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By Alfred Brooks

THE only oil produced in Alaska in 1923 was obtained from a dozen small wells, one of which was drilled during the year on the single patented tract of petroleum land in Alaska, in the Katalla field. These wells are owned by the Chilkat Oil Co., which finds a ready local market for its product in the form of gasoline produced in its own refinery. The search for oil has been continued, but except in the Cold Bay field there has been no drilling in new fields during the year. Drilling was in progress during 1923 at Pearl Dome, 18 miles from the coast. Here the Associated Oil Co. sank two wells, about 500 and 950 ft. deep, but suspended drilling in June. The Standard Oil Co., drilling on the same structure, reached a depth of 300 ft. by August and is reported to have reached about 700 ft. by November. The company began the, installation of a second drill in August. George C. Martin, who hastily examined the Chignik region, south of the Cold Bay field, reports that what is known of the geology near Chignik is not favorable to the presence of petroleum. Kirtley Mather, who made a geologic survey of the northern part of the Alaska Peninsula, reports that some of the structures and the formations are favorable to the presence of oil. A small seepage has long been known in the region near the mouth of Douglas River, which flows into Cook Inlet. Sidney Paige last summer made an examination .of the oil seepages which have long been known near Cape Simpson, 50 miles southeast of Point Barrow, the northernmost point of Alaska. He reports that there are two very large seepages emanating from small ridges about a mile apart and several miles from the coast. These are marked by flowing petroleum, and cover many acres. There is a very large accumulation of petroleum residue at these seepages. The surface exposures consist of clay shales with hard thin partings of calcareous shale which lie nearly horizontal, and are probably, of Jurassic age. Mr. Paige and his two sub-parties explored a large area inland from the Arctic coast between Cape Beaufort on the south and Point Barrow and Cape Simpson on the north. He himself ascended the Meade River for 150 miles. These surveys indicate a wide distribution of what is believed to be the formation from which the Cape Simpson seepages emerge, which is probably of Jurassic age. These rocks near the north Arctic coast are little disturbed, but are increasingly folded to the south, toward the mountains. No other seepages have been found, but what is known of the geology is not unfavorable to an oil field. This entire region lies in Naval Petroleum Reserve No. 4, and it was examined by the Geological Survey at the request and expense of the Department of the Navy.

Jan 3, 1924

By John W. Bell

In the introduction to an excellent pamphlet, John Gross' makes the following statements: Although marked progress has been made along mechanical lines, the theory and conception of underlying principles have not advanced so rapidly. This lack of advance in theory may be attributed to the unfortunate situation resulting from the controversy as to whether the Rittinger or Kick law is applicable to crushing. Many pages of the technical literature are burdened with theoretical discussions in favor of one or the other of these laws, which have tended to cloud rather than to clear the atmosphere. Realizing the situation, the Milling Committee of the American Institute of Mining and Metallurgical Engineers, through its secretary, E. A. Hersam, studied the milling industry and, because of the apparently hopeless lack of agreement in the Kick versus Rittinger dispute (5, 13, 14, 0, i8, 19, 21, 22, 23. 24, 30. 34), concluded that a foundation for the theory of crushing was of primary importance. The numbers in the paragraph quoted refer to papers listed by Mr. Gross in his Bibliography on Crushing and Grinding.1 One is obliged to infer that these papers show "hopeless lack of agreement in the Kick versus Rittinger dispute." No. 22 is a paper by Arthur O. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 23 2,3 the writer will show that their authors formulated identical conclusions, based solely on experimental evidence. Exactly the same defect in the Stadler hypothesis was described in each of the two papers, although entirely different experimental methods were employed in uncovering it. Both of the authors, independently, formulated the conclusion that Rittinger's was a true hypothesis and Stadler's false. The evidence in a third paper4 will be employed to prove that Stadler's hypothesis, when examined experimentally, is worthless. STADLER'S HYPOTHESIS WORTHLESS It is the writer's opinion that the evidence available in 1915 and the corroborating evidence in 1916 was amply sufficient to eliminate the need for any further consideration of Stadler's hypothesis. Certainly the evidence produced in 1918 dealt this hypothesis its coup de grace. It follows, therefore, that in the year 1918 there was only one credible hypothesis left-the Rittinger hypothesis. Gates' Opinion What information is available in Mr. Gates' paper?2 His conclusions are not consecutively summarized, none the less they are perfectly clear. On page 876, Gates says, and it is a very significant utterance: But my results show, among other things, that many more of Stadler's energy units are obtained per foot-pound-applied in coarse than in fine crushing. This suggests a doubt as to the correctness of his unit, since one would expect a foot-pound-applied to produce the same number of "energy units," whether the feed and resulting product were coarse or fine. When Rittinger's theory is applied, the number of mesh-tons (my unit of surface produced) is

