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By William M. Shakespeare

Thin sheet copper in wide widths and long lengths is being made electrolytically on a commercial scale at the Raritan Copper Works, Perth Amboy, N. J. This achievement is the result of three years of intensive development preceded by almost ten years of intermittent research. While the idea of producing sheet copper by the electrolytic process is old, the literature on the subject is limited. It is evident, however, that the idea has been attractive to many investigators for years, if the issued patents in the field may be taken as an indication. These patents may be divided into three groups. In one group metal is deposited on some form of endless cathode, of which a revolving cylinder or drum is the most common. The cathode moves slowly and at such a rate that metal of the desired thickness is deposited during one pass through the electrolyte. The deposited metal is stripped off the cathode at a point above the solution level and, after washing and drying, is wound up immediately. In the next group of patents a band of thin metal is increased in thickness by being guided through an electrolytic tank in opposition to suitable anodes. The third group covers means of producing sheets of definite dimensions where flat plates or revolving cylinders are used as cathodes. With such cathodes, sheets of definite dimensions are built up to the desired thickness and then stripped off. The process as worked out at the Raritan Copper Works combines the methods of the first two groups. In the first stage a sheet weighing 1 oz. to the square foot is produced on a revolving cylinder. This sheet is increased to any desired thickness in the second stage. In the first stage (Fig. 1) a lead-covered copper drum is mounted horizontally on a lead-lined concrete tank so that it. is partly immersed in the electrolyte. The cylindrical surface of the drum immersed in the electrolyte is opposed by lead anodes spaced about 36 in. from it. The electrolyte, made up of appropriate quantities of copper and sulfuric acid, flows into and out of the tank continuously. The ebullition of gas at the surface of the anodes induces a vigorous circulation of electrolyte between the anodes and the drum surface, which is augmented by means of compressed air introduced through a pipe between the ends of the anodes under the center of the drum, not indicated on the diagram. This

Jan 1, 1934

By Parke A. Dickey, R. Fettke

Drilling activity during 1944 increased considerably in the oil fields and slightly in the shallow gas fields of western Pennsylvania. The number of deep tests (middle Devonian or deeper) completed during 1944 was the same as in 1943 and did not result in discoveries of additional reserves of natural gas or oil. Thus far no commercial oil production has been obtained from sands below the Upper Devonian in Pennsylvania. As a result of the lack of new discoveries, oil production in nearly all of the producing areas in the entire Pennsylvania grade region declined 10 to 11 per cent. Stocks of Pennsylvania Oil The Peilnsylvania grade crude-oil stocks in storage on January 1943 were 2,479,000 bbl. In January 1944 crude stocks totaled 1,846,000 bbl., and during 1944 the stocks of Pennsylvania grade crude declined to a year-ending balance of 1,617,000 bbl. Nearly all of this oil is "unavailable for use," representing pipe-line fill, tank bottoms and limited working inventories at the refineries. The storage of Pennsylvania oil has declined to a point where little or no available crude oil is currently in storage. Prices The posted price of Pennsylvania oil remained constant during 1944; however, on Rug. I, 1944, the wartime Government subsidy was instituted at a rate of 75¢ per barrel. The average price in the Bradford area, including the subsidy, was $3.31 per barrel. New Technical Developments The Pennsylvania Grade Crude Oil Association started a rather ambitious program of research in production fundamentals and technology during 1944. Several well-qualified engineers have been employed and a laboratory is to be equipped at Bradford. The group will investigate, in cooperation with the American Petroleum Institute, the possibilities of bacteria as agents in oil recovery, and undertake other studies of both a fundamental and an applied nature. In cooperation with the Pennsylvania State College, the Research Group worked out a method of measuring water-intake rates in different strata. The method involves injecting salt water down a string of tubing and fresh water down the annulus, determining the location of the interface electrically, and measuring the input rates above and below the interface. The United States Bureau of Mines, in cooperation with oil companies of the Venango district, developed a simple gas-lift tubing string, which is operating successfully in some 20 wells. This device uses the air-gas mixture produced with the oil to lift it by the usual gas-lift principles. The Northern Ordnance Co. is experimenting with a chamber lift device that expels the oil in an unbroken column. These devices hold considerable promise

Jan 1, 1945

By A. W. WALKER

All branches of production engineering showed steady and definite progress during 1941. Most of it has been of the slower and more conservative type rather than the sensational. To a large degree the changes have been influenced by the expanding war economy, and this tendency will probably be even more marked during the coming year. The stress of the times is placing more emphasis on the principles of engineering and conservation, with the object of obtaining the greatest possible recovery with a minimum waste or expenditure of equipment. Also available material supplies are often being turned to new uses.

