucose industry. Mr. Matthiessen was educated at Freiberg, and came to this country in 1857 with Mr. E. C. Hegeler, after having, in company with that gentleman, visited the mining districts of Germany, Belgium, and England. They first went to Friedensville, near Bethlehem, Lehigh County, Penn., wherc attempts had already been made to produce zinc from the ore deposits found in that place. The ore was a he silicate but all attempts to extract zinc from it had failed. Mr. Matthiessen and Mr. Hegclcr were successful, but declined to invest their money in the zinc industry at Bethlehem because 01 certain local conditions. Hearing of the discovery of zinc -ore in the West, they visited both Wisconsin and Missouri and thought of settling in the coal region of East St. Louis. Finally La Salle, as being near to the Wisconsin fields and having much coal, was selected ae the site of the new smelting company, and on Dec. 24, 1858, the first shovelful of dirt was turned for the buildings of the new industry. The Civil War at first interfered with the manufacture of zinc, but afterward a lively demand for the metal arose and the future success of the institution was established. A large rolling mill was added in 1866, and in 1881 the manufacture of sulfuric acid as a byproduct was begun. Mr. Matthiessen also developed the Western Clock hlanufacturing Co. of La Salle, which manufactures the Big Ben alarm clock, from a small plant having 25 men to the present large institution which employs some 2000 workers. Mr. Matthiessen's affiliations with other metal industriefi have been numerous. But the contributions of the great zinc manufacturer to the causes of philanthropy and education have made his name known throughout the country, beyond industrial circles. He has given to the city of La Salle probably a sum total of $500,000 in various benefactions. Among the objects of his philanthropy were the La Salle-Peru Township High School, which he has made a model institution of secondary education for the entire United States; also the Tri-City Hygienic Institute of La Salle, Peru, and Oglesby, Ill., which he endowed for $200,000 This Hygienic Institute, which is unique in the United States, has a traincd staff of health officers, bacteriological laboratory, medical library, infant welfare station, milk station, free dental clinic, school nurse, and isolation hospital. Mr. Matthiessen was elected mayor of the City of La Salle three times and gave generously to the municipality of which he was the head. He also saved for the people of Illinois one of the great scenic districts of the state, Deer Park, with its cafions and natural beauties, which he con-verted into a model scenic resort, and the proceeds of which he gave to the charities of the Tri-Cities in which he lived. He died on Feb. 11, 1918.

Jan 1, 1920

By S. Frederick Ravitz

IN most applied sciences there are two distinct methods of carrying out research and development work. One of these, the theoretical, attempts to solve problems that may arise and to predict facts of practical importance by the application of the various laws of science, the other, the experimental method, attempts to each the 'same goal by obtaining data and trying out various processes in the laboratory. Since individuals frequently find themselves better adapted for one or the other type of procedure, it is natural that -two groups should 'arise, both working toward the sane objective but with somewhat different points of view, although many in each group are thoroughly competent in both fields. Metallurgy, in particular, offers an excellent opportunity for a comparison of the two methods.

Jan 1, 1935

By Alan M., Bateman

MR. JORALEMON'S dispassionate discussion of this subject in TECHNICAL PUBLICATION 340 of the Institute shows clearly some of the failures and successes of geology in the discovery of ore deposits. He forces us to recognize, with humility, that chance has played a greater part than geology in the recent discovery of several great copper deposits. I think, however, he wrongly ascribes to unexpected or chance discovery some deposits that were found as a direct result of geological investigation. I refer to the Rhodesian copper deposits,, particularly Roan Antelope and Mufulira, both mentioned by him as having been found by chance. Also,. I think Mr. Joralemon's paper, inadvertently, does not- give the proper credit for the early recognition of the possibilities of these deposits. I feel sure that Mr. Joralemon and the gentlemen given credit by him will welcome any further information on this subject that will throw light on what will some day- be of. historical interest and importance. Consequently, I should like to set forth some additional data that throw light on who should receive the most credit, and' to indicate that, in my opinion, sound. geological reasoning and not chance is to be, credited for the development of the. Roan Antelope and Muf ulira mines.