Jan 1, 1942

By John W. Bell

In the introduction to an excellent pamphlet, John Gross' makes the following statements: Although marked progress has been made along mechanical lines, the theory and conception of underlying principles have not advanced so rapidly. This lack of advance in theory may be attributed to the unfortunate situation resulting from the controversy as to whether the Rittinger or Kick law is applicable to crushing. Many pages of the technical literature are burdened with theoretical discussions in favor of one or the other of these laws, which have tended to cloud rather than to clear the atmosphere. Realizing the situation, the Milling Committee of the American Institute of Mining and Metallurgical Engineers, through its secretary, E. A. Hersam, studied the milling industry and, because of the apparently hopeless lack of agreement in the Kick versus Rittinger dispute (5, 13, 14, 16, 18. 19. 21, 22, 23, 24, 30, 34). concluded that a foundation lor the theory of crushing was of primary importance. The numbers in the paragraph quoted refer to papers listed by Mr. Gross in his Bibliography on Crushing and Grinding.I One is obliged to infer that these papers show "hopeless lack of agreement in the Kick versus Rittinger dispute." No. 22 is a paper by Arthur 0. Gates2 and No. 23 is one by the writer.3 Employing the subject matter contained in papers Nos. 22 and 232,3 the writer will show that their authors formulated identical conclusions, based solely on experimental evidence. Exactly the same defect in the Stadler hypothesis was described in each of the two papers, although entirely different experimental methods were employed in uncovering it. Both of the authors, independently, formulated the conclusion that Rittinger's was a true hypothesis and Stadler's false. The evidence in a third paper4 will be employed to prove that Stadler's hypothesis, when examined experimentally, is worthless. Stadler's Hypothesis Worthless It is the writer's opinion that the evidence available in 1915 and the corroborating evidence in 1916 was amply sufficient to eliminate the need for any further consideration of Stadler's hypothesis. Certainly the evidence produced in 1918 dealt this hypothesis its coup de grace. It follows, therefore, that in the year 1918 there was only one credible hypothesis left—the Rittinger hypothesis. Gates' Opinion What information is available in Mr. Gates' paper?2 His conclusions are not consecutively summarized, none the less they are perfectly clear. On page 876, Gates says, and it is a very significant utterance: But my results show, among other things. that many more of Stadler's energy units are obtained per foot-pound-applied in coarse than in finc crushing. This suggests a doubt as to the correctness of his unit, since one would expect a foot-pound-applied to produce the same number of "energy units," whether the feed and resulting product were coarse or fine. When Rittinger's theory is applied, the number of mesh-tons (my unit of surface produced) is

Jan 1, 1943

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By S. F. Ravitz

In his excellent book on the Principles of Flotation, Warkl makes the following significant statement concerning the theory of flotation: Two questions of first-rate importance must be considered . . . First, it must be decided whether the sulfide minerals require a collector in order that they may float, or whether as has been claimed, they possess an inherent float-ability and need no collector. Second, it must be decided whether the xanthate type of collector can react with or be adsorbed by the unchanged sulfide minerals themselves, or whether, as many claim, reaction with the collector is dependent upon initial oxidation of the mineral surface. In the theory of flotation there are no two questions of greater importance than these. Ravitz and Porter2 presented evidence that galena can be floated without a collector, and even without any reagent, provided oxidation products are thoroughly removed from its surface. They found also that thoroughly cleaned galena floated faster when xanthate was used than when a frother alone or no reagent was used, and that, without a collector, the recovery decreased gradually upon exposure of the mineral to air. Their work, however, has met with considerable criticism.l(127).3-6 This paper is presented to answer the objections that have been raised, and to present additional data that confirin their results. The apparatus used to provide oxygen-free conditions for flotation has been described in detail by Ravitz and Porter. An atmosphere of purified nitrogen was maintained within the apparatus; all water and solutions used were freed of dissolved air, and could be transferred where desired by the use of nitrogen Pressure; nitrogen was used as the gaseous medium for flotation. The flotation cell was essentially a glass cylinder 15 cm. long and 4.6 cm. in diameter. Near the bottom was sealed a fritted glass filter, which broke up the gas into small bubbles, and 6 cm. above this filter was a side arm through which the froth overflowed. The pulp was agitated by a glass impeller, which entered the cell through a mercury seal. For each test 17 grams of galena and 75 ml. of water were placed in the cell, after which the speed of the impeller and the rate of gas flow were adjusted. Water was added gradually during the test to keep the froth level at the proper height. The mineral was prepared and cleaned in the following manner. Exceptionally Pure galena was crushed by hand on a clean bucking board, screened to —I50+ 200 mesh, and deslimed on a 200-mesh sieve with a fine spray of water. It was then boiled on a hot plate for a number of hours with several changes of the desired cleaning solution, after which both the solution and the mineral were placed in the "mineral flask" of the apparatus. The cleaning solution was drained off, and the mineral was boiled with oxygen-free water

Jan 1, 1943

By W. C. Ellis

Growth directions of tin, zinc, and cadmiun whisker crystals grown spontaneously fi-om the solid are summarized. In tin the edge and diagonal directions of the unit cell are jound with the exception oj- the [110] diagonal. In cadmiurn and zinc the usual dip-ections are a or a + c. The c direction has not been ohserwed. The gvouth directions from the solid are for the most part the same as those found for vapor-grozun crystals and are the ohserz:ed glide directions in plastic j-lou. Whisker crystals of these elements grown from the solid are not twinned. Specular lateral reflections are associated with planes of snall crystallographic indices. WHISKER crystals of tin, zinc, or cadmium grow spontaneously at room temperature in periods from months to years from thin electroplates and other prepared films of the respective metals.''2 It has been established that these crystals are single, but the question of twinning, such as occurs in whisker crystals of silicon,3'4 has not been resolved. Studies of morphology have been restricted to determinations of growth directions.5-' Whisker crystals are both straight and kinked. A few studies have been made on kinked whiskers including measurements of kink anglesa and determinations of growth directions,' but unfortunately not both on the same crystals. The kink angles, for the most part, for each species are grouped closely about a few unique values, but these values are not always those between observed growth directions in a single crystal. To supplement existing data, the growth directions, identification of lateral features, and presence or absence of twinning were determined with a Weissenberg goniometer by a method described elsewhereg on eight whisker crystals of tin, five of zinc, and five of cadmium. One kinked cadmium whisker was examined in detail. The specimens were needlelike, a few microns in cross section, and of the order of a millimeter in length. They were removed from electroplated substrates on laboratory specimens and commercial equipment after several years of growth as described in Table I. Since the specimens were less than 10 p in cross section they were ideally strong with no attendant risk of plastic deformation or damage in handling. Only deformation twinning has been reported for these elements. Therefore, twinning probably should not be expected in whisker crystals of these metals, particularly a twin plane containing the growth direction. A twin plane oriented in this direction was the one sought in this study. No examination was made for twinning at the boundary between kink segments except for one cadmium crystal. Well-defined reflections occur as a whisker crystal of tin, zinc, or cadmium is rotated about its growth direction when illuminated by a beam of light Although these reflections had been noted in earlier investigations, no study of them was made, probably because the earlier studies had disclosed also that