Jan 1, 1942

By C. G. Dunn, P. K. Koh

The primary recrystallization textures of 11 single crystals cold rolled 70 pct were determined. The effect of crystal orientation on recrystallization behavior was investigated also and found to be large. An exception was found to the relationship that strong components of the recrystolliza-tion texture are rotated from the cold rolled texture by 30° about <110> axes, but generally this relationship was found to hold quite well. Evidence supporting oriented nucleotion was found and discussed in terms of probable active slip systems and <110> edge dislocations. THE present paper is one of a series&apos;.&apos; on recrys-tallization phenomena in cold rolled 31/4 Si-Fe single crystals and is mainly concerned with the textures produced on recrystallizing from stable end orientations. Data for the cold rolled textures of the crystals used have been reported by Koh and Dunn." Their results showed that the sharpness of the cold rolled textures varied throughout the range of stable end orientations and that the tendency to re-crystallize was related to the sharpness of the cold rolled texture, wherein the rate of recrystallization increased and the size of recrystallization grains decreased with decreasing sharpness of the cold rolled texture. In the project under consideration the cold rolled crystals involved a range of initial orientations that, together with those studied earlier,&apos; provide information on several points. The first point is whether the recrystallization texture from a (111) [112] cold rolled texture, a stable end orientation that can be approached from different directions, correlates in any way with the initial orientation before cold rolling, or whether the recrystallization texture is entirely determined by the observable features of the cold rolled texture. The second point involves a preliminary study to see if the main components of the recrystallization textures correlate with the most probable active slip systems. The third point is to determine what effect departure from slip on one slip system has on the tendency to recrystallize. Experimental Procedure Nine specimens from a previous investigation and two additional ones from thicker stock (lot D) but of similar composition to the others were used in the present work. The latter two crystals, like the nine, were cold rolled to a reduction in thickness of 70 pct and the amount of widening was determined. Depending on their primary recrystallization behavior, the cold rolled crystals were given short or long anneals at temperatures of 980°C or higher. Table I summarizes the data on initial orientations, the textures obtained by cold rolling, and the anneals given to provide samples for the present work. Fine grained specimens were thinned to 0.002 in. thickness and placed in an integrating specimen holder. Chart recordings using a transmission X-ray method and MoKa radiation were obtained for diffraction from (110) planes. This method,&apos;,3 which gives data for all of the pole figure except a central region of 30" radius, proved sufficient to give generally five, and sometimes six, of the six possible pole concentration areas of any one component of the texture. These data, therefore, were generally adequate for a positive identification of each component present. The first texture obtained for specimen B-2 was poorly defined because of too coarse a grain structure and therefore too few crystals in each component for satisfactory X-ray reflections. The same

Jan 1, 1957

Not many accounts of mining methods or labor conditions in the mines in early days can be found; all that have seemed to be authentic have been included in the preceding chapters. In practically every field opened before 18.10 the first coal was recovered by quarrying, or strip mining, the operations, of course, being by hand and on a very small scale; new places were opened as soon as mining became difficult or expensive. Where it became necessary to drift into the hillsides, the first coal was taken out by sleds, and later by small wagons, the power being supplied by the miner, often assisted by dogs. Wooden rails then came into use with small cars running on them. These cars were very small, usually holding only a few bushels; in the Belleville, Illinois field as late as 1863 cars having capacities of 10, 12, 14 and some as large as 30 bushels were in use. It was not until late in the 1830&apos;s that horses or mules came into use for haulage purposes in the Pittsburgh region, although some had been used in the Richmond, Virginia field at an earlier date. Coal was usually undercut by hand and wedged down, until explosives came into general use, and in nearly every field the coal was forked into the cars, the fine coal, or slack, being left in the mine. Hours worked were very long, but few records remain to show what they were at different places and times; there was undoubtedly a great variation between districts and mines. Roy stated a miner could often produce 5 or 6 tons in a day of 8 or 9 hours in a 5 or 6 foot seam, but in another seam possibly could mine only 3 1/2 to 4 tons. He estimated an average day&apos;s work in a four foot seam at about 3 1/2 tons.1 In a strike in the Belleville field in 1863 it was found that an average skilled miner in the field could probably produce 4 to 5 tons a day.2 This was only the coarse coal loaded and did not include slack. In 1873, a coal undercutting machine was tried at a mine about two miles from Brazil, Indiana, said to be the first coal machine ever invented in the United States. It was run by a 5 H.P. steam engine (later to be changed to compressed air), and consisted of an iron rim, four feet in diameter with movable steel teeth placed about one foot apart. The rim

Jan 1, 1942

By Karl F. Peterson, H. K. Najarian, Robert E. Lund

R. E. Powers (Mellon Institute of Industrial Research, University of Pittsburgh, Pittsburgh)—Several of the processing methods at the Josephtown smelter might well be considered by plants producing iron blast furnace sinter. In making blast furnace sinter the objectives are much the same as at Josephtown: To produce a strong yet porous agglomerate which has a size consist in a preferred range. The emphasis however is on production rate rather than on the size or other quality characteristics of the sinter. Therefore the technique is entirely different. Blast furnace sinter plants attempt to make their product in one pass with a minimum recirculating load and without sizing treatment after the sinter leaves the machines. Many plants are beginning to feel that more attention paid to quality characteristics would make the product a more desirable blast furnace material. The pelletizing of very fine material before placing it on the bed is not deliberately done at iron sinter plants, although the installed equipment sometimes gives a simila; action. Such pelletizing seems to be a promising way of accommodating the down-draft sintering Drocess to very fine materials. he recirculating load at Josephtown is extremely high. Two plants&apos; producing blast furnace sinter from raw material of comparable size have a recirculating load of about 100 pct of production. At Josephtown it is notable that the size and the rate of feeding of return fines are under control. This is not the case in most blast furnace sinter plants. The sinter slicer appears to be a worthwhile aid in effecting such control. It is similar in action to rakes and plows used at a number of plants, but it has the advantage that the sinter dislodged is positively removed and not left to the inefficient action of a stationary grizzly screen. Blast furnace size specifications for sinter are not at present so rigid as the Josephtown specifications. If the sizing of blast furnace sinter does become important, then the Josephtown crushing practice, in which a large recirculating load is maintained on the crusher in order that minimum fines be produced, may be of interest. H. K. Najarian (authors&apos; reply)—We note that Mr. Powers appreciates some points that we tried to emphasize in our paper as being of importance in making zinc sinter of good quality and which may well be applied to iron blast furnace sinter. Fig. 2 shows the pelletized nature of the feed going on to a Dwight-Lloyd type sintering machine in our plant. Practically, the entire feed consists of more or less rounded particles which permit more uniform air passage through the charge, while the larger particles in the charge roll on to the surface of the grate, which helps to keep the grate openings from being clogged by excessive fines. An agglomerated charge of this character may preferably be made by rolling action while being mixed in a rotating drum and wetted by water sprays to bind the fine particles together and give the charge the desired moisture content. From our observations of several iron ore and blast furnace flue dust sintering operations, we believe that if the fines from the sinter discharged from the machines were to be recirculated and agqlomerated with the new feed, not only sinter of better quality but larger capacity from the machines may be realized.

