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By John P. Cosgro

THE increasing use of internal furnace-boilers for mining power-plants (doubtless due to the facility with which they may be installed by reason of their portability; the fact that they require no masonry setting, cast-iron fronts, buck-stays, etc., and practically no foundation; and the successful maintenance of their claim to equal and even superior economy, as compared with boilers of other types) warrants this description of a method of remedying an injury which may occur, to such a furnace, even though it satisfies the requirements of the formula adopted by the U. S. Board of Supervising Inspectors of Steam Vessels.2 I refer to the sagging and deformation which may result from overloading the boilers for a considerable period-say, several days. This is sometimes unavoidable under the exacting conditions of continuous day-and-night service, and (in spite of all rules and precautions, the occasional careless or unskillful work of attendants. Such a trouble is by no means unknown on shipboard, where this type of boiler is so largely employed. In the case here described, two Morrison suspension-furnaces, of half-inch steel, 14 ft. long and 45 in. in mean diameter, with ? dry-back ? boilers, were necessarily operated somewhat beyond their rated -normal capacity for several months, while a third similar unit was -in process of installation, and during this period, though the boilers were kept perfectly clean, heat seems 'to have been generated slightly faster than it was absorbed by, the water in the boilers. The corrugations of this.

Mar 1, 1905

Although Australia is the world's second biggest producer of iron ore, the last few years have not been easy for companies in Western Australia's Pilbara region (see map) where more than 90% of the nation's iron ore is produced. For at least two years, the industry has been plagued by dwindling markets abroad, production-crippling strikes at home, skyrocketing costs (well above the official 13% annual inflation rate), an acute shortage of investment capital, increasingly ferocious competition from foreign iron ore producers-and, as if all this weren't enough, by an over-valued currency that has caused incalculable damage not only to iron ore producers but to the entire mining industry of Australia. (This last problem was mitigated less than two months ago, on November 28, 1976, when the Australian dollar was devalued by a whopping 17 1/2%.)

Jan 1, 1977

Discussion of the paper of Robert Henry Lee, presented at the New York Meeting, October, 1913, and printed in Bulletin No. 82, October, 1913, pp. 2515 to 2522. J. E. JOHNSON, JR., New York, N. Y.:-This subject is of great and increasing importance, as the exhaustion of the deposits of ore of ideal character compels the iron industry to turn more and more to those which are less desirable on account of physical structure as well as analysis. Thus there has come about the use of the Mesabi ores in this country and also of the Minette ores in Europe, which produce far greater percentages of flue dust than did the lumpy ores first used. It is a matter of common knowledge that the utilization of this flue dust, containing commonly 50 per. cent. or more of metallic iron, has been a problem very difficult and tedious of solution. The number and variety of the solutions now offered is a sufficient proof of this. . I have no intention of going into the merits or demerits of any of these processes as such, but there are certain considerations concerning the effect of their product on the operation and fuel economy of the blast furnace which I believe to be new and which I desire to bring out. There are two methods, so different as almost to be direct opposites of each other, whereby an ore may exercise an unfavorable effect on the economy of the blast furnace. The first of these is well known and results from the ore being refractory and highly resistant to the action of the reducing gases of the blast furnace. This results in its being incompletely reduced when it reaches the zone of fusion and in the reduction being finished by the direct action of carbon in the lower portion of the furnace. This is so well understood as to need little elaboration or comment. There is, however, another action which may be equally as fatal to the economy of the furnace as the foregoing, and as this point is, to the best of my knowledge, not generally appreciated, it is one upon which I desire to lay particular stress. This consists in the ore being, in a sense, too reducible.