Jan 1, 1967

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 05 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to he used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would he available at each shaft for places where there were no regular blasting stations and could he used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By R. G. Jones, W. L. Kinney, R. E. Schilson, R. S. Wilson, G. A. Clark, H. Suralo

This paper describes the field opeturluns at the Fry in situ combustion project in Crawford County, 111. Field operations may be divided into three phases—-preliminary engineering studies of 1960 and 1961, ignition, and, finally, operation of the test. The Fry combustion project was a pilot test of cocur-rent in situ combustion on a 3.3-acre inverted five-spot. In the summer of 1960, air injection tests were conducted using old wells in the area and a new air injection well drilled especially for the test. In the summer of 1961, four new oil wells were drilled on a 3.3-acre pattern for the pilot test. All old wells in the near vicinity of the pilot test were abandoned to preclude any possibility of gas escaping to adjacent formations. A compressor plant was also installed at the test site in 1961. Air injection tests and other engineering studies were made. At the conclusion of these tests, ignition was accomplished with a 40 kw electrical igniter. Operation of the Fry combustion test was little different from the usual oilfield operation with the exception of the high rate of air injection. No difficulty was found with the openhole completion. There were no unusual corrosion or emulsion problems until bottom-hole temperatures exceeded 150F. Four cores were drilled into the burned-out portion of the reservoir with no mechanical difficulties. INTRODUCTION A cocurrent or conventional in situ combustion test was conducted on the Fry unit of the Marathon Oil Co., in Crawford County, Ill. Besides testing combustion in the Fry type of reservoir, the test was intended to evaluate various operating procedures. This paper presents a discussion of the Fry combustion test field operations. The geology of the reservoir and the test performance are discussed in other papers.'.' HISTORY OF THE FRY UNIT The Fry unit consists of 11 leases totaling 337 acres in an area in which oil was originally discovered in 1906. Production from this reservoir, as from all Robinson sand reservoirs in the area, was very high during the first few years. As the dissolved-gas energy was dissipated, early attempts to retard the decline consisted of vacuum, air, and gas repressuring, and recently, most of the area has been subjected to water flooding. Following early air repressuring, the Fry waterflood unit was formed and water flooding initiated in July, 1952. In the area in which the Fry combustion test was conducted, a peripheral water flood of a 160-acre area was developed as shown in Fig. 1. Between 1956 and 1960, it became apparent that the water flood was unsuccessfd. In looking for a suitable site for an in situ combustion test, the Marathon Oil Co. selected the Fry unit as possessing those features most desirable for such a research project. Field operations associated with the Fry combustion test began in the spring of 1960.

Jan 1, 1966

By F. O. Garrabrant

POWER is furnished to the Climax Molybdenum Co. by the Public Service Co. of Colorado over two 100,000-volt line to a bank of three 3333-kva. Transformers 100/13.8 kv. These transformers arc so designed that with the addition of forced cooling their capacity may be increased 33 1/3 per cent at 12,000-ft. altitude. Under present conditions, these transformers are operated at 140 per cent of their rating with forced air cooling. On the secondary side of this bank there is a 15-kv., 1200- amp. Westinghouse oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station (Fig. I) where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. MINE Power at 13.8 kv. enters the mine through a ventilation raise over a three-conductor, size 1-0 varnished cambric insulated, leaded, steel-wire armored cable, insulated from 17 kv. This power is received and distributed through one switching station and four underground substations, shown in Fig. 2. No. 370 substation (Fig. 3) contains three 250-kva, 13,800/440-volt, single-phase Pyranol transformers protected on the high-voltage side by a 600-amp. manually operated oil circuit breaker. This station serves 20 slusher-hoist motors, a 45-kw., three-phase, 220-volt electric furnace, a Massco hot miller for tempering and milling Timken detachable bits and a hoist with 175-hp. motor on the No. 3 development shaft. Also at this shaft are three pumps. One two-stage pump, 1200 gal, per min., 100-ft. head, 3600 r.p.m., driven by a 150-hp. motor, is at the bottom of the shaft. Two deep-well pumps, 500 gal. per min., 500-ft. head, are at the collar of the shaft. Under normal operation the one pumping 1200 gal. per min. is controlled by a float switch to take care of the water. If this pump fails, or if its capacity is insufficient, the first 500-gal. pump is automatically started, and if still more capacity is needed the second 500-gal. pump automatically starts. Time delay relays are provided on the deep-well pumps, so that if power fails the column will have time to drain before the pumps restart. On failure of any of these, an alarm is sounded in the train dispatcher's office, so that mechanics can be sent to ascertain the cause of failure and make repairs. The cable from No. 370 substation is run to the switching station, in which are four15-kv., 600-amp. direct-current solenoid-operated oil circuit breakers. The current for operating and tripping is furnished by a 125-volt storage battery. All breakers are operated from the switchboard in the compressor room. One circuit from this switching station goes to the