Jan 1, 1931

H. W. MussEn, Collingwood, Ontario, Can. (communication to the Secretary*):—Doubtless all engineers who have paid more than a casual visit to Russia have come into contact with that formidable document, the " smieta," or estimate of the Russian manager. It is an institution which seems to be inseparable from Russian mining-methods. As Mr. Shockley points out, the year's costs a agree very closely" with the figures of the smieta made at the beginning of the year. In cases which have come under my notice, I have been satisfied that there had been a good deal of " rigging" of accounts to this end. The sum spent by the manager was carefully kept down to the total of the estimate, because it was well known that no more money would be forthcoming, and when this is done it is not a difficult matter to see that the right amounts come under the right heads in the final accounts. I do not say that this is always the case, but I have come across notable instances. Looking at it from the Russian owner's point of view, I think that the balance of the argument is usually in favor of the smieta as limiting " impulsive indiscretion" on the part of managers. The Russian Mining Code of to-day leaves much to be desired, but this is thoroughly appreciated by the official world there, and great changes are spoken of for the future. In many, if not in most, cases the laws are interpreted liberally, and a great deal is left to the discretion of the district mining engineer or inspector, with whom it is necessary to keep in close touch. As Mr. Shockley points out, the officials of a mine, who are always Russians, are held personally responsible for accidents resulting to employees, quite independently of any compensation which may be made to the sufferer. This seems startling enough, but I have always found that a penalty is imposed only

Jan 1, 1909

By R. M. Howe, J. Spotts McDowell

Preparation and Use.—Magnesite is an important refractory in open-hearth, heating, and electric furnaces for steel-making and in many of those employed in the metallurgy of copper and lead. It is sold in the form of brick, finely ground "furnace magnesite" for bricklaying, and dead-burned grains for making and repairing furnace bottoms. The latter are a mixture of granules varying in size from pieces of about 5/8 in. (1.6 cm.) diameter to very fine but sandy particles. Dead-burned magnesite results from calcining the crude or lightly burned mineral at a temperaturc that will not merely drive off practically all the CO2, but will also cause sintering of the particles. During this process the pieces shrink considerably and become hard, dense, and inert to atmospheric moisture and CO2; under-burned material, on the other hand, will hydrate on exposure to the air. A small percentage of ferric oxide seems to be necessary for the production of a satisfactory sinter; from 4.5 to 8 per cent. in the dead-burned grains is considered the most desirable amount. Dolomite has been little used for brickmaking in the United States, but it is prepared for use in the granular condition calcined or "double-burned" and is the principal ingredient of several materials offered for sale under various trade names for refractory purposes. Dolomitic refractories are almost wholly confined to the open-hearth and electric furnaces, where they are used for fettling and as substitutes for magnesite. Much more magnesite and dolomite are used for basic open-hearth steel-making than for all other refractory purposes. The hearth of the furnace is usually built up of magnesite brick and dead-burned grain magnesite so laid that the brick base is protected by a working bottom of the granular material. The latter is sintered into place in layers 1/2 to 1 in: (1.25 to 2.5 cm.) thick to a total depth of 12 to 18 in. (30 to 45 cm.) at the center of the furnace. After each heat, burned dolomite is thrown against the banks as high as it will stick, and all holes in the bottom are Glled. At most plants such holes, at the end of each week, are also carefully filled with dead-burned grain magnesite. For temporary

Jan 1, 1920

By William Greenawalt

The Greenawalt electrolytic copper extraction process is applicable to suitable oxide ores, sulfide ores and concentrates, and low-grade matte. The process is self-sustaining in acid on sulfide ores or concentrates, on mixed ores, on high-grade oxidized ores; and on low-grade oxidized ores the amount of acid to be purchased or manufactured is greatly reduced. The ferric iron is under easy control, and copper may be effectively deposited from solutions of any acid or iron content likely to be used in copper leaching. From 60 to 75 per cent. of the copper may be removed from the solution in passing through the electrolytic department, before returning the solution to the ore. The amount of copper requiring chemical precipitation, from the waste and foul solutions, is only about 5.0 per cent, of the total copper produced, and this 5.0 per cent. is converted into the electrolytic metal in the regular operation of the process. The entire process can be made practically automatic, and can be carried out mechanically. There is no uncertain departure of any kind in the application of the process; no unusual or unproved apparatus is used. EVER since electrolytic copper refining gave promise of success, about a half century ago, efforts have been made to apply the idea to the extraction of copper from its ores. The methods of attack have, usually, been to leach with a dilute acid and then to deposit electrolytically the copper from the resulting solution with the simultaneous regeneration of the solvent used in the process. Many efforts have been made, and many patents have been taken out, both in the United States and abroad, to give this idea practical application, but only within the past few years have the difficulties been surmounted. Leaching and electrolysis, as applied to copper ores, are among the most promising fields in modern metallurgy. New conditions are arising, which will bring the wet methods into prominence; while high freight rates and increasing cost of fuel will tend to limit smelting to highly favored localities. The possibilities offered by the installation of hydro-electric plants and the greatly enlarged range of power transmission will greatly widen the field of electrolytic methods.