Jan 12, 1913

By Peter K. Strangway

During 1973, a 20,000 net ton (18 100 metric ton) formcoke test was carried out at Inland's 26.5-foot (8.08-meter) hearth diameter on NO. 5 Blast Furnace. The formcoke briquettes were produced from Elko1 coal by FMC at their pilot plant in Kemmerer, Wyoming. The briquettes were shipped to Inland in open railroad cars and most of them were placed in an outside storage field until a sufficient amount was available for the actual blast furnace test. The test consisted of three phases: an initial four-week base period; the actual 33-day formcoke test period; and a final three-week post-test base period. During the formcoke test period, the regular coke skip loads were gradually replaced with skips of formcoke until, after a three-week period, the replacement ratio reached 80%. The furnace operated smoothly and quality hot metal was produced. The burden remained permeable and it was possible to maintain wind rates 9 of over 100,000 scfm (47 Nm3/s) without significant blast pressure increases. Dusting from the formcoke was not a problem either in the stockhouse or in the top gas. The formcoke briquettes maintained their original shape and density as they descended through the furnace. When the 80 percent replacement level had been reached, the inwall temperatures became higher than normal (about 2000 ºF (1090 ºC)) and followed an unusual steady pattern. In addition, the top gas temperature became higher than normal, and the hot metal analyses and temperatures began to fluctuate unacceptably. In general, it is felt that the high inwall temperatures were a result of channeling of the hot furnace gases along the stack walls. This channeling appears to have been promoted by a change in distribution of the burden in the top of the furnace which, undoubtedly, was caused by the increased increments of formcoke. Since it was felt to be unsafe to operate the furnace for an extended period of time with the inwall temperatures at the high level, a decision was made to remove the formcoke from the furnace. When this was done, the inwall temperatures returned to normal. Once the inwall temperatures had returned to normal, the remaining 3,800 net tons (3450 metric tons) of formcoke were used in the furnace at a nominal 40% replacement level for five continuous days. The furnace operated smoothly during this period, and the results were representative of satisfactory furnace operation with partial replacement of regular coke with formcoke. The test was successful in proving that formcoke can replace up to about 50% of the regular coke in a large furnace operating at high wind rates. However, additional work will be required to develop acceptable charging practices for using higher levels of formcoke replacement.

Jan 1, 1977

By Claude P. Heiner

While the title of this paper embraces the entire West Coast, the author, in the interest of simplification. has confined the discussion to California-particularly the central section. California's population has, risen 45 pct since 1940. Its electric requirements have also increased, not only because of the growth in industry and population but also because of the widespread tendency of all types of electric consumers to use more electricity than ever before. The growth of California's electric utilities is shown by published data covering operations, of three of its large systems and is given in Table 1.

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 N. A. Zeigler

GRaphitization phenomena occurring in solid cast iron have principally been studied from the angle of the reactions taking place during annealing of white castings in the manufacturing of malleable iron. The selected bibliography at the end of this paper presents the high spots of the published information in this field. Graph-itization of gray iron upon reheating has received much less attention, primarily because only on rare occasions is this material subjected to any kind of industrial heat-treating. Moreovcr, articles made from gray iron are seldom used commercially at temperatures when any graphitization can take place. When used in such conditions, it is assumed that the articles can be replaced at frequent intervals, which can be afforded because of the low cost of gray iron. However, there is at least one commercial process in which articles made from gray iron, during their manufacturing processing, are subjectcd to rather high temperatures; that is, manufacturing of cast-iron enamelware, which is glazed at about 1600° to 1700°F. (870' to 930°C.) From practical experience, it has been found that one of the convenient and economical compositions of gray iron suitable for enameling is, in round figures, as follows: Si, 2.50 per cent; Mn, 0.35; S, 0.07; P, 0.80; T.C., 3.35; C.C., 0.70. Since this iron usually is cast into thin-walled sections, the percentage of com- bined carbon in it is relatively high. Upon heating in the enameling furnaces, this combined carbon graphitizes, which causes permanent dimensional changes of the articles. Obviously, from practical considerations, it is desirable to maintain these dimensional changes as constant as possible. In the course of experimentation with this iron, two series of experimental heats were prepared in which the manganese to sulphur ratio of the standard cupola metal was offset by ladle additions of: (I) progressively larger amounts of sulphur, and (2) progressively larger amounts of manganese. Each one of these ladle charges was cast into experimental plates 8 by 4 in. and about 5/16 in. thick. All told, 13 ladle charges were thus treated. All of them are listed in Table I, together with their chemical analyses. It may be noted that the manganese to sulphur ratio in these samples ranges between 1.2 and 30.2. Every one of these compositions was subjected to thermal allalysis, using an automatic, self-recording differential dila-tometer. The specimens used in this instrument are 50 mm. long and 4 mm. in diameter. During the dilatometer cycle they are kept under a vacuum of about 3 microns. Each one of them was subjected to a thermal cycle by heating it to I830°F. (1000°C.) in 3 hr. and cooling to room temperaturc in 5 hr. Figs. I, 2 and 3 present three automatically recorded dilatometer curves for irons: Serial No. I (normal composition), Serial No. 6 (high sulphur), and Serial No. 13 (high manganese). All three of these figures are