Jan 1, 1952

R. E. Powers (Mellon Institute of lndustrial Research, University of Pittsburgh, Pittsburgh)—Several of the processing methods at the Josephtown smelter might well be considered by plants producing iron blast furnace sinter. In making blast furnace sinter the objectives are much the same as at Josephtown: To produce a strong yet porous agglomerate which has a size consist in a preferred range. The emphasis however is on production rate rather than on the size or other quality characteristics of the sinter. Therefore the technique is entirely different. Blast furnace sinter plants attempt to make their product in one pass with a minimum recirculating load and without sizing treatment after the sinter leaves the machines. Many plants are beginning to feel that more attention paid to quality characteristics would make the product a more desirable blast furnace material. The pelletizing of very fine material before placing it on the bed is not deliberately done at iron sinter plants, although the installed equipment sometimes gives a simila; action. Such pelletizing seems to be a promising way of accommodating the down-draft sintering Drocess to very fine materials. he recirculating load at Josephtown is extremely high. Two plants&apos; producing blast furnace sinter from raw material of comparable size have a recirculating load of about 100 pct of production. At Josephtown it is notable that the size and the rate of feeding of return fines are under control. This is not the case in most blast furnace sinter plants. The sinter slicer appears to be a worthwhile aid in effecting such control. It is similar in action to rakes and plows used at a number of plants, but it has the advantage that the sinter dislodged is positively removed and not left to the inefficient action of a stationary grizzly screen. Blast furnace size specifications for sinter are not at present so rigid as the Josephtown specifications. If the sizing of blast furnace sinter does become important, then the Josephtown crushing practice, in which a large recirculating load is maintained on the crusher in order that minimum fines be produced, may be of interest. H. K. Najarian (authors&apos; reply)—We note that Mr. Powers appreciates some points that we tried to emphasize in our paper as being of importance in making zinc sinter of good quality and which may well be applied to iron blast furnace sinter. Fig. 2 shows the pelletized nature of the feed going on to a Dwight-Lloyd type sintering machine in our plant. Practically, the entire feed consists of more ,or less rounded particles which permit more uniform air passage through the charge, while the larger particles in the charge roll on to the surface of the grate, which helps to keep the grate openings from being clogged by excessive fines. An agglomerated charge of this character may preferably be made by rolling action while being mixed in a rotating drum and wetted by water sprays to bind the fine particles together and give the charge the desired moisture content. From our observations of several iron ore and blast furnace flue dust sintering operations, we believe that if the fines from the sinter discharged from the machines were to be recirculated and agqlomerated with the new feed, not only sinter of better quality but larger capacity from the machines may be realized.

Jan 1, 1952

By Jack White, Ralph B. Adair

The Mufulira mine in Northern Rhodesia is 13° south of the Equator and at an altitude of 4100 ft above sea level. The concentrator was planned in 1930 to treat about 10,000 tons of ore per day, but before construction work had gone very far the size of the plant was reduced to 1500 tons daily capacity. Then just as the construction was finished in December 1931, the price of copper did not warrant the commencement of operations and it was not until October 1933 that the mine re-opened. Extensions have been made to the plant in 1934, 1935, 1936, 1937, 1940 and finally in 1945. The capacity of the plant when the 1945 extension is completed will be 9500 short tons per day. The various extensions have been carried out with the minimum confusion and with great skill on the part of the designers and the erection engineers. In dealing with the plant in this paper it is proposed to give an outline of the present practice and only to refer to past practice when it is considered of particular interest. Ore Treated The ore consists of fine grained highly siliceous, felspathic and argillaceous quartz- ites, the latter containing graphitic carbon and locally termed graywacke. These rocks are considerably altered near the surface and on the fringes of the ore body. Despite extensive efforts at mixing, both underground and in the crushing plant, many sudden changes of ore take place in the mill feed. The minerals are chalcocite, bornite and chalcopyrite with relatively unimportant amounts of malachite, native copper, azurite, covellite and chrysocolla. Gold is present at about I oz to a thousand tons, and silver at about 2 dwt. per ton. With increasing depth the nonsulphide content or,, as commonly termed locally, the oxide content of the ore is decreasing and is now less than 0.25 pct against a high of 0.77 pct- Crushing Operations The first stage of crushing is done under. ground where the oversize from I2-in. grizzlies is fed to a 56 X 72-in. jaw crusher. At the Selkirk shaft two 10-ton skips deliver the run of mine ore at a rate of 450 tons per hour to the 400-ton capacity shaft bins. One 6 ft wide Ross chain feeder and one locally made air-operated finger feeder feed the two 30-in. McCully crushers. Ahead of each crusher is a 6 X 9-ft spring grizzley spaced at 5 in. on a 35O slope. Neither the Ross chain feeder nor the finger feeder is entirely satisfactory and for the wet, muddy and sticky ore that has to be handled it is planned to replace these with two 6-ft wide apron pan feeders. The McCully crushers deliver a minus 5-in. product, at a.rate of over 500 tons per