Jan 1, 1924

Discussion of the paper of DOUGLAS BUNTING, presented at the New York meeting, February, 1915, and printed in Bulletin No. 97, January, 1915, pp. 1 to 21. ARTHUR HOVEY STORRS, Scranton, Pa.-I know something about accidents Nos. 4 and 8. At the Fuller mine there was no disturbance of the surface in connection with either of those accidents, there was no surface settling, nor signs of sand in the mine, and seemingly the water all came from a subterranean basin or water-bearing stratum overlying the Six Foot vein. A similar thing was shown, in an accident not referred to in Mr. Bunting's table, at the Avondale, where we had an inflow of water, as high as 12,000 gal. a minute, for a considerable period. Regarding the second accident (No. 8) at the Fuller mine: after the first inflow of water (No. 4 of the table), the mining was abandoned in the Six Foot vein. A long slope was. driven in an underlying seam, and when this had reached a point about 2,000 ft. southeast of the place where the first inflow of water occurred, and where the bore holes show there was apparently sufficient rock covering for the Six Foot vein, a rock plane was driven up to the upper vein (Six Foot) and one afternoon this exposed the coal. During the following night the inflow of the water occurred. There were absolutely no workings at this point so that no disturbance of the overlying rocks could have occurred. The result of the inflow of water was to fill up the entire lower-vein workings in about two or three clays and caused the abandonment of the mine. After several years, considerable time and money were spent in an effort to hoist the water from the shaft and mine workings, with varying, but in the main rather indifferent, success until, clue to the breaking of one hoisting rope, the engineer pulled the other hoisting tank up through the tower and down on to the engine, making a complete wreck of the whole plant. It was then decided to allow the mine to soak and it is still doing so, other portions of the property being mined from adjoining collieries.

Jan 5, 1915

By Thomas Fraser, David R. Mitchell

THE primary purpose of the loading plant is to transfer the finished product from the preparation machines to the railroad car, truck, or barge in which it is to go to market. Secondary purposes of the loading plant are: to prevent excessive degradation of sizes during loading, to reduce segregation as much as possible, and to prevent excessive quantities of very good or off-grade coal from entering any one car. Nonuniformity of consignments is one of the chief sources of com- plaints by consumers and coal merchants. A properly designed loading plant will do much to reduce complaints and ensure uniformity, not only from shipment to shipment but also in each individual consignment along the length of the car and in cross section. DEGRADATION IN STORAGE, LOADING AND SHIPMENT The object in loading is not only to handle the coal without breakage and segregation, but also to produce a well-trimmed load that will ride to its destination without spillage or breakage en route. The small sizes present in a car of prepared coal as received by the purchaser consist of: missed small sizes not screened out in the preparation plant, breakage during loading, and breakage caused by handling en route. For domestic sizes of anthracite shipped from the Pennsylvania anthracite field to Atlantic coast cities, the dealers' loss because of break- age in loading and shipment amounts to from 2 to 5 per cent and for anthracite shipped to Middle West cities, the loss is 8 to 9 per cent. This difference is in part because dealers in the Middle West rescreen their coal more thoroughly than is usual in the East. One dealer handling Illinois No. 6 coal shipped from southern Illinois mines to Chicago reports an average size loss of 10 per cent when lump coal is unloaded with a fork. This runs as low as 5 per cent for carefully loaded coal and as high as 15 per cent for carelessly loaded coal.