Jan 1, 1945

By L. F. Coffin

Monotonic tension and controlled cyclic-strain tests have been conducted on an annealed steel at various strain rates and temperatures in the weep range. Evidence of a strain-induced precipitate reaction abozle 500°C was determined from pronounced cyclic-hardening effects, increase inflow stress, inter granular fracture, drastic decrease in low-cycle fatigue resistance, and loss in tensile fracture ductility, as well as through electron-microscopy examination. It was found that strain rate as well as temperature had a pronounced effect on these observations. A parameter of the form P = T(7.95 - log ip) was determined, which was directly related to the ductility, defined either as the true strain at fracture in the case of simple tension, or as 2N 1/2 ? E p in the case of cyclic plastic strain. A highly satisfactory comparison of the cyclic life as measured and as computed from the ductility predicted by this parameter was obtained for a range of temperatures from 400° to 650°C and for crosshead speeds of 0.002 to 0.2 in. per min. MANY metals subjected to sustained loading at elevated temperatures derive resistance to creep from solid-state reactions of various alloying constituents. These reactions can be many and varied, and occur as a consequence of time and temperature according to reaction-rate principles. Strong evidence exists that many of these reactions are influenced by plastic strain, the argument being put forth that dislocations act as sites for nucleation and growth for the new phase.' Many examples can be cited where plastic strain during transformation influences the structure, and as a consequence affects the properties.2 The low-alloy steels serve as particularly good examples for investigating the effect which aging reactions have on resulting properties. Alloying elements such as carbon and nitrogen, and carbide formers such as manganese, chromium, molybdenum, and so forth, can, by the particular aging reaction which occurs, produce strengthening effects at various temperatures, and so endow strength to the alloy over a broad range of temperatures. These effects have been systematically studied by Glen both during creep3 and for constant strain-rate tension testing.4,5 The work of Glen is of considerable interest in connection with the present study, particularly in regard to the effect of various alloying additions on the true stress-strain curve and on fracture ductility at various elevated temperatures. He has shown that strain-age hardening occurred during plastic straining, which was quite specific both to the temperature of the test and to the alloying element involved. When strain-age hardening was noted at a particular temperature, a .minimum in reduction of area was also obtained. Provided the temperature for the minimum reduction of area exceeded 300°C, all minimum ductility fractures were found to be inter-crystalline in nature. For carbon steels containing molybdenum, two minima in reduction of area were noted, one occurring at 200° to 250°C, the precise temperature depending on the percentage of molyb-

Jan 1, 1964

By Fred C. Bond

MANY grindability tests have been made in the Allis-Chalmers laboratory since the tabulated results were last published.1 The lists of standard ball-mill and standard rod-mill grindability tests are brought up to date in this paper, and considerable new material is added. This includes a tabulation of open-circuit grinding tests, with impact and compression-crushing tests. The standard ball-mill grindability tests (Table I) were made in a standard laboratory ball mill, 12 by 12 in. inside, running at 70 rpm, with a charge of 285 iron balls ranging from 1 1/2 to 3/4 in. in diameter and weighing 20,125 grams. Seven hundred cubic centimeters of minus 6-mesh stage-crushed dry feed was used, and the circulating load was maintained constant at 250 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 28 to 200, and the grindability is the number of net grams of screen undersize produced per revolution. The table includes 479 tests. It has been calculated that the ball mill does 52.3 joules of useful work on the ore per revolution2 with 93 joules input to the mill. The standard rod-mill grindability tests (Table 2) were made in a laboratory tilting rod mill, 12 in. in diameter by 24 in. long inside, with a wave-type lining, running at 46 rpm. The grinding charge consisted of six 1 1/4-in. diameter and two 1 3/4-in. diameter steel rods, all 21 in. long and weighing 33,380 grams. The mill was rotated level for eight revolutions, then tilted up 5° for one revolution, down 5° for one revolution, and returned to level for eight revolutions. It was discharged by tilting downward at 45° for 30 revolutions. Feed consisted of 1250 cc of minus 1/2-in. stone and the circulating load was maintained constant at 100 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 3 to 65, and the rod-mill grindability is the number of net grams of screen undersize produced per revolution. The table includes 105 tests. The impact and compression-crushing tests (Table 3) are given in a table that is an extension of a previous publication3 in which the method of testing was described in detail. Two equal hammers weighing about 30 lb each were arranged to strike simultaneous blows on opposite sides of the test specimen. Broken pieces of rock passing a 3-in. square and retained on a 2-in. square were set up with the smallest dimension between the hammers, and this distance was measured. The height of the hammers was increased slightly before each blow until the specimen broke, and the foot-pounds per inch of thickness was calculated. Ten pieces