Jan 1, 1944

By T. J. Woods

Tars paper deals with the origin, definition and prevention of scabs an semifinished rolled-steel product. Mold coatings, which are considered essential in scab prevention, were found to be effective only in relation to the ingot surface of semikilled and killed steels and not necessary for the rolled product. That is to say, the scabs on ingots caused by having been cast in uncoated molds are eliminated in the soaking and rolling operations, therefore are not present on the rolled product. This fact, which is contrary to the average opinion, was established after extensive and thorough experiments had proved the ordinary conception of the origin of scabs to be incorrect. This subject may be of little or no concern to steel plants producing small ingots for the manufacture of small billets and bars, except as it may be of interest in showing why their surface troubles with scabs do not exist on the product of heats with normal pouring practice. But to the producer of large ingots to be rolled into slabs for large plates used in shipbuilding, it is a very important subject. DEFINITION OF A SCAB The generally accepted opinion is that the individual scabs present on rolled products (as shown by Fig. I) are considered to have been individual splashes on the ingot surface. This opinion was found to be incorrect. Experiments showed that there were no individual scabs on the product that had appeared as individuals on the ingot surface. The single scabs seen are either pieces of a shell that is present near the bottom of all top-cast ingots, or a veneer that may be present on any part of an ingot improperly poured, as when the stream strikes the mold wall. This definition is based on a detailed study of scabs and their characteristics in which the following have been found true: Location.- On the product of heats poured without trouble, scabs, if present, are always on the product that corresponds to the bottom end of the ingot, and usually occur in clusters, This is true whether molds are coated or not. Shape.-There is no regularity as to shape, size or thickness. It is sometimes possible to separate a few pieces and place them together, like pieces of a jigsaw puzzle, indicating that at one time they had been united as one piece. If individual splashes were the origin of scabs, the foregoing would not be true. Instead, the splashes, which occur throughout the pouring of the ingots, would be present uniformly over the entire product and there would be considerable similarity in the shape and thickness of the scabs. EXPERIMENTAL WORK Our experimental work was conducted on 200-ton heats of semikilled plate steel tapped into an oval ladle 15 ft. high. All ingots were top-cast direct into 24 by 54 by 82-in. molds, straight-sided, set on fat stools. Two separate studies were made about one year apart, each study extending

Jan 1, 1943

By B. H. Cody

THE Morenci concentrator of the Phelps Dodge Corporation is a part of the project that was completed after five years had been spent in development of the open-pit mine, equipment and process testing in the former No. 6 concentrator, and construction of the reduction works. The reduction works, of which the concentrator is a part, is about a mile and a quarter southeast of Morenci, on the highway that leads up Morenci Canyon. The concentrator has a nominal capacity for the treatment of 45,000 tons of ore daily. The original plant, which has a rated capacity of 25,000 tons of ore daily and is now in its fifth year of operation, began production on Jan. 30, 1942. The addition, to increase the capacity 8o pct, began operation on Dec. 1, 1943. The equipment of the original plant has been described in MINING AND METALLURGY (May 1942). No changes in equipment were recommended for installation in the concentrator extension and only one major change, forced by a shortage of maintenance labor, has since been made. Metallurgical treatment has been altered in some ways as experiments progressed and as labor and economic conditions dictated. No complete description of the operation of the concentrator has been published, so this paper will record such operating and equipment changes as have been necessary, the operating methods that have been developed and the metallurgical results that have been obtained. Descriptions are given in the order of the processes with a flowsheet and an equipment list of each process for convenient reference. The questions asked by engineer visitors have guided the detail in which operation and maintenance are described. During the period Jan. 30, 1942 to Sept. 1, 1946, ore milled has totaled 43,595,253 dry tons, from which 1,643,885 tons of concentrate have been made, containing an indicated production of 783,164,809 lb of copper. The principal results obtained in operation of the concentrator during the years 1942 to 1945, inclusive, and to Sept. 1, 1946 are shown in Table 1. A graphic general flowsheet of the plant is given in Fig I. ORE The Morenci ore that is now being treated is a medium-hard monzonite porphyry in which the principal sulphide minerals are chalcocite and pyrite. A minor amount of covellite is present. Chalcopyrite and bornite have been observed under the microscope. The chalcocite occurs mainly as a coating on pyrite. Some of the coatings are so thick that the pyrite presents the appearance of being an inclusion in chalcocite. Some are less than one micron thick and are barely visible at 500 diameters magnification of a polished cross section. Some chalcocite is present as a dissemination in the gangue. Various oxidized copper minerals are present as a result of oxidation in place but they represent a minor proportion of the copper minerals of the ore body. Oxidation