Jan 1, 1943

By J. A. Fellows, Earnshaw Cook, H. S. Avery

THE increased use of heat-resistant alloys (26 per cent Cr, 12 per cent Ni; 16 per cent Cr, 35 per cent Ni; 12 per cent Cr, 60 per cent Ni; etc.) in recent years has been accompanied by continued demands by the consumer for improvement in properties. The manufacturers of these materials, of necessity, have resorted to more critical studies of the basic causes of variations in fundamental properties. To keep pace with the need for augmented information, previous techniques of investigation have been refined and new test methods and instruments created. It is the purpose of this paper to describe the apparatus, experimental technique and precision of measurement that were employed in the exploration of properties of heat-resistant alloy presented in a companion paper.? The detailed precautions in experimental procedure essential for attainment of accurate and comparable data at elevated temperatures cannot be too highly emphasized. THE CREEP TEST Since the fundamental concepts of creep testing have been recognized for the past 20 years, present interest is restricted to refinements in equipment and experimental procedure. The creep-test laboratory constructed for the American Brake Shoe and Foundry Co. is designed to permit 1000-hr. tests in the temperature range of 1200° to 2200°F. with a choice of applied stress extending from 50 to 50,000 lb. per sq. in. in increments of 50 lb. per sq. in. To simplify the calculation of weight combinations, the loading beams (Fig. I) are provided with counterweights to eliminate the tare. The precision of the knife-edges (10.1 ratio) is such that a 10-gram weight placed on the upper specimen grip will be exactly balanced by a 1-gram weight attached to the loading shackle. Check of the loading weights proved the maximum error to be 0.06 per cent. Axial alignment is provided by crossed knife-edges in the grip seats. The furnaces are similar in design to those of Crane and Company, Battelle Memorial Institute, and the U.S. Steel Corporation. An 80 per cent Pt, 20 per cent Rh No. 19 B. and S. gauge wire is used as the heating element. The details of construction are shown in the radiograph reproduced in Fig. 2. The windings are held in place on an alundum tube with alundum cement; the distribution of turns is adjusted to give a maximum temperature gradient of ± 1°F. over a 4-in. gauge length (Fig. 3). To avoid sharp temperature differentials at the ends of the test section, the length of the furnace is four times the gauge length. Sixteen-inch coupons are used to reduce the temperature of the threaded

Jan 1, 1942

1913. FRANDKA WBONA DAMS Montreal, Canada. 1921. WILLIAM CUTHBERTB LACKETT Sacriston, Durham, England. 1923. GELASIOC AETANI. Rome, Italy. 1929. TAKUMAD AN Tokyo, Japan. 1920. HENRYS TURGISD RINKER. Merion Station, Pa. 1922. FEDERICGIOO LITTI. Torino, Italy. 1906. SIR ROBERTA . HADFIELD. London, England. 1928. JOHNHA YSH AMMON.D. New York, N. Y. 1921. FRANKW ILLIAMH ARBORD London, England. 1917. HERBERTH OOVER. Washington, D. C. 1906. HENRI LECHATELIER. Paris, France. 1900. WALDEMAR LNDGREN Cambridge, Mass. 1879. JOHNM ARKLE .New York, N. Y. 1921. C. MCDERMID London, England. 1913. EZEQUIELO RDONEZ Mexico City, Mexico. 1909. ALEXANDRPEO URCEL Paris, France. 1911. ROBERT H . RICHARDS Cambridge, Mass.

Jan 1, 1929

By Robert Glass Cleland

BECAUSE of the country's vast mineral resources, Alexander Von Humboldt, the great German scientist who visited Mexico, or the Kingdom of New Spain, a hundred and fifty years ago, very aptly called that country the "Treasure Chest of the World." During the three centuries of Spanish rule, mining was the principal industry of the kingdom, the largest contributor to the country's prosperity, the chief common denominator in Mexican economic life. Public revenues and private fortunes, employment and markets, agriculture and trade, charities and churches, the incentive for frontier settlement and exploration, and the funds for outfitting the expedition of colonizer and conquistador-all were dependent upon the streams of precious metals that flowed century after century from the mines of Guanajuato, Zacatecas, Real del Monte, San Luis Potosi, and scores of other minería that lay scattered everywhere throughout the land. Sonora, one of the oldest and most renowned of the frontier provinces of New Spain, had been the scene of a hundred rich strikes, sudden stampedes, and widely heralded bonanzas long before the famous and frenzied Gold Rush

Jan 1, 1952

By Cushwa, C. C.