Jan 1, 1947

By A. T. English

X-ray dif action techniques were used to investigate ordering and domain coalescence for the B2 (CsC1) superlattice in this alloy. Essentially complete disorder is obtainable by drastic quenching. isothertmal ordering occurs very rapidly at temperatures between 565° and 680°C, only a few seconds being required for development of substantial long-range order. A very fine antiphase dormain structure, found in samples annealed for short times, undergoes coalescence with further annealing. Domain "size" is found proportional to (t)1/2 . This Yate law can be derived by adapting the formalism of Hillert's analysis 01- metallurgical grain growth. For this Purpose the material is descvibed in terms of a "Swiss-cheese " structure, consisting of two multiply connected, mutually intertwined domain volumes. The observed activation enevgy for coalescence, 70 kcal per g-atom, agrees with an estit~lated value ,for volume diffusion in disordered rrzaterial. It is suggesled that the experimental value vejlects the substantially disordered environment of mobile vacancies located in the moving antiphase domain boundary. The alloy Fe-Co-2V is bcc below about 900'C and orders into a B2 (CsC1) superlattice below about 730c It has important applications in magnetic components, including many requiring extensive cold rolling.* example, Marcinkowski and Brown6 observed a re1atively small as-quenched domain size of about 250Å in B2 Fe3A1. Since there appear to be no published data on domain coalescence in B2 alloys, this process was studied in detail. MATERIALS AND EXPERIMENTAL TECHNIQUES A commercial high-purity alloy, nominally 49 Fe-49Co-2V, was obtained from the Arnold Engineering Co. in the form of cold-rolled sheet 0.004 in. thick. Samples were annealed for 1 hr at 800°C in dry H2 to produce disorder and quenched into iced brine. They were then annealed in a nitrate-nitrite salt bath for various times at temperatures between 525° and 625°C. Estimates of the LRO parameter S and the APD size D were made using established X-ray diffraction techniques. Superlattice peaks are best observed using CoKa radiation, for which the value of (fFe-fCo) is maximized. Peak-to-background intensity ratio was improved by filtering CoKß from the incident beam to eliminate sample fluorescence, and by using a pulse-height analyzer. The LRO parameter S was computed from the ratio of integrated intensities I 100/I 200- Owing presumably to systematic errors associated with the presence of crystallographic texture and X-ray extinction, only rough agreement was obtained between experimental and theoretical (absolute) intensity ratios. For this reason certain experimental refinements (e.g., use of balanced filters) were considered superfluous and were omitted. Emphasis was placed on maximizing intensity in order to reduce the time taken to accurately determine the contour of the very small (loo), peak. This was done by point counting 105 counts at appropriate intervals. Background intensity averaged about 1200 cps.

Jan 1, 1967

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

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 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 Fred C. Bond