Jan 1, 1947

By B. H. Cody

The Morenci concentrator of the Phelps Dodge Corporation is a part of the project that was completed after five years had been spent in development of the open-pit mine, equipment and process testing in the former No. 6 concentrator, and construction of the reduction works. The reduction works, of which the concentrator is a part, is about a mile and a quarter southeast of Morenci, on the highway that leads up Morenci Canyon. The concentrator has a nominal capacity for the treatment of 45,000 tons of ore daily. The original plant, which has a rated capacity of 25,000 tons of ore daily and is now in its fifth year of operation, began production on Jan. 30, 1942. The addition, to increase the capacity 80 pct, began operation on Dec. I, 1943. The equipment of the original plant has been described in Mining and Metallurgy (May 1942). NO changes in equipment were recommended for installation in the concentrator extension and only one major change, forced by a shortage of maintenance labor, has since been made. Metallurgical treatment has been altered in some ways as experiments progressed and as labor and economic conditions dictated. No complete description of the operation of the concentrator has been published, so this paper will record such operating and equipment changes as have been necessary, the operating methods that have been developed and the metallurgical results that have been obtained. Descriptions are given in the order of the processes with a flowsheet and an equipment list of each process for convenient reference. The questions asked by engineer visitors have guided the detail in which operation and maintenance are described. During the period Jan. 30, 1942 to Sept. I, 1946, ore milled has totaled 43,595,253 dry tons, from which 1,643,885 tons of concentrate have been made, containing an indicated production of 783,164,-809 lb of copper. The principal results obtained in operation of the concentrator during the years 1942 to 1945, inclusive, and to Sept. I, 1946 are shown in Table I. A graphic general flowsheet of the plant is given in Fig I. Ore The Morenci ore that is now being treated is a medium-hard monzonite porphyry in which the principal sulphide minerals are chalcocite and pyrite. A minor amount of covellite is present. Chalcopyrite and bornite have been observed under the microscope. The chalcocite occurs mainly as a coating on pyrite. Some of the coatings are so thick that the pyrite presents the appearance of being an inclusion in chalcocite. Some are less than one micron thick and are barely visible at 500 diameters magnification of a polished cross section. Some chalcocite is present as a in in the gangue. Various oxidized copper minerals are present as a result of oxidation in place but they represent a minor proportion of the copper minerals of the ore body. Oxidation

Jan 1, 1949

By Albert A. Smith

The creep properties of metals have assumed increasing importance in recent years and many investigations have been made on various phases of the problem. In the past year the annual lectures of the Institute of Metals Division1 and the Iron and Steel Division2 were concerned with this subject, and are excellent reviews. For many years the creep properties of lead and lead alloys have been investigated at the Central Research Laboratories of the American Smelting and Refining Co., and it is hoped that the present paper will contribute knowledge of some of the factors that influence creep. Lead, being a metal that recrys-tallizes readily at room temperature, is an excellent material upon which to investigate creep phenomena because of the simplicity of the apparatus needed, and it might be expected that results would apply to some extent to elevated-temperature tests of other metals. Material The creep tests of this investigation were all made on high-purity lead (99.9998+ Pb) in which the only impurity that could be detected by spectrographic and chemical means was approximately 0.0001 per cent Fe. Flat strips, 0.750 in. wide by 0. 100 in. thick, were prepared by extruding 2-in. dia. cast billets at a temperature of 250°C. This gave a material of 2 to 3 mm. average grain size with a tensile strength of j.400 lb. per sq. in. when tested at a straining rate of 5 per cent per minute. With this extremely soft material, great care was necessary in handling the specimens, to avoid distortion and subsequent recrystallization. The creep tests at 30°C. were conducted in a constant-temperature room, specially built for the purpose in the basement of an office building in which there was no moving machinery, and no vibration was detectable under the most severe conditions of shock in the adjacent surroundings. The room was well insulated and heated electrically, control being by means of a de Khotinsky regulator operating a relay. Maximum temperature variation over a period of several years has been ± 1/2°C. Specimens of the extruded strip, 16 in. long, were used for test purposes with no reduced gauge section, thus eliminating any distortion due to machining. The method of clamping and supporting the specimens is shown diagrammatically in Fig. 1. The gauge marker for the measurement of extension consisted of an aluminum strip, 0.050 in. thick by 0.75 in. wide, one end of which was ground and sanded to a precise knife-edge and fastened to the specimen by suitable pins. After the strip was fastened in place a fine scratch was inscribed just below the knife-edge by means of a sharp teasing needle. Just below this gauge scratch a short vertical mark was made so that the same point could be located and read each time. Several hundred specimens have been prepared by this method, and it was found that with experi-

Jan 1, 1941

By Albert A. Smith

The creep properties of metals have assumed increasing importance in recent years and many investigations have been made on various phases of the problem. In the past year the annual lectures of the Institute of Metals Division1 and the Iron and Steel Division2 were concerned with this subject, and are excellent reviews. For many years the creep properties of lead and lead alloys have been investigated at the Central Research Laboratories of the American Smelting and Refining Co., and it is hoped that the present paper will contribute knowledge of some of the factors that influence creep. Lead, being a metal that recrys-tallizes readily at room temperature, is an excellent material upon which to investigate creep phenomena because of the simplicity of the apparatus needed, and it might be expected that results would apply to some extent to elevated-temperature tests of other metals. Material The creep tests of this investigation were all made on high-purity lead (99.9998+ Pb) in which the only impurity that could be detected by spectrographic and chemical means was approximately 0.0001 per cent Fe. Flat strips, 0.750 in. wide by 0. 100 in. thick, were prepared by extruding 2-in. dia. cast billets at a temperature of 250°C. This gave a material of 2 to 3 mm. average grain size with a tensile strength of j.400 lb. per sq. in. when tested at a straining rate of 5 per cent per minute. With this extremely soft material, great care was necessary in handling the specimens, to avoid distortion and subsequent recrystallization. The creep tests at 30°C. were conducted in a constant-temperature room, specially built for the purpose in the basement of an office building in which there was no moving machinery, and no vibration was detectable under the most severe conditions of shock in the adjacent surroundings. The room was well insulated and heated electrically, control being by means of a de Khotinsky regulator operating a relay. Maximum temperature variation over a period of several years has been ± 1/2°C. Specimens of the extruded strip, 16 in. long, were used for test purposes with no reduced gauge section, thus eliminating any distortion due to machining. The method of clamping and supporting the specimens is shown diagrammatically in Fig. 1. The gauge marker for the measurement of extension consisted of an aluminum strip, 0.050 in. thick by 0.75 in. wide, one end of which was ground and sanded to a precise knife-edge and fastened to the specimen by suitable pins. After the strip was fastened in place a fine scratch was inscribed just below the knife-edge by means of a sharp teasing needle. Just below this gauge scratch a short vertical mark was made so that the same point could be located and read each time. Several hundred specimens have been prepared by this method, and it was found that with experi-