AT the Park Utah Mine, labor costs of stoping A have been reduced from 30 to 40 per cent. by the use of double-drum hoists and scrapers. The application of scrapers varies with the methods of timbering. In stopes, which must be timbered with square-sets or stringer sets, one method is used; another is used in wide orebodies that can be held open until filling is introduced; and still another in narrow flat veins. Nearly all the oreshoots at the Park Utah lie at less than 50" from the horizontal, and most of the ore is difficult to handle through flat chutes. Proper care is probably the most important requisite for successful scraping, as serviceable equipment can be obtained from several manufacturers. One capable mechanic, therefore, was put in charge of the scraper hoists. He visits every stope daily, oiling hoists and tail1 blocks, inspects all equipment, splices cables and changes "set-ups" when necessary; his pay has been a good investment. The Park Utah now has in use ten double-drum Turbinair hoists, and one Sullivan double-drum a. c. hoist of the same capacity and outside design as the air hoists. The electric hoist is slower than the air hoists and does not pass large boulders easily; it is, however, a very effective machine. The air hoists are fast, powerful and

Jan 1, 1925

By AIME AIME

S ULFURIC acid made in the United States during the last four years has averaged approximately 7,000,000 tons of 50" B6 acid a year. This is double the production of the year 1913. About 66 per cent of the acid is made from sulfur, 16 per cent from pyrite, 13 per cent at copper, lead, and zinc smelters, and the remainder from miscellaneous sources it her the chamber process or the contact process, (at some plants both) is used. In 1923, the contact process accounted for nearly 30 per cent of all the sulfuric acid made in this country and since then it has been increasingly employed. It was first introduced here in 1898, and has always .proved less expensive than the chamber process for the manufacture of .acid of high concentration and purity, but the latter produces acid of low strength suitable for many industries.

Jan 1, 1929

By I. R. Kramer, L. J. Demer

Aluminum single crystals were pulled in an electrolytic cell allowing surface removal during the deformation. The extent of Stages I and 11 of the stress-st-aitz curve was increased and the slope decreased as the rate of metal removed from the surface was increased. An increase of the strain rate caused a decrease in the effectiveness of the metal removal. The data indicate that the work-hardening coefficient in Stage I is determined primarily by the conditions which exist on the surface of the crystal. In Stages 11 and 111, both surface effects and internal barriers are important. ALTHOUGH numerous investigations have been conducted on the plastic flow characteristics of metals in an attempt to explain the mechanism of work-hardening, relatively few studies have taken into account the influence of the surface. In all current theories of work-hardening it is assumed that the impediments to the movement of dislocations are within the crystal. The barriers due to the surface and the existence of solid and liquid films have been neglected even though it has been demonstrated that the surface exerts a large effect. A number of investigators1-l8 have shown that solid films on the surface of single crystals markedly affect their mechanical behavior. In general, the presence of a solid film tends to increase the yield stress and increase the work-hardening rate. Often, on single crystals, Stage I and, at times, Stage II regions are completely suppressed. Various mechanisms have been offered for the effects of oxide and metal films as well as the influence of electrolytes. Of these, concepts concerned with the locking of surface dislocation sources and the blocking of dislocations at the surface resulting in pileups appear to be actively considered at present. Barrett,11 Takamura,6 Gilman,19 Lipsett and King,20 Shapiro and Read,10 and Weiner and Gensamer21 are among those who have interpreted their results in terms of piled-up dislocations at the surface, while Adams.22 and Chalmers and Davis23 have explained their experimental observations in terms of locking of surface dislocation sources. In general, the change in plastic flow properties due to electrolytes has been explained in terms of the unblocking or unlocking of dislocations by the removal of the oxide films. In considering the two proposed mechanisms, it appears that the locking of sources of surface dislocations by a solid film should exert a primary influence only on the critical resolved shear stress for flow and not on the slopes of Stages I and 11. However, the blocking at the surface of dislocations from internal sources may also affect the critical resolved stress and furthermore exert an influence throughout the whole plastic range. In certain cases it does not seem feasible to explain the results of experimental observations in terms of locking of surface dislocation sources. The abnormal aftereffects found by Barrettll,12 by removing the oxide by an acid treatment are excellent evidence of the blocking of dislocations at the surface. Additional evidence in favor of a blocking due to a pileup of dislocations at the surface may be found from the observations that the critical resolved shear strength continues to increase with the thickness of the oxide layer until very heavy oxide layers are formed. If the locking of surface dislocations sources were the dominant factor, the critical resolved shear stress would not be expected to increase after all of the surface sources were locked by the formation of the oxide. This may be expected to happen after a few atomic layers of the oxide are formed. In spite of the above evidence on the strong influence of the surface on the plastic flow characteristic, this has been ignored in current theories of work-hardening. Seeger24,25 suggested that most of the dislocations may slip out of the crystal only when the specimen axis is within certain areas of the orientation triangle. In other areas the resolved shear stress in other glide systems is large enough to generate dislocations which can form Lomer-Cottrell locks, thereby decreasing the average slip distance in some directions and causing a larger hardening rate. Friede126 assumed that at the beginning of Stage 11, a large number of Lomer-Cottrell dislocations are formed by a catastrophic process which used up all the Frank-Read sources on the secondary slip-planes. In this manner a fixed number of Lomer-Cottrell locks is formed which act as barriers against which the dislocations can pile up. In Stage 111, Seeger24 and Diehl, Mader, and Seeger25 proposed that Lomer-Cottrell barriers are circumvented by the cross slip of extended screw dislocations. Cottrell and Stokes,27 Friedel,26 Cottrell,26 and stroh29 suggest that the Lomer-Cottrell dislocations collapse under the stress field of the dislocation pileup. It is the purpose of this paper to report the changes in Stages I, II,and III of the deformation process in