Many grindability tests have been made in the Allis-Chalmers laboratory since the tabulated results were last published.' The lists of standard ball-mill and standard rod-mill grindability tests are brought up to date in this paper, and considerable new material is added. This includes a tabulation of open-circuit grinding tests, with impact and compression-crushing tests. The standard ball-mill grindability tests (Table I) were made in a standard laboratory ball mill, 12 by 12 in. inside, running at 70 rpm, with a charge of 285 iron balls ranging from 1½ to 3/4 in. in diameter and weighing 20,125 grams. Seven hundred cubic centimeters of minus 6-mesh stage-crushed dry feed was used, and the circulating load was maintained constant at 250 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 28 to 200, and the grindability is the number of net grams of screen undersize produced per revolution. The table includes 479 tests. It has been calculated that the ball mill does 52.3 joules of useful work on the ore per revolution2 with 93 joules input to the mill. The standard rod-mill grindability tests (Table 2) were made in a laboratory tilting rod mill, 12 in. in diameter by 24 in. long inside, with a wave-type lining, running at 46 rpm. The grinding charge consisted of six I1/4-in. diameter and two 13/4-in. diameter steel rods, all 21 in. long and weighing 33,380 grams. The mill was rotated level for eight revolutions, then tilted up 5' for one revolution, down 5° for one revolution, and returned to level for eight revolutions. It was discharged by tilting downward at 45° for 30 revolutions. Feed consisted of 1250 cc of minus in. stone and the circulating load was maintained constant at 100 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 3 to 65, and the rod-mill grindability is the number of net grams of screen undersize produced per revolution. The table includes 105 tests. The impact and compression-crushing tests (Table 3) are given in a table that is an extension of a previous publication8 in which the method of testing was described in detail. Two equal hammers weighing about 30 lb each were arranged to strike simultaneous blows on opposite sides of the test specimen. Broken pieces of rock passing a 3-in. square and retained on a 2-in. square were set up with the smallest dimension between the hammers, and this distance was measured. The height of the hammers was increased slightly before each blow until the specimen broke, and the foot-pounds per inch of thickness was calculated. Ten pieces

Jan 1, 1949

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

Molybdenum was first discovered on the Continental Divide in Colorado in 19 11; and in 1918 two companies, one of which was the Climax Molybdenum 250 stpd operation, were producing. In 1919 both mines were shut down, but in 1924 the Climax mine reopened and has been producing ever since. Over the years mine production has been increased until today the mill capacity is 48,000 stpd (43,500 mtpd). From 1935 to 1973 all ore was mined underground by block caving with scraper slushing. Since 1973 ore production has been supplemented by open pit mining. As demand for molybdenum increased, upper levels of the mine became worked-out and new lower levels were opened. In late 1972 production began on the 600 level with the mined ore being primary and secondary crushed in the new No. 4 crushing station. Ore is hauled to the primary in 10 st (9. 1 mt) capacity Granby side-dump cars. The trains, carrying 200 st (181 mt), pass throug.h a dump block and discharge directly into a 54 in. by 84 in. Traylor TC gyratory crusher set at an 8 in. (203 mm) open side setting. Since the ore has already been broken to a size which will pass the finger raises in the mining areas, there are only rare oversize boulders to block the crusher opening. Selection of the crusher was based more on its capacity than its receiving opening, as is often the case with gyratories used in open pit operations. Nominal and maximum capacities for this unit are 1250 and 1800 stph (1 135 and 1635 mtph). The ore is often wet and sticky, so a mechanical car cleaning device is utilized. Under the crusher is a 300 st (270 mt) ore pocket equipped with a 60 in. (1.52 m) wide inclined conveyor which delivers the ore to a 2,000 st (1800 mt) surge bin. Near the head pulley of this conveyor is a wood picking station whe re tramp wood and steel are put into cars for removal. Two more apron feeders remove the primary crushed ore from the surge bin, and impact belts and conveyor belts transport the ore to the crushing station for screening ahead of two 7 ft (2. 13 m) standard crushers. Secondary crushed ore is then conveyed to the surface for tertiary crushing. The entire underground crushing room containing the primary, two secondaries, screens and related equipment is serviced by a 100 st (90 mt) crane with a smaller capacity fast hook. A control room is situated so the operator can observe all the crushing operations. Rotoclone collectors remove dust from the entire operation (including the fumes from the zinc pots) and discharge it to the final product belt. The Rotoclones are vented through drifts and raises to the surface. Shop facilities, electrical gear, and lubrication equipment are located