Jan 1, 1941

By Norman P. Goss

What would appear to be a very simple process on paper has proved to be overrun with a multitude of unseen difficulties. For more than 70 years attempts have been made to cast metals in a continuous manner, but it has been only within the last Io years that some degree of success has been achieved.1 Historical Development Bessemer's Experiments—Method I Bessemer was the first to try to cast molten metal in a continuous manner. His method is fully described in U. S. Patent No. 49053, issued in 1865. The patent drawing is shown in Fig. I. In this method the molten metal was poured between a pair of rolls of large diameter, which were cooled by water sprays. Means were provided to prevent moisture from coming into direct coptact with the molten metal. The width of the strip was controlled by the flanges. On paper this should work to perfection, but it is beset with many difficulties still awaiting solution. Bessemer failed to reduce his invention to commercial practice. About 1890 Bessemer's method was revived by Norton and associates (U. S. Patent No. 441375). They formed the Fluid Metal Rolling Co. at Maywood, Ill., and spent $250,000 in endeavoring to develop the technique for this simple method. Finally they had to abandon it. Naturally, they did carry the technical development far beyond anything Bessemer had done, but even so it was far from being a useful commercial success. In recent years it has undergone further development by C. W. Hazelett (U. S. Patent No. 2058447), who has succeeded in obtaining long lengths of copper, brass, aluminum, and even stainless steel. He introduced many' ingenious devices but the greatest obstacle still remained—the disintegration of the roll surface when the surface is heated and chilled in rapid succession. Difficulty is also experienced in maintaining a uniform supply of molten metal to the rolls. The patents pertaining to this method would indicate these to be the major obstacles. Experimental work by Mellen—Method 2 A departure from the method introduced by Bessemer, was tried by Grenville Mellen, of East Orange, K. J., sometime between 1914 and 1920, who obtained a number of patents.2 His method is illustrated in Fig. 2. This method failed to make an impression upon the art; however, it is of value in pointing out the difficulties the inventor ran into and the way in which he tried to meet them. The split mold sections form two endless chains. When going through the center of the apparatus, these sections engage each other to form a complete mold. As the engaged mold section moves downward at a constant speed, molten metal is teemed into it. As this is going on, another set of cooperating mold sections move into position on top of the one being filled. The molten metal is teemed into the mold through a silica tube. According to the teachings of this patent, the level of molten

Jan 1, 1945

By G. Gutierrez, D. A. Elgillani, M. C. Fuerstenau

The zero-points-of-charge of apatite, calcite, and fluorite are pH 6.4, 10.8, and 10.0, respectively. Scheelite is negatively charged above at least pH 3. In this article, the flotation responses of these minerals in the presence of potassium oleate and sodium silicate are described and compared with electrokinetic data. Colloidal silica appears to be the species principally responsible for calcite depression, while silicate anion is the species responsible for fluorite depression. Additions as high as 1 x 10-3 mole/liter silicate have no effect on the flotation responses of apatite and scheelite. Selective flotation of nonmetallic minerals is difficult to achieve with fatty acids or soaps by themselves. As a result, specific reagents are added to aid these separations, and one of the reagents commonly employed for this purpose is sodium silicate. Flotation separations of various calcium-bearing minerals such as fluorite from calcite1-3 and scheelite from calcite2,4 and apatite,2,5 for example, almost always involve the use of sodium silicate. The mechanisms by which sodium silicate functions as a depressant are still not understood, probably for a number of reasons. For one thing, the dissolution process of sodium silicate is complex, giving rise to a number of ionic and colloidal species.' Moreover, the type and concentration of these species depend on the ratio of Na2O to SiO2, the concentration of sodium silicate, and the pH of the system.' At the present time, it is not known which species, colloidal silica or silicate anion, is responsible for depression. If colloidal silica is the species that is adsorbing, then adsorption must occur by electrostatic attraction between the colloid and the mineral surface. Silicate anion, on the other hand, may adsorb either physically or chemically. The objective of this paper is to determine first the active species of sodium silicate and then the conditions under which this species will adsorb and function as a depressant. EXPERIMENTAL MATERIALS AND METHODS Pure samples of apatite (Durango), calcite (Iceland-spar), fluorite, and scheelite were used in this investigation. Pure potassium oleate was used as collector, while reagent-grade HC1 and KOH were employed for pH adjustment. The sodium silicates used were samples obtained from the Philadelphia Quartz Co. In one series of experiments, various sodium silicates containing different ratios of SiO2 to Na2O were added to calcite systems. These sodium silicates contained SiO2-to-Na2O ratios of 3.75 to 1, 3.22 to 1, 2.40 to 1, and 1.60 to 1. All other experiments were conducted with the sodium silicate containing the ratio of 3.22 to 1, which is the one that is normally used in industry. Flotation experiments were conducted with 21/2-g charges of 48 x 150-mesh material in conductivity water with an apparatus and technique described previously. Electrokinetic experiments were conducted with both a Zeta Meter and streaming potential apparatus. Particle size was 48 x 65 mesh for the streaming potential experiments. EXPERIMENTAL RESULTS The first series of experiments involved flotation of scheelite in the absence and presence of sodium silicate. As shown in Fig. 1, flotation response was not affected with even the relatively high addition of 1 x 10-3 mole per liter sodium silicate. Interestingly though, no flotation was effected below pH 6 with this collector addition. The responses of apatite to flotation under these same conditions are given in Fig. 2. Similarly, no depression was obtained in basic media under these conditions. Similar experiments were conducted with fluorite, and in this case, depression was noted above about pH 11 with 1 x 10-3 mole per liter sodium silicate (Fig. 3). When calcite was floated with these same levels of addition of sodium silicate, essentially no flotation was possible above pH 7 (Fig. 4). The effect of collector addition with a constant addition of sodium