Jan 1, 1949

By Frederick V. Lawrence

Broadly speaking, fluxes are substances which promote wetting and spreading or enhance the fluidity and manipulative properties of materials in joining, fusion, and smelting operations. The term most strictly applies to substances used to facilitate soldering and brazing; but in the discussion which follows, the definition will be broadened to include materials used to promote fusion in arc welding and which aid in the smelting of metals. In each of the mentioned applications, the role of flux and its composition differs considerably; consequently in the discussion which follows, fluxes for soldering, brazing, arc welding, and smelting applications will be considered separately. Soldering and Brazing Fluxes Soldering and brazing are metallurgical joining processes in which a joint is formed using a filler metal of composition dissimilar to that of the joined pieces (base metal). The filler metal, solder or braze, is always of lower melting point than the base metal. Consequently with these processes there is no melting of the base metal. The successful creation of the joint depends upon the wetting and spreading of the solder or braze around or throughout the joint. The obstacles to successful soldering and brazing are contaminants and tenacious oxide films on the metal surfaces. In soldering and brazing applications, fluxes act to remove and exclude oxides and other impurities from the joint. There is no fundamental difference between soldering and brazing. The distinction between them is made on the basis of temperature. Join¬ing operations carried out at temperatures above 500° C are termed brazing. The higher temperatures used in brazing enable higher melting point filler metals to be used which generally provide stronger joints. The higher temperatures of brazing processes also require fluxes which are substantially different from those used in soldering and which will remain effective at high temperatures. Consequently the distinction between soldering and brazing can also be made on the basis of the fluxes used. Soldering Fluxes Even thoroughly cleaned and degreased, metal surfaces still possess tenacious oxide films which inhibit the soldering of the joint. The reactions which form these films increase with increasing temperature. The rate of oxide formation and the tenacity of the film varies with each base metal. Aluminum, magnesium, stainless and high alloy steels, aluminum and silicon bronzes when exposed to air form hard, adherent oxide films. These metals require highly active and corrosive fluxes to remove and prevent the reformation of oxide films during soldering. Metals such as copper and silver, on the contrary, form less tenacious films at a lower rate, so that mild fluxes are sufficient. Soldering fluxes must dissolve or remove oxides and promote rapid wetting and spreading of the solder. The flux removes the oxide film by loosening and floating it off into the main body of the flux. The oxide removal is often aided by the alloying of the solder with the base metal, undermining small areas of the oxide and detaching them (Anon., 1959). The flux coagulates and suspends the oxide particles which have been loosened. The flux forms a protective cover preventing the reformation of the oxides, permitting the spreading of the solder and the formation of a metallic bond between the solder and the base metal. Constituents of Soldering Fluxes: Soldering fluxes are commonly grouped into three categories: highly corrosive, intermediate, and noncorrosive. One normally would choose to use the least corrosive flux which will give satisfactory performance.

Jan 1, 1983

By A. D. Hughes

A relatively small area in Malaya, about 200 miles long by 40 miles wide, is the most important source of tin in the world. Some tin is recovered in other parts of the peninsula. Of the tin mined, 98 pct is recovered from alluvial deposits. From 1935 to 1941 the average annual world production of tin was 190,000 tons. The average annual production from Malaya during the same period was 62,000 tons or about one third of the total. Other producing countries, in order of importance, were the Netherlands East Indies 34,000 tons, Bolivia 30,000, Belgian Congo, Nigeria, Siam, Burma, China, and a few others with smaller amounts. The serious shortage of tin during the war period was due to the fact that the Japanese were occupying Malaya, Netherlands East Indies, Siam, Burma, and China which, together, formerly were producing 65 pct of the world supply. Because of the importance of tin in the world economy, and because of the fact that little is generally known in this country regarding the methods of recovering the tin in Malaya, it is believed that a short description of the industry, as carried on there, may be of interest to many of the readers. In addition you may be interested in the effect of the Japanese occupation, subsequent recovery of the properties and conditions existing at that time. The Malay Peninsula lies at the southeast tip of Asia. Singapore, a British colony, is situated on a small island at the southern end of the peninsula. It is a city of nearly a million population, an important world port, the site of a large British naval base and of one of the tin smelters. The Federation of Malaya consists of nine states, plus Penang, Malacca, and other small areas formerly part of the Straits Settlements, and it now covers the entire peninsula. The civil administration is controlled by the British. A constitution was adopted recently after reaching agreement with the Malay Sultans of the various states, who retain control of religious matters. The Capital is at Kuala Lumpur, a city of about 200,000 population, 250 miles north of Singapore and about half way to Penang. The country covers an area of 51,000 square miles, an area about equal to that of Florida and about one third that of California. It has a population of about seven million consisting of Malays, Chinese, East Indians, and a relatively small number of Europeans. The principal products are tin and rubber. There are over three million acres of planted rubber. The country is not self-supporting agriculturally as about one third of the necessary food supplies is imported. Rice is the principal food of the Asiatic population and some of the recent unrest is caused by a shortage of rice and the consequent high price. Good highways have been built through the settled part of the country. A railway, Government owned, extends from Singapore northward through the entire peninsula. The climate is tropical but health conditions are good for a tropical country. I was in Malaya from 1939 until January 1942, during the time of the simultaneous air attacks on Pearl Harbor, Manila, Hongkong, and Singa-