Jan 1, 1962

By Donald F. Hammer, Donald W. Peterson

The Magma mine at Superior, Arizona, has produced over 13 million tons of ore yielding 1.5 billion pounds of copper. It is a mesathermal deposit, and, although the bulk of the ore has come from the Magma vein, most of the recent production has come from replacement deposits in limestone. The mine lies in a Precambrian and Paleozoic sequence of rocks that includes schist, diabase, quartzite, shale, and limestone; nearby, these rocks are intruded by igneous bodies of Mesozoic age that were probably the source of the ore-bearing fluids. Locally, these rocks are overlain by Cenozoic volcanic rocks and conglomerate. Layered rocks at the mine dip eastward about 30°. The area is extensively faulted, and surface faults are grouped into an earlier east-trending set and a later northtrending set. An additional set of diagonal faults that trends northwest and northeast is recognized underground. The east-trending Magma vein-fault, with its splits and branches, is the principal mineralized structure of the district. The bulk of the ore occurs in large ore shoots that plunge steeply westward and are sporadically connected by irregular ore shoots dipping gently eastward. The main ore controls apparently are selective zones of permeability related to the transverse faults. A bed of crystalline limestone near the base of the Martin Limestone (Devonian) is the host for the replacement ore bodies-discontinuous tabular pods dipping eastward parallel to the bedding. Selective permeability is again the apparent ore control, caused both by local hydrothermal solution of limestone and by deformation along faults. The principal copper minerals in the vein are chalcopyrite, bornite, enargite, tennantite, chalcocite, and digenite; these minerals vary in abundance according to a distinct zoning pattern. The upper part of the mine yielded considerable zinc as sphalerite, and the oxidized zone was rich in silver. The chief gangue minerals are pyrite, quartz, and hematite. In the limestone replacement bodies, chalcopyrite and bornite are the chief ore minerals.

Jan 1, 1968

Jan 1, 1925

By John H. Bell

R. B. T. KILIANI, New York, N. Y.-I wish to present some figures based upon actual observation extending over a few months time which seem to prove the author's conclusions. It is a comparison of two machines operating on the same ore, in the same plant, one taking the undersize of ½ -in. screen and the other taking the undersize of 8-mesh. Each of the machines was first operated as a pebble mill, with flint pebbles, in open circuit, then with metal pebbles in open circuit and, finally with metal pebbles in closed circuit. For the first case, the percentage of oversize on 48-mesh, Tyler screen, in the discharge, varied, in the mill taking ½ -in- feed, from 37.2 to 19.9 per cent. The second mill, taking the 8-mesh feed, had a variation in oversize on 48-mesh, from 36.8 per cent: to 0.5 per cent.; so it is seen there is considerable variation in these figures. I have computed the tons per kilowatt-day reduced through 48-mesh for each machine under each one of these conditions, giving 6 sets of figures. The first of these I called 1.0 and determined the ratio of the others to this.. In the snipe way, I figured the relative mechanical efficiency according to Kick's law, using' the figures given by Mr. Taggart in his paper, for ordinal numbers of the Tyler screens. I have also figured the efficiency in mesh-tons per horsepower according to Gates' method by plotting the screen analyses of feed and product and multiplying the area by a constant dependent on the scale used. The commercial efficiency of tons per kilowatt-day reduced through 48-mesh checks very closely with the efficiency by Rittinger's law, within 1 per cent. except in one case. By Kick's law, the efficiency checks fairly closely, within about 5 per cent., with the commercial efficiency of the mill grinding feed, but when we start to grind a finer, 8-mesh material, and especially grinding very fine; with a

Jan 4, 1917