Jan 1, 1969

By E. C. Smith

THE CHAIRMAN.-Mr. Earle Smith- has kindly offered to make some remarks in connection with segregation in the product, Mr. Smith: E. C. SMITH,* Cleveland, Ohio-I will start this off by a story of the code days of the American Iron and Steel Institute. In those days, Republic did not have a big strip mill, and the question came up of the check analysis of rimmed steel. I suggested that they let me be the chairman of the Committee on Accumulating the Information on the Variation of the Composition of the Ingots of Rimmed Steel. Somebody wanted to know why I wanted to be the chairman. I said, "We don't make any of the big ingots of rimmed steel, and no matter who makes them they are so bad that you don't dare publish the information. So if I collect it and hide it you can't charge me with having done anything that really disrupted the commercial situation. So, everybody furnished me with a rather complete story as to the horrible variations in big rimmed ingots. I took it out to my house and put it away, and it is still there. Now, segregation in the big ingot is no longer an "-academic-matter-as far as II am concerned. ¬I am in a situation like that of the German who was told in 1936 about the coming war. He said: "Until it is your throat that is going to be cut, you don't pay any attention to what is happening to the other man." That is the way with the problem that we have in front of us. Recently I have been told that the comptroller was questioning the payment for certain articles of stamped steel that had been made from rimmed steel, because upon check analysis of the rimmed steel it had been determined that these stamped articles were not within the compositions that were included in the specification, and therefore he as the comptroller had no alternative. It was not meeting the specification; therefore, no payment. So I would say to all of you: You can very seriously consider this matter of segregation now and on into the future on the basis of the fact that it is actually going to cost your companies, which are making these products, money unless there is some new idea about the infallibility of quantitative chemists who analyze some of these products or lawyers who read specifications. I don't know which man is going to be the major factor in the future. If the National Bureau of Standards could in some way show the specification writers the variation between good chemists analyzing for what they expect to be umpire results, the writers would certainly realize that some of the specifications to which we have calmly signed our names as being O.K. are well, as one man said to me, they might be filled by the devil, and they, certainly could be filled by the Lord, but they could not be filled by mankind. I would like to inject two or three things into a segregation study which I want you to think about, because I would be a little surprised if there is very much discussion of them. In the first place, we generally make steel in the United States in the basic open hearth, and I do not think there has been nearly enough consideration given to segregation within the bath before it moves into the ladle. I am sure that any of the people who handled the 9400 series of the NE series when it contained the 0.60 per cent silicon have more respect than they had previously for the segregation that is within the ladle; that is, before the metal reaches the ingot mold, the frozen ingots and the segregation in them. Another point to which I think we ought to give serious consideration is the segregation of iron oxide in the ingot-iron type of material, where the distinction between that and the steel segregation is that it is actually segregating iron oxide. Normally we are segregating some form of carbon or sulphur, or something like. that, with the stamping steel industry now coming down into the ranges of chemistry, where the segregation of iron oxide will be just