Jan 1, 1950

By Frank A. Herald

Through the generous cooperation of the authors of the papers in this chapter, a great wealth of fundamental data is being made available. The same plan of securing and submitting data will be carried forward next year, with the hope of making a nearer approach to complete presentation of the data contemplated in the plan. In order to conserve space, we are printing below such footnotes as have general application to Table 1 in the papers in this chapter: Footnotes to Column Headings Table I In areas where both oil and gas are produced, unless gas is marketed outside the field, such areas are included in column headed "Oil." Manufacture of casinghead gasoline and carbon black is interpreted as outside marketing of gas. Production per acre is determined by dividing into the number of barrels of oil the sum of the number of acres assigned to "Oil" plus such number of acres of the total assigned to "Oil and gas" as represents the portion thereof occupied by oil. Wells producing both oil and gas are classified as "Poducing oil only" unless gas from them is marketed off the lease. W. water; G. gas; A, air. Bottom-hole pressures are preceded by "e." All other figures represent pressures at casinghead with well closed. P, paraffin; A, asphalt; M, mixed. 0 Cam, Cambrian; Ord, Ordovician; Sil, Silurian; Dev, Devonian; Mis, Mississippian; MisL, Lower Mississippian; MisU, Upper Mississippian: Pen. Pennsylvanian; Per. Permian; Tri, Triassic; Jur. Jurassic; CreL, Lower Cretaceous; CreU. Upper Cretaceous; Eoc, Eocene; Olig, Oligocene; Mio. Miocene; Pli. Pliocene. 6, sandstone: SH, sandstone, shaly; Ss, soft sand; H, shale; L, limestone; LS, limestone. sandy; C, chalk; A, anhydrite; D, dolomite; Da. arkosic dolomite; GW, granite wash; P, serpentine. Figures are entered only for fields where the reservoir rock is of pore type. Figures represent, ratio of pore space to total volume of net reservoir rock expressed in per cent. "Par" indicates that the reservoir rock is of pore type but said ratio is not known by the author. "Cav" indicates that the reservoir rock is of cavernous type; "Fis," fissure type. A, anticline; AF, anticline with faulting as important feature; Af, anticline with faulting as minor feature; AM, accumulation due to both anticlinal and monoclinal structure; H, strata are horizontal or near horizontal; MF, monocline-fault; MU, monocline-unconformity; ML, monocline-lens; MC, monocline with accumulation due to change in character of stratum; MI, monocline with accumulation against igneous barrier; MUP, monocline with accumulation due to sealing at outcrop by asphalt; D, dome; Ds, salt dome; T, terrace; TF, terrace with faulting as important feature; N, nose; 9, syncline. Information will be found in text as indicated by symbols: A, name of author, other than above. who has compiled the data on the particular field; C, chemical treatment of wells; G, gas-oil ratios; P, proration; U, unit operation; R, references; W, water; 0. other information.

Jan 1, 1934

By Jack White, Ralph B. Adair

THE Mufulira mine in Northern Rhodesia is 13° south of the Equator and at an altitude of 4100 ft above sea level. The concentrator was planned in 1930 to treat about 10,000 tons of ore per day, but before construction work had gone very far the size of the plant was reduced to 1500 tons daily capacity. Then just as the construction was finished in December 1931, the price of copper did not warrant the commencement of operations and it was not until October 1933 that the mine re-opened. Extensions have been made to the plant in 1034, 1935, 1936 1937, 1940 and finally in 1945. The capacity of the plant when the 1945 extension is completed will be 9500 short tons per day. The various extensions have been carried out with the minimum confusion and with great skill on the part of the designers and the erection engineers. In dealing with the plant in this paper it is proposed to give an outline of the present practice and only to refer to past practice when it is considered of particular interest. ORE TREATED The ore consists of fine grained highly siliceous, felspathic and argillaceous quartzites, the latter containing graphitic carbon and locally termed graywacke. These rocks are considerably altered near the surface and on the fringes of the ore body. Despite extensive efforts at mixing, both underground and in the crushing plant, many sudden changes of ore take place in the mill feed. The minerals are chalcocite, bornite and chalcopyrite with relatively unimportant amounts of malachite, native copper, azurite, covellite and chrysocolla. Gold is present at about 1 oz to a thousand tons, and silver at about 2 dwt. per ton. With increasing depth the nonsulphide content or, as commonly termed locally, the oxide content of the ore is decreasing and is now less than 0.25 pct against a high of 0.77 pct. CRUSHING OPERATIONS The first stage of crushing is done underground where the oversize from 12-in. grizzleys is fed to a 56 X 72-in. jaw crusher. At the Selkirk shaft two 10-ton skips deliver the run of mine ore at a rate of 450 tons per hour to the 400-ton capacity shaft bins. One 6 ft wide Ross chain feeder and one locally made air-operated finger feeder feed the two 30-in. McCully crushers. Ahead of each crusher is a 6 X 9-ft spring grizzley spaced at 5 in. on a 35° slope. Neither the Ross chain feeder nor the finger feeder is entirely satisfactory and for the wet, muddy and sticky ore that has to be handled it is planned to replace these with two 6-ft wide apron pan feeders. The McCully crushers deliver a minus 5-in. product, at a rate of over 500 tons per