Jan 1, 1978

By H. R. Hendricks

The ore of the Iron King mine, being very hard and having a very fine crystalline structure, presents many problems in milling that are not present in ordinary lead-zinc ores. This very fine crystallization causes much interlocking of minerals and is most prevalent between the gold and pyrite and the sphalerite and pyrite. The recovery of gold is extremely important in the economic value of the ore and therefore constitutes a large part of the metallurgical problem to be solved. All of the gold is free but it has not been observed in any of the mill products except under very high magnification. Some of the gold is tarnished, and this condition also makes recovery difficult. The gold values have been determined to be highest in association with chalcopyrite and arsenopyrite. The silver occurs mostly in association with galena and tennantite. A small amount of native silver has also been observed occasionally. The galena, being softer than the other minerals, frees much better in grinding. Some oxidation of galena is present in the ore, as the non-sulphide lead content amounts to approximately 10 pct of the total lead. Some oxidation of sphalerite is also present in the ore. Microscopic examination of the zinc concentrate shows many grains in the minus 325-mesh to be half pyrite, together with the fact that some of the zinc is marmatite, makes production of a high-grade zinc concentrate very difficult unless recovery is sacrificed. In the grinding, it has been determined that the flotation feed must be maintained at 85 pet minus 200-mesh to free the minerals sufficiently to make the best economic recoveries. Table I illustrates the effect of fine grinding. It shows that the bulk of the recovered minerals is in the minus 325-mesh size and is the part of the minerals that is more readily freed. The pyrite concentrate is not given in the table because it contains practically no minus 325-mesh. Development of Present Flowsheet and Milling Practice The first milling operation started in 1938, with a mill capacity of 140 tons daily, producing a bulk gold-silver-lead concentrate. After several stages of expansion and changes of the mill flowsheet, the mill is currently treating 380 tons of ore daily and three separate concentrates are produced. In order of their importance and value, these are: (I) lead concentrate, (2) zinc concentrate and (3) pyrite concentrate. The present average grade of mill heads is: gold 0.11 oz, silver 4.00 oz, lead 2.50 pet, zinc 7.60 pct, copper 0.22 pct, iron 22 pct and 30 pct insolubles. Specific gravity of the ore is 3.8. During the period when bulk flotation mas practiced, the flowsheet was very simple. The ore was ground fine enough to liberate as much gold, silver and lead as possible. Necessary collectors and frothers were added in sufficient quantity

Jan 1, 1949

By Donald L. Katz, Marshall B. Standing

Density data are reported on 16 saturated hydrocarbon vapors at Pressures ranging from 1000 to 8220 Ib. per sq. in. and at temperatures ranging from 35° to 250°F. These data have been used to extend the compressibility-factor chart for natural gases up to 10,000 lb. per sq. in. The relatively large quantity of high-boiling constituents present in high-pressure vapors in equilibrium with crude oils makes it necessary to include in the analysis of the gas the molecular weight, density, and possibly boiling range of the heptanes and heavier fraction. Relationships have been presented by which the density of gases may be obtained directly from the temperature, pressure, and gas gravity provided the gases have a common source or are similar in composition. Introduction The densities of natural gases are necessary in many engineering in in petroleum production and utilization, Gas reserves, changes in reservoir pressure, gradients in gas wells, metering of gases, pipeline flow, and compression of gases are typical problems requiring the density of the gas. A decade ago, engineering computations used ideal gas laws with deviations up to 500 Ib. per sq. in.4 Recent discoveries of pools having Pressures UP to 7500 lb. Per sq. in., and installation of pressure-maintenance and recycling plants, have increased the need for data On gas density at high pressures. Methods of Computation The accepted method of computing the density or specific volume of natural gases is the use of the ideal gas law corrected by a compressibility factor. The method proposed by Kay7 of correlating compressibility factors for gaseous mixtures on pseudocritical properties5 appears to be satisfactory for natural gases.2 The meth-me-compressibility factor has been shown to deviate systematically from the behavior of natural gases2 and a chart giving these corrected factors is available. A method of computing specific volume of gaseous mixtures from partial molal volumes has been reported by sage and Lacey.8 Complete data for the computation are not yet available for all hydrocarbon gases. Most- of the reported data on gas densities have been in the single-phase region removed from the dew point or saturation condition. Since most gases in the reservoir Or while on contact with liquid '''ring flow are saturated, the determination of densities under conditions of saturation is particularly important. Although there is no reason to believe that the saturated gases differ in behavior from those consider. ably removed from their dew points, the usual experimental determinations find difficulty on approaching the dew-point conditions, This paper reports the density of I6 saturated gases in equilibrium with crude oils. The relationships developed to compute the density of gases to high pressure' have been modified and enlarged. A simpli-fied method of obtaining gas density is presented in the Form of charts, which apply for specific groups of gases.