Jan 1, 1944

By Earle C. Smith

The Chairman.—Mr. Earle Smith has kindly offered to make some remarks in connection with segregation in the product, Mr. Smith: E. C. Smith,* Cleveland, Ohio—I will start this Off by a story oF the code days of the American Iron and Steel Institute. In those days, Republic did not have a big strip mill, and the question came up of the check analysis of rimmed steel. I suggested that they let me be the chairman of the Committee on Accumulating the Information on the Variation of the Composition of the Ingots of Rimmed Steel. Somebody wanted to know why I wanted to be the chairman. I said, "we don't make any of the big ingots of rimmed steel, and no matter who makes them they are so bad that you don't dare publish the information. So if 1 collect it and hide it you can't charge me with having done anything that really disrupted the commercial situation." So, everybody furnished me with a rather complete story as to the horrible variations in big rimmed ingots. I took it out to my house and put it away, and it is still there. Now, segregation in the big ingot is no longer an "academic " matter as far as I am concerned. I am in a situation like that of the German who was told in 1956 about the coming war. He said: "until it is your throat that is going to be cut, you don't pay any attention to what is happening to the otherman." 'That is the way with the problem that we have in front of us. Recently I have been told that the 'Omptroller was questioning the payment for certain articles of stamped steel that had been made from rimmed steel, because upon check analysis of the rimmed steel it had been determined that these stamped articles were nut within the compositions that were included in the specification, and therefore he as the camptroller had no alternative. It was not meeting the specification; therefore, no payment. So I I say to all of you: You can very seriously consider this matter of segregation now and on into the future on the basis of the fact that it is actually going to cost your companies, which are making these products, money unless there is some new idea about the infallibility of quantitative , chemists who analyze some of these products or lawyers who read specifications, I don't know which man is going to be the major factor in the future. if the National Bureau of Standards could in some way show the specification writers the variation between good chemists analyzing for That they expect to be umpire eel. the writers would certainly realize that some of the to to which we have calmly signed our names as being O.K. are— well, as one man said to me, they might be filled by the devil, and they certainly could be filled by the Lord, but they could not he me by mankind. I would like to inject two or three things into a segregation study which I Want you to think about, because 1 would be a little surprised if there is very much discussion of them, In the first place, we generslly make steel in the United States in the basic open hearth, and I do not think there has bee,, nearly enough consideration given to segregation German the bath before it moves into the ladle. I am sure that any of the people who handled the 9400 series of the NE series when it contained the 0,60 per cent silicon have more respect than they had previously for the segregation that is within, the ladle; that is, before the metal reaches the ingot mold, the frozen ingots and the segregation in them. Another pint to which I think We ought to sis serious consideration is the segregation of iron oxide in the ingot-iron type of material, where the distinction between that and the steel segregation is that it is actually segregat. ing iron oxide, Normally we are segregating some form of carbon or sulphur, or something like that, with the stamping steel industry now coming down into the ranges of chemistry, where the segregation of iron oxide will be just

Jan 1, 1945

By P. R. Hines

Diphenyl guanidine is used as an accelerator in vulcanizing rubber. Other rubber accelerators are also flotation collectors, e.g., dithiocarbamate, thiazole, and the xanthates. Urea and its derivatives are good flotation collectors,' and guanidine is the nitrogen analog of urea. These two characteristics suggested testing diphenyl guanidine as a flotation collector. Diphenyl guanidine has been tried in flotation work previously, but references give no details.'.' Bunker Hill Co. of Kellogg, Idaho, supplied the ore sample used in these tests. The sample was typical of the ore milled in 1949 and contained 4.52 pct Pb and 1.08 pct Zn. A flotation test made by the author with potassium ethyl xanthate as a collector, the same flotation reagents, and the same grind employed in regular Bunker Hill mill practice was the standard for comparing results with the compounds and derivatives of guanidine. The compounds and derivatives of guanidine are only fair collectors of galena. However, if potassium ethyl xanthate were used in the galena float, its presence later in the sphalerite float would make it impossible to introduce another type of collector and be certain which collector contributed the result, so the compound or derivative of guanidine under test was used in both the galena and sphalerite floats in Table I. On both the galena and the sphalerite in the Bunker Hill ore, the depressing action of sodium cyanide and zinc sulfate is much greater with collectors of the guanidine type than with the xanthates. If the depressing agents are left out when a guanidine type of collector is used, but a pH of 8.6 is maintained with either lime or sodium carbonate, the depressing action is approximately the same as in the standard xanthate test. Consequently the depressing agents, sodium cyanide and zinc, were not used in the galena floats of tests 1226, 1228, and 1286, Table I, so the galena floats would be on a comparable basis with the standard xanthate test. As shown by the tests in Table I, the loss in the tail when diphenyl guanidine is used is 8.3 to 9.3 pct, as compared to 17.6 pct with potassium ethyl xanthate, or a recovery of 2.3 to 2.0 lb more zinc. Concentrate grade is 13.7 pct with potassium ethyl xanthate and increases to 22.9 and 31.8 pct with diphenyl guanidine, showing a much higher selectivity. Unless the Barrett's No. 4 coal tar creosote is considered to be a collector, no collector other than diphenyl guanidine is present in the tests, Nos. 1226, 1228, and 1286. Tests 1218, 1253, and 1263 in Table I1 were run to check the effect of the depressing agents used in flotation of the galena (sodium cyanide and zinc sulfate) upon subsequent flotation of sphalerite by diphenyl guanidine. Urea was substituted for potassium ethyl xanthate in tests 1253 and 1263 because it is a good collector for galena and a poor one for sphalerite, and the effect of the depressing agents is more marked with urea than with potassium ethyl xanthate. Test 1205 checks the effect of urea alone without any depressing agents when used for the flotation of galena. Table I1 shows that diphenyl guanidine as a collector for sphalerite is compatible with other flotation reagents used in floating galena. When the depressing agents sodium cyanide and zinc sulfate are used in the galena flotation step, the amount of lime in the sphalerite flotation must be increased (compare 1253 and 1263). Table I11 gives some of the compounds of guanidine which were tested and compared. Some of the compounds and derivatives of guanidine* are solu- * Diphenyl guanidine is in commercial production and is available on the market. ble in water and others are not. Those tested were fed dry. When an organic collector is promising, its compounds and derivatives should be tested and their comparative collecting power recorded. Eventually, in this way, it may be possible to discover some of the essential chemical characteristics of a good collector. For example, it is interesting to compare the diphenyl and dibutyl derivatives in Table IV with those of urea in Table 111, Example 3, U. S. Patent 2,664,198. All are good collectors. Table I shows some of the effects of lime on recovery of sphalerite by diphenyl guanidine. Table V

Jan 1, 1960