Jan 1, 1947

By K. D. Hester, A. W. Fletcher

A nickel-copper sulphide concentrate was treated in a pilot plant at Warren Spring Laboratory during 1961, by a hydrometallurgical roast-leach-solvent extraction process devised to cleanly separate the metal values and produce acid sulphate liquors containing the metals. It is shown that 86 pct of the nickel and 95 pct of the copper could be leached from the calcine, after a sulphatizing roast on finely ground flotation concentrate. Iron was removed from the leach liquor by limestone neutralization with a small loss of values. Copper and nickel were separated by solvent extraction with naphthenic acid in kerosene after partial neutralization of the iron-free liquors with ammonium hydroxide. Hydrogen sulphide and chlorine were used to remove zinc (and residual copper) and cobalt impurities respectively. The recovery of pure copper and pure nickel, by electrowinning, from the acidic solvent extract strip liquors finally produced, appears technically feasible. An economic analysis of production costs shows that small, geographically isolated, nickel-copper sulphide deposits cannot usually be worked profitably by established processes. The requirements for profitable operation on a comparatively small scale could possibly be met by a flotation-roast-leach process. Two difficulties, obtaining satisfactory extraction of nickel from a sulphide concentrate and separating the copper and nickel in solution, have prevented the previous use of this method. A solution to the first problem, dissolution of the nickel, was described in a recent paper by Thornhill.1 It was shown in this paper that copper and nickel in a flotation concentrate could be oxidized to water soluble sulphates by roasting in a fluidized bed with added salt cake. Workers at the Warren Spring Laboratory have shown that the separation of copper and nickel can be accomplished by solvent extraction using a relatively inexpensive carboxylic acid, naphthenic acid. Moreover, the mineral acid required to strip values from the loaded solvent could be produced, and the metals recovered, by electrowinning; thus the latter section of the process was regenerative. The removal of iron from the leach liquor and the elimination of impurities such as zinc and cobalt before nickel electrowinning would be necessary to complete the flowsheet. The process under investigation could therefore be summarized (Fig. 1) as follows: 1) Flotation to produce a bulk sulphide concentrate. 2) Fluidized bed roasting to sulphate the nickel and copper minerals. 3) Leaching of roaster calcine to extract soluble nickel and copper. 4) Precipitation of iron from the separated leach liquor by air oxidation after the adjustment of pH and temperature. 5) Recovery of copper from the iron-free Liquor by pH adjustment and liquid-liquid extraction. 6) Purification of the copper raffinate by removing zinc and residual copper as insoluble sulphides and cobalt as cobaltic hydroxide. 7) Recovery of nickel from the purified liquor by pH adjustment and liquid-liquid extraction. Laboratory and pilot plant studies detailed below, were carried out at the Warren Spring Laboratory, to test the feasibility of each one of the unit processes and to obtain preliminary plant design data. FLOTATION The ore was an altered basic rock containing concentrations of magmatic sulphides.

Jan 1, 1964

By J. E. House, J. L. Drobnick, R. R. Swanson, D. W. Agers

The paper describes an improved process for the recovery, purification, and concentration of copper values from acidic dump leach liquors. The process employs a new liquid ion exchange material (LIX® 64 reagent) capable of separating copper values from iron and other impurities common to dump leach liquors in the typical pH range of 1.4-3.0. Properties and performance of the reagent are discussed. Design scale-up information for both pilot plants and commercial installations, including preliminary economics, are presented. Economic considerations include the production of electrolytic copper from the purified copper aqueous concentrate. INTRODUCTION Since the commercial acceptance of the liquid ion exchange process in the mineral processing industry, it has been predicted that a wide selection of chemical reagents to efficiently and economically separate many metals in solution under a wide range of conditions would be commercially available. Several years ago, General Mills began a basic research and development program to determine if it were possible to develop a reagent to extract economically metallic ion from existing process streams. Progress toward reaching this objective was reported in late 1963, when development of quantities of LIX -63 reagent, a highly selective reagent for the extraction of copper, was introduced. The properties of this reagent are described in a previous paper.&apos; The development work with this reagent indicated that perhaps the largest potential application for a copper extractant, the treatment of acidic dump leach liquors, would not be economically possible with LIX-63 because of limitations of the reagent to extract copper below a pH of 3.0. The relatively small pH adjustment required with LIX-63 would result in a significant processing cost and the undesirable precipitation of ferric hydroxide from the liquor. This pointed to the need for another copper extractant which would extract copper from dump leach liquors without pH adjustment. General Mills was successful in synthesizing a molecule which does not have the pH limitation of the LIX-63, yet possesses the selectivity-solubility properties required of a metallic extractant. The resulting LIX-64 reagent broadens the pH range from which copper can be effectively extracted and separated from other metallic impurities. Of special significance is the fact that the reagent, LIX-64 was developed by combining the technical requirements of a specific problem in hydro-metallurgy with theoretical organic and physical chemical principles.