Jan 1, 1942

By R. P. McLaughlin

Discussion of the paper of R. P. McLaughlin, presented at the San Francisco meeting, September, 1915, and printed in Bulletin 'No. 108, December, 1915, pp. 2313 to 2319. M. E. LOMBARDI, Berkeley, Cal.-Mr. Bell, I would ask you to state before the meeting what you said to me the other day, about the final recovery of oil from oil sands that have become inundated with water. A. F. L. BELL, San. Francisco, Cal: In stating my opinion, I give it with certain qualifications. In the Kern field, and possibly in some others, I think we will eventually get all the oil out, whether the water comes in or not, since it is peculiarly constructed, being in the form of a partial dome. In 1909, while I was field manager of the Associated Oil Co., I installed some large air compressors to handle the water. At that time the oil had migrated to the higher and more shallow portions of the field, because the water had inundated the deeper sections. The result was that the shallow districts within the center of the dome were in many, cases producing as much or more oil per well than the deeper wells. And in the inundated sections the water had slowly spread over large areas, so that nothing but water was pumped from them. Under these conditions, we decided to pump the wet. well with air, and installed four 2,000-cu. ft. compressors. The air was compressed and forced into the wet wells at, about 250 lb. pressure. Shortly the inundated district that had produced almost all water and only about 150 bbl. of oil a day increased its production to over 2,000 bbl. of oil per day. It was not-long before, we noticed that the production of the shallow territory was falling off to such an extent that in many cases the wells were abandoned; at the same time the area of the water-inundated district was greatly reduced, showing that when the water pressure was reduced the oil migrated back to the lower depth. The wells which at first had had the air jets put into them finally produced so much oil that we had to take the 'air jets out and return to the use of the beam pump, because it is not commercially feasible to pump, an oil of 14° gravity with air. To handle such oil profitably by air, it must carry with it 50 to 90 per cent. of water. The results have led me to believe that oil monoclines, where there is a gradual rise to a certain point, or in plunging anticlines, even though water gets into oil sands, eventually we get the larger percentage of the oil. However, the man with a small acreage in the deeper sections may find some morning that his oil has gone, another man in some higher portion of the district having gotten it.

Jan 12, 1915

By Marshall B. Standing, Donald L. Katz

Density data are reported on 16 saturated hydrocarbon vapors at Pressures ranging from 1000 to 8220 Ib. per sq. in. and at temperatures ranging from 35° to 250°F. These data have been used to extend the compressibility-factor chart for natural gases up to 10,000 lb. per sq. in. The relatively large quantity of high-boiling constituents present in high-pressure vapors in equilibrium with crude oils makes it necessary to include in the analysis of the gas the molecular weight, density, and possibly boiling range of the heptanes and heavier fraction. Relationships have been presented by which the density of gases may be obtained directly from the temperature, pressure, and gas gravity provided the gases have a common source or are similar in composition. Introduction The densities of natural gases are necessary in many engineering in in petroleum production and utilization, Gas reserves, changes in reservoir pressure, gradients in gas wells, metering of gases, pipeline flow, and compression of gases are typical problems requiring the density of the gas. A decade ago, engineering computations used ideal gas laws with deviations up to 500 Ib. per sq. in.4 Recent discoveries of pools having Pressures UP to 7500 lb. Per sq. in., and installation of pressure-maintenance and recycling plants, have increased the need for data On gas density at high pressures. Methods of Computation The accepted method of computing the density or specific volume of natural gases is the use of the ideal gas law corrected by a compressibility factor. The method proposed by Kay7 of correlating compressibility factors for gaseous mixtures on pseudocritical properties5 appears to be satisfactory for natural gases.2 The meth-me-compressibility factor has been shown to deviate systematically from the behavior of natural gases2 and a chart giving these corrected factors is available. A method of computing specific volume of gaseous mixtures from partial molal volumes has been reported by sage and Lacey.8 Complete data for the computation are not yet available for all hydrocarbon gases. Most- of the reported data on gas densities have been in the single-phase region removed from the dew point or saturation condition. Since most gases in the reservoir Or while on contact with liquid '''ring flow are saturated, the determination of densities under conditions of saturation is particularly important. Although there is no reason to believe that the saturated gases differ in behavior from those consider. ably removed from their dew points, the usual experimental determinations find difficulty on approaching the dew-point conditions, This paper reports the density of I6 saturated gases in equilibrium with crude oils. The relationships developed to compute the density of gases to high pressure' have been modified and enlarged. A simpli-fied method of obtaining gas density is presented in the Form of charts, which apply for specific groups of gases.

Jan 1, 1942