Jan 1, 1967

In Mr. Weed&apos;s interesting paper, frequent reference is made to the Cananea copper-deposits, which are said to have been so vigorously exploited that they produced 14,000,000 Ib. of copper in 1901. The ore-bodies which yielded this large quantity of metal are described as a product of contact-metamorphism, and in Mr. Weed&apos;s genetic classification of such deposits, they receive a place under the sub-heading, "Deposits impregnating and replacing beds of contact-zone."* They are considered by Mr. Weed to represent a special type under this head, and are further classified as: " l. Chalcopyrite deposits: (a) Pyrrhotite ores; (6) Magnetite ores, Cananea type." In condensing much information into a comparatively small space, the presentation of a subject may be so abbreviated as to detract from its clearness. The situation at La Cananea, considering its importance from an economic as well as scientific standpoint, may he worthy of more detailed description. As Mr. Weed rightly remarks, " structural differences are so important, and have SO marked a bearing, not alone on the theory of their genesis, hut also on the working of the deposits and a determination of their value as mines, that they merit a full discussion."† It is my present purpose: (1) to show that it is doubtful whether the copper of La Cananea is mined, except to a limited

Jan 1, 1903

By J. D. Bateman

At Giant Yellowknife, where high grade gold-bearing orebodies are highly irregular in shape, geology has been applied extensively to the mining of ore. The classical functions of the mine geologist in the fields of exploration and mine development have been extended to guide ore extraction, ensuring "clean" mining and effectively reducing waste dilution. THE property of Giant Yellowknife Gold M.ines Ltd. is situated west of Yellowknife Bay on the north shore of Great Slave Lake, a distance of 600 air miles north of Edmonton, Alberta. The Giant claims were staked in 1935, the company was formed in 1937, and the main orebody system was disclosed by diamond drilling in 1944 following a geological study of the property by A. S. Dadson,* Consulting Geologist for the company. Production began in 1948 at the rate of 200 tons per day and, during 1950, reached a daily rate of 425 tons. During the first 3 years of operation a total of 366,000 tons was milled with an average grade of 0.79 oz of gold per ton. Work is in progress with an expansion to 700 tons per day in view. A descriptive account of the geology and gold-bearing shear zone has appeared previously.* The rock formations in the vicinity of Yellowknife Bay have been subjected to protracted pre-Cambrian tectonic deformation culminating in a series of late faults having a cumulative horizontal displacement exceeding 11 miles. The Giant property is underlain by part of an Archean sequence, several miles thick, consisting of basic volcanic flows and minor intercalated tuffs. The volcanic succession forms the west limb of a major syncline, the flows facing east, but overturned on Giant property to dip west at 65" to 75". Orebodies are confined to shear zones up to 200 ft in width, which were formed along early thrust faults. The shear zones assume fold-like attitudes, the larger of which have an amplitude of several hundred feet. The rock formations beyond the limits of the zone of shearing do not reflect the simulated folds, the axes of which are within a few degrees of the strike of the flows. The schistosity and most of the planar elements in both the shear zones and the orebodies dip west • A. S. Dadson and J. D. Bateman: Structural Geology of Camdian Ore Deposits, Can. Inst. Min. Met. Jubilee Volume (1948), PP 273-283. at angles between 65" and 75", generally corresponding to the dip of the flows. The planar elements within the shear zone system thus dip more or less constantly west whether the shear zone is flat, vertical, or expressed as east or west dipping limbs. The shear zones reflect the deformation and alteration of the basic volcanic flows into chlorite schists which, in most places, have undergone metasomatic replacement to form chlorite-sericite-carbonate schists or sericite schists. The boundaries between the shear zones and country rock, although often gradational, usually can be defined within a few feet or even inches as they are expressed by the limits of metasomatic alteration. Orebodies may occupy a small or large proportion of the shear zone and, although they generally conform to the shape of the zone, their morphology is much more complex. Ore boundaries in some instances are sharp and can be delineated with a chalk line; but more generally, a large proportion of the ore boundaries is not visually obvious and can be determined only by the perception acquired by the geological mapping of ore or study of drill cores. Ore shoots in fold-like attitudes may transect the planar elements of the shear zone at any angle; yet the schistosity within the ore shoot may be coincident with that in the enclosing shear zone. Thus it is clear that problems may arise in the delineation of mining boundaries. Ore generally consists of 20 pct or more quartz with ferruginous carbonates in sericite schist deposited in two dominant stages. The earlier stage limits of quartz with carbonate, pyrite, and very fine-grained arsenopyrite in lenses and bands. The later stage consists of quartz-carbonate lodes, in

Jan 1, 1952

By Suzanne T. Mathis, J. C. Ferm, J. T. Berger, W. Horger

A system of core logging and computerized entry and retrieval is described which permits accurate and rapid logging of core, cuttings and geophysical records derived from exploration and development bore holes. The system is so constructed that older drilling records can also be utilized. Output consists of conventional computer print out and graphic logs with a variety of scales, symbols and formats. Such logs can be used in constructing rock profiles which are necessary in property assessment and mine development.

Jan 1, 1983

By C. M. Schwerin

It is often necessary to calculate the size of a hoisting-plant required to raise a given quantity of material, either as a preliminary to the detail design of the machinery, or to decide whether machinery on hand or offered by a manufacturer is adapted to the work to be done. The first element of the problem to be determined is the load to be raised. In a mine that is already developed, this is limited by the size of car that can be hoisted out of the mine and that will pass through the underground gangways. If these place no limits on the design, the size of the load will depend on the output desired per day, and on the number of hoists that call be made per day. The latter are fixed by the time required per hoist and the number of hours available for hoisting, after deducting from the working-day the time required for raising and lowering men, sending down supplies, and for the many small delays in handling cars. It must also be decided whether the hoisting is to be done in one shift or in all of them.

Jan 1, 1903

By G. C. Bernard, L. W. Bernard, L. W. Holm, W. L. Jacobs

The effect of loam on the permeability of porous media to water was studied as a {unction of foaming agent concentration, specific permeability, pressure gradient, length of a porous medium and its oil saturation. At a given fluid saturation in a porous medium, the permeability to water was found to be the same whether loam was present or not. Foam decreases the permeability to water by developing a higher trapped gas saturation than that obtained by water flooding without loam present, increasing the concentration of foaming agent increased the trapped gas saturation and thereby decreased the permeability to water. The presence of oil reduced the capability of most foaming agents to decrease the permeability of a porous medium to water. A few surfactants were found to be effective loaming agents even in the presence of oil. These results are similar to those reported in a previous paper on the effect of loam on the permeability of porous media to gas. The effect of loam was found to persist in long porous media at moderately high reservoir temperatures and during the passage of many pore volumes of surfactant-free water. INTRODUCTION This paper describes part of a study on a novel approach in the use of surfactants for oil recovery; the use of foam rather than water to displace oil. Previously it was found that foam can displace oil which normally is not displaced by water.4 The foam is formed by successively injecting a suitable surfactant solution and gas into a porous medium. Foam appears to have at least two uses in the field: (1) it shows promise as a superior oil recovery agent, and (2) it shows promise as a selective permeability reducing agent. Foam may be very useful in water floods, or in other oil recovery processes, where highly permeable streaks or unfavorable mobility ratios are a problem. A previous paper reported the effect of foam on the permeability of porous media to gas.1 In the present study the effect of foam on the permeability of porous media to water is reported. The specific objectives of the study were to determine: (1) the effect of foam on the permeability to water in porous media of various specific permeabilities, (2) the effect of foam on the permeability to water in the presence of oil, (3) the effect of foam and crude oil on the trapped gas saturation, (4) the effect of foam on permeability to water at trapped gas saturation, (5) the effect of pressure gradient on the permeability to water under foaming conditions, (6) the persistence of foam during the passage of surfactant-free water through the porous medium, and (7) the effect of various foaming agents, length of the porous medium and temperature on the permeability reduction caused by foam. EXPERIMENTAL PROCEDURES EQUIPMENT AND MATERIALS The experimental apparatus consisted of consolidated and unco n soli dated porous media, wet test meters and constant delivery pumps. The porous media consisted of consolidated sandstone cores (6 to 36 in. long), and unconsolidated sand packs (3 to 30 ft long). The consolidated cores had permeabilities of 32 and 1,000 md and porosities of about 20 per cent. The sand packs had permeabilities of 3,500 to 211,000 md and porosities of about 40 per cent. (Throughout this report a term such as 100 md sand" is used. This term means that the porous medium had a dry, nitrogen permeability of 100 md.) Fluids used in the experiments were distilled water, 1 per cent NaCl solution, aqueous solutions of foaming agents, nitrogen gas, air and crude oil.

Jan 1, 1966

THE annual winter meeting of the A.I.M.E., which is to open in New York on Feb. 14, will be charac-terized by a number of novel features. The most important is that the technical sessions will be dis-tributed over four days instead of three as heretofore. This makes it possible to reduce the number of conflicts between sessions and also will enable each member to attend a larger number. At the same time the ses-sions are so arranged by groups that those interested only in one or more that are related can, if necessary, cover all in that group in three days as heretofore. The general excursion to which the fourth day has, for some years, been devoted has been regretfully given up, mainly because of the uncertainties of the weather in New York in February. In its place a series of group excursions are being offered at various times during the meeting, arranged, so far as possible, so as not to con-flict with any technical session likely to be of interest to those taking part in a particular excursion. Among the places to be shown the visitors are: the construction work on the Eighth Avenue subway, a real mining job; the Vehicular Tunnel, the world&apos;s greatest ventilating project; the manufacture of 42-in. conveying belts; one of the largest and most modern steam plants; a coal terminal; manufacture of mining equipment; a cable-making plant; an oil terminal and refinery. Other en-tertainment features include a smoker Monday evening with a real chance to get together and talk; an informal dance and entertainment given by the Woman&apos;s Auxiliary on Tuesday evening; the dinner-dance at the Waldorf Wednesday evening, with no long speeches; a general session Thursday evening when the Petroleum Division will present a series of important reviews of the year. An extensive program has been arranged for enter-taining the ladies, and members are urged to bring their wives. There will be dinners, teas, theater parties, shopping and sight-seeing excursions as well as the dances. The usual general luncheons for members, guests and ladies will be served each day at headquar-ters, and various committees and groups will hold lunch-eons, dinners, and round table conferences as indicated in the schedule on the page opposite. The Mining and Metallurgical Society of America and the American Society of Steel Treaters are participat-ing in this meeting and their members will be welcome as guests at any session or function. A joint conference of the Mining and Metallurgical Society and the A.I.M.E. has been arranged for Tuesday afternoon, under the chairmanship of Clinton H. Crane, at which the whole matter of the mineral reserves of this country is to be up for discussion.

Jan 2, 1927